yi_ xu_n--n. 
 
 REESE LIBRARY 
 
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 UNIVERSITY -OF. CALIFORNIA. 
 
 ] - '-hrassion No. 
 
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THE METALLURGY OF GOLD, 
 
NEW METALLURGICAL SERIES. 
 
 EDITED BY 
 
 W. C. ROBERTS- AUSTEN, C.B., F.R.S., 
 
 Chemist and Assayer of the Royal Mint ; Prof, of Metallurgy, Royal Cottege of Science. 
 
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 Assoc. Mem. Inst. C.E., of the Royal Mint. 
 
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rS 
 ^S 
 
 I 
 
THE 
 
 METALLURGY OF GOLD. 
 
 T. KI11KE ROSE, D.Sc., 
 
 ASSOCIATE OP THE ROYAL SCHOOL OP MINES ; FELLOW OP THE CHEMICAL SOCIETY 
 
 ASSISTANT ASSAYER OP THE ROYAL MINT. 
 
 BEING ONE OF A SERIES OF TREATISES ON METALLURGY, 
 WRITTEN BY ASSOCIATES 
 
 OF THE 
 
 BOYAL SCHOOL OF MINES. 
 
 EDITED BY 
 
 prot TO. <L 1Roberts*Busten t C.JB., JMR,S> 
 
 WITH NUMEROUS ILLUSTRATIONS. 
 
 THIRD EDITION, REVISED AND ENLARGED. 
 
 LONDON: 
 
 CHARLES GRIFFIN AND COMPANY, LIMITED; 
 
 EXETER STREET, STRAND. 
 
 1898. 
 
 [All Rights Reserved.'} 
 
PBEPACE. 
 
 IN the preparation of the THIRD EDITION, the text has 
 again been revised and considerable additions made to 
 it, in order to keep pace with the progress of the 
 metallurgy of gold. In particular, the chapters on the 
 Cyanide Process will be found to have undergone great 
 changes. The frontispiece is from a photograph kindly 
 lent by Mr. Alfred James, late of Johannesburg, and 
 several other new illustrations have been added. 
 
 ROYAL MINT, 
 
 April, 1898. 
 
PREFACE TO THE SECOND EDITION, 
 
 IN preparing this edition, the whole book has been care- 
 fully revised, due attention being paid to some useful 
 suggestions contained in the various reviews of the first 
 edition. Eight new illustrations have been added, in- 
 cluding the frontispiece, which is from a photograph kindly 
 lent by Messrs. Eraser & Chalmers. The rapid progress 
 made in improving the cyanide process has rendered it 
 necessary to make considerable alterations in, and additions 
 to, the chapters devoted to that subject, and a new chapter 
 on Economic Considerations has also been added. The 
 author desires to express his thanks to the old students 
 of the Royal School of Mines and others, who have kindly 
 sent information on work connected with the metallurgy 
 of gold, now in progress in various parts of the world. 
 
 ROYAL MINT, 
 
 August, 1896. 
 
INTRODUCTION TO THE FIKST EDITION. 
 
 As this is the first of a series of treatises devoted to indi- 
 vidual metals, which is being prepared under my guidance 
 it may be well to offer a few introductory remarks. 
 
 Associates of the Royal School of Mines have taken their 
 full share in conducting mining and metallurgical operations 
 in all parts of the world, but, notwithstanding the wide 
 experience they have gained, no treatise claiming to give a 
 general account of the Metallurgy of Gold and adequately 
 dealing with modern processes could hitherto have been 
 attributed to a Student of the School. It may be claimed 
 that in this country Dr. Percy founded the literature of 
 Metallurgy, but his volume on " Silver and Gold," although 
 unrivalled in accuracy of detail, is only a splendid fragment, 
 and gold is alone dealt with in the sections devoted to the 
 refining of bullion and to assaying. A large amount of 
 valuable information concerning new methods and machinery 
 used in the treatment of gold ores has appeared in various 
 official publications during the last twenty years, and much 
 is either scattered over the pages of the scientific press or 
 incorporated in the proceedings of various learned societies. 
 Any attempt, however, to study the subject as a whole from 
 
Vlll INTRODUCTION. 
 
 these sources of information would be hopeless for those 
 whose time is limited, or who have not good libraries at 
 their disposal. 
 
 Mr. Rose gained his practical experience of gold and 
 silver extraction in the Western States of America, and has 
 written a volume which should prove to be very useful, 
 as its careful and conscientious preparation entitles it to 
 confidence. 
 
 W. C. ROBERTS-AUSTEN. 
 
 ROYAL MINT, March, 1894. 
 
PREFACE TO THE FIRST EDITION. 
 
 IN the present volume an effort has been made to supply 
 a succinct summary of the existing condition of the 
 metallurgy of gold, for the use of students and others 
 who are interested in the industries connected with the 
 precious metals. It has been said that mill-managers as 
 a class have little to learn from books ; but, even if this 
 be so, they will still need to keep themselves acquainted 
 with the progress of their art in far distant countries. 
 To them the bibliography appended will be useful. 
 
 Although, in this book, brief accounts are given of some 
 typical machines, attention has been directed rather to 
 methods of procedure than to details of machinery, which 
 will be dealt with in a separate volume in the series. 
 The practice in particular extraction works has been 
 described at some length, as such information, wherever 
 it is available, is among the most valuable that can be 
 given. Some recently devised methods of great importance 
 in the metallurgy of gold, such as the MacArthur-Forrest 
 cyanide process, the new barrel chlorination process, and 
 the improved Gutzkow parting process, are given for the 
 first time in a manual. Particular attention has been 
 paid to the assay of gold bullion, the system in use at 
 the Royal Mint and the precautions necessary to ensure 
 
X PREFACE. 
 
 the highest attainable accuracy being described. Little 
 space is devoted to the geographical and geological dis- 
 tribution of gold ores, as this ground has been amply 
 covered by other works. Moreover the processes of 
 smelting, leaching, and pan-amalgamation used in the 
 treatment of silver ores, whether they contain gold or not, 
 have either been omitted or merely sketched in outline, 
 as belonging to the metallurgy of silver. 
 
 I arn indebted to the President and Council of the 
 Institute of Civil Engineers for leave to reproduce Figs. 
 Nos. 14, 16, 17, 32, 33, 35, and 36, and to Mr. John 
 Murray, for Figs. 58, 60, 61, 62, 63, 64, 66, and 67, which 
 are copied from Dr. Percy's Metallurgy. I also take 
 this opportunity of expressing my thanks to my colleague, 
 Mr. F. W. Bayly, and to other friends for their kind 
 assistance while certain sections were going through the 
 press. 
 
 In conclusion, I desire gratefully to acknowledge the 
 kindness of Prof. Roberts - Austen in giving valuable 
 advice throughout the progress of the work. It was 
 at his suggestion that the task was undertaken, and it 
 is hoped that it will contribute to the realisation of his 
 wish that the experience of School of Mines men should 
 be collected in a number of works which will together 
 form a comprehensive metallurgical series. 
 
 T. K. ROSE. 
 
 ROYAL MINT, March, 1894. 
 
CONTENTS. 
 
 CHAPTER I. THE PROPERTIES OF GOLD AND ITS ALLOYS. 
 
 PAGE 
 
 INTRODUCTION, . . . 1 
 Physical and chemical proper- 
 ties of gold, ... 2 
 Introduction, ... 2 
 Colour, .... 2 
 Malleability and ductility, . 
 Hardness, 3 
 Tenacity, .... 3 
 Specific gravity, ... 4 
 Cohesion, .... 4 
 Specific heat, ... 4 
 Fusibility, .... 4 
 Spectrum, .... 4 
 Latent heat, . . 4 
 
 Magnetism, . 
 
 Conductivity and expansion 
 
 Atomic weight and volume, 
 
 Volatility, . 
 
 Crystallisation, 
 
 Solubility, . 
 Preparation of pure gold, 
 Allotropic forms of gold, . 
 Alloys of gold, . 
 
 Amalgams, . 
 
 Gold and silver, . 
 copper, . 
 
 Liquation, . 
 
 PAGE 
 
 4 
 
 5 
 
 5 
 
 5 
 
 9 
 
 10 
 
 11 
 
 12 
 
 13 
 
 15 
 
 16 
 
 17 
 
 IS 
 
 CHAPTER II. CHEMISTRY OF THE COMPOUNDS OF GOLD. 
 
 Compounds of gold, 
 Haloid compounds, 
 Chlorides of gold, 
 
 Proto- chloride, 
 
 Auro-aurichloride 
 
 Trichloride, . 
 
 Chlor-auric acid, 
 Bromides of gold, 
 
 Proto-bromide, 
 
 Auro-auric bromide, 
 
 Tribromide, . 
 Iodides of gold, 
 Cyanides of gold, 
 
 20 
 21 
 21 
 21 
 21 
 21 
 27 
 21 
 27 
 28 
 28 
 28 
 28 
 
 Aurous cyanide, . 
 
 Auro- cyanide of potassium, 
 
 Auri-cyanide of potassium, 
 Oxides of gold, 
 
 Aurous oxide, 
 
 Auric oxide, 
 Aurates, . 
 
 Fulminating gold, . 
 Sulphites of gold, 
 Hyposulphites of gold, 
 Silicates of gold. 
 Sulphides of gold, . 
 Purple of Cassius, . 
 
 28 
 28 
 29 
 29 
 29 
 29 
 30 
 30 
 30 
 31 
 33 
 33 
 34 
 
Xll 
 
 CONTENTS. 
 
 CHAPTER III. MODE OF OCCURRENCE AND DISTRIBUTION OF GOLD. 
 
 Forms in which gold occurs in 
 
 nature, .... 35 
 Vein gold, .... 35 
 Placer gold and nuggets, . 37 
 Calaverite, . . . .38 
 Sylvanite, .... 38 
 
 Petzite 
 
 Nagyagite, . 
 
 Composition of native gold, 
 Geographical distribution of 
 
 gold, . . . . 
 Origin of gold ores, . 
 
 PAGE 
 3* 
 38 
 39 
 
 39 
 42. 
 
 CHAPTER IV. PLACER MINING SHALLOW DEPOSITS. 
 
 Placer deposits, ... 43 
 Methods of obtaining gravel 
 
 from shallow placers, . 44 
 Appliances used in washing the 
 
 gravel, . .45 
 
 Pan, ... .45 
 
 Batea, ... .45 
 
 Prospecting trough, . 46 
 
 Horn- spoons, . . 46 
 
 Cradle, . .' .46 
 
 Long-torn, . .47 
 
 Puddling-tub, ." .47 
 
 Siberian trough, . . 48 
 
 Sluice, ... . 50 
 
 Sluicing, . . . .51 
 
 Use of drops, . . 52 
 
 ,, undercurrents, . 52 
 
 ,, mercury, . . 53 
 
 Cleaning-up, . .53 
 
 Tail-race, .... 54 
 
 Ground-sluice, ... 54 
 
 Booming, .... 54 
 
 Dry diggings, . . , .55 
 
 Cement gravels, ... 55 
 
 Tail sluices, .... 55 
 
 Fly catchers, .... 56 
 
 River mining, . . . .56 
 
 1. River mining proper, . 56 
 
 2. Dredging, ... 58 
 
 3. Deep bar mining, . . 58 
 Methods of working Siberian 
 
 placers, .... 59 
 
 1. Siberian sluice, . -.. .. 60 
 
 2. Trommel, . . .61 
 
 3. Pan, . . . . 61 
 Beach mining, .... 62 
 New method of working shallow 
 
 placer deposits, . . .62 
 
 CHAPTER V. DEEP PLACER DEPOSITS. 
 
 Nature and mode of origin of 
 
 deposits, .... 
 
 Distribution of gold in the 
 
 gravels, .... 
 
 Origin of gold in the gravels, 
 
 Minerals occurring in the 
 
 gravels, 
 
 Methods of working, 
 " Hydraulicking," 
 
 Commencement of opera- 
 tions, .... 
 
 Supply of water, 
 
 Breaking down the bank, 
 
 Sluicing gravel, *. . 
 
 Cleaning-up, 
 
 Disposal of tailings, . 
 Drift mining, 
 Shaft ,, 
 Hydraulic elevator, 
 Economic conditions, 
 Shallow placer deposits, 
 Deep 
 
 72 
 75 
 
 77 
 80 
 81 
 
 82 
 82 
 84 
 86 
 86 
 
CONTENTS. 
 
 X11I 
 
 CHAPTER VI. QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 Primitive methods of crushing 
 
 and amalgamating, . 
 Mortar, .... 
 Maray, . . .91 
 
 Chilian mill, . . . 
 Arrastra, 
 
 Iron prospecting arrastra, 
 Stamp battery, 
 Rock-breakers, . 
 
 Blake, 
 
 Dodge, . 
 
 Gates, 
 
 Schranz, 
 Material 
 
 employed for 
 
 crushing surfaces, . 
 
 Position and use of rock- 
 breakers, . . ",.' 
 Battery proper, . ... 
 
 Foundations, . . . 
 
 Framework, 
 
 PAGE 
 89 
 
 91 
 91 
 92 
 92 
 95 
 95 
 101 
 101 
 102 
 103 
 104 
 
 105 
 
 106 
 107 
 108 
 108 
 
 Mortar, , . ' 
 Splash-box, 
 Dies, 
 
 
 PAGE 
 108 
 
 110 
 110 
 
 111 
 111 
 
 112 
 113 
 114 
 114 
 115 
 115- 
 120 
 12a 
 121 
 122" 
 123 
 124 
 126 
 > 
 . 127 
 
 Cam-shaft 
 Cams, 
 Tappets, 
 Stamp, 
 Guides, 
 
 
 
 Height of drop, 
 Screens, . 
 Order of fall of stamps, 
 Automatic feeding, 
 Stanford's feeder, 
 Challenge 
 Water supply, 
 Amalgamated plates, . 
 Mill site, 
 General arrangement of th< 
 stamp mill, . 
 
 CHAPTER VII. AMALGAMATION IN THE STAMP BATTERY. 
 
 Treatment of the amalgamated 
 
 plates, 129 
 
 Discoloration of the copper 
 
 plates, 130 
 
 Chemicals used to promote 
 
 amalgamation, . . 131 
 
 Use of mercury, . . 132 
 Grade of plates, . .'132 
 
 Muntz metal plates, . 133 
 
 Shaking copper plates, . 134 
 
 Corrugated plates, . . 135 
 
 Mercury wells, . . 135 
 
 Galvanic action in amalgama- 
 tion, . . . 
 Designolle process, . 
 Cleaning- up, . 
 
 The clean-up barrel, 
 
 pan, 
 
 Retorting, 
 Loss of mercury, 
 
 ,, gold, . 
 
 Non-amalgamable gold, . 
 Gold in pyrites, 
 
 135 
 13& 
 137 
 13S 
 138 
 140 
 142 
 145 
 151 
 152 
 
 CHAPTER VIII. OTHER FORMS OF CRUSHING AND AMALGAMATING 
 MACHINERY. 
 
 Special forms of stamps, . 
 
 Husband, 
 
 Steam, . _ . . 
 
 Elephant, 
 
 Morison's high-speed stamp 
 Huntington mill, 
 Crawford ball mill, . 
 Other ball mills, 
 Tyrolean mill, . . . 
 
 155 ( Schemnitz mill, . 
 
 155 j Lazzlo amalgamator, 
 
 156 5 Pan-amalgamation, . 
 
 158 
 158 
 159 
 163 
 165 
 165 
 
 The old system, 
 The Boss continuous system, 
 Treatment of concentrates, . 
 Molloy's hydrogen amalgamator, 
 Jordan's amalgamator, . 
 
 165 
 166 
 167 
 168 
 171 
 172 
 173 
 174 
 
XIV 
 
 CONTENTS. 
 
 CHAPTER IX. CONCENTRATION IN STAMP MILLS. 
 
 PAGE 
 
 Concentration, . . .175 
 
 Settling boxes, . . .177 
 
 Classification according to size, 178 
 
 Screens, . . . .178 
 
 Pointed boxes, . . .179 
 
 Early concentrating machinery, 180 
 
 Blanket strakes, . . .181 
 
 Riffled sluices, . . .182 
 
 Raising-gate concentrator, . ] 82 
 
 Round buddle, . . . Ib2 
 
 Centrifugal concentrators, . 183 
 
 Hendy s, . . .183 
 
 PAGE 
 
 Duncan's, . . .183 
 
 Percussion tables, . . ] 84 
 
 Gilpin County concentrator, 184 
 Frue vanner, . . .185 
 
 Method of working, . 189 
 
 Riffle belted, . . 192 
 
 Embrey concentrator, . 192 
 
 Luhrig vanner, . .193 
 
 Hartzjigs, . . , 196 
 
 Pneumatic jig, . . 197 
 
 Clarkson and Stanfield's con- 
 centrator, . . . .197 
 
 CHAPTER X. STAMP BATTERY PRACTICE IN PARTICULAR LOCALITIES. 
 
 In California, . . . .199 
 In Colorado, .... 204 
 On free-milling ores in Australia 
 and New Zealand, . . 208 
 
 In the Thames Valley, New 
 
 Zealand, . . . .212 
 In Dakota, . . , . 215 
 In the Transvaal, . . . 216 
 
 CHAPTER XI. CHLORINATION : THE PREPARATION OF ORE FOR 
 TREATMENT. 
 
 The Plattner process, . . 224 
 Origin, . . . .224 
 
 Method of working at Reich- 
 
 enstein, .... 225 
 
 Modern practice in chlorination, 226 
 
 Crushing, . . . .227 
 
 Krom's rolls, . . . 228 
 
 Rolls at Rapid City, Dakota, 231 
 
 Comparison between rolls 
 
 and stamps, . . . 232 
 
 Drying the ore, . . . 233 
 
 Roasting, .... 234 
 
 Reverberatory furnace, . 235 
 
 Chemistry of oxidisingroasting, 238 
 
 Decomposition of various 
 
 minerals, .... 240 
 Elimination of arsenic and 
 antimony, . . .241 
 
 Use of salt, .... 242 
 
 Losses of gold, . . . 244 
 
 Mechanical furnaces, . . 246 
 
 1. With mechanical stirrers, 247 
 
 (a) O'Hara, . . 247 
 
 (b) Spence, . . 248 
 
 (c) Pearce Turret, 249 
 
 (d) Brown horse shoe, 250 
 (e i Ropp straight line, 252 
 
 2. With rotating bed, 252 
 
 3. Revolving cylinders, 255 
 
 (a) Bruckner, . 255 
 
 (b) Hofmann, . * 256 
 
 (c) White, . . 258 
 
 (d) White-Howell, 259 
 Use of producer gas in roast 
 
 ing, . . . . . 260 
 
 CHAPTER XII. CHLORINATION : THE VAT PROCESS. 
 
 Construction of vats, 
 Charging-in, . 
 Generation of chlorine, 
 Impregnation, 
 Reactions in vat, 
 
 261 
 263 
 264 
 266 
 267 
 
 Amount of chlorine required, . 269 
 Leaching, .... 270 
 Precipitation of gold, . . 271 
 Cost of working, . . .273 
 
CONTENTS. 
 
 XV 
 
 CHAPTER XIII. CHLORINATION : THE BARREL PROCESS. 
 
 PAGE 
 
 History, 274 
 
 Mears process, . . . 275 
 Thies process, .... 277 
 Construction of barrel, . 280 
 Charging-in, . . .280 
 Amounts of chemicals re- 
 quired, . . . .282 
 Method of leaching, . . 283 
 Mechanical difficulties in leach- 
 ing, 284 
 
 Leaching by pressure, . . 284 
 
 Riotte's method, . . . 285 
 
 Precipitation of gold, . . 285 
 
 1. Soluble precipitants, . 286 
 
 Ferrous sulphate, . 286 
 
 PAGE 
 
 Organic substances, . 287 
 
 Sulphuretted hydrogen, 287 
 
 Sulphurous acid, . 288 
 
 2. Solid precipitants, . . 289' 
 
 Charcoal, . . .289 
 
 Insoluble sulphides, . 290 
 
 Metals, . . .291 
 
 Modern patent processes of 
 
 chlorination, . . . 292 
 Newbery-Vautin, . . 292 
 Pollok, .... 292 
 Swedish, .... 29& 
 
 Cassel, 295 
 
 Julian, . . . .295 
 
 Greenwood, . . . 295 
 
 CHAPTER XIV. CHLORINATION : 
 
 Vat process 296 
 
 1. Butters' Mill, Kennel, Cali- 
 
 fornia, .... 296 
 
 2. Plymouth Mill, Amador 
 
 Co., California, . . 297 
 
 3. Treadwell Mine, Alaska, 298 
 Barrel process 
 
 1. Mears process at Deloro, 
 
 Canada, . . .298 
 
 2. Thies process in Carolina, 302 
 
 PRACTICE IN PARTICULAR MILLS. 
 
 Cost of working, . . 
 
 3. Modern barrel process at 
 
 the Golden Reward Works, 
 Deadwood, Dakota, . 
 Cost of working, . . 
 
 4. Bromination Mill, Rapid 
 
 City, Dakota, . . 
 The Cassel -Hinman bromine 
 
 process, .. 
 Future of chlorination, 
 
 303 
 310- 
 
 311 
 314 
 
 CHAPTER XV. THE CYANIDE PROCESS. 
 
 History, 319 
 
 Mac Arthur- Forrest process, . 321 
 
 1. Preparation of ore, . .321 
 
 2. Dissolving the gold, . 322 
 Vats, . . . .322 
 Charging-in, . . . 323 
 Separation of slimes, . 324 
 Leaching, . . .324 
 Agitation, . . .325 
 Strength of cyanide solu- 
 tion, . . . .328 
 
 Disposal of tailings, . 330 
 
 3. Precipitation of gold, . 331 
 Cleaning-up, . . . 332 
 
 4. Production of bullion from 
 
 precipitate, . . . 333 
 
 Plant required, . . . 337 
 
 Treatment of ore slimes, . 337 
 
 ,, accumulated 
 
 slimes, . 340 
 Testing of ores, . . .341 
 Direct treatment of Rand ore, 342 
 Use of cyanide in stamp bat- 
 tery, . . . .343 
 Method of treatment 
 
 At the Robinson mine, . 344 
 
 Sylvia mine, . . 344 
 
 ,, New Primrose mine, 34ti 
 
 Siemens-Halske process, . 347 
 
 Results of process 
 
 In South Africa, . . 351 
 ,, other parts of the world, 353- 
 
-xvi 
 
 CONTENTS. 
 
 CHAPTER XVI. CHEMISTRY or THE CYANIDE PROCESS. 
 
 Action of potassium cyanide on 
 
 gold and other metals, . 354 
 Decomposition of potassium 
 
 cyanide, . . . .360 
 Reactions in the zinc boxes, . 361 
 Action of potassium cyanide 
 On metallic salts and minerals, 364 
 On oxidised pyrites, . . 366 
 The soda solution, . . . 368 
 Re-precipitation of gold and 
 
 silver in leaching vats, . 369 
 Testing strength of solution, . 369 
 Estimation of sulphides in 
 
 potassium cyanide, . .371 
 
 PAOE 
 
 Strength of solution required, 372 
 Consumption of cyanide, . 372 
 Methods of increasing the 
 speed of action of potas- 
 sium cyanide, . . . 373 
 The Hood process, . . . 374 
 ,, Sulman-Teed process, . 376 
 ,, Molloy process, . . 377 
 ,, Pelatan-Clerici process, . 377 
 ,, Johnston process, . . 378 
 Wilde process, . . 378 
 Ores suitable to process, . . 378 
 
 CHAPTER XVII. PYRITIC SMELTING, . 
 
 380 
 
 CHAPTER XVIII. THE REFINING AND PARTING OF GOLD BULLION. 
 
 General considerations, . . 383 
 Refining, . . . .384 
 
 Composition of bullion, . 384 
 Melting furnace, . . . 386 
 Crucibles, . . . .387 
 Melting, . . - . .387 
 Refining, .... 388 
 Toughening, . . . 390 
 Casting, . . . .391 
 Losses of bullion, . . . 393 
 Uetining by sulphur, . . 395 
 Osmiridium in gold bars, . 395 
 ^Parting processes, . . . 396 
 Cementation, . . . 397 
 Parting by sulphide of anti- 
 mony, .... 397 
 Parting by sulphur, . . 397 
 Nitric acid process, . . 398 
 
 1. Granulation of alloys, . 398 
 
 2. Dissolving granulations, 399 
 
 3. Treatment of gold resi- 
 
 dues, . . . .400 
 
 4. Treatment of silver solu - 
 
 tion, . . . .401 
 
 5. Reduction of silver 
 
 chloride, . . . 401 
 Parting by sulphuric acid, . 402 
 
 1. Mixing and granulating 
 
 alloys, . . . 402 
 
 2. Dissolving silver, . . 403 
 
 3. Melting gold residue, . 406 
 
 4. Precipitation of silver, . 407 
 
 5. Crystallisation of sul- 
 
 phate of copper, . 407 
 Combined process, . . 408 
 Gutzkow process, ; . 409 
 New Gutzkow process, . 411 
 
 Miller's chlorine process, . 414 
 Original method, . -415 
 Later improvements, . 417 
 Modern practice at Mel- 
 bourne, . . -420 
 Modern practice at Sydney, 429 
 Refining brittle gold by 
 
 chlorine, . . .431 
 Parting by electrolysis, . 432 
 Parting of gold from plati- 
 num, iridium. &c., . 435 
 
 CHAPTER XIX. THE ASSAY OF GOLD ORES. 
 
 General considerations, . 
 Blowpipe assay, 
 Sampling and crushing the ore 
 
 Metallics, . 
 
 Crucible method of assay, 
 Fusion, 
 
 Assay-ton weights, . 
 General charges, 
 
 437 
 437 
 438 
 439 
 440i 
 440 
 441 
 441 
 
 Fluxes, . 
 Methods of operation, 
 Roasting before fusion, 
 Cleaning slag, . 
 Treatment of base ores, 
 Cupellation, 
 
 Influence of base metals, 
 
 442 
 444 
 447 
 448 
 448 
 448 
 451 
 
CONTENTS. 
 
 XV11 
 
 CHAPTER XIX. Continued. 
 
 Inquartation and parting, . 452 
 
 Examination of assay 
 
 materials, . . . 454 
 
 Examination of cupel, . . 454 
 
 Assay by scorification, . . 455 
 
 Detection of gold in minerals, . 457 
 Estimation of gold in dilute 
 
 solution, .... 458 
 
 Special methods of assay, . 458 
 
 1. Mixed wet and dry method, 458 
 
 PAGE 
 
 459 
 459 
 461 
 461 
 461 
 461 
 
 6. Assay of purple of Cassius, 462 
 
 7. ,, a Mint sweep, . 462 
 
 2. Amalgamation, 
 
 3. Chlorination, . 
 
 4. Whitehead's method, 
 
 5. Assay of pyrites, 
 
 (a) Swartz's method, 
 
 (b) Staptf's 
 
 CHAPTER XX. THE ASSAY OF GOLD BULLION. 
 
 (Parting assay, .... 462 
 
 1. Selection of sample, . 463 
 
 2. Preparation of assay-piece 
 
 for cupellation, . . 464 
 
 3. Cupellation, . . .467 
 Assay furnace, . . . 467 
 Cupels, . . . .469 
 Method of operation, . 469 
 Flashing, . . . .470 
 Temperature of muffle, . 471 
 
 4. Preparation of cupelled 
 
 buttons for parting, . 474 
 
 5. Parting, . . . .475 
 
 (a) In parting flasks, . 475 
 
 (b) In platinum trays, . 476 
 Relative advantages of 
 
 these, . . .477 
 
 6. Weighing the cornets, . 478 
 Losses of gold in bullion assay- 
 ing, 478 
 
 Silver retained in cornets, . 480 
 Occluded gases, . . .481 
 Checks or proofs, . . .481 
 Limits of accuracy in assay, . 482 
 Parting by sulphuric acid, . 482 
 Preliminary assay, . . . 433 
 Assay by cadmium,. . . 433 
 
 Assay of alloys of gold, silver, 
 
 and copper, . . . 483 
 Effects of other metals, . . 484 
 Assay of various gold alloys, . 484 
 
 A. Alloys requiring scorifi- 
 
 cation, . . 484 
 
 Arsenic and antimony, . 484 
 Iron and manganese, . 485 
 Cobalt and nickel, . . 485 
 Zinc, . . . .485 
 Tin, . . . . .485 
 Aluminium, . . . 485 
 
 B. Amalgams, . . . 486 
 
 C. Platinum group, . . 486 
 
 (1) Platinum, . . .486 
 
 (2) Palladium, . . .487 
 
 (3) Rhodium and iridium, 487 
 
 D. Tellurium compounds, . 488 
 Wet methods of assay, . . 488 
 Other methods, . . .489 
 
 1. Touchstone, . . . 489 
 
 2. Colour and hardness, . 490 
 
 3. Density, . . . .490 
 
 4. Spectroscope, . . . 490 
 
 5. Electrolysis, . . . 491 
 
 6. Induction balance, . .491 
 Auxiliary balance for weighing 
 
 cornets, .... 492 
 
 CHAPTER XXI. ECONOMIC CONSIDERATIONS. 
 
 Management of gold mills, . 493 
 
 ost of production of gold, . 493 
 Annual production of gold, past 
 
 and present, . . . 495 
 
 Table of production in different 
 
 countries, . . . 498 
 
 Amount produced by different 
 
 processes, . . . 500 
 
 Consumption of gold, 
 
 501 
 
XVlll 
 
 CONTENTS. 
 
 BIBLIOGRAPHY. 
 
 PAGE 
 
 Periodicals, .... 503 
 
 General, 504 
 
 Properties of gold and its alloys, 505 
 
 Distribution of gold, . . 507 
 
 Placer mining, . . . 508 
 
 Crushing and amalgamation, . 509 
 
 Concentration, . . . 510 
 
 PAGE 
 
 Roasting 511 
 
 Chlorination, . . . .511 
 Cyanide process, . . .512 
 Pyritic smelting, . . . 512 
 Refining and parting of bullion, 512 
 Assaying, . . . .513 
 
 INDEX, 
 
 515 
 
THE METALLURGY OF GOLD. 
 
 CHAPTER I. 
 THE PROPERTIES OF GOLD AND ITS ALLOYS. 
 
 PHYSICAL AND CHEMICAL PROPERTIES OF GOLD. 
 
 Introduction. From very early times the ancients were 
 attracted by the beautiful colour, the brilliant lustre, and the 
 indestructibility of gold, and spared no pains in the endeavour 
 to acquire it. In the code of Menes, who reigned in Egypt in 
 3600 B.C. or about 2000 years before Moses, the ratio of value 
 between gold and silver is mentioned, one part of gold being 
 declared equal in value to two and a half parts of silver, and it 
 is, therefore, clear that the extraction of both metals from the 
 deposits containing them must have been carried on before that 
 time. It is, indeed, probable that gold was the first metal 
 observed and collected, since it occurs in fragments of all sizes 
 in loose sand, and the operations of collecting the larger pieces 
 and melting them together are so simple. Among the rock 
 carvings of Upper Egypt there are several illustrative of the art 
 of washing auriferous sands by stirring and working them up by 
 the hand in hollowed-out stone basins, and subsequently melting 
 the gold in simple furnaces with the aid of mouth blow-pipes. 
 The earliest of these carvings is supposed to date back to about 
 2500 B.C. However, in ancient times gold appears to have been 
 mainly derived from India, and that country continued to supply 
 most of the gold used in Europe until the discovery of America 
 by Columbus. 
 
 In order to collect alluvial gold, the sands were washed down 
 over smooth sloping rocks by means of running water, and the 
 particles of gold, sinking to the bottom of the stream by reason 
 of their high density, were entangled and caught in the hair of 
 raw hides spread on the rocks. Among the hides used were 
 
 1 
 
THE METALLURGY OP GOLD. 
 
 sheepskins, and hence originated the form of the legend of the 
 Golden Fleece. Stripped of its heroic dress, this legend merely 
 describes a successful piratical expedition about 1200 B.C. to win 
 gold, which was being laboriously obtained from streams with 
 the help of sheepskins by the inhabitants of what is now 
 Armenia. Similar expeditions have not been unknown in much 
 later times, and the method of obtaining gold by washing river 
 sands is still practised, with improvements in matters of detail, 
 in many parts of the world. Metallurgists are almost pro- 
 verbially conservative in their methods. Hides are even now 
 occasionally employed to catch the gold, but sheepswool, when 
 used, is generally in the form of blankets. 
 
 At the present day, however, when auriferous sands are 
 washed, the aid also is invoked of what Baron Born called in 
 1786 the "elective affinity" of mercury for gold when mixed 
 with impurities. The ease with which gold-amalgam can be 
 collected, in spite of its being less dense than gold itself, is due 
 to the fact that it is miscible in all proportions with mercury, so 
 that, under proper conditions, large globules of liquid alloy are 
 formed by the running-together of smaller particles, and the 
 former are readily caught in suitable crevices. 
 
 In the history of gold, it is also of interest to the metallurgist 
 to remember that the earliest dawn of the science of chemistry 
 was heralded by the study of the properties of gold, and by the 
 efforts which were made to invest other matter with these 
 properties. From the fourth to the fifteenth century, chemistry, 
 which was first called "chemia" (x^sia), and then "alchemy," 
 was defined as the art of transmuting base metals into gold and 
 silver, almost all the labours of philosophers being intended to 
 aid directly or indirectly in solving this problem. At the end 
 of this period, while Paracelsus was giving to chemistry a new 
 aim that of investigating the composition of drugs, and their 
 effect on the human body Agricola was reducing to order the 
 numerous empirical facts which together made up the art of 
 metallurgy, and although alchemy died hard, its era of usefulness 
 may be said to have ended here. Gold has doubtless been the 
 cause of many of the wars and marauding expeditions from 
 which the world has suffered, but on the other hand it has been 
 instrumental, in a far greater degree than most other commodities, 
 in promoting the growth of civilisation, the efforts of the 
 alchemists having laid the foundations of the science of chemistry, 
 and those of the gold-seekers having resulted in the discovery of 
 new countries, and in the spread of knowledge of all kinds. 
 
 Colour. The lustre and fine colour of gold have given rise 
 to most of the words which are used to denote it in different 
 languages. The word "gold" is probably connected with the 
 Sanscrit word "jvalita," which is derived from the verb "jval" 
 to shine. It is the only metal which has a yellow colour when 
 
TiiE PJROPEKFIES OF GOLD. 3 
 
 in mass and in a state of purity. Impurities greatly modify this 
 colour, small quantities of silver lowering the tint, while copper 
 raises it. In a finely divided state, when prepared by volatilisa- 
 tion or precipitation, gold assumes various colours, such as deep 
 violet, ruby and reddish-purple, the tint varying to brownish- 
 purple and thence to dark brown and black. This purple colour 
 has been supposed by some experimenters (viz., Guy ton de 
 Morveau, Biichner, Desmarest, Creuzbourg and Berzelius) to be 
 due to the formation of a coloured oxide of gold of unknown com- 
 position, but Buisson, Proust, Figuier and, more recently, Kriiss 
 have shown that no oxygen can be obtained from this coloured 
 material, and that it probably consists of metallic gold. Similar 
 colours are seen in purple of Cassius, and in Roberts- Austen's 
 purple alloy of aluminium and gold, the colour in each case being 
 probably due to a particular form of finely divided gold. The 
 ruby coloured gold, suspended in a liquid in which it has been 
 precipitated by ether and phosphorus, is present in the most 
 finely divided state obtainable, and perhaps it then approaches 
 the condition of separate atoms. Such gold shows no tendency 
 to settle by gravity even after being kept undisturbed for several 
 years. Somewhat less finely divided gold gives a faint blue tinge 
 to light transmitted through the liquid in which it is suspended. 
 Gold precipitated from its solution as bromide is in a different 
 molecular condition from that formed from chloride, giving out 
 3-2 calories in passing into the latter state.* The surface colour 
 of small particles of native gold is often apparently reddened by 
 being coated with translucent films of oxides of iron. Very thin 
 plates of gold are translucent, and appear green by transmitted 
 light, while remaining yellow by reflected light. On heating, 
 the green colour changes to ruby red, but is restored by the 
 pressure of a hard substance by which the state of aggregation 
 is again altered (Faraday). Molten gold is green, and its vapour 
 is also probably greenish. 
 
 Malleability and Ductility. Malleability and ductility 
 are possessed by gold at all temperatures to a far higher 
 degree than by any other metal. A single grain of gold can 
 be drawn out into a wire over 500 feet long, and leaves of 
 not more than -g-Q ^ Q Q of an inch in thickness can be obtained 
 by beating. Faraday has shown that the thickness of these 
 leaves may be still further reduced by floating them in a 
 dilute solution of potassic cyanide by which they are partly 
 dissolved. 
 
 Hardness. The hardness of gold lies between that of alu- 
 minium and that of silver, corresponding to the number 979 in 
 Bottone's scale, in which the diamond is 3,010. 
 
 Tenacity. The purest gold obtainable has a tenacity of 7 tons 
 
 * Thomsen, Thermochemische Untersuchungen, vol. iii., p. 412. 
 
4 THE METALLURGY OF GOLD. 
 
 per square inch and an elongation of 30-8 per cent., but the 
 presence of as little as ^rnnj" ^ other elements, especially 
 bismuth, tellurium, lead, and other metals with high atomic 
 volumes, greatly lowers these constants, as well as the malle- 
 ability and ductility of the metal, while its hardness is 
 increased.* Gold containing ^^ of bismuth can almost be 
 crumbled in the fingers. 
 
 Specific Gravity. The specific gravity of gold when pre 
 cipitated from solution by oxalic acid is 19-49 (G. Rose) ; f 
 when cast it varies from 19*29 to 19-37, but this can be raised 
 by compression to over 19-48.| Henry Louis has shown that 
 the specific gravity of unannealed "parted" gold (i.e., the 
 residue left after boiling silver-gold alloys in nitric acid) is 20-3, 
 its density being lowered by the process of annealing. When 
 precipitated by ferrous sulphate, its density may be as high as 
 20-72 (G. Rose). 
 
 Cohesion. On heating, gold can be welded like iron below 
 the point of fusion, and finely divided gold agglomerates on 
 heating without being subjected to pressure. Pressure alone is 
 also sufficient to make gold dust cohere, while a true flow of the 
 particles of gold can be induced in the case of the pure metal 
 and some of its alloys. 
 
 Specific Heat. The specific heat of gold is -0324 (Regnault) 
 or -0316 (Yiolle). 
 
 Fusibility. Gold fuses, after passing through a pasty stage, 
 at a clear cherry-red heat, just below the fusing point of copper 
 and much above that of silver. The metal expands considerably 
 on fusing and contracts again on solidifying. Carnelley gives 
 the temperature of fusion as 1,037, Yiolle as 1,045, whilst 
 recent determinations have given it as 1,061-7 (Hey cock and 
 Neville), 1,050 to 1,060 (Le Chatelier), and 1,072 (Holborn 
 and Wien). 
 
 Spectrum. In the gold spectrum Huggins saw 23 lines, the 
 wave lengths of the most important ones being 523-1, 583*5, and 
 627-6 respectively.)) 
 
 Latent Heat. The latent heat of fusion of gold is 16-3, and 
 the normal lowering of the freezing point for 1 atom of impurity 
 in 100 atoms of gold is 10 -6, but the presence of from O'l to 
 4*0 per cent, silver does not cause any alteration in the freezing, 
 point. 
 
 Magnetism. Gold is diamagnetic, its specific magnetism 
 being 3-47 (Becquerel), if that of iron is taken as 100. 
 
 * Roberts- Austen, Phil. Trans. Royal Soc., vol. clxxix. (1888), p. 339. 
 
 t Pogg. Ann., vol. Ixxiii. (1848), p. 1, and vol. Ixxv. (1848), p. 403. 
 
 t Eighth Report of the Royal Mint, 1877, p. 43. 
 
 Trans. Am. Inst. of<Mng. Eng., Chicago Meeting, 1893. 
 
 || For full information on the spectroscopic characteristics of gold, see- 
 Lockyer and Roberts, Phil. Trans. Royal Soc., vol. clxiv. (1874), part ii. r 
 p. 495; and Fre"my, Ency. Chim., vol. iii. (1888), L'or, p. 40. 
 
UNIVERSITY 
 
 THE PROPERTIES OF GOLIX" 
 
 Conductivity and Expansion. Its electrical conductivity 
 is 76-7, and its thermal conductivity 60 (or, according to 
 Despretz, 103), that of silver being 100 in each case. Its co- 
 efficient of linear expansion is ,0*0000 144 between and 100. 
 
 Atomic Weight and Volume. Its atomic weight, formerly 
 believed to be about 196*2, has been more recently given by 
 Kriiss as 196*64, by Thorpe and Laurie as 196*85, and by Mallet 
 as 196*79. It is probable that 196*8 is not far from the truth. 
 The atomic volume of gold is 10*2.* 
 
 Volatilisation of Gold. The boiling point of pure gold has 
 not been determined ; calculated according to Wiebe's formula it 
 would be about 2,240, or nearly 500 above the melting point of 
 platinum.f However, contrary to the belief of the older experi- 
 menters (Gasto Claveus, and others), it is sensibly volatile in air 
 at far lower temperatures. Robert Boyle was unaware of this 
 fact, but Homburg gilded a silver plate in 1709 by holding it over 
 gold strongly heated in the focus of a burning mirror (Encyclo- 
 paedia Britannica, 1778, and Gmelin's Nandbuch), and St. Claire 
 Deville volatilised and again condensed gold when melting it 
 with platinum. The rapid volatilisation of gold, when heated by 
 an ordinary blowpipe, was first proved in 1802 by Dr. Robert 
 Hare, of Philadelphia (Tilloch's Magazine), a purple stain being 
 thus produced on bone-ash in a few seconds. Lastly, a discharge 
 of high-tension electricity from gold points causes its volatilisa- 
 tion, and if the discharge is sent through a fine gold wire 
 stretched on paper, it converts it into a purple streak of finely 
 divided condensed particles of the metal. The rapid distillation 
 of gold caused by heating it in a current of air of considerable 
 velocity, such as that furnished by a blow-pipe, by which the 
 liquid is thrown into waves, may be shown at any time by 
 heating a fragment of the precious metal of the size of a pin's 
 head 011 a bone-ash cupel in the oxidising flame of a good mouth 
 blowpipe. Almost immediately after the fusion is complete, a 
 purple stain of condensed gold begins to form on the outer 
 margin of the cupel. The author has found that a piece of gold 
 weighing 0'5 gramme loses half its weight in an hour, if heated 
 on a cupel by a foot^blowpipe (the temperature attained being 
 probably less than 1,300), and only a few minute beads are 
 observable, detached from the main button. Alloys of copper 
 and gold disappear much more rapidly. No doubt most of the 
 gold passes off as spray, but perhaps part of the loss may be due 
 to rapid volatilisation, and could not be correctly described as 
 mechanical loss. 
 
 The volatility of gold, both when pure and when alloyed with 
 silver and copper, has been investigated by Napier, J who found 
 
 * Introd. to the. Study of Metallurgy, by Prof. Roberts- Austen, p. 58. 
 
 t L. Meyer, Mod. Theories of Chemistry, p. 134. 
 
 t Chan. &oc. Journ. t vol. x. (1859), p. 229; and vol. xi., p. 168. 
 
D THE METALLURGY OF GOLD. 
 
 that an alloy of 100 parts gold to 12 parts of copper, if kept for 
 six hours at a temperature just high enough to keep it melted, 
 lost 0-234 per cent, of its gold contents, and at the highest tem- 
 perature attainable in an assay muffle, it lost 0'8 per cent, in six 
 hours. An increase in the amount of copper present caused an 
 increase in the loss of gold. In the simple operation of pouring 
 about 30 Ibs. of a gold-copper alloy from a graphite crucible into 
 moulds, fumes were given off, of which the part condensed in a 
 wet glass beaker held above the crucible contained 4 '5 grains of 
 gold. Napier also found that gold does not appear to volatilise 
 so readily when alloyed with silver only, as when copper is 
 also present. 
 
 Makins found that gold volatilises sensibly along with silver 
 and lead, when melted with these metals in a muffle in an 
 ordinary bullion assay. The loss of gold by volatilisation on 
 melting its copper alloy is the common experience of mints. At 
 the Sydney Mint it was estimated to be 0*017 per cent., or 170 
 per million sterling melted, and is probably seldom less than 
 O'Ol per cent., or one part in ten thousand. At the same 
 Mint, Leibius found that the sweepings from the top coping-stone 
 of a chimney 70 feet high contained 1'46 per cent, of gold and 
 6-06 per cent, of silver. Similar results have been obtained at 
 some other mints. 
 
 A number of experiments were recently made at the Royal 
 Mint by the author,* with the view of determining the effect of 
 variations in the temperature and other conditions on the vola- 
 tility of gold and some of its alloys. The test pieces were heated 
 in a muffle furnace, the temperature of which was determined by 
 the optical pyrometer, and by the Le Chatelier thermo-couple, 
 which consists of platinum and rhodio-platinum wires. The 
 results of some of the experiments on fine gold, and certain 
 alloys of gold and copper, are given in the table on the next page. 
 
 The following conclusions may be drawn from the table : 
 
 1. The loss of gold on heating the pure metal rises with the 
 temperature, being four times as great at 1,250 as at 1,100, 
 whilst it is insignificant at 1,075 and probably nil at 1,045, the 
 melting point (Violle). 
 
 2. An atmosphere consisting largely of carbonic oxide is 
 apparently favourable to the volatilisation of gold, the rate being 
 six times greater than in an atmosphere of coal gas at the same 
 temperature. The increased loss when the graphite crucible was 
 substituted for a cupel may be noticed in this connection ; clay 
 crucibles of similar shape to the graphite one have an opposite 
 effect, the loss being less than on cupels. 
 
 3. The increase of loss of gold alloyed with copper, when the 
 percentage of the latter metal is increased, which is observable 
 in one of Napier's experiments, is confirmed. 
 
 * Chem. Soc. Journ., vol. Ixiii. (1893), p. 714. 
 
TIIE PROPERTIES OF GOLD. 
 
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8 THE METALLURGY OF GOLD. 
 
 The volatilisation of gold-copper alloys merits further in- 
 vestigation, as it is of great industrial importance. The high 
 percentage of the losses shown in the table is, no doubt, due to 
 the smallness of the masses treated, and to the comparatively 
 large surfaces exposed in consequence of this. 
 
 A number of experiments were also made at the same time on 
 various gold alloys to determine the relative effect of the presence 
 of other metals on the volatilisation of gold. The temperatures 
 and conditions used were similar to those described above. It 
 was found that the volatility of gold was increased by the pre- 
 sence of any metallic impurity, even by the non-volatile metals, 
 such as platinum. The tellurium alloys suffered the heaviest 
 losses, gold containing five per cent, of tellurium losing from 
 16*5 to 39*5 parts per 1,000 by volatilisation per hour at a 
 temperature of 1,245 (i.e., 200 above the melting point of pure 
 gold), a point of much importance in connection with the metal- 
 lurgy of telluric gold. Lead and platinum had a very slight 
 effect in increasing the volatility of gold ; copper and zinc a more 
 marked effect, while five per cent, of antimony or mercury 
 caused losses amounting to about two parts per 1,000 of gold per 
 hour at 1,245. Further conclusions which were drawn were as 
 follows : 
 
 1. Although the temperatures employed were higher than 
 those at which pure zinc, cadmium and tellurium, and probably 
 antimony and bismuth, are distilled, yet these elements were 
 never completely volatilised. The " temperature of dissociation 
 of the alloys," as in the case of the bismuth-arsenic alloy 
 investigated by Edward Matthey in a paper read before the 
 Royal Society, January 26, 1893, is much higher than the 
 ordinary distillation point of the more volatile constituent. 
 Thus zinc boils at 950, but its gold alloy loses little or no zinc 
 at 1,120, and still retains part at 1,250, whilst antimony was 
 not driven off to any extent by the highest temperature attained. 
 Copper appears to pass off more easily than some of the so-called 
 volatile metals, perhaps because the dissociation of its alloy with 
 gold may not form an initial stage of the operation. 
 
 2. The amount of gold lost depends partly on the volatility of 
 the alloying metal, but perhaps too much stress has been laid on 
 this factor in the past, the results of heating alloys of mercury, 
 zinc, antimony and copper pointing to that conclusion. A metal 
 with a strong attraction for gold, such as copper, may carry it 
 off, perhaps as a vaporised alloy, more easily than one which 
 mixes with it less intimately. 
 
 3. It is observable that those impurities which reduce the 
 surface tension of a button of liquid gold appear to increase the 
 vapour pressure of the metal as indicated by the loss on heating. 
 This was to be expected 
 
 4. A current of air or coal gas insufficient to disturb the 
 
THE PROPERTIES OF GOLD. 9 
 
 surface of the liquid metal does not appear to increase the 
 volatilisation. 
 
 It may be added that Hellot stated that if an alloy of one 
 part of gold and seven parts of zinc is heated in air, the whole 
 of the gold comes off in the fumes. 
 
 Crystallisation of Gold. Gold crystallises in the cubic 
 system, occurring frequently in nature in the form of cubes, 
 octahedra and rhombic dodecahedra. Cleavage is never exhibited. 
 Single detached crystals are comparatively rare, and the crystals 
 are usually attached end to end, forming strings, and branching, 
 arborescent, or moss-like masses, which are composed of micro- 
 
 Fig. 3. 
 
 Fig. 2. 
 
 scopic crystals, usually octahedra. These forms occur frequently 
 in quartz veins, but the single crystals, which are usually of larger 
 size, viz., from inch to 1J inch in diameter, are mainly found 
 in drift deposits. They are rarely perfect or of brilliant lustre, 
 although such crystals were found at the Princeton Gold Mine, 
 Mariposa County, California, but occur more frequently with 
 rounded angles, raised edges, and cavernous faces, which are often 
 marked with parallel striations, and possess little or no lustre 
 (Fig. 1). The octahedra found in California are usually flattened 
 parallel to two opposite faces, or elongated, or otherwise dis- 
 torted. Still more frequently they are only partially developed, 
 as in Figs. 2 and 3. In all these cases " the incomplete crystals 
 
10 THE METALLURGY OF GOLD. 
 
 have the appearance of a failure for lack of material " (W. P. 
 Blake). Crystals of greater complexity, containing many modi- 
 fying faces, occur chiefly in Siberia, Transylvania and Brazil. 
 The most common forms occurring naturally in Australia are 
 the octahedron and the rhombic dodecahedron.* 
 
 Artificial crystals can be obtained in several ways, but with 
 great difficulty. The slow cooling of an ingot of gold from fusion 
 usually gives faces and sometimes angles of octahedra on the 
 surface of the metal, f The presence of small quantities of copper 
 prevents this crystallisation. Feathery crystalline plates are 
 precipitated in the electrolysis of a solution of chloride of gold 
 and ammonium. By keeping an amalgam containing 5 per cent. 
 of gold at a temperature of 80 for eight days, and then di- 
 gesting it at 80 with nitric acid of specific gravity 1'35, 
 and subsequently subjecting the residue to a red heat, bright 
 crystals of gold which are, however, usually microscopic can be 
 obtained. | In the Percy collection are some gold crystals found 
 in the mercury troughs at the foot of the " blanket strakes," in an 
 amalgamation mill. The troughs are placed so as to catch any 
 stray particles of gold that may pass the blankets. As the 
 amount of gold recovered in this way is very small, it is not 
 worth while to clean out the troughs frequently, and in this case 
 they had remained undisturbed for nine months, at the end of 
 which time all the amalgam was found to be crystallised. The 
 mercury has been dissolved off by nitric acid, and the gold 
 crystals remain. The smaller crystals are rather indefinite in 
 shape, but amongst the larger ones (which are about half the 
 size of a pea) are well-defined combinations of the octahedron, 
 rhombic dodecahedron, and cube. 
 
 Solubility of Gold. Gold is readily soluble in aqua regia, 
 or in any other mixture producing nascent chlorine, among such 
 mixtures being solutions of (1) nitrates, chlorides, and sulphates 
 e.g., bisulphate of soda, nitrate of soda, and common salt; 
 (2) chlorides and some sulphates e.g., ferric sulphate ; (3) hy- 
 drochloric acid and potassium chlorate ; (4) bleaching powder 
 and acids, or salts such as bicarbonate of soda. The action 
 is much more rapid if heat is applied or if the gold is alloyed 
 with one of the base inetals than if it is pure. The presence of 
 silver in the gold retards the process, a scale of insoluble chloride 
 of silver being formed over the metal, and the action may even- 
 tually be completely stopped if the percentage of silver present 
 is large. Gold is also dissolved by liquids containing chlorine 
 and bromine, but the action is much slower than that of aqua 
 regia; and subject to the same difficulties if silver is present; 
 
 * For a full account of the crystalline forms of native gold, see a Paper 
 by W. P. Blake, in Precious Metals of the U.S.A., 1884, p. 573. 
 
 t Chester in Am. Journ. of Science and Arts, vol. xvi., July, 1878, p. 29. 
 Krafft, Encyclopaedia. Brit., article "Crystallisation." 
 
THE PROPERTIES OP GOLD. 11 
 
 heat assists the dissolution. Iodine only dissolves gold if it is 
 nascent, or if heated with gold and water in a sealed tube to 50. 
 Metallic gold dissolves in hot strong sulphuric acid, especially 
 if a little nitric acid is added (the precipitated metal dissolving 
 most readily), forming a yellow liquid, which, when diluted with 
 water, deposits the metal as a violet or brown powder. The 
 solution also becomes covered with a shining film of reduced 
 metal on. exposure to moist air. On addition of hydrochloric 
 acid or a metallic chloride, auric chloride is formed, no longer 
 precipitable by water.* Gold is also attacked when used as the 
 positive pole of the battery in the electrolysis of strong sulphuric 
 acid, but is immediately reduced again by the evolved hydrogen, f 
 According to Nickles, J the easily decomposable metallic per- 
 chlorides, perbromides and periodidep are capable of dissolving 
 gold, lower chlorides, <fec., of the base metals being formed, and 
 gold chloride, &c., produced. The higher chlorides and bromides 
 of manganese and cobalt (Co 2 Cl 6 , &c.) act well, and a hot strong 
 aqueous solution of ferric bromide or of ferric chloride will also 
 readily dissolve gold. Ferric iodide is decomposed by gold 
 under ordinary conditions, aurous iodide being produced. Some 
 other haloid compounds only attack gold in the presence of 
 ether, in which case even hydriodic acid itself is decomposed 
 and aurous iodide formed. Selenic and iodic acids have also 
 been mentioned as solvents for gold, and the effect of a mixture 
 of nitric and nitrous acids is described in Chap. xx. Alkaline 
 sulphides attack gold slowly in the cold, and more rapidly if 
 heated, producing sulphide of gold which is subsequently dis- 
 solved. Ditte denies this. Gold is also soluble in the hypo- 
 sulphites of calcium, sodium, potassium, and magnesium. 
 
 Spring has shown that gold is soluble in hydrochloric acid if 
 heated with it to 150 in a closed tube, and is subsequently 
 reduced by the liberated hydrogen and deposited as microscopic 
 crystals on the side of the tube. C. Lossen pointed out in 1895 
 that if a solution of potassium bromide is electrolysed, the re- 
 sulting alkaline solution, containing hypobromite and bromate 
 of potassium, is capable of dissolving gold. Gold is dissolved 
 by aqueous solutions of simple cyanides and by certain double 
 cyanides, such as sulphocyanides and ferrocyanides, which act 
 very slowly except in presence of oxidising agents and with the 
 aid of heat. 
 
 Preparation of Pure Gold. The purest gold obtainable is 
 required for use as standards or check pieces in the assay of 
 gold bullion. The following method of preparing it was adopted 
 by Roberts- Austen in the manufacture of the Trial-Plate, by 
 
 * Watt's Diet. o/Chem., Supplement, p. 652. 
 
 t Spiller, Chem. Newt, vol. x., p. 178. 
 
 J Ann. Ch. Phy*. [4], vol. x. , p. 318. 
 
 Zeitschr. anoryun. Chem., vol. i. (1S93), p. 240. 
 
12 THE METALLURGY OF GOLD. 
 
 which the imperial gold coinage is tested.* Gold assay cornets, 
 irom the purest gold which can be obtained, are dissolved 
 in nitrohydrochloric acid, the excess of acid expelled, and 
 alcohol and chloride of potassium added to precipitate traces of 
 platinum. The chloride of gold is then dissolved in distilled 
 water in the proportion of about half an ounce of the metal to 
 one gallon, and the solution allowed to stand for three weeks. 
 At the end of this time the whole of the precipitated silver 
 chloride will have subsided to the bottom, and the supernatant 
 liquid is removed by a glass siphon. Crystals of oxalic acid 
 are then added from time to time, and the liquid gently warmed 
 until it becomes colourless, when precipitation is complete, a 
 point reached in three or four days if ten-gallon vessels are used. 
 The spongy and scaly gold so obtained is washed repeatedly 
 with hydrochloric acid, distilled water, ammonia, and distilled 
 water again, until no reaction for silver or chlorine can be ob- 
 tained, after which it is melted in a clay crucible with bisulphate 
 of potash and borax, and poured into a stone mould. Lack of 
 care in any one of the operations will result in gold containing 
 one or two parts of impurity in ten thousand. 
 
 With regard to the above method, it may be observed that 
 carefully purified sulphurous acid gas is a more convenient pre- 
 cipitant than oxalic acid, and may be substituted for it without 
 any ill effects, as any foreign metals that may be present are in 
 such small quantities that their sulphites, even if formed, would 
 remain dissolved. It should be added that, according to the 
 recent researches of Kohlrausch, Rose, and Holleman, silver 
 chloride is soluble in 600,000 to 700,000 parts of pure water at 
 the ordinary temperature. It follows that, under the condi- 
 tions given by Roberts-Austen, the solution contains at least 
 0-3 to 0-4 part of silver per 1,000 parts of gold, and this pro- 
 portion is doubtless higher in practice, owing to the greater 
 solubility of silver chloride in solutions of other chlorides than 
 in pure water. There can be no doubt that part of this silver 
 is precipitated with the gold, and that, by redissolving and 
 reprecipitating, a purer product is obtained The amount of 
 silver remaining in solution can, moreover, be reduced to about 
 one-fifth of the amount noted above by adding a small quantity 
 of hydrobromic acid to the solution, silver bromide being far 
 less soluble than silver chloride. Another additional precaution 
 is to remove the gases taken up by the gold during the process 
 of melting by heating it to redness in vacua. 
 
 Allotropic Forms of Gold. Little is known of these. The 
 marked influence of traces of other metals on the properties of 
 gold has already been touched on; from this and from the 
 variations in colour and other properties the existence of 
 several allotropic modifications of gold might be inferred. In 
 * Fourth Annual Report of the Mint, 1873, p. 46. 
 
ALLOYS OP GOLD. 13 
 
 alloys containing appreciable quantities of other metals, 
 evidences of allotropy are not met with so frequently. The 
 potassium alloy, however, containing 10 per cent, of gold, on 
 being attacked by water, leaves a black finely-divided gold 
 powder, and there is reason to believe that this combines with 
 water to form a hydrate.* 
 
 Wilm f states that if gold is dissolved in dilute sodium 
 amalgam under water, the aqueous liquid becomes dark violet, 
 and when this is acidulated with hydrochloric acid, a black pre- 
 cipitate of pure gold is obtained. The black gold differs from 
 the ordinary modifications in its extreme lightness ; moreover, 
 it is soluble in alkaline solutions, and does not amalgamate with 
 mercury or with sodium amalgam. When heated, it yields the 
 ordinary modification as a violet red powder. This form of gold 
 appears, from Wilm's account, to resemble the black precipitate 
 obtained on digesting certain aluminium-gold alloys with hydro- 
 chloric acid. 
 
 ALLOYS OP GOLD. 
 
 Gold can be made to alloy with almost all other metals, but 
 most of the bodies thus formed are of little or no practical 
 importance. Tin, zinc, arsenic and antimony unite with gold 
 with contraction, and form pale yellow or grey coloured, hard, 
 brittle and easily fusible alloys, of which all, except those con- 
 taining zinc, are soluble with difficulty in aqua regia. The 
 arsenic and antimony alloys are slowly decomposed by mercury, 
 the base metal being separated as a black powder, which con- 
 sists in part of arsenide or antimonide of mercury. Lead and 
 iron alloy with gold with expansion, while in the case of copper 
 no change of volume takes place. 
 
 Gold alloyed with a small percentage of lead is a hard, brittle, 
 pale-yellow substance, which can be crumbled with the fingers. 
 If more than about 4 per cent, of lead is present, there is 
 marked segregation on solidification, and this also takes place 
 in the case of the zinc alloy and of some others. 
 
 Heycock and Neville have shown J that the freezing point of 
 lead is lowered by the addition of gold to it in accordance with 
 the general law. Thus, the freezing point of pure lead being 
 327, an addition of 3*8 per cent, of gold reduces it to 301, and 
 Roberts-Austen has recently found that the eutectic alloy of 
 gold and lead, which contains about 13 per cent, of the former 
 metal, melts somewhere between 190 and 198. Similarly, by 
 adding 6-9 per cent, of gold to thallium, the freezing point of 
 the latter is lowered from 301 to 261. 
 
 * Introd. to Study of Met., p. 91. 
 
 t Zeitschr. anorgan. Chem., vol. iv. (1893), p. 325. 
 
 ZChem. Soc. Journ., vol. Ixv. (1894), p. 72. 
 
14 THE METALLURGY OF GOLD. 
 
 Several alloys of gold with other metals in molecular proportions 
 have been isolated, or their existence proved in various ways. 
 Thus, for example, the compound AuSn was recognised by Ma- 
 thiessen, from the curve of electric conductivity of the gold-tin 
 series of alloys, and this substance was more recently detected 
 by A. P. Laurie by observing the electro-motive force developed 
 by alloys of different compositions when dipped into a solution 
 of SnCl 2 . Heycock and Neville succeeded in isolating the com- 
 pound AuCd in 1892, after having suspected its existence for 
 some time. In the same year, Mylius and Fronim prepared a 
 number of gold alloys by precipitation from solution.* Thus 
 a gold-zinc alloy, containing equal weights of the two metals, 
 and approximating in composition to AuZn 3 , was obtained in 
 the form of black, spongy flocks, by adding a solution of gold 
 sulphate to water in which a zinc plate was placed. The gold- 
 zinc slimes obtained in the cyanide process (q.v.} may be com- 
 pared with this. Gold-cadmium, similarly obtained, is a lead- 
 grey crystalline precipitate, having the composition AuCd 3 . On 
 heating, it is converted into gold-mono-cadmium, AuCd (see 
 above). If the gold-zinc alloy is shaken with a solution con- 
 taining a cadmium salt, the gold-cadmium alloy and a zinc salt 
 are obtained. Similarly, a copper plate, acting on solutions 
 containing gold, yields a black, spongy compound of gold and 
 copper ; and gold-lead and gold-tin alloys in the form of black 
 slimes are also readily prepared. 
 
 The aluminium alloys have been investigated by Roberts- 
 Austen, f These allovs are remarkable for their intense colour, 
 varying from yellowish-green to purple, and some of them appear 
 to present the characteristics of true chemical compounds. A 
 white alloy containing 10 per cent, of aluminium is very hard, 
 and has a melting point no less than 417 lower than that of 
 pure gold ; but a deep purple alloy, containing 22 per cent, of 
 aluminium (corresponding to the formula AuAl 2 ), appears to 
 melt at a temperature between 1,065 and 1,070 s , or over 20 
 higher than the melting point of pure gold. It presents, there- 
 fore, the extremely rare case of an alloy, the fusing point of 
 which is higher than that of the least fusible of its constituents, 
 and this fact affords strong evidence that it is a true compound 
 of gold and aluminium. It is hardly necessary to point out that 
 the melting points of ordinary chemical compounds are often 
 much higher than the melting point of the least fusible con- 
 stituent. There is a strong tendency for this purple alloy to be 
 formed when gold is melted with an excess of aluminium, and 
 the result is that the alloys rich in aluminium usually show a 
 marked lack of homogeneity. Evidence of the existence of the 
 compound AuAl 2 was also obtained by Heycock and Neville J 
 
 * Ghem. Soc. Journ., vol. Ixvi., part 2 (1894), p. 236. Nature, vol. xliv. 
 (1891), p. 111. 
 t Proc. Roy. Soc., vol. 1. (1892), p. 307. J Loc. cit. 
 
ALLOYS OP GOLD. 15 
 
 by noting the effect on the freezing point of tin caused by addi- 
 tions of gold and aluminium in various proportions. A crystal- 
 line alloy of gold and bismuth, containing gold 68*22 per cent., 
 bismuth 31-78 per cent., has been prepared by Pearce by means 
 of liquation. By melting this with silver, a crystalline alloy 
 corresponding to the formula AuAg was produced, and crystal- 
 line alloys of gold, silver and copper were also obtained.* 
 Specimens of these products are in the Percy collection. 
 
 The diffusion of gold into other metals, both liquid and solid, 
 has lately been investigated by Roberts- A usten.f The rates of 
 diffusion in liquid metals are of the same order as those of 
 soluble salts in water, but the diffusion in solids is very slow. 
 If gold is placed at the base of a cylinder of solid lead 70 mm. 
 high, some is found to have reached the top in thirty days, the 
 temperature being kept at 251. The rate of diffusion is still 
 measurable at 100, but is almost inappreciable at ordinary 
 temperatures. 
 
 The alloys which are most important to the metallurgist are 
 those which gold forms with mercury, copper, silver, lead, and zinc. J 
 Amalgams. Mercury rapidly dissolves gold at ordinary tem- 
 peratures, forming liquid, pasty, or solid amalgams, according 
 to the proportions of the metals present, and their purity. A 
 piece of gold rubbed with mercury is immediately penetrated by 
 it and becomes exceedingly brittle. The ductility is not always 
 restored when the mercury is removed by distillation, a crystal- 
 line structure being often induced. Some forms of precipitated 
 gold are not readily taken up by mercury, the particles tending 
 to float on the surface of the latter. An amalgam containing 
 90 per cent, of mercury is liquid, that containing 87 '5 per cent, 
 is pasty, and that containing 85 per cent, crystallises in yellowish- 
 white easily fusible prisms. On dissolving precipitated gold in 
 mercury heated to 120, and then cooling the mass, white 
 crystalline plates having a composition corresponding to the 
 formula AuHg 4 separate out.J Amalgams with smaller propor- 
 tions of mercury can be obtained in various ways by heating 
 gold and mercury to different temperatures up to a low red heat, 
 and acting on the products with nitric acid. Gold amalgam 
 dissolves readily in excess of mercury, forming a liquid mass 
 .from which it may be partially separated by straining through 
 chamois leather, when a white pasty amalgam containing about 
 3.3 per cent, gold remains behind, while the mercury which 
 niters through contains some gold, the amount varying with the 
 temperature, but not with the pressure applied. Kasentseff has 
 
 * Trans. Am. Inst. Mng. Eng., 1885. 
 
 t Proc. Roy. Soc., vol. lix., p. 281. Bakerian Lecture, Phil. Trans., 1896. 
 
 Gold is separated from its ores by lead in smelting operations, these two 
 metals mixing in all proportions. It can be separated from molten lead by 
 zinc or aluminium, but these processes are dealt with in other volumes of 
 the series. 
 
 Fre"my, Ency. Chim., vol. iii., L'or, p. 99. 
 
16 THE METALLURGY OP GOLD. 
 
 shown* that this liquid amalgam contains O'll per cent, of gold 
 if it is filtered at 0, O126 per cent, at 20, and 0-650 per cent, 
 at 100C.; these amalgams, therefore, behave like aqueous 
 solutions. 
 
 When amalgams are gradually heated, the mercury is distilled 
 off by degrees, the action soon ceasing if the temperature is 
 allowed to become stationary, and distillation recommencing if 
 it is again raised. At 440 (somewhat below a red heat), an 
 amalgam containing about three parts of gold to one of mercury 
 is obtained, and at a bright red heat almost all the mercury is 
 expelled, and if the heating has not been pushed too rapidly the 
 vapours contain but little gold. The gold obstinately retains 
 about O'l per cent, of mercury, which is not driven off below the 
 melting point of gold. 
 
 Gold and Silver. Gold and silver unite in all proportions, 
 yielding alloys which are harder, more fusible, and more elastic 
 than either metal. The hardest is that containing two parts of 
 gold to one of silver. The colour of gold is sensibly lowered by 
 the addition of very small quantities of silver, and on increasing 
 the proportion of the latter, the colour changes by tints of a 
 greenish-yellow (when from 20 to 40 per cent, of silver is pre- 
 sent) to white, with a scarcely perceptible yellow tinge (when 
 50 per cent, of silver is present), and silver- white (when more 
 than 60 per cent, of silver is present). Pearce has obtained 
 in regular octahedra the alloys corresponding to the formulae 
 Au 8 Ag, Au 6 Ag, and Au 2 Ag by liquation, and Levol has 
 obtained A.uAg, AuAg 2 , and AuAg 5 in perfectly definite 
 homogeneous forms. The alloys containing small quantities 
 (less than 20 per cent.) of gold liquate readily if kept for some 
 time in a state of quiet fusion, an alloy containing one part of 
 gold to five parts of silver (AuAg 9 ) sinking to the bottom, and 
 slightly auriferous silver floating at the top. The silver-gold 
 alloys most used in jewellery are green gold (silver 25, gold 75), 
 dead-leaf gold (silver 30, gold 70), and the alloy containing 40 
 per cent, of silver. Triple alloys of gold, silver, and copper 
 are employed far more frequently by English jewellers than 
 those last mentioned; of these the alloys, consisting of 22-, 18-, 
 15-, 12-, and 9-carat gold respectively, can be hall marked, t 
 Alloys of gold and silver were much used for coinage before the 
 methods of parting became well known and inexpensive. 
 
 Electrum, which includes pale yellow alloys with from 15 
 to 35 per cent, of silver, and which occurs native, was much 
 used for ornaments and coins by the Greeks and Romans, and 
 by the nations which acquired their arts. The use of silver 
 in the gold-copper coinage alloys was not discontinued until 
 quite recently, all English guineas and the Australian sovereigns 
 manufactured at Sydney up to the year 1871 containing some 
 
 * Bull. Soc. Chim. [2], vol xxx., p. 20. 
 
 t 22-carat gold contains || by weight of fine gold, and so on. For 
 details of these alloys see Gee's Goldsmith's Handbook, pp. 41-52. 
 
ALLOYS OP GOLD. 17 
 
 of it. Both nitric and sulphuric acids attack silver-gold alloys, 
 almost completely dissolving out the silver if it is present in 
 amounts variously stated at from 60 to 75 per cent, or over, 
 while, if the proportion falls below 60 per cent., some of the 
 silver is left undissolved with the gold. Hydrochloric acid 
 scarcely attacks these alloys, and the action of aqua regia is soon 
 arrested if the proportion of silver is considerable. 
 
 Gold and Copper. These metals dissolve in one another in 
 all proportions, forming a complete series of homogeneous alloys, 
 which are less malleable, harder, more elastic, and more fusible 
 than gold, and possess a reddish tint. Those with less than 
 12 per cent, of copper are fairly malleable; when more than 
 this is present they are more difficult to work owing to their 
 hardness. Since no change of volume occurs when these alloys 
 are formed, their densities may be calculated from those of gold 
 and copper. The density of standard gold is 1748, and that of 
 the alloy containing gold 900, copper 100 is 17*16. For the den- 
 sities of other gold-copper alloys, see Rigg in the Report of the 
 Mint, 1876, p. 46. Many of the alloys have been used for coinage 
 at various times. The Greeks and Romans, after electrum had 
 fallen into disuse, employed the purest gold they could procure, 
 viz., that from 990 to 997 fine. Under the Roman Emperors, 
 however, copper was intentionally added, and in the two cen- 
 turies preceding the fall of Rome very base alloys were used, 
 some containing only 2 per cent, of gold, or even less.* In the 
 middle ages these base alloys were discarded, and the " byzant " 
 of Constantinople and the " florin " of Florence were both nearly 
 pure gold, while the first gold coins struck by the nations of 
 Western Europe were also intended to be absolutely fine. The 
 standard 916-6 or ij (i.e., 916-6 parts gold in 1,000) was adopted 
 by England in the year 1526, the standard of 994-8, which had 
 been introduced in 1343, being finally abandoned in 1637; the 
 900 standard was introduced in France in 1794, and subsequently 
 adopted in other countries. These two standards are now those 
 most commonly used, the English standard being employed by 
 Russia, Portugal, India, and Turkey, and the French standard 
 by most other civilised countries; the Austrian ducat, however, 
 has a fineness of 986 and that of Holland a fineness of 983, while 
 the Egyptian standard is only 875. Of all these alloys the 
 900 and 916-6 standards are those best adapted for coinage, 
 keeping their colour fairly well, and resisting wear better than 
 richer alloys. The 900-alloy is harder and wears better than 
 the 916'6-alloy, but the difference is not great, the rate of wear 
 depending less on such small differences of composition than on 
 the mechanical and thermal treatment of the alloys during the 
 operation of coining. f The alloys used in coinage generally 
 
 * La Monnaie dans I 1 Antiquity. Paris, 1878. 
 t Fourteenth Report of the Royal Mint, 1884, p. 45 and seq. 
 
 2 
 
18 THE METALLURGY OF GOLD. 
 
 contain from eight to twelve parts of silver per 1000 In addition 
 to the gold. 
 
 Gold-copper alloys tarnish on exposure to air owing to oxida- 
 tion of the copper, and blacken on heating in air from the same 
 cause. This oxidised coating may be removed and the colour of 
 fine gold (not that of the original alloy) produced by plunging 
 the metal into dilute acids or alkaline solutions, the operation 
 being technically known as " blanching." The colour of alloys 
 may be improved without previous oxidation by dissolving out 
 some copper by acids, a film of pure gold being thus left on the 
 outside which can be burnished. French jewellers use a hot 
 solution of two parts of nitrate of potash, one part of alum, 
 and one of common salt for this purpose. 
 
 Nitric or sulphuric acid dissolves out the copper from gold- 
 copper alloys under conditions similar to those under which 
 it removes silver from silver-gold alloys. If the copper falls 
 below 6*5 per cent, the alloy is not attacked by these acids 
 (Pearce). Aqua regia dissolves all the alloys completely. 
 
 Liquation of Gold Alloys. The subject of liquation 
 generally, including that observed in gold alloys, has been dis- 
 cussed in the volume introductory to this series. * Experiments 
 of Roberts- Austen and of Peligot are there described which tend 
 to prove that liquation does not occur in the gold- silver and 
 gold-copper alloys rich in gold and free from all impurities. 
 Levol had previously come to the same conclusion as Peligot 
 with regard to gold-copper alloys, but gave no details of 
 his experiments, remarking, however, that the oxidation of 
 the copper made the research difficult. Roberts-Austen has 
 also cited t the evidence afforded in the preparation of the 
 standard gold trial plate made in the Royal Mint in 1873 as 
 conclusive in proving that considerable masses of standard gold 
 can be obtained of uniform composition. A mass of standard 
 gold weighing 72 ozs. was cast in a suitable mould, and rolled 
 into a plate 37 inches long and 6*5 inches wide; portions of 
 metal were cut from different parts of it, and when these were 
 assayed it was found that the greatest variation between any 
 two assays was xdhj^y, there being no evidence of concentration 
 of the precious metal anywhere. In this case the gold and copper 
 used to make the alloy were of exceptional purity. In 1889, 
 moreover, this chemist f examined the composition of two 
 large ingots, weighing 400 ozs. each, which had been sent to the 
 Mint for coinage by the Bank of England. The fineness of these 
 ingots was 896-2 and 978-5 respectively, and the results showed 
 that no definite liquation had taken place in them. In the case 
 of many of the ordinary trade ingots, however, the discrepancies 
 between assays on pieces taken from opposite ends of the bars 
 
 * Roberts- Austen, Introd. to the Study of Met., p. 72. 
 
 t Nineteenth Report of the Royal Mint, 1888, p. 54. J Loc. cit. 
 
ALLOYS OP GOLD. 19 
 
 prove that the composition is not uniform, but this lack of uni- 
 formity is doubtless due to the presence of some other element 
 in addition to gold, silver and copper. 
 
 James C. Booth, of the U.S. Mint, Philadelphia, has stated* 
 that liquation of gold-copper alloys is induced by the presence 
 of small quantities of base metals, the greatest effect being 
 produced by antimony, bismuth and arsenic, while lead, tin and 
 zinc act in the same way to a less degree, but no direct experi- 
 ments have been adduced in support of this view. Alloys 
 containing much lead, bismuth and zinc cannot be obtained 
 perfectly homogeneous. 
 
 It has been pointed out by the author f that segregation takes 
 place in standard gold rendered crystalline by small quantities 
 of lead or bismuth, the presence of 0-2 per cent, of either of 
 these metals causing the centre of a sphere 3 inches in diameter 
 to be enriched in gold to the extent of about one part per 
 thousand. This is due to the fact, proved by Roberts- Austen, 
 that an eutectic alloy of gold and lead remains molten after the 
 remainder of the mass has solidified, and is consequently driven 
 towards the centre of the sphere, and that this fusible alloy 
 contains much gold but very little copper. Arnold's micro 
 graphic results | tend to confirm the view which follows from 
 this that the brittleness of crystalline gold is due to the presence 
 of films composed of such eutectic alloys separating the crystals 
 of gold from each other, but Osmond and Roberts- Austen proved 
 the non-existence of these films in quickly-cooled ingots. 
 
 It has been shown by Edward Matthey|| that when gold ingots 
 containing members of the platinum group are cooled from a 
 state of fusion an alloy rich in the more fusible element (gold) 
 falls out first, driving the less fusible constituent to the centre. 
 Thus the assay of an outside cut of such an ingot gives a result 
 too high in gold, sometimes by several per cent. It has long 
 been known, moreover, that iridium and osmium become concen- 
 trated towards the bottom of the mass. The reason for this is 
 that, at the temperature of fusion of gold, these refractory ele- 
 ments, either free or alloyed with gold, sink in the molten metal 
 and are left in the state of small crystalline particles. 
 
 Matthey has more recently investigated U the question of 
 segregation in connection with the alloys of gold, silver, lead 
 and zinc produced in cyanide mills. He found that one such 
 ingot weighing about 120 ozs. contained 662 parts per 1,000 of 
 gold at the bottom corner, and only 439 parts at the top. In 
 another case, when 164 per cent, of lead and 9 '5 per cent of zinc 
 were present, the standard fineness of an ingot weighing 400 
 
 * Condemnation of Ingot Melts. Washington, 1875. 
 t Chem. Soc. Journ., vol. Ixvii., 1895, p. 552. 
 J Engineering, vol. Ixi. (1896), p. 176. 
 Phil. Trans. Roy. Soc., vol. clxxxvii. (1896), A. p. 417. 
 I! Proc. Roy. Soc., vol. li. (1892), p. 447, and Phil. Trans. Roy. Soc. t 
 vol. clxxxiii (1892), A. pp. 629-652. 
 IF Proc. Roy. Soc., vol. Ix. (1896), p. 21. 
 
20 THE METALLURGY OF GOLD. 
 
 ozs., as shown by actually separating the whole of the precious 
 metals, was, gold 614-0, silver 75-8, and its true value 1,028, 
 while the value as deduced from the average of fourteen assays 
 made on it (gold 576-0, silver 90) would have been only 965. 
 Seven dip assays made on this ingot varied from 562 to 622 fine. 
 Other cases of irregular distribution were even more remarkable. 
 By experiments with synthetic alloys of gold and zinc, Matthey 
 found that gold tends to liquate towards the centre of the mass 
 but only in a slight degree, the centre of a 3-inch sphere of an 
 alloy containing gold 90, zinc 10, being only about 1 to 2 parts 
 per 1,000 richer than the outside. Lead acts similarly but with 
 greater effect, the centre being about 29 parts per 1,000 richer 
 in gold than the outside when 30 per cent, of lead is present, 
 but the combination of 15 per cent, of lead and 10 per cent, of zino 
 is still more powerful in causing segregation, the sphere being 
 found to contain 657 parts of gold at the top, 785 in the centre, 
 and 790 at the bottom, gravity thus playing a part. The 
 addition of silver, however, if it amounts to not less than about 
 two-thirds of the quantity of zinc and lead taken together, 
 appears absolutely to prevent any liquation from taking place, 
 an alloy containing approximately, gold 55, zinc 7, lead 18 y 
 silver 20, being practically homogeneous. 
 
 CHAPTER II. 
 CHEMISTRY OF THE COMPOUNDS OF GOLD. 
 
 Compounds of Gold. Gold is characterised chemically by an 
 extreme indifference to the action of all bodies usually met with 
 in nature. Its compounds are formed with difficulty, and decom- 
 pose readily, their heat of combination being in general small, 
 while some are formed by endothermic actions. The result of 
 this condition of things is that gold is found in nature chiefly 
 in the metallic form, and the mineralogist has, therefore, few 
 compounds to consider. Nevertheless, the laws governing the 
 formation of artificially-prepared compounds of gold are of great 
 importance to the metallurgist, as the processes of extraction 
 are all based on these laws, and in particular, some knowledge 
 of the reactions and general behaviour of gold in its compounds 
 is essential to those who are engaged in extracting gold by wet 
 processes. A short account of some of the compounds of gold is, 
 therefore, appended, special attention being paid to those bodies- 
 which are most likely to be of interest to the metallurgist. 
 
 Gold forms two series of compounds, having the general 
 formulae AuR and AuR 3 , while doubiful compounds corre- 
 sponding to AuR 2 , AuR 4 , and AuR 5 , have been declared to 
 exist by Thomson, Prat, Figuier, and others. These two series- 
 are denominated aurous and auric respectively. 
 
CHLORIDES OF GOLD. 
 
 21 
 
 Compounds of Gold with, the Halogens. Gold forms two 
 series of compounds with the halogens, the general formulae 
 being AuR and AuR 3 respectively. Supposed compounds having 
 the formula AuR 2 have been described, but are probably mix- 
 tures of the series denoted by AuR and AuR 3 . All these bodies 
 are very unstable, existing throughout only low ranges of tem- 
 perature, whether in the dry state or in aqueous solution. The 
 chlorides are the most easily formed and the least unstable, the 
 bromides coming next, as might be predicted from the heats of 
 combination which are given in the subjoined table in calories: 
 
 
 Chloride. 
 
 Bromide. 
 
 Iodide. 
 
 AuR, solid 
 AuRg, solid, 
 AuR 8 , in solution, 
 
 + 5'8 
 + 22-8 
 + 27-3 
 
 - o-i 
 
 V 
 
 + 5-1 
 
 - 5-5 
 
 CHLORIDES OF GOLD. 
 
 Gold Monochloride or Aurous Chloride, AuCl. This 
 salt is prepared by heating the trichloride to 185 in air for 
 twelve hours. It is non-volatile and unaltered at ordinary tem- 
 peratures and pressures by dry air, even when exposed to light, 
 but begins to decompose at temperatures above 160, and the 
 decomposition is complete if it is heated at 175 to 180 for six 
 days, or at 250 for one hour. Its density is 74. Water con- 
 verts aurous chloride into a mixture of gold and gold trichloride. 
 It is a citron-yellow amorphous powder. 
 
 Auro-Aurichloride, Au 2 Cl 4 . A dark red compound having 
 this composition is said, by Thomsen, to be obtained by heating 
 finely-divided gold in a current of chlorine to 140 to 170. 
 According to Kriiss and Lindet it is merely a mixture of AuCl 
 and AuCl 3 . It yields gold and Au01 3 if brought in contact with 
 water, ether, or alcohol. 
 
 Auric Chloride or Gold Trichloride, Au01 3 . Preparation. 
 Trichloride of gold can be prepared, according to Debray,* 
 by heating finely-divided gold in a current of pure dry chlorine 
 in a glass tube at a temperature of 300. The chloride formed 
 sublimes at this temperature and is deposited in the cooler part 
 of the tube in fine red prisms and needles crystallising in the 
 triclinic system. Kriiss tried to repeat Debray's experiments 
 without confirming his results.! He found that at first brown- 
 ish vapours of AuCl 3 were given off freely at 140 to 150, and 
 were condensed in the cold part of the tube, but that after some 
 time the sublimation ceased, and the whole mass was trans- 
 formed to brownish-red trichloride. On continuing to raise the 
 
 * Compt. fiend., vol. xix., p. 985. t Berichte, vol. xx., p. 212. 
 
 OF THB 
 
 UNIVERSITY 1 
 
22 THE METALLURGY OF GOLD. 
 
 temperature AuCl 3 began to decompose at about 180 in spite of 
 the presence of an excess of chlorine, and protochloride of gold, 
 AuOl, of a greenish-yellow colour was formed. At 220 the 
 AuOl was completely decomposed, leaving metallic gold, while 
 up to this temperature a little AuCl 3 continued to be sublimed, 
 but towards 300 all action ceased, and the gold remained un- 
 attacked by the chlorine at all higher temperatures. On lowering 
 the temperature, Kruss observed in succession the formation of 
 AuCl at 220, and of Au01 3 at 180. 
 
 The exact point at which gold ceases to be attacked by chlorine 
 and the rate of volatilisation of the chloride are of great import- 
 ance in connection with the loss of gold, on roasting auriferous 
 materials with salt.* The matter has, therefore, been lately 
 investigated by the author f with the following results. Gold 
 unites with chlorine if placed in the gas at atmospheric pressure 
 at all temperatures up to a white heat, but the subsequent 
 decomposition of the chloride is rapid above 300. The absorption 
 of chlorine by gold with the formation of chlorides at first in- 
 creases in rapidity as the temperature rises, and reaches its 
 maximum at about 225. The fact that gold is attacked by 
 chlorine, and that the chloride is subsequently volatilised at all 
 temperatures between 180 and 1,100, was proved by means of 
 Deville's hot and cold tubes, which enable part of the sublimed 
 chloride to be collected. The rate of volatilisation at various 
 temperatures is as follows : 
 
 
 
 Gold Volatilised. 
 
 Number. 
 
 Temperature. 
 
 Lost from 
 Porcelain Vessel 
 containing 
 the Gold. 
 
 Recovered from 
 Water Tube. 
 
 Percentage 
 Volatilised in 
 30 minutes. 
 
 1 
 
 180 
 
 Grammes. 
 
 Grammes. 
 0-00105 
 
 0-007 
 
 2 
 
 230 
 
 ... 
 
 0-0522 
 
 0-35 
 
 3 
 
 300 
 
 ... 
 
 0-3473 
 
 2-32 
 
 4 
 
 390 
 
 0-2731 
 
 0-1748 
 
 1-82 
 
 5 
 
 480 
 
 0-1325 
 
 0-0884 
 
 0-88 
 
 6 
 
 580 
 
 0-0907 
 
 0-0624 
 
 0-60 
 
 7 
 
 590 
 
 0-0864 
 
 0-0590 
 
 0-58 
 
 8 
 
 805 
 
 0-0753 
 
 0-0518 
 
 0-50 
 
 9 
 
 965 
 
 0-2441 
 
 0-1672 
 
 1-63 
 
 10 
 
 1100 
 
 0-2895 
 
 0-2037 
 
 1-93 
 
 For purposes of comparison, it may be added that when gold 
 is heated in air or coal gas, no gold is volatilised below 1,050, 
 
 * See Prof. Christy's experiments, Section on " Roasting," Chap. XI. 
 t Chem. Soc. Journ., vol. Ixvii. (1895), p. 881. 
 
CHLORIDES OP GOLD. 
 
 23 
 
 and only about 0*02 per cent, in 30 minutes at 1,100,* or about 
 one-hundredth part of that volatilised in chlorine. 
 
 In the table the amounts " lost " by volatilisation include the 
 amounts recovered from the water tube and the gold condensed 
 on the inside of the outer tube \ the latter sublimate was not 
 recovered separately after each experiment. 
 
 In every case the gold recovered from the water tube was 
 associated with a little less chlorine than was required to form 
 gold trichloride, and therefore presumably contained either some 
 metallic gold or AuCl, or both ; these doubtless resulted from 
 the dissociation of some of the trichloride when in the form of 
 vapour. 
 
 The amounts volatilised vary according to two different factors 
 (1) The vapour pressure of gold trichloride, AuCl 3 , which of 
 
 Fig. 3a. 
 
 course increases continuously as the temperature rises; and (2) 
 the pressure of dissociation of the trichloride, which also rises 
 continuously with the temperature, but not at the same rate as 
 the vapour pressure. The rise of vapour pressure tends to raise, 
 and that of the pressure of dissociation to reduce, the amount of 
 gold volatilised as chloride. The vapour pressure increases more 
 rapidly than the pressure of dissociation at temperatures below 
 300, and also above 900, but less rapidly at intermediate tem- 
 peratures. Hence the curve (Fig. 3a) showing the variation of 
 volatilisation with temperature is irregular, passing through a 
 * Chem. Soc. Journ., vol. Ixiii. (1893), p. 717. 
 
: THE METALLURGY OF GOLD. 
 
 maximum near 300, and a minimum at a point somewhere 
 below the melting point of gold. The first-named change in the 
 direction of the curve possibly occurs at the melting point of the 
 chloride, namely 288. The second change is perhaps caused by 
 the change of sign of the heat of formation of the trichloride 
 AuCl 3 ; when this becomes negative, the pressure of dissociation 
 of the compound would decrease, in accordance with the law of 
 van't Hoff and Le Chatelier. However this may be, it is certain 
 that when gold is heated in chlorine at atmospheric pressure, 
 trichloride of gold is formed and volatilised at all temperatures 
 above 180, up to, and probably far beyond, 1,100. 
 
 The usual method adopted for the preparation of auric chloride 
 is to dissolve gold in aqua regia, and then to evaporate the 
 liquid to dryness, keeping the temperature above 100 to pre- 
 vent the formation of the hydrochloride Au01 3 HCl. A brownish- 
 red mass is thus formed, consisting of AuCl 3 mixed with more 
 or less of the protochloride and of hydrochloric acid. On taking 
 up with water the protochloride is decomposed into gold and 
 trichloride, but the hydrochloric acid can only be eliminated by 
 shaking with ether, which withdraws the trichloride from its solu- 
 tion in water. If an attempt is made to drive of? the hydrochloric 
 acid by heat, a partial decomposition of the trichloride results. 
 
 Auric chloride exists both in the anhydrous state and in 
 combination with two equivalents of water, Au01 3 . 2 H 2 O, when 
 it occurs in orange-red crystals. The anhydrous salt is of a 
 brilliant-red colour, crystallising in needles belonging to the 
 triclinic system and melting at 288 under a pressure of two 
 atmospheres of chlorine. It can be prepared by drying the 
 hydrated salt at 150. The anhydrous and hydrated salts are 
 both hygroscopic, and dissolve readily in water with elevation of 
 temperature ; they are also soluble in alcohol and ether, and in 
 some acid chlorides, such as As01 3 , SbCl 5 , SnCl 2 , Si01 4 , <fec. 
 
 Auric chloride is readily decomposed by heat. Lowe states* 
 that 4 grammes of the trichloride, when heated in a porcelain 
 basin on a boiling water bath, can be completely transformed 
 into the monochloride, although not until after the lapse of 
 several days. On the other hand, as has already been mentioned, 
 Krtiss states that the decomposition of auric chloride, in an 
 atmosphere of chlorine, begins at 180. According to the 
 experiments of the author,! auric chloride is observed to suffer 
 slow decomposition at as low a temperature as 165 in an 
 atmosphere consisting of chlorine, about 1*6 per cent, being 
 converted into monochloride in four hours at this temperature ; 
 the decomposition is about five times more rapid at 190. The 
 decomposition in air can be readily observed at 100, although 
 it does not seem to be so rapid as was indicated by Lowe. In 
 
 * Dingier' s poly t. Journ., 1891, vol. cclxxix., p. 167. 
 t Ohem. Soc. Journ., vol. Ixvii. (1895), p. 902. 
 
CHLORIDES OF GOLD. 25 
 
 seven days only 6-6 per cent, of the trichloride was decomposed, 
 the initial rate of decomposition being 0*041 per cent, per hour. 
 At 165, however, the initial rate of decomposition appeared to 
 be 3 '2 per cent, per hour, and the conversion into monochloride 
 was complete in four or five days at 160 and in ten hours at 190. 
 The rate of decomposition of the trichloride in air at various 
 temperatures can be calculated from the above data by the help 
 
 / / / \m 
 
 of Harcourt and Esson's formula a y a = ( TI /T ) where otj, a 2 
 
 are the rates of decomposition at the absolute temperatures 
 r lt r 2 respectively. The value of the constant m for this 
 chemical action is found to be about 27, and by substituting 
 this number for m in the equation, the rate of decomposition of 
 gold trichloride is calculated to be 0-365 per cent, in a year at 
 15. The decomposition begins to be observable at 70 in air, 
 when the monochloride is formed, about twenty- five years, 
 however, being required for the complete change at this tem- 
 perature. The observed rate of decomposition shows that a 
 similar change would require about 1,000 days at 100, while it 
 results from calculation, using Harcourt and Esson's formula, 
 that at 200, thirty-six hours, and at the melting point namely, 
 288 less than one minute suffices for the complete decomposi- 
 tion of AuCl 3 in air. 
 
 Whether dry or in aqueous solution, the trichloride is also 
 decomposed by light, gold being deposited in scales in the latter 
 case, but the presence of free hydrochloric acid prevents this 
 decomposition. Weak voltaic currents precipitate metallic gold 
 from the solution of the trichloride upon the negative pole. The 
 solution of trichloride of gold is also decomposed by many re- 
 ducing agents, such as most organic substances, metals and 
 protosalts ; heating the solution in every case hastens the 
 decomposition. The reduction by organic matter is assisted by 
 the action qf light, which is especially efficacious in the presence 
 of starchy and saccharine compounds, or of charcoal or ether. 
 In direct sunlight the last reagent deposits a bright mirror of 
 metallic gold, but under ordinary conditions excessively finely 
 divided gold (Faraday's gold) is precipitated. Alkalies also 
 quicken the action of organic matter, and it may be said that 
 all organic compounds reduce gold chloride on boiling in the 
 presence of potash or soda,* while Muller states that a mixture 
 of glycerine and soda lye is one of the best precipitants for 
 gold chloride, separating the metal completely in highly dilute 
 solutions. According to Kriiss, if potash and soda are quite 
 free from organic matter, they have no action on solutions of 
 auric chloride, whether cold or hot. If a small quantity of 
 organic matter is present, sub-oxide of gold is precipitated; if 
 larger quantities are present, both metallic gold and sub-oxide 
 *Fr<hny, Ency. Chim., vol. iii., 16e Cahier, p. 74. 
 
26 THE METALLURGY OF GOLD. 
 
 are precipitated in the cold, but gold alone at boiling point. 
 Alkaline carbonates are without action on cold solutions, but if 
 they are hot, then half the gold is precipitated as hydrate, while 
 the other half remains in solution in the form of a double 
 chloride of gold and the alkali. The precipitation by means 
 of charcoal is of especial importance in view of its adoption in 
 practice. 
 
 It has been stated that a current of hydrogen gas will preci- 
 pitate gold completely, especially on boiling the solution, but 
 Kriiss has proved that, if the hydrogen is quite pure, it has no 
 effect either on cold or hot solutions. Sulphur, selenium, phos- 
 phorus and arsenic all precipitate gold on boiling the solution 
 of the trichloride. Many metals reduce chloride of gold, the 
 action being, of course, most rapid in the case of the most highly 
 electro-positive metals, such as zinc and iron. It seems probable 
 from the author's experiments * that turnings of these metals 
 would make good precipitants for use on a commercial scale in 
 the chlorination process. Lead sometimes gives fine dendritic 
 plates of gold. Sulphuretted hydrogen precipitates sulphide of 
 gold from both neutral and acid solutions, all traces of gold being 
 readily removed from a solution by this reagent, whilst phos- 
 phoretted, arseniuretted, and antimoniuretted hydrogen all 
 precipitate metallic gold. The lower oxides of nitrogen, 
 nitrous acid and many other " ous " acids and oxides effect the 
 same decomposition. Sulphur dioxide is a convenient reagent, 
 and is often used in the laboratory, being almost equally effica- 
 cious in cold and hot solutions. The reaction is 
 
 2AuCl s + 3SO 2 + 6H 2 = 2Au + 6HC1 + 3H 2 S0 4 . 
 
 Various protosalts also reduce trichloride of gold. Ferrous 
 sulphate is often used to detect the presence of gold in solution 
 as chloride ; this reagent gives the solution a pale blue colour 
 by transmitted light, and brown by reflected light, owing to the 
 formation of finely divided precipitated gold. The reaction is 
 represented by the following equation 
 
 2AuCl 3 + 6FeS0 4 = 2Au + 2Fe 2 (S0 4 ) 3 + Fe 2 Cl 6 . 
 
 To test a dilute solution for gold, a test tube filled with the 
 liquid is held in the hand side by side with a test tube filled 
 with distilled water and a few drops of a clear solution of ferrous 
 sulphate are added to each. On looking down through the 
 length of the test tubes from above, with a white surface as 
 background, any slight changes of colour may be detected by 
 comparison and the liquids may also be compared with the 
 original solution in a test tube. In this way, by a little prac- 
 tice, the presence of gold in the proportion of only ^^iy^ 
 '(1 dwt. per ton of water), or even less can be detected. The 
 
 * See Section on " Precipitation of Gold," Chap. xiii. 
 
BROMIDES OF GOLD. 27 
 
 method is often used in the chlorination process, but it is better 
 to use protochloride of tin, SnCl 2 . This substance gives a brown 
 precipitate of variable composition in concentrated solutions, 
 but if mixed with the tetrachloride, SnCl 4 , it gives a precipi- 
 tate of purple of Cassius. The reaction is very sensitive, and 
 by its means a violet coloration by transmitted light can be 
 obtained in a solution containing 1 part of gold in 500,000 parts 
 of water, while by special means the presence of 1 part of 
 gold in 100,000,000 parts of water can be detected, as de- 
 scribed below : * The liquid supposed to contain gold is raised 
 to boiling, and poured suddenly into a large beaker contain- 
 ing 5 to 10 c.c. of saturated solution of stannous chloride, 
 and the liquids agitated so as to effect complete mixture. A 
 yellowish-white precipitate of tin hydrate forms, which settles 
 rapidly, and can be readily separated from the bulk of the liquid 
 by decantation. If the solution originally contained at least 
 1 part of gold in 5,000,000 of water (3 grs. per ton), the 
 precipitate is coloured purplish-red or blackish-purple, accord- 
 ing to the nature of the solution, and the condition of the 
 precipitant. The colour can be seen without comparing it with 
 other precipitates. If less gold than this is present it is better 
 to compare the precipitate with one obtained by the use of 
 boiling distilled water, and to increase the quantity of liquid 
 used while adhering to the same amount of stannous chloride. 
 In this way the presence of 1 part of gold in 100,000,000 
 parts of water (1 grain of gold in 6 tons of water) can be 
 detected, the amount of liquid required in this case being about 
 3 litres. The gold is concentrated in the precipitate in which 
 a distinct colour is caused by less than 0'05 per cent, of the 
 precious metal. 
 
 Color-auric Acid, H AuCl 4 . Gold trichloride in the presence 
 of free hydrochloric acid is supposed to form this compound, 
 which crystallises out on evaporation in vacuo in long, yellow 
 needles, having the composition HAuCl 4 + 4H 2 O, and since 
 gold chloride unites with many other soluble chlorides to form 
 double chlorides, this hydrochloric acid compound is regarded as 
 an acid. It is more stable than gold trichloride. The chlor- 
 aurates, having the general formula M'AuCl 4 , or AuCl 3 . M'Cl, 
 are readily soluble bodies which can be crystallised, and which 
 decompose with about the same readiness as chlor-auric acid. 
 
 BROMIDES OF GOLD. 
 
 Gold Protobromide, AuBr, is a yellowish-green powder ob- 
 tained by heating the tribromide to about 140. It is insoluble 
 in water, but is decomposed by it, metallic gold and the tribro- 
 
 * T. K. Rose, in Chem. News, 1892, vol. Ixvi., p. 271. 
 
28 THE METALLURGY OF GOLD. 
 
 mide being formed ; the change is especially rapid on boiling, 
 and is hastened by the presence of hydrobromic acid. 
 
 Auro-auric Bromide, Au 2 Br 4 , is produced by the action of 
 bromine on finely-divided gold in the cold, some tribromide 
 being simultaneously formed. Water breaks up this bromide 
 into AuBr and AuBr 3 , and, according to some authorities, it is 
 only a mixture of these bodies. 
 
 Gold Tribromide, AuBr 3 , is produced by the action of a 
 mixture of bromine and water on gold, particularly on the 
 application of heat. Auric tribromide resembles the trichloride 
 in most of its properties. It crystallises in blackish needles or 
 scarlet plates. It is deliquescent, and very soluble in water, and 
 suffers decompositions similar to those noted in describing AuCl 3 , 
 its solutions being still less stable than those of the chloride. A 
 solution of gold tribromide is gradually decolorised by sulphur 
 dioxide, being completely reduced to the state of monobromide 
 before any precipitate of metallic gold is formed. It is prepared 
 in a pure state by heating finely-divided gold in sealed tubes 
 with bromine and arsenic bromide, AsBr 3 , to 126. Gold tri- 
 bromide forms intensely coloured brownish-red aqueous solutions, 
 the presence of a mere trace of the salt in a solution being 
 observable in this way. Double bromides exist analogous to 
 the chlor-aurates. The bromides are not volatile. 
 
 The Iodides of Gold are of little interest to the metallurgist. 
 They are prepared with difficulty, and decompose more readily 
 than the chlorides and bromides and are not likely to become of 
 practical importance. The tri-iodide is formed if gold is heated 
 with water and iodine to 50, particularly in direct sunshine. 
 
 CYANIDES OF GOLD.* 
 
 Cyanogen and gold unite in two proportions forming aurous 
 and auric cyanides, but the latter is only known with certainty 
 in combination. 
 
 Aurous Cyanide, AuCy, is obtained by heating aurocyanide 
 of potassium, KAuCy 2 , with hydrochloric or nitric acid. It is 
 a lemon-yellow crystalline powder, insoluble in water, and 
 unaltered by exposure to air. It is decomposed by heat, yield- 
 ing metallic gold and cyanogen, and is soluble in ammonia, in 
 alkaline cyanides, and in hyposulphite of soda. It is unattacked 
 by the mineral acids, except by aqua regia, but is decomposed 
 when boiled with potash, metallic gold being thrown down. 
 
 Aurocyanide of Potassium, KAuCy 2 , is obtained by crystal- 
 lisation from its solution, which is prepared by dissolving 
 metallic gold, auric oxide or aurous cyanide in a solution of 
 potassium cyanide. It is slightly soluble in water and the 
 
 * See also chapter xvi. 
 
OXIDES OF GOLD. 29 
 
 aqueous solution, especially if hot, gilds copper or silver without 
 the agency of a battery. Gold is also precipitated from the 
 solution by zinc and the alkali metals. Precipitates are also 
 formed on the addition of salts of zinc, tin, iron, or silver, no pre- 
 cipitates being formed if potassium cyanide is present in excess. 
 Auricyanide of Potassium, AuCy 3 . KOy, is formed by adding 
 potassium cyanide to a solution of trichloride of gold, the pre- 
 cipitate first formed being redissolved. The solution is com- 
 pletely decolorised, and on cooling deposits colourless crystals of 
 AuCy 3 . KCy + 3H 2 O. These effloresce in air, giving up two 
 molecules of water ; and, on heating, the third molecule of water 
 and some cyanogen are given off, aurocyanide of potassium being 
 formed, and this in its turn is decomposed at a slightly higher 
 temperature. 
 
 OXIDES OP GOLD. 
 
 ATITOUS Oxide, Au 2 O. This oxide is prepared by decomposing 
 aurous chloride, AuCl, or the corresponding bromide by potash 
 in the cold (Berzelius), when a violet precipitate forms which 
 is blackish when moist, but greyish when dry. When freshly 
 precipitated it is soluble both in alkalies and in cold water, 
 forming an indigo blue solution, with brownish fluorescence, and 
 on warming the solution slightly the corresponding hydrate is pre- 
 cipitated. It is also prepared by the action of nitrate of mercury 
 on the trichloride, and by boiling aurate of potash with organic 
 compounds, such as citrates or tartrates, or by boiling a solution 
 of the trichloride with the potassium salts of these acids. When 
 prepared according to these methods, aurous oxide always 
 contains a certain proportion of metallic gold. Kriiss obtained 
 the oxide pure by reducing brom-aurate of potassium at by 
 SO 2 , passing in the gas only until the solution became colourless, 
 after which an excess of gas would have precipitated metallic 
 gold. Aurous hydrate is then precipitated by potash, and, after 
 being agglomerated by boiling, it is filtered, washed with cold 
 water, dried, and heated to 200 to expel the water of hydration. 
 At 250 it is resolved into gold and oxygen. Hydrochloric acid 
 decomposes aurous oxide into metallic gold and auric salts, slowly 
 in the cold, quickly at a bpiling temperature; aqua regia dis- 
 solves the oxide, but sulphuric and nitric acids are without action 
 on it, while weak bases at once decompose it. 
 
 An intermediate oxide, AuO, is prepared as a black powder 
 by dissolving metallic gold in aqua regia containing an excess of 
 hydrochloric acid, then adding an excess of carbonate of potash, 
 and afterwards filtering and drying the precipitate. It has been 
 little studied, but the temperature at which it decomposes has 
 been fixed at 205 and its hydrate has been prepared. 
 
 Auric Oxide, Au 2 O 3 . This, the best known oxide, is a black 
 
30 THE METALLURGY OF GOLD. 
 
 powder when anhydrous, and is precipitated from solutions of 
 auric chloride in the form of a hydrate by the caustic alkalies, the 
 carbonates of the alkalies, and hydrates of the alkaline earths or 
 zinc. The readiest method of preparation of this compound is 
 to add caustic potash, little by little, to a hot solution of gold 
 chloride, until the yellow precipitate first formed is dissolved to 
 a brown liquid. Then a slight excess of sulphuric acid or some 
 Glauber's salt is added, the precipitate filtered off, washed and 
 purified from potash by being redissolved in concentrated nitric 
 acid, and reprecipitated by dilution with water. On drying this 
 precipitate in vacuo, the hydrate Au 2 O 3 . H 2 O, an ochreous 
 powder, results. If it is heated to 110, oxygen begins to be 
 given off. At 160, AuO remains, and on heating for some time 
 at 250, metallic gold remains. Trioxide of gold dissolves in 
 concentrated sulphuric and nitric acids, from which it is partly 
 reprecipitated on boiling or on dilution, and these solutions are 
 supposed to contain sulphates and nitrates of gold respectively. 
 Double nitrates of gold and the alkalies have been obtained as 
 crystals. Hydrochloric and hydrobromic acids dissolve the 
 trioxide forming the haloid salts, but hydriodic acid decomposes 
 it on boiling, giving iodine and metallic gold. Gold trioxide 
 dissolves in boiling solutions of alkaline chlorides, giving aurates 
 and chlor-aurates, while it also combines with metallic oxides to 
 form aurates. 
 
 It is easily reduced by hydrogen, carbon and carbonic oxide, 
 with the aid of very gentle heat. Boiling alcohol reduces it, 
 yielding minute spangles of gold which were formerly used in 
 miniature painting. 
 
 Aurates. The aurates of potash and soda have the general 
 formula Au 2 O 3 . R' 2 O or K' 2 Au 2 O 4 assigned to them. They 
 are readily soluble, crystallisable compounds, and are formed 
 when alkalies are added in excess to solutions of gold chloride. 
 The aurates of calcium, magnesium and zinc are insoluble in 
 water, but soluble in hydrochloric acid. 
 
 Fulminating Gold is a compound of auric oxide with am- 
 monia, Au 2 O 3 (NH 3 ) 4 , which is formed by precipitating gold 
 chloride with ammonia or its carbonate, or by the action of 
 ammonia on gold trioxide. When prepared by the former 
 method its composition is variable, but the fulminate is always 
 a fearful explosive, decomposing with violence at 145, or on 
 being struck, and sometimes even spontaneously. It is de- 
 composed without explosion by sulphuretted hydrogen, and 
 by protochloride of tin. It is a grey pulverulent powder, 
 insoluble in water, but soluble in potassium cyanide, auricyanide 
 of potassium being formed. 
 
 Sulphites of Gold. Alkaline sulphites, or sulphur dioxide, 
 which reduce gold trichloride easily, do not produce the same 
 effect on a solution of an alkaline aurate. If sodic bisulphite is 
 
OXIDES OF GOLD. 31 
 
 added to a boiling solution of sodic aurate (NaAuO 2 ) a yellowish 
 precipitate is formed, soluble in excess of sodic bisulphite, and 
 consisting of a double sulphite of gold and sodium, or sodic 
 auro-sulphite, having the composition 3Na 2 SO 3 . Au 2 SO 3 + 3H 2 O. 
 It is obtained pure by precipitating the corresponding baric salt 
 with BaCl 2 , and decomposing the precipitate with the minimum 
 quantity of sodic carbonate. Double sulphites of potassium and 
 ammonium with gold also exist. These salts are decomposed by 
 acids, sulphite of gold being deposited, and also on boiling their 
 aqueous solutions, but the addition of sulphuretted hydrogen or 
 alkaline sulphides has no effect on them. 
 
 Hyposulphites of Gold. These compounds are now called 
 thiosulphates by chemists, but the old name is retained as being 
 universally employed by metallurgists. They are especially 
 interesting to the metallurgist, as on their formation depends the 
 extraction of gold from auriferous silver ores, when these are 
 treated by the ordinary hyposulphite, or by the Russell process. 
 The soluble double hyposulphites of gold with the alkalies and 
 alkaline earths have the general formula 3R"S 2 O 3 . Au 2 S 2 O 3 + 
 4H 2 O. The double compounds of gold with sodium, potassium, 
 calcium, magnesium and barium, are all known. The sodic 
 salt is prepared by adding a dilute solution of gold trichloride 
 little by little to a concentrated solution of sodic hyposulphite, 
 when the following reaction occurs : 
 
 8Na 2 S 2 8 + 2AuCl 3 = Au 2 S 2 3 . 3Na 2 S 2 8 + 2Na 2 S 4 6 + 6NaCl. 
 
 The double hyposulphite may be separated by precipitation with 
 strong alcohol, with which it is also washed, or it may be 
 purified by repeated solution in water and precipitation with 
 alcohol. Thus prepared, it consists of colourless crystalline 
 needles, highly soluble in water, but almost insoluble in alcohol. 
 The solution, which possesses a sweetish taste, decomposes under 
 the influence of heat, the action being much more rapid when 
 nitric acid is present; metallic gold and sulphate of soda are 
 formed. Gold, however, is not reduced from its solution as double 
 hyposulphite by either stannous chloride, ferrous sulphate or 
 oxalic acid, although sulphuretted hydrogen and alkaline sulphides 
 give a black precipitate of Au 2 S 3 . The addition of hydrochloric 
 acid or of dilute sulphuric acid does not immediately cause an 
 evolution of sulphur dioxide and a deposit of sulphur, as in the 
 case of ordinary hyposulphites. Since, therefore, the double 
 sulphite of soda and gold does not present the characteristics of 
 either aurous salts or of hyposulphurous acid, it has been sug- 
 gested that it contains a compound radical, and has a composition 
 expressed by either Na 3 S 4 O 6 Au or Na 3 S 4 O 6 Au + 2H 2 O. The 
 addition of any dilute acid soon effects the decomposition of this 
 body in solution, gold sulphide being precipitated j the reaction 
 is accelerated by heat. 
 
32 THE METALLURGY OF GOLD. 
 
 The double hyposulphites of potassium, calcium, barium and 
 magnesium present similar characteristics. If the barium salt 
 is treated with the amount of sulphuric acid required by theory, 
 a solution of the acid auro-hyposulphite, 3H 2 S 2 O 3 . Au 2 S 2 O 3 , is 
 obtained, but it cannot be crystallised. It has been supposed 
 that the calcium salt is more easily formed than the sodic salt, 
 and, therefore, that calcium hyposulphite is more suitable than 
 sodium hyposulphite for use in the leaching process, whenever 
 gold is present in perceptible quantities. According to a series 
 of experiments conducted by Russell,* this is not the case, very 
 little difference existing in the ease of formation of the calcic, 
 sodic, magnesic and potassic salts. 
 
 Russell has demonstrated f that finely divided gold is soluble 
 to a limited extent (i.e., 0*002 gramme in 1,000 c.c. in 48 hours), 
 in solutions of sodic hyposulphite of all degrees of concentra- 
 tion. The action depends on the oxidation of the gold by the 
 air present in the solution, the soluble double hyposulphite, 
 Au 2 S 2 O 3 . 3Na 2 S 2 O 3 + 4H 2 O, and caustic soda being formed. 
 
 The formation of this hyposulphite by the action of the sodic 
 salt on gold sulphide is far more complete and rapid. In 24 
 hours in the cold, 0*066 gramme of gold, and in 2 hours at 
 65, 0*117 gramme of gold were dissolved in dilute solutions. 
 Since an alkaline sulphide will again precipitate Au 2 S 3 from 
 the solution, these facts seem to be at variance, and Stetefeldt 
 suggests that the gold sulphide originally contained a small 
 quantity of metallic gold in an excessively fine state of division, 
 and that on heating more gold was set free. He instances Level's 
 statement that sulphuretted hydrogen precipitates metallic gold, 
 and not gold sulphide, from a boiling solution of the trichloride. 
 He believes J that it is only the free gold which is attacked, and 
 that the results obtained on attacking the sulphide were better 
 than those in which metallic gold was used, owing to the much 
 finer state of division in which the gold existed in the former 
 case. Another reason for the inverse reactions may of course 
 exist in the influence of mass, the small quantity of sodic 
 sulphide which is formed by the reaction being insufficient to 
 precipitate any gold, which comes down on further additions of 
 the same reagent. 
 
 In some further experiments which Russell conducted, it was 
 proved that gold sulphide is far more soluble in a solution 
 containing the molecular quantities, 4Na. 2 S 2 O 3 . 3Cu 2 S 2 O 3 (hypo- 
 sulphite of copper and sodium), than in sodic hyposulphite alone, 
 especially if the solutions are cold. If the solution is heated 
 the reduction of the gold sulphide to metallic gold renders it less 
 soluble, and less rapid dissolution occurs. The composition of 
 the double salts obtained by the use of this " extra " solution 
 has not been worked out. 
 
 * Stetefeldt's Lixiviation of Silver Ores, Netf York, 1888, p. 90. 
 ^ Ibid., p. 19. tlbid., p. 22. 
 
SULPHIDES OF GOLD. 33 
 
 SILICATES OF GOLD. 
 
 The existence of auro-silicates is now admitted without dis- 
 pute, and gold has for centuries been used to impart colour to 
 glasses, the method used being as follows : A solution of chloride 
 of gold is added to a mixture of sand with alkalies and alkaline 
 earths or lead, and the whole is then fused, and colourless or 
 yellow transparent silicates of gold thus formed. These are 
 decomposed by being reheated gently to low redness, oxides of 
 gold, or more probably metallic gold, being set free, and red or 
 purple colorations thus obtained. The occurrence of silicates of 
 gold in nature seems to be doubtful. 
 
 Experiments conducted by E. Cumenge* tend to show that 
 the alkaline auro-silicates, obtained in the wet way, may have 
 played an important part in the formation of auriferous quartz. 
 The following conclusions have been established by these 
 investigations : 
 
 1. If an alkaline aurate, obtained by dissolving auric ses- 
 quioxide in caustic soda, is mixed with an alkaline solution of 
 silicate of soda (soluble glass), the mixture may be concentrated 
 by evaporation until it has attained a syrupy consistency with- 
 out being decomposed. Auro-silicate of soda is, therefore, fairly 
 stable, so long as there is an excess of alkali present. 
 
 2. The decomposition of this auro-silicate is effected by the 
 addition of hydrochloric acid to it, by which gelatinous silica is 
 precipitated. This carries down a certain proportion of gold 
 which gives a rose colour to the white magma. 
 
 3. This decomposition may also be completely effected by the 
 action of an aqueous solution of carbonic acid under pressure. 
 Thus, if the syrupy, alkaline auro-silicate is introduced into a 
 bottle of seltzer water, which is then hermetically closed, the 
 decomposition can be seen to be gradually going on without 
 the semi-fluid mass being dissolved, and the latter is replaced 
 at the end of some days by coherent silica, which, on exposure 
 to the air, assumes a white opaline appearance tinged with rose 
 colour. 
 
 4. When gelatinous silica, obtained by the decomposition of 
 an alkaline auro-silicate, is heated to redness in a current of 
 steam, it assumes either a beautiful, unalterable rose colour, or a 
 reddish tint with visible grains of gold, according to the pro- 
 portion of precious metal present, and the conditions under 
 which the precipitation has been effected. 
 
 SULPHIDES OF GOLD. 
 
 These compounds are prepared as brown or black precipitates 
 by passing sulphuretted hydrogen through a solution of gold 
 
 * Fr6my, Ency. Chim., vol. iii., L'or, p. 62. 
 
 3 
 
34. THE.. METALLURGY OF GOLD. 
 
 chloride. The exact composition of the precipitate varies with 
 the temperature and degree of concentration of the solution, and 
 the amount of free acid present. Levol and Kriiss state that 
 Au 2 S is precipitated in the cold, but that only metallic gold and 
 free sulphur are thrown down from boiling solutions. According 
 to others, Au 2 S 3 is precipitated from cold and Au 2 S from boiling 
 solutions. It seems probable, however, that free sulphur is always 
 formed in considerable quantities, and whether the solutions are 
 hot or cold, dilute or concentrated, definite compounds are not 
 precipitated, variable mixtures of the two sulphides with free 
 sulphur and metaUic gold being formed. Similar precipitates 
 are formed by alkaline sulphides and by sulphides of most of the 
 heavy metals. The sulphides are soluble to some extent in a 
 saturated solution of sulphuretted hydrogen, and are easily 
 soluble in hot solutions of alkaline sulphides or alkalies, forming 
 double salts so that precipitation from alkaline solutions is 
 never complete. The sulphides are readily decomposed into 
 gold and sulphur by the action of heat, the decomposition 
 being complete below 270. Sulphide of gold is also dissolved at 
 ordinary temperatures by potassic cyanide, and is slowly attacked 
 by mercury with formation of mercury sulphide. 
 
 Purple of Cassius. This body was discovered by Cassius 
 of Ley den in 1683. It contains gold and oxide of tin, and is 
 used to colour glass and glazes, various shades of violet, red 
 and purple being thus obtainable. Several methods of prepara- 
 tion are used, of which the following is that employed at the 
 factory at Sevres* : Half a gramme of gold is dissolved in aqua 
 regia composed of 16 '8 grammes of hydrochloric and 10*2 grammes 
 of nitric acid, and the solution is then diluted with 14 litres of 
 water- To this solution is added, drop by drop, a solution of a 
 mixture of protochloride and tetrachloride of tin, prepared as 
 follows : 3 grammes of finely divided tin is dropped, little by 
 little, into 18 grammes of aqua regia (constituted as above, with 
 the addition of 5 c.c. water), the reaction is checked by cooling 
 if it is too violent, and the solution of chloride of tin formed 
 is allowed to cool. The precipitate of purple oxide thus obtained 
 is finely coloured when it has been washed with boiling water. 
 The purple precipitate obtained by M uller, by reducing chloride 
 of gold with glucose in an alkaline solution containing tin oxide 
 in suspension, differs from that prepared by the foregoing method 
 in losing its colour at a red heat, while the true purple of Cassius 
 becomes brick-red under such conditions. The true colour is 
 seen when metallic tin acts on trichloride of gold, or when alloys 
 of gold and tin are attacked with nitric acid. 
 
 The composition of purple of Cassius has been the subject of 
 much discussion. Some chemists have considered it to be a 
 compound of aurous oxide with the oxides of tin. Debray 
 * Fr^my's Ency. Chim., vol. v., L'or, p. 63. 
 
OCCURRENCE AND DISTRIBUTION OP GOLD. 35 
 
 regards it as a lake of stannic acid coloured by finely divided 
 gold. If this latter view is correct then the gold may be present 
 in an allo tropic modification, since (1) the purple of Cassius is 
 completely dissolved by ammonia, a purple solution being formed, 
 and the solution may be kept unaltered for weeks although it 
 is decomposed by light, or on heating, becoming bluish, and 
 finally depositing metallic gold; the ammonia may be removed 
 by dialysis, leaving a purple aqueous solution which contains 
 both gold and stannic oxide;* (2) the purple of Cassius does 
 not yield any gold on treatment with mercury. Muller confirms 
 Debray's views, showing that fine purple compounds can be made 
 with gold and magnesia, lime, baryta, sulphate of barium, &c., 
 the colour depending on the presence of finely divided gold and 
 not on the other constituent. The purple colour possessed by 
 (possibly allotropic) gold, when in a finely divided condition, 
 is further attested by the purple stain given to the fingers by 
 a solution of gold trichloride, and by the colour of Roberts- 
 Austen's aluminium alloy, AuAl 2 . 
 
 CHAPTER III. 
 
 MODE OF OCCURRENCE AND DISTRIBUTION 
 OF GOLD. 
 
 ITorms in which Gold occurs in Nature. Gold is obtained 
 from two very different sources, viz., (1) from, veins in rock for- 
 mations, and (2) from placers, or the alluvial deposits of ancient 
 <md modern streams. 
 
 Vein Gold. In this case the metal, whenever it is present 
 in visible grains or masses, has sharp angular edges, and although 
 usually not distinctly crystalline, it frequently penetrates the 
 rock irregularly in various directions, and is completely inter- 
 woven with, and attached to the matrix, usually quartz, so that 
 the metal cannot be separated from the rock without crushing 
 the latter. 
 
 The gold in lodes is sometimes in the form of crystallisations, 
 which are, however, exceedingly rare, and crystals of gold are still 
 probably unknown to most miners, although they occur more fre- 
 quently in placer deposits. Arborescent branching and dendritic 
 masses of crystalline gold are more common than single crystals 
 in both quartz lodes and placer deposits. The crystalline forms 
 met with have already been described, p. 9. In the Transyl- 
 vanian lodes, gold occurs chiefly in thin sheets or plates, often as 
 
 * Chtm. Soc. Journ.y vol. Ixiv. (1893) p. 575. 
 
36 THE METALLURGY OF GOLD. 
 
 much as from half an inch to two or more inches in breadth. 
 Such plates are rarely thicker than a visiting card, and are- 
 generally covered with crystalline lines and markings, revealing 
 a distinct geometrical structure. Gold also occurs in wire-like- 
 forms, sometimes penetrating crystals of other minerals, such 
 as calcite and dolomite. 
 
 The matrix in which the gold is contained is usually quartz, 
 intersecting as veins or interlaminated with sub-crystalline, slaty r 
 or schistose rocks, especially hydromica and chloritic slates. 
 Gold also occurs sparingly in similar veins in granite and gneiss,. 
 and has been detected in the trachytes of Colorado, and in 
 silurian and carboniferous trachytes, as well as in some lime- 
 stones. 
 
 With regard to the distribution of gold, W. P. Blake ob- 
 serves* " There is a much greater dissemination of gold in a 
 ragged granular condition, in situ, in fine particles in the midst 
 of rock formations, and without any obvious connection with 
 veins, than is generally supposed. Prominent examples are- 
 found in the belts or zones of layers of soft slate in Georgia, and 
 in North Carolina. ... The Boly- Fields gold vein, Lumpkin 
 County, Georgia, is an example of the occurrence of coarse ragged 
 gold in the midst of a mass of slate, without any defined quartz 
 vein. The gold is closely associated with bornite, pyrites, and 
 dolomite." The dissemination of gold in the schistose rocks of 
 North Carolina has also been noted by Professor Kerr,f and by 
 Dr. Emmons, and similar occurrences have been observed in Texas, 
 Nova Scotia, and in other parts of the world. At the Contention 
 Mine, Tombstone, Arizona, free metallic gold is found in the thin, 
 cracks and cleavage surfaces of partially decayed porphyry, and 
 appears to have been deposited there from solution, and not 
 mechanically. It occurs in thin subcrystalline flakes and scales, 
 and may have been derived from the decomposition of the iron 
 pyrites with which the adjoining sedimentary formations are 
 charged. Gold also occurs in small quantities (1 part in 
 1,124,000) in the bed of clay on which the city of Philadelphia 
 is built. Its occurrence in solution in sea water has been proved 
 by Sonstadt.J 
 
 The wide distribution of gold in minute quantities throughout 
 the world was pointed out by W. E. Dubois, an Assayer in the 
 United States Mint, in 1861, and is further attested by a large 
 number of specimens now in the Percy collection. These con- 
 sist of small specks of gold of different sizes which have been 
 obtained from the most varied sources. Thus, samples of Pattin- 
 son's crystallised and uncrystallised lead, pig lead from all 
 
 * Prod. Gold and Silver in United States, 1884, p. -581. 
 t Trans. Am. Inst. Mng. Eng., vol. x., p. 475. 
 $ Chem. News, vol. xxvi., p. 159. 
 Journ. Am. Phil. Soc., June 1861. 
 
UNIVERSITY 
 
 OCCURRENCE AND DISTRIBU 
 
 countries, lead fume, i-ed lead, litharge, white lead, precipitated 
 carbonate of lead and acetate of lead were all found to contain 
 gold, which seems to be invariably present in galena. Moveover, 
 it appears to be impossible to procure samples of copper in 
 which gold cannot be detected, although the Lake Superior 
 copper contains less than 1 part in 1,000,000 ; the bronze and 
 copper coins of all nations are usually found to contain much 
 greater quantities of gold than this. Similar evidence has been 
 adduced which tends to show that all ores of silver, antimony, 
 .and bismuth contain gold. * 
 
 Placer Gold and. Nuggets. Placer gold is usually in the 
 form of 'small scales, but pellets or rounded grains also occur, 
 while larger masses or nuggets are usually of a rounded mammil- 
 lated form. The chief difference between the appearance of 
 placer and vein gold lies in the fact that the former is always 
 rounded, showing no sharp edges, even the crystals having their 
 angles smoothed and rounded off. This has been pointed to by 
 the advocates of the erosion theory of the origin of placer gold, as 
 evidence in favour of their views, the roundness of the fragments 
 being taken to prove that abrasion of the gold has been effected 
 by attrition with water and grains of sand. The largest masses 
 of gold yet discovered have been found in auriferous gravel. 
 The " Blanch Barkley " nugget, found in South Australia, 
 weighed 146 pounds, and only six ounces of it were gangue; 
 -and one still larger, the " Welcome " nugget, from Victoria, 
 weighed 2,195 ounces, or 183 pounds, and yielded gold to the 
 value of 8,376 10s. 6d. In Russia a mass was found in 1842 
 near Miask, weighing 96 pounds troy. The largest mass from 
 California is given in the State Mineralogist's report as weigh- 
 ing 2,340 ounces, or 195 pounds, but no authentic cases seem 
 to be on record of nuggets from this State weighing more than 
 20 pounds. 
 
 The minerals most common in auriferous quartz lodes or in 
 the placer deposits are platinum, iridosmine, magnetite, iron 
 pyrites, galena, ilmenite, copper pyrites, blende, tetradymite, 
 zircon, garnets, rutile and barytes ; wolfram, scheelite, brookite 
 and diamonds are less common. Diamonds are associated with 
 gold in Brazil, and also occasionally in the Urals and in the 
 United States. The sulphides present in auriferous quartz 
 frequently contain gold ; the gold in such an ore is usually in 
 part quite free, disseminated through the quartz, in which visible 
 grains of the metal often occur, and in part locked up in the 
 pyrites, whence but little can in general be extracted by mercury. 
 It is, however, in all probability in the metallic state in pyrites, 
 although this is not completely established. The subject is 
 discussed in Section on " Gold in Pyrites," Chap. vii. 
 
 * Loc. cit., and E. A. Smith on Bismuth, &c., Journ. Soc. Chem. Ind. t 
 vol. xii. (1893), p. 216. 
 
38 THE METALLURGY OP GOLD. 
 
 Among minerals other than sulphides which contain gold, 
 the following may be mentioned, although none are of great 
 importance in metallurgy : Calaverite is a bronze - yellow 
 gold telluride, usually containing a little silver, occurring in 
 certain mines in California and Colorado. One analysis gave 
 tellurium 55-5 per cent., gold 44-5 per cent., corresponding to 
 the formula AuTe 2 . Sylvanite, called also graphic tellurium, 
 is a telluride of gold and silver, supposed to correspond to- 
 (AuAg)Te 3 . It sometimes contains antimony and lead in ad- 
 dition. It is from white to brass-yellow in colour, and the 
 arrangement of the crystals sometimes bears a resemblance to- 
 writing characters, whence the name graphic. It occurs in 
 Transylvania, in Calaveras County, California, and in Colorado. 
 
 Petzite is a telluride of silver, Ag Te, in which the silver is 
 partly replaced by gold. A specimen from the Golden Rule 
 Mine, according to Genth, contained tellurium 32-68 per cent.,, 
 silver 41-86 per cent., and gold 25-60 per cent. It occurs in 
 Transylvania, Chili, California., Colorado, and Utah. 
 
 Nagyagite or Foliated Tellurium is remarkable for being 
 foliated like graphite, which it also resembles in its colour, a 
 blackish lead-grey, and in having a hardness of from 1 to 1-5 only. 
 Its density, however, is above 7. It occurs in Transylvania, and 
 contains tellurium 32-2, lead 54-0, gold 9-0 to 13-0 per cent., some- 
 times with silver, copper and sulphur in addition. Other gold 
 tellurides and some native gold amalgams are occasionally met 
 with, but these minerals in which gold is an essential part are 
 rarely of much importance as ores, as they seldom occur in 
 sufficient abundance to be regarded as anything but specimens 
 for collectors. In some few mines, however, notably at the 
 Cripple Creek district, Colorado, in Transylvania, in Boulder 
 County, Colorado, and at the Bassick mine in the same State, 
 the value of the ore depends on the tellurides of gold con- 
 tained in it. 
 
 One of the most striking differences between the ores of gold 
 and those of all other metals lies in the extremely small propor- 
 tion which the desired material bears to the worthless gangue 
 with which it is accompanied. Occasionally hand specimens 
 of vein stuff are found containing several per cent, of the 
 precious metal, but these are of quite exceptional occurrence 
 and have not the slightest economic importance. The greater 
 part of the vein gold now being produced is derived from ores 
 containing only about one part of gold in seventy or eighty 
 thousand, whilst, under exceptional circumstances, a yield of one 
 part in half a million parts of gangue may give handsome profits. 
 Placer deposits are usually much less rich than this; the average 
 amount of gold contained in those now worked does not exceed 
 one part in one million, and in California deposits of gravel with 
 only one part of gold in fifteen millions have proved susceptible 
 of successful treatment by the " hydraulicking " method on a. 
 large scale. 
 
OCCURRENCE AND DISTRIBUTION OP GOLD. 39 
 
 Composition of Native Gold. Native gold usually contains 
 silver, which occurs in varying proportions, the colour becoming 
 paler with the increase of silver. The finest native gold yet 
 found is that from the Mount Morgan Mine, Queensland, which 
 is 997 fine. The finest Russian gold was that formerly obtained 
 at Katerineburg in the Urals, and yielded gold 989 '6, silver 
 1*6, copper 3-5, and iron 0*5 (G. Rose). Gold dust from West 
 Africa has been found to contain 978*1 of fine gold, the remainder 
 being silver. The gold found in the British Isles varies from 
 800 to 900 fine, the remainder being silver ; the specimens from 
 the district around Dolgelly are sometimes a little over 900 fine. 
 Particulars of the fineness of gold from Australia, California, 
 &c., are given in the chapter on Refining, Chap, xviii. Gold is 
 occasionally found alloyed with copper, and sometimes also with 
 iron, bismuth, palladium, or rhodium. Rhodic-gold from Mexico 
 was found to be of the specific gravity 15-5 to 16 '8 and contained 
 34 to 43 per cent, of rhodium. Bisrnuthic-gold has been called 
 Maldonite* 
 
 Geographical Distribution of Gold. Gold occurs in lodes 
 in many districts composed of partially metamorphosed rocks 
 such as slates or schists, while its occurrence in holocrystalline- 
 metamorphic, or igneous rocks is comparativaly rare. Among 
 sedimentary rocks, its occurrence is almost confined to the sands 
 of rivers which run for a part of their course through crystalline 
 formations, or more particularly through districts in which gold 
 occurs in quartz veins. Such river sands are rarely quite free 
 from gold. The beds of ancient rivers no longer existent are 
 also frequently auriferous. In spite of the fact that the sea 
 contains gold in solution, the aggregate amount perhaps 
 exceeding that contained in the accessible portion of the earth's 
 crust, nevertheless, unaltered marine deposits seldom or never 
 contain a perceptible quantity of the metal, except in one or two 
 cases of beach deposits formed by the erosion of auriferous land- 
 formed gravels. 
 
 In the British Isles, gold is found in some of the streams of 
 Cornwall and in lodes and river gravels near Dolgelly and in 
 other parts of Wales, in Sutherland shire and near Leadhills in 
 Scotland, and in the County of Wicklow. The total amount which 
 has been obtained from these localities is small, probably not 
 exceeding 40,000 ounces, and little is now being produced. On 
 the Continent of Europe, gold is most abundant in Hungary and 
 Transylvania, where the gold occurs in quartz lodes contained in 
 eruptive rocks of tertiary age, chiefly propylite, porphyry, diorite 
 and granite. The minerals occurring with the gold are galena, 
 blende and pyrites. In the German Empire, the gold obtained 
 is chiefly derived from the smelting of argentiferous galena in 
 which small quantities of the more precious metal are contained. 
 * Dana's Mineralogy, p. 108. 
 
40 
 
 THE METALLURGY OF GOLD. 
 
 In Italy the only important mines are those of Pestarena and 
 Yal Toppa in North Piedmont, near Monte Rosa. The pyritic 
 ores from these mines are treated by amalgamation. Gold is 
 also found in the sands of the Rhine, the Reuss, the Aar, and 
 other rivers, and in small quantities in Sweden. 
 
 The gold-bearing districts of Russia are (1 ) the Urals, (2) Siberia, 
 Eastern and Western, whilst an insignificant amount is also 
 derived from Finland and from the Caucasus. The gold was for- 
 merly derived chiefly from lodes both in the Urals and in Western 
 Siberia. In the Urals, quartz-mining began in 1745, and the 
 output from this source continually increased up to the year 
 1810, when it began to fall off, and has been trifling since 1838. 
 The placers, which are mainly situated on the eastern slope of 
 the range, were discovered in 1774, and the yield has continually 
 increased since then up to the present time. The Siberian placers 
 were discovered in 1829, although quartz-mining had been prose- 
 cuted since 1704.* The output from these continually increases, 
 in spite of the exhaustion of the old placers and the reduction in 
 the percentage of gold contained in those still worked ; this 
 increase of output is in consequence of the discovery of new 
 placers further east. The auriferous gravels are all thin, shallow 
 deposits, ranging from 3 to 20 feet in thickness, and as they are 
 worked out, other gravels are opened-up further east, so that 
 operations are being gradually transferred from west to east. 
 When the exhaustion of these placers has proceeded further it 
 may be expected that more attention will again be paid to the 
 quartz lodes. The relative amounts produced by the different 
 districts are given as follows : f 
 
 
 Ural, 
 per Cent. 
 
 Western Siberia, 
 per Cent. 
 
 Eastern Siberia, 
 per Cent. 
 
 Finland, 
 per Cent. 
 
 187680, . 
 
 20'0 
 
 6-0 
 
 74-0 
 
 0-02 
 
 1890, . 
 
 28-0 
 
 6-4 
 
 65-4 
 
 o-04 
 
 Almost the whole of the production is now derived from 
 shallow alluvial deposits by means of sluicing operations, as 
 described on pp. 59-62. 
 
 In India, almost all the gold now being produced is derived from 
 the quartz lodes of the Colar gold-field, Mysore, in Southern 
 India, in which work was begun in the year 1880. The ore is 
 free milling, and is treated by amalgamation in the stamp battery. 
 A little gold also comes from the Presidency of Madras. In 
 China and Japan considerable quantities of gold are produced, 
 
 * Mintral Industry for 1892, p. 203. 
 
 t Prod, of Gold and Silver in U.S.A., 1886, p. 81: Mineral Ind., 1892, 
 p. 203. 
 
OCCURRENCE AND DISTRIBUTION OP GOLD. 41 
 
 chiefly by various primitive methods; little is known of the 
 methods used and of the amount produced in the former country. 
 
 From the United States a large percentage of the total gold 
 production of the world is obtained. The chief producing States 
 are California, Colorado, Dakota, Montana, Nevada, Idaho and 
 Oregon, but smaller amounts come from many other States. The 
 produce is now far more from lodes than from placer deposits, 
 and in the treatment of auriferous quartz and pyritic ores almost 
 all the known methods of treatment are applied in different 
 localities.* Mexico produces considerable quantities of gold, 
 and gold ores are also found in various parts of Canada, Colombia, 
 Bolivia, Chili, Venezuela, Brazil, Peru, and the small States of 
 Central America. The production of several of these countries 
 was formerly much larger than it is at the present day, the 
 reduction being especially marked in the cases of Brazil and 
 Venezuela. On the other hand, the gold districts of British 
 and French Guiana, of the Argentine Republic and of Uruguay 
 are now being developed, and may eventually prove to be of 
 considerable importance. 
 
 Gold is somewhat widely distributed in Africa, the chief 
 sources of production in former times being the placer deposits 
 of the Gold Coast and Abyssinia. The discoveries of auriferous 
 quartz deposits in the Transvaal since 1884 have converted that 
 region into one of the most important among gold producing 
 countries. The richest deposits are in the Banket formation on 
 the Witwatersrand, whence about nine-tenths of the gold 
 obtained in the Transvaal is derived, the De Kaap field being 
 next in importance. No placer deposits of value have been 
 discovered, and the gold is mainly obtained by stamp-battery 
 amalgamation, followed by treatment of the tailings by means 
 of the cyanide process. The produce from both sources is being 
 continually increased. The production by means of treatment 
 with cyanide is now about 30 per cent, of the whole. 
 
 Gold is found in all the Colonies of Australia, together with 
 Tasmania and New Zealand ; the largest amount is now pro- 
 duced by Victoria and Queensland, Western Australia and New 
 South Wales follow in the order named. The chief gold pro 
 ducing districts of Queensland are Charters-Towers, Rockhampton 
 (where the Mount Morgan mine is situated), Croydon, and 
 Gympie. Almost all the gold is produced from the quartz 
 mines, the placers having been practically exhausted ; thus in 
 1891; out of a total production of 576,439 ozs., 560,418 ozs. 
 were produced from quartz, and only 16,021 ozs. from alluvial 
 deposits, while in 1877 the amounts were Placer gold, 164,778 
 ozs.; quartz gold, 188,488 ozs. total, 353,266 ozs.f 
 
 * For full details as to the geographical and geological distribution and 
 the nature of the gold ores found in the United States, the student is 
 referred to the Annual Reports of the Director of the United States Mint, 
 and of the Californian State Mineralogist. 
 
 t Mineral Industry, 1892, p. 192. 
 
42 THE METALLURGY OF GOLD. 
 
 The production of gold in Victoria is now increasing, but is far 
 less than formerly, owing to the exhaustion of the alluvial deposits, 
 although the yield of the quartz mines has continuously increased. 
 
 In 1894 the production was divided as follows : Alluvial 
 gold, 254,308 ozs.; reef gold, 419,371 ozs. total, 673,680 ozs. 
 The chief producing districts are Ballarat, Sandhurst (Bendigo) r 
 Beechworth, Maryborough, Castlemaine, Gippsland and Ararat. 
 In 1894 in Victoria the average yield per ton of quartz crushed 
 was 8 dwts. 8 grains of gold, this being slightly less than the- 
 mean yield during the last decade. 
 
 In New South Wales about half the gold is obtained from 
 quartz lodes ; the pyritic ores are not yet effectively treated, and 
 the river bank deposits have not up to the present been exploited 
 on a large scale.* In New Zealand, only quartz lodes are worked 
 in the North Island, alluvial workings being confined to the 
 Middle Island. The most important districts are those of 
 Kuaotuna, Thames, Coromandel, Waihi and Reefton in the 
 North Island, and Ross and Kumara in the Middle Island. f 
 
 Origin of G-old Ores. The origin of mineral veins, includ- 
 ing those in which gold is contained, has long been discussed 
 by geologists. The old theory that the quartz of veins was 
 originally in a molten condition and was ejected from below into 
 fissures is no longer maintained, although in 1860 H. Resales 
 brought forward evidence in its favour as far as the Victorian 
 lodes are concerned. It is now admitted by all that the 
 materials forming the veins have been transported in aqueoua 
 solution and precipitated where they occur. In a few excep- 
 tional cases, sublimation may have played a part. The view 
 that the solutions found their way downwards from above has- 
 been abandoned, but the ascensional theory and the lateral 
 secretion theory both have many adherents. The last-named 
 theory has found its principal supporter during the last twenty 
 years in Prof. F. von Sandberger, who pointed out that the 
 gangue of many lodes varies in composition if the nature of 
 the rocks through which they pass is changed, and claimed to 
 have proved by analysis that the materials forming vein-stone 
 are derived from the adjacent country rocks. He stated, more- 
 over, that such minerals as augite, hornblende, mica, and 
 olivine, which are essential constituents of crystalline rocks, 
 contain small quantities of the heavy metals occurring in veins. J 
 Although Sandberger did not try to detect gold in the silicates, 
 this metal is not likely to be an exception. Prof. A. Stelzner 
 objected to these conclusions, urging that small quantities of the 
 sulphides of the heavy metals were probably mechanically mixed 
 
 * Report of the Dept. of Mines and Agriculture, N.S. W., 1894. 
 
 t Annual Report of the Mining Commissioner of New Zealand, 1891. 
 
 Untersuchungen fiber Erzgange. Wiesbaden, 1882 and 1885. Useful 
 abstracts are given in Phillips' Ore Deposits and in Le Neve Foster's Ore 
 and btone Mining. 
 
SHALLOW PLACER DEPOSITS. 43- 
 
 with the crystals of minerals which Sandberger analysed in the 
 belief that they were pure. Stelzner advocated the retention of 
 the ascentional theory, which alone affords a satisfactory expla- 
 nation of the difference in composition observable in neighbour- 
 ing lodes passing through the same rocks, and apparently formed 
 at different periods. The two theories are, however, not con- 
 tradictory, and perhaps neither need be entirely rejected, the- 
 solutions being supposed to pass more or less freely in the plane 
 of the lode alter they have been impregnated. For the origin, 
 of placer gold, see p. 68. 
 
 CHAPTER IV. 
 TREATMENT OF SHALLOW PLACER DEPOSITS. 
 
 THE deposits grouped together under the name of " placers " 
 comprise sands, gravels, or any loosely coherent or non-co- 
 herent alluvial beds containing gold. They have accumulated, 
 owing to the action of running water, in the beds of rivers, or on 
 the adjoining inundation plains, or on sea beaches. They fall 
 naturally into two groups, between which no strict line of 
 demarcation exists. These are 
 
 (1) Shallow or modern placers, which are in or near existing 
 rivers, and have not yet been covered by other deposits. 
 
 (2) Deep level or ancient placers, which now lie buried beneath 
 an accumulation of debris or coherent rock, the rivers by which 
 they were formed having often been deflected into other channels 
 by more or less extensive changes in the physical geography of 
 the district in which they existed. Beach deposits occur in each 
 subdivision. 
 
 In this chapter, the first of these groups will be considered. 
 
 From the earliest times to the present day, shallow placer 
 deposits have probably yielded more gold in the aggregate than 
 has been derived from all other sources put together. 
 
 Shallow placers consist of loose aggregations of sand, gravel, 
 loam and clay, accumulated by the action of existing rivers and 
 streams, and not extending to a greater depth than 10 or 15 feet 
 from the surface. They contain metallic gold in fragments of 
 all sizes, ranging from the finest dust to nuggets weighing 
 thousands of ounces. Auriferous sands are found in the beds of 
 most rivers which flow during any part of their course through a 
 region composed of crystalline rocks. If the rivers have rocky 
 beds, gold may be found in the crevices, caught in natural riffles, 
 and the whole may subsequently be covered by beds of sand. 
 
 Gold also occurs in river bars and banks, in river " flats," or 
 inundation plains, in the dry bed of streams which only flow 
 
44 THE METALLURGY OF GOLD. 
 
 after heavy rains (" gulch diggings "), in terrace gravels on the 
 sides of valleys high above the present level of the water ("bench 
 diggings "), and on the sides and tops of hills (" hill diggings "). 
 The last two subdivisions are evidently ancient rather than 
 modern deposits. The gravels may contain boulders of any 
 size, up to several feet in diameter, or may shade off into fine 
 sand, while sandy clays, especially if on the bed-rock, are 
 frequently very rich. In the Urals, the placer deposits often 
 consist of heavy clays, while others are formed of waterworn 
 fragments of auriferous quartz, talcose and chloritic schists, 
 serpentine, greenstone, &c. Gold occurs under very various 
 conditions in these deposits. It may occur in the grass roots on 
 inundation plains, or near the surface of the gravels in river 
 beds, or dispersed through the whole thickness of a stratum. 
 More commonly, however, the lowest part of the superficial beds, 
 just above " bed-rock " (the country rock of the district), is 
 richest. In hollows, cracks, and crevices of the bed-rock, or, 
 if it is soft and decomposed, in the substance of the upper part 
 of the rock itself, to the depth of 1 or 2 feet, gold occurs in 
 the greatest quantities. In pipeclays just above bed-rock, in 
 Victoria, it was not uncommon to find 12 ozs. of gold or more in 
 a single tubful of "dirt," and similar rich bed-rock deposits 
 occur in California. The depth at which bed-rock is found 
 varies greatly ; it may crop out at the surface, or it may be 
 buried beneath hundreds of feet of gravel, and great variations 
 occur even in a single district. In the Urals, however, the 
 thickness of the gravel is usually less than 3 feet thick, and is 
 rarely more than 12 feet. 
 
 Methods of obtaining Gravel from Shallow Placers. 
 The gravel obtained from any placer deposit is, with exceptions, 
 ultimately all treated alike, but the mode of winning the dirt 
 varies with the necessities of the case. On flats and bars, the 
 surface gravel, if rich enough, is loosened with pick and shovel, 
 and then washed. If only the part just above the bed-rock will 
 pay for treatment, it is reached by " stripping," or, if covered by 
 too great a thickness of barren material, shafts are sunk, and 
 short levels run from the bottoms of them in all directions. If 
 the nature of the ground permits of it, tunnels are run without 
 shafts, and the rich gravel is then followed from the surface, 
 wherever it is found. This system was much practised in the 
 early days in California, although now seldom to be seen in 
 operation ; it was called " coyoting," from the coyote, which lives 
 in holes in the ground. When water was encountered in the 
 shaft, it was drawn out by a bucket, until it came in too fast, 
 when the claim was abandoned.* In California, in somewhat 
 later times, efforts were made to reach the gold in the river beds. 
 
 * In Siberia and in the Yukon district the gravels are perennially frozen, 
 and are thawed with wood fires set at the ends of the drifts, and then taken 
 out with pick and shovel. 
 
SHALLOW PLACER DEPOSITS. 45 
 
 by deflecting the streams of water from their courses, and in 
 other ways. These methods will be briefly described under the 
 head " River Mining." 
 
 Methods of Washing the Gravel. The Pan. When the 
 existence of gold in the placers of California and Australia first 
 became known, the diggers were not acquainted with any appar- 
 atus which was well adapted to extract the metal. The house- 
 hold pan was used everywhere to wash the gravel, and though,. 
 in its original form, it was a difficult implement to use efficiently, 
 it has retained its place in both countries for prospecting and also- 
 for washing small quantities of rich material, however they may 
 have been obtained. The pan is usually made of stiff sheet-iron, 
 is flat-bottomed and circular, and at bottom about 14 inches 
 in diameter. The sides slope outwards at an angle of about 30 
 to the bottom, and are about 5 inches wide. A riffle is a useful 
 addition, formed by the thickening or bulging inwards of the 
 side, situated about half-way up the latter and running about 
 half-way round the pan. The method of using this pan embraces 
 several operations. First, it is filled to about two-thirds of its 
 capacity with pay-dirt, of which it then contains from one- 
 fifth to one-quarter of a cubic foot. It is then placed at the 
 bottom of a water-hole or convenient stream, and the dirt is 
 thoroughly broken up with both hands, care being taken not to 
 leave any lumps of clay. As soon as the contents of the pan are 
 reduced to the consistency of soft mud, the pan is grasped with 
 both hands a little behind its greater diameter, inclined away 
 from the operator, raised until the dirt is only just covered with 
 water, and shaken from side to side, while a slight oscillatory 
 circular motion is also imparted to it. The mud and fine sand is 
 soon obtained in suspension in the water, and gradually passes 
 over the far edge, which is lowered more and more, until little 
 but the stones, coarse particles of sand, black sand and gold is. 
 left. The larger stones lie on the top and are removed by hand. 
 The final stage consists in lifting the pan with a little water in 
 it, and by a movement of the wrist, something like that used in 
 vanning, causing the material to be spread out by the water, in 
 a comet shape, in the angle of the pan. The "colours" i.e., 
 yellow specks of gold are seen at the extreme head of the comet, 
 and also occur in the succeeding inch or two, mixed with the 
 black sand, while the quartz-sand forms the remainder of fche 
 tail and is scraped or washed off. The gold is separated from 
 the black sand by (a] amalgamation with mercury, or (b) drying 
 and blowing away the black sand, a wasteful process. Liquid 
 amalgam is readily separated from sand, and the mercury is 
 then driven off by heat. 
 
 The Batea differs from the miner's pan in not having a flat 
 bottom. It is of wood turned in a lathe, about 20 inches in 
 diameter, conical, or more rarely basin shaped, and about 3- 
 
46 THE METALLURGY OP GOLD. 
 
 inches deep in the centre, so that the angle at the apex is about 
 160. The gold collects at the lowest point and clings to the 
 wooden surface under conditions when it would slide over iron. 
 The batea consequently is more rapid and effective in obtaining 
 .a " prospect" than the pan, especially when the gold is fine, but is 
 less frequently used in the United States and Australia. It had 
 its origin in South' America. It is usually now made of enam- 
 elled iron with a hole in the centre fitted with a cork, when it 
 is for use in countries other than South America. It is con- 
 sidered by MM. Oumenge and Fuchs to be especially favoured 
 by the negro race. 
 
 Prospecting Trough. This instrument is used in the far east, 
 especially by the Chinese, Malays, Annamites, &c. It is made 
 of wood, and is shaped in the form of a very flat reversed roof- 
 top, the angle between the long sides being about 150. In 
 place of a circular movement of the water an alternating rocking 
 motion is used, the water flowing up and down. The instrument 
 is easily handled, but is very slow. 
 
 Horn Spoons cut out of black ox-horns have been used by 
 prospectors, especially to finish the work begun by the pan. 
 The surface holds the gold well and shows " colour " very 
 readily. 
 
 The Cradle or Rocker was introduced in California soon after 
 the first rush to the diggings took place in 1849. It consists of 
 .a rectangular wooden box, about 3 feet long and 18 inches wide, 
 resting on two rockers (D, Fig. 4) similar to those used for 
 infants' cradles. The 
 shape of the walls is 
 shown in Fig. 4, which 
 is a section of the ap- 
 paratus. The method of 
 using it is as follows : 
 
 The gravel is shovel- 
 led into the riddle-box, 
 A, the bottom of which ~ ^~ 
 
 consists of J inch mesh Fig. 4. 
 
 screen; theworkman sits s ale * in - = 18 inches - 
 
 by the side of the machine and rocks it with one hand, while he 
 pours on water by means of a dipper filled from a water-hole with 
 the other. The dirt is disintegrated and carried through the riddle, 
 and falls on the apron, B, which consists of blanketing. Here 
 the fine gold is caught, and the dirt then passes out from back 
 to front over the bottom, which is slightly inclined towards the 
 front, and the coarse gold, black sand, &c., is caught in two or 
 three riffles, C, each of about 1 inch in height, to which mercury 
 is sometimes added to assist in retaining the gold. The rocking 
 motion not only assists in the disintegration of the dirt, which 
 .is effected by the water, aided by the stones, but also prevents 
 
SHALLOW PLACER DEPOSITS. 47 
 
 the sand from packing behind the riffles ; in the event of this 
 happening, gold would pass over the surface of the sand and be 
 lost. Consequently the rocking should be quite continuous, 
 since, after every pause, the sand in the riffles must be stirred 
 up before recommencing. It is, therefore, desirable for two men 
 to work together at the cradle, one to carry the gravel and 
 charge it into the hopper, and to remove the large stones from 
 the latter by hand, while the other man rocks the cradle and 
 pours on water. It requires three or four parts of water to wash 
 one part of gravel, and it is, therefore, better to carry the ore to 
 water than to carry water to the ore. When a clean-up of the 
 cradle is desirable, the riddle is removed, the apron is taken out 
 and washed in a bucket, and the accumulations behind the "riffles 
 are scraped out and panned. Most of the fine gold in the dirt 
 is lost by the cradle, and two men working together can only 
 wash from 3 to 5 cubic yards per day, according to the nature of 
 the dirt. 
 
 The Long Tom replaced the cradle in California after a short 
 time, and was used there for some years, while it is still in 
 operation in parts of Australia and Dutch Guiana. It consists 
 of a sluice-box or trough (A, Fig. 5) about 12 feet long, 20 inches 
 
 Fig. 5. 
 
 at the upper end, and 30 inches at the lower end, and 9 inches 
 deep, with an inclination of about 1 inch to the foot. The lower 
 end of the trough is cut off at an angle of about 45 and closed 
 by a screen of punched sheet iron, B, which prevents large stones 
 from passing through it. Below the screen is the upper end of 
 the riffle-box, C, which is about 12 feet long, 3 feet wide, and at 
 about the same inclination as the upper trough. It is fitted 
 with several riffles, which are sometimes supplied with mercury. 
 In working, a stream of water enters at the upper end of the 
 sluice-box, into which gravel is continually shovelled, while a 
 man breaks up the lumps with a fork, removes the large stones, 
 and puddles the lumps of clay. Two to four men can work at 
 one torn, and wash about five times as much in a day as can be 
 done by one or two men with the cradle. Only the coarse gold 
 is caught, and the machine is only suitable for washing small 
 quantities of rich dirt where there is a plentiful supply of water. 
 The material caught by the riffles is scraped out occasionally and 
 panned, but the riffle-box is too short for close saving of the gold. 
 The Puddling-tub. When water is scarce, as was the case in 
 many places in Australia where rich gravels were found, the 
 
48 
 
 THE METALLURGY OP GOLD. 
 
 long-torn is inadmissible, and the puddling-tub is resorted to. 
 This is particularly well adapted for washing clays, and is still 
 used to disintegrate lumps of clay encountered in sluicing opera- 
 tions. It consists of one half of a barrel which has been sawn 
 in two ; into this the dirt is dumped and stirred up with water 
 by means of a rake, until all the clay is held in suspension in the 
 water, when a plug a few inches from the bottom is removed, 
 and the slime run off. The operation is repeated until the tub 
 is filled with gravel and sand to the level of the plug-hole, and 
 this residue is then shovelled out and washed by the pan, the 
 cradle, or by sluicing. Large boxes were used in Australia in 
 this way in early days, the rakes being worked by horse- or 
 steam-power; in 1860 no less than 3,958 boxes, worked by 
 horses, were in use in Victoria alone.* 
 
 The cradle, long-torn and puddling-tub are now little used 
 except by the Chinese, who gain a precarious livelihood with 
 their help by washing over the heaps of tailings accumulated from 
 sluicing or hydraulicking operations in Australia and California. 
 
 The Siberian Trough.^ In Siberia, in the Urals and in the 
 valleys of the Obi, the Yenisei and the Lena, individual workers- 
 
 A 
 
 C 
 
 u 
 
 T7 
 
 Fig. 6. 
 
 still exclusively use a trough, which differs from the long-torn 
 mainly in requiring more constant attention on the part of the 
 operator, and which resembles the old German buddle. The 
 trough consists of a rectangular box open above and at one end. 
 When sandy gravels are being treated, the bottom of the disinte- 
 gration box (A, Fig. 6) which is about 40 inches square, is 
 made of a perforated screen of wood or sheet-iron, having holes 
 of from | inch to 1 inch in diameter. The dirt is shovelled 
 into this box, and, contrary to cradle-practice (see p. 45) if 
 the gold is present in fine flakes, mercury is added here also, 
 the amount depending on the richness of the auriferous ma- 
 terial as determined by assay, the proportion used, however, 
 being never more than 10 of mercury to 1 of gold. Water is 
 directed upon the charge in the box, either by pipes from a 
 
 * Philips' Met. of Gold and Silver, 1867, p. 139. 
 
 t For further particulars see the account given by Cumenge and Fuchs 
 in Fr^my's Ency. Chim., vol. v., L'or, 1st Section, p. 12; and also Levat's- 
 L'or en Siberie Orientate, Paris, 1897. 
 
SHALLOW PLACER DEPOSITS. 49 
 
 reservoir or more often by pumping, and the fine material is 
 carried through the screen and falls on to the table, B, while the 
 pebbles are collected by hand and thrown away. If clay is being 
 treated, no screen is used; the lumps are puddled in the box, and 
 the mud carried over by an overflow of water. The table is 
 slightly inclined, about 20 feet long, and, for the greater part of 
 its length, is about 20 inches wide. It is furnished with five 
 riffles, of which two (E, C) near the top are about 2 inches high, 
 while the others (D, E)are of less height. The disintegration of 
 the sand is completed on the table with the aid of a small rake con- 
 tinually used by the workman. When disintegration is complete, 
 the stream of water is diminished in amount, and the workman 
 continues to rabble the sands which have accumulated above the 
 riffles, pushing the contents of the lower riffles up the table again, 
 until the water runs clear, and little except pyrites is left 
 behind the riffles where the so-called " grey concentrates " 
 accumulate. These are either concentrated further on the same 
 table, or removed and worked on a smaller table. In either 
 case the stream of water is still further reduced, being graduated 
 so as to carry away the last particles of quartz, together with all 
 materials of moderate weight, such as garnets, rutile, tourmaline, 
 &c., and even all the fine pyrites. If mercury has not been 
 added previously, it is sprinkled on before this last operation, 
 unless the gold is very coarse, when no mercury is added at any 
 stage of the proceedings. 
 
 The "black concentrates," thus obtained, consist entirely of 
 amalgam, magnetite, and the large grains of pyrites. The final 
 operation, by which the amalgam is separated, is the most difficult, 
 and requires the greatest amount of skill on the part of the 
 operator. The material is worked on the same table with very 
 little water, with the aid of a small rake, or more often with the 
 hand of the workman, who kneels down by the trough for the 
 purpose. Finally, all the pyrites having been washed away, the 
 magnetite is removed with a magnet, and the amalgam collected. 
 The tailings from the black concentrates are treated over again, 
 together with the grey concentrates. 
 
 The apparatus just described treats about 500 Ibs. of sand at 
 one time, and can be worked by one man, but usually gives 
 employment to four people (of whom three are frequently women), 
 who can treat about 5 tons of sand per day. The degree of 
 success attained depends largely on the skill of the workman ; 
 in Siberia and Russia the art is handed down from father to son, 
 certain families devoting their whole lives to the work during 
 many generations. These workmen attain such a degree of 
 dexterity in the use of the trough, that practically the whole of 
 the valuable contents of the gravels treated are extracted by 
 them, but the work is only suited to those who are content with 
 small earnings. 
 
50 
 
 THE METALLURGY OF GOLD. 
 
 The Sluice. The sluice has replaced all these implements for 
 washing the gravel from shallow placers, where water is abundant. 
 Sluices are constructed of " boxes," each of which resembles the 
 upper part of the long-torn. The bottom of each box is made of 
 rough boards, 'about 12 feet long and 1 J inches thick, cut 4 inches 
 wider at one end than at .the other; the total width is usually 
 from 16 to 18 inches, while the sides are 8 or 10 inches high. 
 It is considered by some New Zealand experts that sluice boxes 
 are usually made too narrow, so that the current is too deep and 
 fine gold is lost, while the flow is unduly checked by the sides, 
 thus creating undesirable eddies. The box is held together with 
 nails, and no attempt is made to render it watertight, as the 
 swelling, caused by the absorption of water, and the filling up of 
 chinks with sand and clay soon stops all leakages. The narrow 
 end of the box fits into the wide end of that next below it, and 
 so a sluice, made up of hundreds of boxes, can be rapidly put up 
 or taken down and moved to another locality. 
 
 In all but the smallest and cheapest sluice boxes, extra strips 
 of wood are affixed to the sides so as to protect them from the wear 
 
 Fig. 7. 
 Scale = &. 
 
 caused by the attrition of the stones and gravel carried through 
 by the current. When worn thin, these strips are replaced. 
 The bottom is similarly protected by the riffle bars, whose main 
 function is to catch the gold. These riffle bars are usually placed 
 longitudinally, and are strips of rough wood from 2 to 4 inches 
 thick, from 3 to 7 inches wide, and about 5J feet long. They are 
 wedged in the boxes at a distance of 1 or 2 inches apart by 
 means of transverse bars., so that two sets of riffles are placed in 
 each box in the manner shown in Fig. 7, which represents the 
 whole of one box and parts of two others. The rectangular 
 depressions thus formed between the bars are well adapted to 
 
SHALLOW PLACER DEPOSITS. .-6-1 
 
 Intercept all heavy particles that pass down the sluice, such as 
 gold, mercury, amalgam, pyrites, &c., which gravitate to the 
 bottom of the stream. Sometimes the riffles are placed trans- 
 versely, and sometimes for a short distance in 2jig-zag fashion. 
 This arrangement does not retain anything, but affords a better 
 chance of amalgamating the gold, which, together with the 
 mercury (in this case fed in constantly), slides down the inclined 
 riffles from side to side of the sluice, having sunk to the bottom 
 .by virtue of its high specific gravity.-. {. : 
 
 When all is ready, a stream of water is turned into the head 
 of the sluice, where the dirt is shovelled in also. The amount 
 of dirt shovelled in per man depends on the height of the 
 lift and the nature of the soil, as well as on the labourer. 
 It varies from 3 to 7 cubic yards per diem, and should average 
 5 or 6. The first gravel sluiced fills up most of the riffle depres- 
 sions, leaving enough inequalities of surface, however, to inter- 
 cept and retain the mercury, &c. 
 
 The length of the sluice varies with the consistency of the 
 gravel, the fineness of the gold, the capital available, and the 
 fall of the ground. It must be sufficient to complete the dis- 
 integration and then to catch the gold. The length may be 
 adjusted by experiment; if, in the clean-up, the lowest boxes 
 yield much amalgam, an addition to the length is necessary, 
 while if they yield none, the sluice may be shortened by the 
 removal of one or more boxes. Even in the latter case, however, 
 the tailings would almost certainly contain some fine gold. The 
 grade of the sluice is measured by inches per box, so that a grade 
 of " 12 inches" means one of 12 inches in 12 feet. The usual 
 grades are from 8 inches to 20 inches per box, depending 
 
 (1) On the fall of the ground, since the sluice cannot be raised 
 far above it, nor sunk deep into it, owing to the increased 
 expense thereby occasioned. 
 
 (2) On the nature of the dirt to be washed. Tough, tenacious, 
 clayey, or cemented gravels require higher grades to effect their 
 disintegration than loose material. Instead of being disintegrated, 
 clay sometimes becomes aggregated into balls, which roll down 
 the sluices, picking up particles of gold previously caught in the 
 riffles, and these lumps of clay must be removed by hand and 
 puddled. There must be sufficient grade to enable the water to 
 carry away all but the largest stones, so as to avoid unnecessary 
 hand-picking, but on the other hand, while coarse gold is readily 
 caught, fine particles are lost if the current is too rapid. 
 
 (3) On the quantity of water available. The reduction of the 
 grade lessens the duty of the water, so that if the supply of the 
 latter is short or costly, the grade is made as steep as possible, 
 consistent with saving a fair proportion of the gold. A steep 
 grade reduces the necessary length of the sluice, as disintegration 
 takes place sooner. Since a steep grade, a rapid flow, and deep 
 
52 THE METALLURGY OP GOLD. 
 
 currents are best suited to effect speedy and thorough disintegra- 
 tion of the gravel, while a low grade, and slow and shallow currents 
 are best adapted for saving the gold, the upper part of a sluice, 
 for a sufficient distance to effect the complete disintegration of" 
 the gravel, is sometimes made of higher grades, or with narrower 
 boxes than the lower part, which is occupied solely in catching 
 the gold. When this is done additional supplies of water should 
 be introduced at the point where the change is made, otherwise, 
 the duty of the water being reduced, the sand packs in the angle 
 where the grade is altered, and constant attention is required to- 
 prevent the stream from overflowing. 
 
 The opposite requirements of disintegration and gold-saving 
 are more often supplied by "drops" and "undercurrents." A 
 vertical fall of the pulp constitutes a drop, which is arranged as- 
 follows: The sluice terminates in a "grizzly," or inclined grating; 
 made of parallel iron bars placed longitudinally to the stream 
 and from 1 to 6 inches or more apart, according to the exigencies 
 of the case. All the water and fine stuff pass through the 
 grizzly and fall a distance of from 1 to 10 feet into a sluice 
 below. The larger stones or boulders roll down the inclined 
 bars, and are shot over a precipice (if possible) or on to a steep 
 slope outside the sluice, as, unless some arrangement for remov- 
 ing these rocks is made, they will accumulate until they can no 
 longer roll off the grizzly. The higher the fall, the more 
 effectively it acts in causing disintegration. Sometimes, near 
 the head of a sluice, the grizzly is omitted from a fall, and the 
 boulders are retained to help in breaking up the gravel. The 
 chief disadvantage in permitting them to remain with the rest of 
 the gravel lies in the fact that they wear out the sluice, and that 
 much water is required to wash them down. 
 
 The undercurrent is often used in conjunction with a drop. A 
 grizzly with bars placed close together allows most of the water 
 and fine material to pass through, while the coarse stuff is carried 
 over and falls into the main sluice below. The fine material is 
 carried off by a short sluice placed at right angles to the general 
 direction of the main sluice, and is discharged into the upper end 
 of a large broad box from three to ten times as wide as the 
 sluice, and with its long diameter parallel to the main sluice. 
 A number of check-boards help to distribute the stream evenly 
 over the whole width of the box. This box, to which the name 
 undercurrent is often given, although it properly belongs to the 
 whole arrangement, is usually of lower grade than the sluice, 
 and is in that case supplied with additional water. A broad, 
 shallow stream is thus made to flow with reduced velocity over 
 the surface of the box, and, as the latter is plentifully supplied 
 with riffles and mercury, considerable quantities of fine and 
 " rusty " gold and fine particles of amalgam, which would otherwise 
 be swept away and lost, are retained by it. The tailings from the- 
 
SHALLOW PLACER DEPOSITS. 53 
 
 undercurrent are discharged into the main sluice below the 
 -drop. 
 
 Both grizzlies and undercurrents are used more frequently in 
 hydraulicking than in shallow placer sluicing, in which the large 
 stones are usually removed by a man with a blunt pronged fork, 
 who also either breaks up the lumps of clay or removes them 
 .and puddles them in tubs. 
 
 The Use of Mercury in Sluicing. Mercury is added at the 
 head of the sluice after washing has been in progress for a suffi- 
 ciently long time for all leakages to have been stopped, and for 
 the lowest depressions to have been filled in with sand. The 
 mercury is broken up into small particles either by pressing it 
 through chamois leather above the sluice, or by letting a thin 
 stream fall on to a man's hand, by which it is scattered, or by 
 some similar method. The amount added varies with the rich- 
 ness of the dirt and the magnitude of the operations, enough 
 being added to dissolve the amalgam formed. It is carried down 
 the sluice and lodges in the riffles, the greater part being retained 
 in the first few boxes. Fresh supplies are introduced every few 
 hours at the head of the sluice, and sometimes at various points 
 lower down the sluice also ; in particular, mercury is added to 
 the undercurrents, as it is especially valuable in catching the 
 finer particles of gold which would otherwise be lost, whilst coarse 
 gold can in great part be saved without mercury. Sometimes 
 the latter is forced into the substance of the wooden riffles by 
 driving an iron gas pipe into the wood, and filling it up with 
 mercury, which is forced by the pressure of the column through 
 the pores of the wood. The amalgam then forms on the surface 
 and in the interstices of the blocks, and in cleaning-up this is 
 scraped off. A better plan is to use amalgamated copper plates, 
 which is now often done. These resemble the plates used in 
 stamp batteries, described on p. 124. 
 
 The Clean-up. The length of each " run," at the end of which 
 the boxes are cleaned-up, varies, according to the richness of the 
 gravel, from a day to a whole season, and is usually a week. 
 The upper part of the sluice, which retains most of the gold, is 
 usually cleaned-up more frequently than the remainder. A 
 clean-up is begun by discontinuing the supply of gravel, and 
 letting the water continue to flow until it passes through the 
 sluice quite clear. The first six or eight sets of riffle bars are 
 then taken up, and the sand, mercury and amalgam washed 
 down, all the latter being caught by the first riffle left in. It is 
 scooped out thence by a wooden ladle or iron spoon into a 
 bucket, and the rich sand is collected and panned. The next 
 few riffle bars are now taken up, and so on, or alternatively 
 the work may be begun on several sections at the same time. 
 Lastly, the whole sluice is carefully searched over, and particles 
 of amalgam or mercury picked out of every crevice where they 
 -have lodged by spoons, penknives, &c. 
 
54 THE METALLURGY OF GOLD. 
 
 The mercury thus collected is quite liquid, and is strained 
 through chamois leather to separate the solid amalgam, which 
 is then retorted. The retorts used in well conducted enterprises 
 are similar to those in use in stamp mills, described on pp. 140-142.. 
 Amalgam obtained as the result of operations on a small scale, 
 however, is often merely heated on a shovel over an ordinary 
 fire, the mercury being driven off and lost. 
 
 Care must, of course, be taken to prevent partial clean-ups by 
 unauthorised persons. A man with a shot-gun stationed by the- 
 sluice is a frequent preventive measure. 
 
 Tail Race. The tailings from sluicing operations on low 
 ground which has not much fall are removed through a covered-in 
 wooden sluice, or, better still, through a large iron pipe. The work 
 proceeds in the up-stream direction, and the worthless material 
 stripped from above the pay-gravel is thrown on the top of the- 
 tail race, which thus passes through a mound of earth and dis- 
 charges into the open air lower down the valley. As the digging 
 and sluicing progresses up-stream, the tail race is lengthened and 
 the sluice boxes proper are conveyed further up the valley so as 
 always to be near the auriferous material last uncovered. This- 
 method originated with the Chinese. 
 
 Ground Sluice. In some cases boarded sluice boxes are not 
 used, but a stream of water is conducted to a little trench cut in 
 the pay-dirt, which is soon enlarged by the action of the water, 
 while the banks are at the same time shovelled or prised by the pick 
 or crowbar into the sluice. The gold is caught in the natural 
 riffles afforded by the uneven wearing of the bed, or rocks may 
 be added to arrest the gold, no mercury being used. Ground 
 sluicing is only adopted where the water supply is precarious, or 
 the season very short, so that violent rains cause floods that 
 would sweep away sluice boxes, and then are succeeded by dry 
 intervals during which the boxes would warp and crack. Only 
 the coarse gold is saved, while the duty of the water is usually 
 much less than in wooden sluices. After a time, usually when 
 the water gives out, the auriferous material is collected from the 
 sluice and washed in a long- torn or cradle. 
 
 Booming. This method of sluicing, which originated in the 
 United States, is adopted when the water supply is insufficient 
 for continuous operations. A dam with a light gate, capable of 
 being easily lifted, is built just above the part of the valley 
 where the auriferous gravel is situated. The water trickling 
 down the valley accumulates behind the dam, and finally over- 
 flows at one point into a small rectangular box fastened to the 
 end of a long lever. When full of water this box depresses its 
 end of the lever and raises the dam-gate, so that all the 
 accumulated water rushes out at once and scours the valley 
 bottom. As the box falls it empties itself of water, and the- 
 dam-gate returns to. its original position by its own weight. 
 
SHALLOW PLACER DEPOSITS. 55 
 
 This device is usually employed in connection with ground- 
 sluicing, but a line of sluice-boxes might be used through which 
 the sudden flood could carry gravel piled just above the head of 
 the series. 
 
 Dry Diggings. If no water can be obtained, it is sometimes, 
 profitable to concentrate very rich pay-dirt by winnowing, 
 tossing it in a blanket until the lighter particles have been 
 blown away, and finishing on a batea with or without mouth- 
 blowing. It is of course a wasteful method of concentration. 
 Several machines have been invented for dry concentration, 
 some of which have attained a fair measure of success. These 
 machines all use a blast of air by w T hich the sand is kept partly 
 in suspension, while it is moved by gravity down an inclined 
 table which is furnished with riffles. The auriferous material must 
 be quite dry or perfect disintegration cannot be accomplished. 
 In a typical dry washer, the gravel is first made to pass through 
 shaking screens, the mesh of which is adapted to the character 
 of the material. The object of the screen is to eliminate the 
 larger fragments, which are usually barren. The shaking screen 
 delivers the material on to an inclined table formed of a wire- 
 screen, covered with light canvas or some similar material 
 through which are forced pulsating blasts of air. These sudden 
 puffs throw up the sand and let it settle again alternately, and 
 as a result the light material works down the table, while the 
 gold is retained by the riffles, being too heavy to be tossed over 
 them by the air. 
 
 Cement Gravels. When gravels are cemented by iron oxides 
 or clay so as to be too hard for disintegration in the sluice, 
 except with the aid of a great number of drops, they are either 
 sent to the stamp-mill or, with more advantage, treated by a 
 cement-mill before being washed in the sluice. The treatment 
 of cemented gravel is best preceded by exposure to the air 
 for a time, frost, sun and rain having a very rapid effect on it. 
 The cement-mill is a large iron pan fitted with coarse screens ; 
 it has revolving arms carrying plough-shares, which puddle 
 the clay and break up lumps of gravel, but leave boulders 
 untouched. Water, fed in with the dirt, carries the fine stuff 
 through the screens, and at intervals the boulders are removed 
 by various contrivances. 
 
 Tail Sluices are sometimes erected to intercept the tailings 
 from one or several sluices with the object of collecting a further 
 percentage of gold from the waste material. These tail sluices 
 are made of much greater size than those described above, and 
 in some cases pay for construction. The Kumara sludge channel, 
 erected by the New Zealand Government to carry the tailings 
 from the sluicing works of the district into the river, caught 
 957 ozs. of gold in the four years ending in 1891. This sluice is 
 3 feet 6 inches wide, and has a grade of 1 in 28, while the boxes 
 
56 THE METALLURGY OP GOLD. 
 
 which discharge into it are only 18 inches to 22 inches wide. 
 Usually, in good work, the sluices are long enough to make the 
 tailings too poor to be worked over again at a profit, except by 
 the Chinese, until after they have been enriched by natural 
 concentration in the rivers. 
 
 My Catchers were invented in Australia for the purpose of 
 catching the fine particles of gold, which, successfully evading 
 the riffles of all sluices, float down on the surface of the rivers. 
 These devices consist of weirs constructed on piles driven into 
 the river bed, and stretching across from bank to bank of the 
 river. Boards covered with blanketing or coarse gunny -sacking 
 are attached to the weirs and collect all particles floating on the 
 surface of the water. At intervals the blankets are taken up 
 and washed in a tank. These fly catchers soon pay for their 
 cost of construction on many rivers, but are liable to be 
 damaged by floods, and by being used as bridges by men and 
 animals. 
 
 River Mining. River mining consists in the working of 
 auriferous gravels in the channels and beds of existing rivers, 
 and may with convenience be made to include the exploitation 
 of deep bars below the level of the water. It is conducted in 
 three different ways : 
 
 1. A portion of the river bed is laid dry by damming and 
 fluming. 
 
 2. The gravel is raised by dredges or similar contrivances, 
 operated from a boat. 
 
 3. Pits are sunk on the bank, and the gravels below water 
 level excavated and raised to the surface by ordinary mining 
 methods, or by the hydraulic elevator. The gravel won by 
 either of these three methods is washed in the usual way by 
 sluicing. 
 
 1. River Mining Proper. This method is carried on chiefly 
 in California. In the case of the larger rivers, or where only a 
 small capital is available (from 100 to 1,000), wing-dams are 
 built out from the bank, above and below the part of the river 
 which it is desired to work, and a third dam, built parallel to 
 the direction of the current, connects their mid-stream ends. 
 The dams are usually constructed of boulders, with their inter- 
 stices filled with gravel, and are only slightly higher than the 
 river surface at low-water (summer) level. The water-flow is 
 forced to the other side of the river channel, its level being little 
 altered. The space thus cut off is pumped dry by a Chinese 
 pump, worked by an undershot wheel placed in the current by 
 the side of the mid-stream dam. The pay-dirt is usually covered 
 by a mass of barren boulders and gravel, and is exposed by 
 stripping or by sinking shafts and running drifts. The dirt is 
 taken out down to the level of the bed-rock, which is cleaned 
 thoroughly, all crevices being well searched. The pay-dirt is 
 
SHALLOW PLACER DEPOSITS. 57 
 
 washed by sluicing, the water being supplied directly from the 
 river, or, if necessary, raised to the required height by a dip- 
 wheel (an undershot water wheel with small buckets placed near 
 its periphery, which empty their contents into a flume). The 
 boulders are removed by a derrick worked by water-power. As 
 the construction of the dam can only be begun when the water 
 is low, the season is usually very short ; sometimes only a few 
 days or hours are left, after the draining and stripping have been 
 finished, for the actual collection and washing of the pay-dirt. 
 Operations are stopped by the autumnal rise of the river, which 
 overtops the dam and fills the pit, often coming so suddenly as 
 to leave no time to remove any of the tools and machinery. 
 Sometimes large returns of from 100 to ,1,000 per day are 
 obtained in this short time, and these may be sufficient to yield 
 handsome profits on the undertaking, but the dams are always 
 swept away in winter. 
 
 ^Vhen enough capital is available, or the river to be operated 
 on is small, head- and foot-dams are made, stretching across from 
 bank to bank, and the water is carried off in a wooden flume and 
 delivered into the river bed below the foot-dam. These dams 
 are usually built of timber, faced with planks, and supported by 
 earth and stones. The portion of the bed thus isolated is drained 
 and worked as in wing-dam practice, the source of power being 
 the water-race in the flume. The work is also in this case often 
 cut short by the rising waters, which carry away the dams, 
 flume, sluices, wheels, &c. 
 
 Of late years efforts have been made in two or three places in 
 California to prevent the winter floods from stopping work. On 
 the Feather River a permanent head-dam and flume has been 
 constructed by an English company to drain a part of the bed, 
 and tunnels have been made on the American and Feather 
 Rivers to permanently drain large reaches and deliver the water 
 at a lower point. None of these undertakings have as yet been 
 strikingly successful, from various causes. 
 
 River mining is probably subject to more uncertainty than any 
 other branch of gold mining. The whole capital invested may be 
 lost, and all works and machinery swept away by a flood before 
 the pay-dirt is sighted, while numerous instances are on record in 
 which the alluvium on the river-bed, after having been laid bare 
 at great expense, was not rich enough to pay for sluicing. 
 
 Between the years 1857 and 1880 the Californian river-beds 
 were covered so deep by the tailings from hydraulicking that 
 they could not be worked with advantage. Since the suspension 
 of hydraulicking, however, and the gradual working down of the 
 debris, some places have again become worthy of attention.* 
 
 * For a full account of river mining, with details of dam-construction, 
 Ac. , the student is referred to the article 011 the subject by R. L. Dunn in 
 the Ninth Annual Report of the Californian State Mineralogist, 1889, pp. 
 262-281. See also the Eleventh Report, 1892, pp. 150-153. 
 
58 THE METALLURGY OP GOLD. 
 
 2. Dredging. This method of winning the gravel has seldom? 
 met with any success, for the reasons that the gold, even when it 
 exists in the river bed, cannot be got at in such a simple way. 
 The bed-rock cannot by this means be cleaned and creviced, and 
 if the poor gravels on the surface are thick, they cave in and 
 slide down into the pit made by the dredge, so that the pay-dirt 
 is never reached. Moreover, boulders impede or prevent the 
 work. The favourite implements are various forms of vacuum 
 lifts and pumps, the power being sometimes supplied by a hy- 
 draulic elevator. Water, gravel-and stones are brought into the 
 barge together, and the stuff is usually then and there washed and 
 dumped into the river ' again. The system ' has made somewhat 
 more progress in New Zealand than elsewhere, and the following 
 details of recent work .there are taken from a Government report 
 on mines."* On the Molyneux River in New Zealand the centre 
 bucket dredger has done good service, the difficulty of raising 
 big stones being that most severely felt. Here it is found that 
 little of the fine scaly gold of the rivers is caught in ordinary 
 sluices, and the washing is done by separating all the coarser 
 material by means of trommels, and passing the' fine sand by 
 itself over wide tables in a very shallow stream. The volume of 
 water must be just enough to keep the tables free from sand,' as if 
 the latter begins to collect, it is believed that the fine gold is not 
 being caught. The tables are covered with cocoa-nut matting, 
 baize, blanketing, or even plush, each material having its advo- 
 cates. On the Shotover River, the Sand Hills Dredging Company 
 separates the stones by passing the gravel through a revolving 
 cylinder 3 feet in diameter and 10 feet long, set with fine holes ;. 
 30 tons per hour are treated by this trommel. The washing is 
 done on inclined tables which are 64 inches wide, and about 
 27 feet long, having a grade of 18 inches in 12 feet. These 
 tables have a superficial area of 144 square feet only, which is 
 insufficient for the treatment of 30 tons per hour. At the 
 Waipapa Creek Dredging Company's works, which are situated 
 on the sea coast, a Welman dredge is used, which acts on the 
 centrifugal suction system. The pump is 3 feet 6 inches in 
 diameter, and the delivery tube 13 inches in diameter, and the 
 pump raises gravel from an area having a radius of 40 feet. All 
 large stones are caught and separated from the fine stuff by a 
 riddled hopper-plate. The gold is very finely divided, and is 
 caught on plush mats which are washed every eight hours. The 
 water for washing is supplied from a reservoir by means of an 
 18-inch pipe. Stones of 56 Ibs. weight are lifted by this pump, 
 which may be used with advantage on all low wet ground. 
 
 In Northern Italy a dredging plant has also been successfully 
 at work for some years.f 
 
 3. Deep Bar Mining. Deep bar mining, as conducted by 
 
 * New Zealand Mng. Comm. Report, 1891. 
 
 f In several places in the United States bucket-elevator dredges and single 
 shovels are in operation. Eng. and Mng. Jour. vol. Ixiii. (1897), pp. 211 
 and 259. 
 
SHALLOW PLACER DEPOSITS. 59) 
 
 the old methods, consisted in sinking shafts in the bank and 
 running drifts below water level, either under the river bed 
 itself or under deep bars, for the purpose of digging out the pay- 
 dirt from the surface of the bed-rock and taking it to the surface 
 t'o wash. It was mostly unsuccessful owing to the amount of 
 water, quicksands, &c., encountered, although to-day these diffi- 
 culties could be often overcome by the present improved practice 
 in quicksand digging. As, however, this kind of digging is 
 better suited to the efforts of a few individuals than to under- 
 takings on a large scale, it is not likely to be largely reverted to. 
 In particular, working in loose detrital matter under the river 
 bed itself must always be hazardous to 'the workers, and the 
 chances of success very dubious, although profits have been, 
 obtained where the pay-dirt was only worth 70 cents per cubia 
 yard. Deep bars, however, are now being worked profitably 
 by the hydraulic elevator, which has become of great import- 
 ance in placer mining. This machine may be more conveniently 
 described after hydraulicking has been considered. 
 
 Method of Working Siberian Placers. The methods and 
 apparatus employed in Siberia differ so markedly from those 
 which have been adopted elsewhere, that they are well worth 
 a special description, although they cannot usually be applied to 
 placers found in other parts of the world, owing to the difference 
 in economic conditions. In California the valleys are narrow 
 and the grade steep, so that watercourses are usually close to- 
 the auriferous deposits, and the sluices can be made of almost 
 any length, while still conforming to the general slope of the soil. 
 In Siberia the slope of the valleys is so gradual that the flow of 
 the water is almost imperceptible, and takes place through wide 
 marshy tracts. The result of this is that the sluices must be 
 short, being usually less than 100 feet long, and their upper ends 
 must be raised on trestles. As a further consequence, also, the 
 gravel must be excavated by hand and carried in waggons to the 
 sluice, and the tailings removed in a similar way, the flow of 
 water acting by gravity not being available for these purposes. 
 The excavation is made in benches or terraces, working up the 
 valley. The height of each bench above the lower one is about 
 5 feet, and the gravel is picked down from the face of each bank 
 and shovelled into carts by which it is conveyed to the washing 
 establishments. The additional expense entailed by these causes 
 is balanced by the low cost of labour in the country, and an inci- 
 dental advantage lies in the fact that the washing apparatus can 
 be placed outside the limits of the river during flood time, and 
 so may remain for a number of years undisturbed, while all the 
 gravel in the district is being washed. Only surface deposits 
 are exploited, no deep workings existing in the country. 
 
 There are three types of apparatus employed, each designed 
 for washing a particular kind of deposit. They are : 
 
60 THE METALLURGY OF GOLD. 
 
 1. The Siberian sluice, which is used to wash light sand. 
 
 2. The Trommel, used for loamy sands. 
 
 3. The Pan, used for gravel which is cemented together by 
 means of compact clay. 
 
 1. The Siberian Sluice. The apparatus at Voltchanka, which 
 may be taken as a type, consists of a head sluice and three 
 secondary sluices, which are placed at right angles to the head 
 sluice, and which leave it at different points and converge to a 
 common centre, where the tailings are discharged. The head sluice 
 begins at a height of 13 feet from the ground ; it is about 90 feet 
 long by 2 feet wide, and has a fall of about 1 in 14. The sands 
 are dumped from the waggons on to a wooden platform situated 
 above the sluice-head, and shovelled into the latter, a stream of 
 water being turned in at the same time. After passing through 
 a grizzly, the gravel runs over a series of cast-iron cross-bar 
 riffles, which form a number of rectangular depressions (or pigeon- 
 holes) in the bed of the sluice, by which the disintegration is 
 favoured. The stream then flows over an iron screen, through 
 which a part of it falls into the first secondary sluice, while the 
 remainder continues its course over more pigeon-hole riffles. 
 This arrangement resembles the Californian undercurrent. A 
 second and a third screen open on to the other secondary sluices, 
 and the part of the gravel (consisting chiefly of small stones) 
 which has resisted disintegration, and has not passed through the 
 screens, is then let fall into a hopper, whence it is removed to 
 the tailings waggons. 
 
 The secondary sluices are wider than the head sluice, and have 
 & steeper grade, and the amount of water and auriferous 
 material passed is of course much less in each one than on the 
 principal sluice. The sands first pass over a number of trans- 
 verse riffles, and then over about 30 feet of blanketing, the sluice 
 being widened at the same time, and subdivided by longitudinal 
 wooden partitions, and the grade being as much as 1 in 6, while 
 ismall drops are introduced at intervals of a few feet. The third 
 sluice has a more gentle inclination than the others. The tail- 
 ings fall into a shallow sump, and are instantly lifted out of it 
 by a bucket elevator or " tailings-wheel " operated by water 
 power, and stored in a hopper, whence they fall into waggons, 
 by which they are removed to the dumping ground. 
 
 No mercury is used in these sluices, and only the production 
 of " grey concentrates " is attempted, this work being continued 
 from 6 a.m. to 7.30 p.m. every day, after which the concentrates 
 .are collected from the riffles. They consist of gold in scales and 
 plates, magnetic iron oxide, pyrites, rutile, together with some 
 quartz, &c. They are treated with mercury in the Siberian 
 trough or on inclined tables, the method being that described 
 on p. 48. 
 
 The apparatus described above treats 500 tons of gravel per 
 
SHALLOW PLACER DEPOSITS. 61 
 
 day, the labour required being furnished by twenty men and ten 
 horses. The gravel treated contains an average of from 12 to 15^ 
 grains of gold per ton, rarely falling below 6 grains per ton ; the 
 exceptional richness of 3} dwts. per ton has been observed. The 
 gold is chiefly found in the head sluice, where 70 per cent, is 
 caught, 30 per cent, being caught on the secondary sluices. At 
 a similar establishment at Tchernaia-Retchka, however, where 
 the gold is less finely divided, 97 per cent, was caught on the 
 head sluice, and only 3 per cent, on the secondary sluices. The 
 amount of water used at Voltchanka is about six times the 
 weight of the gravel. The cost of construction of the works 
 was 70,000 roubles, or about 7,000. 
 
 2. The Trommel. Gravels which are too compact for satisfac- 
 tory disintegration in the short sluices described above are sub- 
 jected to a preliminary treatment by a trommel. At Berezovsk 
 the trommel is of sheet iron of 9 mm. thick, having holes in it of 
 about 1 mm. in diameter. The trommel is about 12 feet long,. 
 3^ feet in diameter at one end, and 4i feet at the other, and is 
 set inside with denticulated plates of iron to assist in the disin- 
 tegration effected by the water. The machine is driven by a 
 water wheel, and is sufficient for the disintegration of from 400 
 to 500 tons of gravel per day, requiring the expenditure of about 
 3 horse-power to drive it. The amount of water used in the 
 trommel and on the tables is 67*5 litres per second, or about 
 seven and a-half times the weight of the ore. The washing is 
 effected on inclined tables only 30 feet long and 12 feet wide, 
 and with a grade of about 1 in 4, placed with the incline at right 
 angles to the length of the trommel. Near the head of the 
 table, and stretching across it, is a deep trough-like depression, 
 and below this there are a number of transverse riffles, in which 
 grey concentrates are caught and treated as usual. The Bere- 
 zovsk establishment employs twenty-five men and fourteen horses 
 constantly ; the trommel usually lasts for two seasons. 
 
 3. Pan Washings. Sandy clays cannot be economically 
 disintegrated in a trommel, and are, therefore, treated in a 
 washing pan, which bears a strong resemblance to the cement 
 pans employed in California. The pan usually consists of cast 
 iron, and is from 8 to 16 feet in diameter, with vertical sides 
 from 1 to 5 feet high. The bottom is of cast iron or sheet iron, 
 and has numerous holes in it of about T ^ inch in diameter, 
 widening downwards. The bottom is divided into 25 sectors, 
 between which are deep grooves for the collection of the pebbles. 
 Through a circular opening in the centre of the pan there 
 passes a revolving axis to which are suspended eight horizontal 
 arms studded with vertical iron teeth, some of these being shaped 
 like plough-shares. The revolution of these arms effects the 
 disintegration of the sandy clays, which are fed into the pan 
 together with water and puddled until fine enough to pass- 
 
^62 THE METALLURGY OF GOLD. 
 
 through the holes in the bottom, and the stones are removed 
 at intervals by opening little gates placed opposite the radial 
 grooves. The disintegrated gravel falls from the pan on to 
 concentration tables, similar to those used after disintegration 
 in the trommel. At Berezovsk the pan is 11 J feet in diameter 
 and 5 feet deep, and the arms revolve at the rate of 25 turns per 
 minute. From 50 to 55 tons of material are treated in twelve 
 hours, the water consumed, including that required for power, 
 being ten times the volume of the sand. 
 
 Beach Mining. Beach mining is a comparatively unimport- 
 ant form of shallow placer mining. The sea beaches on parts 
 of the coasts of California, Australia and New Zealand contain 
 small quantities of gold, which has been proved, in all cases in 
 which the matter has been investigated, to be derived from the 
 cliffs, which mostly contain a still smaller quantity. Some 
 streaks of black sand, however, in the " Gold Bluff," California, 
 have yielded $135 or 6| ozs. per ton by actual working.* The 
 waves of the sea wash down and partially concentrate the poor 
 sands, and, under certain rather exceptional circumstances, as 
 the tide goes out the surface of the beach is left covered with 
 black sand, in which numerous specks of gold occur. This is 
 carefully scraped up and transported inland to be washed, as sea 
 water is not well adapted for the purpose, although it is used by 
 one Californian company. The next tide usually washes away all 
 the valuable material which has not been collected, or else covers 
 it with barren sand. There is great difficulty in washing the 
 black sand in California, as it consists largely of rounded grains 
 ^of magnetite, the density of which is about 5-0, while the gold 
 is in minute flakes and scales, which can be seen under the 
 microscope to be oblong in shape, and thicker at the sides than 
 in the middle (a shape due to continued pounding of a malleable 
 material). This form is so easily moved and buoyed up by water 
 that it is difficult to get a "colour" with the pan, and the 
 amount caught by the mercury in sluices or long-toms is 
 usually an insignificant proportion of the total assay value of 
 the sand. The industry is generally a languishing one. 
 
 Treatment of Shallow Placer Gravels by Steam Shovels 
 and Amalgamated Plates. A plant for the treatment of placer 
 deposits, which are similar in nature to those worked in Siberia, 
 has been devised in America, where labour-saving contrivances are 
 indispensable in all such cases, owing to the high rate of wages. 
 The apparatus consists essentially of a combination of a steam 
 shovel or navvy, and an amalgamator consisting of a wide waggon- 
 shaped wrought-iron trough loosely lined with silver-plated amal- 
 gamated copper plates, which form a series of steps or riffles at 
 the sides of the trough. The material is elevated into a hopper 
 >by the excavator, and thence is charged into a revolving trommel 
 * Prod. Free. Met., U.S.A., 1884, p. 557. 
 
DEEP PLACER DEPOSITS. .. 63 
 
 placed inside the trough. Here the disintegration of the gravel 
 is effected, and the fine material falls through into the trough, 
 while the stones are discharged outside at the end of the 
 trommel. At the bottom of the trough there is a water pipe, 
 carrying water at high pressure, and in that pipe a series ot jets 
 pointing alternately forwards and backwards. The result is to 
 give a series of eddies or whirlpools in the water in the trough, 
 and the sand and fine gold is continually carried up to the 
 surface near the middle line of the trough, and in descending 
 again near the sides it comes in successive contact with several 
 of the plates, which form a series of steps. The fine sand is 
 eventually discharged at the end of the trough. It is stated by 
 the Bucyrus Steam Shovel and Dredge Company, by which the 
 machinery is manufactured, that such a machine erected in 
 Montana has a capacity of from 600 to 800 cubic yards of gravel 
 per day, all the water required being supplied by a pump raising 
 500 gallons per minute, the proportion being only three of water 
 to one of gravel, which seems much too small. It is stated that 
 gravel containing only 12 cents of gold (i.e., about 3 grains) to 
 the ton has been treated successfully by this machine, but exact 
 records of continuous work done by it are wanting. It will be 
 noted that special means must be adopted in each case for the 
 handling of the tailings. 
 
 CHAPTER V. 
 DEEP PLACER DEPOSITS. 
 
 Nature and Mode of Origin of Deposits. This discussion is 
 necessary in order that the description of the methods of treating 
 the deep placer gravels may be intelligible. Both, in Australia 
 and California, besides the superficial placer deposits situated in 
 or near the existing rivers, which in the deep canons of the 
 Klamath and other rivers in the extreme north of California 
 attain a thickness of 250 feet, there exist auriferous gravels 
 which bear no apparent relation to the present drainage of .the 
 country. These gravels often attain enormous thicknesses,,. and 
 are in many places covered by volcanic rocks, consisting of 
 basaltic lavas and tuffs, which are sometimes; interbedded with 
 gravel and loam This latter circumstance shows that inter,- 
 mittent action of the volcanic vents, with long intervals of 
 repose, has taken place. There has been some difficulty in 
 accounting for the origin of these deep placers, and it has been 
 ascribed in succession to the agency of the sea, of ice, and (for 
 California) of a huge river flowing from north to south at right 
 angles to the direction of flow of the existing rivers. Kone of 
 these views are now entertained, and the " fluviatile " theory is 
 
64 THE METALLURGY OF GOLD. 
 
 generally accepted, the origin of the gravel being ascribed to the 
 depositions of ancient rivers flowing in courses roughly parallel 
 to those of existing rivers. The geological age of these ancient 
 rivers has not yet been determined with certainty, but though 
 they may be Pleistocene, the balance of palaeontological evidence 
 is perhaps in favour of Whitney's view that the deposits were 
 formed in the Pliocene period. 
 
 The ancient Californian rivers probably had their sources at 
 somewhat higher altitudes than those now existing, and had 
 more uniform general grades, the slope of their beds corre- 
 sponding more nearly to the general slope of the country. The 
 existing rivers, on the other hand, have steep grades in the upper 
 parts of their courses, followed by comparatively level stretches 
 below. The old rivers, however, like their successors, had rapids, 
 falls and level stretches, the grade varying from 5 feet to- 
 250 feet or more per mile. The Pliocene rivers ran in valleys 
 which were broad and shallow in comparison with the present 
 deep precipitous canons, and the volume of water was in general 
 much greater than that delivered by their representatives of 
 to-day. The width of the valleys varied from 100 feet to fully 
 1J miles (which is the width at Columbia Hill), and the depth 
 must have been often over 1,000 feet. These valleys were 
 already partly filled up by accumulations of gravel, when the 
 outbreak of volcanic activity in many cases filled up the remain- 
 der, and the streams were deflected into other channels, which 
 often lie close alongside the old canons. These new channels 
 have been excavated by the running water until they now lie 
 much below the level of the beds of the Pliocene rivers, and 
 consequently the gravels which were deposited in the old valleys- 
 now sometimes crown the highest ground in the district, the 
 general level of the country having been greatly reduced in 
 height. The new channels have been cut partly in the old 
 country rock and partly in the Pliocene auriferous gravels and 
 their covering of volcanic rocks. Sometimes the course of the 
 present canons cuts that of the 
 old at several points owing to 
 the sinuosity of both, see Fig. 
 8, in which A represents the 
 modern river, and B the 
 ancient one. The result is 
 that sections of the old valley 
 from bed-rock to surface are exposed in the sides of the 
 caftons, usually at some height above the present level of the 
 water, and it was at such points as these that the discovery 
 of the existence of the deep placers was first made. The hard 
 covering of basalt has served to protect the more friable gravels, 
 which have been for the most part removed in those places 
 where the lava has been worn away or has never existed, so that 
 
DEEP PLACER DEPOSITS. 
 
 65 
 
 the largest tracts of gravels still existent lie beneath the volcanic 
 rocks. 
 
 Fig. 9 represents a section across two ancient channels (B, B) 
 and a modern canon, that of the American river. Here, A is the 
 volcanic capping, which is 800 feet thick above the Red Point 
 channel; B, B are the auriferous gravel channels; C, C are 
 deposits of gravel on the " rims," containing gold in places ; D is 
 the bed-rock, consisting of dark-blue slates ; E is a barren deposit 
 of angular debris and boulders ; F F are prospecting tunnels, 
 which were put in at too high an altitude ; F' is the tunnel 
 bored with the object of reach mg the bottom of the gravel de- 
 posit ; H are prospecting winzes sunk in order to discover the 
 position of the gravel. The space included within the dotted 
 lines N M Y M' N' has been obviously denuded since the deposi- 
 tion of the volcanic cappings, the soft slate rims N M R and N' M' 
 
 Fig. 9. 
 
 having been worn away, while the hard lava has resisted erosion. 
 The vertical depth from M to the American river is about 1,800 
 or 2,000 feet. 
 
 This condition of things is that prevailing in California, but in 
 Victoria the structure closely resembles that just described, with 
 the exceptions that the old valleys were smaller and that the 
 erosive action of the rivers since the deposition of the basalt has 
 been comparatively slight, owing to the slight grades of the 
 streams caused by the low elevation of the country and to the 
 small amount of the rainfall. In consequence of this the basalt 
 has usually not been worn through, and the " deep leads " or old 
 river bottoms are often below the level of the present streams, 
 so that although a larger proportion of the Pliocene gravel 
 remains, it is more difficult and expensive to mine. 
 
 The shallow placers, at any rate in California, have resulted in 
 
66 THE METALLURGY OF GOLD. 
 
 the main from the erosion of these deep placers, the materials of 
 which, having undergone a natural concentration in the ground 
 sluices afforded by the river beds, furnished the wonderfully 
 rich river-bed and bar deposits, which yielded so much gold 
 between 1848 and 1860. The deep level gravels vary greatly in 
 thickness, as has already been stated, being only 2 feet thick at 
 Table Mountain, Tuolumne County, and over 600 feet thick at 
 Columbia Hill, Nevada County, California, and averaging from 
 100 to 300 feet, the thickness varying with the nature of the 
 river bed, and the subsequent erosion. The gravel consists in slaty 
 districts chiefly of quartzose sand, the fine materials furnished 
 by the disintegration of the slate having been for the most part 
 swept away, and the products of the quartz veins contained in 
 the slate being left. Near bed-rock, but at no higher level, 
 there is often a collection of large boulders, varying in size up 
 to 10 feet in diameter, consisting mainly of quartz, but sub- 
 ordinate to these are others usually similar in character to the 
 bed-rock. These boulders, though rounded, are too large to have 
 been transported far by running water, and have probably been 
 polished by the attrition of the sand carried over them. At the 
 Forest Hill Divide, Placer County, California, the gravels consist 
 almost entirely of pure white quartz, but at other places such 
 quartz is rare. It is to be noted that it is only in slaty districts, 
 where the gravels are mainly quartzose, that rich auriferous de- 
 posits occur. In granite districts, where the gravels are composed 
 of more heterogeneous materials, and in cases where they consist 
 of volcanic boulders and detritus, little or no gold is found. The 
 lower parts of the gravels are often cemented into a conglomerate, 
 called " cement," by infiltration of silica, oxides or sulphides of 
 iron, or, rarely, carbonate of lime ; when the gravels are covered 
 with lava, the whole thickness is in some cases converted into 
 cement. The upper parts of the gravels often contain pipe-clay 
 in greater or less quantity, either in pure beds or mixed with 
 sand. Fossil leaves occur in the clays, and drift wood occurs 
 throughout the whole of the deposits in extraordinary abund- 
 ance, particularly in Australia ; this wood is for the most part 
 silicified or replaced by sulphide of iron. The higher portions of 
 the gravels are often altered by the action of air and water on 
 the iron sulphide, thus forming ferrous salts, haematite and 
 hydrous sesquioxides, which colour the gravels red and brown 
 respectively. The upper gravels are hence called " red gravels." 
 The lowest layers, being protected from alteration from above, 
 are coloured dark blue -grey by the ferrous sulphide contained in 
 them, and are hence called "blue gravels," their occurrence 
 giving origin to the old "blue lead" theory, owing to their 
 uniformity of colour over wide areas in California. The sul- 
 phide of iron which incrusts fossil bones and teeth found in the 
 gravel, and replaces the substance of drift wood, was formerly 
 
DEEP PLACER DEPOSITS. 67 
 
 believed to be derived from below, as the result of metamorphic 
 action going on in the country rock. 
 
 Distribution of Gold in the Gravels. The gold is found 
 chiefly either in contact with or just above bed-rock. If this con- 
 sists of soft slate, and especially if the planes of cleavage are at a 
 high angle to the horizon, particles of gold are often found in the 
 natural riffles thus formed, and are disseminated through the 
 rock to the depth of a foot or two. If depressions, pot-holes, or 
 fissures exist in the old river bottom, they are usually very 
 rich in gold. Where, as often happens, there is a channel, or 
 " gutter," to adopt the Australian expression, cut by the stream 
 in the lowest part of the valley, the gravel filling it is usually 
 much richer than that found elsewhere. Such rich portions, 
 often only a few feet wide, and of insignificant depth, but extend- 
 ing to considerable distances in the direction of the stream, are 
 called "leads." As a result of these circumstances, the "blue 
 gravels " happen to be richer in general than the " red gravels," 
 from which arose the old theory that only blue gravel pays to 
 work. Coarse gold and nuggets chiefly occur near bed-rock 
 in the deeper parts of the channel, but the " rim-rock " gravels 
 are also often rich. The " rim-rock " is that portion of the bed- 
 rock which forms the sides of the old valley, thus lying consider- 
 ably higher than the central channel. The richness of gravels here 
 doubtless arises from the existence of old bench or terrace gravels, 
 which are consequently the oldest of the whole series, being 
 formed before even the gutter gravels. Rich streaks also occur 
 at various levels in the gravels, often resting on "false bottoms," 
 which consist of impermeable beds of clay or some similar 
 material. Sometimes these streaks are richer than those encoun- 
 tered at bed-rock, as, for example, at the Paragon Mine, Placer 
 County, California. Although it is concentrated in this manner 
 at various points, gold nevertheless occurs disseminated through 
 the greater portion of the red gravel, where, however, it is in a 
 finer state of division and less abundant. Besides existing as 
 free particles, gold may occur in quartz boulders, although this 
 is rare. For instance at the Polar Star Mine, Dutch Flat, 
 Oalifornia, a white quartz boulder was found, which contained 
 288 ozs. of gold. Gold may also occur, together with pyrites, 
 replacing the substance of drift wood. 
 
 The amount and position of the gold varies, as in the case of 
 the present rivers, with the grade, the shape of the valley, the 
 volume of water, the amount of gravel being carried down, &c. 
 "An underloaded current i.e., a current charged with less 
 detritus than it is well able to carry is apt to cut its bed, and 
 prevent the accumulation of gravel. A greatly overloaded 
 current will deposit too rapidly to admit of the concentration of 
 the gold dust " * Under conditions intermediate between these 
 
 * Ross E. Browne, Tenth Report, of the Gal. State Mineralogist, p. 448. 
 
68 THE METALLURGY OF GOLD. 
 
 extreme states, the current may be just strong enough to keep its 
 bed clear from all accumulations except a small quantity of coarse 
 gravel and the coarse gold, which is caught in the natural riffles, 
 and thus all the conditions necessary to form a rich bed of pay-dirt 
 may be present. If, however, the bed consist of granite or other 
 rock which wears in smooth and rounded shapes, little gold will 
 be caught. Slates, consisting of layers of uneven hardness, wear 
 irregularly, and afford a good gold catching surface. The condi- 
 tions noted above as necessary to form rich gravels cannot be 
 expected to have been prevalent over great distances. " An 
 increase of grade or narrowing of the channel will cause an 
 increase of velocity, and the same stream may be underloaded 
 in a narrow steep section, and overloaded in a broad flat 
 section." * The difference of velocity between the middle and 
 sides of a stream, and between the inside and outside of a bend, 
 may give the right conditions in one part of a river bed and not 
 in another. Thus with high grades, rich gravels should occur 
 in the less rapid, and with low grades, in the more rapid parts 
 of a stream. Having regard to such considerations, the richer 
 parts of existing rivers can be pointed out with little trouble. 
 When, as in the Pliocene rivers, the beds are buried to a depth 
 of hundreds of feet, the richer parts are more difficult to find. 
 
 The history of the Pliocene rivers began with a period when ex- 
 cavation exceeded deposition, when the rivers were underloaded 
 for at least a portion of each year, and the channel was constantly 
 being deepened. Some bench or terrace gravels were formed at 
 this time, and being at the sides of an underloaded river tended 
 to be rich. The river bed, although rocky and comparatively free 
 from sand, would perhaps accumulate some coarse gold, which as 
 the channel deepened was no doubt in part ground up into fine 
 particles and carried off, but at the time when the excavation had 
 reached its lowest point, some of this coarse gold would certainly 
 be present. When the underloading of the stream ceased, what- 
 ever caused the cessation, a pause must in many cases have occurred 
 before the gravel proved too much for the stream to carry. During 
 this pause the conditions for gold catching were favourable, and 
 hence rich gravels were formed on bed-rock in the gutters or 
 channels. Then, as the streams became overloaded, sand and 
 gravel accumulated rapidly, so that little concentration of the 
 gold in them could take place. The rivers flowed over thick 
 sand banks and, in consequence, frequently changed their courses. 
 The sands, being deposited by overloaded rivers, of course con- 
 tained fewer and smaller boulders, and the thick masses of poor 
 sand thus went on accumulating until the volcanic outbursts 
 put an end to the process. 
 
 Origin of the Gold in the Placers. The origin of the 
 gold in deep placers has long been a vexed question. It was 
 
 * Ross E. Browne, Tenth Report of the Col. State Mineralogist, p. 448. 
 
DEEP PLACER DEPOSITS. 69 
 
 formerly accepted without question that the erosion of auriferous 
 quartz lodes existing at higher altitudes furnished both gravel 
 .and gold. In support of this it w<*s urged that the same districts 
 which furnished auriferous gravels abounded in quartz veins at 
 higher levels, while "Whitney pointed out that numerous lodes 
 were intersected by the valleys and were still to be seen in the 
 bed-rock. On the other hand, the fact that the nuggets found 
 in the drift are much larger than any masses of gold encoun- 
 tered in veins, and that the placer gold is of superior fineness, 
 .are difficulties in the way of accepting this theory. Moreover, 
 Egleston states that nuggets as large as a man's fist have been 
 found embedded in the midst of fine sand, whither they could 
 not have been carried by the action of running water, but authentic 
 instances of such finds seem to be lacking, nuggets usually occur- 
 ring in coarse gravel, among boulders. It is further declared by 
 the opponents of the erosion theory that if a small quantity of 
 soft material like gold, mixed with lumps of hard quartz, were 
 washed down by water, then, long before the quartz could be 
 reduced by grinding to the condition of grains of sand, the gold 
 would be worn down to such a fine state of division that none of 
 it could lodge in the river bed at all. In opposition to this con- 
 tention, it may be urged that the extreme malleability of fine 
 gold would make this comminution very slow, and scales of the 
 metal are said to have their edges blunted and thickened by the 
 pounding action of dry sand moved by the wind, instead of having 
 them worn away. However, it is certain that the form of placer 
 gold is different from what might have been expected if it con- 
 sisted of water-worn vein gold. Nuggets are mammillary in 
 form, and generally appear more like concretions than water- 
 worn fragments, in spite of the fact that they often include 
 more or less quartz. 
 
 In 1864, in order to account for these and other facts, Mr. A. 
 C. Selwyn, of Victoria, suggested a theory of solution in which it 
 is supposed that the gold disseminated through the rocks and 
 drifts is dissolved by percolating waters which contain acids and 
 salts in solution, and is reprecipitated around certain centres. 
 Selwyn considered that the waters capable of dissolving gold, must 
 have acquired this property by passing through the beds of basalt, 
 &c., overlying the drifts, inasmuch as large nuggets occur in 
 districts where basaltic eruptions have taken place, while, where 
 these are absent, the gold is very fine, and nuggets can scarcely 
 be said to exist. The fact has long been known that gold is 
 soluble in certain dilute solutions of salts, likely to be met with 
 in nature, such as a mixture of nitrates with chlorides, bromides 
 or iodides, or as the haloid ferric salts. This has been firmly 
 established by the researches of Skey, Daintree, Egleston and 
 others. Also, the precipitation of gold from these solutions around 
 nuclei consisting of particles of gold, pyrites, <fec., by organic matter 
 
70 THE METALLURGY OP GOLD. 
 
 present in the liquid, has been studied, and efforts made to form 
 nuggets similar to those found in nature, without much success. 
 This, however, is not surprising since the conditions in nature, 
 including almost unlimited time and immense quantities of ex- 
 ceedingly dilute solutions, cannot be reproduced in the laboratory. 
 Among other pieces of evidence against the erosion theory which 
 have been cited, may be mentioned the fact that some gold placers- 
 occur at higher levels than any quartz veins yet discovered or 
 likely to be discovered; also that nuggets have been found 
 embedded in decomposed rocks in positions to which they could 
 not possibly have been carried by running water, so that these 
 nuggets at least must have been formed by accretion. Some 
 regard must also be paid to the prevalent belief among diggers 
 that if a little " seed gold " is left in the tailings from sluicing 
 operations, the deposit will grow in richness so as to be worth 
 working over again after a few years. 
 
 Some of these arguments have been met by the exponents of 
 the erosion theory. The fineness of the placer gold has been 
 accounted for by supposing that the impurities (silver, copper, 
 &c.) formerly present in the native gold have been dissolved 
 away by meteoric water, in which they are much more soluble 
 than gold is. The existence of large masses of gold in placer 
 deposits was accounted for by Whitney by assuming that the 
 upper portions of the lodes, now washed away, were richer, 
 and contained larger masses of gold than the remains of the 
 lodes now lefc, but Liversedge has shown* that this assumption 
 is not necessary. Some nuggets too have been found showing 
 undoubted signs of erosion by water, but these are rare. Liver- 
 sedge has recently adduced evidence (loc. cit.) that, even if the 
 small particles of gold found in placers have grown by accretion, 
 nuggets cannot have appreciably increased in size. The sug- 
 gestions made to account for the great richness at bed-rock 
 viz., that gold has "settled" through the quicksands, or that the 
 gold solution has remained longest in contact with the sand 
 nearest bed-rock are not wholly satisfactory, and must be 
 supplemented by some such explanation as that given above,. 
 p. 68. 
 
 Minerals occurring in the Placer Deposits. In Califor- 
 nia, if quartz grains and silicified wood are excepted, the most- 
 abundant mineral is black iron-sand, which usually consists of 
 magnetite, although menaecanite, a form of hematite in which part 
 of the iron is replaced by titanium, also occurs. These minerals 
 must have been derived from the lavas, as neither of them are 
 known to occur in the quartz veins of the country. Platinum and 
 its allies are usually present in more or less abundance; thus iridos- 
 mine occurred to the extent of 1 in 100,000 of the gold in the 
 early days, and increased afterwards to 15 or 16 times that pro- 
 
 * Proc. of the Royal Soc. oftf.S. Wales, Sept., 1893. 
 
DEEP PLACER DEPOSITS. 71 
 
 portion. It is still abundant in the beach sands of Northern 
 California. Grains of native copper, nickel, and perhaps lead 
 have also been detected, and a few diamonds occur, while garnets, 
 small crystals of zircon and cinnabar are very abundant. 
 
 METHODS OF TREATING DEEP PLACER GRAVELS. 
 
 Both in California and Australia, when the first gold dis- 
 coveries were made, the river beds and bars were at once 
 explored, and soon afterwards the flats closely adjoining them. 
 Subsequently, the bench gravels situated in the same valleys, 
 and the side ravines and gulches, which remained dry during 
 most of the year, were prospected and worked, owing to the 
 rapid growth of the mining population and to the feet that 
 the exhaustion of the shallow placers was already beginning 
 to make itself felt. The result was that the exposed edges 
 of the outcrop of some of the deep leads were found, and 
 the pay-dirt followed into the hill-side by drifting. Then, as in 
 many cases it was found that the gravels overlying the pay-dirt 
 on hill-sides, although poor by comparison with the earth below 
 them, nevertheless contained a small quantity of gold, the idea 
 was evolved of breaking down the whole bank by jets of water, 
 and passing all the material through the sluices. Hence arose 
 the practices of drift mining and hydraulicking,* of which the 
 former is largely used in California, while the latter, though 
 much used in California prior to 1884 and in New Zealand, 
 cannot be applied in Australia or Siberia owing to the general 
 flatness of the country. In Siberia, only shallow placer deposits 
 are worked. In Australia, the deep leads are usually reached by 
 shafts, since the surface of the country is not intersected by deep 
 canons, as in California. 
 
 Hydraulicking. This method of working consists, as has 
 been already stated, in breaking down banks of gravel by the 
 impact of powerful jets of water, and passing the disintegrated 
 material through a line of sluices, without the agency of hand 
 labour. The chief requisites for the successful application of 
 hydraulicking are 
 
 1. Large quantities of auriferous gravel, not less than 30 feet 
 in thickness, and not overlaid by any appreciable thickness 
 of barren material, which would necessarily be passed through 
 the sluices with the pay-dirt. The gravel treated need not be 
 rich, a mean yield of less than 1 grain of gold per cubic yard 
 being often enough to furnish profits if the operations are on a 
 sufficiently large scale. 
 
 2. A plentiful and uninterrupted supply of water throughout 
 considerable portions of the year. 
 
 * The invention of hydraulicking is ascribed to Edward Mattison, of 
 Sterling, Connecticut, who used the method in 1851, on a small scale. 
 
72 THE METALLURGY OP GOLD. 
 
 3. Sufficient fall in the ground so that (a) the water may be 
 delivered under the pressure of a head of from 100 to 300 feet, 
 (6) the tailings can be easily carried away to a large dumping 
 ground, which is most conveniently either the sea, or a large 
 and rapid river. 
 
 Commencement of Operations. In California, the naturally 
 occurring banks or cliffs in the gravels in the sides of the gulches 
 were first selected for attack. Later, some of the deposits occur- 
 ring in those channels which are not intersected at favourable 
 points by the present system of drainage were operated on. It is 
 necessary in such cases to run a tunnel from the nearest canon in 
 the bed-rock to the lowest point in the gravel, this point being 
 found or guessed at by prospecting operations. The tunnels are 
 often of great length, that at the North Bloomfield mine, Nevada 
 Co., Cal., for example, being 7,874 feet or 1| miles long. One or 
 more shafts are then sunk from the surface through the gravel 
 to the tunnel, and washing operations are begun by ground- 
 sluicing, letting the water and gravel fall down the shaft and 
 run through the tunnel, in which the sluices are sometimes laid. 
 The surface near the shaft is thus gradually lowered, or it may 
 be terraced by hand labour, until an excavation is made of 
 sufficient size to enable the ground to be attacked with the hose. 
 Washing then proceeds regularly in this manner, the bank being 
 broken down by jets of water, and the products being allowed to 
 fall down the shaft and pass through the tunnel. Care must, 
 of course, be taken in the initial stages of the work to prevent 
 the blocking of the shaft, either by the caving in of its walls, 
 which are held apart by heavy timbering, or by runs of ground, 
 which may occur when the upper terrace is undermined by the 
 jets. After work has proceeded for some time, however, the 
 gravel immediately round the shaft is all washed away down to 
 bed-rock, and the shaft is then almost or entirely obliterated 
 (according to the location of the tunnel), giving place to a huge 
 open cut or funnel-shaped excavation, at the bottom of which 
 is the upper end of the tunnel. Having thus briefly indicated 
 the usual sequence of events in opening a hydraulic mine, the 
 plant and the ordinary method of working may be considered 
 under four heads. 
 
 1. The supply of water. 
 
 2. Breaking down the banks. 
 
 3. The washing proper, or sluicing. 
 
 4. Disposal of the tailings. 
 
 1. The Supply of Water. The amount of water required by 
 large undertakings is far more than can be obtained from the 
 rainfall on the hills immediately round the mine. Thus the 
 North Bloomfield Mine, in the season of 1877-8, used between 
 sixty and seventy millions of gallons per day, or enough for the 
 total supply of a city of three million inhabitants. The workings, 
 
 
DEEP PLACER DEPOSITS. 73 
 
 moreover, must necessarily be considerably above the level of 
 any large rivers in the neighbourhood, so that it is often neces- 
 sary to build huge reservoirs at convenient spots to store up the 
 rainfall and melted snows of large districts, and to convey the 
 water thence to the mine by ditches, flumes, or pipes. In Cali- 
 fornia the reservoirs have an aggregate capacity of 8,000,000,000 
 cubic feet, the largest, that of the South Yuba Water Company, 
 having a capacity of 1,800,000,000 cubic feet. 
 
 The ditches conveying the water from these reservoirs to the 
 mines pursue a course determined by the contour of the country. 
 They are usually cut in the sides of hills, and are given as far 
 as possible a uniform grade, which in different ditches varies 
 from 5 feet to 40 feet per mile. With the higher grades more 
 care in the lining of the ditch is necessary to prevent erosion of 
 the banks. The ditch may be lined by stones or boards, or left 
 unlined, but the loss from leakage amounts to several per cent, 
 of the delivery of unlined ditches. The ditch is sometimes 
 buried in the ground to prevent damage by avalanches, to keep 
 the water from freezing, and to reduce the loss by evaporation, 
 which, in the case of one ditch in California, was found to be 
 about 1 2 per cent, of the amount delivered from the reservoir. 
 Wooden flumes are also used instead of ditches, but though 
 cheaper, are more liable to be damaged. The dimensions of the 
 ditches vary enormously ; they range up to 100 miles in length in 
 California, and are from 2 to 15 feet wide, and from 1 to 5 feet deep. 
 In New Zealand ditches have been used less than wooden flumes, 
 but latterly, the use of sheet-iron pipes has been greatly extended 
 there, and many flumes have already been replaced by them. 
 These pipes are made as much as 30 inches in diameter, and are 
 cheap, durable and easily repaired; they entirely prevent losses by 
 leakage and evaporation, and deliver more water under similar 
 circumstances than flumes, as they offer less frictional resist- 
 ance to the flow. 
 
 Where it is necessary to cross side cafions or gulches the 
 flume or iron pipe is carried over on a light tressle bridge, or 
 the water is passed through an inverted siphon formed by a 
 wrought-iron pipe which passes down one side of the valley and 
 up the other, being filled from a head-box or reservoir, and 
 delivering the water at a lower level on the other side. At 
 Cherokee, Butte County, an inverted siphon pipe is used to 
 carry the water across a ravine 873 feet deep. The diameter of 
 the pipe is from 30 to 34 inches, and its greatest thickness (where 
 there is a pressure of 384 Ibs. per square inch) is 0-375 inch. * 
 The Miocene Ditch Company, operating in the same County, 
 carried their flume, of 4 feet wide and 3 feet deep, round the face 
 of a bluff 350 feet high, supported on L-shaped, iron brackets 
 made of bent T rails, soldered into holes previously drilled by 
 * Ninth Report, Gal. State Min. (1889), p. 125. 
 
74 THE METALLURGY OF GOLD. 
 
 men let down the face of the cliif by ropes. Successful ditch 
 construction often depends on such engineering feats.* 
 
 The water is delivered at a convenient height above the work- 
 ings into a head-box consisting of a small wooden reservoir from 
 which the pipes take their origin. In many cases the reservoirs 
 and ditches are owned by separate companies, or, as in New 
 Zealand, by Government, and the water is sold to the miners by 
 measure. The unit in New Zealand is a " government head," 
 and in the United States a "miner's inch," following the system 
 in vogue in Spain and Italy. The amount of water that will 
 flow through an orifice 1 inch square, cut in a board of 1 inch 
 thick, under a head of water that varies with the custom of the 
 locality but is usually from 4 to 8 inches, is called a miner's 
 inch. The amount of flow in twenty-four hours is called a 
 " twenty-four hour inch," and similarly there are ten-hour and 
 twelve- hour inches. The quantity of water in a miner's inch 
 varies with the head of water used and the form and size of the 
 orifice for delivery. Thus the amount delivered from an orifice 
 25 inches long and 2 inches wide is reckoned as 50 inches, 
 although it will be more than fifty times as much as the delivery 
 from an orifice 1 inch square. The twenty-four hour inch under 
 a head of 7 inches amounts to about 2,230 cubic feet.f 
 
 The water is conveyed from the head-box by pipes, which were 
 formerly made of canvas hose, to which iron rings 3 inches apart 
 were added for pressures of over 100-feet head. These latter 
 are called " crinoline hose," and were generally made of from 
 6 to 8 inches in diameter. They were used for some time, but 
 were found to suffer from rapid rotting, and to be liable to burst 
 at pressures of above a 200-feet head, and are now completely 
 replaced in the United States by sheet iron. In New Zealand 
 the canvas hose still lingers, but is now being rapidly replaced. 
 The iron feed-pipes are made from 10 to 15 inches in diameter, 
 and the thickness varies according to the pressure which they 
 may be called on to withstand. Sharp bends in them are avoided 
 as the flow of water is checked thereby. They are liable to 
 collapse if the level of the water in them is reduced, and a partial 
 vacuum formed inside ; hence, as in the case of all other sheet- 
 iron pipes used in hydraulicking, they are fitted with valves, 
 which are constructed so as to freely admit air from without. 
 The best and cheapest form of these is that used in New Zealand, 
 which has a 2-inch hole and a rubber clack like that used in 
 pumps. The valve is always open until the rising water lifts it 
 up on its surface, and closes the orifice. The water is discharged 
 
 * For details of the cost of construction, &c., of ditches, the student is 
 referred to Egleston's Mining and Metallurgy of Silver, Gold and Mercury 
 in the United States, vol. ii., pp. 120-180. 
 
 t For further details concerning miner's inch vide Art. by P. M. Randall 
 in Precious Metals in the United States, 1884, pp. 558-572. 
 
DEEP PLACER DEPOSITS. 1& 
 
 through a nozzle called a "giant" or "monitor." The nozzle 
 was at first a sheet-iron tube, having an aperture 1 inch in 
 diameter, and was held in the hand. The size of the nozzle was 
 gradually increased, until it has now reached a diameter of 
 11 inches. Such a stream, under a head of 200 feet, requires 
 special appliances to control it, deflect it at will, and prevent the 
 nozzle from " bucking." A number of ingenious contrivances 
 have been devised both in America and New Zealand, but all the 
 monitors bear a general resemblance to the one shown in Pig. 10. 
 2. Breaking Down the Bank. When the jet is first directed 
 against the bank, the water spatters in all directions, then buries 
 itself a little, and after a time, in loose ground, a " cave " takes 
 place, the undermined bank falling down. By the method of 
 undermining, the power of the giant is much increased, especially 
 where hard and soft layers alternate. When large caves are 
 about to take place the water is turned off, as otherwise the 
 
 Fig. 10. 
 Scale, 1 in. = 4 ft. 
 
 ground may run so far as to overwhelm the monitor and the 
 workman directing it. The nozzle is placed as near to the bank 
 as possible, consistent with the safety of the workers, so as not 
 to waste too much of the initial velocity of the stream of water. 
 Consequently, lofty banks are not advantageous, and if they 
 exceed 200 feet they are usually worked in terraces of 100 feet 
 or so in height. In some parts of the Spring Valley Mine, how- 
 ever (see Fig. 10a), a bank of 450 feet high was worked in a 
 single bench, and it was then not unusual for the runs of ground 
 to bury pipes which were throwing 7-inch streams from a dis- 
 tance of 400 feet from the face of the bank. 
 
 The jet is, if possible, delivered unbroken against the face of 
 the bank, as its disintegrating power is thus kept at its maximum. 
 However, in some cases, where it is cemented, the gravel is too 
 hard to be economically broken down by the water alone, and 
 blasting is then resorted to, a drift being run into the bank, and 
 cross-cuts made at the end in which the powder is placed ; the 
 
76 
 
 THE METALLURGY OF GOLD. 
 
DEEP PLACER DEPOSITS. 77 
 
 drift is then filled up, and the charge exploded by electricity. 
 It is more economical to blow out the base of the bank, as the 
 upper part then falls by its own weight and can be broken up 
 by the water. Sometimes arrangements are made to explode 
 very large blasts : thus, at the Blue Point Gravel Mine, a 
 charge of 50,000 Ibs. of powder was exploded at the end of a 
 drift 275 feet long in the year 1870, and 150,000 cubic yards of 
 gravel were brought down, while at another mine, 3,500 Ibs. of 
 dynamite were exploded in 1872, and 200,000 cubic yards of 
 gravel disintegrated at once. 
 
 3. Washing the Gravel in the Sluices. The sluices in which the 
 gold is caught are constructed on exactly the same principles 
 as those already described, but are larger and, though usually 
 made of wood, are of more massive construction, in accordance 
 with the great quantities of gravel to be handled and the con- 
 tinuous nature of the work. The sluices are commonly called 
 " flumes," but it is better to restrict the use of this word to a 
 conduit for carrying water only. The sluice boxes used in 
 hydraulicking, though, as usual, only 12 feet long, are as much 
 as from 3 to 6 feet wide and from 2 to 3 feet deep ; they are 
 lined with heavy planks on the sides, and the pavements are 
 made of more durable materials than is usual in shallow placer 
 sluicing, wooden blocks, rocks, or T railroad iron being most 
 usually employed. The wooden blocks are from 12 to 30 inches 
 square, and from 8 to 18 inches deep. They are usually 
 made of one of the softer varieties of pine (e.g., the " digger ; ' 
 pine, Pinus sabiniana, and others), which "brooms up" under 
 friction, and thus presents a better catching surface. The blocks 
 are cut across the grain of the wood, and are set side by side 
 across the sluice, each row separated from the next by strips of 
 wood to which they are nailed, while they are also kept in posi- 
 tion by the side lining which is placed upon them. The inter- 
 stices in the block pavement act as gold catchers, and are filled 
 with small stones, or, with less advantage, allowed to fill up 
 with gravel when washing begins. On account of the rapid 
 wearing away of the wood, much of the gold and amalgam caught 
 is scooped out and carried off again. Wooden block riffles only 
 last from two to four weeks when in heavy work, but are easy to 
 take up and put down again in cleaning up ; they are discarded 
 when worn so as to be only 4 to 6 inches thick. 
 
 Rock pavements are made of those boulders which are most 
 easily obtained in the particular district. Basalt is generally 
 used, oval stones of 15 or 18 inches long and from 9 to 12 inches 
 thick being selected and placed on end, with a slight slant in 
 the direction in which the current flows. They are held in place 
 by wooden planks which divide the sluice into compartments, so- 
 that if one stone works loose the pavement as a whole is not 
 affected. The interstices as before are filled with gravel. Rock 
 
78 THE METALLURGY OP GOLD. 
 
 pavements are very durable, lasting from three to six months, 
 but require more grade to the sluice, and occasion loss of time 
 in cleaning-up and re-paving the sluices. Consequently, they 
 are never used near the head of a sluice, where cleaning-up is a 
 frequent operation, but are often used for the lower parts 
 of sluices, where they sometimes alternate with block riffles, and 
 are especially suited for tail-sluices which are only cleaned-up 
 once a year. Rock pavements cost less than other forms of 
 riffles. 
 
 Iron riffles, which usually consist of T-iron rails, are placed 
 longitudinally in the sluice, closely packed side by side. They 
 present a large amount of space available for catching the gold 
 and amalgam, last well, present little resistance to the current (so 
 that the grade may be low while the duty of the water remains 
 high), and are easily taken up and put down. They are, therefore, 
 generally used at the head of the sluices. Though their first 
 -cost is higher than that of wooden blocks, they are more 
 economical in the end, owing to the saving of time in cleaning- 
 up and to their longer life. Egleston * instances the results of 
 experiments made at the Morning Star Claim, California, where 
 three sections of sluice, each 65 feet long, were laid at a distance 
 of 300 feet from the face of the bank which was being worked. 
 The first section was as usual laid with wooden blocks, the 
 second with old rails, and the third with rocks. When the 
 clean-up was made the middle section gave 9 ozs. more of gold 
 bullion than both the others combined. If old rails cannot be had, 
 strips of wood bound with iron are used, but are less durable 
 and satisfactory. 
 
 At the Blue Spur Consolidated Gold Company's plant, Gabriel's 
 Gully, New Zealand,! where the sluice is necessarily very short, 
 most of the stones are first separated from the gravel, and the 
 .finer material is then passed over a sluice paved with trans- 
 
 Fig. 11. 
 
 Scale = J. 
 
 -verse angle-iron riffles, placed with the hollow side facing down 
 stream (Fig. 11, in which the arrow shows the direction of the 
 stream); these iron riffles are placed 2 or 3 inches apart. Below 
 the section containing these riffles, there is a false bottom to the 
 .sluice, formed by an iron plate perforated with small round holes, 
 
 * Gold, Silver and Mercury in the United States, vol. ii. , p. 218. 
 t Report of the Mining Commissioner of New Zealand, 1891. 
 
DEEP PLACER DEPOSITS. 79 
 
 through which some of the water and the finest particles of the 
 gravel fall on to cocoa-nut fibre matting, laid on the true bottom 
 of the sluice. Here the fine gold is caught, the principle being 
 similar to that used in undercurrents. 
 
 The sluice is often divided into two by a median longitudinal 
 partition, so that one side may be at work while the other is 
 being cleaned-up or repaired, both sides being sometimes worked 
 when water is very plentiful. There are usually unpaved rock- 
 cuts above the sluice, leading to it from the places undergoing 
 the process of piping. These rock-cuts are rarely supplied with 
 mercury, and very little gold is usually caught there. The sluice 
 may be placed above the tunnel, or in the tunnel itself (one way 
 of preventing unlawful cleaning-up), or below it. In the case of 
 the North Bloomfield Mine, the irregular slates, dipping at a high 
 angle, forming the floor of the tunnel, were used as natural riffles 
 for the lower part of the sluice, and thus all cost of wooden 
 frames, pavements, &c., was saved, but the floor of the tunnel was 
 lowered by 3 feet, and deep holes worn in it, after 22,000,000 
 cubic yards of gravel had passed through it. 
 
 The length of the sluice, if capital is not lacking, depends on 
 the cost of construction and of the maintenance, as compared 
 with the value of the gold saved owing to the increased length of 
 the system. The length may be diminished by a plentiful use of 
 drops, grizzlies and undercurrents, all of which are described above 
 under the head of shallow placer sluicing ; they are made of pro- 
 portionately large size in hydraulicking. Coarse gold is of course 
 soon caught, but fine gold may successfully evade all the riffles of a 
 long sluice. The Spring Valley Mine has three parallel lines of 
 sluices, each 2 miles in length, and it is estimated that 95 per cent, 
 of the gold contents of the gravel is caught.* This length is un- 
 usual, the average not exceeding about 1,000 feet. The cost of 
 construction is from $25 to $35 per box (of 12 feet long) at the 
 larger Californian mines, and less in New Zealand. Here, by a 
 number of contrivances, especially by eliminating all but the 
 smallest stones by means of grizzlies or trommels, the length of 
 the sluices has been greatly diminished, but American engineers 
 regard the retention of the stones as desirable, owing to their 
 action in assisting disintegration. 
 
 Grade of the Sluices. The grade depends on the available fall 
 of the ground and on the character of the material to be washed. 
 The minimum is from 2 to 4 inches per box, such low grades 
 being sometimes enforced by the nature of the ground, some- 
 times adopted from choice if the gravel is light, the gold fine, 
 and water plentiful. With these low grades, however, disin- 
 tegration is slow and incomplete ; stones, unless they are small, 
 cannot be sluiced ; large ones block the sluices and must be re- 
 moved by hand, and the " duty " of the water, as regards sand, is 
 * Ninth Report Gal. State Min. (1889), p. 129. 
 
80 THE METALLURGY OF GOLD. 
 
 greatly decreased. The 6-inch grade is that most generally used, 
 but as much as 12 inches per box, where it can be obtained, 
 is regarded as desirable in good American practice. Higher 
 grades than this are unusual. Steep grades effect disintegration 
 rapidly, thus shortening the length of the sluice, and enable 
 all but the largest rocks to be sluiced, but less gold is then 
 caught and a more plentiful use of undercurrents is necessary. 
 In New Zealand an idea is gaining ground that sluices have been 
 made too narrow, and that if the width is greatly increased and 
 the grades diminished, and the depth of the current thus reduced 
 to a minimum, the losses of gold will be reduced. Under such 
 conditions all rocks must be removed by grizzlies, and in any case 
 the cemented gravels, so common in the Californian blue leads, 
 could not be treated in this way. It is considered necessary to 
 have a sufficient depth of water to cover the largest boulders to 
 be sluiced, but it undoubtedly diminishes the amount of gold 
 saved if the water is more than two inches deep. Nevertheless, 
 a depth of water of from 6 to 9 inches is not unusual. Where 
 poor or top gravel is being "piped," it is worked off as rapidly 
 as possible, and with less regard to the percentage of gold saved 
 than when rich stuff is treated, but with more regard to the 
 number of cubic yards treated. 
 
 The Use of Mercury. Mercury is added in great quantities 
 several times a day in America on the principle that " the more 
 quicksilver added the greater are the chances of catching the 
 gold," but less frequently and in less quantities in New Zealand. 
 In some Caliioruian sluices from 2 to 4 tons of mercury are in use 
 at once. The feeding is regulated by the appearance of the 
 amalgam in the sluice, the additions being made near the head- 
 box and in the undercurrents. The loss of mercury is usually 
 from 10 to 15 per cent, of the amount used per run. When 
 cemented gravels are being treated, owing to the extra amount of 
 trituration required, the loss may be as high as 30 per cent. 
 These losses are the more serious, for the reason that amalgam 
 is more easily lost than pure mercury, so that a heavy loss of 
 mercury denotes a heavy loss of gold. 
 
 Cleaninq-Up. The process does not differ from that described 
 under the heading of shallow placer mining. It is advisable not 
 to defer the clean-up too long as losses of amalgam are caused by 
 the wearing of the riffles. Usually from 50 to 95 per cent, of 
 the total yield of amalgam is caught in the first twenty or thirty 
 boxes, which are cleaned-up frequently. The following table* 
 shows the percentage yield of the various sections of the slu ices, 
 <fec., at the North Bloomfield mine, California, for the year 
 1877-8 : 
 
 Total yield, . . . . $311,276.20. 
 
 * Ninth Report Gal. State Min., p. 131. 
 
DEEP PLACER DEPOSITS. 81 
 
 Near bank, from rock cuts in mine (all in gold dust, no 
 
 quicksilver being added in the rock cuts), . . 4*57 per cent. 
 
 Sluice in tunnel (1,800 feet), 86'26 
 
 Tunnel below sluice (6,000 feet), see p. 79, . . . 4 '50 
 
 Cut below tunnel (200 feet), 081 
 
 Tail sluices (300 feet), 1-21 
 
 From seven undercurrents, . . . . . . 2 '65 
 
 100-00 
 
 The first undercurrent caught five times as much as the sixth, 
 and nearly three times as much as the seventh, which was of 
 double size. The yield of the seventh ($947) induced the Com- 
 pany to add another undercurrent. This mine affords an example 
 of the difficulty of catching fine gold. The gold loss was un- 
 known, but was believed not to exceed 5 per cent, of the contents 
 of the gravel. 
 
 The bullion obtained by retorting the amalgam from the 
 sluices is finer than that from quartz mills, and is sometimes 
 990 fine in Australia, although Californian placer gold is often 
 as low as 850 fine. The remainder is mainly silver, but copper, 
 lead, iron, and some of the minerals existing in the gravel also 
 occur. The amalgam from the head of the sluices yields finer 
 gold than that caught lower down and in the undercurrents. 
 
 Tailings. The disposal of the tailings is one of the most 
 important points to be considered in hydraulicking. Where 
 there is not sufficient fall to enable the tailings to be removed 
 from the lower end of the sluice without pumping, hydraulicking 
 is impossible. The tail-sluices usually terminate on the side of 
 a canon, in a river, or in the sea. The enormous amount of loose 
 sand and gravel, delivered from the hydraulic mines of Placer 
 County, California, and the neighbouring counties into the Yuba 
 and Feather rivers prior to 1880, filled up their beds to such an 
 extent that in rainy weather disastrous floods ensued, and much 
 valuable agricultural land was buried beneath sterile drift de- 
 posits and rendered worthless. The farmers thereupon took action 
 against the Mining Companies and obtained a perpetual injunc- 
 tion forbidding them to discharge their tailings into these rivers. 
 The result has been to stop hydraulicking in these districts, and 
 the efforts to work the deep leads more extensively by drifting, 
 or on the other hand, to impound the tailings by dams made of 
 brushwood, or to return them to their original position, have not 
 resulted in unqualified success. Consequently the gold winning 
 industry has not been maintained on the extensive scale it had 
 assumed prior to the action of the courts. 
 
 The Yuba and Feather rivers, in which it was estimated that 
 from 750,000,000 to 800,000,000 cubic yards of gravel were con- 
 tained, have continued during the last ten years to carry this 
 material lower down and to distribute it over the level ground of 
 the plains. The result has been that ground -sluicing on a 
 tremendous scale has been carried on, and the tailings in the 
 
 6 
 
82 THE METALLURGY OF GOLD. 
 
 upper parts, enriched by the removal of the valueless sands, will 
 probably pay to work over again in the immediate future. The 
 impounding of tailings behind brushwood dams is now believed 
 to afford a solution of the difficulty in California, and hydraulicking 
 will probably be recommenced on an extensive scale before long. 
 
 Drift and Shaft Mining. The problem of reaching the rich 
 leads lying in the gutters of the old river channels, without re- 
 moving the superincumbent masses of poor material, was attacked 
 in California before hydraulicking had been invented. The 
 method of following rich beds which passed into the hill-side by 
 means of timbered tunnels was practised before the upper beds 
 were suspected of containing any gold. The rise and rapid 
 progress of hydraulicking probably checked the development of 
 this tunnel or drift method between the years 1855 and 1880, 
 but it is now the most important branch of the placer mining 
 industry in California, while, as already stated, the analogous 
 method of shaft mining has always been almost exclusively in 
 use in Australia. Although hydraulicking was used wherever 
 practicable after its introduction, many deposits occur where it 
 cannot be employed. Where the upper gravels are thick and 
 almost sterile, and in particular where the pay-gravel is covered 
 by a great depth of lava (as for example at Table Mountain, 
 California), drift mining must be resorted to. Want of sufficient 
 grade, or of dumping ground, or of water supply, may also render 
 hydraulicking impossible. 
 
 At first work was done in California only on those leads which 
 were intersected by one of the existing canons, so that the drift 
 could be carried in paying ground from the start. In later 
 times, when the country had been more carefully surveyed, and 
 the probable courses of the Pliocene rivers roughly indicated, 
 efforts were made to reach leads at points far from any outcrop. 
 In order to do this, a tunnel is driven in the bed-rock from 
 the nearest canon to a point just below the gravel which it 
 is proposed to mine (Fig. 9, p. 65). Formerly tunnels were 
 often driven when the proofs of the position, depth and value of 
 the channel rented on evidence which would now be deemed 
 insufficient. Consequently they sometimes completely failed in 
 their object. Instances are on record of tunnels being driven 
 hundreds of yards in directions in which there was absolutely 
 no chance of encountering a river channel. The tunnel was 
 frequently put in too high, so that it was necessary to sink a 
 shaft at the inner end of the drift in order to reach the pay- 
 gravel, and thus the extra expense of raising water and gravel 
 vertically for some distance was incurred. During the past few 
 years, the location of the ancient channels, even at distances of 
 several miles from the nearest outcrop, has been determined 
 with some precision by engineers who have made a special study 
 of the subject. 
 
DEEP PLACER DEPOSITS. 83 
 
 If the bed-rock tunnel is successfully placed, so that, while 
 there is a fair grade outwards throughout its length, its inner end 
 is a few feet below the lowest point of the channel, the succeed- 
 ing operations are much cheapened and simplified, as " upraises " 
 to the gravel then serve to drain the workings and to deliver 
 the pay-dirt into cars, which convey it to the tunnel mouth. The 
 methods of cutting out the gravel bear a strong resemblance to 
 those used in coal mining, if the channel is wide and of fairly 
 uniform value, but the gravel is of course frequently tested by 
 panning. A detailed description of these methods of mining the 
 gravel, however, may with more propriety be given in treatises 
 on mining, than in this volume.* 
 
 The pay-gravel is carried to the tunnel mouth in cars, which 
 are propelled by hand, by horses, or by steam power, according 
 to the magnitude of the work. At Bald Mountain, California,! 
 the gravel is brought through a tunnel 7,000 feet long in a 
 train of eighteen cars, each holding 2 tons, by a locomotive 
 which performs the trip in five minutes. The cost of transpor- 
 tation at this mine is stated to be, by man-power 21 cents per 
 carload (of 2 tons), by mule-power 9 cents per carload, by steam 
 4f cents per carload. Outside the tunnel the gravel is dumped 
 into bins, whence it is delivered, by gravity if possible, either to 
 the sluices or to the batteries. In many cases the gravel is 
 cemented into a conglomerate which is too tough to be easily 
 disintegrated in the sluices. It is then passed through "cement 
 mills," which closely resemble the stamp battery to be described 
 in the next chapter, the chief differences to be noted being 
 in the facilities for delivery. Double discharge mortars are 
 used, and the screens are very coarse, the mesh being usually 
 about T 3 ^-inch, but varying up to J-inch in diameter. One 
 battery of ten stamps, each weighing 950 Ibs., making 94 drops 
 of 9 inches in height per minute, will crush about 40 or 50 
 tons of gravel in ten hours so that it will pass through a T ^--inch 
 mesh screen. Mercury is put into the mortar, and most of 
 the gold is usually caught there on amalgamated copper plates, 
 but copper plates outside the mortar are also used as in quartz- 
 milling, and rubbers are employed to brighten the gold. If 
 well-arranged plates are laid down, the number of sluice boxes 
 which can be added with advantage is very small, a length of 
 from 50 to 300 feet being used, the former limit being most 
 common. No attempt is made to save the auriferous magnetic 
 sands and sulphides which these conglomerates usually contain. 
 
 In cases where cement mills are not required, the gravel is 
 washed in sluices which differ little from those already described. 
 The boxes are not more than from 18 inches to 24 inches wide 
 
 * A detailed description is given by Russell L. Dunn, in Eighth Report of 
 al. Mate Min., 1888, p. 736. 
 
 t Ninth Report of Cal. State Min., 1889, p. 118. 
 
84 THE METALLURGY OP GOLD. 
 
 and deep, and the series is seldom more than 300 or 400 feet- 
 long Iron riffles are most in favour. Where the amount of 
 gravel to be washed is small, or the water is scarce, the gravel is- 
 allowed to accumulate for some time and the water stored in a 
 tank or reservoir. It is in some cases a great advantage to keep 
 compacted gravels exposed to the air during a few months before 
 washing them, as they " slack " and disintegrate under the in- 
 fluence of the weather, and subsequently are more easily treated,, 
 while for a similar reason, tailings are sometimes impounded, 
 and re- washed after some time has elapsed. The disintegration 
 of cemented material, which has been " slacked " by exposure to 
 the weather, is usually completed in a cement-pan. This is a 
 cast-iron pan with perforated bottom, and with a gate in the side 
 for the removal of boulders, which are mostly barren and are 
 separated from the auriferous material by this system, instead of 
 being crushed and mixed with it, as is the case when stamp-mills 
 are used. In the pan, four revolving arms, furnished with 
 plough-shares, break up the gravel, which is carried through the 
 apertures in the bottom by a stream of water, and falls into the 
 sluice. A pan of 5 feet in diameter and 2 feet in depth will treat 
 from 40 to 120 tons per day, according to the nature of the 
 gravel. (See also p. 61.) 
 
 The Hydraulic Elevator. In this machine a jet of water 
 under high pressure forces water, gravel, and boulders up an 
 inclined plane, and delivers them all at the head of the sluice, 
 which may be as much as 100 feet above bed-rock. The differences 
 in construction between the machines made in Australia, New 
 Zealand, and the United States are only matters of detail. 
 They consist essentially of an upraise pipe, usually of wrought 
 iron, having a diameter of from 12 to 24 inches, which terminates 
 below in an open conical funnel ; a hydraulic nozzle, delivering 
 water under the pressure given by a head of from 100 to 500 
 feet, projects into this funnel, and sand and gravel can also enter 
 round the sides or through a special orifice. The inclination of 
 the upraise pipe is usually from 45 to 65. The top of the 
 upraise pipe is turned over and terminates above a sluice, into- 
 which the gravel falls and is washed in the ordinary way. The 
 subjoined figure shows the arrangements at the base of the 
 upraise pipe of the elevator manufactured by Mr. J. Henry of 
 San Francisco. In Fig. 12 Nos. 1 to 13 are castings, No. 14 
 consisting of wrought iron ; a ball joint is formed by Nos. 3, 4,. 
 and 5, enabling the pipe bringing the water to be moved. The 
 nozzle and the lower part of the upraise pump are sunk in a 
 sump excavated in the bed-rock, and the gravel is washed down 
 by any means (usually by a jet from an ordinary hydraulic 
 nozzle) into this sump. The entrance to the upraise pipe is 
 protected by a coarse grating which prevents large stones, pieces 
 of wood, &c., from entering it. The force of water is enough to- 
 
DEEP PLACER DEPOSITS. 85 
 
 complete the disintegration of the gravel during its passage 
 through the upraise pipe, so that a sLort sluice is enough to 
 effect the washing proper. If the excavation is carefully arranged, 
 it may be kept funnel-shaped, so that the elevator, once placed in 
 a sump, may be worked there permanently without being moved. 
 When the pit is large enough, the washing may be done inside 
 it, only the tailings being raised to the surface by the hydraulic 
 elevator. This was done at Oroville in 1880,* the flume to 
 deliver the tailings into the river being about 500 feet long. 
 The head of water required varies according to the vertical 
 height through which the gravel must be raised ; a head of 
 about 70 feet is required for every 10 feet of vertical upraise. 
 
 Since it is just as expensive to raise water as gravel, arrange- 
 ments must be made to deliver as much gravel into the sump as 
 can possibly be raised by the jet, otherwise the expense per 
 cubic yard will be greater, and there will be too much water 
 with the gravel for satisfactory treatment in the sluices. 
 
 Wherever the necessary head of water is available, the 
 hydraulic elevator is now recognised as the best method of 
 working flat placers, river-bars, &c., or any deposits which are 
 either below the water level of the district, or which have not 
 sufficient fall for the disposal of the tailings by gravity. It is 
 in wide use in California and New Zealand. The following 
 instances of work in both countries may be given: At the 
 Blue Spur Consolidated Gold Mining Company, Gabriel's Gully, 
 New Zealand,! tailings, which have accumulated close to the 
 ea on the foreshore, are sluiced in this manner. The vertical 
 upraise is 60 feet, the angle of inclination of the upraise pipe 
 
 * Prod. Gold and Silver in the U.S.A., 1880, p. 15. 
 
 t New Zealand Mining Commissioners' Report, 1891, p. 65. 
 
86 THE METALLURGY OF GOLD. 
 
 being 6 3 -5 ; about 480 tons of gravel are raised per shift, the 
 head of water used being 400 feet, while the amount of water used 
 in each elevator is seventeen government heads. The sluice is 
 short, and has an inclination of only 3| inches in 12 feet; the 
 upper parts are fitted with transverse, patent |~"- shaped, angle 
 iron riffles, in which the angle faces uj) stream (see Fig. 11). The 
 lower parts of the sluice have a false bottom of wrought-iron 
 plates, perforated with round holes ; beneath these plates is the 
 true bottom of the sluice, covered with cocoanut matting in 
 which fine gold is caught. The tailings are discharged into the sea. 
 
 At Quartz Valley, Siskiyou County, California,* on hard 
 ground, where the elevator was first used, it took forty-three 
 days to work out a piece of ground 300 feet by 250 feet, which 
 was of an average depth of 18 feet. The bank was washed down 
 with 600 miner's inches of water, and went to the elevator 
 through a 30-inch bed-rock flume, which had a grade of 5 inches 
 in 12 feet. The water and gravel were raised through a 20-inch 
 elevator pipe without any contraction at the throat. It was 
 set at an angle of 40, and the pipe was 42 feet long, the vertical 
 upraise being thus 28 feet. The force used was 1,000 inches of 
 water with a head of 230 feet, delivered through a 6-inch nozzle, 
 and the gravel was emptied into a sluice 6 feet by 3 feet, with 
 a grade of 1J inches in 12 feet. When 3,500 inches of water 
 were running in this sluice, they could not carry off all the 
 gravel raised by the elevator. The work was done without any 
 delay from stoppage of the machine, and there were no repairs, 
 the wear of the elevator being very little. 
 
 Economic Conditions in Placer Working Shallow Placer 
 Deposits. In work by individuals the results differ greatly, 
 both according to the strength and skill of the worker and to 
 the contents of the gravel. Under the best conditions of climate,, 
 a strong, well-nourished, American digger may be able to raise 
 by the shovel from 10 to 12 cubic yards of gravel per day and 
 throw it into a receptacle 3 feet above the ground. Native 
 labour cannot be expected to effect so much, and in French 
 Guiana it is reckoned that only about half a cubic yard of earth 
 per man per day can be shovelled into the sluice. If the work- 
 man must wash the gravel, as well as raise it, much less can be 
 accomplished. For an active man, it is a fair day's work to 
 dig and wash from fifteen to twenty pans of dirt, the amount 
 treated thus not exceeding about 10 cubic feet. On the other 
 hand, with the cradle, the output may be from 1| to 2 cubic 
 yards per man in a day, while with the long- torn it may rise to 
 3 or 4 cubic yards per man. In the Siberian trough, only from 
 1 to 1J cubic yards can be treated by one worker per day, but 
 a large percentage of the gold is believed to be saved in this 
 
 * Egleston's Silver, Gold and Mercury in the United States, 1890, vol. ii.,. 
 p. 307. 
 
DEEP PLACER DEPOSITS. 87 
 
 apparatus. The minimum contents in gold, which will make 
 the gravel worth treating, depends, of course, on the cost of 
 labour in the country in which the deposit exists, since few men 
 will continue to work for less reward than they could obtain in 
 other employments. The wages of placer miners in various 
 countries is as follows : In California, from $2J to $4 per day ; 
 in Colorado, Montana, Dakota, and the neighbouring States, from 
 $2 to $3 1 per day ; in Australia, from 6s. to 10s. per day ; in 
 French Guiana, the negroes and coolies are paid about 3s. 4d. 
 per day, exclusive of rations. The wages in Siberia are given 
 below. In California, the Chinese are content to earn about 
 75 cents or 3s. per day. In early times, in California and 
 Australia, when the virgin shallow deposits were being worked, 
 large sums were often realised by individual diggers, cases being 
 on record in which 5 ozs. of gold were obtained from one pan of 
 bedrock scrapings lying under heavy gravel, and earnings of 
 several hundred dollars per day were not uncommon. The re- 
 sults obtained on the Klondike in Canada are still more remark- 
 able. Authenticated instances of 15 oz. of gold per pan are 
 recorded, and an average of $5 to $7 per pan was obtained on 
 certain claims in 1897. 
 
 Concerted work, with the aid of the sluice, is much more 
 effective ; in California gravel containing about 1 pennyweight 
 of gold per cubic yard is worked at a profit, the dirt being lifted 
 into the sluice by hand-labour, and the tailings removed by 
 sluicing with water ; at Ballarat, in Australia, where the gravel 
 is raised to the surface from underground workings through a 
 vertical shaft several hundred feet deep, and subsequently 
 washed, 12 grains of gold per cubic yard of material pay for the 
 treatment, while in Siberia, as stated below, the cost is even less. 
 In French Guiana, the unhealthiness of the climate and the cost 
 of supplies render it impossible to work gravel containing less 
 than about 3 pennyweights of gold per cubic yard.* 
 
 In Siberia,! the distance of the workings from the nearest town, 
 and the traditions of the industry, require workmen to be hired 
 by the year, or, in cases where no work is attempted in winter, 
 for the season. The best workmen, employed in actual washing, 
 receive 1 rouble (about 2s.) per day; the diggers and carters 
 75 copecks (Is. 6d.), and the women and boys from 40 to 50 
 copecks (9d. to Is.) per day : rations are also distributed in mines 
 far from towns. Sometimes piece-work is done, the price paid 
 being from 2^ to 2^ roubles per cubic sagen, or from 4d. to 5^d. 
 per cubic yard (!); the horses and men are then both supplied by 
 the workman. The total cost of treatment of the gravel varies 
 greatly with its geographical position. On the banks of the 
 Lena, where the season only lasts for five months, it is estimated 
 that the gravel must contain 2 zollatniks of gold per 100 pouds, 
 or 4^ dwts. per cubic yard ; but in the neighbourhood of Ekater- 
 * Frmy, Ency. Chim., vol. v., L'or, p. 48. t Ibid, p. 47. 
 
88 THE METALLURGY OP GOLD. 
 
 ineberg, the deposits in the bed of the Pechma, a tributary of the 
 Obi, are worked when they only contain from 8 to 9 grains per 
 cubic yard. This is less than that required in other countries 
 in order that a profit may be made from similar deposits, viz., 
 those in which the gravel must be handled at least twice. 
 
 The cost of working the perennially frozen placers is much 
 more, but Levat gives an instance in which a bed of gravel, 
 9 feet thick, yielding about 1J dwts. per yard, and covered by 
 100 feet of barren material, was worked at Malamalski in the 
 Trans-Baikal.* 
 
 Deep Placer Deposits. The cost of hydraulicking and drift 
 mining necessarily varies enormously with the conditions. In 
 hydraulicking, it depends largely on the magnitude of the opera- 
 tions. With large quantities of water available at a cheap rate, 
 big banks of soft gravel, and a well-constructed and convenient 
 taflmgs-tunnel, the cost has been reduced to from 2 to 3 cents 
 per cubic yard in California, while the average cost there in 1884 
 was about 10 cents, and only in exceptional cases amounted 
 to as much as 50 cents per cubic yard. This is without reckon- 
 ing interest on capital expended. As already stated, hydraulick- 
 ing is rarely possible in Australia, and then only on a small 
 scale. A case in Australia is cited, however, by Mr. G. Warn- 
 ford Lock, in which three workmen washed 150 cubic yards of 
 gravel per day with 1,875 cubic yards of water, the jet being 
 directed against a bank about 30 feet high. The cost of work- 
 ing was 2 per day, or a little more than 3d. per cubic yard, 
 and a saving of 1 J grains of gold per cubic yard would, in this 
 case, pay expenses. The cost of production in the year 1875 
 of 1 ounce troy of gold in hydraulicking is given on the next 
 page for two Californian mines, f viz., the La Grange Company, 
 which worked on a thin bed of gravel, and the North Bloomfield 
 Mine, in which the bank was 250 feet high. 
 
 In drift mining, the cost of treating uncemented gravel is less 
 than that of treating conglomerate. The cost at the Hidden 
 Treasure Mine, on the Damascus Channel, Placer County, Cali- 
 fornia, was $0-9236 per carload of 1 ton in the spring of 1888, 
 the yield being $1*2347 per ton. Here the ground could all be 
 " picked " down, requiring no blasting, and the gravel could be 
 sluiced without crushing, but the timbering was costly. The 
 working of this mine was exceptionally cheap, as, under ordinary 
 conditions, a cost of from $1-50 to f 1'75 per carload is as low as 
 can be expected, and in many mines the cost rises to $3 per 
 carload. The cost of milling cemented gravel is from 20 to 40 
 cents per ton, the capacity of the mills being from 5 to 12 tons 
 per stamp in twenty-four hours. 
 
 * See L'Or en Siberie Orientate, vol. i. , p. 146. 
 t Ninth Report Cod. State Min., 1889, p. 133. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 89 
 
 
 LA GRANGE. 
 
 NORTH BLOOMFIELD 
 
 
 Per cubic yard 
 of Gravel. 
 
 Per ounce 
 of Gold. 
 
 Per cubic yard 
 of Gravel. 
 
 Per ounce 
 of Gold. 
 
 Water, 
 
 0'8 cents. 
 
 $1-43 
 
 0'75 cents. 
 
 $2-09 
 
 Labour, 
 
 3-6 
 
 6-85 
 
 1-40 
 
 3-93 
 
 Materials, 
 
 1-0 
 
 1-81 
 
 0-32 
 
 0-88 
 
 Explosives, 
 
 
 ... 
 
 0-35 
 
 0-98 
 
 Blocks, 
 
 ... 
 
 ... 
 
 0-18 
 
 0-50 
 
 General expenses, 
 
 
 
 0-25 
 
 0-70 
 
 Contingent, . 
 Salaries of officials, 
 
 0-6 
 
 0-26 
 094 
 
 
 .. 
 
 Taxes, . 
 
 
 0-09 
 
 ... 
 
 ... 
 
 
 6'0 cents. 
 
 $11-38 
 
 3'25 cents. 
 
 $9-08 
 
 One cubic yard usually contains from 1^ to If tons of gravel. 
 
 The relative costs of working the various classes of gold 
 deposits in California, by methods applicable to the respective 
 classes, are given by John Hays Hammond,* as follows: 
 
 1. Auriferous vein deposits (mining and milling), 
 
 2. Drift mining, ....... 
 
 3. Miner's pan, . . . . . 
 
 4. Rocker or cradle 
 
 5. Sluicing, 
 
 6. Hydraulic method, ...... 
 
 Cost per Ton of 
 Material treated. 
 
 $3 to $10. 
 
 75 cents to $4. 
 
 &5 to $8. 
 
 $2 to $3. 
 
 75 cents to $1. 
 
 If to 8 cents. 
 
 CHAPTER VI. 
 
 QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 Primitive Methods of Crushing and Amalgamation. In all 
 
 countries, when the richest alluvial deposits have been worked 
 out or have all been taken possession of, efforts have been 
 made to extract the gold from the various hard, auriferous 
 materials met with in veins. The earliest machines used for the 
 purpose seem to have been stone mortars, in which the gold 
 quartz was crushed by stone-hammers, or by large rocks raised 
 by levers and let fall, whilst the fine material was subsequently 
 washed out with water. The heavy particles of gold then settled 
 to the bottom of the mortar, the lighter, worthless material 
 being washed away. Hollo wed-out stone mortars, in which this 
 
 * Ninth Report Gal. State Mine, 1889, p. 105. 
 
90 THE METALLURGY OP GOLD. 
 
 operation was conducted, have been found in Wales, in Central 
 America, in the Pyrenees, and in Transylvania. Diodorus 
 Siculus, the Greek historian, has given a detailed description of 
 the method of gold-quartz reduction employed in the mines of 
 Upper Egypt 1,900 years ago.* The ore was first reduced to- 
 coarse powder in mortars, then finely crushed in hand-mills 
 resembling the flour-mills of the present day, and finally washed. 
 In order to separate the pulp from the uncrushed lumps, the 
 Egyptians, in common with other races in ancient times, employed 
 sieves, but in extracting the gold from auriferous sands they used 
 raw hides, on which the flakes of gold were entangled. f These 
 devices closely resemble those still in use in many parts of the 
 world. I 
 
 The date of the first use of mercury for amalgamation is 
 unknown, but it has no doubt been used for this purpose for 
 the last 2,000 years. The earliest mention of quicksilver itself 
 appears to occur in the works of Theophrastus, about B.C. 300 ; 
 but Diodorus, in the account just mentioned, does not refer to 
 its use. Only a few years later, however, Vitruvius, about 
 B.C. 13, described the manner in whicji, by the help of mercury, 
 gold was recovered from cloth in which it had been interwoven, 
 and in Pliny's time the separation of gold from its impurities 
 generally by the same means was well known. || It is probable 
 that this knowledge was never afterwards entirely lost, although 
 the references to it in the Middle Ages are very scanty. For 
 example, Geber,U in the eighth century, was aware that mercury 
 would dissolve considerable quantities of gold and silver, but 
 not earthy materials, and Theophilus, the monk,**" in the 
 eleventh century, carefully described the method of washing 
 the sands of the Rhine on wooden tables, the final operation 
 consisting in treating the concentrates with quicksilver for the 
 removal of the gold. The extraction of gold by mercury is 
 described by Biringuccio in his treatise (which was published in 
 Italian in 1540), ft and had apparently been a secret art for 
 some time previously. 
 
 Mercury was introduced into Mexico as a means of extracting 
 
 * Diodor., iii., 13. A full translation is given by B. H. Brough in his 
 Cantor Lectures on Mine Surveying, Journ. Soc. Arts, 1892. 
 
 t Beckmann's Hist, of Inventions, vol. ii., p. 334. 
 
 $ An interesting account of some primitive methods of treating gold 
 quartz now employed by the Chinese is given by Henry Louis in the Eng. 
 and Mny. Journ., Dec. 31, 1892, p. 629, from which it appears that these 
 methods bear points of resemblance to those of the Egyptians. 
 
 Vit., lib. vii., cap. viii. 
 
 || Nat. Hist., lib. xxx., cap. vi., sect. 32. Quoted in full in Percy's 
 Metallurgy of Gold and Silver, p. 559. 
 
 IT Salmon's Geber, cap. xlvii. 
 
 ** Theophili, lib. iii., cap. xlix. 
 
 tt De la Pirotcchnia, Venice, 1540. Lib. ix., cap. xi., fol. 142. 
 
QUARTZ CRUSHING IN THE 
 
 the precious metals by Bartolome Medina in the year 1557,* 
 and its use doubtless soon spread to Peru and the neighbouring 
 countries. When Barba wrote in 1639 there were three amal- 
 gamation machines in use in Peru,f viz., the mortar (used in the 
 Tintin process), the trapiche or Chilian mill, and the maray. 
 The mortar was hollowed out in a hard stone, the cavity being 
 about 9 inches both in diameter and in depth. The ore was 
 triturated with water and mercury in this mortar with the aid of 
 an iron pestle, while a stream of water, flowing through, carried 
 away the crushed particles. The slimes were roughly concen- 
 trated, but most of the gold remained in the morfear with the 
 mercury. 
 
 The maray, although equally primitive, was probably of greater 
 capacity : it makes use of the principle employed in the bucking 
 hammer. About the year 1825 Miers \ saw it in Chili, where ic 
 is probably still in use. It consists of a flat or slightly concave 
 stone, 3 feet in diameter, on which lies a spherical boulder of 
 granite about 2 feet in diameter. This is rolled to and fro by- 
 two men seated on the ends of a long, wooden pole, which is 
 firmly fixed to the boulder. Ore, water and mercury are ground 
 together in this machine, and then washed down. 
 
 The Chilian mill closely resembles the edge-runner mill of the 
 present day, which is used for grinding and mixing mortar, &c. 
 The Peruvian trapiche had a similar circular bed of hard stone, 
 but only one stone runner, which was driven by mules. The 
 Chilian mill is still used to prepare ores for treatment in the 
 arrastra, which was not mentioned by Barba, and may perhaps 
 be regarded as an outcome of the trapiche. 
 
 Among other ancient methods of amalgamating ores, the tina 
 and cazo systems may be mentioned. The tina system, largely 
 used in Chili, at least as late as the year 1853, is a modifica- 
 tion ot that still at work in the Tyrol (see p. 165). The old 
 and long-abandoned Norwegian method of treatment, described 
 by Schliiter in 1738,|| and formerly practised at Kongsberg, waa 
 also, no doubt, derived from the same source. In Norway and 
 Chili, however, the working was not continuous. Charges of 
 crushed ore were introduced into wooden tubs (tinas] furnished 
 with cast-iron bottoms and stirred up with water and mercury 
 by means of revolving iron agitators, comparable to the mullers 
 of the modern amalgamating pan. At Kongsberg only 80 to 
 100 Ibs. of ore were treated at one time with about 40 Ibs. of 
 mercury, but in Chili a tina contained from 4 to 6 cwt. of ore 
 
 * Georg Agricola's Bermannus . . . ubersetzt von Friedrich August 
 Schmid, 1806, p. 49. 
 t Arte de losmetales, lib. iii., cap. xxi. 
 
 J Travels in Chile and La Plata, by John Miers, vol. ii. , p. 398. 
 Percy's Metallurgy of Silver and Gold, p. 567. 
 II Schliiter's Grundticher Unterricht von Rutte- Werken, p. 211. 
 
92 
 
 THE METALLURGY OF GOLD. 
 
 and 150 Ibs. of mercury or more. After the agitators had been 
 at work for about four hours, making sixteen revolutions per 
 minute, the mercury was drawn off, squeezed through calf-skin 
 or canvas bags, and the amalgam removed from the bags and 
 retorted. The cazo process, introduced in Peru in 1609,* was 
 similar, but the bottom of the vessel was of copper and the 
 pulp was boiled. These processes were used in treating ores 
 rich in silver. 
 
 The Arrastra was also one of the earliest crushing machines 
 in use in America, being introduced at the same time as the 
 Patio process i.e., about 1557 and is still in wide use in Mexico, 
 although chiefly in the treatment of silver ores by the Patio 
 process. It is a circular, shallow, flat-bottomed pit, 10 to 20 feet 
 in diameter, and paved with hard, uncut stones. Granite, basalt, 
 
 Fig. 13. 
 
 and compact quartz are all used for the rock pavement, which is 
 made 12 inches thick, and is either placed on a bed of well- 
 puddled clay from 3 to 6 inches thick, or set in hydraulic cement, 
 so that no chink or cranny remains, into which the mercury or 
 amalgam can find its way. In the centre a vertical shaft revolves, 
 carrying two or four horizontal arms, to each of which is attached 
 a heavy stone by thongs of bullock hide, or by chains. These 
 grinding stones weigh from 400 to 1,000 Ibs. each, their forward 
 ends being about 2 inches above the floor, whilst their other ends 
 drag on it. They are moved by mules walking round outside 
 the arrastra, or by water- or steam-power, the speed varying from 
 four to eighteen turns per minute. Fig. 13 represents an 
 arrastra of the simplest description ; at the front the stones 
 forming the edge have been removed, so as to expose a section 
 of the rock pavement. 
 
 Ore of about the size of pigeons' eggs is introduced, enough 
 
 water being added to make the pulp of the consistency of cream, 
 
 and mercury is sprinkled over it after most of the grinding has 
 
 been done. When the ore cannot be ground any finer, more 
 
 * Barba, Arte de los Metales. Loc. cit. ant. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 93 
 
 water is added to dilute the pulp, the mule is driven slowly for 
 half an hour to collect the mercury, and the pulp is then run 
 out and another charge shovelled in. An arrastra of 10 feet in 
 diameter takes a charge of 500 to 600 Ibs. of ore, and treats about 
 1 ton per day of twenty-four hours. The amount of wear sus- 
 tained by the grinding stones is equal to from 6 to 10 per cent. 
 of the ore crushed. The output is so limited that the use of the 
 arrastra has never been general outside Mexico, although it has 
 been used in almost every district in the United States for a 
 short time after the commencement of quartz mining in the 
 particular locality. 
 
 It is often stated that the results obtained, so far as the 
 percentage of gold extracted from refractory ores is concerned, 
 cannot be equalled by any other amalgamating appliance, and 
 that the Mexicans, using the arrastra, formerly treated at a profit 
 ores which hardly yield any gold to the stamp mill, or even to 
 the amalgamating pan. In consequence of its power of saving 
 gold and the cheapness with which it can be erected and worked, 
 the arrastra is still valuable for prospecting. In preparing the 
 bed for this purpose, every care must be taken that the surface is 
 even, and all joints properly cemented, or else the mercury and 
 amalgam will in great part find their way into the founda- 
 tions. 
 
 The reasons for the high extractive power of the arrastra, when 
 treating certain ores, are no doubt to be found in the extreme 
 fineness to which the ore is reduced, and in the prolonged contact 
 between the ore and mercury, which is maintained while they are 
 being ground together. Moreover, the grinding action of the 
 dragged stones keeps the particles of gold bright and in a suitable 
 condition for amalgamation, without exercising force enough to 
 flatten and harden them. The relative advantages of grinding 
 surfaces consisting of iron and stone are less certain, and probably 
 vary with the nature of the ore in course of treatment. An 
 instance is given by Colonel Harris,* in which stone was better 
 than iron. In this case, at Cerro de Pasco, Peru, the old method 
 of grinding ores in circular pans, by edge-runners of stone or 
 granite, was found to entail a rapid wearing of the edge-runners, 
 and, in order to remedy this, the runners were shod with iron. 
 The returns at once fell off, and on careful trial it was found that 
 the yield in the old machines was 15 per cent, more than in the 
 new ones. Other similar instances are on record, but it must 
 nevertheless be conceded that, with some ores, the presence of 
 iron is necessary for good work, chlorides of silver and mercury 
 being reduced by it, which would otherwise be lost in the 
 tailings. The somewhat extravagant views often expressed, 
 upholding the great superiority of the arrastra over its more 
 modern rivals as an amalgamating machine for gold ores, are 
 * Min. Journ., December 24, 1892, p. 1466. 
 
THE METALLURGY OF GOLD. 
 
 perhaps hardly justified, since a direct comparison with stamps 
 or pans has been possible in only a few instances. 
 
 One of the most remarkable of these instances is that afforded 
 by the experience at the Pestarena Mines, Val Anzasca, Italy. 
 The ore of this mine contains about 20 per cent, of pyrites, and 
 is of somewhat low grade, rarely containing as much as 15 dwts. 
 of gold per ton. Efforts were at first made to treat it by means 
 of stamps and amalgamated plates, with the result that only 
 65 per cent, of the gold was extracted. Better results attended 
 the introduction of the Francfort mill, a modified arrastra, driven 
 in this instance by steam ; this mill is substantially a wooden 
 pan, with dies and shoes of stone. The mercury was added to 
 the pulp after it had been finely ground, and the amalgamation 
 -and grinding of the pulp subsequently kept up for seven hours. 
 From ore containing 12-3 dwts. per ton, the mill extracted 10 "2 
 dwts., or 83 per cent., for a considerable period of time, while in 
 the year 1890 the average extraction was 794 per cent. There 
 were twenty-eight mills, each of which treated 1,200 Ibs. of ore 
 per day. The stone bed is said to last ten months, and the shoes 
 from six to eight weeks, the total cost of treatment being very 
 low. 
 
 Successful work by the arrastra on gold ores is also being done 
 at the present day at the Bote Mine, Zacatecas. The material 
 treated is an auriferous silver ore, the gold being extracted in 
 the arrastra, and the silver subsequently obtained by means of 
 the Patio process. Here the closely-fitting cut-stones forming the 
 bottom of the arrastra are first coated with an amalgam of silver, 
 and the ore then ground for twelve hours, mercury being added 
 to it to the extent of one and a-half times the weight of the gold 
 present, so as to form a "dry" amalgam. The extraction of 
 gold in this case, however, is only 60 per cent., a result which is 
 probably not better than that which would be obtained in pans 
 and settlers, if economical conditions admitted of their employment. 
 The necessity of keeping the amount of mercury low so as to 
 obtain " dry" amalgam, and thus to prevent loss by leakage into 
 the foundations, is one of the objections to the use of a roughly 
 made stone-floor for amalgamation, the percentage of extraction 
 being necessarily reduced by this precaution. There were about 
 100 arrastras at work in California in 1889, each treating from 
 1 to 3 tons of ore per day. They are used where only small 
 quantities of high-grade ore are available for treatment. 
 
 At Smartsville, and in other places in California, the arrastra 
 is now applied to crushing and amalgamating hard cement- 
 gravel obtained from a drift mine.* At Smartsville the cement 
 is coarsely crushed by be ng passed through a Gates rock-breaker, 
 and is then charged into four arrastras, each of which is 12 feet 
 in diameter and 3 feet deep and capable of containing from 5 to 
 * Eleventh Report Cal. State Min., 1892, p. 315. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 95 
 
 9 tons of gravel. The grinding is effected by four blocks of diabase, 
 ach weighing from 600 to 1,000 Ibs.; the rate of revolution is 
 14 times per minute, the time of grinding being one hour. A 
 tablespoonful of mercury is fed in with each charge, and the 
 total loss is only 10 per cent, of this. The pavement costs $40 
 and lasts for six months. At the end of the hour, a gate is 
 opened in the side of the arrastra and the charge run out into a 
 sluice, 200 feet long, where the mercury and amalgam is caught 
 by means of riffles. The capacity of one arrastra is 50 tons of 
 hard cement per day, or 75 to 90 tons of soft surface-gravel per 
 day. The cost is from 6 to 8 cents per ton, and more gold is 
 extracted than when a stamp battery was used at a cost of from 
 20 to 40 cents per ton. One man attends to the four arrastras, 
 and another to the Gates crusher. 
 
 Iron Prospecting Arrastra. The difficulties of getting suit- 
 able stone for the arrastra in some localities, and of constructing 
 the pavement so as to prevent loss of mercury, have led to the 
 introduction, for prospecting purposes, of wrought-iron pans with 
 steel bottoms on which stone drags are used. One of these so- 
 called arrastras is 3 feet in diameter and 12 inches deep, with an 
 axis revolved by bevel gear, which is placed below the bottom. 
 The upper part of the axis carries a yoke to which stone drags 
 are attached by chains. It is doubtful if the steel bottom is as 
 good for amalgamation as the stone pavements, but the losses 
 of mercury by leakage are certainly avoided. Perhaps a thin 
 pavement of flat stones might be put inside such a prospecting 
 pan with advantage, when working on some ores. 
 
 The Stamp Battery. The stamp battery evolved, no doubt, 
 from the pestle and mortar was not introduced until a com- 
 paratively recent date. Beckmann* states that mortars, mills, 
 and sieves were used exclusively in Germany throughout the 
 whole of the 15th century, and in France stamps were unknown 
 as late as the year 1579. Brough has suggested that the origin 
 of stamp mills was probably due to the manufacture of gun- 
 powder. It seems certain that in 1340 a stamp mill, used in 
 connection with this industry, existed in Augsburg, and that 
 Conrad Harscher, of Nuremburg, owned one in 1435. They 
 were first applied to the gold industry at the beginning of the 
 16th century, a doubtful record stating that they were intro- 
 duced into Saxony by Count von Maltitz in 1505, whilst in 1519 
 the processes of wet-stamping and sifting were established in Joa- 
 chimsthal by Paul Grommestetter, who had some time previously 
 introduced them at Schneeberg. The improvements gradually 
 spread through Germany, and detailed descriptions and drawings 
 of the apparatus are given by Agricola f in 1556, from which it 
 appears that the earliest battery consisted of a single stamp, 
 
 * Hist, of Inventions, vol. ii., p. 334. 
 
 t De re Metallica, vol. viii., pp. 233 and 247. 
 
96 THE METALLURGY OP GOLD. 
 
 raised by means of two levers fixed to the axle of a wheel. Dry- 
 crushing was at first employed, but the great production of dust 
 soon led to the use of water in the mortar. In some Hungarian 
 mines, Bennett H. B rough* recently saw some primitive stamps 
 in use, resembling those drawn by Agricola, weighing only 
 100 Ibs. each, and having their heads made in some cases of 
 hard blocks of quartzite. At that time, in cases where the 
 conditions of water supply were favourable, these stamps were 
 able to treat with profit an ore containing as little as 2^ ozs. of 
 gold to 50 tons of ore, and at Zell, in the Tyrol, they were 
 able to treat a slaty material containing 1 oz. of gold to 50 tons 
 of ore. Such economical work is seldom possible with the 
 modern Californian stamp under the most favourable circum- 
 stances. 
 
 Agricola's exact description of the treatment of auriferous 
 quartz in Germany in 1556 shows that the methods in use at 
 that time were strikingly similar to those still employed in 
 Transylvania and the Tyrol, which were among the districts 
 of which he wrote. Doubtless in these districts, the methods 
 have been handed down from generation to generation with 
 little change, while in other countries, where tiiey were intro- 
 duced hundreds of years later, the changes have been rapid and 
 striking. In those points in which the older differed from 
 the modern Tyrolean practice, it resembled the practice of the 
 ancient Egyptians, so that the origin of the methods may, 
 perhaps, be traced back to them. The wooden stamps, shod 
 with hard stone or iron, were arranged in sets of three, and 
 raised by cams to fall by gravity when released. The rock was 
 shovelled dry into the mortar, and coarsely crushed by the 
 blows of the stamps. 
 
 Next, it was ground as fine as flour in a stone mill supplied 
 with water, and carried by the stream of water into the upper- 
 most of three wooden tubs, whence it overflowed in succession 
 into the other two. Revolving mechanical stirrers, furnished 
 with six paddles, kept in agitation the contents of the tubs and 
 " separated even very minute flakes of gold from the crushed 
 ore. These flakes, settling to the bottom are drawn to itself 
 and cleansed by the quicksilver (lying in the tubs), but the 
 water carries off the dross."f Agricola here expounds the 
 theory of amalgamation still adhered to in Austria, where 
 mercury is regarded merely as a useful means of collecting 
 particles of gold, which have already been separated from the 
 crushed ore by their great density. The Tyrolean bowls, still 
 in use at Vorospatack in Hungary and in a few retired valleys 
 in the Eastern Alps, do not differ essentially from the tubs 
 
 * Proc. Inst Civil Eng., Session 1891-1892, part ii. 
 t Agricola, loc. cit. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 97 
 
 drawn and described by Agricola ; and, although wet crushing 
 by the stamps has been introduced, the mortar is even now 
 seldom furnished with screens in these mills. 
 
 Elsewhere the changes in stamp battery practice, introduced 
 since Agricola wrote his treatise, have been many and great. 
 One of the first was the insertion of screens in the side of the 
 mortar, so that the two operations of crushing and sifting were 
 united. In 1767 M. Jars saw these in use in the Hartz,* 
 though even then only a single screen of brass wire 12 inches 
 square delivered the product of three stamps, and in several 
 other districts in Germany screens had not been adopted. The 
 screen was completely protected from the splash of the stamps, 
 so that the sieving of the ore was very slowly effected by a 
 current of water flowing through. 
 
 The most important improvement, however, has undoubtedly 
 consisted in combining the operations of crushing and amalgama- 
 tion by charging mercury with the ore into the battery, and 
 placing amalgamated copper plates in the mortar, so as to catch 
 gold. No mention of these practices appears to have been placed 
 on record before stamp batteries began to work in California 
 in 1850, although they had possibly been adopted in Georgia 
 before that time. 
 
 The use of the copper plate was probably suggested by the 
 experience in Mexico and South America of the working of 
 the Cazo process, in which it was well known that amalgam 
 tended to adhere to the copper sides of the vessel unless the 
 proportion of mercury to gold and silver present was less than 
 4 to 1. Thus Baron Born wrote in 1786, f "In new kettles 
 . . . the inside becomes wholly and so perfectly silvered 
 that it never can be cleaned. . . . The silvery coat is 
 daily increased by slow and gradual apposition, and the crusts 
 of amalgama, accumulating on the bottom and sides of the 
 vessels, become gradually so thick that on emptying them they 
 often fall off by their own weight as silver plates, which, when 
 dry, show the laminated texture of their daily augmentation." 
 
 The knowledge of this behaviour of amalgam in the Cazo 
 process must have been common to many who were engaged in 
 exploiting the quartz veins of the West soon after their dis- 
 covery, and the speedy application of this knowledge is exactly 
 what might be expected from those quick-witted pioneers. 
 Nevertheless, the exact date and locality of the introduction of 
 the copper plate remains a matter for conjecture. 
 
 The copper plates fixed in the battery in the early fifties were 
 about 4 inches wide and as long as the mortar, and were placed 
 one on the "feed" side and one on the discharge side just 
 
 * Voyages Melallurgiques, vol. ii., p. 309. Paris, 1780. 
 t Baron Inigo Bern's New Process of Amalgamation, p. 122, translated 
 by Raspe. London, 1791. 
 
 7 
 
98 
 
 THE METALLURGY OP GOLD. 
 
 underneath the screens. It was soon found that the plates 
 worked better from the start if they were coated with mercury 
 before they were placed in position, and this has now been 
 invariably done for the last forty years. Crushed ore, stones, 
 water, and amalgam are flung violently against the plates, and 
 the amalgam is retained in great part. 
 
 The scouring action of the pulp on the plates is, however, 
 always great, and becomes more violent in proportion as the 
 stamps are larger. These are now as much as 1,250 Ibs. in 
 weight in the Transvaal, an enormous mass when compared 
 with the 750-lb. stamp of a few years ago, and the 120-lb. stamp 
 of the last century. With this increase in weight it has become 
 desirable to modify the use of the copper plate. The plate on 
 the feed side, long ago condemned by many, is accordingly being 
 discarded more and more, and that below the screens is curved 
 away in such a manner that it cannot be struck directly by the 
 splash from the stamp (see Fig. 21, p. 109). The latest step in 
 this direction has been to line the mortar with cast-steel plates, 
 furnished with slots for the purpose of catching amalgam. (See 
 Fig. 45a). 
 
 After leaving the mortar, the pulp was treated forty years 
 ago, in districts other than Transylvania, mainly by passing it 
 over inclined tables covered with blankets, much in the same 
 way as Jason may have seen the golden sands worked in Asia 
 Minor. The sands accumulating on the blankets were washed 
 off at intervals and ground in mills with mercury. In addition, 
 the ore was frequently passed over or through baths of mercury, 
 in imitation of the old Tyrolean practice, still adhered to, with 
 modifications, in many Australian mills. 
 
 Amalgamated copper plates over which the pulp flowed were 
 tried, after those placed in the mortars had been proved to be 
 beneficial, but in Western America were at first almost every- 
 where rejected,* probably owing to the great depth of the 
 stream of ore and water made to flow over them. When long 
 afterwards, about the year 1870, this mistake began to be recti- 
 fied, the value of the plates was soon recognised in California. 
 
 The pulp is now led over the surface of these inclined plates 
 in a very thin stream, not more than a quarter of an inch deep. 
 The pulp does not run down in a regular stream, but in a series 
 of little wavelets which tumble over and over, and are supposed 
 to bring every part of the pulp successively in contact with the 
 amalgamated surface. The catching powers of the plates are 
 thus supposed to be practically independent of the tendency of 
 the particles of gold or amalgam to sink to the bottom of the 
 stream. This theory is not accepted by the Austrian school (see 
 
 * See Nevada and California Processes of Silver and Gold Extraction, 
 p. 61, by G. Kustel. San Francisco, 1863. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 99 
 
 above, p. 96), and it is certain that native gold is caught more 
 easily in proportion as it contains less silver (and is of higher 
 density), so that when the particles of metal consist of an alloy 
 containing a large proportion of silver, and are, therefore, of 
 comparatively low density, the yield on the plates is generally 
 poor. In any case, however, whatever be the working hypo- 
 thesis adopted, the amalgamated plate should theoretically be 
 better adapted for its work than the Tyrolean mill and other 
 machines using mercury baths, owing to the slight depth of the 
 pulp on the plates, and the short distance through which the 
 gold particles are compelled to settle before reaching a catching 
 surface. The plates are wiped down with rubber or brashes 
 about once a day, and the gold separated in the usual way from 
 the excess amalgam thus collected. 
 
 The German stamp has a rectangular stem made of wood or, 
 latterly, of iron, with an iron head, the total weight never 
 exceeding 300 to 400 Ibs. It was introduced almost unchanged 
 into France, Cornwall, and, after the discovery of gold in 1849, 
 into the United States, but has given place in all new districts 
 to the Californian stamp, and need not be fully described here. 
 In 1850, the first stamp-mill seen in California was erected at 
 Boston Ravine, Grass Valley. The stamps consisted of tree- 
 trunks shod with iron, and the framework was constructed 
 of logs. 
 
 The ordinary method of reduction and amalgamation of gold 
 quartz in a stamp battery now consists of the following opera- 
 tions : 
 
 1. The ore is broken down to a moderate size, varying from 
 that of a man's fist to a nut, by passing through the jaws of a 
 stone -breaker, or by hand hammers. 
 
 2. The ore is then fed into the mortar-box of a stamp mill, 
 where it is pulverised to any required degree of fineness. In 
 wet crushing, a stream of water is introduced also, and the blows 
 of the stamps splash the water and pulp against screens set in 
 the side of the mortar, the finely-divided ore being ejected in 
 this way. In some cases the mortar-box is partly lined with 
 amalgamated copper plates, by which some of the gold is caught 
 and retained, mercury being in this case usually fed into the 
 mortar-box with the ore and water. 
 
 3. On issuing from the battery, the pulp is allowed to run 
 slowly over a series of inclined, amalgamated, copper plates, by 
 which a further percentage of the gold is amalgamated and 
 retained. 
 
 4. The tailings are sometimes further treated by running 
 over rough hides or blankets, by which some particles of gold 
 and pyrites are retained, or the pyrites are separated from the 
 valueless sands by concentration on some form of vanner or jig. 
 These concentrates are subjected to further treatment, usually 
 either by smelting or by chlorination. 
 
100 THE METALLURGY OF GOLD. 
 
 5. At intervals the gold amalgam is wiped off the copper 
 plates, the excess of mercury separated by squeezing through 
 filter bags of chamois leather, buckskin, or canvas, and the solid 
 amalgam thus obtained is retorted so as to distil off the mercury, 
 and the gold is then melted. 
 
 In the following pages much information has been derived 
 from (1) "The Milling of Gold Ores in California," by J. H. 
 Hammond, published in the Eighth Report of the Californian 
 State Mineralogist, 1888; (2) the series of articles on "The 
 Variations in the Milling of Gold Ores," by T. A. Rickard, 
 published in the Engineering and Mining Journal in the years 
 1892 to 1895; and (3) " The Amalgamation of Free-Milling Gold 
 Ores," by Louis Janin, Jun., published in the Mineral Industry 
 for 1894. 
 
 The stamp battery must be regarded from two different points 
 of view viz., (a) as a crushing machine, (b) as an amalgamating 
 machine, and it should be remembered that the modifications- 
 designed to make it a more efficient crusher often reduce its 
 power as an amalgamator, and vice versa. Rickard has pointed 
 out that two typical methods are the Californian practice, for 
 the treatment of "free-milling" gold ores, and the Colorado 
 practice, for the treatment of ores, sometimes called "refractory,"' 
 which, however, yield most of their gold when carefully treated 
 in the stamp battery. The word " refractory " is better reserved 
 for those ores which cannot be satisfactorily treated by direct- 
 amalgamation, whatever be the method adopted. The stamp 
 battery is also used purely as a crushing machine in preparing 
 certain silver ores (whether gold is present in addition or not) 
 for treatment by other processes (e.^.,theWashoe, Reese River, and 
 lixiviation processes), which properly appertain to the metallurgy 
 of silver and will not be described in this volume, although a. 
 short account of pan-amalgamation is given on pp. 167-173. 
 
 The Californian practice consists briefly in crushing the ore and 
 effecting its discharge from the battery as rapidly as possible. 
 With this object in view heavy stamps are used, running very 
 fast, with a small drop ; the screen area is large and placed as 
 low down as possible, and the mortar is made narrow, with 
 nearly vertical sides. These arrangements all increase the out- 
 put of the battery. There are usually amalgamated copper plates 
 inside the mortar on the discharge side only. The method is 
 suitable for ores containing coarse, free gold which is easily 
 amalgamated, and which is caught largely on the inside plates, 
 in spite of the short time during which the ore remains in the 
 mortar. A small amount (1 to 5 per cent.) of pyrites, especially 
 of clean iron pyrites, does not interfere with rapid amalgamation; 
 but if the percentage of this mineral is high (10 to 20 per cent.), 
 and especially if the easily decomposable variety of sulphide of 
 iron (marcasite), or some sulphides of lead and zinc, or if com- 
 pounds of arsenic or antimony or other base minerals are present, 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 101 
 
 the amalgamation is greatly retarded or prevented, and the 
 Colorado practice is resorted to. If the gold is fine, or if, for any 
 other reason whatever, amalgamation is difficult, the Californian 
 practice must be modified in the same direction. 
 
 The Colorado practice was first devised in Gilpin County, 
 Colorado. In the early days of this gold-field, about the year 
 1860, after the arrastra had been discarded as being too slow, 
 the fast-drop and shallow-discharge batteries, like those used in 
 California, were introduced, and gave good results in working on 
 the oxidised surface-quartz, 60 per cent, to 75 per cent, of the 
 gold being extracted. As the mine workings got lower, however, 
 the percentage of pyrites steadily increased, and the mills gave 
 poorer and poorer results, until a return of only 30 per cent, to 
 40 per cent, was obtained. A return to efficient working was 
 only made after a long series of costly experiments, which 
 resulted in the present slow-working, long-drop stamps with a 
 wide, roomy mortar and a very deep discharge, the lowest part 
 of the screen being more than a foot above the dies, instead of 
 only 6 inches as in California. Amalgamated plates were put 
 inside the mortar on both feed and discharge sides. The object 
 of these arrangements was to keep the ore in the mortar for a 
 long time, so as to increase the chance of catching the gold 
 on the inside plates. The duty of the stamps was of course 
 greatly diminished, the output of a typical Colorado battery 
 being only about 1 ton per stamp per day of twenty-four hours, 
 while with Californian practice it is from 2 to 5 tons. Never- 
 theless the ultimate object is equally well attained by both 
 types of battery viz., the extraction of from 75 per cent, to 95 
 per cent, of the gold present, including that saved in the concen- 
 trates. 
 
 Ores containing more than 20 per cent, of sulphides, unless 
 they are clean, cubical iron pyrites, usually come under the 
 heading of true refractory ores, and cannot be treated by any kind 
 of stamp-battery amalgamation. Their treatment is considered 
 in succeeding chapters. The following is a general description 
 of the machinery employed in stamp-battery practice. In the 
 sequel the modifications adopted in different districts will be 
 discussed, and the conditions for successful amalgamation of 
 different classes of ore referred to. 
 
 1. Rock -breakers. There are two classes of these machines 
 in general use, viz.: (a) Those constructed on the reciprocating- 
 jaw principle, and (b) gyratory crushers. The Blake and the 
 Dodge crushers are representative of the former class, and the 
 Gates crusher of the latter. 
 
 The Make Crusher is showli in section in Fig. 14. The rock 
 is crushed between the stationary jaw, BC 1 , and the swinging 
 jaw, D, which is pivoted at E, and moved by the eccentric, F, 
 through the toggles, J K. The swinging jaw, in order to be 
 
102 
 
 THE METALLURGY OF GOLD. 
 
 as light as possible, should be made of steel, cast hollow and 
 braced by ribs, but is usually composed of cast iron. The 
 machine works at about 250 revolutions per minute. At each 
 revolution the moving jaw is advanced about |-inch towards the 
 other, and the lumps of rock which have dropped down between 
 the jaws are broken ; as the moving jaw recedes, the fragments 
 slip lower down and are further crushed at the next advance, 
 and this process is repeated until the ore is small enough to- 
 pass out at the opening at the bottom. The distance between 
 the jaws at the bottom limits the size of the fragments, and 
 this distance may be regulated at will by moving the wedge, 
 L, or by changing the length of the toggles, J K. The capacity 
 of the machine is great, being about 300 tons of ordinary rock 
 per day of twenty-four hours in the case of the machine whose- 
 
 Fig. 14. 
 
 dimensions at the mouth are 20 inches by 10 inches, when th& 
 lower edge of the jaws are set to approach within 1 inches of each 
 other. The power required for this is stated to be 14 H.P. 
 
 Many modifications have been made in this machine by 
 various makers since the patent expired, and of these the Blake- 
 Marsden machine has come into more extended use than any 
 other reciprocating-jaw crusher. The most important alteration 
 is the pivoting of the moving jaw below instead of above, which 
 is adopted in the Dodge crusher shown in Fig. 15. The effect of 
 this arrangement is to make the product more uniform in size, 
 and as there is little or no motion of the movable jaw at the 
 delivery aperture, this may be made as narrow as desired, sa 
 that a finer product can be obtained, although it is at the ex- 
 pense of capacity. The Dodge crusher is more particularly 
 recommended for fine crushing in concentration works, or where- 
 the product is to be subsequently passed through rolls. A pre- 
 liminary treatment of the ore in a Blake crusher is desirable,, 
 where fine crushing by a Dodge is resorted to. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 103 
 
 Authorities differ as to the relative advantages of the two 
 positions of the pivot. A. Sahlin decides * in favour of placing 
 the pivot below, assigning as a reason that more work is neces- 
 sary to crush comparatively fine material than to break down 
 large pieces of rock in the upper part, where the points of 
 contact between the crushing jaws and the rock are few. The 
 shortest stroke and great leverage should, therefore, he con- 
 siders, be at the bottom, where the work is heaviest. On the 
 other hand, W. P. Blake dissents from all these views. 
 
 Fig. 15. 
 
 Gates Crusher. This machine has been in use for fifteen years 
 in crushing macadam, ballast, and iron ore, chiefly in the United 
 States, but has not long been applied to crush gold ores. It is 
 now largely used, however, both in America and in South Africa, 
 being probably the most economical rock-breaker where large 
 quantities of ore are being treated. It is shown in Fig. 16, and 
 consists of a vertical shaft of forged steel, G, rotated at the 
 bottom by a bevelled wheel, L, placed J inch out of centre. At 
 the top of the shaft is a chilled-iron breaking head, F, and the 
 shell surrounding this is lined with twelve chilled-iron, concave 
 pieces, E. These form the crushing faces. The shaft, G, has a 
 gyratory motion imparted to it by the eccentric box, D, attached 
 to L, and the rock is thus crushed, without grinding, between 
 the head and liners. The distance between the crushing surfaces 
 at the bottom may be regulated by set-screws. With dry ore 
 this distance may be as low as || inch, no pieces larger than this 
 being allowed to pass. It is stated that this machine works 
 with less expenditure of power than the Blake crusher, and that 
 its product is more uniform and can be made finer. Its first 
 cost, however, is higher, and, what is of moi-e importance, the 
 
 * Trans. Am. Inst. Mng. Eng., 1892. 
 
104 
 
 THE METALLURGY OF GOLD. 
 
 repairs are troublesome and expensive, but it certainly works 
 well in many places where it has been adopted. 
 
 . 16. 
 
 The Schranz Stone-breaker. This machine (Fig. 17) is 
 of interest, as it forms a link between the ordinary stone -breaker, 
 with reciprocating jaws, and the crushing rolls. The movable jaw, 
 instead of having a reciprocating motion, has a rocking one, some- 
 what similar to the motion of a circular blotting-pad. The jaws 
 are of cast steel, and the space between them is regulated by 
 means of the vertical screw, I, which adjusts the wedge, c. The 
 short connecting-rod, b, made of cast iron is made with a narrow 
 section, so that it is broken before any other part of the stone- 
 breaker, if some very hard material should find its way between 
 the jaws. The machine has the advantage, in common with the 
 Dodge crusher, that the opening at the bottom does not vary, so 
 that a uniform-sized product is secured, the maximum diameter 
 of which can readily be reduced to 8 mm. (0'3 inch). The 
 machine can thus take the place of roughing-rolls. The large 
 sized machine, working at 250 revolutions per minute, and with 
 an expenditure of 10 to 12 horse-power, is stated to crush from 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 105 
 
 4 to 5 tons of rock per hour to the size given above, and thus 
 compares favourably with the Dodge crusher. It is in use at 
 Laurenburg, on the Lahn, and at other places in Germany, 
 and gives great satisfaction. 
 
 Material Employed for the Crushing Surfaces. The 
 selection of the most suitable material for the working parts, 
 and especially for the crushing surfaces, of reduction machinery 
 is a matter of the greatest possible importance, as the economy 
 effected by using a durable material, which seldom requires 
 renewal, is very great. When rock is crushed by repeated blows, 
 as in the case of ordinary rock-breakers, stamps, and, perhaps, 
 
 Fig. 17. 
 
 rolls, it does not follow that the hardest material is 
 the most durable. A substance is wanted that will not be 
 broken or caused to crystallise unduly by a blow, but which, 
 at the same time, will show neither signs of deformation 
 nor rapid attrition by the impact of fragments of quartz. 
 Both chilled iron and hard chromium-steel have been proved to 
 be very useful for the surfaces of impact machines, but the softer 
 and tougher open-hearth mild steel has also been found advan- 
 tageous for the surfaces of rolls. The quality of toughness seems 
 on the whole to be of even more importance than mere hardness. 
 
106 THE METALLURGY OF GOLD. 
 
 No doubt different qualities are required for the crushing surfaces 
 in different machines, and even in the same machine when 
 operating on different classes of ore. Thus chromium - steel, 
 which has been found so useful for the manufacture of stamp 
 heads and dies, was not a success when used . in the Huntington 
 mill, where the action is rather one of grinding than of impact. 
 
 Perhaps Hadfield's manganese steel is the best material for the 
 jaws of stone-breakers. At the Penmaenmawr granite quarries,, 
 the plates inserted at the sides of the jaws to protect the outer 
 casing of a Blake crusher were worn out in two months, when 
 ordinary steel was used, while similar plates made of cast 
 manganese steel, after being in service for more than twelve 
 months, had only been worn to the extent of J inch, and were 
 still at work. In stamp-mills, the crushing faces, if made of chilled 
 iron, last for six or eight months, while steel faces last twice 
 as long. The whole question of the nature of the materials to be 
 used for different purposes is one for the metallurgist, and too- 
 much attention can hardly be paid to it in the future. 
 
 Position and Use of Bock-Breakers. The aperture of a 
 rock-breaker should be placed on a level with the floor, so that 
 the ore can be dumped down by the side and shovelled into the 
 jaws. As noted below (p. 128), it is now becoming customary 
 to place the rock-breaker in a separate building distinct from 
 the battery house. 
 
 A grizzly is employed to separate the fine material, which is 
 passed straight to the stamps, while in some cases the material 
 from the rock-breaker is also screened and part returned to be put 
 through again. Revolving trommels and flat shaking screens 
 have both been used for this purpose. 
 
 The efficiency and economy in crushing, attained by the 
 machines on the reciprocating-jaw principle, are so fully recog- 
 nised that there has been a tendency of late years to use them 
 to reduce gold-quartz to a very small size before feeding it into 
 the stamp batteries, the capacity of which is greatly increased in 
 this way. For fine crushing, multiple-jaw crushers, on the 
 same principle as the Blake, have been constructed, but have 
 not passed into general use ; the use of a pair of rolls between 
 the rock-breaker and the stamp battery has also been advocated. 
 This is done by the Huanchaca Mining Company at Antofagasta, 
 Chili, and gives an increase of capacity to the stamps of over 
 20 per cent., while the cost is trifling. The size of the ore 
 best suited for feeding into a stamp battery may be roughly 
 put down as about J inch for light stamps and \ inch for heavy 
 stamps. If the size of the material is much smaller than this^ 
 no advantage in speed is gained, while the jar given to the stamps 
 and framework is greatly increased. At the present day few 
 large mills are erected without rock-breakers, which have also- 
 been successfully added to many old mills. Nevertheless, they 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 107 
 
 are absent or rarely seen in several of the richest gold-fields of 
 the world. In Gilpin County, Colorado, they are not used ; in 
 Victoria the extent of their use may be judged from the fact 
 that there were 5,901 stamp-heads in operation in 1891, and 
 only twelve stone-breakers, and in other parts of Australia they 
 have been similarly neglected. 
 
 The Stamp Battery. Californian " gravitation " stamps are 
 
108 THE METALLURGY OF GOLD. 
 
 in general use at the present day for crushing gold ores. A stamp 
 is a heavy iron, or iron and steel, pestle, raised by a cam keyed 
 on to a horizontal revolving shaft, and let fall by its own weight. 
 Stamps are ranged in line in groups of five stamps each, which 
 have a mortar-box in common. Fig. 18 represents the side view, 
 And Fig. 19 the front view, of a ten-stamp battery, with the 
 .amalgamating tables removed to show the foundation timbers or 
 mortar-blocks, A. 
 
 The foundations are of the highest importance, as, if they are 
 badly made through carelessness or false economy, the efficiency 
 of the battery is greatly decreased, and it soon shakes itself to 
 pieces. The blow of a stamp is partly employed in crushing the 
 ore, and is partly expended in producing a concussion or jar 
 acting on the framework and foundations. The amount of energy 
 used up in the latter way depends largely on construction, for 
 details of which the student is referred to the volume on Metal- 
 lurgical Machinery. In preparing the ground for the foundations, 
 the earth is removed until bed-rock is reached if possible, 
 and the latter is then carefully smoothed and sometimes covered 
 with a layer of cement. The wooden mortar-blocks of from 6 to 
 14 feet long are placed upright in this trench, and the space round 
 filled up with sand, or, as in the Transvaal, solid masonry is 
 built round the blocks. The framework is now usually made of 
 wood, which is a far more satisfactory material for the purpose 
 than iron. It consists of the massive battery sills, B, on which 
 rest the battery posts, C, and the braces, E. The posts are held 
 together by the stamp-guides or tie-timbers, D. Wooden braces 
 have completely replaced iron rods, which allow the battery to 
 spring. It is better to place the braces on the discharge side 
 alone, thus leaving more room to work on the feed side. 
 
 Tlie Mortar. The mortars are made of cast iron, but differ in 
 shape according to the nature of the ore and the corresponding 
 modifications made in the course of treatment. They weigh 
 from 1 to 3 tons, being especially thick at the bottom where 
 there is the greatest strain. An ordinary mortar is about 4 feet 
 7 inches long, 50 inches high, and 12 inches wide on the inside 
 at the level at which the dies are set. The bottom is from 3 to 
 f* inches in thickness, and has a heavy flange cast on it, by which 
 it is bolted to the mortar-blocks. These are tarred over, all 
 cracks in them having been filled with sulphur, and are then 
 covered with three thicknesses of blanket, carefully coated with 
 tar on both sides. The mortar is placed on these blankets and 
 securely bolted down. This arrangement lessens the chance of 
 the mortar working loose, the jar being diminished. A sheet of 
 rubber, J inch thick, may be used instead of the blankets. Figs. 
 20 and 21 represent sectional elevations of the two chief types 
 of mortars, Fig. 20 showing a mortar intended to be supplied 
 with an inside amalgamated copper lining plate, , on the screen 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 109* 
 
 side only, and Fig. 21 a mortar designed to have copper plates, e, e, 
 placed both at the front and back. In both figures, b is the feed- 
 opening through which the ore is introduced into the mortar ; 
 c is the bed on which the die is placed ; d is the screen-opening. 
 The chief difference between them is in the feeding arrangement; 
 in the latter case the back plate is put in a recess, and is pro- 
 tected from the falling rock fed into the battery. As stamps 
 continue to increase in weight, and the scouring effect of the 
 pulp becomes in consequence more marked, the tendency grows to 
 discard the type shown in Fig. 21. The plates catch the coarse 
 gold inside the mortar when the pulp is flung against them. 
 Occasionally copper plates are also put at the ends of the 
 mortar. All the plates, which are bolted through the mortar 
 itself or to its lining plates, must be so arranged that they can. 
 
 Fig. 20. 
 
 Fig. 21. 
 
 easily be taken out and cleaned, as, when very rich ores are 
 being treated, precious metal accumulates on them very fast. 
 
 The width of the mortar varies from 10 to 14 inches at the 
 level of the bottom of the screens. As has been already men- 
 tioned, narrow mortars are best fitted for rapid discharge, but, if 
 hard flinty ores are to be crushed, a narrow mortar causes 
 frequent breakage of the screens, unless the discharge is deep 
 i.e., unless the bottom of the screens is a considerable distance 
 above the surface of the dies. By this latter arrangement the out- 
 put is reduced, since, the nearer the screens are to the dies, the 
 more rapid is the discharge. The depth of the discharge is only 
 about 6 or 7 inches in California, where an adjustable battery- 
 screen has been introduced to keep this depth constant, in spite 
 of the wearing of the dies. In the battery thus modified, the 
 screen-frame is supported on a wooden block, which is easily 
 removable, and to which the copper plate is bolted. When the 
 dies wear down, this block is replaced by one of less height, to- 
 which a suitable plate has already been fixed. 
 
 Mortars often have a lining of cast-iron plates, bolted to them, 
 
110 
 
 THE METALLURGY OF GOLD. 
 
 near the bottom, to protect them from the rapid wear due to the 
 splashing of the pulp. These plates last from six to nine months, 
 -and can be replaced when worn out. 
 
 The splaslirbox, not shown in the figures, and now often 
 omitted, is bolted to the outside of the mortar just below the 
 .screens. It is rectangular, consists of wood or iron, and is of the 
 same length as the mortar. It receives the pulp as it passes 
 through the screens, and distributes it evenly over the amalga- 
 mating tables by a number of spouts, usually three. Instead of 
 the splash-box, a splash-board is now almost universally employed, 
 the usual material for it being heavy canvas. The old form of 
 mortar had its upper part, or housing, of wood (see Fig. 44, 
 p. 204), but, as mercury is lost through the smallest aperture, 
 .and it was difficult to make these wooden housings quite tight, 
 mortars are now cast in one piece, including the housings. The 
 roof of the mortar is made of 2-inch planking, through which holes 
 are cut to admit the stems of the stamps and the water pipes. 
 
 When the mortar is in place, the dies are put into it, a layer of 
 sand being often introduced first. The dies consist of two parts, 
 the footplate and the die proper, or boss. 
 Fig. 22 shows, in plan and elevation, 
 one of the many forms of dies in use ; 
 here the footplate is almost square, so 
 as to fit the mortar; it is 1 or 2 inches 
 thick, and 10 or 12 inches square. The 
 boss is cylindrical, from 3 to 6 inches 
 high, and of the same diameter as the 
 shoe. Shoes and dies are made either of 
 iron or steel, hard chilled-iron being used 
 for wet crushing, and soft iron for dry 
 crushing. Cast-steel dies and shoes have 
 been often tried, but owing to their 
 tendency to chip and to their uneven 
 wear, their introduction formerly met 
 with little success. Nevertheless, in 
 most mills remote from foundries, where 
 transportation is an important item in 
 ^g cos o f SU ppii eS) steel shoes and dies 
 ,have now replaced those of iron, as the life of steel is from two 
 and a-half to three times that of iron, and the cost only about 
 twice as great. * About seven years ago, chrome-steel shoes and 
 dies were introduced into California, and proved their superiority 
 over those manufactured from most other kinds of steel ; forged 
 steel and, according to Janin, manganese steel have also been 
 used with success. Sometimes steel shoes and iron dies are used, 
 the wear being more even than when both consist of steel. At 
 .the Patterson mine in 1888, a set of iron shoes and dies lasted 
 
 * Eighth Report Gal. State Min., 1888, p. 708. 
 
 A 
 
 
 f 
 
 1 
 
 
 
 i 
 
 
 
 t > 
 
 
 
 
 
 IY 
 
 lg> * 
 Scale = T 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. Ill 
 
 on an average six weeks, crushing 1,680 tons of ore, while one 
 set of chrome-steel shoes is said to have lasted fourteen months, 
 crushing 16,800 tons, or ten times the amount crushed by iron.* 
 The wear of iron shoes and dies is stated to be about 2 or 3 Ibs. 
 per ton of ore crushed in California. In Colorado, at the New 
 California Mine, the wear of shoes was 11-3 ozs. of iron per ton 
 of ore, and that of the dies 4'5 ozs. At the Robinson Mine, 
 South Africa, the wear of steel shoes and dies is, according to 
 Mr. Harland, 0'45 Ib. per ton crushed for shoes and 0*30 Ib. per 
 ton for dies. When the boss is worn down to within from J inch 
 to 1 inch of the footplate, the die is replaced. Dies wear more 
 slowly than shoes since they are protected by a layer of pulp, 
 which is from 1J to 3 inches thick. The dies should all be 
 renewed together, as it is important that those in the same 
 battery should be of equal height, otherwise one or more will 
 become almost bare of ore, and a disastrous pounding result. 
 If a die breaks, it should not be replaced by a new one, but by 
 one worn to the same extent as the others in the battery. Iron 
 false-bottoms or chuck-blocks are placed beneath partially-worn 
 dies, so as to keep the depth of discharge constant. 
 
 The cam-shaft, H, Fig. 19, is of wrought iron, and about 
 5 inches in diameter (A, Fig. 23). It is now usual to have a 
 separate cam-shaft for each five or ten stamps, which have thus a 
 separate driving wheel. The advantage of this arrangement is 
 that repairs can be done to one or more stamps without necessi- 
 tating the stoppage of the whole mill, as used to be the case when 
 there was only one cam-shaft. The cam-shaft is placed at a 
 'distance of from 5 to 10 inches from the stem-centre, and is 9 to 
 10 feet above the mortar bed. The bearings rest on supports 
 attached to the battery posts on the discharge side. 
 
 The cams are made of cast iron, with chilled faces, which 
 are 2 to 3 inches wide (B, Fig. 23). The double cam, two views 
 of which are shown in Fig. 23, is now in almost universal use, 
 though single and treble forms have been employed. Sometimes 
 cams are cast in two pieces which are bolted together, so that 
 when one is worn out, it can be replaced without first removing 
 the other cams on the shaft. It is stated, however, that these 
 sectional cams work loose, and are not much used in conse- 
 quence. The hub is always strengthened by a band of wrought 
 iron shrunk on it. The shape of the cam face is the involute 
 of a circle slightly modified at the end so as to stop the upward 
 motion gradually. The radius of this circle is equal to the 
 distance between the centres of the cam-shaft and stem, which 
 depends on the height to which the stamp is to be lifted, so that 
 the curve of the cam varies with the drop. A cam should last 
 .several years unless broken through being a faulty casting, or 
 .through carelessness in letting the stamp fall when hung up. 
 
 The cam-face works against the iron collar or tappet, shown in 
 * Eighth Report Col. State Min., 1888, p. 656. 
 
112 
 
 THE METALLURGY OF GOLD. 
 
 plan and section in Fig. 24, which is bored out at A to fit the 
 stem of the stamp. The tappet is fitted with a wrought- 
 iron gib, which is pressed against the stem by two or three 
 keys behind it, thus binding the tappet firmly on the stem 
 while, at the same time, admitting of rapid adjustment 
 to another position. The entire end surface of the tappet 
 comes in contact with the cam-face, by which the stamp is 
 raised and, at the same time, rotated, all its parts being round. 
 The effect of this is, that the shoe does not strike the ore in the 
 mortar in exactly the same place twice in succession, and the 
 wear of its face is made more uniform. The greater part of the 
 revolution takes place during the raising of the stamp, but the 
 latter does not quite cease to rotate as it falls, and a slight grind- 
 
 u 
 
 J 
 
 ! 
 
 c 
 
 
 1 i 
 
 V 
 
 AKeywatf 
 
 _! __.-' 
 
 \ 
 
 V 
 
 
 Fig. 24. 
 
 Fig. 23. 
 
 ing action on the ore has been noticed by many observers. The 
 amount of rotation varies with the fall, the extent to which the 
 cam and tappet are greased, and the state of wear of their sur- 
 faces. A little grease is always added to reduce wear, but, if too 
 much is present, the stamp does not revolve at all, while, accord- 
 ing to J. H. Hammond, when the tappet is in the right condi- 
 tion, one revolution is effected in from four to eight blows, with 
 a 6- or 8-inch drop. Other observers find the usual rate of 
 rotation more rapid, and in Gilpin Co., Colorado, where the 
 average drop is from 16 to 18 inches, the stamp makes from 1 J 
 to 1 J revolutions at each blow. Tappets should last for four or 
 five years ; and, having both ends alike, they can be reversed 
 when one end is worn out, and their worn and grooved faces 
 can be planed down when necessary. Some millmen assert that 
 tappets may be broken by the cam if keyed too tightly to the 
 stem. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 113 
 
 As the cam-thrust is not applied at the centre of the stamp, 
 there is always a considerable side pressure, which greatly in- 
 creases the friction in the guides and wears the latter out, 
 besides causing a loss of power. Moreover, another result of 
 this is that the stamp tends to be inclined (not vertical) when it 
 is released, and so the blow on the die is given slightly to one 
 side i.e., the side of the die on which the cam works. Conse- 
 quently, there is a tendency for this side to wear down more 
 quickly than the other. To obviate these disadvantages, cams 
 have been introduced at Johannesburg with a wide hub, and 
 the two blades set one at each end of the hub, so that they work 
 on opposite sides of the stamp and cause it to revolve in different 
 directions at each successive uplift. The Blanton cam, now 
 generally used on the Rand, is fastened to the shaft by a semi- 
 circular wedge pinned to the cam-shaft, no keys being necessary. 
 
 The pulley on the cam-shaft (F, Fig. 19) is made of wood on 
 cast-iron flanges ; if iron alone were used, the rapid succession 
 of jars, caused by the dropping of the stamps, would soon cause 
 the material to crystallise and break. A tightener pulley on the 
 belt driving the cam-shaft is often used, by which the stamps 
 can be put in motion or stopped without interfering with the 
 driving power. 
 
 The stamp itself consists of three parts, the stem, the head, 
 and the shoe. The stem (G, Fig. 19) is from 12 to 15 feet 
 
 Fig. 26. 
 
 long, and from 3 to 3J inches in diameter; it is made of wrought 
 iron, and has both ends tapered for a length of 6 or 8 inches to 
 fit the heads, so that, if one end is broken off, the stem can 
 be inverted and the other end used. The head (Fig. 25) and 
 shoe (Fig. 26) are made of equal diameter viz., about 8 to 10 
 inches. The head is of tough cast-iron, about 15 to 20 inches 
 high, and has a tapered socket at each end, the upper one (A B) 
 for the stem and the lower (E F) for the tapered shank of the 
 shoe. When these are driven into their respective sockets, into 
 
 8 
 
114 THE METALLURGY OF GOLD. 
 
 which a few strips of wood are inserted to keep the two metal 
 surfaces from touching each other, a few blows by the stamp 
 bind them securely together, no other fastening being necessary. 
 Slots are provided (shown in the figure), at the base of the two 
 sockets, through which wedges may be driven to force out the 
 shoe or stem when necessary. The head is often strengthened 
 by bands of wrought iron, shrunk on at each end, to prevent 
 splitting by the wedge-like action of the tapering stem and shoe. 
 The head lasts several years, being rarely ruptured. The shoe 
 (Fig. 26, which is on a larger scale than Fig. 25) consists of two 
 parts, the shank, which fits into the head, and the shoe proper 
 or butt. The latter is made of very hard white iron, and the 
 shank of softer iron; steel is also used, as has been already 
 mentioned. The diameter of the shank is about half that of the 
 butt. The shoe is replaced when the butt, which is from 6 to 
 12 inches in length when new, has been worn down to about 
 1 inch in length. To keep the total weight of the stamp con- 
 stant, several sizes of heads are sometimes used in one mill, 
 the heavier heads taking partly-worn shoes. "Chuck-shoes" 
 are inserted between heads and shoes with the same object. 
 
 The relative weights of tappet, stem, head and shoe, which 
 together make up the stamp, vary considerably. There is an 
 advantage in increasing the weight of the stem, as one of small 
 diameter tends to spring and bend from the blow of the cam, 
 or when the stamp falls, and to wear the guides rapidly. The 
 stem weighs from 250 to 475 Ibs., the tappet from 80 to 130 Ibs., 
 the head from 175 to 370 Ibs., and the shoe from 100 to 230 Ibs. 
 The total weight of the stamp is usually from 650 to 1,150 Ibs., 
 but is sometimes as low as 450 Ibs., and, for prospecting purposes, 
 the weight is only from 100 to 300 Ibs. Old dies weigh from 
 20 to 50 Ibs. when they are discarded, and old shoes from 25 to 
 40 Ibs. A steel tappet on a 900-lb. stamp weighs 112 Ibs. 
 
 The stamp stems are guided in boxes bolted to the wooden 
 cross-ties, which also serve to hold the battery posts together. 
 There are two of these guides (D, Fig. 19), one within a foot or 
 so of the top of the battery posts, and the other as low as the 
 raising of the stamp head will allow. The depth of each guide 
 is about 15 inches, and the stems are fitted closely to the guides, 
 metal boxes being used occasionally, although wood is much more 
 general. The guide-boards are sometimes pierced with large 
 square holes in which bushes of wood, with the grain parallel to 
 the length of the stamp, are placed fitting the stem exactly. In 
 this way, the guide-boards themselves are preserved from wearing 
 out. Sectional guides, consisting of a series of iron keys enclosing 
 wooden bushings, are also used. In this case each stem has a guide 
 to itself, and the bushings can be renewed by hanging-up the one 
 stamp without stopping the other stamps in the battery. 
 
 Each stamp is provided with a lever or jack (I. Fig. 18) made 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 115 
 
 of wrought iron, or of wood protected by iron. The jack is for the 
 purpose of raising the stamp and hanging it up out of reach of 
 the cam. When this is to be done, a strip of wood, an inch or 
 more thick, is laid with one hand on the cam as it rises, and the 
 stamp is thus raised an inch higher than usual, so that the jack 
 can be slipped in under the tappet with the other hand. The 
 stamp is thus suspended above the cam and can be repaired 
 without stopping the others, while it can only be released in a 
 manner similar to that in which it was hung up. Above the 
 stamps there is a double rail, on which is a movable pulley; by 
 this the stamps, <fec., can be lifted up for repairs. When the 
 stamps are to be set up, the head is put on the die and the stem 
 dropped into it, iron being left sometimes on iron, but more 
 usually some canvas or other packing being put into the head 
 socket and the stem dropped into that. The stamp is then 
 raised and dropped into the shoe, the shank of which is often 
 surrounded by strips of wood for packing. As already stated, 
 the parts are soon wedged firmly together by raising and letting 
 fall the stamp a few times. 
 
 The height of the "drop" of the stamps varies from 4 to 18 
 inches, and the number of drops per minute varies from 30 to 
 over 100. These depend on one another to a great extent, 
 >aii increase in the height of the drop being necessarily accom- 
 panied by a diminution in the number of drops per minute. 
 With a drop of 8J inches, about 95 blows can be obtained, the 
 tappet then just having time to fall after leaving one face of the 
 cam, before the other begins to raise it. As, within certain 
 limits and under certain conditions, an increase of speed results 
 in an increase of yield of pulverised ore, efforts have been made 
 to raise the number of blows per minute. It has been proposed 
 to use two cam-shafts, one above and one below the gib-tappets, 
 a speed of 250 blows per minute being thus attained, according 
 to the statements of the manufacturer, but this device has 
 apparently not passed into use. The subject will be returned to 
 when the conditions for successful amalgamation are discussed. 
 
 D. B. Morison discusses the question of height of drop from 
 the point of view of speed of crushing in a valuable paper, which 
 should be consulted.* The time occupied in seconds in the fall of 
 the stamp, if all friction, including the resistance of the air, is left 
 2H 
 
 out of account, is A/-TT~ where H = height in feet, and G = 32-2. 
 
 For a fall of 8 inches this amounts to about 0'2 second, but 
 owing to the friction between the stamp and the guides and the 
 resistance due to the water, the actual time taken, according to 
 Morison's experiments,t is about 0'225 second (see Fig. 26a). 
 
 * " Gravitation Stamp Mills for Quartz Crushing," by D. B. Morison, 
 Proc. N. E.C.I. ofE. and S., 1896-7, Session xiii. 
 t Loc. cit. 
 
116 THE METALLURGY OF GOLD. 
 
 The velocity at the time of impact and the force of the blow are- 
 correspondingly diminished, the latter being 17 per cent, less than 
 it would be if the work were done without friction in vacuo, a 
 mean result of twenty-four experiments. The time required in 
 raising the stamp is somewhat greater, owing to the imperfect 
 method necessarily employed. The shape of the cam causes the 
 stamp to start upwards at a certain velocity immediately the cam 
 meets the tappet, and to maintain the same velocity until near 
 the end of the stroke. In a properly constructed cam, designed 
 to give a drop of 8 inches, the vertical component of the velocity 
 of the cam is, for ninety drops per minute, about 2-5 feet per 
 second. If the cam were suddenly removed from the tappet 
 when moving at this velocity, the stamp would continue to rise 
 against gravity for about 1^-inch neglecting friction. 
 
 The inherent defect of the cam is that it strikes the tappet a 
 tremendous blow in the effort to lift the stamp at the full 
 velocity from the start, and it is this blow which is the cause of 
 much of the intense noise and vibration felt in every stamp mill. 
 In spite of the power of the blow, which tries the cam-shaft 
 severely, the dotted curve in Fig. 26a shows that at first the 
 movement of the cam is checked a little, probably in part by 
 distortion of the cam, and that it rapidly recovers itself, and 
 by elastic distortion in the opposite direction lifts the stamp 
 faster than the normal rate, and these alterations succeed each 
 other during the whole lift, after which the cam presumably 
 recovers itself in the time that elapses before it strikes the next 
 blow. 
 
 Fig. 26a shows a curve traced out by a pencil attached to a 
 900-lb. Sandycroft stamp, the ordinates representing vertical 
 movement of the stamp and the abscissae being time in tenths of 
 a second. The figure was traced out during one of a number of 
 experiments made by Morison, who subsequently devised the 
 high-speed stamp described on p. 158. 
 
 Screens. The screens are set in iron frames, which now usually 
 slide in grooves cast in the mortar, and are keyed to it, but 
 were formerly fitted into recesses and bolted. They are made 
 either of wire-cloth, or of Russian sheet-iron or steel, in which 
 holes are punched. The sheet-iron is about -% inch thick and 
 weighs about 1 Ib. per square foot. The rough (" burred") side of 
 the punched plate is put on the inside, so that the holes widen 
 outwards, and are thus prevented from becoming clogged. The 
 holes are round, oval, or consist of long slots (from J to J inch 
 long) ranged parallel or inclined to each other. The shortest dia- 
 meter of these holes ranges from about ^ to -^ inch or more, 
 according to the nature of the ore and the method employed 
 in its treatment; the usual size is from about -^ to ^ inch. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 117 
 
118 
 
 THE METALLURGY OF GOLD. 
 
 The relative advantages of wire-cloth and sheet-iron are not yet 
 beyond dispute, and vary with the nature of the ore. Slots- 
 appear to be better suited for discharge than meshes, but, on 
 the other hand, there is a great loss of discharge area in the use 
 of punched iron. Thus a wire mesh screen, containing 18 holes 
 to the linear inch, has 324 holes to the square inch, while a 
 round-punched sheet-iron screen has only 140 holes of the same 
 size per square inch. At Otago, N.Z., where friable quartzose 
 ores are treated, both kinds are used, but the wire-cloth (No. 18) 
 lasts a little longer, and costs nearly 25 per cent, less than the 
 round-punched Russia iron. In Gilpin County, Colorado, the 
 " burr-slot " screen is used, having horizontal slots placed alter- 
 nately. The screens are inverted after a time, as the lower part 
 wears out soonest. 
 
 In California, both steel and brass wire, and slot- and needle- 
 punched tinned- and sheet-iron are used as screen materials, steel 
 wire screens, however, being seldom used, owing to their rapid cor- 
 rosion by rusting. The sheet-iron consists of the best soft Russia 
 iron and has a smooth glossy surface, and the slots and needle 
 holes are almost universally burred. The size is named according 
 to the number of the sewing-machine needle by which the holes 
 were punched, the sizes numbered 5, 6, 7, 8, and 9 being those 
 most used. The following table gives details of these screens : 
 
 No. of Needle. 
 
 Corresponding 
 Mesh. 
 
 Width of Slot. 
 
 Weight 
 per Sq. Foot. 
 
 5 
 
 20 
 
 0-029 in. 
 
 1-15 Ibs. 
 
 6 
 
 25 
 
 027 
 
 1-08 ,, 
 
 7 
 
 30 
 
 024 
 
 0-99 
 
 8 
 
 35 
 
 0-22 ,, 
 
 0-92 
 
 9 
 
 40 
 
 020 
 
 0-83 ,. 
 
 The burred horizontal or horizontal-alternate slots are those 
 most in use. Tinned-iron screens are made thinner and, there- 
 fore, have the advantage of allowing a more rapid discharge. 
 The brass-wire screens used vary from 16 to 60 mesh, 30 to 
 40 being the most common ; with their use the pulp discharged 
 is more uniform in size than if slots are employed. Although 
 sheet-iron is the material usually employed for screens, it is 
 often preferable to use copper, as pyritic ores, if kept for any 
 length of time after being mined, soon become oxidised and acidi- 
 fied, and the ferrous sulphate thus formed corrodes iron rapidly, 
 whilst the water used is often more or less acid if it comes from 
 mines. Copper is not attacked in the same way. T. A. Rickard 
 has shown * that the life of screens in the mills at Blackhawk r 
 Colorado, diminishes as the creek is descended, the water 
 becoming more and more impure. At the South Clunes United 
 * Eng. and Mng. Journ., Sept. 3, 1892. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 119 
 
 Mill, Victoria, iron punched gratings lasted less than a week, 
 but, on introducing copper plates containing 100 holes per square 
 inch, the life of the screen was increased to a month, and 
 275 tons of ore were passed through it. At Blackhawk, the iron 
 screens last while from 80 to 430 tons are passed through them, 
 according to the position of the mill. At Grass Valley, the 
 average is 200 tons, at Bendigo, 134 tons, and at Otago, N.Z., 
 only 40 tons. In California the brass-wire screens last from 
 10 to 14 days, corresponding to a passage of 120 to 140 tons, and 
 the Russia-plate lasts from 15 to 40 days, the average being 
 30 days, corresponding to the passage of about 330 tons. The 
 rate of wear of the screens depends greatly on their position, 
 being more rapid with a shallow than a deep discharge, and 
 more rapid in a narrow than a wide mortar. Pieces of iron 
 or wood in the ore may cause the screen to break, and these 
 should, consequently, be removed from the ore as far as possible. 
 The battery is hung up now arid then, so that a thorough inspec- 
 tion of all the screens may be made, and those that are broken 
 replaced. 
 
 The screens were formerly set vertical, and this system still 
 prevails generally in Australia. In the United States, however, 
 they are now placed at an angle which varies somewhat but is 
 never far from 10, and this has been found to facilitate discharge. 
 In wet stamping, screens are usually placed on one side of the 
 mortar only viz., that opposite the feeding side. In cases where 
 the discharge is required to be as rapid as possible, the screen 
 area is increased, and double discharge (front and back) mortars 
 have been made, but have not been used much, except for dry 
 crushing, and screens at the ends of the mortars are used at 
 Harriettville, Victoria (Blckard). The area of the screen is from 
 3 to 4 square feet per battery in California, the height from 
 the bottom to the top of the screen being from 8 to 10 inches ; 
 in Gilpin County, the screens are usually 8 inches high and 
 4J feet long. Opinions differ as to the necessary amount of 
 screen area to be used. By some experts it is contended that the 
 capacity of a battery is really limited, not by its crushing, but 
 by its discharging power. Thus, by a number of experiments 
 conducted some years ago at the Metacom Mill, California, 
 it was shown that when crushed pulp instead of the unbroken 
 ore was fed into the mortar, the rate of discharge was not 
 increased. This was taken as a convincing proof that the 
 discharge area is usually far too low. Nevertheless, double dis- 
 charge is hardly ever used for wet-crushing mills, the various 
 objections that are made to it being summarised as follows : 
 
 1. Inconvenience is caused in the arrangement of the copper 
 plates, both inside and outside the battery. 
 
 2. So great a quantity of battery water must be used, that the 
 pulp is too thin for efficient amalgamation on the plates. 
 
120 THE METALLURGY OF GOLD. 
 
 3. The ore does not stay long enough in the battery to be 
 effectively amalgamated. 
 
 When, however, battery amalgamation is not attempted, double 
 discharge might be advantageous, and, in particular, it is to be 
 recommended with heavily sulphuretted ores, containing brittle 
 pyrites. The object to be attained in this case is to leave these 
 sulphides unbroken as far as possible, so as to facilitate concen- 
 tration, since great losses from sliming will inevitably result if 
 they are subjected to repeated blows in the battery before being 
 discharged. 
 
 Order of Fall of Stamps. The order in which the stamps 
 drop is of considerable importance. If they were let fall 
 in succession from one end to the other of the mortar, the pulp 
 would be driven before them, so that the stamp which fell last 
 would have its die covered by too deep a cushion of ore, while 
 that at the other end would be almost bare. The result to be 
 obtained is to keep the ore equally distributed through the 
 mortar, so that each stamp shall do the same amount of crushing, 
 although it is inevitable that the middle stamps should be more 
 efficient than the end ones in discharging the ore. The order 
 most favoured in California is 1, 4, 2, 5, 3, and that on the Rand 
 is 1, 3, 5, 2, 4, whilst the orders 1, 5, 2, 4, 3 and 1, 5, 3, 2, 4 are 
 also often used. Several other orders have their advocates, 
 and are probably little inferior to the above for the particular 
 ores on which they are employed. Since the end stamps are of 
 less efficiency than the others, it has been argued that a larger 
 number of stamps in one mortar would be advantageous, and at 
 Clausthal, in the Hartz Mountains, there are usually from nine 
 to eleven stamps in a battery,* placed close together, space being 
 greatly economised in this way. Long and wide experience has, 
 however, proved that the best number is five. 
 
 Feeding. Ore is fed into the battery either by hand or by 
 automatic machines. It is often asserted that really intelligent 
 hand-feeding is better than the automatic method, since the stamps 
 are not all equally efficient. At any rate, hand-feeding is still 
 persisted in on some Australian gold-fields, although, perhaps, 
 not always performed with intelligence. The feeder on small 
 mills is often expected to break down the big pieces of ore with a 
 sledge hammer, a rock-breaker not being used, but this method of 
 working may be safely set down as irrational and uneconomical, 
 and the result usually is that large and small pieces go into the 
 mortar together. In the United States and in the Transvaal, 
 self-feeders are almost universally employed in modern mills. 
 The art of feeding consists in keeping the depth of pulp on the 
 dies constant throughout the battery, as long as the work is 
 carried on. This is much better done by automatic machinery 
 than by hand, and it was found that by the introduction of the 
 * Meinecke, Proe. Inst. C. Eng., Session 1891-1892, part ii. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 121 
 
 former in California the capacity of the stamps was increased bv 
 15 to 20 per cent., while the wear of shoes and dies was decreased 
 by 25 per cent., and that of the screens by 50 per cent. It is 
 not difficult to discern the cause of the advantages, for, if the 
 dies are insufficiently covered with ore, less crushing is done, while 
 a greater concussion must be taken up by the stamp and by the 
 die, mortar, &c. If the die is quite bare this concussion is so 
 great that the stem may be bent or broken, and the shoe and die 
 battered. On the other hand, if the ore is too deep in the 
 mortar, there is so thick a cushion that much of the force is 
 taken up in compressing without crushing it ; whilst, besides 
 the reduction of output, the head, under these circumstances, 
 sometimes becomes detached from the stem, which is broken or 
 
 Fig. 27. 
 
 Scale, 1 in. = 2 ft. 
 
 Fig. 27a. 
 
 battered by the next blow. The maximum capacity is obtained 
 with "low feeding," the depth of pulp on the dies being about 
 2 inches or less. 
 
 One advantage of even feeding is that a larger proportion of 
 gold is caught, owing to the more regular and even flow of the 
 pulp over the plates, the danger of scouring being diminished. 
 
 There are two classes of automatic feeders, each consisting of 
 a pyramidal hopper with inclined floor, connected with the feed- 
 opening of the mortar-box by a spout or shoot. Stanford's 
 automatic feeder, which was the first invented (in California, 
 about 1870), and is still used, is a typical example of the one 
 class, and Hendy's Challenge feeder, the best machine devised, 
 and the most largely used, is a good example of the other class. 
 
 Stanford's feeder, shown in Fig. 27, consists of a hopper, A, 
 
122 
 
 THE METALLURGY OF GOLD. 
 
 with an adjustable spout, B, which is hinged at C. Fig. 27a 
 is a front view of B, showing how it is suspended. The spout 
 is attached to the vertical rod, E, which is hung to the lever, G, 
 and this has its fulcrum at D. Near the top of a stamp-stem 
 (usually the middle one in the battery), the feeding-tappet, F, 
 is keyed. When the battery is full of ore, the tappet does not 
 come down far enough to strike the lever, G, but, when the ore 
 gets low, the lever is struck, and the result is that the spout, B, is 
 jerked up and down again, the ore being thus thrown forward* 
 
 Fig. 28. 
 
 This machine works well with dry ores which are moderately 
 fine, but, if the ore is wet, and especially if it is argillaceous, it 
 sticks in the hopper until at last a powerful jerk brings it down 
 with a run, the mortar-box is filled up, and all the evils of 
 over-feeding result. The consequence is that the battery is fed 
 as irregularly as by the worst hand-feeding. 
 
 The Challenge feeder (Fig. 28) is constructed so that the 
 tray, A, below the sheet-iron hopper, B, is revolved in a hori- 
 zontal plane by means of a gear-wheel below it, shown in the 
 figure, and this gears with teeth set in the bottom of the tray, A. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 12$ 
 
 The gear-wheel is set in motion by a friction grip, D, placed on 
 the outside of the frame, and actuated through the lever, E, by 
 the bumper-rod, G, against which the tappet of the stamp strikes. 
 At each partial rotation a given quantity of ore is scraped off 
 by the stationary wings or side plates, H, resting on the tray, A. 
 This amount of ore is regulated by the condition of the mortar. 
 The machine is especially adapted for very wet or sticky ores. 
 It weighs, with the frame, about 900 Ibs., and is the most 
 expensive automatic feeder in the market. Other automatic 
 feeders, such as the Tulloch and the Boiler machines, constructed 
 on a similar principle, the ore being scraped and not shaken into 
 the mortar, are cheaper than the Challenge, and are doing excel- 
 lent work on certain kinds of ore. In all cases one feeder is 
 sufficient for a battery of five stamps, more ore being fed to the 
 middle stamps, where the most work is done, than to the end ones. 
 The feeders are now sometimes suspended from above instead of 
 being supported from below. 
 
 Water Supply. The water is supplied to the stamps by 
 horizontal pipes passing just above the top oi the housing of the 
 mortar-box, with an opening supplied with a stop-cock opposite 
 each stamp. The water is commonly made to impinge against the 
 stem of the stamp, either just above or below the wooden casing, 
 and thence run down into the mortar, thus keeping the stem clean. 
 In front of the battery there is also another pipe of about half 
 the size, to supply water to the tables to help carry off the pulp. 
 This pipe is often pierced with pin-holes, so that the water is 
 supplied as a number of fine jets. The water is warmed by 
 steam in winter in many mills, from a belief that amalgamation 
 is promoted by warmth. This belief is founded on general 
 experience rather than on exact experiments, and some results 
 appear to point to a different conclusion. 
 
 Thus, Professor Le Neve Foster obtained the following results* 
 at the Pestarena Mill in Italy, during the years 1869-70 : 
 The average temperature of the water supplied to the mill during 
 the six summer months was 52 R, and the average temperature of 
 the water supplied in the winter months was 39-4 F., and yet, 
 in spite of that, and of the fact that the average temperature, for 
 instance, of the month of January, 1870, was as low as 33 -6 F., 
 he extracted 3'1 per cent, more gold with the cold water than 
 with the warm. These figures do not necessarily prove that cold 
 water is better for amalgamation, as there were in this instance 
 other matters to be taken into consideration, but they show that 
 amalgamation is possible even when the temperature of the 
 water is on an average only 39. The difference in the results in 
 this case might have been due to the turbidity of the water 
 (which was derived from the glaciers at Monte Rosa) in the 
 summer, and its clearness in winter, or to the fact that the 
 pyrites were more liable to decompose in warm weather, and so 
 additional sickening of the mercury was caused in summer. 
 
 * Loc. tit. 
 
124 THE METALLURGY OF GOLD. 
 
 The amount of water used varies from 1J to 6J gallons 
 per stamp per minute, the average in California being about 
 2J gallons, on the Rand about 5J gallons, and in Colorado 
 only about If gallons. In California, with fast-running rapid- 
 discharge batteries, the amount of water per ton of rock crushed 
 varies from 1,000 to 2,400 gallons, the mean being about 1,700 
 .gallons, while in Colorado the average amount is as high as 
 2,500 gallons. Besides varying with the method of crushing 
 adopted, the amount of water varies with the nature of the 
 gangue, clayey ores requiring more, while the large quantity 
 required by sulphide ores is due to the deep discharge necessitated 
 by the difficulty of catching the gold, as well as to the high 
 density of the pulverised material, which renders it more difficult 
 to convey in suspension over the plates. As a rule, the more 
 rapid the output, the less water per ton of ore is required in the 
 battery. Coarse crushing requires less water in the battery, but, 
 on the other hand, more has to be added on the plates. The 
 amount of battery water per ton is increased by over 20 per cent, 
 by the employment of double-discharge mortars. The amount of 
 water to be added on the plates varies with their grade, as well 
 as with the density and size of the particles of crushed ore. It 
 should be only just enough to prevent the pulp from accumu- 
 lating on the plates, as any excess over this tends to check 
 amalgamation and to scour the plates. The average duty of a 
 miner's inch in a gold stamp-mill is given by P. M. .Randall * as 
 12 tons of quartz, if the head under which the water is supplied 
 is 4 inches, and 15*88 tons, if the head is 7 inches. This gives 
 the proportion of the volume of water to that of ore as 11*1 to 1. 
 This may be compared with the proportion of between 7 and 10 
 to 1 in Siberian placer working, and as much as 30, or even 50 
 to 1 in hydraulicking. It may be mentioned that a ton of 2,000 
 Ibs. of quartz occupies about 13 cubic feet when unbroken, and 
 about 20 cubic feet after having been broken up, so that in a 
 lode a cubic yard contains about 2 tons, and in a tailings heap 
 only about 1^ tons. 
 
 Amalgamated Plates. These plates are usually made of 
 copper, and are as much as f inch thick for the inside of the 
 battery and -fa to J inch thick for the outside. The average 
 weight used in California is 3 Ibs. per square foot. It was 
 formerly laid down as a general rule that the heaviest plates 
 were the best, as they last longer and are not so easily dented, 
 but comparatively light plates are now used. The copper should 
 be of the best quality, and, if it is hard, it must be annealed 
 before applying the mercury, so as to make it absorbent. This 
 was formerly done by heating it from below as uniformly as pos- 
 sible until sawdust laid on the upper surface was ignited. The 
 plate is then straightened by blows of a wooden mallet, striking 
 
 * Article on Practical Hydraulics in Sixth Report Gal. State Min., 1886. 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 125 
 
 a block of wood laid on it, and the surface is carefully cleaned by 
 scouring with sand or fine emery paper until quite bright, and 
 washed with strong soda to remove all traces of grease. Cleaning 
 may also be effected by nitric acid diluted with 9 parts of water, 
 or by a 2^ per cent, solution of potassium cyanide, rubbed on with 
 a woollen rag and carefully washed off with water. As soon as 
 the plate is clean, it is rubbed with a mixture of fine sand, sal- 
 ammoniac and mercury by means of a brush, the sal-ammoniac 
 preventing the recommencement of oxidation. More mercury 
 is sprinkled on and wiped over with a piece of rubber until no- 
 more can be absorbed, the whole surface being now thoroughly 
 coated and bright. After having been left for an hour, the plate 
 is finally washed with clean water. 
 
 If the plates were used in this condition they would not 
 catch the gold very well at first, but would continually improve 
 until they had become coated with gold amalgam. In order to 
 make them efficient from the start, they are usually coated with 
 gold amalgam or silver amalgam before being laid down in the 
 mill. Gold amalgam is most effective but is seldom used, as it is 
 so much more expensive than silver. The amalgam is rubbed 
 on with a piece of india-rubber, the plate being wetted with a 
 solution of sal-ammoniac to keep it bright, and such plates last for 
 years without further treatment. A more usual method of 
 preparing the plates in California is to coat them with electro- 
 deposited silver. This plating is done by certain establishments 
 in California for most of the mills, but it can be done on the 
 spot without much difficulty, the plant required being inexpen- 
 sive. After being silvered, the plates have the mercury applied 
 to them. They absorb a large amount of mercury, catch gold 
 well, and are little trouble to keep clean. The plates need not 
 be re-silvered, except after scraping and sweating (see p. 140), 
 as they become coated with gold amalgam in the course of time. 
 About 1 oz. of electro-deposited silver is required per square foot 
 of copper plate. Silvered plates are not used inside the mortar. 
 
 The position of the plates is as follows : The lower edges of 
 the inside plates are level with the upper surface of the pulp, 
 when the battery is working properly i.e., they are usually at 
 1J or 2 inches above the surface of the dies. The plate on the 
 feed side is generally about 9 to 12 inches wide, and is of the 
 same length as the battery ; it is bolted to the mortar itself, and 
 its angle of inclination varies with the shape of the latter, so 
 that the angle of inclination is sometimes 40, and it is sometimes 
 nearly vertical. The plate on the discharge side is inclined at 
 an angle of 10 or 20 to the vertical, and is as wide as the space 
 below the screen permits, being usually from 3 to 6 inches wide. 
 It is fixed to a wooden chuck-block, which has its top bevelled 
 off so as not to obstruct the screen opening. The block is bolted 
 to the mortar with some thicknesses of blanketing between, ia 
 
126 THE METALLURGY OF GOLD. 
 
 order to make a tight joint. Several sizes of these chuck-bleaks, 
 with their copper plates attached, are kept in one and the same 
 mill, a wider block being substituted for a narrower one when 
 the wear of the dies has proceeded to a certain extent. 
 
 At the Alaska Treadwell Mill, and at one or two mills on the 
 Rand, the cast-steel linings of the mortars are furnished with 
 several horizontal slots or recesses for the collection of amalgam, 
 and these take the place of the back copper plates. The front 
 or chuck-plate is, however, retained in these mortars. This is 
 shown in Fig. 45a, Chap. X. 
 
 The outside plates are fastened to a wooden table with copper 
 nails, or wooden damps, or by wedges driven into the raised 
 edges of the table. The table is as wide as the battery (4 feet 
 7 inches) and usually from 6 feet to 8 feet long. In California a 
 length of 2 or 3 feet of plates of the same width, the apron plates, 
 are interposed between the battery and the tables proper, on to 
 which there is a drop of 2 or 3 inches. Below the tables there is 
 .a succession of four or five sluice plates, each about 30 inches long, 
 with drops of 1 to 3 inches between them. They are usually 
 made narrower than the others, and are frequently only 1 2 or 
 18 inches wide, but this practice is not to be commended, as the 
 stream of ore and water, forced into a narrower channel, becomes 
 deeper and flows more rapidly and tumultuously, with the result 
 that the contact between the ore and amalgamated plates is much 
 reduced, and very little gold is caught. The use of the drops is 
 to assist in catching the float gold and to separate the amalgam 
 which has become floured and mixed with the pulp. In place of 
 the composite arrangement described above, a single length of 
 12 or 14 feet of plates is often used, especially in the Transvaal. 
 
 To aid in catching the float gold, swinging amalgamated plates 
 have been introduced, and are in use in the sluices below the 
 batteries of many Californian mills. They are also used in 
 hydraulicking. The swinging plate consists of a curved strip of 
 silver-plated amalgamated plate about 3 inches deep, and of the 
 same width as the sluice in which it is hung; it is suspended 
 on eyes through which wires pass. The plate thus hangs, half 
 submerged, with its concave side up-stream, and is kept swinging 
 by the current, so that all floating particles of gold must come in 
 contact with it. It is found in practice that, immediately under 
 each plate, across the sluice, a line of amalgam which has dropped 
 from the plate accumulates. The plates are placed a few feet 
 apart. They cost little and are very effective. 
 
 Mill Site. This should be easily accessible by road, rail, and 
 water, if possible ; moreover, it should be near both wood and 
 water, and there should be a good fall of the ground. The least 
 fall that is considered sufficient in California is 33 feet from the 
 mouth of the rock-breaker to the floor on which the concentrators 
 .are placed, when rock-breakers are used, followed by stamps, 
 .copper-plated tables, sluice plates, and two successive concentra- 
 
QUARTZ CRUSHING IN THE STAMP BATTERY. 
 
 127 
 
 tion tables. If a second concentrator is dispensed with, however, 
 and space otherwise economised as far as possible, 29 J feet may 
 be enough. 
 
 ji(W*u * '"I 
 
 Arrangement of the Mill. The general disposition of the 
 machinery is shown in section in Fig. 29. This represents a 
 mill in which the ore is delivered from the ore-cars through a 
 
128 THE METALLURGY OF GOLD. 
 
 grizzly on to the rock-breaker floor, and thence by a shoot to the 
 automatic feeders of the stamp battery ; the pulp, after passing 
 over the plates, is conveyed by sluices to the double row of 
 " frue-vanners " (described on p. 185), which are shown standing 
 back to back on the lowest floor. The ore-bins should be 
 of sufficient capacity to hold several weeks' supply for the 
 stamps, in case of breakdowns or other delays in the mine. 
 Their floors are made of planking, which is laid with the 
 lengths in the direction of the slope, for, if placed trans- 
 versely, the boards wear fast, and the ore packs at the edge 
 of each one, with the result that its movement is impeded 
 and must be assisted by shovelling. The slope should be- 
 at least 45 in order to enable the ore to move downwards by 
 gravity, when the lowest portion is drawn from the shoot. Ore- 
 bins with flat bottoms have greater capacity, but necessitate an 
 additional handling of the ore. The sills of the bins should be 
 placed horizontally on terraced ground, not on the slope of 
 the hill. A shoot from the ore-bin door leads to a grizzly, 
 through which the fine ore drops into the main battery ore-bin. 
 The larger pieces of rock are discharged either into a coarse ore- 
 bin or else upon the platform by the side of the rock-breaker and 
 on a level with its mouth, into which it is shovelled by hand. 
 The former course is preferable, as in that case the rock-breaker 
 can be fed continuously by a gate in the coarse ore-bin, which is- 
 opened and shut by a rack and pinion. By this arrangement 
 there is a saving of labour, but the chief advantage is that the 
 rock-breaker is thereby kept constantly at work. At the North 
 Star Mill, California, it was found that when, by arranging for 
 a continuous feed from the coarse ore-bin down a shoot leading 
 direct to the rock-breaker, the latter was in constant work, it 
 absorbed 12 horse-power, as against 8 horse-power when in inter- 
 mittent work, but its output was over 50 per cent. more. When 
 the stamp battery is used only to crush the ore, which is subse- 
 quently treated in pans or other amalgamators, or by concentra- 
 tion, it is of great advantage to separate the fine product of 
 the rock -breaker by sieving, instead of passing the whole through 
 the stamps. This arrangement increases the output and pre- 
 vents unnecessary sliming of the ore, thus greatly reducing the 
 loss of sulphides when an attempt is made to save these by 
 concentration. As observed on p. 106, it is becoming cus- 
 tomary in large mills to place the rock-breakers in separate 
 buildings. 
 
 The product of the rock-breaker is mixed a"s thoroughly as 
 possible with the ore, which originally passed through the grizzly, 
 and led by means of a shoot direct to the automatic feeders, 
 which should be mounted on wheels, so as to be readily movable. 
 Plenty of space must be left behind the feeders for convenience in 
 repairing and in the exercise of supervision. The amalgamating, 
 
AMALGAMATION IN THE STAMP BATTERY. 129 
 
 tables should also be easily accessible, space being left to pass be- 
 tween them, and the same remark applies to the sluices and the 
 tables or other appliances for concentration. All shafts, bearings, 
 &c., should also be easily accessible, so that oiling, re-lining, and 
 repairs may be readily done. . 
 
 The tailings are discharged into a large sluice by which they 
 are carried into a river, or into the sea, or run into settling pits, 
 or impounded behind dams. One of the two latter courses is 
 adopted, either if water is scarce, so that it is necessary to use it 
 over again, or if the discharge of tailings is forbidden by law, or 
 if the tailings are rich enough to be subjected to further treat- 
 ment, at once or at some future time. 
 
 The whole of the machinery is contained in a strong building 
 of timber or corrugated iron, to protect it and the workmen from 
 the weather, and to prevent theft of amalgam. The interior 
 should be lighted by as many windows as possible, in order to 
 facilitate superintendence and repairs. In cold climates, the 
 mill buildings are sometimes warmed by passing the flues from 
 the boiler fires through them from end to end, before leading the 
 products of combustion to the stack, or by steam-pipes. 
 
 CHAPTER VII. 
 AMALGAMATION IN THE STAMP BATTERY. 
 
 Treatment of the Plates. In order to keep the plates in proper 
 condition so that successful amalgamation may be maintained, 
 they must be prepared carefully, and the closest watch kept over 
 them. The silver-plated copper table is preferred in California 
 from the ease with which it is kept clean, but is not used in the 
 Transvaal. It is not considered desirable to put on it as much 
 mercury as it will hold, since, if the amalgam is too fluid, losses 
 are sustained by scouring, but, on the other hand, if the amalgam 
 becomes too hard and dry from absorption of gold and silver, 
 further amalgamation is checked and fresh mercury must be 
 added. 
 
 The condition of the inside plates is regulated by the amount 
 of mercury supplied to the mortar. In Colorado there is an 
 opening at the front of the battery and above the screen frame, 
 ordinarily covered by canvas, which can be lifted up by the mill- 
 man, who introduces his arm, and determines by passing his hand 
 over the front plate whether the right amount of mercury is being 
 added by the feeder. The regulation of the addition of mercury 
 
130 THE METALLURGY OF GOLD. 
 
 is thus effected without removing the screen frame. The amount 
 added varies with the conditions of crushing and the richness 
 of the ore, but in general from 1 to 2 ounces of mercury are fed 
 in for every ounce of gold contained in the ore. The finer the 
 state of division of the gold and the more sulphides there are 
 contained in the ore, the more mercury is required. It is fed 
 into the battery at stated intervals of from half an hour to an 
 hour. In some mills, particularly in Australia, amalgamation in 
 the battery is not attempted, no inside copper plates being 
 provided, and under these circumstances it is not usual to feed 
 mercury into the battery. 
 
 The practice of feeding mercury into the battery, although 
 almost universally pursued in the best mills in the United 
 States and South Africa, still meets with opposition from certain 
 experienced millmen. The objections urged against it are mainly 
 that the mercury so introduced, and the amalgam formed through 
 its agency, tend to become so excessively subdivided that a high 
 percentage is lost through flouring; moreover the mercury is liable 
 to sicken when the ores contain sulphurets. These evils, no doubt, 
 exist, and tend to increase with the percentage of sulphides present, 
 while arsenic and antimony in particular cause heavy losses of 
 both mercury and amalgam if battery amalgamation is attempted, 
 but with ordinary free-milling ores such losses are not serious. 
 
 The amalgamated plates are dressed as frequently as is neces- 
 sary, the length of time allowed to elapse between two operations 
 depending partly on the richness of the ore. To dress the 
 plates, the battery is stopped, and the amalgam wiped oif the 
 plates with a brush or piece of indiarubber held between two 
 pieces of wood, so that only about half an inch of it projects ; 
 and, if necessary, fresh mercury is added. The amount of 
 mercury put on the plates should be enough to keep their 
 surfaces in a pasty condition, but not enough to gather into 
 liquid drops or to run off. The plates are wiped three or four 
 times a day, or as often as every two hours when rich ores are 
 being treated, whilst, with exceptionally valuable materials, the 
 operation may be necessary at intervals of a few minutes only. 
 
 Discoloration of the Copper Plates. The plates often 
 become stained by the formation on them of oxides, carbonates, 
 or other compounds of copper through the corrosive action of the 
 water and pulp. Ores containing decomposing sulphides acidify 
 the water and thus cause the corrosion of the plates, a yellow 
 film being formed on the surface of the metal. The presence of 
 carbonic acid in the water is equally harmful, but Aaron points 
 out that the addition of slaked lime to the water neutralises 
 the acid substances and diminishes the tarnishing. The yellow, 
 brownish, or greenish discoloration, the so-called " verdigris," 
 appears in spots and spreads quickly, especially on new 
 plates, those which have been silver-plated being less liable 
 
AMALGAMATION IN THE STAMP BATTERY. 131 
 
 to become dirty than the others ; whilst when a plate has become 
 covered with a thick layer of amalgam it is not readily dis- 
 coloured. When these stains appear the plate must be at once 
 cleaned, as the stained part catches little or no gold. The 
 chemicals used for the purpose are generally sal-ammoniac and 
 potassic cyanide, the operation being conducted as follows : 
 The battery is stopped, the plates rinsed with clean water, and 
 a solution of sal-ammoniac applied to the stained parts with a 
 scrubbing brush, and left covering them for a few minutes in 
 order to dissolve the oxides. It is then washed off, a solution 
 of potassic cyanide rubbed on to brighten the plate, and almost 
 instantly washed off, fresh mercury being then added if neces- 
 sary. Janin states (Mineral Industry, 1894, p. 332) that long 
 brushing with potassium cyanide is necessary, as otherwise the 
 spots reappear when the water is turned on. 
 
 Whisk brooms are perhaps better than indiarubber for brush- 
 ing the plates ; these brooms are cut down to a short length so 
 as to be stiff enough. The plate is brushed all over and the 
 amalgam thus thoroughly loosened from it, after which, com- 
 mencing at the top where the amalgam is thickest, it is sub- 
 jected to a systematic stiff brushing, each stroke being directed 
 longitudinally down the table, and not towards the centre. The 
 surplus amalgam is thus brushed to the lower end of the plate, 
 whence it is removed, and a thin coating of amalgam is left over 
 the whole surface of the plate, excluding the air and preventing 
 the formation of " verdigris." 
 
 Chemicals Used to Promote Amalgamation. The use of 
 chemicals to aid amalgamation was formerly much more general 
 than at present, although battery-men were less afflicted by the 
 rage for nostrums than pan-amalgamators. Almost the only 
 chemical now used on a battery in California or Colorado, 
 both to promote amalgamation and to clean the plates, is cyanide 
 of potassium, and the use even of this reagent is becoming less 
 general every year. A dilute solution is believed to promote amal- 
 gamation, but probably its action consists merely in thoroughly 
 -cleaning the plates and the mercury from all trace of oil, grease, 
 and base metallic oxides. 1 or 2 Ibs. of potassic cyanide should 
 be enough to supply a 40-stamp battery for twelve months 
 when treating free-milling ores, by which the mercury is not 
 much affected. The difference between such a mill and one 
 running on base ores may be judged from the fact that at the 
 Hidden Treasure Mill, Colorado, where there are seventy-five 
 stamps, 260 Ibs. of cyanide were used in a year. Here the plates 
 are dressed every twelve hours with a weak solution containing 
 2 oz. of cyanide in 3 gallons of water, the operation being neces- 
 sitated by the acidity of the water which comes from the mine, 
 and is further contaminated by sulphates formed in the ore. 
 Sodium is now used mainly to clean the mercury, after it has 
 
132 THE METALLURGY OP GOLD. 
 
 been retorted, before again using it in the battery. A weak 
 solution of caustic soda is used to remove grease from the plates. 
 
 The merest trace of any kind of grease or oil is very prejudi- 
 cial to successful amalgamation, forming a film over the plates 
 and over the little globules of mercury, and thus preventing 
 contact between them and the particles of gold. Ammonia or 
 potassic cyanide removes the film, but a saturated solution of 
 wood-ashes, or of soda carbonate or caustic soda is more gener- 
 ally used than these reagents. Grease may consist of the tallow 
 dropped from the miner's candles, or the oil from the loose steam 
 which is sometimes used to warm the battery feed-water, or from 
 the bearings of the machinery. 
 
 Mercury. The mercury used for the battery and the plates 
 should be quite free from base metals, such as copper, zinc, lead, 
 or tin. If these are dissolved in the mercury they become 
 rapidly oxidised and soon cause it to " sicken," so that it- 
 breaks up into a number of minute globules, each coated with 
 a layer of base metallic oxide. These globules are not only 
 useless for amalgamation, but as they are very minute and refuse 
 to coalesce, they are carried away in the tailings, together with 
 such gold as may have been taken up before the oxidation had 
 proceeded far. Such impure mercury, when used to coat the 
 plates, also causes their discoloration by oxidation of the dis- 
 solved base metals. Mercury acts best when it already contains 
 gold and silver, and it is customary to dissolve some silver in 
 the new mercury applied at the starting of a mill, if no old 
 mercury can be had to mix with it. When work has been some 
 time in progress there is no difficulty on this account, as the 
 quicksilver, which has been squeezed through a filter to separate 
 the amalgam, carries some dissolved silver and gold with it. 
 
 Grade of Plates. The grade or inclination of the plates 
 varies with the nature of the ore to be treated, heavy pyritic ores 
 requiring a higher grade than light quartz, while the coarser the 
 crushing the steeper must be the grade. In California the 
 copper tables have an inclination of from 1^ to 2 inches per 
 foot, the apron plates from J to lj inches per foot, with an 
 average of 1 J inches, and the sluice plates from 1J to 1J inches. 
 With heavily sulphuretted ores a grade of from 2 to 2 J inches per 
 foot is used. The steepness of the grade is of great importance, as 
 on it, and on the amount of water supplied, the attainment of the 
 necessary contact between the ore and the plate depends. When 
 the pulp is flowing properly, it travels down in a series of little 
 waves and ripples, and, in consequence of the friction between 
 the plate and the film of water in contact with it, the upper 
 portions of these little waves travel faster than the lower parts, 
 so that the motion becomes one of tumbling over and over. As 
 a result of this, if the plate is long enough, every particle of 
 pulp comes in contact with the amalgamated surface, and the 
 
AMALGAMATION IN THE STAMP BATTERY. 133 
 
 perfect extraction of amalgamable gold, mercury and amalgam is 
 obtained. 
 
 Muntz Metal Plates. The use of Muntz metal (which con- 
 sists of copper 60 per cent., zinc 40 per cent.) for amalgamated 
 plates is of great interest. It differs from copper in catching 
 gold well as soon as the plate is amalgamated, not requiring to be 
 covered with gold- or silver-amalgam before it begins to do good 
 work. Moreover, the amalgamated surface is very superficial, 
 since the mercury does not sink in so far as it does into a plate 
 composed of pure copper, so that only a small quantity of 
 mercury is required to cover it. The result is that cleaning 
 up is easy and rapid, no iron instrument being necessary, but 
 rubber being always sufficient. These properties make it par- 
 ticularly valuable for custom mills, where it is desirable to catch 
 AS much as possible without mixing the amalgam obtained from 
 two parcels of ore crushed in succession. On the other hand, as 
 it holds little mercury, it cannot absorb much gold, and must be 
 <jleaned-up at frequent intervals. 
 
 The mercury on Muntz metal plates does not suffer so easily 
 from "sickening" as that on copper plates ; it has been suggested 
 that this is due to the electrolytic action of the copper-zinc 
 couple, which sets free nascent hydrogen, and so reduces the 
 compounds of mercury and other metals which have been formed. 
 It follows that Muntz metal plates are preferable for ores con- 
 taining large amounts of heavy sulphides or arsenides. The 
 greenish-yellow stains (called " verdigris " by millmen) which are 
 formed on copper plates when grease and other impurities are 
 present in the battery water, do not appear when Muntz metal is 
 used, and such discolorations as occur on these plates can be 
 better removed by dilute sulphuric acid than by potassium 
 cyanide. At the Saxon Mill, New Zealand, the copper plates 
 formerly required 7 Ibs. of cyanide, costing 23s., per month to 
 keep them in order, while the Muntz metal plates, by which 
 they were replaced, could be kept clean by 5 Ibs. of sulphuric 
 acid per month, the cost being 3s. 4d. It is stated, however, 
 that, in the treatment of highly acid ores, which have been 
 weathered for some time so that they contain large quantities 
 of soluble sulphates, or in cases where the battery water contains 
 -acids, copper plates are less affected than Muntz metal, over 
 which a scum is rapidly formed. This is not the experience in 
 the Thames Valley, N.Z., where Muntz metal is preferred in 
 spite of the extremely acid nature of the water and ore. 
 
 Generally, it may be stated that Muntz metal plates are 
 cheaper, wear better, and require less attention than copper. In 
 dressing new Muntz metal plates the following method is 
 adopted in New Zealand : The surface of the plate is scoured 
 with fine, clean sand ; then it is rinsed with water, and washed 
 with a dilute (1 to 6) solution of sulphuric acid. Mercury is 
 
134 THE METALLURGY OF GOLD. 
 
 then applied and rubbed in with a flannel mop until it wets the 
 surface of the plate (i.e., amalgamates with it) in one or more 
 places, after which the mop is given a circular movement, passing 
 through these spots, until the amalgamation of the surface 
 spreads from them over the whole plate. 
 
 The discoloration of the Muntz metal plates is prevented by 
 the weak electric current produced, as has been already stated. 
 The same effect can, according to Aaron, be obtained when 
 ordinary copper plates are in use, by placing them in contact 
 with iron or some other metal which is positive to copper. Strips- 
 ot iron bolted to the top and sides of the plate are said to be suffi- 
 cient tor the purpose, the copper being in that case unaffected by 
 the acidity of the water, which causes oxidation and dissolution of 
 the iron only. Jaiiin's experience does not support these views* 
 
 Shaking Copper Plates. A shaking copper plate is recom- 
 mended by some of the best authorities to be used either below 
 or in place of the ordinary amalgamating tables, especially in 
 cases where these do not appear to give good results. An 
 ordinary fixed copper plate requires an inclination of from 1J to 
 2 inches per foot, in order to keep it clear of sand, when the plate 
 is of the same width as the battery. If, however, the plate is 
 subjected to a short rapid shake, the sand is kept from packing, 
 and amalgamation is well performed with a grade of only J to J 
 inch per foot, or the amount of water needed with the pulp may 
 be greatly reduced and better contact thus obtained. For these 
 plates, silver-plated copper is the material employed. They are 
 affixed to a light wooden frame which is moved by a crank- shaft, 
 revolving 180 to 200 times per minute, placed on one side, with 
 a throw of 1 inch at right angles to the direction of the flow of 
 the pulp. In some mills, a longitudinal shake is given to the 
 plate instead of this side shake. The frame may be hung on 
 rods from above, but is more conveniently supported on four 
 short iron springs, forming rocking legs. The width of the tables 
 should be made as great as possible, while the length is of less 
 importance, as, the thinner the current of pulp flowing over them, 
 the better the chance of the gold particles coining in contact 
 with the plates and being retained. These shaking plates were 
 first used in Montana in 1878, and have since been employed 
 at several Californian mills, giving satisfactory results. It is- 
 advantageous to add to them an amalgam- and mercury-saver. 
 A simple device for this purpose is to nail a strip of wood, 
 half an inch thick, across the copper plate near the top, thus 
 forming a shallow riffle, the angle of which is soon filled with 
 sulphides and coarse sand, which are kept in agitation by the 
 movement of the table. This is stated by W. M'Dermott and 
 P. W. Duffield* to be the most effective contrivance yet devised 
 for catching quicksilver and hard amalgam. If the inside copper 
 * Gold Amalgamation, London and New York, 1890, p. 16. 
 
AMALGAMATION IN THE STAMP BATTERY. 135 
 
 plate should become hard by accident or neglect, chips of amal- 
 gam escaping through the screens are retained in this riffle, and, 
 becoming spherical by rolling up and down under the effect of 
 the shaking motion, increase in size just as a snowball does when 
 rolled in snow. 
 
 Corrugated Plates. Another device to increase the catching 
 power of copper plates is to make them in a corrugated form so 
 that they will hold little pools of mercury. Where much float 
 gold is believed to be passing off, this is supplemented by the 
 addition of a second plate placed above and facing the other with 
 its amalgamated surface downwards. The pulp is allowed just 
 sufficient room to pass between them, in contact with both, and 
 there is little doubt that in this way some fine gold is caught 
 that would otherwise escape. 
 
 Mercury Wells. Another method of saving mercury and 
 amalgam, which would otherwise be lost in the tailings, consists 
 in the application of mercury wells or riffles. A mercury well 
 or riffle consists of a shallow gutter filled with mercury, over the 
 surface of which the pulp flows or through which it is forced to 
 pass by suitable machinery. Attwood's amalgamator, formerly 
 much used in California, was a machine of the latter class. 
 Such wells are even now often used in modern mills. It is 
 believed by some of the best authorities that such devices catch 
 no more mercury than could be secured by either well-arranged 
 stationary or shaking plates, while the practice of placing a well 
 between the screens and the amalgamating tables is especially to 
 be condemned, as it prevents proper supervision being kept over 
 the feeding of mercury into the battery, over-feeding being 
 difficult to detect under these circumstances. 
 
 Galvanic Action in Amalgamation. In amalgamation in 
 the mortar, on plates, or in pans, not only are free metals 
 absorbed, but the dissolved salts, and, to a less extent, the 
 insoluble compounds of the heavy metals, are reduced and 
 amalgamated, chiefly by galvanic action. The copper of the 
 plates, or the iron of the mortar or pan, constitutes the positive 
 element, and all metals less oxidisable than this reacting metal 
 are reduced by it, and are then amalgamated by the mercury. 
 In this way iron reduces both lead and copper, although, if these 
 are present in the form of undecomposed sulphides, this action 
 will be very slight. Now, if lead is introduced into the amalgam, 
 the latter becomes pasty, and is subjected to considerable losses, 
 and copper has an equally harmful effect. It is for this reason that 
 the arrastra is found to be better than the stamp battery or even 
 the pan for certain plumbiferous ores. This action of iron is of 
 course enormously increased if the ores are subjected to a 
 chloridising roast before being amalgamated, as in the Reese 
 River process for the treatment of auriferous silver ores. 
 
 In a number of mills, this galvanic action is increased by the 
 
136 THE METALLURGY OP GOLD. 
 
 passage of a weak electric current through the charge by means 
 of a dynamo. The amalgamated plates or the walls of the pan 
 are connected with the negative pole, while the positive pole is 
 formed of a plate of lead or iron dipping into the pulp. Under 
 these conditions the mercury is still further protected from 
 attack, and remains bright and lively, but the deposition of base 
 metals in it is favoured, and the stronger the current the more 
 this action is induced. Consequently, such methods are attended 
 with the best results when dealing with ores containing little or 
 no copper, lead, &c., since in these cases the strength of current 
 can be increased, and the mercury kept clean, without any ill 
 effects. The principle is made use of in Bazin's centrifugal 
 amalgamator, Molloy's hydrogen amalgamator, and other similar 
 machines. 
 
 Designolle Process of Amalgamation.* In this process a 
 solution of mercuric chloride is used. It was tried at the Haile 
 Mine, South Carolina, the method being as follows : Charges of 
 600 Ibs. of roasted ore were placed in cast-iron barrels with 
 1,000 Ibs. of cast-iron balls, or in pans. The barrel was partly 
 filled with water, and 1 gallon was added of a solution containing 
 1'7 per cent, of mercuric chloride, the same amount of hydro- 
 chloric acid, and twice as much salt, if the ore contained less 
 than 15 dwts. of gold per ton. This would be equivalent to 
 about 10 parts of mercury to 1 part of gold. The barrel was 
 rotated for twenty minutes and then discharged into a settler, 
 and the suspended amalgam caught on copper plates. The 
 mercury was supposed to be reduced by the iron thus- 
 
 HgCl 2 + Fe = FeCl 2 + Hg, 
 
 and the metallic mercury thus freed amalgamated with the gold. 
 If no common salt was present, some mercurous chloride was 
 formed according to the equation 
 
 2HgCl 2 + Fe = Hg 2 Cl 2 + FeCl 2 , 
 
 and the subsequent reduction of the insoluble calomel by iron 
 was not complete. Hydrochloric acid was supposed to hasten 
 the amalgamation by setting up some electrolytic action. 
 
 The total cost of the process at the Haile Mine was said to be 
 only 35 cents, or Is. 7Jd., per ton, but it was abandoned when 
 the percentage of iron in the material under treatment increased, 
 owing to improved methods of concentration. Large quantities 
 of oxide of iron were then amalgamated, and rendered the 
 resulting mass harder to treat than the ore itself. By repeated 
 washing, settling, and regrinding with fresh mercury, it could 
 be partially purified, but not without a loss of gold. It is stated 
 that 87 per cent, of the gold in the ore was extracted. 
 
 * This account is abridged from that given by Louis Janin, Junr., in 
 Mineral Industry in 1894, p. 346. 
 
AMALGAMATION IN THE STAMP BATTERY. 137 
 
 The Clean-up. The amalgam, both on the inside and outside 
 plates, does not accumulate evenly, but in ridges and knots 
 which serve as nuclei for the collection of more. It is not 
 advisable to allow the coating of amalgam to become very thick, 
 since, although the plates catch better as the amalgam accumu- 
 lates, losses may be experienced by scouring. The wiping of the 
 outside plates usually takes from 10 to 15 minutes for each 
 battery. The amalgam so obtained is ground with more mercury 
 in a clean-up pan in order to soften it, the skimmings from the 
 mercury wells, &c., being added to the charge. The inside plates 
 are not wiped until the amalgam stands up in ridges on it ; the 
 operation may be necessary as often as twice a week, but it 
 usually takes place twice a month, when a general clean-up 
 is made. All amalgam, however obtained, unless it is already 
 hard and dry, is usually at once separated from the excess of 
 mercury contained in it by being squeezed in filter-bags, and the 
 pasty residue alone passed to the clean-up pan or the retort. 
 
 In cleaning-up, the stamps are hung up, two batteries at a time, 
 the screens, inside plates, and dies are all taken out, and the 
 " headings," or contents of the mortar, consisting of the pulp, 
 mercury, sulphides, and pieces of iron and steel, amounting in all 
 to a quantity sufficient to fill two or three buckets, are carefully 
 scraped out and fed into the mortar of one of the other batteries, 
 which has not yet been cleaned up. In California, the headings 
 from the last batteries are panned, the iron removed by a magnet, 
 and the remainder ground with mercury in the clean-up pan 
 In Gilpin County, however, it is not panned, but merely 
 returned to the battery on restarting. Amalgam is found 
 adhering to the inside of the mortar and to the dies, and is 
 carefully detached and added to the clean-up pan, and all the 
 plates are well scraped with a sharp-edged piece of hard rubber, 
 care being taken not to scratch them. The plates are then 
 redressed and put back, the batteries restarted, and the next ones 
 stopped and cleaned up. Three men can clean up forty stamps 
 in from five to seven hours, ten stamps being thus idle for the 
 whole of this time. 
 
 The amalgam obtained from the batteries, outside plates, 
 mercury wells, or sluices, is rarely clean enough for immediate 
 retorting ; it is usually found to contain mixed with it grains 
 of sand, pyrites, magnetite and other minerals, together with 
 fragments of iron and other foreign substances. The skim- 
 mings from mercury wells are still more impure. These materials 
 must be purified by grinding with fresh mercury and washing 
 before they cm be passed to the retort. The scraps of iron 
 consist of fragments of shoes, dies, shovels, picks, hammers, and 
 drills: they are knocked about in the mortar until a quantity 
 of gold and amalgam has been driven into their interstices. At 
 the Jefferson Mill, Yuba County, California, about ^ ton of such 
 
138 THE METALLURGY OF GOLD. 
 
 scrap, picked out by hand or by a magnet, had accumulated in 
 1885. It was attacked by warm dilute sulphuric acid until the 
 surface had all been dissolved oft', and the residue was then well 
 washed, and gold to the value of $3,000 thus recovered. The 
 shoes, dies, &c., which were too large for this treatment, were 
 boiled in water for half an hour, and then struck by a hammer, 
 when the gold dropped out.* In Colorado, and in small mills 
 generally, the dirty amalgam is ground in a mortar by hand with 
 fresh mercury and hot water, until it is reduced to an even thin 
 consistency, when the dirty water is poured off, and the mercury 
 poured backwards and forwards from one clean porcelain basin to 
 another until the pyrites, dirt, &c., have risen to the surface, when 
 they are skimmed off. The clean mercury is then squeezed 
 through canvas or wash-leather, when the greater part of the gold 
 and silver contained in it, together with about one and a-half 
 times its weight in mercury, remains in the bag, the rest of the 
 mercury, with a small quantity of the precious metals dissolved 
 in it, passing through. The skimmings obtained are put back 
 into the mortar, and re-ground by themselves with fresh mercury. 
 
 In large mills a clean-up revolving barrel is often employed to- 
 mix the amalgam. This is made of iron, and is similar in con- 
 struction to chlorinating barrels, but without the lead lining. 
 At the Plumas Eureka Mill, the barrel is 3 feet in diameter and 
 4 feet long, and revolves twenty times a minute ; the charge is 
 700 Ibs. of amalgam and 20 Ibs. of mercury, or more if the 
 amalgam is very rich. A dozen or more pieces of iron, such as 
 worn out battery shoe shanks, are put into the barrel, which is 
 filled nearly up to the top with water, and then revolved for 
 from six to twelve hours. The use of the iron is to help to mix 
 the amalgam and mercury, but it causes some loss by flouring, 
 and is omitted with advantage in Australian mills. The barrel 
 is then opened and washed out with water, the tailings being 
 run over amalgamated plates and through a mercury well or 
 some other form of amalgam-saver, after which the amalgam is 
 scooped out of the barrel and squeezed again in wash leather. 
 
 The Clean-up Pan is even more extensively used than the 
 barrel. One of the oldest in use in the United States, the 
 Knox Pan, is still a great favourite, especially for treating 
 battery sands, skimmings, &c. It consists (Fig. 30) of an iron 
 pan, 5 feet in diameter and 14 inches deep. Wooden or iron 
 shoes are attached to the arms, g, which make from twelve to 
 fourteen revolutions per minute. Iron shoes are considered 
 better for brightening or polishing the particles of gold con- 
 tained in the pyrites, and so rendering them fit for amalgama- 
 tion. The charge for this pan is about 300 Ibs. of concentrates, 
 skimmings, &c. Concentrates are treated in this way only when 
 the quantities of them are small, while at the same time they 
 * Sixth Report Gal. State Min., 1886. 
 
AMALGAMATION IN THE STAMP BATTERY. 
 
 139 
 
 contain mercury and amalgam. The charge is made into a pulp 
 with water and ground for three or four hours, after which 50 Ibs. 
 of mercury are added, and mixing is carried on for a few hours 
 longer, before the pulp is diluted, settled, and discharged. The 
 tailings suspended in the water are often caught in settling pits, 
 and either sold or subjected to further treatment on the mill, as 
 they are frequently of high value. When the Knox pan is 
 required for mixing dirty or impure amalgam with additional 
 mercury, the wooden shoes are used, as shown in the figure. 
 The waste matter is then washed off with a stream of water as 
 before, and the amalgam retorted. 
 
 Fig. 30. 
 
 The position from which the greater part of the gold is 
 obtained in a clean-up varies according to the ore and the 
 method of treatment. In California from 50 to 80 per cent, of 
 the gold saved is caught on the single plate inside the battery, 
 the remainder being caught on the outside plates and the sluice 
 plates, or being contained in the concentrates. In the Grass 
 Valley, at the Original Empire and the North Star Mills, from 
 70 to 85 per cent, is caught inside the battery. The amalgam 
 from the battery plates is usually richer than that from the 
 outside plates, especially if the gold is coarse. The reason for 
 this is that coarse gold, being easily amalgamable, is almost all 
 caught on the inside plates, while fine gold, even if amalgamated 
 in the battery, forms a more fluid amalgam which passes through 
 
140 THE METALLURGY OF GOLD. 
 
 the screens and is caught outside. Coarse gold forms a richer 
 and stiffer amalgam than tine gold, because it is more easily 
 taken up by amalgam which is already almost saturated with 
 gold. At the two mills last named the value of the plate 
 amalgam is $4.50 per oz., and that of the battery amalgam 
 -$8.50 per oz. 
 
 About every three or six months the plates are scraped with 
 & sharp iron chisel or palette knife, or scaled by chipping with a 
 hammer and chisel. At the North Star Mill, California, the 
 plates are immersed in boiling water so as to soften the amalgam 
 before they are scraped. They are sometimes "sweated" in 
 California by heating them over a wood fire before scraping 
 them. At the Empire Mill, Grass Valley, California, the sweat- 
 ing of the outside and apron plates of four batteries produced 
 bullion to the value of $19,000.* These plates had been down 
 for eighteen months, and the ore which had been run over them 
 averaged 18 dwts. of free gold per ton. After scaling and 
 sweating, the plates may require replating. In course of time 
 they are worn out, the copper becoming brittle and worn into 
 holes, but they usually contain gold enough when discarded to 
 pay for a new set. 
 
 Several methods are in use for recovering the gold from old 
 plates. For example, they may be dissolved in nitric acid, 
 when the gold is left nearly pure. A more economical method 
 of detaching the gold, much used in Australia, is described by 
 Mr. M'Cutcheon as follows : The plates are placed on the 
 hearth of a reverberatory furnace, or on a fire made with logs 
 in the open air, and the mercury expelled at a gentle heat. If 
 the temperature is too high, the gold sinks into the copper at 
 once, and the copper must then be dissolved. After the mer- 
 cury has been driven off, the plate appears to be more or less 
 coated with gold on one side. This surface is treated with 
 hydrochloric acid for eight or ten hours, and the plate is then 
 replaced on the hearth and exposed to a dull red heat until well 
 blackened. On plunging it into cold water, the gold now scales 
 off, and is collected and freed from copper by boiling in nitric acid. 
 
 Retorting. The solid amalgam, which is retained in the 
 canvas or wash-leather filters, usually contains from 30 to 45 per 
 cent, of gold and silver, according to the state of division of 
 the gold present in the ore, and also to the degree of care exer- 
 cised in squeezing out the excess of mercury. For separating the 
 gold from the mercury there are two kinds of retorts in general 
 use the pot-shaped retort, which is sometimes cast with trun- 
 nions to swing on supports, in small mills ; and the cylindrical 
 retort, shown in Fig. 31, in larger mills. The pasty amalgam 
 is rolled up into balls or kneaded into cakes, and squeezed 
 into the pot -shaped retort, and often rammed down with a 
 * Eighth Report Gal. State Min., 1888, p. 714. 
 
AMALGAMATION IN THE STAMP BATTERY. 
 
 141 
 
 bolt-head, although this course is deprecated by some managers, 
 who prefer to leave the amalgam as spongy and open in 
 texture as possible, believing that a uniform product is thus 
 obtained more rapidly at a lower temperature and without so 
 much loss. In the horizontal retort the amalgam is placed in 
 iron trays divided into compartments by partitions. In either 
 case, the retorted metal is prevented from adhering to the iron, 
 either by laying it on three or four thicknesses of paper, the 
 ashes of which remain beneath the amalgam, or by covering the 
 iron trays with a coating of whitewash. The former plan is 
 preferred. The mercury is condensed in cooling tubes passed 
 through water ; the loss through volatilisation is usually very 
 
 Fig. 31. 
 
 small, and may be taken as being about one grain of gold per 
 pound of mercury. 
 
 The charge is heated slowly until the boiling point of mercury 
 is reached, when the fire is checked, and the retort kept at an 
 even temperature for one or two hours, or until the bulk of the 
 mercury has been driven off. The retort is then raised gradually 
 to a bright red heat to expel the remainder, and after cooling 
 it is opened, the trays are withdrawn, and the retorted metal is 
 loosened by a chisel, if necessary, and turned out on a table. 
 
 In retorting amalgam containing considerable quantities of 
 base materials, there is a danger of the vent being choked up by 
 condensation of solid material. The retort should be so arranged 
 
142 THE METALLURGY OF GOLD. 
 
 that a rod can be passed through the condensing pipe so as to 
 clear it of obstructions, if necessary. The front of the retort is 
 luted on carefully with chalk or wood-ashes and salt, and firmly 
 clamped so as to be quite tight, otherwise a loss of mercury 
 is incurred. In all retorts the lid is turned and ground so as to 
 fit on perfectly. The condensing pipe should not have an open 
 end dipping freely into water, as in that case a sudden cooling 
 of the retort would cause the water to be sucked in, and an 
 explosion would occur. The open end of the pipe may be 
 enclosed in a rubber bag immersed in water, in which case the 
 . amount of distention of the bag will indicate the progress of the 
 operation. 
 
 The pot-shaped retort requires no brick fittings, and can be 
 heated over an assay furnace or forge fire, or in a fire built on 
 the ground, when it is placed on a tripod stand. In the latter 
 case the fire is lit at the top and burns slowly downwards. The 
 pot-shaped retort is not filled to more than two- thirds its capacity, 
 and must be heated very gradually at first. 
 
 The retorted metal is porous and from 500 to 950 fine in gold, 
 the remainder being in general chiefly silver, with base metals 
 and sulphides in smaller quantities. The gold is melted in 
 crucibles with carbonate of soda and borax, and suffers a further 
 loss in weight, due to the volatilisation of a small quantity of 
 mercury, which is obstinately retained until the melting takes 
 place. This amount is usually not more than from 0*5 to 
 1 part per 1,000 of the retorted metal, if the retorting has been 
 carefully performed. 
 
 Loss of Mercury. The loss of mercury is due to two causes, 
 "flouring," or minute mechanical sub-division, due to excessive 
 stamping or grinding, and "sickening," or extreme sub-division 
 caused by chemical means. In the latter case, a coating of some 
 impurity is formed over the minute globules of mercury, which 
 are thereby prevented from coalescing, from taking up gold 
 and silver, or from being caught by the plates and wells, as the 
 coating prevents all contact between the mercury and other 
 bodies. The impurity may be an oxide, sulphate, sulphide, or 
 arsenide of some base metal, either originally present in the 
 mercury, or taken up from the ore by it ; occasionally the 
 mercury itself may be partly converted into a sulphate or other 
 salt, although this latter condition is not common. The employ- 
 ment of mercury, which contains no base metals dissolved in it, 
 will reduce the loss due to sickening, but such pure mercury 
 is not always obtainable (except by very careful distillation, in 
 which the first and last portions condensed are rejected), and it 
 soon takes up fresh impurities when used with sulphuretted 
 ores. The base metals usually present in mercury are rapidly 
 oxidised in the air, especially in contact with water ; the oxida- 
 tion is made much more rapid by the presence of any acid in the 
 
AMALGAMATION IN THE STAMP BvTTERY. 8 143 
 
 water, and this acidity (due to the presence of acid sulphates 
 from decomposing pyrites) is rarely quite absent from battery and 
 mine waters. The metallic oxides thus formed are not soluble in 
 mercury, and they float on its surface in the form of little black 
 scales, which soon form a coating. A surface coating may also 
 be formed of grease. The use of sodium amalgam in preventing 
 the sickening of mercury is noticed below (p. 144). One of the 
 impurities in mercury most to be feared is lead, as the amalgam of 
 this metal tends to separate out of the bath of mercury in which 
 it is dissolved. According to Prof. J. Cosmo ISTewbery, it rises 
 to the surface by degrees, taking with it any gold amalgam that 
 may have been formed, and floats as a frothy scum, coating the 
 mercury and preventing any further action by it, whilst it is 
 readily powdered and carried away in suspension by a current of 
 water flowing over it, so that the gold contained in it is lost. 
 
 Sickening of the mercury is also promoted by base minerals 
 present in the ore. Most gangues, except heavy spar, hydrous 
 silicates, &c., have no action on the quicksilver ; even clean 
 cubical iron pyrites, and other iron and copper pyrites, if 
 they are undecomposed, are harmless, although the materials 
 last named cause sickening when partly decomposed. The 
 other sulphides are all more or less harmful, their action 
 being, however, much less energetic than the compounds of 
 arsenic and antimony. J. Cosmo Newbery conducted a number 
 of experiments in Australia many years ago, to determine the 
 action of some of the base metals on mercury, and found that 
 compounds of arsenic and antimony are particularly harmful, and 
 that if gold containing metallic arsenic is amalgamated, the result- 
 ing amalgam is black and powdery, and floats on the mercury, 
 being coated by black metallic arsenic, which separates out and 
 refuses to unite with mercury. Arsenical pyrites seems to act 
 in the same way as metallic arsenic, a large amount of black 
 sickened mercury being produced by it, the action being especially 
 energetic if the pyrites is partly decomposed. The black coating 
 is, in this case, a mixture of pyrites, arsenic, and mercury, in a 
 very finely divided state. Sodium amalgam acts beneficially 
 when arsenic is causing loss of mercury. 
 
 Sulphide of antimony breaks the mercury into black powder 
 even more quickly than arsenic, some sulphide of mercury being 
 formed if there is any trituration, whilst the antimony forms an 
 amalgam. The action of sodium amalgam on this mixture is of 
 no avail, as sodic sulphide is formed, more antimony amalgam 
 produced, and sulphuretted hydrogen set free, the results on 
 the amalgamation of the gold being very disastrous. Bismuth 
 sulphide acts similarly, but with less rapidity. 
 
 Floured mercury is perfectly white in appearance, like flour, 
 sickened mercury, as already stated, being blackish. If this 
 floured mercury is examined with a lens, it is seen to consist of 
 
144 THE METALLURGY OF GOLD. 
 
 a number of minute particles many of them microscopic each 
 of which is perfectly bright and pure, shining like a mirror. 
 They are prevented from coming into contact and coalescing 
 by being surrounded by films of air, concerning which Prof 
 Huntington has remarked " If you scratch the film of air, the 
 particles run together and form one globule." Floured mercury 
 is readily carried away and lost in the tailings, but if passed 
 through and agitated with a large body of clean mercury much 
 of it is at once absorbed in the mass. The loss through n curing 
 is experienced in the milling of refractory and free-milling ores 
 alike. 
 
 In California the total loss of mercury varies from i to 1 oz. of 
 mercury per ton of ore crushed, the mean being about J oz. per 
 ton. Most of the mercury is lost as such and not in the form 
 of amalgam, as is proved by the fact that where the largest 
 proportion of mercury is fed into the battery the greatest loss 
 takes place but the highest percentage of gold is recovered. 
 Thus, at the North Star and Empire Mills the greatest loss in the 
 State occurs, 1 oz. of mercury being lost per ton, but over 90 per 
 cent, of the gold is extracted. In the Blackhawk Mills, Colorado, 
 where base ores are crushed, containing from 12 to 20 percent, of 
 pyrites, the loss of mercury is from i to | oz. of mercury per ton. 
 The mills in which the greatest loss of mercury occurs have the 
 deepest discharge, the ore and mercury being in these cases 
 pounded together for a greater length of time before being 
 ejected from the mortar, so that more flouring takes place. 
 
 Many suggestions have been made at various times for the 
 reduction of the loss of mercury. The use of certain chemicals 
 in keeping it clean and lively and in neutralising the bad effect of 
 base minerals has already been noticed. Sodium amalgam was. 
 su ggested by Sir Wm. Crookes, F.R.S., many years ago, as 
 likely to effect this purpose. It is prepared by heating a basin 
 or iron flask of mercury to about 300 F., and dropping in little 
 pieces of sodium, not larger than a pea, one by one. Each 
 addition causes a slight explosion and a bright flash of flame. 
 The sodium may be added with less loss and less danger to the 
 operator if the mercury is kept at a somewhat lower temperature 
 and the sodium stirred into the mercury with an iron pestle or 
 pressed below its surface with a spatula. When about 3 per 
 cent, of sodium has been added to the mercury, the reaction 
 becomes less active and the amalgam is then poured out upon d 
 slab or shallow dish, allowed to cool and solidify, and then 
 broken up and kept in stoppered bottles under naphtha. When 
 it is necessary to revivify a quantity of mercury, a few small 
 pieces of the amalgam are added to it and stirred in, or are 
 previously dissolved in clean mercury before being added to the 
 impure stuff. The strong affinity of sodium for oxygen enables 
 it to reduce the oxides of the base metals which are coating the 
 
AMALGAMATION IN THE STAMP BATTERY. 145 
 
 mercury globules, and as sodic oxide is very soluble in water it 
 is at once removed in solution, neutralising part of the acidity of 
 the water at the same time. The base metals are redissolved by 
 the mercury which is then in good condition to take up the 
 precious metals or to be caught on amalgamated surfaces or in 
 riffles. Sodium amalgam is not much used except in amal- 
 gamating-pans or in mercury wells or riffles i.e., wherever large 
 bodies of mercury can be directly acted on by it. It is of 
 comparatively small value when added to the mortar of a 
 stamp battery, although this use of it is not unknown. 
 
 Loss of Gold. The losses of gold in amalgamation may be 
 ranged under the following heads : 
 
 1. Loss of free gold or amalgam due to a want of proper care 
 in amalgamation. 
 
 2. Loss of gold or sulphurets imbedded in particles of rock. 
 
 3. Loss of gold which "floats" in water and is carried away 
 with the slimes. 
 
 4. Loss of gold which is not in a condition to be directly 
 amalgamated. 
 
 The latter heading may be subdivided into two, viz. : 
 
 (a) Loss of gold contained in sulphurets. 
 
 (b) Loss of free gold, which is prevented from being amal- 
 gamated by being coated with a film of some mineral ("rusty" 
 gold), or with grease, or by being in an unsuitable physical 
 condition (hammered gold). 
 
 In the preparation of gold ores for amalgamation every care 
 must be taken that the course best suited to each particular case 
 is being pursued. In some instances, which are not common, 
 the whole of the gold may be present in a form in which it can 
 be directly amalgamated. In general, however, the gold is pre- 
 sent in two or more forms, one capable and the others not 
 capable of amalgamation. In such cases there is no reason to 
 be dissatisfied with the action of an amalgamating machine if it 
 extracts a high percentage of the free gold, even though the 
 total extraction obtained by it is comparatively low. It has 
 frequently been declared that the greater part of the loss of 
 free gold, which really takes place, is not due so much to the 
 imperfection of the various amalgamating machines which are 
 now at the disposal of the metallurgist, as to the lack of care 
 and facility of resource on the part of the men in charge of them. 
 It is probable that no two ores can be treated to the best advan- 
 tage under exactly the same conditions. A millman experienced 
 in the treatment of the ores of one district may be quite at fault 
 when attempting to amalgamate an ore unlike those to which he 
 has been accustomed. A silver mill, in particular, has been pro- 
 nounced to be the worst possible school for a gold amalgamator, 
 whose work must be closer in proportion as his amalgam is richer 
 than that obtained from silver or^s. At the risk of recapitulating 
 
 10 
 
146 THE METALLURGY OF GOLD. 
 
 much that has been already said, it may be worth while to direct 
 the student's attention to a few general rules which can be 
 applied in many cases. In the first place the stamps and 
 screens must be such as are calculated to produce the largest 
 possible output, without rendering the pulp unsuitable for the 
 processes of amalgamation or of concentration, or both, which 
 are to follow. The ideal crushing would be to " crack the nut 
 and leave the kernel entire," or in other words, to liberate the 
 particles of gold without breaking them. It was formerly be- 
 lieved that a light stamp working rapidly was best adapted for 
 this, and even quite recently light fast stamps, with narrow 
 heads, only 4 inches across, have been re-introduced experimen- 
 tally. On the whole, however, the tendency is in favour of 
 heavy stamps working fast with a low drop, as producing the 
 maximum output with the minimum amount of slimes. Jn con- 
 nection with this, the small production of slimes by the steam 
 stamp (described on p. 156) should be noted. 
 
 The subject of delivery is closely connected with that of 
 crushing and must be considered at the same time. The screens 
 are not usually placed quite close to the level of the pulp in the 
 mortar on account of the rapid wear caused by the violent pro- 
 jection of pulp against them when in that position. Their 
 height above the dies is varied according to the ore, the 
 delivery being slower in proportion to this depth of discharge. 
 The best size for the mesh of the screens must be determined by 
 direct experiment. It has often been contended that, as the 
 crushing must be fine enough to liberate the particles of free gold 
 from their matrix, therefore the size of the screens depends on 
 the state of division of the precious metal in the ore. Never- 
 theless it does not follow that an ore must be reduced so that 
 the gangue is as fine as the gold particles contained in it, 
 although this is sometimes erroneously assumed. On the con- 
 trary, even when the gold is of extreme fineness, coarse crushing 
 through 20 or 30-mesh screens, may be the best practical method 
 to adopt, since otherwise the sliming of the ore may cause the 
 loss of valuable mineral which should be obtainable by concen- 
 tration, besides reducing the yield on the plates. In the course 
 of the crushing, much of the ore will have been reduced to a 
 comparatively fine state of division, and probably this portion 
 will be found after amalgamation to contain but little gold ; from 
 this, the coarser material, in which the gold is still locked up, 
 may be separated by sizing in " pointed boxes " (see p. 179) and 
 reground in an amalgamating mill or pan. If, on the other hand, 
 the slimes are found to be as rich as the coarse sand, it is obvious 
 that no finer crushing is required, as the output would be 
 thereby diminished without any corresponding increase in the 
 yield per ton. If the slimes, after separation of the sulphides, are 
 found to contain more free gold than the coarse sand does, this 
 
AMALGAMATION IN THE STAMP BATTERY. 147 
 
 fact points to the conclusion that a coarser screen might be used 
 without detriment, and experiments in this direction should be 
 made, and the limit of economy thus found by trial. 
 
 The evils of overstamping, due to slowness of discharge, have 
 been often dwelt on, and probably are frequently exaggerated. It 
 is true that the excessive production of slime thus caused is an 
 undoubted evil, but, setting this aside, it has been frequently 
 asserted that particles of gold are reduced in size by overstamping, 
 so that they will float off in suspension in water, whilst, even 
 if not reduced in size, they are hammered and flattened so as 
 to be rendered incapable of amalgamation. Other authorities, 
 however, consider it doubtful whether much loss is due to 
 -either of these causes. Small particles of gold, it is contended, 
 are not likely to suffer further comminution when mixed with 
 la/rger particles of gangue, which would usually prevent them 
 from coming into contact with the shoes and dies ; these small 
 particles, too, would probably only receive a single blow, which 
 would throw them off the die, and a large piece of gold, which might 
 remain on the die while several blows were struck, would not be 
 easily lost even if partially subdivided. As regards hammered 
 gold, it does not seem to be beyond doubt that flattening and 
 hardening alone will prevent gold from being amalgamated. 
 Prof. T. Egleston has described a number of experiments* which 
 tend to show that amalgamation is retarded by this treatment of 
 gold, but, on attempting to repeat these experiments at the Royal 
 Mint, the author could not obtain similar results to those of Pro- 
 fessor Egleston. Pieces of pure gold when subjected to repeated 
 blows with a clean 7-lb. hammer on a clean anvil occasionally 
 showed a disinclination to amalgamate, but, if these pieces were 
 washed with dilute ammonia, so as to remove any grease that 
 might be adhering to them, they were instantly wetted by mercury 
 and were dissolved by it at about the same rate as clean annealed 
 gold. It thus appeared that, in these cases at least, grease from 
 the fingers of the operator formed a more potent preventive 
 of amalgamation than the hardness of the gold. Moreover, 
 gold-leaf which has been subjected to an extended course of 
 hammering is readily amalgamable. The matter may perhaps 
 be left as an open question for the present. 
 
 There is still much difference of opinion as to the desirability 
 of attempting to amalgamate the gold inside the battery. It is 
 considered by some authorities that the addition of mercury to 
 the mortar is a mistake, and that no copper plates should be put 
 inside. In spite of numerous practical illustrations of the 
 soundness of the contrary view, they hold that no machine can 
 be successful at once as a crusher and an amalgamator. The 
 practice of adding mercury to the mortar when no inside plates 
 
 * Metallurgy of Gold, Silver, and Mercury in the United States, 1887, 
 -vol. ii., p. 586. 
 
148 THE METALLURGY OP GOLD. 
 
 are used is certainly not now much favoured, although it is still 
 adhered to in Ballarat and some other districts in Australia. 
 In treating rich ores, however, when the gold is coarse and of high 
 standard, there does not seem to be any valid objection to be 
 raised against catching the gold on inside plates in a concen- 
 trated form, instead of letting it all go to the outside. In this 
 case mercury must be added to the mortar, and this is probably 
 not so important a cause of loss of mercury by flouring and 
 sickening as is often assumed. If a decomposing ore is mixed 
 with lime beforehand, and the acidity of the water used is 
 corrected, the conditions do not appear to be favourable to the 
 production of salts injurious to the mercury, and the latter when 
 charged in is probably almost instantly washed through the 
 screens or else dashed against and retained by the plates, the 
 condition of which is thereby improved by the softening of the- 
 amalgam, so that gold is more readily caught on it. The mercury 
 does not remain on the die, subjected to repeated blows, which 
 would no doubt cause much flouring. Finely divided gold, how- 
 ever, particularly if it is of low standard, containing much silver, 
 requires more mercury for its amalgamation than coarse gold, and 
 in this case it is difficult to keep the plates in good order, so that 
 it is usually advantageous to save the extra trouble and labour, 
 caused by looking after and cleaning-up inside plates, by putting 
 all the plates outside. This has been the experience in a number 
 of mills, including that of the Montana Company, where the 
 inside plates have been entirely discarded. 
 
 Amalgamation outside the battery has also been the subject 
 of much discussion and some careful investigation. Numerous 
 amalgamating machines have been patented, the inventors in 
 every case praising their own contrivances and decrying the 
 copper plate, but the latter has not as yet been superseded, and 
 its principle is applied to almost all its more promising rivals. 
 
 To secure successful amalgamation it is necessary that the 
 particles of gold should be brought into absolute contact with the 
 mercury. This contact is obtained in one of three ways, viz. : 
 
 1. The mercury and ore are ground together in pans, arrastras 
 and similar machines, contact being secured by pressing the gold 
 and mercury together. 
 
 2. The ore is allowed to flow over or even through a bath of 
 liquid mercury, or the endeavour is made to ensure contact by 
 letting the pulp fall from a height upon the mercury. 
 
 3. The ore is allowed to flow over either stationary or shak- 
 ing amalgamated copper plates, drops being introduced between 
 the plates to break up the pulp and to assist in catching the 
 amalgam. 
 
 The first method is undoubtedly the best for ensuring contact, 
 but the operation is tedious and, in most cases, unnecessary. 
 Opinions are divided as to the relative merits of the two latter 
 
AMALGAMATION IN THE STAMP BATTERY. 
 
 methods, but the great majority of metallurgists are advocates 
 of the superiority of the plates. They point out that in a mer- 
 cury bath, in spite of the first impression to the contrary felt 
 by every one on approaching the subject, contact with the ore is 
 very difficult to obtain. If wet pulp is introduced at the bottom 
 of a bath of mercury, it rises to the surface in lumps, surrounded 
 by films of water, and dry pulp is still more effectually pro- 
 tected from contact with the mercury by films of air. If a thin 
 stream of pulp is run over the surface of a bath of mercury of 
 sufficient size, the chances of the particles of gold coming in 
 contact with the quicksilver (by settling through the stream) 
 are greatly improved, but in this case the more convenient 
 copper plate could be substituted for the bath. Moreover, it is 
 well known that rich gold or silver amalgam catches gold more 
 readily than pure mercury, and, whilst the surface of the plates 
 can readily be covered with such pasty amalgam, it would 
 involve a large and unnecessary sinking of capital to keep any 
 considerable percentage of precious metals in the baths. For 
 these reasons, and for the practical reason that plates are found 
 to work better than baths, the use of the latter has been 
 gradually more and more restricted, until they are now only to 
 be found in narrow wells and riffles for the purpose of catching 
 hard amalgam and floured mercury, and as purely supplementary 
 aids to the plates, whilst in the most modern mills they are 
 often dispensed with altogether. When a proper disposition of 
 the plates is made, it is rare to find amalgamable gold escaping 
 into the tailings. Even if the latter contain several penny- 
 weights of gold to the ton, this does not show that the 
 amalgamation effected on the plates is unsatisfactory, and, where 
 tailings have to be reground or roasted, or treated in any way 
 that may alter the condition of the gold, before a further extrac- 
 tion by mercury is obtained, no proof is afforded that the plates 
 are not doing their work. 
 
 The loss of finely divided or " float " gold, particularly when 
 it cannot be checked by the use of swinging plates, is often 
 another name for the loss of slimed sulphides. Many examples 
 have been adduced by the vendors of patent amalgamating or 
 chlorinating machinery of the large percentage of the gold in 
 the ores crushed in particular mills, which has been carried 
 away suspended in water in a form not easily recoverable by 
 settling. In the majority of these cases, however, no attempt 
 seems to have been made to distinguish between the values 
 contained in slimed sulphides and those existing as particles 
 of free gold. Where this is not done there are no grounds for 
 the assumption that any free gold is escaping at all. Thus 
 G. M'Dougal, of Grass Valley, California, found* that a gallon 
 of water in the stream, J mile below two mills, contained on 
 * Mines West of the Rocky Mountains, by R. W. Raymond. 
 
150 THE METALLURGY OF GOLD. 
 
 an average 1-18 cents worth of gold. He called this " float" gold,, 
 but did not try to find out its physical condition, and it was- 
 very likely contained in sulphides. Again at the Spring Gully 
 Mine, in Queensland, the tailings from the battery, if settled in 
 the ordinary way by running off the water, were found to con- 
 tain 7 dwts. of gold per ton, but, if carefully filtered, assayed 15 
 dwts. All such examples prove only that the slimes are rich, 
 not that " float " gold is being lost, and although it is of course 
 likely that some finely divided gold is carried away in suspen- 
 sion in water during the treatment of many ores, nevertheless, 
 if sufficient care were taken in ascertaining this loss, it would 
 probably prove to be less than is generally believed. 
 
 The following scheme of examining tailings with a view 
 to determine the causes and amount of loss is given by 
 M'Dermott & Duffield* : 
 
 Small samples are taken at intervals from the waste outflow 
 of the mill, until a bucketful is collected ; this is allowed to 
 settle for several hours, the clear water is decanted, preferably 
 through a filter, and the remainder evaporated to dryness. Care 
 must be taken to avoid spilling anything out of the vessels 
 containing the samples. The sample having been well mixed, 
 portions are treated as follows 
 
 A. One part is panned and examined for free gold, amalgam, 
 and quicksilver. If these are present, it is probably the fault of 
 the millman, and nothing further need be done until this state of 
 things is remedied. 
 
 B. The tailings are sized on brass screens, and the coarse, 
 medium, and tine materials (the latter consisting, say, of that 
 portion which passes a 100-mesh screen) are weighed and assayed 
 separately, the coarser portions being reground and panned to 
 find whether their values are in free gold or in sulphides. 
 
 C. The sulphides are separated from the tailings on a vanning 
 shovel or batea, and are weighed and assayed. It may thus be 
 determined whether they are worth saving, and the size of the 
 mesh used for the screens will depend largely on this, and on 
 the nature of the sulphides, which will in many cases be badly 
 slimed and difficult to catch if the ore is finely crushed. 
 
 D. The loss due to fine or "float" gold maybe determined 
 by assaying the slimes after the sulphides have been carefully 
 removed by concentration. This requires much skill and patience, 
 but can in almost all cases be successfully accomplished by the 
 vanning shovel. The concentrates may be examined under the 
 microscope for fine specks of gold, but these and the fine sulphides 
 can be recovered by concentration on suitable machinery. The 
 assay value of the tailings from the vanning shovel will give 
 some idea of the amount of float gold which is being lost. It 
 will usually be found to be smaller than may be expected. If it 
 
 * Gold Amalgamation, London and New York, 1890, p. 7. 
 
AMALGAMATION IN THE STAMP BATTERY 151 
 
 is large, the use of some system of amalgamation more perfect 
 than that by copper plates (such as pan-amalgamation) or of a 
 method of smelting, or of a wet method, may be considered, if 
 the advantages appear sufficient to pay for the presumably 
 increased cost. 
 
 Non-Amalgamable Gold. The appearance in the tailings 
 of free gold, which is not especially finely divided, but, neverthe- 
 less, is not in a condition to be amalgamated, may be regarded 
 as a rare occurrence, but deserves some consideration. Amal- 
 gamation is in these cases prevented by the existence of a 
 thin film of some neutral substance over the surface of the 
 gold. The film may be so thin as to be transparent, but it is 
 enough to prevent contact between the gold and the mercury. 
 The disastrous effect of a film of grease covering gold particles 
 has already been remarked upon. It is said to have been a 
 fruitful source of loss in the treatment of certain ores in the 
 Transvaal that they were impregnated with mineral oil. The 
 effects of grease may be combated by the use of chemicals (caustic 
 alkalies, potassic cyanide, &c.), but it is, of course, better to use 
 every precaution to avoid the introduction into the pulp of oil 
 from bearings, guides, &c., or contained in steam from the boiler. 
 Losses in amalgamation are also caused by the greasy sub- 
 stances contained in some ores, such as the powdered hydrated 
 silicates of magnesia and of alumina, which cause frothing, and 
 coat the gold with a slime which prevents the action of the 
 mercury. 
 
 Other films are formed of oxide of iron, compounds of sulphur, 
 arsenic, &c., or of silica. Some years ago J. Hankey, of San 
 Francisco, had a collection of particles of native gold which 
 appeared as bright and lustrous as usual, but were coated by 
 thin translucent films of red oxide of iron. These particles of 
 " rusty " gold could not be wetted by mercury, but, if a piece 
 were snipped off one end, the mercury seized on the fractured 
 surface at once. Such gold seems to be rare in nature. 
 
 In 18b7, William Skey, of the Geological Survey of New 
 Zealand, after a series of experiments on the ores and tailings of 
 the Thames Valley, came to the conclusion that the bright gold 
 particles which refused to amalgamate were always coated by 
 some compound of sulphur. He found that gold takes up 
 sulphur from sulphides of ammonium or sodium, or from sul- 
 phuretted hydrogen, when brought in contact with their solu- 
 tions, and that after this the gold refuses to amalgamate. He 
 supposed that these compounds of sulphur were often formed by 
 the action of acidulated water on the minerals in ores, and that 
 consequently " a large area of the natural surfaces of native gold 
 is covered with a thin film of auriferous sulphide, and that the 
 greater part of the gold which escapes amalgamation at the 
 battery consists of this sulphurised gold." 
 
152 THE METALLURGY OP GOLD. 
 
 Gold in Pyrites. In the preceding pages no account has 
 been taken of the loss of gold which is contained in pyrites, as 
 it has been assumed that these are saved by concentration if 
 they are valuable, and this subject is dealt with in Chapter ix. 
 Nevertheless, as this gold comes under the head of non-amal- 
 gamable gold, its physical state and the causes of its disinclination 
 to unite with mercury may conveniently be considered here. In 
 general, pyrites yields an extremely low percentage of its gold 
 contents if it is run over the amalgamated plates, and if it is 
 ground very fine in a pan with mercury the percentage extraction 
 is better. However, in general, pyrites, even if it is reduced to 
 the very finest slimes, and its prolonged contact with mercury 
 ensured by continuous grinding, cannot be made to yield more 
 than about 40 per cent, of its gold, and this is at the cost of much 
 mercury, which is floured or sickened during the process. Iso- 
 lated instances of better results are on record, but must be 
 regarded as exceptional cases. Among the old processes used 
 for the amalgamation of the gold in pyrites may be mentioned 
 the treatment in revolving wooden barrels with mercury, as 
 practised at the St. John del Rey Mine, and the practice of 
 leaving the pyrites to be decomposed by weathering before 
 grinding it with mercury. This method of oxidation seems to 
 be decidedly inferior to the alternative plan of roasting the 
 sulphides, by which the oxidation is rendered more complete 
 and the particles of gold agglomerated to some extent. However, 
 the amalgamation of pyrites, even when roasted, is far from 
 perfect, part of the gold still remaining in a condition unfit for 
 extraction in this way. The ores, which have been met with in 
 various parts of the world, consisting mainly of limonite or 
 hydrated oxide of iron, and in most cases believed to be the 
 result of decomposition of pyritic ores by atmospheric agencies, 
 are also extremely refractory, causing the mercury to sicken 
 rapidly, and yielding only about the same percentage of gold as 
 can be obtained from unoxidised pyrites. 
 
 The most celebrated case of this kind is that of the Mount 
 Morgan ore, in Queensland, which is an ironstone gossan con- 
 sisting of silicious brown iron ore, derived according to one view 
 from the decomposition of pyrites. Although the gold appears 
 to be free, it cannot be amalgamated, yielding only about 30 per 
 cent, when crushed in batteries and subjected to prolonged 
 grinding in pans with mercury. When the ore is dehydrated by 
 roasting in reverberatory furnaces the extremely fine particles of 
 gold are agglomerated, and between 80 and 90 per cent, can then 
 be extracted by amalgamation, the remainder being presumably 
 coated by oxides of iron. The richness of the ore, however, 
 makes even this result unsatisfactory, and a process of chlorina- 
 tion has been adopted in practice. A similar case was noticed 
 by Mr. Mactear * in South America, where a limonite ore which 
 * Mining Journal, Jan. 24, 1893, p. 70. 
 
AMALGAMATION IN THE STAMP BATTERY. 153 
 
 only yielded from 35 to 40 per cent, of its gold when treated in 
 Huntington pans, was made to yield between 85 and 90 per 
 cent, by merely subjecting it to a dehydrating calcination before 
 amalgamating it. Louis Janin, Jr., mentions another case* 
 in the ores of the Southern Cross Mine, Deer Lodge County, 
 Montana, which consist of limonite derived from the alteration 
 of pyrites. In panning large samples, only one or two specks of 
 gold could be seen, although the ore contained from 1 to 2 ounces 
 per ton. This ore yielded only about 40 per cent, on being 
 amalgamated, but over 90 per cent, was dissolved out by leach- 
 ing the raw ore with cyanide of potassium, and similar results 
 were obtained by chlorination. Here the ore was thoroughly 
 decomposed, but yet the gold would not amalgamate to a much 
 greater extent than if it were still contained in the original 
 pyrites, whilst the chemicals at once dissolved it. 
 
 In seeking to explain the behaviour of gold in pyrites, various 
 theories have been propounded. According to one of these, the 
 gold is supposed to exist in the pyrites in the form of sulphide 
 combined with sulphides of iron, silver, copper, <fec., and the 
 refusal of the gold to amalgamate is explained in this way, auric 
 sulphide not being acted on by mercury. Some observers have 
 endeavoured to dissolve gold out of pyrites by the action of 
 alkaline sulphides, and when, after many attempts, this was at 
 length successfully accomplished, it was put forward as additional 
 evidence that the gold must have been in the state of sulphide, 
 although metallic gold is known to be soluble in these menstrua. 
 
 The balance of evidence, however, seems to be in favour of the 
 theory that gold, at any rate in great part, exists in pyrites in 
 the metallic state. Although the metal is generally invisible in 
 un decomposed crystals of pyrites, it becomes visible when such 
 crystals are oxidised either by air and water in nature, or by 
 means of nitric acid, or by being roasted or subjected to 
 deflagration with nitre. As a result of such decomposition, 
 particles of bright lustrous gold, angular and ragged in shape, 
 but of considerable size, often become apparent. These particles 
 may be separated from the oxides of iron by washing, and the use 
 of nitric acid, followed by panning, is frequently resorted to in 
 order to detect gold in pyrites. Moreover, although usually 
 invisible, gold can sometimes be seen in unroasted pyrites. As 
 long ago as the year 1874, Richard Daintree and Latta found 
 specimens of cubical pyrites,! in which gold could be seen 
 under a microscope gilding the cleavage planes of the crystals. 
 Again, G. Melville-Attwood, on examining crystals of auriferous 
 pyrites from California in 1881,| found that the faces of the 
 crystals were gilded in some places, and that here and there 
 
 * Mineral Industry for 1892, p. 249. 
 
 t Proc. Royal #or. of New South Wales, 1874; paper on " Iron Pyrites." 
 
 J Precious Metals in the United States, 1881, p. 604. 
 
154 THE METALLURGY OP GOLD. 
 
 little specks or drops of gold occurred, partially imbedded in the 
 pyrites. These films were too thin to be detected by an ordinary 
 lens, so that it did not seem surprising that such impalpable 
 material could not be taken up by mercury. Louis Janin, 
 Jr., more recently* found crystals of pyrites in a porphyritic 
 gangue from the Republic of Colombia, which had gold in small 
 globules on their surfaces. Lastly, it has long been known that 
 crystals of pyrites are often found adhering to an amalgamated 
 plate, the particles of gold on their surfaces having been amalga- 
 mated. It seems likely, in view of all these facts, that some of 
 the gold at any rate is in the metallic state, and its refusal to 
 amalgamate is not very surprising, when it is remembered how 
 completely a thin coating of certain sulphurised compounds 
 prevents amalgamation, and how readily sulphuretted hydrogen 
 would be evolved from decomposing pyrites, Some authorities 
 have contended that the metallic gold is disseminated mechani- 
 cally through the mass of pyrites, but the action of potassic 
 cyanide, in dissolving the whole of the gold out of comparatively 
 coarsely crushed pyrites, seems to point to the correctness of the 
 view that the interior of the crystal is not auriferous, the 
 deposition of the gold being superficial, so that the enrichment 
 of the pyrites is confined to its crystalline faces, and possibly, but 
 not probably, to its cleavage planes. 
 
 The following details f of a microscopical examination by Prof. 
 Morton, of the condition in which pyrites is left after being 
 leached with cyanide, confirms this view to some extent : 
 
 " Upon the ordinary auriferous sulphide of iron, or arsenical 
 pyrites, the solution of potassium cyanide acts readily, not by 
 dissolving the sulphuret, but by attacking the gold upon its 
 exposed edges, and eating its way into the cubes by a slow 
 advance, dissolving out the gold as it goes. An examination 
 with the microscope of the pyrites after the gold has been 
 removed, suggests the method of the operation. A sample of 
 very rich pyrites, from a mine north of Bedding, was treated 
 with a weak solution, containing less than two-tenths of 1 per 
 cent, of cyanide, for 168 hours; the assay showed a complete 
 extraction of the gold ; as the sulphurets showed no change in 
 their appearance to the naked eye, some of them were placed 
 under the microscope. 
 
 " There is no change visible in the form of the crystals as a 
 whole ; along the fractured faces the mispickel looks clean and 
 unaltered, showing the silvery- white colour and intense refraction 
 of the arseno-pyrite. Upon the faces of the crystals appear dark 
 lines, short, and parallel to each other. In places they are 
 crowded close together ; in other parts they are at considerable 
 distances, but always in parallel lines. The lines vary in length, 
 being from four or five to over a hundred times their width ; the 
 * Mineral Industry, 1892, p. 249. t Loc. cit. 
 
CRUSHING AND AMALGAMATING MACHINERY. 155 
 
 lines are very irregular and often broken. These lines are fissures 
 in the pyrites, and extend so deep into it that the microscope does 
 not reveal their depth. By using the higher powers the walls of 
 one of the fissures were seen to be completely honeycombed, 
 looking somewhat like two empty honeycombs set opposite each 
 other ; evidently the mineral removed was crystallised along its 
 contact walls at least. As the raw or untreated pyrites does not 
 show any such fissuring, but, upon the contrary, shows a surface 
 marked only by striation lines common to pyrites, I assume that 
 the fissuring in the treated sample is caused by the solution 
 acting upon some soluble mineral, probably gold, arranged in 
 plates, occurring in groups, but which, by its colour and iso- 
 morphism and the extreme tenuity of its lines, is undistinguish- 
 able from the mass of pyrites enclosing it." 
 
 CHAPTER VIII. 
 
 OTHER FORMS OF CRUSHING AND AMALGAMATING 
 MACHINERY. 
 
 Special Forms of Stamps. Since within certain limits and 
 under certain conditions the capacity of a stamp battery depends 
 on the number of blows given per minute and on the momentum 
 of the fall, various contrivances have been suggested with a view 
 to increase both of these. Among these special stamps Husband's 
 pneumatic stamp, the Ball Steam stamp, and the Elephant stamp 
 may be noticed. 
 
 Husband's Pneumatic Stamp consists usually of two stamps, the 
 stems of which are attached to pistons of small diameter working 
 in pneumatic cylinders, I) (Fig. 32). These cylinders have a reci- 
 procating up and down motion given to them by the revolution of 
 a crank shaft, C, above them, with which they are connected by 
 iron rods, E, affixed to trunnions, cZ 1 , on the cylinders. When a 
 cylinder is raised by the crank shaft, the air in it, below the 
 piston, is compressed and the stamp stem thus lifted. Similarly, 
 the downward motion of the cylinder causes a compression of the 
 air above the piston, which urges the stamp downwards with 
 greater velocity than it would have by virtue of its weight alone. 
 There is also a contrivance for rotating the stamps so as to give 
 even wearing of shoes and dies. These stamps have not been 
 much used except on tin-ores in Cornwall. Their output is from 
 20 to 30 tons per head per day through a 36-mesh screen, the 
 power required being about 20 to 25 H.P. per head. One of 
 
156 
 
 THE METALLURGY OF GOLD. 
 
 the chief difficulties encountered in attempting fine crushing 
 with these stamps of enormous capacity is the effect on the 
 screens. Wire screens do not stand the excessive impact in a 
 satisfactory manner, while Russia-iron screens, if punched for 
 
 HARVEY'S PATENT HUSBAND'S STAMPS 
 
 Seal* Yi in<*.-l foot. 
 
 Fig. 32. 
 
 fine crushing, are more effective when thin, and are then of little 
 durability. 
 
 Steam Stamps. The ordinary form of steam stamp consists of 
 a direct-acting vertical engine, having a steam cylinder and 
 
CRUSHING AND AMALGAMATING MACHINERY 
 
 157 
 
 slide valve at the top, the piston-rod being rigidly connected 
 with the stamp. Each stamp head works in a separate rectan- 
 gular cast-iron mortar, with screens both at the front and the 
 back, and sometimes all round. The screens are of Russia sheet- 
 iron with punched holes of about -J- to -^ inch in diameter, the 
 steam stamp being best adapted for coarse crushing. The speed 
 
 Fig. 33. 
 
 of working is from 90 to 100 blows per minute, and the out-turn* 
 is from 100 tons to as much as 225 tons of ore per head in twenty- 
 four hours. It is obvious that gold could not be economically 
 saved on plates inside the mortar of one of these stamps, aud, as 
 a matter of fact, until recently they were only employed in 
 coarsely crushing the copper ores of the Lake Superior region. 
 Nevertheless, as the curious fact seems to be well established 
 that these stamps, with their heavy blow, do not make so much 
 slimes as the ordinary gravitation stamp, it is to be expected that 
 
158 THE METALLURGY OF GOLD. 
 
 they will be largely used in the future for crushing before con. 
 centration. They have been tried in crushing silver ores in 
 Montana, and gold ores at the Homestake mine, in the Black 
 Hills, through a 30-mesh screen. As their capacity is so great 
 their use is limited to cases in which large quantities of ore are 
 available, one Ball steam stamp, such as is used in the Lake 
 Superior district, being equal to at least fifty head of gravitation 
 stamps. 
 
 The chief advantages of the steam stamp are economy of space 
 and labour. With gravitation stamps more work is thrown on 
 the rock-breaker, but it is doubtful whether there is any waste 
 of power, as the whole of the work can be done by means of a 
 single steam engine. The advantage of subdividing the work 
 among a number of batteries is that stoppages for repairs and 
 breakages affect only a small part of the crushing capacity at one 
 time. Also, it has been pointed out that water-power can be 
 directly applied to gravitation stamps with little loss of efficiency, 
 whilst it is more difficult and wasteful to apply water-power to 
 compress air for use in place of steam in the Ball stamp. 
 
 TJie Elephant Stamp consists of a bent or compound lever of 
 hammered-iron or steel (a, a, Fig. 33), one end of which is 
 pivoted on a fulcrum-pin, 6, and the other end is fitted with 
 stamp shoes. The figure represents a battery of two stamps. 
 The levers are connected with the crank-shaft, c, by strong semi- 
 circular springs, c?, secured to short connecting rods, e. The 
 stamp heads work in a mortar, f, fitted with dies in the usual 
 way, and having discharge screens on three sides, and coarsely 
 crushed quartz is fed in through an inclined shoot at the back. 
 The action of the bow-springs, d, introduced between the crank 
 shaft and the lever, is to receive and store up the force of the 
 recoil from the blows of the stamps, and by giving it out at the 
 next descent of the crank to effect a certain saving of power ; in 
 addition to this the springs act as cushions, taking up the jarring 
 effect of the blows, and so diminishing the wear and tear of the 
 machine. A mill, consisting of two heads of Elephant stamps, is 
 said to be capable of stamping from 12 to 15 tons of hard gold- 
 quartz in twenty-four hours through a 30-mesh screen. The 
 advantages over gravitation stamps consist in the lightness, 
 cheapness, ease of transportation and erection of Elephant 
 stamps, and the driving power required, and they are conse- 
 quently especially suitable for prospecting purposes, and in pre- 
 liminary work during the development of new mines. It was 
 found, however, in India that the lateral wear and tear at the 
 top of the mortar-box by the levers which carried the crushing 
 heads was very great, and frequent renewals were necessary. 
 The constant position of the heads, moreover, makes the wear 
 very uneven, and the machines require much oil, whilst their 
 capacity rapidly falls off" as they are being used. 
 
 Morison's high-speed stamp* differs from the gravitation 
 * Proc. N.E.C.I. ofE. and S. 1896. 
 
CRUSHING AND AMALGAMATING MACHINERY. 
 
 159 
 
 stamp only in the absence of the cam and ordinary tappet. A 
 modified tappet is enclosed in a tight-fitting cylinder, and is 
 raised and lowered by the pressure of water in the cylinder. 
 The number of drops in a minute is increased by 25 or 30 per 
 cent. (See reference on p. 116.) 
 
 The Huntington Mill. The Huntington Roller Mill, having 
 now been in steady use for several years in the United States 
 and elsewhere, has conclusively proved its claim to rank with 
 stamps as a practical machine for fine crushing. It consists of 
 an iron pan, at the top of which a ring, B (Fig. 34), is set, and 
 attached to this are three stems, D, each of which has a steel 
 shoe, E, fastened to it. The stems are suspended from the ring 
 and are free to swing in a radial direction, as well as to rotate 
 round their own axes, whilst the whole ring, B, with the stems 
 
 Fig. 34. 
 
 and shoes, revolves round the central shaft, G. The shoes or 
 rollers, as they are called, are thus driven outwards by centri- 
 fugal force and press against the replaceable ring die, C. The 
 most modern yoke for the suspension of the rollers is shown in 
 Fig. 35.* In front of each roller is a scraper, F, which keeps 
 the ore from packing. The rollers are suspended with their 
 bases at the distance of 1 inch from the bottom of the pan, 
 which can also be replaced when worn. The lowest part of the 
 screen is situated a little above the top of the rollers, and out- 
 side it there is a deep gutter into which the ore is discharged, 
 
 * This figure is taken from Mr. A. Harper Curtis' s paper on " Gold 
 Quartz Reduction." Proc. Civil Engineers, vol. cviii., part ii., 1892. 
 
160 
 
 THE METALLURGY OF GOLD. 
 
 and from which a passage leads to the copper-plate tables. The 
 ore and water being fed into the mill through the hopper, A, 
 generally by an automatic feeder, the rotating rollers and the 
 scrapers throw the ore against the sides where it is crushed to 
 any required degree of fineness by the centrifugal force of the 
 rollers acting against the ring die. From 17 to 25 Ibs. of 
 mercury are placed in the bottom of the pan, the clearance below 
 the rollers permitting them to pass freely over the mercury 
 without coming in contact with it, so that it is not stirred up and 
 
 PI. AM. 
 
 Fig. 35. 
 
 Scale, | inch = 1 foot. 
 
 " floured," but the motion is such as to bring the pulp in contact- 
 with the quicksilver. The speed of the mill is from 45 to 75 
 revolutions per minute. The ore should be broken in rock- 
 breakers to a maximum size equal to that of a walnut, or, better 
 still, of a cobnut, before being fed in. The action of the rollers is 
 one of impact rather than of grinding, the ore being granulated 
 without the production of much slimes. The free gold, as soon 
 as it is liberated from its matrix, is in great part amalgamated 
 and retained by the mercury at the bottom of the pan, the 
 remainder being caught on the plates outside the mill. Coarse 
 
CRUSHING AND AMALGAMATING MACHINERY. 161 
 
 gold is caught inside and fine gold outside the mill, but the yield 
 inside is comparatively small when ores with high percentages of 
 sulphides are in course of treatment. 
 
 The mill is particularly adapted for the treatment of ores con- 
 taining brittle sulphides, which, if pulverised by stamps, are 
 liable to become "slimed," and so to be in an unsuitable condition 
 for concentration. It is also suitable for argillaceous quartzes, 
 which yield their gold more readily under the " puddling " action 
 of the rollers than when pounded by stamps. Moreover, the 
 Huntington mill does much more satisfactory work than stamps 
 on soft ores or in regrinding coarse tailings. The reason for this 
 lies, of course, in the relatively large amount of screen area in 
 the mill and its consequent high efficiency of discharge, a point 
 in which stamps are decidedly inferior to it. In the experi- 
 ments at the Metacom Mill, California, already quoted, it was 
 proved that stamps will take as long to pass a ton of tailings 
 through a screen as a ton of the original rock, but with the 
 Huntington mill the finer and softer the material, the more 
 rapidly it is passed through. As the splash is heavy against 
 the sides the wear of the screens is somewhat rapid, but they 
 can be very quickly replaced. 
 
 The mill is made in three sizes viz., 3^, 5, and 6 feet in 
 diameter respectively and the capacity of a 5 feet mill, the one 
 which is most commonly in use, is from 10 to 20 tons of rock 
 per day through a 30-mesh screen, the power required being 
 from 10 to 12 H.P. The weight of the ring die in the mill is 
 611 Ibs., and of each of the roller shells, which are replaceable, 
 about 140 Ibs. The wear and tear on these replaceable parts 
 is very great, amounting to about 14 oz. of steel per ton of rock 
 crushed when soft ores, previously broken small, are being treated. 
 If large pieces of hard quartz are fed into the mill or if the mill 
 is overfed, the mercury is splashed against the scrftens and 
 passes through with the pulp, and when by accident pieces of 
 iron or steel are introduced, the ring die is occasionally broken. 
 Another source of disaster in Huntington mills lies in the use 
 of acidulated water, such as that derived from mines or encoun- 
 tered when decomposing pyritic ores are treated ; the mill is 
 rapidly corroded and rendered unfit for work by such water. 
 
 The chief advantages supposed to be gained by the use of 
 Huntington mills instead of stamps may be thus epitomised: 
 
 1. JReduced First Cost. The cost for the same capacity is not 
 more than two-thirds that of stamps, even at the manufacturers' 
 shops, while the difference in favour of the mill is even more 
 in outlying districts from its light weight, and corresponding 
 low freight, and from the cheapness of its erection. 
 
 2. Saving of Power. The mill is said to run with about one- 
 half the power per ton of ore crushed. 
 
 3. Tlie wear and cost of renewals if less for the mill than for 
 
 U 
 
162 THE METALLURGY OP GOLD. 
 
 stamps, the cost being from twopence to threepence per ton of 
 ore for the former, against about fivepence or sixpence for 
 the latter. 
 
 4. There is less loss by flouring of amalgam and quicksilver, 
 while the good discharge and absence of grinding leaves the pulp 
 in a better condition for concentration. 
 
 The first three advantages appear to refer only to such soft 
 and brittle ores as are especially suited to the Huntington mill. 
 The mill requires to be set to work in an intelligent manner by 
 experienced and skilful hands, and watched carefully. The 
 dangers of over-feeding have been already alluded to. One 
 difficulty in automatic feeding is that self-feeding, such as is 
 carried on by stamps, is impossible. The automatic feeder must 
 work separately and be set to feed a certain weight of ore per 
 hour, this weight having been determined by trial. If, after 
 this, there is any change in the hardness of the rock, no automatic 
 change in the rate of feeding takes place, and the machine may 
 be choked up or run at below its maximum capacity unless 
 watched and the feeder regulated. Another difficulty is in the 
 quantity of water to be added. An excess of water, making 
 thin pulp, does not favour internal amalgamation, and it may 
 be stated that, in general, the pulp should be kept as thick as 
 possible, consistent with its prompt discharge through the screens 
 when sufficiently fine. If the pulp is too thick to run easily over 
 the copper plates outside the mill, water may be added there by 
 means of a perforated pipe. The rate of running should be as high 
 as possible, since, if the other conditions are the same, the crush- 
 ing power varies as the cube of the number of revolutions per 
 minute. 
 
 A few examples are appended of the results obtained in 
 actual practice by this excellent machine. At the Spanish 
 Mine, Nevada County, California, there are four Huntington 
 mills, three of 5 feet diameter and one of 4 feet diameter. The 
 ore is free-milling and is passed through a Blake stone-breaker 
 and thence to the mills. The four mills run at 58 revolutions 
 per minute, and pulverise 35 tons of ore each in twenty-four 
 hours, to pass through a slot screen equal to 20-mesh. The 
 pulp is passed over the usual amalgamated plates after leaving 
 the mills, and J oz. of mercury is added with each ton of ore. 
 Forty-five per cent, of the gold recovered comes from the inside 
 of the mill, where the amalgam obtained is much richer than 
 that from the plates. The loss of quicksilver is from Jg- to $y oz. 
 per ton of ore, and the total cost of milling is about one shilling 
 per ton, while for the month of November, 1887, it was only 
 tenpence per ton. The ore is a soft talcose slate, containing 
 streaks and veins of ferruginous quartz, carrying gold. The 
 chief trouble in working lies in the frequent re-adjustment of 
 feed which is found necessary. In a special test run of one 
 
CRUSHING AND AMALGAMATING MACHINERY. 163 
 
 month 42 '4 per cent, of the gold contents was extracted, and 
 the remainder lost in the tailings. This poor result was pro- 
 bably due to over-feeding, but profits were made, nevertheless, 
 although the ore yielded only a little over 1 dwt. of gold per ton 
 in 1887 and 1888. Twenty- two horse-power were used by the 
 Huntington mills in crushing from 120 to 140 tons per day. 
 
 At the Shaw Mine, El Dorado County, a 5-foot mill, making 
 50 revolutions per minute, pulverised 10 to 12 tons per day, so 
 as to pass through a 25- to 30-mesh screen. At the Mathines- 
 Oreek Mine, in the same county, a 5-foofc mill pulverised 9 to 
 10 tons per twenty-four hours, so as to pass a screen equal to a 
 40-mesh. At the Monte Ohristo Mine, Mono County, two 5-foot 
 mills, running at from 65 to 75 revolutions per minute, pulverised 
 2 tons of ore per hour to pass a screen equal to a 40-mesh ; 25 
 Ibs. of quicksilver were charged into each mill at the commence- 
 ment of the run, and about oz. more added each half hour. 
 
 Although cases have been adduced in which 90 per cent, of 
 the gold contents of an ore crushed in the Huntington mill was 
 retained inside the machine, this is decidedly exceptional, and 
 the mill is probably inferior to the stamp mill as an amalgamator 
 on many ores, although its product is better adapted for treat- 
 ment on copper plates and by concentration. 
 
 Crawford Ball MiU. This mill (see Fig. 36 *) consists of an 
 annular trough, divided concentrically by a vertical slit, which 
 passes through its deepest part all round ; the outer part of the 
 trough is fixed to the framework, whilst the inner portion, 
 having a circular spindle passing up through it, is capable of 
 being revolved by bevel gearing as in the Huntington mill. In 
 the annular trough, which is 4 feet in diameter, are placed eight 
 chilled cast-iron or hard-steel balls, each about 8 inches in dia- 
 meter : two of these are shown in position in the figure. Below 
 the vertical slit in the trough there is another vessel, in which the 
 mercury for amalgamation is placed ; this vessel communicates 
 with the trough by the narrow continuous slit passing all round 
 the machine. The ore, previously partly crushed, is fed into 
 the mill through the hopper, and the central spindle, carrying 
 with it the inner half of the annular trough, is revolved about 
 180 times per minute. This causes the 8-inch balls to revolve, 
 each on its own axis, and all of them around the annular trough, 
 and the ore is ground by them. A current of water is intro- 
 duced into the lower cavity of the machine, and passing over 
 the surface of the mercury moves up through the vertical slit 
 in the bottom of the trough, and through the ore which is in 
 process of being crushed, and overflows at the top. The direction 
 of movement of the pulp is shown by the arrows. 
 
 The idea is that, as the ore is crushed sufficiently fine, it will 
 rise and pass away with the current of water, and that the 
 particles of gold will fall down through the slit at the bottom 
 * From Proc. InsL Civil Eng., vol cviii., part ii., 1892. 
 
164 
 
 THE METALLURGY OP GOLD. 
 
 of the trough, and coming in contact with the mercury, will be 
 amalgamated there. It is claimed that the current of water 
 may be regulated so that gold only can iall through it, while- 
 not only quartz but sulphides also are carried up by it, and not 
 permitted to come in contact with the mercury, which is thereby 
 kept quite clean and preserved from agitation, so that the loss. 
 due to flouring and sickening is entirely obviated. 
 
 Pig. 36. 
 
 It is obvious that rather too much is expected from this 
 machine. Since the effect of the current of water is exactly 
 the same as that in the ascending - current classifiers (see 
 p. 179), it" is clear that " equal- falling " particles will not be 
 separated from each other by its action. Consequently, only 
 the very coarsest particles of gold, which could not escape 
 amalgamation in any machine yet devised, can be separated 
 from sulphides in this way. Moreover, it seems likely that- 
 
CRUSHING AND AMALGAMATING MACHINERY. 165 
 
 the agitation produced by the movement of the balls -would 
 occasionally permit the escape of large particles of ore before 
 they were finally comminuted, and, what is of much greater 
 importance, the rapid swirl of the water and pulp, travelling 
 in places at little less than 24 feet per second, which is the 
 speed of the periphery of the moving part of the trough, would 
 appear to effectually prevent that perfect separation by gravity 
 of heavy and light particles, which is essential to the successful 
 working of the machine. The machine is of little practical value, 
 and has only been described as a type of ball mills. 
 
 Other Ball Mills. Among other mills in which the crushing 
 is done by iron balls may be mentioned the Cyclops mill, in 
 which a single large ball, 12 inches in diameter, is carried 
 round in a vertical plane by frictional contact with a pair of 
 flexible discs fitted on a horizontal revolving shaft. The grind- 
 ing is done between the ball and a grooved circular path in a 
 vertical plane, against which it is pressed by centrifugal force. 
 The ball makes from 250 to 600 revolutions per minute, and the 
 largest sized mill is said to crush over 40 tons per day. It is 
 fitted for both wet and dry crushing, but records do not seem 
 to exist of successful work done on a large scale during any 
 considerable length of time. 
 
 Speaking generally, most of the ball machines which have 
 been devised have already been proved to be very uneconomical 
 owing chiefly to the great wear and tear of the grinding 
 surfaces, and the enormous driving power required. It is 
 difficult to obtain a ball which can be relied on to wear evenly, 
 and in many cases the effectiveness of the machine is materially 
 reduced after it has been, in operation for a very short space of 
 time. 
 
 The Tyrolean Mill. In Hungary, in Transylvania, and 
 in the Tyrol, the principle of separation of the operations of 
 crushing and amalgamation is still successfully used, the mills 
 employed being undoubtedly the progenitors of the amalgama- 
 tion pans described below (pp. 167-173). In some districts of 
 Hungary and Transylvania, Californian stamps have been re- 
 cently introduced, displacing, to some extent, the square German 
 stamps, but the amalgamation is still effected in bowls or pans, 
 which are the modern equivalents of the old Tyrolean mills still 
 to be found at work in certain retired valleys in the Eastern 
 Alps (see pp. 91 and 96). The best known of these modern 
 machines are the Schemnitz mill and the Lazzlo amalgamator. 
 
 The Schemnitz mill, shown in Fig. 36a, closely resembles the 
 ancient forms and is always used in pairs, one overflowing into 
 the other. It consists of a cast-iron bowl, A, about 2 feet in 
 diameter at the top and 18 inches at the bottom, with an internal 
 depth of 7 inches. A massive wooden nmller, B, hollowed out 
 inside in the form of a cone, is suspended in the bowl by the 
 iron rods, C, and is revolved at the rate of about twenty turns 
 
166 
 
 THE METALLURGY OF GOLD. 
 
 er minute. About 26 Ibs. of mercury are charged into the 
 wl, and the lower face of the muller is set with twenty iron 
 teeth, D, placed radially, which almost touch the mercury and 
 force the pulp (delivered into the mill from the trough, E) out- 
 wards towards the periphery of the bowl, whence it overflows 
 into the second mill.* 
 
 In the neighbourhood of Schemnitz, in Hungary, lead ores 
 containing 2 to 4 dwts. of gold per ton are crushed by stamps 
 and treated in these mills, the tailings being concentrated for 
 smelting. At Nagybanya, in Hungary, the loss of mercury in 
 these mills amounts to from J to 1 part for each part of bullion 
 
 Fig. 36a. 
 
 recovered. Thus at the Kreuzberg works, near Nagybanya, in 
 one month 155 metric tons of ore were treated, and 1*957 kilo- 
 grams (62-9 oz.) of gold were recovered with the loss of 1*9 kilo- 
 grams of mercury, and in another month, 754 '5 metric tons of 
 ore yielded 5-756 kilograms (185-8 oz.) of gold, with the loss of 
 2-9 kilograms of mercury, f The clean-up usually takes place 
 once or twice a month, and the capacity of each pair of mills is 
 from f to 1 ton of ore per day. 
 
 The Lazzlo amalgamator j has only recently come into use, 
 but has already proved itself to be of value in Transylvania. It 
 
 * TraiU de Metallurgie, by C. Schnabel, translated by Gautier, p, 738. 
 t Loc. cit. $ Ibid., p. 739. 
 
CRUSHING AND AMALGAMATING W 167 
 
 differs from the Schemnitz mill mainly in having a flat-bottomed 
 bowl, which is furnished with two circular iron partitions divid- 
 ing the bowl into three concentric compartments. The ore is fed 
 into the centre one and overflows into the others in succession, 
 and thence into a second smaller bowl. The muller is of iron, 
 and dips into each compartment of the bowl, compelling the 
 pulp to pass down and come in contact with the mercury three 
 times before it escapes, and the gold, owing to its density, does 
 not readily pass upwards and over the partitions. 
 
 At the Fiizesd Dreifaltigkeit mine, near Boicza, in Hungary, 
 the ore is crushed by Californian stamps, the capacity of which 
 is 0-8 ton per head in 24 hours, and then passed through Lazzlo 
 amalgamators. The amalgam is collected in settling pans, which 
 resemble the amalgamators, but have no iron teeth, and the 
 tailings are classified in Spitzkasten, and concentrated by passing 
 over buddies and then over canvas tables. The auriferous pyrites 
 is caught on the buddies and sold to smelters, but the product 
 of the canvas tables is very rich in free gold, and is ground with 
 mercury in iron mortars by hand with pestles. The Lazzlo 
 amalgamators are about 25 inches in diameter, and each pair 
 treats from 1'7 to 2 tons of ore in twenty -four hours, 75 to 
 80 per cent, of the gold being s-aved. The loss of mercury is 
 about 1 oz. per ton of ore, and the power required for twenty- 
 four pairs of amalgamators and eight settlers is 4 H.P. 
 
 At a mine near Brad, in Transylvania, each pair of amalga- 
 mators treats from 3 to 3| tons of ore per day, but only 55 per 
 cent, of the gold is extracted, the tailings being treated in 
 American pans, and concentrated on Bilharz tables for smelting. 
 Specimens of the ore containing visible gold are amalgamated by 
 hand in mortars, and the rest contains about 8 dwts. per ton. 
 
 Amalgamation Pans. An amalgamation pan consists of a 
 circular cast-iron pan, provided on the inside with a renewable 
 false bottom of cast iron constituting the lower grinding sur- 
 face and a "muller," or upper grinding surface (d, Fig. 37), 
 attached to a vertical revolving spindle, <?, which is set in motion 
 by bevel wheels, t, placed below the pan. The muller grinds to 
 impalpable pulp ore which has been already reduced to a coarse 
 powder by stamps, and also mixes the ore with mercury, intro- 
 duced into the bottom of the pan, and so amalgamates the gold 
 and silver. The origin of the pan is clearly to be traced to the 
 Mexican arrastra, and some of the varieties of the pan are merely 
 slightly modified arrastras. One variety consists of a sectional 
 wrought-iron pan, fitted with a granite grinding bottom and 
 with granite mullers, which are attached to a vertical spindle 
 rotated by hand or by animal power. The Berdan pan also 
 forms a transition between the arrastra and the modern pan, 
 see p. 214. 
 
 In work on gold ores the use of amalgamating pans is mainly 
 limited to regrinding skimmings, blanket sands, and concentrates 
 
168 THE METALLURGY OF GOLD. 
 
 obtained in working a stamp mill. Silver ores in many cases do 
 not readily yield their precious metal if merely treated in a stamp 
 battery and run over copper plates. These ores are then crushed 
 in a battery, roasted with salt if necessary, and then amalgamated 
 in pans. Silver ores containing considerable quantities of gold 
 are often similarly treated, but with purely gold ores it is seldom 
 necessary to resort to this process, and a detailed description 
 will, therefore, be more in place in the volume devoted to the 
 metallurgy of silver, only a brief account being given below. 
 
 Gold ores which do not yield a fair percentage of their values 
 by copper-plate amalgamation are occasionally treated in pans. 
 In such cases the ore may be roasted or treated raw. As already 
 stated, p. 152. it is seldom advantageous to roast a gold ore 
 before amalgamation, since, although in a roasted pyritic ore 
 specks of free gold may often be detected where none were 
 visible in the raw ore, a part of the precious metal usually 
 appears after roasting to be difficult to bring in contact with 
 mercury. The cause of this is not always easy to discover, but 
 it may sometimes be due to the coating of gold by thin films of 
 iron oxide or other mineral. Moreover, the addition of salt to 
 a gold ore in the roasting furnace, as is pointed out in the 
 chapter on chlorination, is often attended by appreciable losses 
 by volatilisation. These two causes are sufficient to account for 
 the low percentage of gold usually extracted when an auriferous 
 silver ore is treated by roasting with salt and pan-amalgamation. 
 When no base metals are present a gold ore may sometimes 
 prove to be satisfactorily handled by roasting and amalgamation, 
 but such cases are exceptional. 
 
 Pan-amalgamation, whether the ores treated are raw or roasted, 
 may be conducted in one of two ways. The older system is to 
 crush wet in the stamp mill, and collect the ore in large shallow 
 settling pits or pointed boxes (see p. 177). A sufficiently dry 
 pulp having been obtained by draining, it is dug out by hand 
 and charged into the pans. The modern system, which is already 
 extensively in use, is called the Boss Continuous System, after 
 the name of its inventor, M. P. Boss. It is briefly described 
 on p. 171. 
 
 Old System. The amalgamating pans in use are very 
 numerous, and vary greatly in form. The shape of the bottom 
 was formerly much in dispute, flat, cone-shaped, and hemi- 
 spherical bottoms each having its advocates, but it is now 
 generally believed that flat-bottomed pans are the best, wearing 
 more evenly and doing more work. The pans are often heated, 
 so as to increase the rate of amalgamation by means of steam 
 led through a chamber below a false bottom in the pan. but the 
 more economical device of introducing steam into the pulp 
 itself has also at all times been in use. The objections to the 
 latter course are that the pulp may be so much diluted that 
 
CRUSHING AND AMALGAMATING MACHINERY. 
 
 169 
 
 amalgamation is checked, and that oil is liable to be introduced 
 with the steam with equally disastrous results. When the ore 
 is roasted before being treated in the pan, it is in some mills 
 charged in hot, hot water being added also, and as the pan is 
 covered up and is still warm from the previous charge, it 
 remains at a sufficiently high temperature throughout the 
 operation without further treatment. The grinding of the ore 
 by the muller is an additional source of heat. 
 
 One of the common forms of amalgamating pans is shown in 
 Fig. 37. This pan is 5 feet in diameter, with cast-iron bottom, a, 
 
 Fig. 37. 
 
 and wooden sides, h. The mullers are shown resting on the cast- 
 iron dies, c, which protect the bottom from wear, whilst re- 
 placeable shoes attached to the lower surface of the mullers 
 are also shown. The shoes and dies can be kept in contact 
 while the spindle, g, is rotated, so that the ore can be ground, 
 or the muller can be raised by rotating the hand -.wheel and 
 centre screw, j, on the top of the spindle, so that only circulation 
 and mixing of the charge take place. In some pans copper plates, 
 I, are introduced, being attached to the side walls and projecting 
 into the interior. These plates are intended both to mix the 
 
170 THE METALLURGY OF GOLD. 
 
 pulp and to catch the amalgam, much of which is retained on 
 them. The more usual system is to employ separate vessels 
 called settlers for the collection of the quicksilver and amalgam, 
 after the pans are discharged. The speed of the muller is 
 usually from 65 to 75 revolutions per minute ; below the muller 
 the pulp is continually worked from the centre of the pan to the 
 circumference, being returned towards the centre above the 
 muller and passing down through the latter by inclined slots 
 which terminate near the centre. In Fig. 37, which represents 
 the form known as the Patton pan, n is the main through 
 which steam is passed into the chamber, b, to heat the pulp, and 
 m is the outlet pipe. 
 
 The method of operation is as follows : The charge of ore is 
 introduced with the mullers raised slightly and kept revolving, 
 water being added at the same time in quantities sufficient to 
 make the pulp of a pasty consistency, so that globules of mercury 
 remain suspended in it without subsiding. The mullers are then 
 lowered and the ore ground for from two to four hours, after 
 which the mullers are raised and the mercury added gradually, 
 and thoroughly mixed with the pulp for 6 to 8 hours longer. The 
 object in raising the mullers is to prevent the sulphides from 
 being ground up with mercury, which would cause considerable 
 losses by flouring and sickening. Nevertheless this raising of the 
 mullers is not an invariable practice. When the amalgamation 
 is thought to be complete, water is introduced to dilute the pulp, 
 and the whole is discharged into a settler ; or else the diluted 
 pulp is stirred by the raised muller at a reduced rate of speed 
 until the globules of mercury have re-united and sunk to the 
 bottom, when the pulp is gradually run off, beginning at the top, 
 usually by pulling out in succession plugs set in the side of the 
 pan at different levels. The discharge takes place into a bucket 
 or tub, where some of the mercury accidentally carried over is 
 caught. The bulk of the mercury in some mills is drawn off 
 from the bottom of the pan before the pulp is discharged. 
 
 In order to facilitate amalgamation various chemicals have 
 been recommended as desirable additions to the charge. At the 
 present day it is recognised that most of these are either useless 
 or absolutely harmful, and only salt, sulphate of copper, nitre, 
 cyanide of potassium, lime, and sodium amalgam are now used. 
 In treating gold ores, cyanide of potassium and sodium amalgam 
 are added to keep the mercury clean and lively, but the latter 
 chemical is now comparatively rarely resorted to. Salt and 
 sulphate of copper are chiefly added to silver ores, their use 
 having been suggested by the Patio process. They are believed to 
 decompose certain base minerals, and so to prevent the sickening 
 of mercury, which would otherwise be caused by their presence, 
 and also to liberate silver from some of its compounds and thus 
 render it capable of amalgamation. The use of lime is of course 
 
CRUSHING AND AMALGAMATING MACHINERY. 171 
 
 to neutralise any acid sulphates of iron, &c., which may be- 
 formed by the partial decomposition of the ore, and so to prevent 
 the sickening of the mercury. If added when the pulp is- 
 diluted, lime is said to be efficacious in assisting the mercury to- 
 collect together and settle.* 
 
 The Boss Continuous System. In this system of pan-amalgama- 
 tion the pulp is continuously run direct from the stamp battery 
 through a series of pans arranged so that each overflows into the 
 next one, which is placed at a slightly lower level. The first 
 
 Fig. 38. 
 
 Scale, 1 inch = 3 feet. 
 
 two or three pans are arranged as grinders, the battery pulp not 
 being fine enough for complete amalgamation, and the pulp is- 
 then passed through a series of amalgamating-pans supplied with 
 mercury, after which the mercury and amalgam are separated 
 from the ore in settlers, which are larger pans in which the pulp 
 is diluted and stirred less vigorously. The tailings overflow 
 
 * For a full account of the chemical reactions involved in pan-amalgamation, 
 the student is referred to the Report of the United States Survey of the 
 Fortieth Parallel, vol. iii., chap. v. 
 
172 THE METALLURGY OF GOLD. 
 
 from the settlers, and are run to waste or led over concentrators. 
 The number of pans arranged in series through which the pulp 
 must pass, in order to yield a fair percentage of its precious 
 metals, is determined by experiment for each particular ore. It 
 is obvious that the consistency of the pulp must be thinner than 
 has usually been considered desirable for successful amalgama- 
 tion, and, as a matter of fact, its volume is usually doubled by 
 the introduction of the continuous process, but in spite of this 
 the percentage of extraction is not lower than by the old method. 
 By the Boss system there is a large saving in labour, fuel, and in 
 wear and tear ; the settling pits or pointed boxes are dispensed 
 with, and no movement of the pulp by hand is needed. The 
 mercury is collected in wells and pumped up into tanks, whence 
 it is fed automatically into the amalgamating pans. One of the 
 pans in use in this process is the Boss Standard Pan, shown in 
 Fig. 38. It will be noticed that there is a steam chamber below 
 the false bottom of the pan, extending up into the conical space 
 in the centre, for warming the pulp. 
 
 Concentration before Pan- Amalgamation. Some refer- 
 ence must be made to this subject and also to that of concentra- 
 tion after pan-amalgamation, as much attention will probably be 
 directed to them in the near future. In most cases it is desirable 
 first to remove the sulphides by concentration and then to treat 
 the tailings in pans. In this way the objectionable, but often 
 valuable, minerals, which complicate the reactions in the pan and 
 cause losses in mercury and amalgam by sickening, are prevented 
 from doing mischief, whilst they are saved by concentration 
 after the preliminary crushing in the battery, more effectually 
 than if they are also subjected to the grinding and sliming action 
 of the mullers. After concentration, the tailings of course con- 
 tain too much water for immediate pan-amalgamation, and rhe 
 excess of water must be removed by settling in pointed boxes. 
 The chief disadvantage in this course is that, where the ore con- 
 tains chlorides, sulphides and other compounds of silver, the 
 slimes which remain in suspension in the water are the richest 
 part of the pulp. It is, therefore, necessary to collect them as 
 perfectly as possible by settling, but, in these cases, a certain 
 amount of loss is unavoidable. The settled tailings are now in 
 good condition for amalgamation, and can be treated expedi- 
 tiously and effectively. If, on the other hand, the ore is very 
 rich in silver, so that the loss caused by immediate concentration 
 is heavy, it is first treated in pans, and the product is then passed 
 over Erue vanners or other slime tables. 
 
 Treatment of Concentrates in the Pan. Some details of 
 the treatment of concentrates by pan-amalgamation are given in 
 Chap. x. in the accounts of work in special localities. It is, how- 
 ever, a survival of old methods, and does not represent the most 
 modern practice, in which concentrates are either smelted or 
 
CRUSHING AND AMALGAMATING MACHINERY. 173" 
 
 treated by wet methods. Whether they have been previously 
 roasted or not, the treatment of concentrates in pans is seldom 
 attended by the successful extraction of a high percentage of 
 the gold. A recourse to this method is justifiable only when the 
 concentrates are not in sufficient quantity to warrant th.3 erection 
 of a chloriiiation plant, and when there is no smelting works near 
 where they can be sold. Under these circumstances a stone 
 arrastra usually gives better results in treating roasted concen- 
 trates than an iron pan. In Australia the method is often 
 adopted of employing a large excess of mercury and little water, 
 and of keeping the roasted material from contact with iron, and 
 in some experiments conducted in Mexico, C. A. Stetefeldt found 
 that by the use of gold amalgam instead of mercury, and by 
 grinding in stone vessels, a high percentage of gold was extracted, 
 from low grade ores. The chief advantage in roasting lies in 
 the reduction of the loss of mercury which is effected by the 
 elimination of the compounds of sulphur, the percentage of 
 extraction of gold not being, as a rule, much affected. This 
 subject is discussed under the heading Gold in Pyrites, 011 p. 152. 
 Molloy's Hydrogen Amalgamator. A description may be 
 given of this machine as it has gained the approval of some 
 well-known metallurgists, although it has never been highly 
 esteemed by practical men, and, in spite of the fact that it has 
 been before the public for several years, has not yet had any 
 success in prolonged operations on a large scale. It consists 
 of a circular pan of cast iron, about 40 inches in diameter and 
 4 inches in depth, which is nearly filled with a bath of mercury. 
 In the centre is a bottomless ebonite box, the sides of which dip 
 into the mercury for about J inch. This is called the anode box, 
 and is filled with an aqueous solution, which in the earlier 
 form contained sulphate, but in the later form carbonate of soda. 
 A leaden plate dips into this solution, and is connected with the 
 positive pole of a battery, the negative pole of which is con- 
 nected with the mercury bath. A wooden disc, nearly as large 
 as the pan, touching the mercury near the periphery of the basin, 
 but riding free above it towards the centre, is rotated, when the 
 machine is at work, at a speed of not more than sixteen revolu- 
 tions per minute, and crushed pulp and water is fed in near the 
 centre of the machine, but outside the anode box. This pulp 
 travels outwards in widening circles under the influence of the 
 rotating disc, which causes it to come in contact with the 
 mercury, and overflows at the edge, leaving its precious metals 
 amalgamated with the bath of quicksilver. An electric current 
 passed through the machine is supposed to decompose the car- 
 bonate of soda, liberating free sodium and hydrogen over the 
 whole surface of the mercury, which is thereby kept clean and 
 lively. It is doubtful, however, if any free sodium is ever 
 present in the bath. If formed it would be at once oxidised 
 
174 THE METALLURGY OP GOLD. 
 
 by the water, liberating hydrogen ; but this element, when 
 nascent, is doubtless of great efficacy in reducing base oxides 
 which may be contaminating the mercury, and so in keeping 
 the latter from sickening. The absence of all grinding action 
 prevents any considerable loss from occurring by flouring, 
 consequently, the loss of mercury in operating the machine 
 would probably be small. 
 
 The value of this machine depends on the success of the efforts 
 to induce perfect contact between the mercury and the gold 
 particles contained in the ore. It does not seem likely that this 
 contact is obtained. The machine is said to be able to treat 
 from 10 to 20 tons per day, and, therefore, to deal with the 
 product of from five to ten head of stamps. The area of the 
 amalgamating surface, however, over which the pulp passes is 
 only about 8 square feet, whilst the area of the copper plates 
 used in the stamp battery in the ordinary way would be from, say, 
 32 to over 100 square feet. The layer of pulp between the disc 
 .and the mercury, if this quantity is treated, must therefore be 
 much thicker than the corresponding layer flowing over copper 
 plates, so that there is less chance of every part of the pulp 
 being brought in contact with the amalgamating surface in the 
 former case. As already intimated, the machine has not yet 
 been proved a success in prolonged operations on a large scale. 
 
 Jordan's Amalgamator. This machine is intended for the 
 treatment of auriferous tailings. It consists of a number of 
 shallow amalgamated copper pans fixed one below the other on 
 .a vertical axis, which is revolved slowly by bevelled gearing 
 placed at the top. Between each two pans is a circular shelf, 
 projecting from the wall of the surrounding casing. The tailings, 
 in the form of fine pulp, are fed in to the first pan, and, by the 
 centrifugal force exerted by the revolution of this pan, the pulp 
 is caused to pursue a spiral course, until, after several revolu- 
 tions, it falls over the edge on to the shelf below. The pulp 
 then flows back on the shelf to the centre of the machine and into 
 the second pan. In this way it traverses the surfaces of all the 
 pans, one after the other, and finally falls into a conical separator, 
 which, it is claimed, saves any heavy material and also any mer- 
 cury that has escaped. The machine has not yet passed into 
 use on any extended scale, although it has been tried at a few 
 mills for short periods of time. It seems well adapted to catch 
 fine gold if the amount treated is not too large to admit of 
 perfect and prolonged contact with the plates. 
 
CONCENTRATION IN STAMP MILLS. 175 
 
 CHAPTER IX. 
 CONCENTRATION IN STAMP MILLS. 
 
 Concentration. The object of concentration is the separation 
 of the heavy valuable mineral from the light worthless gangue. 
 In Europe, where careful attention has been paid to the 
 mechanical treatment of ores for some hundreds of years, compli- 
 cations are introduced by the fact that various base minerals 
 must be separated from one another, an ore being subdivided 
 into several products. Almost all gold ores, however, only re- 
 quire separation into two parts the "concentrates," in which 
 the precious metal is contained, and the " tailings," which are 
 thrown away. Consequently, the German machinery has not in 
 general proved applicable to gold ores without modification, and 
 most of the best appliances now in use have had their origin in 
 the United States. The German system of coarse crushing, sizing 
 by means of screens, and concentrating on jigs, is not, as a rule, 
 applicable to gold ores proper, although much used on auriferous 
 lead, zinc, and copper ores. This system will not be described 
 here, where only the methods in use for the treatment of battery 
 sands, after passing over the amalgamated plates, will be con- 
 sidered, and some reference made to the machines in use for 
 the mechanical treatment of pulp, either before or after pan- 
 amalgamation. 
 
 All the concentrating machines depend for their action on the 
 effect of a difference of densities on the fall of bodies in a fluid. 
 The fluid employed in almost every instance is water, although 
 several machines have been devised in which air is used as the 
 concentrating medium. The use of any other fluid than these 
 may be safely condemned as chimerical. It has often been pro- 
 posed to use some fluid which shall have a lower density than 
 the valuable mineral to be saved, but a higher density than the 
 worthless gangue ; the mineral would then sink, while the gangue 
 would remain floating. The high cost of any such fluid is suffi- 
 cient to put this method out of the question, without discussing 
 any further disadvantages. 
 
 The fall in water of solid materials takes place according to 
 two laws, one applicable to very shallow strata, through which 
 the particles fall with increasing velocity, while the other is true 
 when the depth of the water is considerable, so that the particles 
 for the greater part of their course proceed at their maximum 
 velocity. In shallow water the fall is almost entirely according 
 
176 THE METALLURGY OP GOLD. 
 
 to density,* so that those machines which utilise only the first in- 
 stants of the fall will have great efficacy in concentrating. It is 
 this fact which has necessitated the use of shallow currents in 
 concentrating tables, sluices, &c. In almost all these machines 
 the fine sand and slime is brought into suspension in water, and 
 the liquid is then run over an inclined surface. The deposit of 
 sand, which is thus formed on the table, tends to become enriched 
 in heavy minerals, because the stream moves faster at the surface 
 of the water, where the lighter particles remain, than it does 
 next the bed, where the heavy particles have settled. The 
 deposit is continually worked up and brought again into sus- 
 pension by a rake or broom, or by a series of shakes or blows 
 imparted to the apparatus, so that the effect mentioned above 
 is repeated frequently. If the stirring up is violently performed 
 all the slime and very fine particles are kept in suspension in 
 the water and carried away and lost, a slow stream of water 
 and very slight agitation being favourable to their retention in 
 the deposit which is formed. When the stream of water is rapid 
 and voluminous, fine material, whether heavy or light, is swept 
 away and lost in the tailings, whilst if too small a stream of 
 water is used, much worthless sand is deposited with the con- 
 centrates. It follows that the amount of water used must be 
 regulated according to the work which it is proposed to do. 
 Frequently, it happens that clear water must be added to the 
 pulp to dilute it sufficiently. On the other hand, it occasionally 
 happens that the pulp is too thin, especially after hydraulic 
 sizing, and water is then removed by means of settling boxes, 
 described below. 
 
 Another operation, which it is often of the utmost importance 
 to perform before concentration, is that of classification by size. 
 The necessity of this is obvious, when it is remembered that a 
 shallow stream of water swift enough to carry down fine sul- 
 phides, might be powerless to move a pebble of quartz. The 
 usual method of classification is based on the varying rates of 
 fall of particles through a deep column of water. In this way 
 equal-falling particles are obtained together, and since a sphere 
 of galena is equal-falling with a sphere of quartz of four times the 
 diameter,! it follows that the sizing cannot be perfectly performed 
 
 *At the very beginning of the fall, before the velocity has had time 
 enough to become great, the motion varies nearly as g ^1 -g j, where S is 
 
 the density of the medium (in this case, water), and D the density of the 
 solid particle. When the depth is sufficient for the velocity to attain its 
 maximum, this maximum velocity will be the same for all particles for 
 which a (D - <5) has the same value, where a 2 is the area of the section of 
 the particle at right angles to the line of fall. Such particles are called 
 equai-f ailing. For full investigations of these formula? the student is- 
 referred to Gallon's Lectures on Mining, English edition, vol. iii., p. 47. 
 
 t Richards states (Mineral Industry for 1896, p. 705) that m practice, 
 with particles between 10- and 60-mesh screens, the diameter of the quartz. 
 is from 2 to 31 times that of the galena. 
 
CONCENTRATION IN STAMP MILLS. 
 
 177 
 
 by this method. Nevertheless, such sizing as is possible in this 
 way, when efficiently performed, is of great assistance as a pre- 
 paration for treatment by shallow-stream concentrators. Some 
 sizing machines are described below, p. 178. 
 
 Settling Boxes. These are for the purpose of allowing the 
 sand and mineral in suspension in a flowing current to settle, so 
 that the part of the water not needed in the subsequent treat- 
 ment of the ore may be run off; if it is desirable, this excess of 
 water is available for use over again. These boxes will deliver a 
 small stream of thick concentrated pulp at the bottom and give 
 an overflow of almost clear water at the top. The boxes manu- 
 factured by Messrs. Fraser & Chalmers are shown in plan and 
 
 m m _//m m 
 
 Fig. 39. 
 
 elevation in Fig. 39. The main points observed in the con- 
 struction of these boxes are as follows : 
 
 The sides are at an angle of at least 50 to the horizontal, to 
 insuie the uninterrupted descent of the slimes. The pulp is 
 delivered evenly across the end of the box by means of some 
 such arrangement as the movable tongues (a, Fig. 39). Surface 
 currents are prevented and the incoming pulp thoroughly mixed 
 with the mass of the water in the box by the partition, 6, which 
 extends across the box near the inflow end. The discharge is 
 made by cutting the other end of the box from 2 to 3 inches lower 
 than the sides ; this overflow is made perfectly level so that the 
 water flows out in an even sheet. The discharge pipe for the 
 pulp is at the bottom of the box. The siphon discharge shown 
 
 12 
 
178 THE METALLURGY OF GOLD. 
 
 is the most convenient, as the aperture need not be so much con- 
 tracted as if direct discharge is used; and, besides this, less fall 
 in the mill site is required. A contracted orifice is liable to be 
 choked up, and a smaller diameter than 1 to 2 inches is to be 
 deprecated. 
 
 These boxes have the advantage over the ordinary settling -pits 
 used to retain tailings and to catch pulp for pan-amalgamation, 
 that they do not require to be dug out, while the settling, owing 
 to the elimination of surface currents, is more perfect. One 
 large box of 12 feet by 6 feet at the top and 7 feet 6 inches deep 
 is sufficient to settle the product from five or ten stamps, accord- 
 ing to the degree of separation of water and ore required. 
 
 If the pulp is required for treatment in amalgamating pans, 
 such settling boxes do not give material of sufficiently thick 
 consistency for the purpose. To adapt them to this use 
 M'Dermott suggests that the pulp should be allowed to flow 
 through two small square tanks placed above the pointed boxes, 
 so as to catch the heavy sand. As soon as one tank is full, the 
 pulp is turned to pass through the other, while the first is dug 
 out after a short draining. The slimes are caught in the pointed 
 boxes, and after mixture with the drier heavy sand the pulp is 
 not too thin for amalgamation in pans. Concentration before 
 amalgamation is, however, as yet seldom resorted to. 
 
 Classification according to Size. Classification according 
 to size, although seldom attempted by mill-managers in the 
 treatment of battery pulp, is desirable as a preliminary to con- 
 centration by almost every machine ; these work more efficiently 
 if engaged in the separation of equal sized particles of different 
 specific gravities than if pieces of all sizes are mixed up together. 
 
 The sizing machines most commonly employed are of two 
 kinds, viz. : 
 
 1. Revolving or flat shaking screens, which are suitable 
 
 for coarsely ground ores. 
 
 2, Hydraulic classifiers i.e., boxes containing ascending 
 
 currents of water ; these are suitable for finely pul- 
 verised ores. 
 
 1. Sizing by screens is almost always conducted on wet 
 material, spraying jets of water being used to carry the small ore 
 through. Flat shaking screens offer advantages in cheapness 
 and simplicity, but revolving cylindrical screens, inclined at a 
 slight angle to facilitate discharge, are in more general use. It 
 is customary to employ screens of the minimum fineness of 8 to 
 16 mesh. Where the sizing of finer particles than could be 
 treated by this mesh is required, hydraulic classifiers are used, 
 since their cost in repairs and renewals is far less than that of 
 screens. The latter are seldom used in the sizing of battery 
 sands, but are recommended for the purpose by Resales. * 
 
 * Loss of Gold in the Reduction of Auriferous Veinstone in Victoria. By 
 H. Resales, Melbourne, 1895. This exhaustive aud closely reasoned 
 treatise will well repay a careful study. 
 
CONCENTRATION IN STAMP MILLS. 179 
 
 2. Hydraulic sizers were introduced by Prof. P. von Rittinger 
 more than forty years ago, for use in the Hartz, and from their 
 shape were known as Spitzkasten or pointed boxes. These boxes 
 have the shape of inverted pyramids, the stream of unsized pulp 
 entering at one side and flowing out at the other, whilst there is 
 also a small discharge at the apex. The current slackening on 
 entering the box, the heavier and larger particles in suspension 
 at once begin to settle, and, escaping the influence of the current, 
 fall quietly to the bottom of the box where they are discharged. 
 It is evident that in this classification the particles which reach 
 the bottom will not all be of the same size. A piece of galena 
 of specific gravity 7 '5 will fall through water at a faster rate 
 than a piece of quartz of the same size, but only one-third of 
 the weight. The rate of fall is dependent both on the size 
 of the particles and on their densities, according to a well-known 
 law.* By applying this law it is easily shown that in water a 
 sphere of iron pyrites (specific gravity 4*9) of If mm. in diameter 
 will fall at the same rate as a sphere of quartz (specific gravity 
 2 -6) of 4 mm. in diameter, and a sphere of galena of 1 mm. in 
 diameter. In each successive box, then, a number of approxi- 
 mately equal-falling particles are removed, the closeness with 
 which the subdivision is made varying with the size of the box, 
 and the corresponding extent to which the current carrying the 
 pulp is checked on entering it. The larger the box the more 
 material is collected by it, and, therefore, the more heterogeneous 
 the particles caught. In a small box the material collected con- 
 sists more nearly of truly equal-falling particles. 
 
 These equal-falling particles however are, in any box, whatever 
 its size may be, carried away through the aperture in its apex 
 by muddy water containing material of all sizes down to the 
 finest slime, and it was to eliminate this material that the 
 ascending current was introduced. This is a current of clear 
 water, which enters at the apex of the box in greater quantity 
 than can be discharged by the outflow near the same spot, so that 
 there is an upward current of water into the box. The result is 
 that no muddy water is discharged below, but only the particles 
 of ore which have weight enough to drop through the ascending 
 current, so that by regulating the strength of this, any desired 
 class of ore can be obtained. In Fig. 40, a form much used in 
 America is shown. The sliding partition assists the settling, by 
 causing the pulp to pass downwards, rapid surface currents 
 across the box to the overflow being thus prevented. The 
 discharge of the heavy particles is effected through A, the 
 clear water pipe itself, by the arrangement shown. The launder 
 above the box supplies the clear water current, and shows the 
 head of water used, which must be kept constant to ensure 
 uniformity of results. 
 
 * See footnote on p. 176. 
 
180 
 
 THE METALLURGY OF GOLD. 
 
 The chief defect of the early forms of pyramidal boxes was 
 that, as the area of the vat became larger and larger towards the 
 top, the velocity of the ascending water naturally became less 
 and less, so that many particles were able to settle down below 
 the level of the overflow, but were stopped by the increasing 
 force of the current, so that an accumulation of ore took place 
 half way up the box, and ultimately became so great as to inter- 
 fere with the classification. Several remedies have been devised 
 for this defect, of which one of the simplest and most effectual is 
 to make the box pyramidal below, but with vertical sides in the 
 
 Fig. 40. 
 
 upper part. This construction is partly carried out in the box 
 described above, which, however, though given as a type, is not 
 theoretically the most perfect of its kind. A slime-pit is added 
 to catch the stuff which is too light to settle in the boxes, in all 
 cases in which these slimes are of sufficient value to pay for 
 treatment. The number of boxes used depends on the number 
 of classes of material which it is deemed advisable to make. 
 Usually two or three classes are sufficient, and, as already stated, 
 in the majority of cases no sizing is attempted. 
 
 Early Concentrating Machinery. One of the oldest and 
 most primitive machines employed in concentration of fine sand 
 by means of a shallow stream of water was the German buddlc, 
 
CONCENTRATION IN STAMP MILLS. 181 
 
 which has a distinct but imperfect resemblance to the Long-Tom 
 described on p. 47.* Canvas tables were used below these 
 buddies in Germany, and probably suggested the use of blanket- 
 brakes or tables, which were adopted in the early days of the 
 gold fields of the United States and Australia, and are still 
 retained in some places. The rough surface of the blanketing 
 seems to be particularly efficacious in catching and holding thin 
 plates and spangles of free gold or of sulphides, which are readily 
 washed off smooth surfaces by a current of water, and these 
 rough appliances, although almost useless for catching slimes, still 
 find favour where considerations of economy prevent the purchase 
 of modern high-priced concentrators. The blanketing is usually 
 in strips of 16 or 18 inches wide, and several feet long, and is 
 nailed or stretched on wooden frames, which have an inclination 
 of about J inch per foot. 
 
 At intervals of about half an hour, when a quantity of mineral 
 has already been collected on the rough surface, the blankets are 
 taken off and washed in tubs of water, where a deposit collects 
 which is afterwards dug out. At the St. John del Rey Mine, 
 the framework supporting the blankets was hung on pivots above 
 a shallow tank. When it was necessary to clean them, the 
 framework was turned so that the upper surface of the blanket 
 was inclined downwards, and the mineral washed off its surface 
 by a hose, much time and labour being thus saved. 
 
 At this mine the trays supporting the blankets were 18 inches 
 wide and 30 feet long, with a fall of 1 inch per foot. The upper 
 16 feet were covered with bullock's skins, tanned with the hair 
 on them, and in lengths of 26 inches ; below these were a series 
 of blankets or baize cloths of the same length, made of coarse 
 wool with a long nap. The fall from the battery box upon the 
 tray was 4 inches, a screen being placed across the end to break 
 the fall of the water, and cause it to strike the tray nearly at 
 right angles. About 90 per cent, of the gold contained in the 
 ore was caught on these blankets. The blanket sands contained 
 95 per cent, of sulphides, and were so fine that 90 per cent, of 
 them passed through a 100-mesh sieve. They were amalgamated 
 in revolving wooden barrels, yielding 96 per cent, of their assay 
 value, but this was due to the fact that very little gold was 
 contained in the pyrites, most of it being present in the form of 
 free particles. 
 
 Blanket sluices have been declared unsuitable for catching fine 
 sulphides, and their concentrates are usually contaminated by 
 admixture with much sand. If set at a proper inclination they 
 will save fine amalgam and free gold, but even in this respect 
 they are less satisfactory than shaking copper plates and riffles. 
 
 * For a description of the similar huddle formerly used in Colorado for 
 the treatment of battery sands, see Raymond's Mines, Mills, and Furnaces 
 of the Pacific States. New York, 1871, p. 357. 
 
182 THE METALLURGY OF GOLD. 
 
 Riffled sluices were employed at the same time as blankets for 
 effecting rough concentration. The riffles were formed of half- 
 inch strips of wood nailed across the sluice box, the grade ot 
 which was about three-quarters of an inch to the foot. As 
 soon as the concentrates had accumulated until they reached 
 the top of the riffle, another strip was nailed on, and the process 
 was repeated until the bed of concentrates was several inches 
 thick, when they were scraped out and a fresh start made. 
 Similar to this device was the raising-gate concentrator, which 
 was practically a riffled sluice in which the riffle was raised 
 continuously by machinery, instead of being adjusted at intervals 
 by hand. 
 
 The Round Buddie was invented in Cornwall, where it is still 
 used in dressing the tin ores to the exclusion of almost every 
 other concentrator. There are two varieties, 
 
 1. The convex round buddle, in which the ore and water are 
 added at the centre of the machine, and flow down over the 
 surface to the periphery. 
 
 2. The concave buddle, in which the pulp is added at the- 
 periphery of the machine and flows down to the centre. 
 
 In both cases revolving arms, carrying brushes, pass over the 
 surface and stir the deposit as it is being formed, and the spouts 
 distributing the ore also rotate so as to deliver the pulp evenly. 
 The buddies are from 12 to 18 feet in diameter. 
 
 These machines are not continuous in action, and after the 
 deposit has accumulated to a depth of a few inches the operation 
 is suspended and the deposit dug out. The "headings" or 
 material on the upper 12 or 18 inches of the inclined surface are 
 kept separate, and the stuff near the bottom of the slope is called 
 the "tailings." Eound buddies are not adapted to obtain a 
 finished product in one operation. The headings and tailings 
 must, as a rule, be subjected to farther treatment, and between 
 them is a large quantity of material which differs little from the 
 original ore. Thus handling and re-handling of the stuff is 
 necessitated, and it is on this account that these machines are 
 not now much used on gold ores. The principle on which they 
 depend is favourable to the collection of slimes, and the modern 
 improved buddies are perhaps better adapted for fine ores than 
 any other machines, except those employing a travelling belt. 
 Modifications have been proposed to adapt these buddies to the 
 treatment of gold ores by adding riffles containing mercury, and 
 by other devices, but have not found much favour. One of the 
 chief changes, which was proposed in the United States, is to 
 keep the brushes and ore-spouts stationary, and to rotate the 
 inclined bed. During the earlier part of the revolution an 
 attempt is made to separate the pulp into headings, middlings, 
 and tailings, and each of these is washed into a separate launder 
 by strong jets of water before the completion of one turn, the- 
 
CONCENTRATION IN STAMP MILLS. 183 
 
 table being left clean for further work. A saving of labour in 
 digging out the material is thus effected. The inclination of a 
 buddle is usually from 1 to 1J inches per foot. A device 
 formerly much in vogue in Western America consisted in the 
 use of a cylinder of iron in the centre of a concave buddle, 
 through which the discharge took place. This cylinder was 
 made to rise slowly as the concentrates accumulated on the 
 buddle, the effect being similar to that obtained in the sluices 
 with self-raising gates. 
 
 For working large quantities daily of low grade ore containing 
 a high percentage of pyrites, the revolving buddle is still used. 
 When the percentage of pyrites is low, or if the ore is rich, so 
 that closer work is desirable, or if the slimes are worth catching, 
 revolving-belt tables are substituted for the older machines. The 
 advantages of the revolving buddies consist in low first cost, 
 large capacity, and ability to stand large quantities of water, 
 which are always present in the pulp after the operations of 
 screening and hydraulic sizing. 
 
 Centrifugal Concentrators. Among concentrators with 
 circular revolving beds, which were apparently suggested by the 
 revolving buddle, may be mentioned Hendy's and Duncan's 
 machines. Hendy's concentrator was one of the earliest forms 
 employed in California, having been patented in 1868, and is 
 still retained in some mills. It consists of a shallow cast-iron 
 pan, 5 feet or 6 feet in diameter, supported in the centre by a 
 vertical shaft. A rapid horizontal oscillating motion is given to 
 the pan by the revolution of a crank shaft, which is joined to 
 the periphery of the pan by a connecting-rod. The bottom of 
 the pan slopes gently downwards from the centre to the peri- 
 phery, and there is a basin-shaped depression in the centre, and 
 an annular gutter running round the outside. The pulp is fed 
 into the pan near the periphery, and a depth of about 3 inches 
 of material is maintained. In consequence of the rapid oscillating 
 motion the heavy particles move outwards by centrifugal force 
 and accumulate in the peripheral gutter, while the lighter 
 particles are discharged into the central circular basin and are 
 removed by opening discharge gates at intervals. 
 
 The Duncan concentrator resembles the Hendy pan in con- 
 sisting of a circular iron pan, in which the heavy sulphides are 
 moved to the periphery by centrifugal force. The pan, however, 
 revolves, making 8J turns per minute, while the pulp is supposed 
 to make about 3 revolutions in the pan, thus allowing time for 
 the pyrites and gangue to separate before the former settles at 
 the bottom towards the sides, where it remains until discharged. 
 The lighter particles remain more or less in suspension in the 
 water and are carried to the centre, where their removal from 
 the pan takes place. This machine is used to some extent in 
 California, and, like the Hendy pan, is effective in the separation 
 
184 THE METALLURGY OF GOLD. 
 
 of coarse sulphides. Neither of these concentrators do good 
 work in saving slimes, for the reason that the movement of the 
 concentrates towards the periphery, depending on centrifugal 
 force, takes place in consequence of the superior mass of the 
 heavy minerals. With slimes, the mass of the particles of pyrites 
 is so small that it is not enough to enable them to overcome the 
 effects of the stream of water which carries them towards the 
 centre, and they are, consequently, swept away with the tailings. 
 Percussion Tables. In these machines the work of keeping 
 the pulp in a state of agitation, done by the rakes or brushes 
 in the German and Cornish buddies described above, is effected 
 by longitudinal shakes imparted to the table. The table is 
 made of wood or sheet iron, the surface being as smooth as 
 possible, and the sides being flanged. It is hung by chains or 
 in some similar manner, so as to be capable of limited movement, 
 and receives a number of blows delivered on its upper end. 
 These blows are given by cams acting through rods, or else the 
 table is pushed forward against the action of strong springs by 
 cams on a revolving shaft, and then being suddenly released is 
 thrown back violently by the springs against a fixed horizontal 
 beam. The movement of the pulp depends on the inertia of the 
 particles, which are thrown backward by the blow given to the 
 table, the amount of movement varying with their mass, and 
 depending, therefore, both on their size and density. The vibra- 
 tions produced by the percussion also perform the work of the 
 rakes in destroying the cohesion between the particles, and a 
 stream of water washes them down. The result is that the 
 larger and heavier particles may be made to travel up the table 
 in the direction in which they are thrown by the blow, by 
 regulating the quantity of water, while the smaller and lighter 
 particles are carried down. It is obvious, however, that the 
 inertia of the fine particles of heavy minerals will not be sufficient 
 to move them against the current, and, in consequence, all the 
 slimes, both rich and poor, will be carried down and lost in the 
 tailings. 
 
 Gilpin County " Gilt Edge Concentrator" This variety of 
 shaking table was devised in Colorado, and has displaced the 
 blanket sluices at almost all the mills at Blackhawk. It consists 
 (Fig. 41) essentially of a cast-iron or copper table, 7 feet long and 
 3 feet wide, divided into two equal sections by a 4-inch square 
 bumping-beam. The table has raised edges, and its inclination is 
 about 4 inches in 5J feet at its lower end, the remaining 1 J feet 
 at the head having a somewhat steeper grade. The table is 
 hung by iron rods to an iron frame, the length of the rods being 
 altered by screw threads, so as to regulate the inclination to the 
 required amount. A shaft with double cams, A, making 65 
 revolutions per minute, enables 130 blows per minute to be 
 given to the table in the following manner ; on being released 
 
CONCENTRATION IN STAMP MILLS. 
 
 185 
 
 by the cam, the table is forced forward by the strong spring, 
 B, so that its head strikes against the solid beam, C, which 
 is firmly united to the rest of the frame. The pulp coming 
 from the copper plates is fed on to the table near its upper 
 end by a distributing box, I), and is spread out and kept in 
 agitation by the rapid blows. The sulphides settle to the bottom 
 of the pulp, and are thrown forward by the shock, and eventually 
 discharged over the head of the table at the left hand of the 
 figure, while the gangue is carried down by the water and dis- 
 charged at the other end. One machine is enough to concentrate 
 the pulp from five stamps. If the table consists of amalgamated 
 copper plates, it is of some use for catching free gold also, 
 treating about 8 cwts. per hour. This machine, like other 
 
 Fig. 41. 
 
 percussion tables, is very effective for separating coarse pyrites, 
 but almost all the slimes are lost in the tailings. This is of 
 less importance, as the sulphides with which it has to deal 
 are not of high value, containing usually only from 10 to 12 
 dwts. of gold per ton. The high price of the True vanner and 
 other slime tables prevents their introduction on this field, as 
 they cost more than three times as much as the Gilt Edge con- 
 centrator for equal capacity. 
 
 The Frue Vanner. This machine is described in detail as 
 being typical of the shaking travelling-belt concentrators. Brief 
 accounts of some other forms are given subsequently. Machines 
 of this class are especially adapted for treating finely-crushed 
 battery sands which do not contain a large percentage of 
 " mineral" (that is, sulphides and other heavy materials). They 
 are frequently set to concentrate unsized pulp coming straight 
 from the amalgamating tables, and although they often do good 
 work under these circumstances, when the screens used are 
 equal to 40-mesh sieves or finer, nevertheless if the pulp con- 
 tains coarse stuff the results are not so satisfactory. Rosaies 
 
186 THE METALLURGY OP GOLD. 
 
 recommends * that the tailings from the plates should in general 
 be passed over percussion tables, and should then be separated 
 into four classes, viz. : Coarse sand, medium sand (which is 
 retained on a 40-mesh sieve), fine sand (which will all pass 
 through a 40-mesh sieve), and slimes. The coarse sand may 
 often be thrown away as worthless, although, if necessary, it is 
 reground, and the other classes may be advantageously treated 
 separately on some form of travelling-belt table or round buddle. 
 
 The Frue vanner (Fig. 42) consists essentially of an endless 
 rubber belt, mounted on a frame, with its upper surface slightly 
 inclined to the horizon, and subjected to two movements, a slow 
 constant longitudinal movement and a slight and rapid side shake. 
 The belt forms the bed or plane on which the dressing of the ore 
 is effected, being an inclined plane, 12 feet long, and bounded 
 down the two sides by projecting rubber flanges, which prevent 
 the water and sand from dropping over the sides. An arrange- 
 ment of rollers permits of the belt being slowly revolved in the 
 direction of its length and up the incline; thus, though the 
 dimensions of the working plane remain always the same, its 
 surface is constantly travelling. The crushed rock in a stream 
 of water is delivered near the upper end of the belt by means of 
 the sand distributor, No. 1, Fig. 42, and flows down the belt 
 towards its lower end. Now, as the inclination at which the 
 belt is set is only from 3 to 6 inches on the 12 feet, and as 
 the stream of water is not large and spreads over the whole 
 width of 4 feet, it is obvious that, if it were not for the 
 movements of the belt, much of the crushed rock contained in 
 the water would settle on the belt, while the water and the finer 
 and lighter particles of sand would alone reach the foot of the 
 table and drop over into a waste launder. 
 
 In order to separate the heavy metallic minerals from the 
 accompanying gangue or rock, a second stream of water is 
 applied, whilst a gentle side shake is given to the belt, to keep 
 the sand in a state of agitation, prevent it from "packing," and 
 facilitate the sorting process. The water distributor, 2, is placed 
 about 1 foot above the pulp distributor, and delivers small jets 
 of water, 3 inches apart, over the entire width of the belt. The 
 side shake thoroughly mixes this water with the pulp, spreads 
 the whole uniformly on the belt, and enables the heavy particles 
 of mineral to settle through the sand and cling to the belt, when 
 they are carried up by it past the small jets of water and 
 deposited in the collecting tank, while the lighter gangue is 
 carried down by the stream and delivered into the tailings 
 launder. 
 
 The Frue vanner is shown in plan in Fig. 42&, in side eleva 
 tion in Fig. 426, and in end elevation in Fig. 42c. The following 
 description is abridged from that given by the manufacturers : 
 
 * Loc. dt. 
 
CONCENTRATION IN STAMP MILLS. 
 
 187 
 
THE METALLURGY OF GOLD. 
 
 A A are the main rollers that carry the belt and form the ends 
 of the table. Each roller is 50 inches long and 13 inches in 
 diameter, and is made of galvanised sheet iron. B and C are of 
 the same diameter, and are made in the same way as A A. The 
 roller part of C is shorter than that of A A and B, and also has 
 rounded edges, the upper surface of the belt with its flanges 
 passing over it. The belt E passes through water underneath B, 
 depositing its concentrations in the box, No. 4 ; and then, pass- 
 ing out of the water, the belt, E, passes over C, the tightening 
 roller. By means of the hand screws, B and C can be adjusted 
 on either side, thus tightening and also controlling the belt. 
 
 The boxes holding A A in place have slots and adjusting 
 screws, so that, by moving them out or in, A A can be made to 
 exercise an influence on the travel of the belt, E ; when, as 
 sometimes happens, it travels too much towards one side, this 
 tendency can be stopped most quickly by altering the screws 
 on one end or the other of A A ; the change of position of B or 
 C also controls the belt. 
 
 D D, &c., are small galvanised iron rollers, which support 
 the belt, E, and cause it to form the surface of an evenly 
 inclined plane table. This moving and shaking table has a 
 frame, F, of ash, bolted together, with A and A as its ex- 
 tremities. The frame is braced by five cross-pieces. 
 
 The belt, E, is 4 feet wide, and 27J feet in entire length ; it is 
 an endless belt of rubber with high flanges at the sides. 
 
 G G is the stationary frame. This is bound together by three 
 cross-timbers, which are extended on one side to support the 
 crank shaft, H. 
 
 F is supported on G G by uprights, N, &c., four on each 
 side. The bearings of A, the upper or head roller, are higher 
 than those of A, the foot roller, so that the former is a trifle 
 higher than the regular plane of the table, and the first small 
 roller, D, should be raised by a corresponding amount. 
 
 The shape of the lower or bottom bearings of the uprights, 1ST, 
 <kc., can be understood by examining b, as shown in the end 
 elevation and also partly in Fig. 426. This lower bearing, b, 
 extends across G, underneath, and is supported by a bolt passing 
 through G. A lug on the upper side and on the outside end of 
 b rests on G; and b hangs on the head of the bolt, and is kept 
 stationary by the weight of ]S" and its load. By striking with a 
 hammer the face of b shown in the elevation, b is moved, chang- 
 ing the position of the lower bearing, and thus making N more 
 or less vertical. By thus moving the lower supports of N, &c., 
 the sand corners on the belt hereafter explained are regulated. 
 
 The cranks attached to the crank shaft, H, are J inch out of 
 centre, thus giving a throw of 1 inch, which is the amount of 
 the lateral throw. I is the driving pulley that forms with its 
 belt the entire connection with the power. J is a cone pulley 
 
CONCENTRATION IN STAMP MILLS. 189- 
 
 on the crank shaft, H. By shifting the small leather belt 
 connecting J and W, the uphill travel of the main belt, E, is- 
 increased or diminished at will; the pulley, W, is moved by 
 the hand-screw, m. R, R, R are three flat steel spring connec- 
 tions bolted underneath the cross-pieces of the frame, F, and 
 attached to the cranks of the shaft, H. These springs give the 
 quick lateral motion about 200 per minute. 
 
 No. 2 is the clear water distributor, and is a wooden trough 
 which is supplied with water by a pipe, and the water discharges 
 on the belt in drops through grooves 3 inches apart. 
 
 No. 1 is the ore-spreader, which moves with F, and delivers 
 the ore and water evenly on the belt. 
 
 There is also a copper well that fits in (and shakes with) the ore 
 spreader as shown in the drawing. This is used in concentrating, 
 gold ores, for saving amalgam and quicksilver escaping from 
 the silvered plates above, and can be taken out and emptied at 
 any time. Into this well falls all the pulp from the battery. 
 Its ends are lower than the wooden blocks of the spreader, so 
 that the pulp passes over the ends of the well and is evenly 
 distributed. 
 
 For some gold ores it is desirable to use on the ore-spreader a 
 silvered copper plate the size of the spreader, and, when this is 
 used, the wooden blocks of the spreader are fastened to a movable- 
 frame on top, so that they can be removed when the plate is 
 cleaned-up, once or twice a month. No. 4 is the concentration 
 box, in which the water is kept at the right height to wash the 
 surface of the belt as it passes through. No. 8 is a section of 
 the launder to carry off the tailings. 
 
 Method of Working. The ore is fed with water on the belt, E,, 
 by means of the spreader, No. 1, which distributes it uniformly. 
 A small amount of clear water is added by No. 2. The depth of 
 the sand and water is kept constant at from to J inch. 
 
 The main shaft, H, is given the proper speed for each kind of 
 ore ; this usually varies from 180 to 200 revolutions of the crank 
 shaft per minute, the former speed being for light, fine " slimes,^ 
 the latter for somewhat coarser materials. The effect of increas- 
 ing the speed of the side shake is to increase the percentage of the 
 material which is being discharged as tailings, and so to tend to 
 loss of pyrites. The diminution of the speed, on the other hand, 
 tends to the production of concentrates containing much sand. 
 
 The speed of the uphill travel of the belt varies from 2 to 12 
 feet per minute, and the grade or inclination from 3 to 6 inches 
 in 12 feet, according to the ore. If the ore treated be poor in 
 pyrites, the upward motion of the belt should not exceed 20 inches 
 per minute ; if richer, the speed is increased accordingly, and in 
 agreement with the inclination of the belt, being greater as this 
 inclination increases, but usually not exceeding 3| feet per 
 minute. The inclination can be changed at will by wedges at- 
 
190 THE METALLURGY OF GOLD. 
 
 the foot of the machine, these wedges being under the lower end 
 of G, G, and resting on shoulders of uprights from the main 
 timber of the mill. The motion, the water used, the grade, and 
 the uphill travel is regulated for every ore individually, and 
 must unfortunately be adjusted with every change in the pulp, 
 if good work is to be maintained. 
 
 In treating ore coming direct from the stamps, too much 
 "water may possibly be present in the sand for proper treatment 
 by the machine. In such a case there should be a box between 
 the stamps and the concentrator, from which the sand with the 
 proper amount of water can be drawn from the bottom, whilst 
 the superfluous water will pass away from the top of the box ; 
 but as sulphides will also pass away with this water, settling 
 tanks should be provided, and the settlings can be worked from 
 time to time as they accumulate. 
 
 The main body of the belt suffers hardly any wear at all, 
 since it merely moves its own weight slowly around the freely 
 revolving rollers ; the life of the belt is lengthened by taking 
 the precaution of keeping it clean from sand at every point 
 except the working surface, so that sand cannot come between the 
 belt and the various rollers. 
 
 The concentration box, No. 4, which is kept full of water, and 
 through which E passes, may be of any size or depth desired. 
 Though not indispensable, it is best to have a few jets of water 
 playing above and underneath on the belt as it emerges from the 
 water in No. 4, so as to wash back any fine material adhering to 
 the belt, and as such a method will cause an overflow in No. 4, 
 the waste water, being full of finely divided mineral, should be 
 settled carefully in the boxes, Nos. 7, 7, 7. Every few hours the 
 concentrations may be scraped out with a hoe into the box, No. 
 9, and if this box be on wheels, it can be readily run on a track 
 to the place where the concentrations are stored. 
 
 The amount of water used on the machine is from 1 to 1 
 gallons per minute of clear water at the head, and from 1|- to 3 
 gallons per minute with the pulp. 
 
 The capacity per day of twenty-four hours is usually put 
 down as about 6 tons of material, fine enough to pass a 40-mesh 
 screen, but the machine does better work when set to treat a 
 smaller amount. In California, in general, two True vanners 
 treat the product of each battery of five stamps. Where the 
 gangue is light, only one Frue vanner is sometimes used for five 
 stamps. Sizing of the material is frequently omitted, the pulp 
 passing direct from the stamps to the copper plates, and thence 
 to the vanners. Resales, however, points out * that this is a 
 mistake ; the coarse sand should be removed by sizing, as other- 
 -wise it interferes with the successful conduct of the work, caus- 
 
 * Reduction of Auriferous Veinstone in Victoria, 1895. 
 
 . 
 
CONCENTRATION IN STAMP MILLS. 191 
 
 ing serious losses of sulphides. If there are large quantities of 
 slimes in the pulp, the capacity of the machines is proportion- 
 ately reduced. The machine requires careful watching by a 
 skilled workman, as any change in the condition of the pulp 
 necessitates readjustment of the belt. 
 
 It is very important to use the proper quantity of water 
 with the pulp from the stamps, and this should be carefully 
 regulated. There should be formed on each side of the belt a 
 slight corner of sand i.e., there should be, on each side, sand 
 with less water in it than there is in the remainder of the pulp 
 on the belt. If there is not a slight sand corner, the corner will 
 be sloppy, and there will be a loss. Sloppy corners are caused 
 by using too much water with the pulp from the stamps passing 
 on to No. 1. 
 
 Frequently, on the other hand, there may not be enough water 
 with the pulp from the stamps, and as a result the sand corners 
 formed will be too wide. The remedy for this is to use more 
 water in the pulp coming on to No. 1. 
 
 As regards the proper amount of water to be used in the 
 water spreader, No. 2, use just enough (no more) to keep 
 the field between No. 1 and No. 2 covered, so that no points 
 (or fingers) of sand shall show on the surface. The whole width 
 of the belt between the water spreader and ore spreader should 
 be kept quite wet. If dry streaks or points occur, and water, 
 as a consequence, runs in streaks at the junction of the wet and 
 dry channels, sulphides will be picked up and " floated " away on 
 the surface of the water; this "floating" of pyrites is caused 
 by its dryness, not by its lightness ; it has been coated with a 
 film of air. When the proper amount of water has been fixed, 
 the carrying over of the clean concentrates past the jets of .No. 
 2 should be accomplished and regulated by the uphill travel only. 
 Frequently, the sand and water on the belt will be distributed 
 unevenly, the sand working to one side of the belt, and making 
 a broad heavy corner, while the other is sloppy. This is caused 
 by the belt not being horizontal from side to side ; it is levelled 
 by raising or lowering one side, by altering the position of the 
 supports, N. 
 
 W. M'Dermott makes the following observations on the 
 appearance of the ore on the belt : 
 
 " Should the discharge of concentrates exceed the quantity of 
 sulphides falling on the belt, sand or rock will be found close 
 up to the jets of water, and by and by passing them. If the 
 uphill travel be too slow, the sulphides collect below No. 2, form- 
 ing a great "head," extending towards No. 1, and even below, 
 in which latter case an increased loss of sulphides will assuredly 
 take place in the waste. When working properly, a small head 
 is always kept below the jets of clear water, and the sulphides 
 oorne over clean and regularly." 
 
192 THE METALLURGY OF GOLD. 
 
 Riffle Surfaced Belts for the Frue Vanner. The smooth belt 
 vanner, just described, is essentially a slime machine adapted to 
 the very finest material. The special value of the vanner in 
 recovering fine sulphides without previous sizing has been illus- 
 trated as follows : * 
 
 "If a flat heavy object say a sixpence is put on a Frue 
 vanner while in operation, it will not pass over with the fine 
 material steadily delivered past the water jets. The shaking 
 motion, by cutting away the supporting adhesion of the belt 
 surface, owing to the inertia of the coin, allows the light current 
 of water on the edge of the coin to wash it backwards with the 
 sand. The same action applies to coarse particles of sand mixed 
 with fine particles of mineral. Thus the effect of mere mass is 
 such that the Frue vanner with smooth belt is not adapted to 
 saving coarse mineral j an excess of wash water at the head 
 drives down the coarser mineral, but the finest clings to the 
 belts and safely passes the water jets." 
 
 In order to overcome this defect a roughened belt was intro- 
 duced, having a number of depressions or riffles on its surface. 
 In moving down this belt the coarse sulphides and gangue pass 
 into these depressions, and then are compelled to proceed over a 
 rising surface. At these points the separation of the coarse 
 sulphides and coarse sand is perfectly performed, the heavy par- 
 ticles remaining in the depressions, while the lighter are washed 
 away. The belt then in moving upwards carries the pyrites 
 with it. In other respects the machine closely resembles the 
 older smooth belt form, to which it is inferior in its power of 
 catching slimes, although better adapted for coarse material. 
 The riffle-belted Frue vanner is worked at a greater inclination, 
 with a faster upward motion, a slightly faster shake, and with 
 more water than the ordinary vanner, and consequently it 
 treats more pulp, one machine usually taking the product of 
 five stamps. It offers great advantages over the smooth belt 
 whenever the percentage of sulphides is very high, as the riffles 
 collect and carry up much more material. 
 
 The Embrey Concentrator. This machine, like the Frue 
 vanner, consists of an endless belt with flanges along its edges ; 
 it differs mainly in having an end-shake instead of a side-shake, 
 the motion being parallel to the length and travel of the belt r 
 instead of at right angles to it. It is shown in Fig. 43, where 
 the light wooden-framed form is shown. This is cheaper but less 
 compact than the iron-framed variety, and is recommended for 
 use wherever floor space is not of the first importance. It is 
 unnecessary to describe it in detail, as it differs little from the 
 Frue vanner. It is necessary to shake this machine at a faster 
 rate than the side-shaking machines, the usual speed being from 
 230 to 240 revolutions per minute, although 200 per minute ia 
 * Gold Amalgamation, p. 71. M'Dermott and Duffield, New York, 1890. 
 
CONCENTRATION IN STAMP MILLS. 193 
 
 occasionally enough. If an attempt is made to treat an ore at 
 too low a rate of speed, complete separation cannot be effected. 
 More water is used than on the Frue vanner, while the inclina- 
 tion and rate of upward movement of the belt are both greater. 
 In consequence of this, more pulp is treated, the amount varying 
 from 6 to 10 tons per day of twenty-four hours, and, therefore, 
 either one or two machines are used to each five stamps, accord- 
 ing to the character of the ore and the closeness of concentration 
 required. Fewer Embrey concentrators than Frue vanners are 
 in use, owing to the fact that the sale of the latter is pushed 
 more by the company which owns both the patents. 
 
 Among other concentrators with endless inclined rubber belts 
 may be mentioned the Triumph, which, like the Embrey, has a 
 longitudinal shaking motion applied to the belt, the crank shaft 
 revolving at the rate of 235 to 240 times per minute. The 
 
 Fig. 43. 
 
 forward motion of the belt is regulated by a friction roller, 
 instead of a cone pulley, which is used in the Frue vanner and 
 the Embrey concentrator. It has also an amalgam saver, con- 
 sisting of an iron trough containing quicksilver, which is stirred 
 by iron teeth attached to a slowly revolving horizontal shaft. 
 It is in wide use in California (where it seems to be preferred to 
 the Embrey machine) and in Australia. 
 
 The Liihrig Vanner. This machine was invented about 
 five years ago, and has already met with great success in many 
 instances in the concentration of auriferous tailings from stamp 
 batteries. It bears a considerable resemblance to the Frue 
 vanner and is thought by many experts to be destined to 
 supersede it. In the Liihrig vanner the endless travelling 
 india-rubber band is not flanged at the sides, and has a slight 
 side inclination, which is, therefore, at right angles to the 
 direction of travel ; the latter is horizontal, not inclined ; by 
 an arrangement of cams and springs, end-blows are given to the 
 
 13 
 
194 THE METALLURGY OF GOLD. 
 
 framework carrying the belt. The pulp from the batteries is 
 collected in settling boxes placed overhead and delivered on to 
 the belt through a small distributing box situated above its 
 right-hand upper corner, the belt being driven from right to lelt. 
 Clear water is supplied through a perforated pipe fixed diagonally 
 across the belt. The pulp moves across the belt from the higher 
 side to the lower, this motion being assisted by the clear water ; 
 the light particles of gangue are washed down in this direction 
 at a faster rate than the heavy particles of pyrites. On the 
 other hand, the travel of the belt and the end-blows move the 
 ore in the direction of the length of the belt. The results of 
 the combined motions are as follows : 
 
 1. Tailings pass off the table nearly opposite the distributing 
 box at the right-hand end. 
 
 2. A middle product, containing both sulphides and gangue, 
 is delivered near the middle of the side of the belt. 
 
 3. Clean concentrates are delivered near the left-hand end of 
 the belt, having travelled the greatest distance before being 
 washed off at the side. 
 
 Each of these three products is delivered into a separate hopper, 
 and by a simple arrangement of sliding plates the exact points 
 on the belt at which the delivery of the middle product is 
 divided from those of the headings and tailings can be altered 
 so that the percentage of each product can be regulated. 
 
 Details of the dimensions, capacity, &c., of these machines are 
 as follows : 
 
 The india-rubber belt is about 4 feet wide and 19 feet in 
 total length ; the belt travels from 18 to 20 feet per minute ; 
 the total quantity of clean water used is from 4J to 5 gallons 
 per minute; the supporting framework is 5 feet 6 inches high, 
 6 feet broad, and 16 feet long; the number of percussions is 
 from 150 to 210 per minute; the extent of motion to and fro is 
 from | inch to 1J inches, according to the nature of the ore. 
 The amount of ore capable of being treated in twenty-four hours 
 is stated to be 4 to 5 tons, if the concentrates are iron pyrites, 
 and 3 to 4 tons if they are galena and blende. At the May 
 Consolidated Mine, 7 to 8 tons per day were being treated on 
 each machine in 1893. At a test run in Glasgow, 2,898 Ibs. of 
 a pyritic gold ore were treated in four hours five minutes, and 
 the re-treatment of the "middlings" occupied one and a-half hours 
 more ; this is equal to a rate of 6J tons per day. In the com- 
 pound system the middlings from several machines are delivered 
 at once to a vanner placed at a lower level, and time thus saved. 
 The power required is given by the Liihrig Company as T ^ H.P. 
 per table, and by Mr. H. D. Griffiths, A.R.S.M., of Johannesburg, 
 .as \ H.P. in the case of the machines at the May Consolidated 
 Mine.* Three single tables are required for each battery oi five 
 * South African Mng. Journ., August 19, 1893. 
 
CONCENTRATION IN STAMP MILLS. 
 
 195 
 
 stamps on the Witwatersrand, this being the same as in the case 
 of the Frue vanners on the same field. 
 
 In comparing the Liihrig vanners with other travelling belt 
 machines, the chief points to be noted are that in the former 
 case : 
 
 1. The direction of travel is horizontal. 
 
 2. Three or more different products are drawn off continuously 
 instead of only two. 
 
 3. The water and pulp flow across the belt instead of in the 
 direction of its length. 
 
 The machine resembles the Frue vanner in that alterations 
 .are required in the inclination of the belt, the rate of travel, 
 the number of blows per minute, the force of each blow, and 
 the amount of clean water and pulp, with every change in the 
 constitution of the ore, so that the machines require constant 
 supervision. The regulation of these is, however, very simple 
 in the Liihrig vanner. 
 
 The results obtained in various countries have been excellent. 
 The following is the result of a trial run on an American gold 
 ore from the Eureka and Excelsior Mine, Oregon : 
 
 
 Weight 
 Treated. 
 
 Value. 
 
 Percentage. 
 
 Raw Ore, . . 
 
 Lbs. 
 760-00 
 
 @ $10-50$ ton = $3 '99 
 
 
 I. Concentrates Heads, 
 
 II. 
 
 Tailings, . . 
 Slimes, . . . 
 Unaccounted for, 
 
 38-33 
 14-51 
 523-08 
 12-00 
 172-08 
 
 , 178-57 , = 3-42 
 2693 = -19 
 20 = -052 
 2-70 = -016 
 
 8571 1 0.0-47 
 4 . 76 }9047 
 
 1*1 
 
 40 
 
 7-82 
 
 760-00 
 
 100-00 
 
 It will be observed that 90*47 per cent, of the gold and silver in 
 the ore is saved in the concentrates ; the Frue vanner is stated 
 to have saved only 60 per cent, in this case, and, as a result of 
 the run, the Liihrig vanner was adopted at the mine. At the 
 May Consolidated Mine, Johannesburg, the best result obtained 
 on unsized pulp was the saving of 77 per cent, of the gold in the 
 concentrates, but the tailings seldom contained much more than 
 2 dwts. of gold, and sometimes only 1 dwt. This result would 
 have been better if the material had been previously classified. 
 
 In 1892, after exhaustive investigation, Prof. J. Cosmo 
 Newbery reported to the Victorian Government that many 
 of the heaps of tailings in Australia, which contained from 
 1J to 3 dwts. of gold per ton, could be concentrated by this 
 vanner, and the concentrates roasted and chlorinated at a profit. 
 Many of the vanners are now in use in Victoria. 
 
196 THE METALLURGY OF GOLD. 
 
 Hartz Jigs. These machines differ in principle from all those- 
 previously described, inasmuch as the particles are separated by 
 their fall through a somewhat deeper column of water than is 
 the case on inclined tables, while a series of blows from below, 
 causing waves moving upwards, continually brings the particles 
 into suspension, and allows them to drop again. The initial 
 period of the fall in water, during which the motion depends 
 chiefly on density, is thus continually reproduced, and the result 
 is a perfect separation of heavy from light particles of ore when 
 working on any materials except the finest pulp. Jigs consist 
 of sieves supporting beds of ore, which are completely immersed 
 in water; the ore is raised and allowed to fall by a quick 
 succession of currents of water caused by the sudden action of 
 a piston below, which is so worked that the upward movement 
 or pulsation resembles that produced by a blow, while the 
 downward movement is gradual. Under these conditions the 
 heavy particles work downwards and pass through the sieve, 
 while the lighter gangue is carried away horizontally by a. 
 stream of water introduced either from below or from above. 
 Such machines are especially suitable for coarse ores. 
 
 In the Hartz jig a layer of coarse heavy particles are spread 
 on the sieve to prevent too much of the ore from passing 
 through. The stuff is fed in regularly at the head of the 
 jig, and the strokes of the piston raise both the bed of heavy 
 particles and the ore. The heaviest grains of ore find their way 
 during the downstroke into the interstices of the bed, gradually 
 pass through it, and coming to the screen, fall through into the 
 tank below. The lighter particles cannot descend, and are 
 gradually washed over the end partition by the continuous 
 supply of water. Two products are, therefore, given, neither 
 requiring further treatment if the conditions are favourable, and 
 the machine properly adjusted. The wire meshes of the screen 
 are always much larger than the ore treated, and the bed is 
 composed of material of as nearly as possible the same density as 
 the concentrates to be obtained, and is usually from ^ to 1 inch 
 in depth. The number of strokes of the piston per minute is 
 from 60 to 80 with coarse sand of -f% inch in diam-ter, and 200, 
 300, or even 400 with very fine sand, approaching slimes. The 
 length of stroke varies under the same conditions from f to 1 inch, 
 and in the case of very fine, almost impalpable sand, the stroke 
 may be diminished till it becomes a mere tremor. Some of the 
 highest authorities on concentration have stated their belief that 
 for enriching even very fine sand, the Hartz jig is the simplest 
 and most economical machine yet invented, and requires the 
 least amount of labour.* It is obvious that it is not adapted to- 
 save rich slimes, and up to the present it has not been widely 
 used for treating gold ores. 
 
 Gallon's Lectures on Mining. Eng. edition, vol. iii., p. 103. 
 
CONCENTRATION IN STAMP MILLS. 197 
 
 Pneumatic Jig. In this machine, which was devised by 
 S. R. Krom, air is used instead of water for the separating 
 medium. The ore is fed into a tank, the discharge from the 
 bottom of which is regulated by a revolving wheel. The puffs of 
 air are delivered through a number of vertical wire gauze tubes, 
 which pass through the ore bed nearly to its top. No less than 
 450 to 500 puffs of air per minute are given, and the result is 
 that the heavy particles alone descend through the interspaces 
 between the wire gauze tubes, while the lighter material is dis- 
 charged horizontally. It is necessary for the success of this 
 machine that the ore should be quite dry, so that the particles 
 may be completely independent of each other ; as a rule the ore 
 would have to be dried in special furnaces. The air jig has not 
 passed into any extensive use, although it might succeed in hot 
 climates, where water is scarce, and where only a pneumatic 
 machine could be used. Careful sizing is necessary as a pre- 
 liminary to successful working of this machine. 
 
 Clarkson & Stanfield's Concentrator. Another dry con- 
 centrator, although one constructed on a totally different principle, 
 is the centrifugal machine, which was devised by T. Clarkson & 
 R. Stanfield. The action of this machine is based upon the 
 joint operation of the three powers centrifugal force, "atmo- 
 spheric resistance, and gravitation and the machine some- 
 what resembles a " Catherine-wheel " working horizontally. 
 Pulverised ore is fed on to the surface of a rapidly rotating disc 
 about 20 inches in diameter, provided at the periphery with a 
 raised rim, which is perforated by a multitude of small radial holes, 
 through which the ore is thrown by centrifugal force. The par- 
 ticles of rich and heavy ore, by reason of their superior inertia, 
 are thrown to a greater distance, while the worthless particles, 
 being lighter, are more quickly overpowered by the forces of 
 atmospheric resistance and gravitation, and thus fall short. 
 The ejected ore is then collected in annular compartments 
 arranged concentrically round the central disc, the process going 
 on continuously while the ore is being shot out. Means are 
 also provided for regulating either the centrifugal force or the 
 Atmospheric resistance, according to the nature of the ore. 
 
 It is stated by the inventors that in one of these machines, 
 5 feet in diameter, 50 tons of ore can be concentrated in twenty- 
 four hours, only 3 H.P. being required. In this machine, the 
 centrifugal force is * proportional to the mass of the particles, 
 whilst the atmospheric resistance is proportional to their cross 
 section. It may be shown, therefore, that all particles for which 
 the product a D (diameter x density) is approximately the same 
 will fall into the same compartment. But the value of a T) is 
 nearly the same for all particles that are equal-falling in air ; for 
 in this medium, d is so small in comparison with D (see footnote 
 on p. 176), that it may be neglected, and, therefore, the formula 
 
198 THE METALLURGY OF GOLD. 
 
 a(D 5) becomes a D. Now, the difference in size betweea 
 equal-falling bodies of different specific gravity is less for air 
 than for water ; for example, the diameters of spheres of galena- 
 and of quartz, which are equal-falling in water and in air, are in 
 the ratio of 1 : 4'1 and of 1:2-9 respectively. Hence it would 
 follow that, according to theory, this machine would be less 
 effective as a concentrator than the wet sizing machine described 
 on p. 179. However, the inventors have obtained good results 
 in trial runs on both a large and a small scale. 
 
 A complete mill on the Clarkson-Stanfield system has been 
 erected at the Castell Carndochan Mine, near Bala, North 
 Wales, for the treatment of a low-grade gold ore, containing 
 about 3 dwts. gold per ton. The mine owners have reported 
 that, in the treatment of a parcel of 700 tons of ore, about 
 12 tons of rich concentrates of two grades were obtained, con- 
 taining about 72 per cent, of the gold in the ore. The total 
 cost of drying, milling, classifying by sieves, concentrating, 
 cartage, Ac., was certified as being 5s. 6 -83d. per ton. The 
 fuel used in drying the ore was 40 Ibs. of coal per ton, and 
 the motive power is water, and therefore costs little. The 
 concentrates are sold to smelters. 
 
 Samples of ore from Coolgardie have also been successfully 
 treated, very rich concentrates and tailings suitable for cyanid- 
 ing being obtained. 
 
STAMP BATTERY PRACTICE. 199 
 
 CHAPTER X. 
 
 STAMP BATTERY PRACTICE IN PARTICULAR 
 LOCALITIES. 
 
 Battery Practice in California. The ores treated are quartz- 
 ose, containing from 3J dwts. to 1J oz. of free gold per ton, the 
 average being from 10 to 12 dwts. Some iron pyrites is also 
 present, sometimes containing traces of other sulphides. The 
 amount of pyrites varies from 1 to 5 per cent. ; it is usually 
 massive, and contains from 4 to 4J ozs. of gold per ton. The 
 pyrites is in somewhat rare instances crystalline, and is then 
 of little value. The gold is seldom visible, and is in such a 
 finely divided state that it cannot be saved by mere gravitation 
 methods, the use of mercury being essential. 
 
 The obsolete practice formerly pursued at Grass Valley, and in 
 other districts of California, has been fully described by G. F. 
 Deetken.* Up to a recent date it had not been entirely super- 
 seded,! and may be briefly described as follows : The ore is fed 
 into the batteries by hand, and no rock-breakers are used. The 
 stamps weigh 800 Ibs., and have a fall of 10 inches, the depth 
 of discharge being only 3 inches. No mercury is added to the 
 battery. After crushing, the pulp is run over a set of blanket- 
 strakes, where the ore is subjected to a rough concentration. 
 The tailings are passed through the so-called Eureka rubber, 
 where the "rusty" gold, ground between sliding plates of iron, 
 is brightened and rendered .araalgamable. Subsequently the 
 sand is run through narrow sluices, lined with amalgamated 
 copper plates, where an effort is made to catch the fine free gold, 
 although the depth and speed of the current renders the opera- 
 tion very incomplete. The tailings are then concentrated in 
 round buddies, tossing tubs, &c., and the concentrates subjected 
 to chloriuation. 
 
 The blanket concentrates are treated in an Atwood's amalga- 
 mator, in which they are made to pass through two mercury wells 
 or baths, being forced under the surface by revolving paddle 
 wheels, and then they are treated with the ordinary tailings in 
 the Eureka rubber. The skimmings of the mercury wells are 
 treated in a small slowly revolving pan, the Knox pan, in which 
 
 * Mineral Resources West of the Rocky Mountains, 1873, pp. 319-345. 
 t According to the latest reports, this method has now been abandoned 
 everywhere. 
 
200 THE METALLURGY OF GOLD. 
 
 they are ground with mercury between iron mullers and dies for 
 several hours, and the tailings from it are roasted and chlorinated. 
 
 The whole process bears a strong resemblance to the present 
 Australian practice on free-milling ores containing coarse gold. 
 It was uneconomical, costing about $2 or 8s. per ton, without 
 reckoning the cost of chlorinating the concentrates, and only 
 from 70 to 75 per cent, of the gold was extracted. 
 
 The present method of treatment* consists briefly in rapid 
 crushing in narrow mortars with heavy stamps working fast with 
 a low drop, the depth of discharge being moderate. Mercury is 
 fed into the mortar box, and the gold is saved on copper plates 
 situated both on the inside and the outside of the mortar, whilst 
 the tailings are concentrated on shaking tables. Mercury wells 
 and the various accessory amalgamating machines formerly 
 attached to most mills are now falling into disuse, and do not 
 form part of the most approved modern machinery. 
 
 The stamps weigh from 750 Ibs. to as much as 1,100 Ibs., the 
 later practice being to make them of a weight of not less than 
 950 Ibs., whilst the height of the drop has been gradually 
 diminished, until the average is now only about 6 inches. 
 According to the most modern view, from 4 to 4J inches is 
 quite enough for ordinary ores ; this height was first used at 
 the Pacific Mill of the Plymouth Consolidated Mining Company, 
 some three years ago, but other companies have been quick to 
 follow their example. The heavy low-drop stamp, delivering over 
 100 blows per minute, has a duty much higher than the more 
 old-fashioned stamps, which still form a great majority of the 
 3,500 head now in use in the State. The softer the ore the 
 lower the drop should be, the limit being passed only when a 
 good splash of the pulp can no longer be obtained, or when the 
 momentum becomes insufficient to break down the larger lumps 
 of quartz rapidly. 
 
 The mortars are narrower and less roomy than those in use in 
 Colorado, and the depth of discharge is kept constant at about 
 6 inches. Below the screen is a single amalgamated copper plate 
 about 4J inches wide, inclined outwards at an angle of about 
 45. Mercury is fed into the mortar at intervals of one hour. 
 The Blake rock-breaker and some form of ore feeder are in 
 almost universal use ; among the varieties of the latter, the 
 Champion and Tulloch's machines are most favoured, although 
 Stanford's and the roller feeders are still to be seen. The 
 screens are inclined outwards at an angle of about 10, and 
 consist either of Russia sheet iron or of brass wire cloth. The 
 horizontal burr-slot screen is perhaps that most in use, but 
 others are equally suitable for certain classes of ore. The size 
 
 * The following is a brief digest of the full account given by J. H. 
 Hammond, M.E., entitled "The Milling of Gold Ores in California," 
 published in the Eighth Report of the California State Mineralogist, 1888. 
 
STAMP BATTKRY PK AC I ICE. 201 
 
 of slot most in use is No. 6, which is '027 inch in width, with 
 about 180 holes to the square inch. The stamping is finer with 
 low grade ores, because the gold in these is in a finer state of 
 division, and the object of the stamping is to release the particles 
 of gold from the rock in which they are enclosed, without re- 
 ducing the ore to an unnecessarily finely divided condition. The 
 crushing is kept coarse if the sulphides contain an appreciable 
 percentage of the gold, since the finer the stamping the more 
 difficult close concentration becomes, and, as the sulphides con- 
 stitute the most brittle portion of the ore, they are always 
 reduced to a finer state of division than the gangue in which 
 they are enclosed. In general, it is better to crush too coarsely 
 than too finely, for if the latter error is fallen into, not only is 
 the output diminished and a larger proportion of the sulphides 
 reduced to slimes, with the result that more is lost in the tail- 
 ings, but the particles of gold are supposed to be subjected to 
 repeated hammering, and converted partly into non-amalgam able 
 and partly into "float" gold, and so lost in great measure. This 
 question of over-stamping has already been discussed, p. 147, and 
 an opinion there stated that the evils caused by it have been 
 greatly overrated. Most of the ores are crushed much finer 
 than might be supposed from the size of the meshes of the 
 screen. If the ores are made to pass through a 30-mesh screen, 
 it is found on trial that more than 80 per cent, of the material 
 will usually pass through a 60-mesh screen also, and 50 per cent, 
 will pass through a 120-mesh screen, much of it being impalpable 
 slime. The low drop, already referred to, acts in the direction 
 of minimising the quantity of slime, and has a most beneficial 
 effect in facilitating the concentration of the sulphides. 
 
 The high speed of the stamps, besides increasing the crushing 
 capacity, is advantageous on account of the good splash thus 
 created, which is beneficial to the battery amalgamation, besides 
 being essential to rapid discharge. The duty of the stamps is 
 on an average 2^ tons per head per day, but is considerably 
 more in the most modern mills. It is lower on clayey or on very 
 hard ores, and cannot conveniently be increased beyond about 
 4 tons, since the battery is an amalgamating as well as a crush- 
 ing machine, and the greatest output may not be compatible 
 with the highest percentage of extraction. The amount of water 
 used is from 1,000 to 2,400 gallons per ton, most being required 
 with clayey and heavily sulphuretted ores. In addition to this 
 from 1 to 2 gallons per minute are supplied to each concentrator 
 for feed water, and from J to 1 gallon per minute for wash water. 
 
 The amalgamating plates outside the battery are in three sets, 
 the apron plates, the tables, and the sluice plates ; the united 
 length of these varies from 10 to 30 feet, according to the re- 
 quirements of the ore. They are almost invariably made of 
 copper which is electro-plated with silver. The details con- 
 
202 THE METALLURGY OF GOLD. 
 
 cerning the preparation and care of these plates have already 
 been given, pp. 124-127. The coarse gold is chiefly saved 
 inside the battery ; the finely divided particles, forming a fluid 
 amalgam, are splashed through the screen and caught on the 
 outside plates. From 50 to 80 per cent, of the gold is caught 
 inside the battery, and the amount of free gold lost in the 
 tailings is not more than 5 to 8 per cent, in the best mills. The 
 tendency is in the direction of using more feed water with a 
 lower inclination to the plates, as by this means the ore is 
 brought into more intimate contact with them, being rolled 
 over and over instead of swimming down them, and the danger 
 of scouring is diminished. Other details concerning Californian 
 stamp mills have been already given in the general description 
 in Chapters vi. and vii. 
 
 Sizing of the tailings before concentration is not attempted, and 
 this point deserves more attention than it has hitherto received 
 at the hands of Californian millmen. Moreover, means should 
 be adopted to get rid of the free gold and floured amalgam con- 
 tained in the sands before passing them over the shaking tables, 
 which are ill adapted for catching these materials. The two 
 concentrators in most general use in the State are the Frue 
 vanner and the Triumph concentrator, both of which have been 
 already described. Next to these the Golden Gate and the 
 Duncan machines are favoured, while Hendy's pan is still to 
 be found doing good work, although it was one of the earliest 
 contrivances used in the Pacific States. 
 
 The concentrates are treated by roasting and chlorination, 
 generally by the vat process. Barrel chlorination has as yet 
 made little headway in California, owing in great part to the 
 fact that only small quantities of sulphides are produced at any 
 one mill, whilst the distances between the establishments are 
 very great. Chlorination costs about $10 per ton, and an 
 average of 92 per cent, of the gold is thus extracted from the 
 concentrates, which are usually worth from $80 to $100 per ton. 
 After concentration, the tailings are in many cases run through 
 a length of 100 to 200 feet of blanket sluices, where an effort 
 is made to catch the free gold, amalgam, and sulphides still 
 remaining in the tailings. These sluices are cheap and require 
 little attention, but only account for a small percentage of the 
 total gold recovered. The blanket sands are ground with 
 mercury in a Chilian mill or arrastra, driven by water or steam 
 power. 
 
 An excellent addition to a mill is afforded by the automatic 
 tailings sampler erected at the original Empire Mill. This 
 machine, by an arrangement of cogs, dips a bucket into the 
 falling stream of tailings at stated intervals and takes a sample ; 
 all these samples are kept separate. They are usually taken at 
 intervals of an hour. Most mills sample their tailings on very 
 
STAMP BATTERY PRACTICE. 203 
 
 rare occasions. The average percentage of extraction does not 
 fall below 85 per cent, in any but the oldest and worst 
 appointed mills in the State, or in those running on unusually 
 refractory ore. 
 
 The usual cost of milling in a 40-stamp mill, treating 80 tons- 
 per day, is as follows : 
 
 Cost per Ton. 
 1. Labour 
 
 (a) In the mill, 20 cents. 
 
 (b) Assaying, retorting, melting, and 
 superintendence, . . . . 24 ,, 
 
 II. Supplies 
 
 Castings, . . . . . . 7 to 10 cents. 
 
 Mercury, . . . . . . 1} to 4 
 
 Lubricants, screens, illuminants, and 
 
 miscellaneous, . . . . 4 to 8 ,, 
 
 Total, . . 354 to 45 
 
 or, Is. 6d. to Is. 10d. 
 
 The cost of water power, if it is purchased, must be added to this, and, if 
 steam is used, about 11 cents per ton must be added for additional labour, 
 oil, &c., besides the cost of fuel. Interest on capital, and the cost of 
 chlorination are not included in this estimate. 
 
 The power required in such a mill is distributed as follows : 
 
 Rock-breakers, 12 H. P. 
 
 40 stamps, 66 
 
 16 concentrators, ....... 8 
 
 8 shaking tables, 2 
 
 1 clean-up pan, ....... l 
 
 1 clean-up amalgamating barrel and batea, . 2 
 
 Total, . . . 92 
 
 The poorest ore which can be mined and milled at a profit 
 must contain from 2 to 5 dwts. of gold per ton. 
 
 In 1889 there were about 3,000 stamp heads existing in 
 California, of which not more than about 2,000 were in active 
 operation, crushing 4,000 tons of ore per day. Of these stamps 
 about 60 per cent, were run by water power, and the remainder 
 by steam. The cost of milling varied from 39 cents to $2 per 
 ton, the mean cost being about $1 per ton.* In 1892, about 
 4,000 head of stamps were in position in California, 3,500 of 
 which were at work. The lowest cost of milling was that at the 
 Dalmatia Mine, El Dorado County, where the total cost of 
 mining and milling was 43 cents per ton. In 1851, when quartz 
 mining was first begun at Grass Valley, no ore was considered 
 worth treating unless it yielded $40 per ton by the wasteful 
 process then in vogue, f In 1873, according to Deetken, the cost 
 had fallen to a minimum of $2 per ton. 
 
 * Ninth Report Gal. State Min., 1889, p. 25. 
 t Seventh Report Cal. State Min., 1692, p. 19. 
 
~204 
 
 THE METALLURGY OP GOLD. 
 
 Method of Working employed in the Blackhawk Mills, 
 <3-ilpin Co., Colorado. The ore obtained in this district and 
 supplied to the mills is often called refractory, although the 
 greater part of the gold contained in it is extracted by amal- 
 gamation. It is refractory in the sense that it yields its gold 
 less readily than the typical Californian free-milling ores. It 
 contains, in general, from 10 to 20 per cent, of metallic sul- 
 phides, and from 15 to 20 per cent, of quartz, the remainder 
 
 Fig. 44. 
 
 of the gangue being chiefly felspathic matter derived from 
 the country rock. The metallic sulphides are chiefly copper 
 and iron pyrites, but antimonial grey copper, arsenical pyrites, 
 and galena are also present, and blende and carbonate of iron 
 occur sometimes. The greater part of the gold is free, being 
 enclosed in the quartz, but about 25 per cent, of the gold and 
 33 per cent, of the silver are locked up in the sulphides, which 
 are saved by concentration, and shipped to the neighbouring 
 
STAMP BATTERY PRACTICE. 205 
 
 smelters in Denver. The gold is in a very finely divided state,, 
 the ore often giving no "colours" when subjected to ordinary 
 panning, although in the mill it will yield a good return. Th& 
 average milling ore contains about J oz. of gold and 2 or 3 ozs. 
 of silver per ton. 
 
 The method used consists briefly in slow crushing and 
 amalgamation in the battery, where as much gold is saved as 
 possible, followed by the passage of the pulp successively over 
 a set of amalgamated plates, then over blankets, and finally over 
 concentrating machines, the tailings being allowed to run into 
 the stream. A sectional view of the battery used is. shown in 
 Fig. 44. It will be observed that only a shallow light iron, 
 mortar, a, is used, the sides consisting of a housing of wood. 
 The water is introduced by the pipe, b, and allowed to run down 
 the stem of the stamp. The ore after passing the screens flows 
 over the amalgamated plates fixed on the tables, c. The stamps 
 are light, weighing from 500 to 600 Ibs. each, and fall at the 
 extremely slow rate of from 26 to 32 drops per minute, the 
 height of the drop being from 16 to 18 inches. This is greater 
 than that adopted in any other part of the world, being four 
 times as great as that adopted in the latest Californian practice. 
 The depth of discharge is 13 inches when new dies have just 
 been put in, and increases to 15 or 16 inches as they wear down. 
 According to the most modern view, however, this depth should 
 be diminished, and at the mill latest erected it is only 11 inches. 
 The shoes and dies are small (8 inches in diameter) owing to the 
 lightness of the stamps, and the screens are fine, a burr slot 
 being used which corresponds to a 50-mesh wire screen. The 
 screen surface is normal, the dimensions being 4J feet by 8 inches 
 for each battery of five stamps. Under these conditions the 
 output is of course low, being on an average about 1 ton per 
 stamp head in twenty-four hours. The feeding is done by hand, 
 one man feeding twenty-five stamps, and there are no rock- 
 breakers, their absence being accounted for by the age of the 
 mills and the lack of fall in the ground in the rear of the 
 batteries, which interferes with their introduction at the present 
 day. Nevertheless, rock-breakers and automatic feeding would 
 reduce the cost of crushing considerably, as the cost of hand- 
 feeding alone is $450 per battery in a year, while the first cost 
 of automatic feeding machines would only be from $500 to 
 $1,000, and the cost for maintenance, power, and repairs nominal. 
 The amount of water added to the battery is from 1J to 2 
 gallons per stamp per minute. The amount of mercury used 
 varies of course with the richness of the ore. "When ore con- 
 taining \ oz. of gold per ton is being crushed, about 2J ozs. of 
 mercury are added to each battery per day in small quantities 
 at intervals of an hour, while about an equal amount (2 ozs.) is 
 required to dress the plates, both inside and outside. Of this 
 
"206 THE METALLURGY OF GOLD. 
 
 quantity one-third is put on the back inside plate, one-fourth on 
 the front inside plate, and the remainder on the outside plates. 
 The loss of mercury from all sources is about 20 or 25 per cent, 
 of that added each day, and, with ^ oz. ore, is about -J- oz. of 
 mercury per ton crushed. 
 
 Two sets of amalgamated plates are used inside the battery, 
 each 4^ feet long, the front one being 6 inches wide, and the 
 back one 12 inches. The front plate is nearly vertical, but the 
 back one is set at an angle of 40. There is only a single set of 
 plates outside, which are 4 feet wide by 12 feet long, and have 
 the high grade of 2J- inches per foot. The plates are dressed 
 every twelve hours with a weak solution of cyanide of potassium 
 (2 ozs. in 3 gallons of water), this being the only chemical used 
 on the mill. The consumption of cyanide is about 2 ozs. per 
 battery of five stamps in three days. Its use is necessitated by 
 the acid nature of the water, which already contains sulphates 
 : and carbonates when it comes from the mines, and is further 
 contaminated by ferrous sulphate, formed by the oxidation of the 
 pyrites in the ore and then dissolved. This water soon forms a 
 Him on the surface of the copper plates, and also rapidly corrodes 
 the screens, which otherwise, from their unusual height above 
 the dies, would last a long time. Nevertheless no effort is made 
 to counteract the bad influence of this water by adding alkalies, 
 limestone, or lime water to it or to the ore. 
 
 After leaving the plates the ore passes over the "blanket- 
 strips," which are 3 feet long and 18 inches wide. They are 
 washed every four hours, or every two hours if the ore is rich. 
 The blankets serve to catch any escaping mercury, amalgam, 
 "rusty" gold, and the heaviest pyrites, together with pieces of 
 rock to which gold is attached. Probably this work would be 
 done equally well by the concentrators, by which an operation 
 would be saved. The blanket-sands are sold to the smelters with 
 the other concentrates. The concentrators consist of shaking 
 tables, constructed in the locality ; they have been already 
 described at p. 184. The concentrates usually contain about 
 15 dwts. of gold and 5 ozs. of silver per ton. 
 
 The amalgam obtained is divided fairly equally between the 
 front inside plate, the back inside plate, and the outside plates. 
 In retorting it the interior of the iron retort is chalked or lined 
 with paper to prevent the gold sticking to it, the latter method 
 being considered the better. The balls of dry squeezed amalgam 
 are put into the retort, broken with an iron rod and then 
 pressed down until hard and uniform. A large bolt with 
 a nut at the end is often used for the purpose. The cover is 
 then put on and luted down with clay. The variety of retort 
 used is similar to the bell-shaped one mentioned on p. 140. One 
 hundred parts of amalgam yield from 33 to 50 parts of retorted 
 metal, the average being about 40 ; the amount depends on the 
 
STAMP BATTERY PRACTICE. 207 
 
 richness of the ore and the coarseness of the gold. The bullion 
 contains from 750 to 850 parts of fine gold, and almost all the 
 remainder is silver, about 10 parts per 1,000 being base metal. 
 According to T. A. Rickard,* the late manager of the New 
 California Mine and Mill, from whose descriptions many of the 
 details already given have been taken, the results obtained by 
 his mill on an ore containing 10 dwts. of gold and 2J ozs. of 
 silver per ton, were as follows : 
 
 
 Gold. 
 
 Silver. 
 
 Saved on the plates, 
 Saved in the concentrates, 
 Lost in the tailings, 
 
 70 '4 per cent. 
 23-4 
 6-2 
 
 42 '6 percent. 
 31-4 
 26 
 
 The cost of milling was 84 cents per ton in 1890, and 78 cents 
 per ton in 1891. This did not include the cost of shipping and 
 smelting the concentrates. Power is supplied for four months 
 in the year by water, for four months (in the dead of winter) 
 by steam, and for the remaining four months by the two com- 
 bined. Labour in 1890 cost $3 per day per man ; in 1893, 
 about $2.50. 
 
 In this locality, the slow drop and deep discharge are more 
 accentuated than in any other part of the world, although, in 
 Calaveras County, California, an almost identical course was 
 formerly pursued with the refractory ores found there. 
 
 The pulverisation is even finer than is indicated by the size of 
 the screen slots, owing to the retention of the pulp in the mortar 
 after it has been crushed, which is due to the small area of dis- 
 charge as compared with that of wire screens, and to the depth 
 of discharge and the slowness of the drop. A large percentage 
 of the pulp is fine enough to pass through a 100-mesh sieve, the 
 pyrites being especially finely crushed. This fact is of import- 
 ance from two points of view. On the one hand, as the fine free 
 gold is chiefly in intimate association with the pyrites, the fine 
 crushing of the latter enables the gold to be separated from it by 
 the mercury, and this result is advantageous ; but, on the other 
 hand, the excessive fineness of the pyrites makes it exceedingly 
 difficult to separate from the ore by concentration, and, although 
 the shaking tables do excellent work and save a fair proportion, 
 a somewhat serious percentage is perforce allowed to run to 
 waste. So difficult to catch is it, that it is carried many miles 
 by the stream, and can always be panned out of the sand 
 which is deposited where Clear Creek debouches on the plains, 
 at a distance of over 30 miles from the mills. 
 
 * Eng. and Mny. Journ., Sept. 10, 1892. 
 
4 J08 THE METALLURGY OF GOLD. 
 
 The use of the slow drop is intended to enable the gold and' 
 pyrites to settle a little by gravity through the lighter parti cles- 
 of gangue between each blow, and so to keep them longer in the 
 battery, and thus increase the chances of amalgamating the gold. 
 The wide roomy mortar prevents a violent splash against the 
 screens and plates, so that little scouring of the latter takes 
 place. There can be no doubt that, on the whole, the method of 
 working in Gilpin County, which was slowly and painfully 
 elaborated about twenty years ago, is the best that could be 
 applied to the particular ores which occur in the district. 
 
 Battery Practice on Free-Milling Ores in Australia and 
 New Zealand. Australian practice as a whole, if some of the 
 new mills in Queensland are excepted. owes little to the experi- 
 ence gained in the United States, and pursues widely different 
 methods from those in use there. Modern methods in both 
 countries are based almost entirely on the processes in use in 
 Central Europe in the first half of this century, and have been 
 modified to some extent in different directions in various 
 localities, the changes being in most cases beneficial, and 
 especially suited to the class of ore in course of treatment. The 
 process employed on the free-milling coarse-gold ores in Victoria, 
 New South Wales, and New Zealand has perhaps undergone less 
 variation from the original Transylvanian type than that 
 employed in any other part of the world, chiefly owing to the 
 extreme simplicity of these ores, and the ease with which a high 
 percentage of the gold can be extracted. 
 
 The ores consist of white quartz, containing large grains and 
 plates of remarkably pure gold, which is thus often visible, 
 and sometimes occurs in big pieces weighing several pennyweights. 
 Inclusions of country rock (slate, &c.) are common, and in these 
 cases the sulphides often slightly increase in quantity, but rarely 
 form more than from ^ to J per cent, of the rock, and consist 
 chiefly of iron or arsenical iron pyrites. The free gold is riot 
 intimately associated with these pyrites, and the ore may even 
 fall off in value when the quantity of sulphides increases. The 
 poorest ores treated contain 4 or 5 dwts. of free gold per ton, an 
 amount which includes from 85 to over 99 per cent, of the total 
 gold contents, and the concentrates contain from 10 dwts. to 
 4 ozs. or more of gold per ton. These concentrates are saved in 
 some districts and treated by roasting and chlorination, while in 
 other places they are allowed to run to waste, especially if they 
 contain less than 1 oz. of gold. 
 
 The method of treatment consists in crushing the ore somewhat 
 coarsely in rectangular mortar boxes, with stamps of medium 
 weight, running at a medium rate of speed, and then in passing 
 the pulp through mercury wells and over blanket strakes. The 
 mortar sand, which is collected periodically, is amalgamated in 
 revolving barrels, together with the blanket concentrates, but no* 
 
STAMP BATTERY PRACTICE. 209 
 
 mercury is added to the mortars, and amalgamated plates are 
 conspicuous by their absence. The following details concerning 
 the machinery, <fec., employed at certain mills in Victoria and 
 New Zealand are chiefly derived from T. A. Rickard's 
 writings* on the subject, to which the reader is referred for a 
 more complete account. 
 
 Rock-breakers and automatic feeding machines are seldom 
 used, and the hand-feeding is often done very carelessly. The 
 mortar boxes are of rectangular shape, have no amalgamated 
 plates inside them, and are considered by American experts to be, 
 in many cases, unsuited for internal amalgamation, as the mer- 
 cury and amalgam do not collect out of reach of the falling 
 stamps, which would, therefore, cause flouring. It is in this 
 way that the fact may be explained that a smaller saving of 
 gold is effected when mercury is fed into the mortar boxes, than 
 when the customary practice is adhered to, since the mercury and 
 amalgam are subjected in the former case to conditions under 
 which excessive flouring takes place, and the mercury is lost in 
 the tailings, together with such gold as it may have taken up. 
 On the other hand, Australian millmen point out that the 
 mortar boxes are spacious enough for the retention of the coarse 
 gold which collects by gravity in them, without being pounded 
 to fragments by the stamps, and is removed at intervals by a 
 scoop, and that the collection of this gold is retarded and not aided 
 by the addition of mercury. The mortar boxes are from 14 to 
 16 inches wide, and spaces are usually left between the dies, 
 and between these and the ends and sides of the boxes, which 
 would certainly be far too roomy if the battery were intended 
 only for a crushing machine and not as an apparatus for collecting 
 coarse gold also. Australian managers prefer blankets to copper 
 plates, their reason for this preference, as well as for the reten- 
 tion of their mortar boxes, being based, no doubt, on the coarse- 
 ness of the gold to be saved. 
 
 The weight of the stamps is usually about 800 or 900 Ibs., 
 8 cwts. being a favourite weight ; the diameter of the shoe is 
 10 inches, the height of the drop is 7| or 8 inches, and the speed 
 about eighty blows per minute. The depth of discharge is low, 
 being usually about 4 inches, and this is kept constant by packing 
 up the die with sand or by putting in thin iron plates below it, 
 as it is worn down. The screens are round-punched Russia iron 
 or sheet copper, the latter lasting much longer ; they are pierced 
 with from 81 to 180 holes per square inch. The screens are of 
 greater area than those used in America, being about 12 inches 
 in vertical height instead of 8 inches. Double discharge is often 
 used and would always be advantageous, the duty being from 
 2J to 3J tons per stamp per day at the mills where double dis- 
 
 * "Variations in Gold Milling," Enq. and Mng. Journ., Jan., Feb.. and 
 March, 1893. 
 
 14 
 
210 
 
 THE METALLURGY OF GOLD. 
 
 charge is used, and little more than half this amount in the 
 single discharge batteries. The amount of water used is very 
 large e.g., 8 gallons per minute per stamp head at Clunes, 
 where there is a double discharge and where a low inclination 
 is given to the blankets. This quantity of water is four or five 
 times as much as that supplied in California. 
 
 A double discharge battery in use at the South Clunes United 
 Mill is shown in section in Fig. 45 ; A and B are the screens, 
 and the launder, C, conveys the discharge from the back to mix 
 with that coming from the 
 front screen. The perfor- 
 ated iron plate, D, serves 
 to distribute the ore evenly 
 over the apron, E, over 
 which the pulp flows ; the 
 latter subsequently passes 
 through the mercury wells, 
 F and G, being compelled 
 to pass through the bath of 
 quicksilver by the vertical 
 boards, H and K. The 
 wells are made of iron, 
 which keeps the mercury 
 lively, and are 3 inches in 
 width and 4 inches deep ; 
 the ore has a fall of 10 
 inches upon the surface of 
 the quicksilver. Each well 
 contains about 50 Ibs. of 
 mercury, and is usually 
 cleaned up once a week by 
 passing its contents through 
 canvas and so separating the 
 amalgam ; the wells are also 
 skimmed at short intervals. 
 They are omitted altogether 
 in some mills, gold being 
 
 Fig. 45. 
 
 saved only by gravity in the mortar and by concentration on 
 blanket strakes. 
 
 The blanket tables, L, should have a united width equal to 
 that of the screens ; they are usually divided into widths of 18 
 inches, and consist of three or four lengths of 6 feet each, over 
 which the ore passes in succession ; the grade at which they are 
 placed is from 1 to 1 inches per foot. Green baize is the material 
 most commonly used, the colour enabling specks of gold to be 
 readily seen. The blankets in the top row are washed in tubs 
 every half hour or hour, according to the richness of the ore, and 
 those in the lower rows at longer intervals. The blankets are 
 
STAMP BATTERY PRACTICE. 211 
 
 sometimes succeeded by concentrators, which usually consist of 
 some modification of the Cornish huddle. At some mills the 
 experiment has been made, with considerable success, of replacing 
 the lowest row of blankets by amalgamated plates. The blankets 
 are well adapted to catch coarse gold, but the fine gold escapes 
 them, and is caught on the plates, which are efficient for this 
 purpose, though perhaps of no greater utility than blankets 
 in arresting coarse particles. 
 
 In cleaning-up, the screens are taken out, the contents of 
 the mortars are passed through a No. 4 sieve, and the material 
 that remains on this is returned to the battery and used to pack 
 round the dies. The fine material one or two buckets full from 
 each mortar box at each fortnightly clean-up contains a quantity 
 of free gold, and is treated with mercury in the amalgamating 
 barrel, together with the blanket concentrates. The skimmings 
 from the wells, and occasionally the blanket concentrates also, 
 are treated in Berdan pans, and the concentrates from the 
 buddies are either roasted and amalgamated in Chilian mills, 
 or, more frequently, roasted and chlorinated in revolving barrels. 
 
 The amalgamating barrel, mentioned above, is hung on trun- 
 nions, and revolved at the rate of about 16 revolutions per 
 minute. From 5 to 20 cwts. of concentrates, battery sand, &c., 
 according to the size of the barrel, are charged in together with 
 75 Ibs. of mercury and some cold water, the chemicals used 
 being sulphate of copper and wood ashes. After running for 
 about ten hours, the barrel is discharged into a tank and its 
 contents allowed to settle, the process being sometimes aided by 
 passing the pulp through a series of wells, as, for example, at the 
 South Chines United Mill where there are three drops, the 
 heights of which are 12, 9, and 6 inches respectively, the 
 material falling each time into a bath of mercury. The battery 
 sand is usually kept separate from the blanket concentrates, as 
 it is much richer than the latter. The tailings from the barrel 
 and the Berdan pans are sometimes run over a few feet of copper 
 plates to catch any finely divided amalgam that may be con- 
 tained in them, and are sometimes concentrated on Eittinger 
 shaking tables. 
 
 The percentage of gold recovered varies greatly according to 
 the nature of the ore and to the method of treatment adopted. 
 At many of the mills 110 steps are taken to ascertain the extent 
 of the loss, and, in the best mills, only from 80 to 85 per cent, of 
 the gold is saved. More than half of the gold which is recovered 
 comes from he mortar boxes, where it is merely collected by 
 gravity, and the remainder is derived mainly from treating the 
 blanket concentrates. At Clunes, where mercury wells are used, 
 the relative amounts of gold obtained from the various appliances 
 are approximately as follows : From the ; mortar box, 45 per 
 cent.; from the wells, 30 per cent.; from the blankets, 15 per cent.; 
 
212 THE METALLURGY OF GOLD. 
 
 and from the skimmings of wells, 10 per cent. At Otago, where- 
 mercury wells are not used, the distribution is as follows : 
 Mortar box, 60 per cent. ; blankets, 33 per cent. ; copper plates,, 
 7 per cent. 
 
 The loss of mercury would be very small, if it were not for 
 the flouring that takes place in the treatment of the skimmings, 
 ' &c., in Berdan pans. Loss by flouring also takes place if the 
 amalgamating barrel is run too rapidly. Owing chiefly to this 
 latter cause, the loss per ton of ore crushed is 8 dwts. of mer- 
 cury at Otago, while at Climes it is about 3 dwts. At the 
 South Clunes United Mill the loss for the eight years (1884-92) 
 averaged only 5J grains of mercury per ton, this being probably 
 the smallest loss on record in a stamp mill. This fact proves the 
 often-repeated assertion that a bath of mercury can be mani- 
 pulated with very little loss. If metallic copper be present in 
 the ore, as is sometimes the case, the copper amalgam formed 
 floats on the surface of the mercury, a.nd is readily carried away 
 with the tailings, the loss being thus increased to 2 or 3 ozs. of 
 mercury per ton of rock crushed. The cost of treatment is, in 
 some well appointed mills, only from 2s. to 2s. 6d. per ton. 
 
 Method of Working in the Thames Valley, N.Z.* The 
 ore in this district may be regarded as on the border line 
 between "refractory" and "free-milling" ores. It varies greatly 
 in hardness and composition, and consists of " stringers " of 
 quartz running through a more or less decomposed and brecci- 
 ated hornblende-andesite. The gold is often in visible specks 
 and threads in the quartz, but is also largely associated with 
 copper and iron pyrites, blende, galena, stibnite, &c., the quan- 
 tity of sulphides varying from J to 10 per cent, of the ore, and 
 averaging from 2 to 3 per cent. Silver occurs as argentite, fcc.,, 
 while tellurides of both gold and silver are sometimes found. 
 The average amounts extracted from the ordinary ore are gold, 
 5 to 10 dwts. per ton, silver, 2 to 5 dwts.; but a considerable 
 percentage of the gold recovered comes from " specimen " ore, 
 which is often worth some thousands of pounds per ton. 
 
 No rock-breakers are used, grizzlies or sizing screens being also 
 unknown, and the feeding is done by hand. The small stuff is 
 shovelled into the mortar, and big pieces of rocks are rolled into 
 the feed opening, and rammed in by a few blows of the sledge 
 hammer, if they stick there. The results of this primitive prac- 
 tice are seen in the excessive rate of wear of the shoes and dies 
 i.e., 22 ozs. of iron per ton of ore crushed, 14-5 ozs. for the shoe, 
 and 7-5 ozs. for the die. The feeding is very high, the mortar 
 boxes being choked up with ore, so that the output is still further 
 reduced from this cause. The mortar box is spacious, approaching 
 in design the Gilpin County pattern, but, as no attempt is made 
 
 * The following account is partly based on the description given by T. A- 
 Rickard in the Engineering and Mining Journal, Dec. 3 and 10, 1892. 
 
STAMP BATTERY PRACTICE. 213 
 
 to amalgamate the ore before it leaves the battery, a narrower 
 shape might be used with advantage. The stamps weigh from 
 620 to 840 Ibs., the faces of the shoes and dies being 9 J inches in 
 diameter, while the dies are 4 inches deep. Both are made 
 of white cast iron, only the shoes being chilled. The height of 
 the drop is 8 or 9 inches, and the number of drops per minute 
 is 75. The depth of discharge, though varying as the dies wear, 
 is on an average only from 2 to 3 inches, the bottom of the screen 
 being placed on a level with the surface of the pulp in the mortar, 
 or even below it, the bottom being on a level with the top of the 
 dies in one mill. The screens are placed in a vertical position, 
 and are made of round-punched Russia sheet iron, having from 
 148 to 180 holes per square inch. 
 
 The wearing of the screens is very rapid, owing partly to the 
 small depth of discharge, the ore being flung almost horizontally 
 from the surface of the dies and striking the screens with great 
 force, but still more owing to the acidity of the battery water. 
 A grating usually lasts about six days, or while about 50 tons of 
 ore are crushed. Copper gratings might be used with advantage. 
 The mine waters contain an unusual quantity of the proto- 
 sulphates of iron, copper, manganese, and aluminium, and, 
 consequently, when the millstuff is very wet, the screens are 
 corroded with great rapidity, while the soft surface ore, in which 
 the sulphides have been largely decomposed, wears out the 
 screens faster than the hard deep-level ore. It is unfortunate 
 that lime water and limestone are not readily obtainable in the 
 district, as an addition of either of these would undoubtedly 
 lengthen the life of the screens. The output is from 1^ to If 
 tons per stamp per day, which is low, considering that no battery 
 amalgamation is attempted. A fast-running heavy stamp, with 
 double discharge, would probably double the capacity of the 
 mills, and leave the gold in a better condition for amalgamation. 
 
 The pulp, after being ejected from the mortar, is run over 
 .amalgamated tables, 7 feet long and 4J feet wide. On the 
 upper part of these tables, next the battery, is a plate 2J feet 
 long, succeeded by a well 2J inches wide and 2 inches deep, filled 
 with mercury. Below this is another length of 18 inches of 
 amalgamated plate, and then a succession of three more "ripples" 
 or wells, not filled with mercury. The total length of the plates 
 is thus only 4 feet, and the only other means adopted to catch 
 the gold consist in the use of the ripples and some blanket 
 strakes, by which some of the free gold and pyrites are caught. 
 After passing over the blankets, the tailings are allowed to run 
 into the sea. The " blind ripples " i.e., those not containing 
 mercury are cleaned with a scoop every half hour, the heavy 
 sand and pyrites so obtained going to the pans together with the 
 material caught on the blankets, which are washed every hour 
 or even oftener. The blankets cost 6 shillings per square 
 
214 THE METALLURGY OP GOLD. 
 
 yard, and last for three months. The plates consist sometime* 
 of copper, but more usually of Muntz metal ; they are cleaned, 
 wiped, and dressed every four hours, the wells being skimmed 
 with a cloth at the same time. The mercury in the wells is 
 squeezed through wash leather once a week. 
 
 The use of the mercury wells serves but little purpose as, in 
 accordance with general experience elsewhere, the sulphides 
 cause a scum to form over the surface of the bath after a few 
 minutes, and no further amalgamation takes place until it is- 
 skimmed off. 
 
 The pans used for the treatment of the blanket concentrates 
 are chiefly Berdans, with a few Watson and Denny's. The 
 Berdan pans are furnished with a drag, fixed to one side 
 of the sloping bottom, instead of a ball, as is usual. This is a 
 useful modification, as grinding and amalgamation are kept 
 separate, while in the ordinary Berdan pan, the ball remains at 
 the bottom of the cone where the mercury collects, and is th& 
 cause of a considerable loss by flouring. The drag consists of two 
 parts, the lower part being the slipper or shoe, weighing 196 Ibs., 
 to which the boss or top, weighing 230 Ibs., is rigidly keyed. The 
 shoe wears out in four months. The Berdan pans are 4| feet in 
 diameter, with a depth of 9 inches, the bottom having an inclina- 
 tion towards the centre of 1 in 2 ; they work at 23 revolutions 
 per minute, and the amalgam is removed and strained every 
 twenty-four hours. 
 
 The greater part of the amalgam is obtained from the plates, 
 the percentages being from the plates about 75 per cent., from 
 the wells 5 to 10 per cent., from the pans about 15 per cent. 
 One of the features of the district is the " specimen " stamp 
 attached to each mill, by which the valuable specimen ore is 
 crushed, the gold being caught chiefly in the mortar box, while 
 all the tailings are saved and treated in the pans. 
 
 The amalgam yields about 40 to 50 per cent, of retorted metal,, 
 which on melting yields bullion about 600 to 675 fine in gold, 
 the remainder being almost entirely silver. The loss of mercury 
 is high, owing to excessive flouring in the pans ; it varies from 
 15 dwts. to 1 oz. per ton of ore crushed. The cost of working 
 is about four shillings per ton, water power being used and paid 
 for. 
 
 The absence of concentrators to save the sulphides renders 
 the process somewhat wasteful. The sulphides from these tail- 
 ings often contain from 15 to 25 dwts. of gold per ton, and a 
 much greater amount of silver. Even the free gold is stated to 
 be by no means completely extracted, nearly 50 per cent, passing 
 away in the tailings. There is certainly an accumulation of 
 tailings on the foreshore estimated to amount to over 1,000,000 
 tons, which are believed to average about J oz. of bullion per 
 ton. A few years ago a large number of samples were collected 
 
STAMP BATTERY PRACTICE. 215 
 
 from various parts of the foreshore and sent to England, where 
 they were assayed by the author, and were found to contain on 
 an average about 2 dwts. of gold and 6 dwts. of silver per ton. 
 A large proportion of the sulphides was in the state of slimes 
 and difficult to save by concentration, but, on reducing the 
 samples in the ratio of 10 to 1 by panning, the residue was 
 found to contain ^ oz. of gold per ton, while if reduced in the 
 ratio of 28 to 1 the residue of pure sulphides contained 1 oz. 6 
 dwts. of gold and 14 dwts. of silver per ton. These results 
 seem to point to the fact that the extraction of the free gold is 
 not so incomplete as is usually believed, although the total loss 
 is considerable, owing to the value of the sulphides. 
 
 Battery Practice in Dakota. The method adopted in the 
 Black Hills consists in battery and outside plate amalgamation, 
 followed by treatment by a few mercury traps and riffles, and 
 by blanket sluices, no systematic concentration being attempted. 
 The reasons for this are that the ore is free milling and the 
 greater part of the gold is sufficiently coarse to be caught inside 
 the battery, whilst the sulphides from most of the lodes are 
 not rich enough to pay for treatment, and so are allowed to 
 run to waste, although they could be saved at a trifling cost. 
 The sulphides from the Homestake Mine contain gold and silver 
 to the value of only $5 or $6 per ton, whilst the poorest ores 
 which can be treated at a profit, either by the pyritic smelters or 
 the barrel chlorination works in the neighbourhood, yield from 
 $8 to $12 in gold per ton. At the Caledonia Mill, however, 
 the ore treated contains about 4 per cent, of pyrites, which assay 
 about $90 to the ton, and concentration is absolutely necessary for 
 successful working. There are 740 stamps running continuously 
 in the Black Hills, most of which are under the management of 
 the Homestake Company ; 3,000 tons of ore are crushed per day. 
 The yield in bullion is from 3 to 4 dwts. of gold and nearly 1 
 dwt. of silver per ton. "When surface ore was being treated the 
 tailings often contained less than ^ dwt. of gold, and even as 
 late as the year 1889, from 70 to 80 per-cent. of the total contents 
 were extracted, but the increase in the amount of pyrites in the 
 ore, due to the greater depth of the workings, has of late years 
 resulted in an increased loss in the tailings and a diminished 
 yield on the plates, so that the extraction probably does not 
 average more than from 50 to 70 per cent., and the tailings 
 often contain as much as 2 to 2 J dwts. of gold per ton. 
 
 The Homestake Mill is arranged in such a way that the ore 
 need not be touched by hand after it is let fall into the gigantic 
 ore-bins ; consequently eight men can do all the work in con- 
 nection with the battery of 200 head of stamps. The ore 
 is passed over grizzlies, thence through Blake's or Gates 
 crushers, each Blake (of which the dimensions of the month 
 are 9 x 15 inches) being set to supply 20 stamps with ore 
 
216 THE METALLURGY OF GOLD. 
 
 passing through a 1^ inch ring. From the crushers the ore 
 passes through Challenge ore feeders into the batteries, which 
 are arranged in two rows set back to back with the ore 
 feeders between them, the distance between the faces being 44 
 feet. The stamps weigh 850 Ibs., their fall is 8 to 9 inches, and 
 they make 90 drops per minute. The screen used is No. 7, and 
 the output is about 4 to 4f tons per stamp per day. There is 
 one copper plate inside the mortar below the screens, and about 
 1 oz. of mercury per day per stamp head is added with the ore. 
 The shoes and dies are of cast iron and last about two months. 
 It has been shown by experiments in the Father de Smet Mill* 
 that for crushing low grade ores a square-bottomed die is the 
 best, and that there is no economy in using the dies after their 
 surfaces have become perceptibly irregular by wearing. It is 
 accordingly desirable to cast them as thin as practicable and to 
 replace them often at least once a month. 
 
 The amalgamated plates placed outside the mortar are about 
 10 feet long and 4^ feet wide ; they have an inclination of 2 
 inches to the foot, the grade being kept as high as is practicable 
 consistent with good amalgamation, owing to the scarcity and 
 high cost of the water. At the lower end of these plates the 
 ore is passed through a mercury trap or well, and the discharge 
 is then narrowed and the pulp is passed through sluices from 
 60 to 100 feet long and 2 feet wide, lined with copper plates, 
 and broken at intervals by mercury traps. Hungarian riffles 
 are also used, and at the Caledonia Mill, where the sulphides 
 are valuable, an imperfect concentration is effected by means 
 of blankets. The battery plates are dressed every day, and any 
 excess of amalgam then found is removed; the sluice plates 
 are dressed every four days. The battery is cleaned up fort- 
 nightly, and the sluices, mercury wells, riffles, &c., monthly. 
 The cost of milling is 83 cents (3s. 8d.) per ton, the water alone 
 costing from 50 to 57 cents per stamp per day, or 12 to 13 cents 
 per ton of ore. The pulp is run through settling pits and the 
 water used over again during the winter months, when the 
 scarcity of water is felt most severely. It is then warmed before 
 use. The fineness of the bullion is, in general, 820 in gold and 
 170 in silver. 
 
 Stamp Battery Practice in the Transvaal, t On the 
 Witwatersrand the ores consist almost entirely of the now 
 familiar " banket " formation, a conglomerate consisting of 
 water- worn pebbles of translucent quartz set in a matrix 
 composed mainly of oxides of iron and silica. In the ore near 
 the surface the gold is almost entirely free, existing iu the 
 cement, the pebbles being generally barren. At a moderate 
 
 * Trans. Am. Inst. Min. Eny., vol. x., p. 97. 
 
 t This account is mainly based on information kindly supplied by Mr. 
 S. H. Farrar, M.I.C.E., F.G.S., M.I.M.E., of Johannesburg. 
 
STAMP BATTERY PRACTICE. 217 
 
 depth, however, the oxides of iron gradually give place to 
 sulphides, but even then the greater part of the gold can be 
 extracted by amalgamation. The ores vary greatly in richness, 
 some containing less than 1 dwt. of gold per ton, and others as 
 much as 2 or 3 ozs., the average being about 15 dwts. 
 
 The method of treatment consists in crushing through some- 
 what coarse screens with a small depth of discharge, thus 
 obtaining a very large output; the gold is caught partly on 
 plates inside the mortar boxes, partly on copper tables placed 
 outside. The pulp is concentrated in various ways, blanket 
 strakes, Frue vanners, and spitzluten being in common use. 
 The Liihrig vanner is being introduced, and has already met 
 with a considerable amount of success. The concentrates, if 
 they contain little or no sulphides, are ground in pans with 
 mercury ; if pyritic, they are either roasted and chlorinated or 
 treated with cyanide. The tailings from the batteries or the 
 concentrators are treated by the cyanide process. 
 
 There are two types of stamp mills in existence on the Rand, 
 both of which have been very successful, although differing 
 somewhat in their construction. These are the American mills, 
 constructed chiefly by Messrs. Fraser and Chalmers, and the 
 English mills sent out by the Sandycroft manufactory, and by 
 other well-known English firms. A brief account of some of the 
 details of these mills is given below. 
 
 Rock-breakers are in use at almost all the mills on the Rand, 
 both the Blake-Marsden and the Gates crushers being in wide 
 use. In general, one Blake-Marsden machine with a feed- 
 opening of 15 by 9 inches is required for every twenty stamps. 
 The rock-breakers are generally contained in separate rock 
 houses, placed at the mine. The advantages are that the rock- 
 breakers can be placed nearer the ground, and are less expensive 
 to erect, while the stuff in the battery ore-bins is more completely 
 mixed. Self-feeders to the batteries are universal, the Challenge, 
 Tulloch, and Sandycroft machines being chiefly used ; they are 
 arranged so as to keep the ore at a depth of 1 to 2 inches on the dies. 
 
 The mortar blocks are usually of 14-inch square timber, from 
 9 to 15 feet long, and are laid on concrete. It has been found 
 that the usual practice of filling the space between the blocks 
 under one mortar, and those under the next, with well-tamped 
 earth is less satisfactory in giving stability than that of filling 
 the space with solid masonry. The latter course has accordingly 
 been adopted at the best mills. The mortar blocks are covered 
 with blanketing or sheets of india-rubber or lead before the 
 mortars are bolted to them. The English mortars are cast about 
 4 feet high and 4 feet 8 inches long ; the width inside at the 
 level of the dies is lOf inches, that of the American mortar being 
 11 inches. Both mortars widen considerably towards the upper 
 part, the walls not being vertical at any point. The thickness 
 
218 
 
 THE METALLURGY OF GOLD. 
 
 of the mortar at the bottom is 6 to 9 inches, and the inside i 
 usually lined with wrought-iron plates to protect it from wear. 
 The American mortars are furnished with amalgamated copper 
 plates inside, both on the feed and discharge sides, but the 
 earlier English mortars on the discharge side only. The inclina- 
 tion of these plates to the horizon is usually from 70 to 80. 
 Steel wire mesh screens are almost invariably used, punched 
 Russia iron plates being rare. There are from 100 to 1,200 
 holes per square inch of the screen, the usual number being 600 
 to 900 (i.e., 25 to 30 holes to the linear inch). The depth of 
 discharge is about 3 to 5 inches, and the scouring action on the 
 front inside plates is consequently so violent that but little 
 amalgam accumulates there, and, in the opinion of many 
 experts, these plates might advantageously be dispensed with. 
 
 Fig. 45a. 
 
 A section of the mortar used at the Croesus mine is shown in> 
 Fig. 45a. a ir \ cast-steel lining plate with slots or recesses in 
 it for collecting the amalgam; b is the feed opening; c the 
 wooden blocks for carrying ^, the copper plate ; d is the screen 
 opening ; and f a steel plate. 
 
 The weight of the stamp in the earlier mills was usually 
 about 850 Ibs. in the American, and from 700 to 750 Ibs. in 
 the English mills. In the later mills, which represent in every 
 respect the very best modern practice, the stamps are of 1,000 
 Ibs. weight or upwards, while in the Modderfontein 40-stamp- 
 battery the weight is 1,250 Ibs., the heaviest gravitation stamp- 
 yet in action. The actual weights of the parts of two English 
 stamps in use were found to be as follows : 
 
STAMP BATTERY PRACTICE. 219" 
 
 Nominal weight of Stamp, . . 700 Ibs. 750 Ibs. 
 
 Actual ,, Stem, . . 238 ,, 287 ,, 
 
 Head, . . 197 197 
 
 Tappet, . . 115 115 
 
 Shoe, . . 180 180 
 
 Total actual weight of Stamp, . 730 779 ,, 
 
 The usual height of drop varies from 6 to 9 inches, and the- 
 speed is usually from 90 to 95 drops per minute. The shoes 
 and dies are 9 inches in diameter, and the shank of the shoe in 
 English mills is 5 inches long and 3 inches wide. The shoe is 
 often kept in work until the butt is worn to about 1 inch thick. 
 The dies are now sometimes made in two pieces, a new boss 
 being fitted into the footplate as soon as the old one has been 
 worn out. Steel is almost exclusively used for shoes and dies,, 
 manganese steel being preferred. The order of drop is usually 
 1.3.5.2.4 in the English mills, but by some authorities the orders- 
 1.4.2.5.3 and 1.5.3.2.4 are considered better. 
 
 The cam-shaft is from 4J to 6 inches in diameter, according 
 to the weight of the stamps. A separate cam-shaft is used for 
 each battery of five stamps. The cam-pulleys in the American 
 mills resemble that described on p. 113; in the earlier English 
 mills they were often of wrought iron, 6 or 7 feet in diameter,, 
 with round arms. Many of these have proved defective, be- 
 coming crystalline and breaking after running for a short time 
 only. Wooden pulleys are now generally used. 
 
 The water for ten stamps is supplied through a 3-inch gas 
 pipe, from which two 1^-inch pipes branch to each mortar. The 
 pipes are sometimes suspended from the roof instead of being 
 attached to the battery posts ; the objection to the latter course 
 is that the pipes work loose owing to vibration. The amount 
 of water supplied varies in the De Kaap gold field from 100 to 
 200 gallons per stamp per hour, or from 1,600 to 3,200 gallons 
 per ton of ore crushed. On the Witwatersrand the amount of 
 water per ton of ore varies from 1,600 to 3,500 gallons, and, 
 according to Hatch and Chalmers, averages about 2,000 gallons 
 per ton crushed. Mercury is added to the mortal boxes in small 
 quantities every few minutes, the total addition amounting to 
 a spoonful in two hours. 
 
 The arrangement of the amalgamated copper plates outside 
 the mortars varies somewhat. In some mills a single plate of 
 best Lake Superior copper is used, 4J to 5 ieet wide and 12 feet 
 long, with a fall of 1 in 12 and a mercury well at the lower 
 end. In others, the lip of the mortar box is covered with a 
 sloping copper plate from 12 to 14 inches wide, running the 
 whole length of the mortar. The pulp flows over this " lip- 
 plate " on to a second copper strip, the " splash-plate," 14 inches 
 wide, and as long as the lip-plate, but inclined in the reverse 
 
220 THE METALLURGY OP GOLD. 
 
 direction so as to throw the pulp back towards the mortar. It 
 then falls into an iron trough or distributor of semi-circular 
 cross-section, perforated with small holes for the purpose of 
 distributing the pulp over the plates. The latter are about 4i 
 feet wide and | inch thick ; the pulp flows over 4 or 5 plates in 
 succession, each 2^ feet long. They consist of copper, which is 
 .seldom coated with electro-deposited silver. There is a shallow 
 riffle at the lower end of the first plate, and a drop of 1 inch 
 between each pair of the rest ; sometimes the plates simply 
 overlap, so that the drops are of only -J inch. At the lower end 
 of the series of plates is placed a mercury trough. The tables 
 for the support of the plates are made of heavy timber so as to 
 be steady ; these timbers are not fastened to any of the battery 
 posts or sills, as the vibration caused by the stamps would inter- 
 fere with amalgamation. The grade of the plates varies from 
 1 to 2 inches per foot, and can be altered by means of wooden 
 wedges or of adjusting screws. 
 
 The plates at the Sheba mill are treated as follows, according 
 to an account furnished by Mr. H. Nichols, A.R.S.M. : The 
 outside plates are usually dressed every hour if the rock is of 
 ordinary richness (i.e., containing from 10 to 15 dwts. of gold 
 per ton). In dressing, the amalgamator brushes all the quick- 
 silver and thin amalgam up to the top of the plate and distributes 
 it there. If the rock crushed has been rich (containing say 
 1 oz. per ton), quicksilver must be sprinkled on the plates and 
 worked into the hard amalgam with a brush, but if it has been 
 only 10-dwt. ore, the plates only require fresh mercury once in 
 two or three hours. The outside plates may require scraping 
 once every day or every other day if the ore is rich ; the inside 
 plates are scraped only on the occasion of the monthly clean-up. 
 The yield from each of the two sets of plates is about the same. 
 A retort of 2,500 ozs. of squeezed amalgam yields about 750 ozs. 
 of gold. Cyanide of potassium is added in the mortar boxes in 
 small quantities when the plates are discoloured by salts of 
 copper ; and soda is used to remove grease. When a screen is 
 badly broken, the tive stamps connected with it are at once hung 
 up, as much loss may be incurred by allowing coarsely-ground 
 xock to pass over the plates. 
 
 On the Rand, the amount of gold caught on the plates is 
 now only from 45 to 65 per cent, of the gold in the ore, and 
 the average is about 55 per cent. A higher percentage can be 
 easily caught, but only if the output is decreased, and this 
 sacrifice is less worth making, for the reason that most of the 
 gold which is allowed to escape from the tables is extracted 
 subsequently by cyanide. The percentage extracted from pyritic 
 ore is but little less than that from the oxidised banket, the 
 more refractory minerals being in general absent from the ore. 
 'The amount of gold caught inside the battery is from 10 to 
 
, 
 
 6 r-, 
 
 1 
 
STAMP BATTERY PRACTICE. 22 1 
 
 40 per cent, of the total saved by amalgamation if inside plates- 
 are used, and 4 or 5 per cent, if there are none, and the greater 
 part of the remainder is retained on the first two feet of tables. 
 This part is scraped every day, but the lower parts are left 
 undisturbed for longer periods, varying up to a week. At the 
 Robinson mine the plates are dressed every four hours, but 
 amalgam is removed only every other morning (Hatch and 
 Chalmers). 
 
 The method of concentration at different mills varies; the 
 following details refer to the Simmer & Jack mill of 100 stamps, 
 when the ore was still free-milling, and a rough method of con- 
 centration sufficed. Here blanket tables, 16 feet long and of the 
 same width as the copper plates, are placed immediately below 
 the latter. The blanket tables are divided longitudinally into 
 three parts by wooden "fillets." The blankets are washed 
 every two hours and the concentrates collected in wooden boxes 
 running on wheels. One division of each table is washed while 
 the pulp is deflected so as to run over the other two divisions. 
 The blanket-sands amount to about 6 per cent, of the tailings ; 
 they are treated by grinding with mercury in four pans, which 
 together treat 24 tons per day of twenty-four hours. The 
 charge for each pan is 2,600 Ibs., five charges being treated in 
 twenty-four hours. The pans are of cast iron, 5 feet in diameter, 
 and have double bottoms into which steam can be passed for 
 heating purposes ; the mullers make from 50 to 60 revolutions 
 per minute. The discharge takes place through a pipe at the 
 bottom of the side of the pan, and the pulp is run into settlers, 
 of which there are two, each 7 feet in diameter. The pulp is 
 here diluted with water and stirred by four arms revolving at 
 the rate of fifteen turns per minute. There are a number of 
 holes at different levels in the side of the settler and the diluted 
 pulp is run off through these as soon as the mercury has settled 
 sufficiently. 
 
 Where the ore is pyritic, this method does not suffice, and 
 vanners or hydraulic classifiers (Spitzlutvn and Spitzkasten) are 
 employed. It is usual to set three Frue vanners to treat the 
 pulp coming from five stamps. There has been much discussion 
 on the respective merits of these two classes of machines in 
 treating battery pulp on the Rand. Where vanners are used, 
 the concentrates have usually been roasted and chlorinated, and 
 the tailings treated by cyanide. The growing tendency to treat 
 even concentrates by cyanide, however, throws doubt on the 
 advantage of using vanners at all, for if both classes of materials 
 into which the pulp is separated are ultimately treated by the 
 same process, there is no prima facie reason for effecting the 
 separation at all, and, on the other hand, spitzluten afford a 
 product in an excellent condition for cyaniding. The pulp 
 coming from the plates has about 40 per cent, oi its gold locked 
 
THE METALLURGY OP GOLD. 
 
 up in the pyrites, and 60 per cent, in the form of finely-divided 
 free gold which cannot be removed by the vanning machines, 
 but is readily extracted by cyanide. It follows that concentra- 
 tion and chlorination are not sufficient of themselves, although 
 the cyanide process may be, and for this reason a Committee of 
 the Johannesburg Chamber of Mines reported in May, 1895, 
 that close concentration is usually quite unnecessary. Doubtless 
 the best percentage extraction would be obtained in almost every 
 case by classifying the pulp, concentrating the various classes 
 separately, chlorinating the concentrates, and cyaniding the 
 tailings. It would appear, however, that, under present con- 
 ditions, the greatest profit is gained by rough classification in 
 spitzluten, followed merely by cyanide treatment. 
 
 The treatment of amalgam presents no unusual features. At 
 the Simmer & Jack mill there is a cast-iron clean-up pan, 3J feet 
 in diameter, with mullers making 60 revolutions per minute. 
 The hard amalgam, skimmings, and headings from the battery 
 are ground in this pan, with fresh mercury, for three or four 
 hours ; after this the mercury is strained through chamois 
 leather, and hard balls of amalgam, which contain from 33 to 
 .35 per cent, of gold, are thus obtained ready for retorting. 
 
 The life of the various parts of the mills is approximately as 
 follows : The shoes last for two to four months, the dies from 
 three to four months, and the screens about fourteen days ; when 
 banket is being crushed the wear is less than if quartz is in 
 course of treatment. Tappets have been in use for three years 
 without showing any signs of wear. The stems of the stamps 
 become crystalline and the ends usually break off after about 
 twelve or eighteen months' use, but this can be avoided by 
 annealing at regular intervals. The mortar boxes last at least 
 four years, the cam-shafts about three years, and the guides 
 twelve months. 
 
 The amount of ore crushed is usually over 4 tons of banket 
 per stamp per diem. In 1891 the average output was from 2J 
 to 3J tons per day ; but is now greater (viz., about 4J tons) 
 owing to the greater average weight of the stamps. At the 
 May Consolidated Mine the 1,150-lb. stamps crush 5|- tons per 
 day, and Hatch and Chalmers consider that, in a few years time, 
 from 5 to 6 tons will be a common result. The large output is 
 due partly to the nature of the rock, banket being softer than 
 ordinary quartz, and partly to the small depth of discharge and 
 the coarseness of the screens. 
 
 In present practice there is really no difference between 
 English and American mills as regards design. It is generally 
 acknowledged that stamps of less weight than 950 Ibs. should 
 not be used, and the tendency is to increase this weight. The 
 fittings, such as shoes and dies, are almost universally made of 
 JEnglish steel. It may safely be said that English mills made 
 
STAMP BATTERY PRACTICE. 
 
 223 
 
 by the leading makers compare favourably with the best 
 American mills. 
 
 Taking an average of the whole of the operations on the 
 Rand, the yield of gold on the amalgamated plates is gradually 
 diminishing, having been 13'5 dwts. per ton in 1890, 11 '3 dwts. 
 in 1891, 9-77 dwts. in 1892, 9-59 dwts. in 1893, 9-52 dwts. in 
 1894, and less than 8 dwts. at the end of 1895. In the last- 
 named year the thirty-seven leading companies on the Rand 
 crushed 2,485,311 tons of ore, and obtained 1,182,794 ozs. of 
 gold on the plates (or 64 per cent, of the total gold contents), 
 37,138 ozs. by concentration and 486,082 ozs. from the tailings 
 by cyanide treatment. 
 
 The cost at the best mills is from 3s. to 5s. per ton for 
 crushing and amalgamating, and about 4s. to 6s. per ton for 
 treatment of tailings by the cyanide process. Concentration 
 and chlorination probably cost about 3s. to 3s. 6d. per ton of 
 ore. The following details of the work done in two mills in 
 twelve months in 1894-95 are given as examples of what is 
 possible with good management : 
 
 
 Tons of 
 Ore. 
 
 GOLD RECOVERED. 
 
 Cost per 
 Ton 
 
 Total. 
 
 Per Ton. 
 
 ROBINSON COY.'S MILL. 
 Treated in Stamp Battery, 
 > by cyanide, . 
 
 107,935 
 70,526 
 
 110, 962 ozs. 
 20,213 
 
 1 oz. 04 dwt. 
 5'7 dwts. 
 
 3s. lid. 
 3s. 10d.* 
 
 CROWN MILL. 
 Treated in Stamp Battery, 
 by cyanide, . 
 
 200,785 
 138,800 
 
 73,339 
 42,749 
 
 7'3 
 6-16 
 
 3s. Od. 
 2s. ll'47d. 
 
 At the Robinson Mill, 4-60 tons of ore were crushed per 
 stamp per day, and at the Crown Mill 5 tons. 
 
 * The cost for June, 1896, was 2s. 7'25d. 
 
224 THE METALLURGY OF GOLD. 
 
 CHAPTER XL 
 
 CHLORINATION : THE PREPARATION OF ORE 
 FOR TREATMENT. 
 
 The Plattner Process. The value of chlorine, as an agent for 
 the extraction of gold from certain ores and from almost all 
 concentrates, has now been recognised for many years, and its 
 use is gradually extending, although it is doubtful whether its- 
 application will ever be as general as appeared probable before 
 the introduction of the cyanide process. Its use was first 
 suggested by Dr. John Percy, F.R.S., at the Swansea meeting of 
 the British Association, held in August, 1848, in a paper* em- 
 bodying the results of experiments carried out in the year 1846. 
 Simultaneously, in 1848, Prof. C. F. Plattner, Assay Master at the 
 Royal Freiberg Smelting Works, applied chlorine gas to the assay 
 of the Reichenstein residues, and proposed that a similar method 
 of treatment should be adopted on a large scale. These 
 residues were the result of treating the Reichenstein ore with 
 the object of extracting the arsenic. They consisted chiefly of 
 oxides of iron and oxidised arsenical compounds, and had been 
 roasted in the course of the process for the extraction ot the 
 arsenic. The residues had been accumulating for more than a 
 century, and contained from 15 dwts. to 1 oz. of gold per ton;, 
 they were considered too poor to smelt, while they could not be 
 made to yield the gold contained in them by amalgamation. 
 
 Prof. Plattner's suggestion was followed up by investigations 
 made by Dr. Duflos in 1848,f and by Lange in 18494 Dr. Duflos 
 compared the results obtained by treating the residues with 
 chlorine water by percolation in a stationary vat, and by agita- 
 tion in a revolving barrel; and as these results were the same, 
 he recommended the stationary vat as being more economical. 
 He also obtained identical results with chlorine-water and with 
 dilute solutions of chloride of lime and hydrochloric acid mixed 
 together. On the other hand, Lange found that gaseous chlorine,, 
 applied to the ore in the same manner as had been used by 
 Plattner in assaying, was a more efficient agent than a solution 
 of chlorine in water, and it seems to have been in accordance 
 with his advice that the first chlorination works, that at 
 Reichenstein, was established in 1849. The chlorinating vessels- 
 
 * Phil. Mag., 1853, vol. xxxvi., pp. 1-8. 
 
 t Erdm. Journ. Prak. Chem., vol. xlviii., 1849, pp. 65-70. 
 
 J Karsten's Archiv., vol. xxiv., 1851, pp. 396-429. 
 
UNIVERSITY 
 
 CHLORINATION. 225 
 
 were small earthenware pots, and the precipitant employed was 
 sulphuretted hydrogen. Plattner subsequently recommended 
 wooden vats coated with pitch, and ferrous sulphate as a pre- 
 cipitant, and although these were not at first used at Reichen- 
 stein, they were adopted by Mr. G. F. Deetken in 1 857, when he 
 introduced the system into California. The system ascribed to 
 Plattner consists of the following operations : 
 
 1. The concentrates or residues are subjected to a " dead " roast 
 in a reverberatory furnace. 
 
 2. The roasted ore is slightly damped with water and charged 
 into wooden vats, holding from 1 ton upwards. The vats have 
 false bottoms consisting of filter beds of gravel or of cloth. 
 Chlorine gas, generated in another vessel, is introduced at the 
 bottom of the vat, and rises through the ore, permeating every 
 part of it. The vat is then closed up and left undisturbed for 
 twenty-four hours or more, by which time all the gold is con- 
 verted into soluble chloride of gold. 
 
 3. The soluble salts are then washed out by water, which is 
 allowed to flow on to the surface of the ore, and, passing through 
 it, drains through the filter bed. When all the gold has thus been 
 removed in solution, the tailings are thrown away. 
 
 4. The solution of gold chloride is acted on by ferrous 
 sulphate, or some other suitable reagent, by which the gold is 
 precipitated ; the particles of the precious metal are allowed to 
 settle, and then are collected and melted down. 
 
 Plattner Process at Reichenstein. The following descrip- 
 tion of the original process employed at Reichenstein is an 
 abstract of the account given in Kerl's ffuttenkunde, vol. iv., 
 p. 372, 1865. The material treated consisted of the residues 
 obtained by roasting arsenical iron pyrites for the production of 
 white arsenic, which was volatilised and condensed in brick 
 chambers. There were forty-eight earthen chlorination pots, 
 each holding 150 Ibs. of ore. These pots were strengthened with 
 iron hoops, and suspended on two journals, so that they could 
 be discharged by inverting them. 
 
 The lower part of the pots was of a conical shape, and this 
 part was filled with pebbles and sand covered by a perforated 
 earthen plate, the function of which was to prevent the ore from 
 mixing with the filter bed. The ore filled the cylindrical part of 
 the pot above the earthen plate. The chlorine was generated by 
 the action of hydrochloric and sulphuric acids on manganese 
 dioxide in earthenware vessels, and was conveyed thence to the 
 ore-pots through leaden pipes. The gas was introduced below 
 the filter bed, and passed upwards through the ore for an hour ; 
 a wooden cover was then fitted on, but not luted down until 
 chlorine had been passed for from six to seven hours longer, 
 after which all joints were luted down with dough, and the vat 
 left until the next day. The cover was then removed and water, 
 
 15 
 
226 THE METALLURGY OF GOLD. 
 
 at a temperature of from 64 to 77 F., poured on, and allowed 
 to percolate through the ore and filter bed by gravity. The 
 liquid coming from twenty-four pots was conveyed to four vats, 
 the first one being filled with solution before the second was 
 used, and so on ; the contents of the fourth vat, being too poor 
 for precipitation, was used over again tor leaching. The leaching 
 was stopped when 90 cubic feet of water had passed through the 
 total charge of 3,600 Ibs., this being at the rate of 312 gallons 
 per ton of 2,000 Ibs. The liquid from the first three vats was 
 drawn off into twenty glass globes, which were heated on a sand 
 bath so as to raise the temperature to 77 F. Sulphuretted 
 hydrogen, obtained from fused sulphide of lead and sulphuric 
 acid, was passed through until the saturation point was reached, 
 when the liquid was left to settle until the next day ; after this 
 the clear supernatant liquid was passed through sawdust filters 
 to catch all the sulphide of gold, which might still have been in 
 suspension. The sulphides were refined by dissolving them in 
 acids, precipitating the metallic gold by ferrous sulphate, and 
 melting it in clay crucibles with nitre and borax. The 
 amount of arsenides treated daily was 3,600 Ibs., containing 
 about |- oz. of gold per ton, so that only about 250 ozs. of gold 
 were extracted yearly, and it is difficult to believe that the 
 enterprise could have been a commercial success. 
 
 MODERN PRACTICE IN CHLORINATION. 
 
 The original vat process as described above has been subjected 
 to great changes, many of which originated in the United States. 
 The most notable alterations are an increase in the size of the 
 vats and the substitution of wood for earthenware in their con- 
 struction, and the use of ferrous sulphate instead of sulphuretted 
 hydrogen to precipitate the filtered liquors. This process, as 
 altered, is still in very extended use for dealing with roasted 
 concentrates, but in a few special cases, where large quantities 
 of material are available for treatment, the barrel process has 
 been adopted. The barrel process of chlorination, which was 
 first used on the large scale in the United States, differs from 
 the vat or Plattner process chiefly in the fact that the chlorination 
 is effected in revolving barrels, instead of in stationary vats, 
 while the operations of chlorination and lixiviation are usually, 
 though not invariably, performed in different vessels. The 
 two processes will be described under the following headings : 
 
 1. Crushing. 
 
 2. Roasting. 
 
 3. Chlorination and Lixiviation. 
 
 4. Precipitation of the gold from its solution and production of bullion. 
 
 Vat and barrel chlorination differ essentially only in the 
 methods described under the third heading, although the differ- 
 
CRUSHING. 227 
 
 ences in the conditions under which they are applied involve 
 variations in the treatment under the other headings. The 
 sections devoted to crushing and roasting include descriptions 
 of the methods used to prepare ore for chlorination both in the 
 vat and the barrel. The remaining operations are described 
 separately for each of the two processes, and a general account 
 of the precipitation of gold is added. 
 
 CRUSHING. 
 
 In those cases in which gold ores are treated by crushing and 
 amalgamation, and the whole of the tailings, or only the con- 
 centrates obtained from them, are subsequently chlorinated, 
 the method of crushing will be determined by the considerations 
 .already discussed in the chapters on amalgamation and concen- 
 tration, and will depend partly on the state of aggregation of 
 the free gold. If, on the other hand, ores are to be treated in 
 the first instance by chlorination, special regard may be paid to 
 the method of crushing as affecting the suitability of the crushed 
 product for leaching. There are two somewhat opposite con- 
 ditions to be fulfilled, viz.: 
 
 1. The crushed product should be fine enough to admit of the 
 Tvhole of the gold being laid open to the attack of the chlorine. 
 
 2. It must be coarse enough to allow all the soluble chloride 
 of gold thus formed to be washed out of the ore rapidly and 
 easily. 
 
 It is quite impossible to fulfil completely both of these con- 
 ditions on a large scale, although in the laboratory some ores, 
 when treated with the greatest care and patience, may be made 
 to yield the whole of their gold. In practice it is better to aim 
 at an extraction of from 8j to 95 per cent. (98 to 99 per cent. 
 Taeing attained in some rare cases), and to do this at as low a 
 cost as possible. With this object in view, the ore is kept as 
 coarse as possible, and is usually passed through screens with 
 only from 8 to 30 holes to the linear inch. The chief point to 
 be attended to is the attainment of uniformity in the product, 
 as any considerable proportion of slimes enormously increases 
 the difficulties of leaching. Moreover, since the crushed ore 
 must almost invariably be roasted, it is a great advantage to 
 adopt some method of dry crushing. Proposals have been made 
 to subject the ore to wet-crushing by stamps or other machines, 
 and to collect the pulp in settling-pits, and dry it, either in 
 the roasting furnace or in a separate furnace. It appears that 
 this method has never been tried, and there are several practical 
 objections to it. 
 
 Dry crushing by means ot stamps has been found to answer 
 well at Park City, Utah, and at other places in preparing silver 
 
228 THE METALLURGY OF GOLD. 
 
 ores for treatment by the hyposulphite leaching process, but 
 many ores would form too much impalpable powder for this- 
 system to be applicable. In order to obtain a perfectly uniform 
 product in any machine, it would be necessary for every particle 
 that had been reduced to the required degree of fineness to be 
 instantly separated from those which were still too large. If 
 the screening is not instantaneous and thorough, the small 
 particles are ground still finer while the larger pieces are in 
 process of reduction to the required fineness. Of all the 
 machines for dry crushing which have been designed with a 
 view to effect this object, rolls have come into the most general 
 use, and give the most satisfaction. They are used, among 
 other places, at Mount Morgan, the largest chlorination mill in 
 the world, and at the Dakota Mills. Rock-breakers are now 
 always employed to reduce the ore to about H-inch cubes, or 
 smaller, before feeding it to the rolls. 
 
 The following description of Krom's rolls, the best known 
 modern make of high-speed fine-crushing rolls, is chiefly derived 
 from Mr. A. H. Curtis' paper on "Gold Quartz Reduction."* 
 Crushing rolls made by several other manufacturers in various 
 countries might be described almost in the same words, and 
 give nearly if not quite equally good results in practice, while 
 they are much lower in price. 
 
 The rolls are placed with their faces a short distance apart,, 
 this distance varying with the degree of fineness to which the 
 ore must be reduced. The main driving power is applied to the 
 shaft of only one roll, by means of belt pulleys, the other roll 
 being driven only with sufficient force to ensure that the rollers 
 will always take hold of the ore, and also to keep them in motion 
 when no ore is passing between them. If all the power is 
 applied to turn one roll only, the feeding must be continuous, or 
 the free-running roll would come to a standstill. On recom- 
 mencing the feeding, a blow would then be given to the stationary 
 roll, in the attempt made by the other roll to set it instan- 
 taneously in motion at full speed, and this jar would have & 
 disastrous effect on the machinery. 
 
 In the older forms, tooth-gearing was used instead of belt 
 pulleys for the application of the power, and the two rolls were 
 compelled to revolve at an equal rate of speed by gear-wheels, 
 connecting together, placed on the axles. The advantages of 
 belted rolls are, that a higher speed can be easily attained, and 
 also that if the rolls were to become jammed from any cause, the- 
 belts would slip or be thrown off, while the tooth-gearing would 
 be broken. Geared rolls, however, are still largely used for 
 coarse crushing. The rolls have crushing tires made of forged 
 steel, and are firmly and simply secured to the shafts by means 
 of cone-shaped heads. Chilled-iron rolls are still often used,. 
 * Proc. Inst. Civil Eng., session 1891-92, vol. cviii., part ii. 
 
CRUSHING. 229 
 
 being much cheaper ; but the wearing of the faces is more rapid 
 and less uniform than in the case of steel rolls. Emery wheels 
 for levelling the unevenly worn faces of chilled-iron rolls are 
 recommended by some makers. These crushing tires can be 
 taken off and replaced when they are worn out. In the older 
 forms of Krom's rolls the crushing strain is taken up by powerful 
 springs, which press the rolls towards one another ; when par- 
 ticularly hard fragments are passing through the rolls, they are 
 forced apart against the action of the springs. These springs are 
 now dispensed with. It is desirable, in order to keep the wear 
 of the faces even, that the rolls should always be kept parallel, 
 and special appliances, such as Krom's swinging pillow blocks, 
 Jiave been introduced by some makers to ensure this. The hopper 
 is specially designed to spread the ore evenly across the crushing 
 face, and the rolls, screens, elevators, &c., are all securely boxed 
 in with a wooden housing. This last precaution is necessary in 
 order to prevent loss by floating dust, which otherwise may be 
 large, the richest part of the ore thus passing off, and not only 
 making the atmosphere of the mill insupportable, but having a 
 disastrous effect on the bearings of the machinery. Rolls are 
 usually from 12 to 16 inches across the face, and from 22 to 36 
 inches in diameter. 
 
 Corrugated rolls are used to some extent in Australia, but 
 though the crushing surface is increased in this way, they are 
 considered in the United States to be of little value, wearing 
 unevenly and soon getting out of order. 
 
 The method of crushing usually adopted in mills where rolls 
 are employed may be described in general terms as follows. It 
 may be supposed that the ore is to be passed through a 20-mesh 
 screen : 
 
 The ore is put through a rock- breaker, and passed at once to 
 classifying screens, by which it is divided into four classes of 
 material. One of these consists of ore that will pass a 20-mesh 
 screen, and this amount (usually small) is separated as finished 
 product ; another small portion is returned to the rock-breaker, 
 being larger than, say, ^-inch cubes, and too coarse for the rolls, 
 and the greater part is passed to one or other of two pairs of 
 rolls. One of these pairs of rolls takes the J-inch material and 
 passes it through, say, a 6-mesh screen. The other receives the 
 product and passes it through a 20-mesh screen. After each 
 passage through the coarse rolls, a certain amount of finished 
 product is obtained, and the remainder is classified, part being 
 returned to the same rolls, and part being sent on to ,the 
 fine rolls, the product from which is also classified. Revolving 
 screens are in general use for classification. The use of two 
 rock-breakers, one for coarse crushing, and the other to take its 
 product and reduce it to the size of broad beans, is often recom- 
 mended. The number of pairs of rolls to be used in succession 
 
230 THE METALLURGY OF GOLD. 
 
 depends on circumstances : two is the usual number, a third pair 
 being occasionally added. One rock-breaker and two pairs of 
 rolls will suffice for the reduction of most ores to a size sufficiently 
 fine to pass through a 30-mesh screen, but the crushing can be 
 done more economically in the long run by still more gradual 
 reduction. 
 
 Mr. John E. Rothwell, late manager of the Golden Reward 
 Chlorination Mill, Dakota, gives it as his opinion* that "two seta 
 of rolls are sufficient, but three will do better. The ore should 
 come to the coarse rolls not coarser than f-inch mesh, and these 
 rolls should be set about f inch apart. The middle rolls are set 
 about ^ inch or less apart, and the fine rolls about as far apart 
 as the size to which the ore has to be crushed. If only two sets 
 are used, the coarse are set a little closer than with three, and 
 the fine remain the same. 
 
 " The springs should be set up so tight that they will not give 
 to the hardest pieces of ore, but will allow a piece of steel or 
 iron to pass through without throwing the belts. The periphery 
 speed of the rolls should be about the same as, or a little faster 
 than, the falling speed of the ore, and the ore should be fed in 
 an even sheet across the surface of the roll ; little trouble will 
 then be experienced in keeping the surfaces true and in pro- 
 ducing a granular pulp, carrying but a small percentage of dust. 
 If rolls were made of larger diameter and narrower, the result 
 would be a still more gradual reduction, and possibly a greater 
 capacity. I have used those of 39 J inches (1 meter) in diameter 
 and 12 and 15 inches face." 
 
 The rolls need to be frequently examined. When any uneven- 
 ness of the faces can be detected, Rothwell suggests t putting a 
 piece of sheet iron in the hopper to deflect the ore away from the 
 more worn portion to the less. 
 
 "At the Bertrand Mining Coy. 's Lixiviation Mill, Nevada, 
 where Krom's rolls are used for pulverising silver ores with a 
 quartz gangue, 15,000 tons of ore were crushed before it was 
 found necessary to put new tires on the finishing rolls ; while 
 after a further crushing of 5,000 tons of ore, the tires of the 
 roughing rolls were still expected to be good for two or three 
 months' work. In neither case were the tires found to be at all 
 grooved, the reason for their renewal being that they had 
 become worn too thin for further work. The rolls in the 
 Bertrand Mill are stated to crush 50 tons of hard quartzose 
 ore in twelve hours to pass through a 16-inesh screen ; while in 
 the Mount Cory Mill, Nevada, 50 tons are reduced to 30-mesh 
 in the same time." J In these mills silver ores are crushed for 
 treatment by roasting and lixiviation with a hyposulphite solu- 
 tion. It is possible to crush quartz to pass through screens of 
 from 30- to 70-mesh by rolls, the degree ot fineness to which the 
 
 * Eng. and Mng. Journ., Feb. 7, 1891. 
 
 t Mineral Industry for 1897, p. 263. 
 
 Curtis on "Gold Quartz Reduction." Proc. Inst. Civil Eng., 1892. 
 
CRUSHING. 231 
 
 rock can be pulverised depending on the number of times the 
 product, after screening, is returned to the hopper. The very 
 fine dust, which is inevitably produced in greater or less quantity, 
 according to the character of the rock, may be separated by 
 exhaustion with a fan, driven at carefully regulated speed, and 
 in this way the sliming is reduced and the resulting loss of gold 
 in the subsequent operations minimised. The average speed of 
 fine-crushing rolls is from 80 to 100 revolutions per minute, the 
 two rolls being driven at equal rates of speed. 
 
 Crushing by rolls at Rapid City, Dakota. The following 
 details concerning the crushing plant at the Rapid City Chlor- 
 ination Works, S. Dakota, which was fitted up in 1891, have been 
 supplied by Mr. D. Dennes, formerly the foreman of the mill. 
 " Gates crushers were used for some months, after which they 
 were replaced by a large Blake crusher, which is found to work 
 more economically and with less cost for repairs. The crushed 
 material from the Blake is passed through a rotating cylindrical 
 drying furnace, and then screened, part being passed to the ore- 
 bin set aside for finished product, a little going back to the rock- 
 breaker, and the remainder going to the first pair of rolls. These 
 have steel tires and are set at T \ inch apart ; they are driven 
 at a speed of 90 revolutions per minute by means of toothed 
 gearing, both rolls being driven independently. The objections 
 to the use of belts for the coarse rolls are stated to be that a 
 specially hard piece of ore may throw the belt off and so 
 interrupt the work. The peripheral speed of the rolls is 14 feet 
 per second. The product of the coarse rolls is classified by 
 screening, and the bulk of it sent at once to the fine rolls, which 
 are steel-faced belted rolls driven at 155 revolutions per minute. 
 The two faces just touch, but each roll is driven by a separate 
 belt, and one is revolved at the rate of two revolutions per 
 minute faster than the other. This arrangement has a remark- 
 ably beneficial effect in keeping the wearing surfaces even. The 
 tires are replaced alternately, so that there is always one new 
 and one old tire working together. Almost all the product of 
 these rolls passes at once through a 16-mesh wire screen, which 
 is the finest screen used. 
 
 " The revolving screens of Messrs. Fraser & Chalmers were 
 used for six months, and then discarded. Flat rectangular 
 screens are now used, measuring 6 feet by 12 feet, set in a 
 wooden frame which is subjected to a reciprocating motion, 
 striking against a rubber pad placed close to one of its long 
 sides. The screen is slightly inclined, so that the ore, which is 
 fed evenly across the screen, travels down it. This apparently 
 retrograde step in screening was justified by a considerable 
 improvement in efficiency and a reduction of expense. The 
 revolving screens were found to wear out very rapidly, and the 
 repairs were costly. The weight of ore causes the metal filter- 
 
232 THE METALLURGY OP GOLD. 
 
 cloth to sag down between the bars constituting the framework ; 
 when this part of the cylinder reaches the top of its course, the 
 cloth sags in the opposite direction from its own weight. The 
 result of this double motion is that the sieve breaks along a 
 line close by and parallel to the framework bars. When this 
 breakage occurs, repairs are tedious and expensive, and the mill 
 must be stopped while they are being done, while flat screens 
 can be lifted out, and a new one put in, by two men in a lew 
 minutes." 
 
 From 100 to 120 tons of ore are stated to have been crushed 
 in this mill per day, but as the mill was in continuous operation 
 for a short time only, and is now shut down, it is not possible to 
 draw definite conclusions from the results obtained. The speed 
 of the rolls seems to have been excessive, and the difficulties 
 with the revolving screens may be partly met by stretching the 
 wire cloth very tightly. It is more common now, however, to 
 use immovable screen plates set at an angle of 45 (Berthelot 
 system) no shaking being required.* 
 
 Comparison between Rolls and Stamps. As the subse- 
 quent treatment of an ore determines its method of crushing, no 
 accurate general comparison of stamps and rolls can be made. 
 A comparison is only possible in the special cases where both 
 methods of crushing are applicable. Wet crushing by rolls need 
 not be considered, as it is not practised ; even where the advo- 
 cates of rolls wish to replace stamps by rolls for wet crushing and 
 amalgamation, they propose that the ore should be crushed dry 
 and then wetted down. Dry crushing by stamps is usually 
 about one-third more expensive than wet crushing, as the 
 capacity falls off to that extent, in spite of double discharge, 
 assisted occasionally by currents of air given by blowers, which 
 are designed to carry the crushed ore against the screens. The 
 amount of slimes made is also large. Advantages in the use of 
 rolls for dry crushing have been stated to be " the fewness of the 
 wearing parts, and consequent small cost of repairs ; the great 
 efficiency of the process, in that the ore escapes from the rolls 
 immediately it is crushed, so that over-crushing is unlikely to 
 occur, and the great capacity of rolls, effective work being 
 constantly done, and the amount of crushing surface brought into 
 contact with the ore per minute being very large. The prime 
 cost of the rolls is considerably less than that of stamps of the 
 same capacity." t 
 
 The capacity of rolls has perhaps been frequently over- 
 estimated owing to the assumption having been made, that the 
 product of equal crushing surfaces must be the same. Thus, 
 Mr. Curtis observes in the paper already cited : " As an 
 index of the capacity of Krorn's rolls, it may be stated 
 that two sets of 26-inch (diameter) rolls, with faces 15 inches 
 
 * Roth well in Mineral Industry for 1897, p. 266. t Curtis, Ice. cit. 
 
CRUSHING. 233 
 
 long, give rather more effective crushing surface than fifty 
 gravitation stamps, each 8 inches in diameter, falling at the rate 
 of ninety drops per minute. In making this calculation, the 
 average diameter of the rolls is taken as only 24 inches, so as to 
 allow for their gradual wearing, while their speed is taken at 
 100 revolutions per minute." 
 
 Although the calculated crushing surface is as stated, it does 
 not follow that the capacity of the two sets of machinery com- 
 pared with one another is the same, since it will depend on the 
 pressure as well as the crushing surface. No doubt the pressure 
 .at the moment of impact of a 900-lb. stamp is enormously greater 
 than that exercised by the faces of the rolls. 
 
 On this point Mr. ,T. Richards, M.E.. writes:* " The co- 
 efficient of (crushing) effect in such machines (as revolving rollers 
 and reciprocating machines) is as the area of the acting surfaces, 
 and the speed with which they approach each other. The 
 mistake in respect of the crushing power of rollers comes from 
 confounding their circumferential velocity with the working one 
 that is, with the parallel velocity at which the surfaces approach 
 and leave each other," that is, the velocity at right angles to 
 the crushing surface. Mr. Richards elsewhere shows that a 
 Blake machine in this way actually outruns a revolving roller, 
 although its crushing face travels more slowly than the periphery 
 of the roll. He proceeds : " The same principle holds good in 
 respect to stamps, the crushing surfaces having a parallel 
 approach, while with rotary machinery the approach is not 
 parallel, except on an imaginary line at the centre." As a 
 matter of fact, the product of two pairs of rolls amounts to only 
 about 40 to 50 per cent, of that which would be given by a fifty- 
 stamp mill on the same ore. In the absence of exact comparative 
 data regarding the same ore, no precise figures regarding the 
 relative capacities can be given. 
 
 A point to be noticed is that, in machines which act by 
 pressure applied on the principle of the lever, such as recipro- 
 cating-jaw crushers or rolls, the whole force necessary to crush 
 the lumps of quartz is transmitted to the fulcrum, this fulcrum 
 being represented in the case of rolls by the bearing surface at 
 the axle. The consequence is that the frame must be made very 
 strong and heavy, and the axle bearings attended to with great 
 care, if rapid wear is to be avoided. 
 
 Drying the Ore. Although rolls are essentially dry crushers, 
 nevertheless, if the ore is nearly pure quartz, a small percentage 
 of moisture in it is not a serious disadvantage. It, however, 
 the moist ore is argillaceous, the product from the rolls will not 
 readily pass through the screen, and the latter soon becomes 
 clogged, frequent stoppages for cleaning being thus necessitated. 
 It has been stated that 5 or 6 per cent, of moisture present in an 
 * Prod, of Gold and Silver in the United States, 1880, p. 369. 
 
234 THE METALLURGY OF GOLD. 
 
 ore does not seriously interfere with crushing, but this does not 
 accord with the experience in some mills. At any rate, in every 
 mill where dry crushing is used, some means of artificially drying 
 the ore is adopted. 
 
 The oldest method was to spread the ore, after the large lumps 
 had been removed by a grizzly and crushed to 1^ inch size, on 
 large flat areas heated from below by flues from the roasting 
 furnace. The floor was usually covered with iron plates. After 
 being dried, the ore was shovelled up and passed to the crushing 
 mill. This plan involved much additional handling of the ore, 
 and was a source of ill-health among labourers, besides requiring 
 great floor space. It has been superseded by the adoption of 
 inclined continuous-discharge, revolving iron cylinders, similar 
 to the Howell-White furnace, biit not lined with bricks. The 
 ore is passed through these, and is dried by the products of com- 
 bustion of a tire, which are also passed through it. One such 
 cylinder, of 3 feet in diameter and 18 feet long, will dry from 30 
 to 40 tons of ore per day, at a small cost for fuel and power. 
 
 An alternative furnace viz., Stetefeldt's shelf drying kiln 
 was described in a paper read by the inventor at the meeting of 
 the American Institute of Mining Engineers, held at Roanoke, 
 Virginia, in June, 1883. In principle it resembles the Hasen- 
 clever furnace, a number of shelves being arranged zig-zag above 
 each other in a vertical shaft, down which the ore slides, falling 
 from shelf to shelf, while the products of combustion from a 
 furnace rise through it. It is 21 feet high, and dries from 30 to 
 50 tons of ore per day. It is in wide use in the United States, 
 and although its first cost is considerable, its working expenses 
 are said to be lower than those of the rotary furnaces. 
 
 ROASTING. 
 
 The operation of roasting, as a preliminary to chlorination, 
 has for its object the expulsion of the sulphur, arsenic, 
 antimony and other volatile substances existing in the ore, 
 and the oxidation of the metals left behind, so as to leave 
 nothing (except metallic gold) which can combine with chlorine 
 when the ore is subsequently treated with it in aqueous solution. 
 For this purpose the ore is heated in a furnace, through which a 
 current of air is passed, salt being added if oxide of copper, lime, 
 magnesia, &c., are present. Ores containing much pyrites might 
 be freed from most of their sulphur by pile roasting, and then 
 subjected to fine crushing and a dead roast in a reverberatory 
 furnace, but the extra cost of handling would probably exceed 
 the saving due to the smaller consumption of fuel. This system 
 has not been tried in chlorination mills on an extensive scale. 
 The ordinary reverberatory furnace, worked by hand labour, is 
 
ROASTING. 235 
 
 in wider use than any other, especially where only a few tons, 
 or less, of concentrates are to be treated per day. Various 
 mechanical furnaces, capable of handling large quantities of ore, 
 have been devised to supersede the old-fashioned contrivance,, 
 and some of these will be described in the sequel. 
 
 Reverberaton i Furnaces The construction of the ordinary 
 reverberatory furnace is too well known to need detailed descrip- 
 tion here.* It consists of a vaulted chamber, containing the 
 ore ; through this chamber, the flames and products of combus- 
 tion from a furnace and a current of air are made to pass in a 
 horizontal direction above the ore, which is thus heated. The 
 ore is also stirred by hand with iron rakes, which are passed 
 through small working doors. The hearth of the vault (alsa 
 called the "laboratory" of the furnace) is formed of bricks placed 
 on edge (not flatwise, except where economy is studied rather 
 than durability), as close together as possible. No mortar is 
 used, but a little clay is plastered between the bricks. The 
 height of the furnace hearth is about 3| feet above the floor of 
 the building, on which the labourers stand, and the space under- 
 neath the hearth is either occupied by vaults or tilled with well 
 tamped rubble. The arch is usually one coui'se of bricks- 
 (8 inches) thick ; the height between it and the hearth is, in 
 long furnaces, about 24 inches near the bridge, and gradually 
 diminishes towards the other end. This height is less in. 
 short furnaces. The best fire-bricks are used for the fire-box and 
 bridge, and for the hearth and arch of the first few feet of the 
 "laboratory." The remainder is made of common brick. It is 
 necessary to have a damper in the flue to regulate the draught; 
 the aperture of the flue should not be on a level with the hearth,, 
 as in that case the loss by dusting is increased. The brickwork 
 of the furnace is supported by longitudinal and transverse iron 
 braces. The working doors have cast-iron frames, and are about 
 15 inches wide and 9 inches high. The fuel used must of course 
 be a long-flame coal, or wood; short-flame coal and coke are 
 inadmissible. 
 
 For the particular purpose of roasting pyritic ores before 
 chlorination, the temperature on the working floor of the furnace 
 must be low when the ore is first charged in, and high in the 
 later stages. If a single small floor is used, the fire must be 
 alternately checked and urged to secure these conditions. More- 
 over, when the roasting is nearly complete, the high temperature 
 required renders the gases passing into the flue very hot, and so- 
 a corresponding waste of fuel results. To prevent this waste, it 
 is customary for roasting furnaces to be built with a very long- 
 hearth, or to have several successive hearths, so as to utilise the 
 
 * See the Introduction to the Study of Metallurgy, by Prof. Roberts- 
 Austen, 1898, p. 291. Also Kustel's Roasting of Gold and Silver Ores, 1880,. 
 pp. 80-89 ; and Gruner's Traite de Metalluryie, p. 265. 
 
236 THE METALLURGY OF GOLD. 
 
 waste heat from the portion of the working floor next the fire. 
 Furnaces with three floors at slightly different levels are much 
 favoured ; in these a charge remains for a few hours on each of 
 the floors in succession. It is first placed on the floor farthest 
 removed from the fire, and, after a time, is raked down on to the 
 middle hearth, and thence to that nearest the fire, fresh charges 
 being put on the spaces just cleared, so that there are always 
 three charges in the furnace in various stages of oxidation. 
 
 The most usual form in the Western States of America is the 
 "4-hearth," in which the length of the hearth is four times its 
 width, so that the dimensions are, say, width 15 feet, length 
 60 feet. In this case there should be eight working doors on 
 ach side. Instead of three floors at different levels, a single 
 continuous floor, gently sloping from the flue towards the fire, is 
 in use at many works in Australia, Mexico and the United 
 States. At Suter Creek a continuous-hearth furnace is 12 feet 
 wide and 80 feet long, and the mineral is worked in three 
 distinct parts, as though there were three floors. The angle of 
 slope is made large in some Australian furnaces, so that in the 
 course of the " rabbling " or stirring, the ore continually travels 
 towards the fire-box. Furnaces with two or three superposed 
 floors are also used to a limited extent ; the lowest floor is next 
 the fire-box, and communicates by a vertical flue with the floor 
 above, and so on. The ore is charged-in on the top floor, and 
 after a time is raked down through the vertical flue on to the 
 next floor. In this case the floors are heated by the gases passing 
 below them as well as above them, and fuel is economised, but 
 the furnaces are costly to build and to keep in repair. 
 
 It was proposed many years ago to insert a drop of 10 feet 
 between the finishing floor and the floor next to it. The charge, 
 when already red hot, would thus fall vertically downwards 
 in a thin shower against a current of hot air. Some furnaces 
 are said to have been built in this way, but it seems that none 
 are now in existence. The principle is excellent, and is utilised 
 in the Stetefeldt furnace. 
 
 When the pyrites to be roasted are rich, it may be an ad- 
 vantage to build dust chambers to ordinary reverberatory 
 furnaces. The amount of gold contained in the dust thus 
 recovered is usually only 1 or 2 per cent, of that contained in 
 the ore, so that in some cases it may be a long while before the 
 dust chamber pays for itself, even if that point is ever reached. 
 
 The operation of roasting pyrites in an ordinary furnace with 
 three floors may be described as follows : The furnace being 
 hot, and the flame from the fire-box reaching completely across 
 the first floor, the ore is charged-in on the third floor and spread 
 out by the rabbling tool. The weight of the charge may be 
 taken as from 12 to 18 pounds per square foot of floor space, 
 varying according to the nature of the ore, a high percentage 
 
ROASTING. 237 
 
 of sulphur necessitating small charges. The layer of ore is 2 
 or 3 inches deep. It is not spread quite evenly, but made 
 to form a series of parallel ridges by means of the rabbling tool, 
 so as to increase the surface exposed to the air. The working 
 doors may be closed at first to heat the ore quickly. Moisture 
 is at once given off in great quantities, and the sulphur soon 
 begins to burn with a blue flame. When this is seen to take- 
 place, all the working doors are opened, and the charge is ener- 
 getically rabbled, with little intermission, until the sulphur 
 flame disappears. If this is not done, clots are formed which 
 are afterwards difficult to break up. The air for the combustion 
 of the sulphur is supplied by holes in the fire-bridge and from 
 the working doors. The flames and heated products of com- 
 bustion from the fire tend to rise above the colder air, and 
 move along next to the arch of the furnace, while the air forms 
 a sheet between these gases and the ore. In practice, although 
 the air is introduced below them, nevertheless the reducing 
 gases from the fire partially mix with the air, and greatly 
 reduce its oxidising power. Moreover, the combustion of the- 
 sulphur in the ore on the first two floors further reduces the 
 amount of free oxygen present in the current of air, and 
 roasting on the third floor is, therefore, largely dependent on 
 air derived from the working doors. 
 
 When the sulphur flames have abated, and the charge has- 
 been heated almost to redness, it is transferred to the middle 
 floor, where it is raised to a dull red heat and most of the oxida- 
 tion is performed ; during this stage the ore swells considerably, 
 so as to occupy much more than its original bulk. All the lumps 
 previously formed should be broken up on this floor. Rabbling 
 is continued until the ore is uniformly dull throughout, so that, 
 on turning it over, the fresh surfaces appear but little brighter 
 than that which has been exposed for some time. The charge is 
 then transferred to the floor next the fire. There is now little 
 risk of the formation of lumps, and the charge may be allowed 
 to reach a bright red heat. Rabbling is of less importance than 
 before, as little oxidation takes place, the chief reaction which 
 occurs being the decomposition of the sulphates already formed. 
 As long as this is still going on. the ore emits the odour of 
 sulphur dioxide. When no further odour can be detected, and 
 the ore can be piled up so as to maintain a vertical face, shows 
 no brijrht specks on its glowing surface, emits no sparks if 
 some of it is tossed up by the working tool, and is inclined to 
 become black very readily from cooling, the charge is said to be 
 " dead " or " sweet," and is ready to be withdrawn. It should 
 be observed that, when the ore Contains much sulphur, its- 
 particles at a low red heat appear less coherent than when cold, 
 and flow almost like water so that the charge cannot be made 
 to form a heap with steep sides. Care must therefore be taken. 
 
238 THE METALLURGY OP GOLD. 
 
 when the ore is on the middle floor to prevent any part of the 
 charge from flowing out of the working doors, which it is very 
 liable to do when being rabbled. 
 
 Kiistel states * that the best means of rapidly ascertaining 
 whether a charge is completely roasted is to throw a little of 
 the ore into some water, and then to plunge a bright iron rod 
 into the liquid. If the rod remains bright the ore is ready 
 for withdrawal, but if sulphates still remain undecomposed, 
 the surface of the iron will be darkened. This is not a safe test 
 with all classes of ore, as the presence of sulphate of iron in the 
 water would not be detected in this way. A more trustworthy 
 ^ind equally simple test is to add a few drops of chloride of 
 barium to the water. A white cloud, which consists of BaSO 4 , 
 indicates the presence of soluble sulphates. In most cases 
 the water need not be filtered before it is tested, but even 
 when filtration is necessary the whole operation can be per- 
 formed in two or three minutes. The charge is withdrawn 
 by a scraper, and falls by gravity through a hole in the floor 
 of the furnace near the working door into a pit below. This 
 hole is covered by a plate while roasting is being performed. 
 
 The time of roasting depends chiefly on the ore, but may be 
 ishortened by more continuous rabbling than most workmen can 
 perform, the work being somewhat exhausting. In a three- 
 floor furnace, concentrates with 15 per cent, of sulphur usually 
 remain eight hours on each of the three floors. The fuel used 
 is either wood or flaming-coal. If the flame from the coal is not 
 long enough to reach across the first floor, this will not be heated 
 uniformly; in that case the part of the charge next the fire-bridge 
 is finished first and must be moved away, while that from 
 the other end of the floor is brought up nearer to the tire. 
 This causes a great increase in the labour, besides occasionally 
 leading to the withdrawal of a charge of which a part is not quite 
 "dead." The draught is regulated by the damper in the flue 
 leading to the dust chamber, and by opening and closing the 
 working doors. 
 
 Chemistry of Oxidising Roasting. Professor Roberts- 
 Austen discusses as follows f the roasting of a " mixture 
 consisting of sulphides mainly of iron and copper, with some 
 sulphide of lead, small quantities of arsenic and antimony as 
 arsenides, antimonides, and sulpho-salts, usually with copper as 
 .a base. The temperature of the furnace in which the operation 
 is to be performed is gradually raised, the atmosphere being an 
 oxidising one. The first effect of the elevation of the tempera- 
 ture is to distil off sulphur, reducing the sulphides to a lower 
 stage of sulphurisation. This sulphur burns in the furnace 
 
 * JRoastiny of Gold and Silver Ores. San Francisco, 1880. 
 t Presidential address to the Chemical Section, British Association, 
 ^Cardiff Meeting, 1891. 
 
ROASTING. 239 
 
 atmosphere to sulphurous anhydride (SO 2 ), and coming in 
 contact with the material undergoing oxidation is converted into 
 sulphuric anhydride (SO 3 ). It should be noted that the material 
 of the brickwork does not intervene in the reactions, except by 
 its presence as a hot porous mass, but its influence is, neverthe- 
 less, considerable. The roasting of these sulphides presents a 
 good case for the study of chemical equilibrium. As soon as the 
 sulphurous anhydride reaches a certain tension, the oxidation of 
 the sulphide is arrested, even though an excess of oxygen be 
 present, and the oxidation is not resumed until the actions of 
 the draught change the conditions of the atmosphere of the fur- 
 nace, when the lower sulphides remaining are slowly oxidised, 
 the copper sulphide being converted into copper sulphate, mainly 
 by the intervention of the sulphuric anhydride, formed as in- 
 dicated. Probably by far the greater part of the iron sulphide 
 only becomes sulphate for a very brief period, being decomposed 
 into the oxides of iron, mainly ferric oxide, the sulphur passing 
 off. Any silver sulphide that is present would have been 
 converted into metallic silver at the outset were it not for the 
 simultaneous presence of other sulphides, notably those of copper 
 and of iron, which enables the silver sulphide to become con- 
 verted into sulphate. The lead sulphide is also converted into - 
 sulphate at this low temperature (viz., about 500). The heat is 
 now raised still further with a view to split up the sulphate of 
 copper, the decomposition of which leaves oxide of copper. If, 
 as in this case, the bases are weak, the sulphuric anhydride 
 escapes mainly as such ; but when the sulphates of stronger 
 bases are decomposed the sulphuric anhydride is to a great 
 extent decomposed into a mixture of sulphurous anhydride and 
 oxygen. The sulphuric anhydride, resulting from the decom- 
 position of this copper sulphate, converts the silver into sulphate, 
 .and maintains it as such, just as, in turn, at a lower temperature, 
 the copper itself had been maintained in the form of sulphate 
 by the sulphuric anhydride eliminated from the iron sulphide. 
 "When only a little of the copper sulphate remains undecom posed, 
 the silver sulphate begins to split up (viz., at about 700) . . . 
 partly by the direct action of heat alone, and partly by reactions 
 such as those shown in the following equations : 
 
 Ag 2 S0 4 + 4Fe 8 O 4 = 2Ag + 6Fe 2 3 + S0 2 
 Ag 2 S0 4 + Cu 2 O = 2Ag + CuS0 4 + CuO 
 
 The charge still contains lead sulphate, which cannot be com- 
 pletely decomposed at any temperature attainable in the roasting " 
 furnace except in the presence of silica. . . . The elimination 
 of arsenic and antimony gives rise to problems of much interest, 
 and again confronts the smelter with a case of chemical equi- 
 librium. For the sake of brevity it will be well for the present 
 to limit the consideration to the removal of antimony, which 
 
240 THE METALLURGY OF GOLD. 
 
 may be supposed to be present as sulphide. Some sulphide of 
 antimony is distilled off, but this is not its only mode of escape. 
 An attempt to remove antimony by rapid oxidation would be 
 attended with the danger of converting it into insoluble anti- 
 moniates of the metals present in the charge. In the early 
 stages of the roasting it is, therefore, necessary to employ a very 
 low temperature, and the presence of steam is found to be useful 
 as a source of hydrogen, which removes sulphur as hydrogen 
 sulphide, the gas being freely evolved. The reaction 
 
 Sb 2 S 8 + 3H 2 = 3H,S + 2Sb 
 
 between hydrogen and sulphide of antimony is, however, endo- 
 thermic, and could not, therefore, take place without the aid 
 which is afforded by external heat. The facts appear to be as 
 follows : Sulphide of antimony, when heated, dissociates, and 
 the tension of the sulphur vapour would produce a state of 
 equilibrium if the sulphur thus liberated were not seized by the 
 hydrogen, and removed from the system. The equilibrium is 
 thus destroyed, and fresh sulphide is dissociated. The general 
 result being that the equilibrium is continually restored and 
 destroyed until the sulphide is decomposed. The antimony com- 
 bines with oxygen and escapes as volatile oxide, as does also 
 the arsenic, a portion of which is volatilised as sulphide. 
 
 " The main object of the process which has been considered 
 is the formation of soluble sulphate of silver." The reactions, 
 however, are precisely similar in an ordinary oxidising roast. 
 
 The following remarks on the decomposition of the various 
 minerals present in complex ores may be of use in assisting the 
 student to understand the reactions which proceed in the roasting 
 furnace : 
 
 1. Iron Pyrites, FeS 2 . On heating this compound, sulphur is 
 volatilised, the reactions being probably expressed thus : 
 
 3FeS 2 = Fe 3 S 4 + S 2 
 7FeS 2 = Fe 7 S 8 + 3S 2 
 
 The sulphur burns to SO 2 , which is partly converted by the 
 heated quartz, &c.,* into SO 3 , uniting with the free oxygen 
 present. The ferrous sulphate formed by this sulphuric acid is 
 split up by the heat and the ferrous oxide (FeO) converted into 
 ferric oxide (Fe 2 O 3 ) which gives the ore a red colour when cold. 
 If the temperature of the part of the charge next the fire-bridge 
 has been too high, or if the charge is kept too long in the 
 furnace, some magnetic oxide is formed, thus : 
 
 3Fe 2 3 = 2Fe 3 4 + 
 
 This is an undesirable change, as the magnetic oxide is acted on. 
 by chlorine far more readily than the sesquioxide. 
 
 * Plattner's MetcUlurgische Ittistprozesse, Freiburg, 1856. 
 
ROASTING. 241 
 
 2. Copper Pyrites. The decomposition of the copper sulphate 
 formed in the furnace leaves a mixture of cuprous and cupric 
 oxides, both soluble in chlorine. 
 
 3. Galena, PbS. The presence of this mineral in any but 
 small quantities is very detrimental, as both lead sulphate and 
 lead silicate (formed by its decomposition in the presence of 
 silica) are very fusible, and, at the temperature required to split 
 up copper sulphate, cause the ore to become pasty and form 
 lumps. Roasting must be performed very slowly and cautiously 
 to avoid this effect. 
 
 4. Arsenical Pyrites, FeAsS. Arseniates of iron, copper, lead, 
 <fec., when formed are not easily decomposed, as they resist a 
 high temperature, and are only slowly converted into sulphates 
 by sulphuric acid at a red heat. It is, therefore, desirable to 
 avoid their formation, and with this end in view the precautions 
 which have been already mentioned in the extract from Prof. 
 Roberts- Austen's address are taken. 
 
 5. Antimonial Sulphides are still more difficult to deal with, 
 the antimoniates formed being less easily decomposed than ar- 
 seniates. Their formation is avoided in the manner already 
 described. 
 
 6. Blende, ZnS, forms oxide and sulphate of zinc, of which the 
 latter can only be split up by a very high temperature. At a 
 bright red heat a basic sulphate is formed which is converted 
 to oxide at a white heat. If blende is roasted at a high tem- 
 perature and with a plentiful supply of air, sulphate of zinc is 
 not formed to a large extent. 
 
 Elimination of Arsenic and Antimony. Mr. H. M. Howe, in 
 explaining how this is effected, distinguishes three horizontal 
 zones in the ore :* (1) the upper surface, where oxidation is only 
 slightly hindered by sulphurous and sulphuric acids and by the 
 products of combustion of the fuel ; (2) the middle layers, where 
 oxidation proceeds to a very limited extent ; (3) the lowest 
 layers, where " a pellet of ore is simply exposed to the action of 
 the other pellets with which it is in contact, of volatilised 
 sulphur, and of sulphurous and sulphuric anhydrides generated 
 by the action of sulphur on previously formed metallic oxides." 
 He proceeds "The expulsion of arsenic and antimony as 
 sulphides is favoured in the middle and lower zones by the 
 presence of volatilised sulphur, mixed with sulphurous acid and 
 at most a very limited supply of free oxygen and sulphuric acid. 
 In the upper part of the middle layer, to which a small amount 
 of free oxygen penetrates, we have the gently oxidising conditions 
 favourable to the formation of arsenious acid and trioxide of 
 antimony. In the upper zone the stronger oxidising conditions 
 rather favour the formation of fixed arseniates and antimoniates, 
 though, even here, part of the arsenic and antimony may 
 * Copper Smelting, U.S.A. Geol. Survey, Washington, 1885. 
 
 16 
 
242 THE METALLURGY OF GOLD. 
 
 volatilise and escape while passing through their intermediate 
 volatile condition of arsenious acid and trioxide of antimony." 
 On stirring the mass, these arseniates and antimoniates, being 
 exposed to the reducing action of volatilised sulphur and unde- 
 com posed sulphides in the lower zones, may again be converted 
 into volatile oxides. Protoxide of iron, suboxide of copper, and 
 sulphurous acid are also efficacious in reducing arsenic acid, 
 higher oxides of iron and copper, and sulphuric acid being 
 formed. "Thus, every individual atom of arsenic may travel 
 forth and back many times through the volatile condition, being 
 oxidised at the surface and reduced below the surface, . . . 
 and every time it arrives at this volatile condition an oppor- 
 tunity is offered it to volatilise and escape." If a small quantity of 
 coal or coke dust is mixed with the ore, after it has been completely 
 oxidised, and the air excluded, the arseniates and antimoniates 
 are again reduced to the lower oxides, and, if they are "carried 
 past the volatile state," i.e., reduced to metals, they may be 
 again passed through it by an oxidising atmosphere. " Of course 
 the expulsion of arsenic and antimony is favoured by the presence 
 of a large proportion of pyrites, both because the sulphur distilled 
 from the pyrites tends to drag them off as sulphides, and because 
 the presence of the pyrites prolongs the roasting, and thus 
 increases the number of times which the arsenic and antimony 
 pass back and forth past their volatile conditions ; hence, it is 
 sometimes desirable to mix pyrites with impure ores to further 
 the expulsion of their impurities." 
 
 The Use of Salt in Boasting. Certain ores require the 
 addition of salt in roasting in order to chloridise material which 
 would otherwise absorb chlorine when the ore came to be 
 " gassed," and so cause additional expense as well as incon- 
 venience. If silver as well as gold is to be extracted from 
 the ore, the addition of salt is necessary in order to form 
 chloride of silver in the furnace, since metallic silver is not 
 attacked by chlorine at the highest temperature ever employed 
 in the leaching vat. The silver chloride is then dissolved out 
 by hyposulphite of soda or some other solvent either before or 
 after the extraction of the gold. 
 
 Even if no silver is present, an ore must be roasted with salt 
 if it contains much copper (as sulphide, or as an oxidised salt), 
 lime, magnesia, or other substance which, after being subjected to ^ 
 an oxidising roasting, is rapidly attacked by chlorine at ordinary 
 temperatures. The salt is usually added towards the end of 
 the operation, when no sulphides and only a small percentage of 
 sulphates are left undecomposed ; sometimes, however, the ore 
 and salt are mixed before charging in. To some sulphides only 
 5 pounds of salt per ton of ore are added, but others require as 
 much as 90 pounds per ton. The weight of salt added must be 
 at least six to eight times that of the silver present in the ore. 
 
BOASTING. 243 
 
 If a large amount of salt is used, it is desirable to leach the 
 roasted ore with water, before treating it with chlorine gas, in 
 order to remove the coating of soluble sulphates and chlorides 
 remaining on the surface of the granules of ore. 
 
 The chemical action of the salt is due to a double decomposi- 
 tion between it and the sulphates of the heavy metals, by which 
 sulphate of soda and the chlorides of the heavy metals are pro- 
 duced. The following general equation approximately represents 
 the reaction : 
 
 2NaCl + RS0 4 = RC1 2 + Na 2 S0 4 
 
 Chlorine is also set free by the action of sulphuric anhydride 
 on salt, and the presence of water vapour induces the formation 
 of much hydrochloric acid. These gases act directly on the 
 several constituents in the ore, forming chlorides and oxy- 
 chlorides. The metallic chlorides and oxychlorides formed are 
 in many cases volatile (e.g.j the compounds of copper, iron, lead, 
 arsenic, antimony, &c.), and, in passing off, the volatile compounds 
 carry away with them varying proportions of gold and silver, 
 which, as a rule, are not recoverable in the dust chambers. The 
 chloride of copper is especially active in causing these losses. 
 
 Other reactions which probably take place are as follows : 
 
 1. Ferrous sulphate, acted on by salt at a red heat in presence 
 of air, yields hydrochloric acid and chlorine, which act on the 
 gold and silver, while ferric sesquioxide and sodic sulphate are 
 produced. 
 
 2. Ferric chloride, Fe 2 01 6 , is also produced at the same time. 
 This is volatile, and chloridises silver with great energy at a red 
 heat, sesquioxide of iron being produced. 
 
 3. Cupric chloride, CuCl 2 , is easily decomposed into cuprous 
 chloride, Cu 2 Cl 2 , and free chlorine, or into the oxychloride, 
 Ou 2 O . C1 2 , and free chlorine. The vapours of CuCl 2 thus give 
 rise to further supplies of nascent chlorine available for the 
 chlorination of the silver. 
 
 4. Arsenic and antimony form volatile chlorides which are 
 decomposed by means of oxygen and water vapour, yielding 
 arsenious and antimonious acids and nascent chlorine or hydro- 
 chloric acid. 
 
 It is thus obvious that the presence of base minerals is 
 advantageous in that they may cause nascent chlorine to be set 
 free in the presence of silver in all parts of the furnace. On the 
 other hand, the loss of gold is increased by any increase in the 
 quantities either of silver or of the base metals, since in the 
 former case the time of the roasting is prolonged. The best 
 chloridising effect is obtained in a highly oxidising atmosphere, 
 so that very little sulphur is required in the ore, and, if much 
 is present, the practice of eliminating the greater part before add- 
 ing the salt is not likely to be attended with any diminution in 
 
244 THE METALLURGY OF GOLD. 
 
 the percentage of silver chloride formed. Moreover, the water of 
 crystallisation in the salt promotes the formation of hydrochloric 
 acid. When salt is used in roasting, the ore should be allowed 
 to cool slowly in heaps after being withdrawn from the furnace. 
 If treated in this way, a higher percentage of the silver, &c., is 
 found to be chloridised than if the ore is wetted down at once, 
 or even spread out to cool in a thin laver. Chlorine continues 
 to be evolved for a long time after the withdrawal of the charge 
 has taken place, the heaps smelling strongly of the gas. 
 
 Losses of Gold in Boasting. Plattner proved in 1856* that 
 in the oxidising roasting of ordinary auriferous pyrites, a loss of 
 gold can take place only when the operation is carried on so 
 rapidly that fine particles are carried off mechanically by the 
 draught. This conclusion, as far as sulphides and arsenides are 
 concerned, has been confirmed by Kustel,f and by Prof. S. B. 
 Christy,| but the latter adds that it is extremely difficult to 
 prevent all mechanical loss by dusting, which is caused by even 
 a moderate draught. Kiistel records the loss of 20 per cent, of 
 the gold present during the oxidising roasting of certain 
 tellurides of gold and silver, and states that this is not a 
 mechanical loss, but is due to volatilisation. The effect of 
 tellurium on the volatilisation of metallic gold is shown on p. 7. 
 
 The losses of gold which are sustained when salt is added to 
 the furnace charge have been fully investigated by Prof. S. B. 
 Christy, and the following account is mainly derived from his 
 paper on the subject. Kiistel had previously found that a 
 telluride ore, on being roasted with 4 per cent, of salt, lost 8 per 
 cent, of its gold before the ore was red hot. Aaron || found that 
 certain ores, consisting of simple pyrites, suffered great loss of 
 gold in roasting with salt which had been added at the com- 
 mencement of the operation ; only a small part of this gold was 
 condensed in the flue, in which was found a yellowish fluffy 
 precipitate, consisting largely of chlorides of copper and iron, 
 and containing nearly 30 ozs. of gold to the ton. He found 
 that the loss was greatly reduced by diminishing the quantity 
 of salt, and by reserving it until the dead roasting was nearly 
 complete. 
 
 In the chloridising roasting of a Mexican ore, consisting mainly 
 of magnetite and pyrites with 3-5 to 7 per cent, of chalcopyrite, 
 Mr. 0. A. Stetefeldt found the losses of gold to be from 42-8 to 
 93 per cent, of the total gold contained. He stateslf that "there 
 is no doubt that the volatilisation of the gold takes place with 
 
 * Metallurgische Rdstprozesse, Freiburg, p. 128. 
 t Roasting of Gold and Silver Ores, 1880, p. 56. 
 $ Tram. Am. Insi. Mng. Eng., 1888. 
 Loc. cit. 
 
 || Leaching of Gold and Silver Ores, 1881, p. 121. 
 1 Trans. Am. Inst. Mng. Eng., vol. xiv., p. 339. 
 
BOASTING. 245 
 
 that of the copper chlorides. The loss increased with the 
 quantity of these chlorides formed and volatilised." He further 
 shows, however, that the presence of copper chloride is not the 
 only possible cause of loss, since an ore consisting of hard white 
 quartz, intimately mixed with about 7 per cent, of calcite and a 
 -iittle pyrites, lost 70 to 80 per cent, of its silver, and 68 to 85 per 
 cent, of its gold, when roasted with 5 per cent, of salt. When 
 subjected to an oxidising roast, no loss of gold took place. The 
 reason for the extraordinary behaviour of this ore was not 
 discovered. 
 
 Prof. Christy found that, in the ores on which he experimented 
 on a small scale in a muffle furnace, a greater loss was sustained 
 by adding the salt near the end of the roasting operation, than 
 by mixing the same weight of salt with the ore at the start. He 
 explained that this is due to the fact that the amount of gold 
 volatilised varies with the amount of chlorine which comes in 
 contact with it. When the salt is added at the start, the 
 chlorine is at first removed by the sulphur as fast as it is formed, 
 escaping as chloride of sulphur, and thus the gold is protected 
 from attack. When the salt is added after a long oxidising 
 roast, the chlorine is rapidly generated (the ore being red hot 
 and containing large quantities of sulphates), and the gold 
 is no longer protected from attack by the sulphur. The loss of 
 gold is also in all cases increased by working at a higher tem- 
 perature, owing to the larger amount of chlorine generated, and 
 to the increase in the volatility of the gold. It is apparent from 
 the results given on p. 22 that the temperature used in chlorid- 
 ising roasting must be very carefully regulated, the loss of gold 
 being increased far more by high temperature than by a length- 
 ening of the time in the furnace. Moreover, the salt must be 
 reduced to the least possible quantity. 
 
 The advantage found to be gained in practice by adding salt 
 near the end of the operation is due to the fact that, in the 
 continuous roasting of ores in long-bedded furnaces, the gases 
 given off from the finishing floor pass over a great length of 
 comparatively cold, unsalted, and unoxidised ore before reaching 
 the flue. The quantity of gold chloride mixed with the chlorine 
 which is evolved from the red-hot ore as soon as the salt is added / 
 is no doubt large, but the S0 2 from the colder ore, and the steam y 
 from the fuel, "offer excellent means for the reduction of the 
 chloride of gold right within the furnace, while the most efficient 
 means probably is the pyrites themselves," which have been 
 proved to be readily capable of condensing gold on their surface. 
 If all the salt is added at the start, there is a continued vola- 
 tilisation of chloride of gold throughout the furnace, and a less 
 favourable opportunity for it to condense. The difference 
 between the results in the muffle and in the reverberatory 
 furnace is thus explained. 
 
246 THE METALLURGY OF GOLD. 
 
 At Nevada City, at the Merrifield Mine, and in other works 
 in the neighbourhood, the old-fashioned long furnace, with a 
 single step separating the finishing hearth from the rest-ef the 
 furnace, was still used in 1888.* These furnaces are from 
 55 to 65 feet long, holding from 6 to 9 tons, and producing 
 about 3 tons of roasted ore per day, so that the ore remains 
 in the furnace from two to three days. The custom there 
 was to give the ore a long oxidising roast at a low red 
 heat, ending at a low cherry-red heat, and then, when the 
 ore reached the finishing floor, the temperature was slightly 
 lowered, and the salt added. The salt was stirred thoroughly 
 into the ore, and as soon as it was " dissolved " by the roasted 
 ore i.e., in about half an hour the charge was drawn into 
 the cooling pit. This lowering of the temperature is evidently 
 of great importance in reducing the loss, while the dura- 
 tion of the roasting is regarded as less material, so long as 
 no salt is present. These mills are on custom work, charging 
 $15 to $17 per ton of ore for treatment, and guaranteeing a 
 yield of 90 per cent, of the gold and 60 per cent, of the silver. 
 Their method of roasting seems to be considered in California as 
 that best suited to concentrates containing a high percentage of 
 sulphur, but their loss in roasting has not been ascertained. 
 The best method of roasting any particular ore, however, cannot 
 be determined by any general rule, and exhaustive experiments 
 must be made in every case before a definite course of procedure 
 is finally adopted. 
 
 At one of the Californian chlorination mills it was found by 
 experiment in 1882 that nearly 50 per cent, of the gold and 
 28 per cent, of the silver was being lost by volatilisation. In 
 this case the pyrites was roasted on two hearths for thirty-six 
 hours, 1 per cent, of salt being added four hours before the 
 charge was drawn. The reason for the great loss was thought 
 by Professor Christy to be the high temperature of roasting, 
 particularly on the charging-in floor. 
 
 The variation of the loss in different ores which are treated 
 precisely alike is doubtless due partly to the presence or absence 
 of metals forming volatile chlorides which carry off the gold, and 
 partly to the physical condition of the latter, the volatilisation, 
 being greater if it is in a state of minute subdivision. 
 
 MECHANICAL FURNACES. 
 
 The furnaces which have been designed with the object of 
 saving the labour necessary to work the reverberatory furnaces 
 may be divided into four classes, viz. : 
 
 * Trans. Am. Inst. Mng. Eng., vol. xiv., p. 340. 
 
MECHANICAL FURNACES. 247 
 
 1. Stationary hearth furnaces, supplied with iron hoes moved 
 by machinery by which the ore is rabbled. The O'Hara, Spence, 
 and Pearce Turret furnaces are examples of this class. 
 
 2. Rotating- bed furnaces, in which the hoes or stirrers are 
 stationary, while the bed supporting the ore revolves, so that 
 the latter is stirred by the hoes. An example of this class used 
 to roast gold ores as a preliminary to chlorination is afforded by 
 the furnace used at the Bunker Hill Mine, California. 
 
 3. Rotating cylindrical furnaces, which consist of brick-lined 
 iron cylinders capable of being rotated, so that the ore is 
 tumbled over and over by their motion while it is being roasted. 
 Examples are the Bruckner, the White-Howell, and the Hof- 
 mann furnaces. 
 
 4. Shaft furnaces, in which the powdered ore falls by gravity, 
 in a shower, through an ascending column of hot air, the oxida- 
 tion being effected in the course of the fall. The Stetefeldt 
 furnace, which is the only one based on this principle, is not 
 used for dead roasting, as it is not adapted for the purpose. It 
 is used for the chloridising roasting of silver ores, and will not 
 be described in this volume. 
 
 In mechanical furnaces the consumption of fuel is often much 
 greater than that in the long-bedded reverberatory furnace, 
 where it is usually from 10 to 20 per cent, of the weight of the 
 ore if flaming coal is used. 
 
 1. Furnaces with Mechanical Stirrers. O'Hara Furnace. 
 This is the oldest mechanical furnace, and it bears a great 
 resemblance to the old-fashioned reverberatory furnace. It has 
 two superposed hearths, in each of which the arch is very low, 
 so as to confine the heat close to the ore. An endless chain, set 
 in motion by suitable machinery, passes through the furnace, 
 resting on the upper hearth, and returns along the lower hearth. 
 Attached to the chain at proper intervals are iron frames of 
 a triangular shape : on these frames are a number of ploughs 
 or hoes set at an angle, so that one set of hoes turns the ore to 
 the centre, and the next set turns it in an opposite direction 
 towards the walls. The ploughs thus stir the ore thoroughly, 
 and at the same time move it gradually towards the fire. The 
 ore falls from the upper to the lower hearth by gravity, and 
 similarly falls from the lower hearth into a pit when it arrives 
 at the hottest place in the furnace. The ore is from five to ten 
 hours in the furnace, according to the amount of sulphur contained 
 in it. In the modern form the hearths are each 8 feet wide and 
 90 feet long, and the capacity is about 35 tons per day, at an 
 expenditure of about 2J H.P. The ore is not roasted dead, 
 however, in this case, about 6 per cent, of sulphur remaining 
 in it. No less than twenty- three of these furnaces, with Brown 
 and Allen's improvements, were in operation in the Uuited 
 States in 1893.* 
 
 * Eng. and Mng. Journ., May 20, 1893, p. 463. 
 
248 THE METALLURGY OF GOLD. 
 
 Spence Furnace. This is perhaps the best shelf furnace yet 
 devised, and has been used successfully in the roasting of copper 
 sulphides in the United States. It consists of a series of four 
 or five superposed hearths, communicating with one another 
 by means of vertical passages at alternate ends of the furnace. 
 The rabbling is done by rakes having a reciprocating motion 
 longitudinally in the furnace and armed with teeth of triangular 
 section having the apices pointing in the opposite direction to 
 that in which the ore travels. When the rakes move in the 
 direction in which these apices point, the ore is only stirred, but 
 when the return movement is made, the flat sides of the teeth 
 push some of the ore along the floor of the furnace, and a part 
 falls through the vertical passages on to the lower hearths. The 
 ore is charged in through a hopper and discharged roasted into 
 a pit. The number of floors may be varied according to the 
 character of the ore to be treated. The motion of the rakes 
 should not be continuous, as in that case wearing of the teeth 
 becomes very rapid. It is better to set them in motion (by 
 racks and pinions) for a few minutes, and then to withdraw 
 them for a while in order to allow them to cool. It is claimed 
 for this furnace that if once made hot, the combustion of the 
 sulphur in pyritic ores supplies the place of other fuel, so that 
 the roasting is perfectly performed with no further cost than 
 that of power for the rakes. It is of course obvious that 
 without an extra fire, the lowest hearth, which is not heated 
 from below and which is the place where the air is admitted, 
 would tend to become the coldest, and would then be in no way 
 adapted for the decomposition of the sulphates formed on the 
 upper hearths, a process which involves endothermic reactions, 
 and, therefore, requires external heat. The Spence furnace 
 without a fire may, however, be valuable for the production of 
 sulphates of iron or copper. 
 
 At the Treadwell Mine, Alaska, BrUckner cylinders were 
 formerly used,* but discarded on account of the amount of dust 
 made and carried into the flue, and the large amount of fuel 
 consumed, and Spence furnaces introduced. Each furnace had 
 four hearths, the lowest being strongly heated to decompose the 
 sulphates. The ore remained in the furnace for sixteen hours, 
 and 3 per cent, of salt was added on the hearth next above the 
 lowest one. The rakes of the upper hearths lasted a long time, 
 but those of the finishing hearth were worn out in three months, 
 and were often broken sooner. These furnaces were not found 
 to be of any use until converted from muffle into reverberatory 
 furnaces so that the products of combustion of the fire passed 
 directly over the ore to be roasted. Six of the double Speuce 
 furnaces were built at a great expense, and the cost of roasting 
 in them was found to be less than half that in the Briickner, but 
 * Eng. and Mng. Journ., April 11, 1891. 
 
MECHANICAL FURNACES. 249 
 
 their capacity was small and the amount of fuel required was 
 found to be very great. An ordinary reverberatory furnace 
 was, therefore, built, and the results obtained were so satisfactory 
 that three others were added, and the Spence furnaces thrown 
 out of work. They are apparently not now used at any chlorin- 
 ation mill, but might possibly be adapted to some ores. 
 
 One of the chief causes of difficulty and expense in working 
 furnaces with mechanical stirring apparatus is that the iron hoes 
 gradually become heated, and they are then rapidly corroded by 
 the sulphur in the ore. In hand stirring, the rabbling tool is 
 withdrawn as soon as it is hot, and allowed to cool, while 
 another is substituted for it meanwhile, thus prolonging its life. 
 To imitate this action, the Spence hoes are sometimes arranged 
 to work for a while and then to be withdrawn completely to 
 cool. In the O'Hara furnace, also, it is better to have only one 
 hearth, the hoes passing completely outside the furnace on the 
 return journey. In spite of such arrangements, however, the 
 trouble and expense caused by the wearing of the hoes was very 
 great until the device was adopted of placing the moving parts 
 of the mechanism outside the ore-hearth with the stirring hoes 
 passing through a slot in the furnace wall. This device is used 
 in the three furnaces described below, all of which are in suc- 
 cessful operation in the United States. Since their introduction 
 the cojjst of roasting has been much reduced and their use is 
 likely to be widely extended. 
 
 Pearce Turret Furnace. This consists of an ordinary rever- 
 beratory hearth built in an annular form (see Fig. 45/). In 
 the centre of the circular space surrounded by the hearth is 
 a vertical iron column carrying four hollow horizontal arms 
 projecting through a slot into the reverberatory hearth which 
 they cross transversely. The column revolves and the arms 
 carry rabble blades which traverse the hearth, stirring the ore 
 and moving it round the circle by degrees. "Air is forced 
 through the hollow arms and is discharged against the rabble 
 blades, performing the double duty of cooling the iron work 
 and of furnishing heated air for the oxidation of the ore." The 
 ore is discharged automatically after passing once round the 
 furnace. Two or more fireplaces are used. The hearth is 6, 7, 
 or 8 feet wide, and is now sometimes built in two floors, one 
 above the other. Moreover, the top of the arch is sometimes 
 used as a drying hearth. An improvement on the air blast for 
 cooling the rabble arms consists in supplying them with water- 
 jackets, by which their life is lengthened, the air being let in 
 through the fireboxes and the apertures in the outer wall. 
 Moreover, in imitation of the intermittent movement in the 
 Brown furnace, a single arm is now used with some pyritic ores, 
 running round the hearth at intervals and stopping automati- 
 cally for a specified period in the open space above the discharge 
 
250 THE METALLURGY OF GOLD. 
 
 hopper. These furnaces are very economical, and are now largely 
 used in Western America for roasting. 
 
 H 1 
 
 Fig. 45/. The tunnel is shown at A A', and the rabbles at R R'. These 
 rabbles are moved from the central support S by the gearing G. 
 Hoppers are indicated at H H' ; BB' are the supporting girders of the 
 tunnel. (From Prof. Roberts- Austen's Introduction to Metallurgy.] 
 
 The Brown Horse-shoe Furnace. In this furnace, varieties of 
 which are called the " Elliptical " and the " Parallel," the ore- 
 hearth (L, Fig 45c) has a narrow chamber, M, on each side of it 
 and separated from it by tiles projecting downwards from the 
 arch and upwards from the hearth, so that only a narrow slot is 
 left between the two series. In the lateral chambers, M, are 
 laid rails on which the carriages supporting the stirrers are 
 moved by a cable running continuously on grooved wheels. 
 These are placed in doorways in the wall of the furnace (see 
 Fig. 45e) to keep them cool. The carriage spans the ore-hearth, 
 passing through the slots connecting L and M, and is arranged 
 to grip the cable or release it. The carriages support steel 
 stirrers which just reach the ore-hearth and stir and move on 
 the ore. In operation half the carriages are traversing the 
 furnace while half are resting in a cooling space (O, Fig. 45<?) 
 outside the furnace. When two carriages are used, the one 
 moving out of the furnace strikes the other, which is at rest in 
 the cooling space, and pushes it forward until it grips the cable 
 automatically and starts off to take up the stirring, and the hot 
 moving carriage is at the same time automatically released and 
 comes to rest. The stirrers are thus prevented from becoming 
 overheated. 
 
 There are one, two, or more fireboxes (P, Fig. 45e) according 
 to requirements, and there is generally a prolongation of the ore- 
 hearth outside the furnace walls, used as a cooling hearth, over 
 which the ore is pushed by the stirrers until it falls into a hopper 
 ready for chlorination. Fig. 45e is a plan of the furnace erected 
 in 1896 at the Golden Reward Chlorination Mill, at Dead wood, 
 Dakota (see p. 303) ; Fig. 456 is a longitudinal section through 
 one of the fireboxes ; Fig. 45c, a transverse section through the 
 
MECHANICAL FURNACES. 
 
 251 
 
252 THE METALLURGY OF GOLD. 
 
 ore-hearth and firebox, and Fig. 45e/, a transverse section through 
 a part of the furnace where there is no firebox. The scale is 
 fa in. per foot in Fig. 45e and ^ in. per foot in the others. In 
 the Golden Reward furnace the roasting hearth is 180 feet long 
 and 8 feet wide, and the cooling hearth 78 feet long. Three 
 rabbling carriages are used, following one another at intervals 
 of 90 seconds, and moving the ore forward at the rate of 22 feet 
 per hour. The ore contains from 2J to 8 per cent, of sulphur, 
 and it is stated that 65 tons of ore crushed through 30-mesh 
 sieves are roasted per day, the product containing 0'3 per cent, 
 of sulphur. The capacity is here 1 ton to every 22 sq. ft. of 
 hearth per day, and the total cost is less than 50 cents per ton 
 for roasting, cooling, and conveying the ore to the barrels,* the 
 <jost per day of twenty-four hours being made up as follows : 
 
 Fuel, 6 cords of wood, $19.50 
 
 Labour, two men, one each shift, . . . . 5.00 
 
 Power, 5H.P., 1.25 
 
 Oil, lights, repairs, &c., 1.50 
 
 $27.25 
 
 The Brown furnace is also able to treat concentrates, and 
 several are at work in various mills in Western America. 
 
 The Ropp Straight-line Furnace differs from the Brown furnace 
 in having the cable placed underneath the hearth, the slot being 
 in the hearth itself. Through the slot an arm, connected below 
 with a four-wheeled truck, extends upwards, and across the 
 upper end of this arm is the supporting bar from which the 
 stirring teeth project. The chamber beneath the hearth is 
 easily accessible for repairs, and the trucks and rabbles are 
 oooled outside the furnace while traversing the track leading 
 back to the starting point. The hearth of the standard furnace 
 is 105 feet long and 11 feet wide, and its capacity seems to be 
 about the same as the Brown furnace of equal area. Several of 
 these furnaces are in use in Western America. 
 
 2. Rotating Bed Furnaces. Rotary Pan Furnace. This is 
 used at the Bunker Hill Mill, California, to desulphurise concen- 
 trates containing much sulphur and small quantities of arsenic, 
 antimony, and lead. The stationary hearth is 7 feet wide and 
 18 feet long, and has two working doors. At the end of the 
 stationary hearth is a drop of 6 inches on to a horizontal revolv- 
 ing hearth, made of iron lined with fire-brick, 12 feet in diameter, 
 with a discharge hole in the centre. This is next the fire-place. 
 The hearth revolves by means of gear-wheels placed beneath it, 
 at the rate of one turn per minute. The charge remains on it 
 for eight hours, and is then discharged through the central 
 aperture. The capacity of the furnace is 2 tons per day, the 
 fuel required being 1 J cords of wood. In Fig. 46, a similar but 
 larger furnace is shown. 
 
 * Eng. and Mng. Journ., July 4, 1896. 
 
MECHANICAL FURNACES. 
 
 253 
 
254 
 
 THE METALLURGY OP GOLD. 
 
MECHANICAL FURNACES. 255 
 
 3. Revolving Cylindrical Furnaces. These furnaces came 
 into more general use than other mechanical furnaces before 
 the Pearce turret and Brown furnaces were invented, but were 
 never largely employed at gold chlorination mills, since their 
 capacity was too large for the modest requirements of most of 
 these establishments until recently. Revolving cylinders treat 
 from 10 to 50 tons of ore per day. 
 
 The Bruckner Cylinder. This furnace was first introduced in 
 Colorado in 1867, and is now, after alterations and improve- 
 ments, extensively used in the United States, chiefly in the 
 chloridising roasting of silver ores, although it is also suitable 
 for dead roasting. In its latest form it consists (Fig. 47) of a 
 cylinder of boiler plate iron, lined with fire bricks. The ends 
 are partly closed, leaving central apertures 2 feet in diameter. 
 There are also four receiving and discharging openings (a, a), 
 closed by hinged doors. The discharge is effected into a pit, or 
 into hot-ore cars placed underneath. The cylinder revolves on 
 two chilled-iron friction rings (6, b), resting upon four chilled- 
 iron carrying rollers (c, c). The rotation is caused by friction 
 between the rings and the rollers, which are driven by a belt 
 .and pulley. The older method of turning the cylinder by means 
 of a gear-wheel, connecting with teeth set on the cylinder, has 
 been abandoned. 
 
 The lining is one brick (2J inch) thick in the middle of the 
 cylinder, but additional layers are added near the ends until the 
 circle is contracted down to the size of the openings in the ends, 
 which are also lined ; each layer Jails short of the preceding one 
 by about 2 inches, so that the lined ends have a conical form. 
 The mortar used consists of one-third fire-clay and two-thirds 
 crushed fire-brick. In the old form six iron pipes passed through 
 the cylinder in a plane at an angle of 15 to the axis, and per- 
 forated plates uniting them formed a diaphragm which assisted 
 in stirring the ore. The circular flange surrounding the opening 
 at one end of the cylinder connects loosely with the fire-box, d\ 
 the other end connects with an opening leading to dust chambers 
 and the stack. The dust chambers (not shown in the figure) are 
 large, and must be carefully attended to, as the amount carried 
 over is usually 10 per cent, of the ore, and may rise to 25 per 
 cent, when the mineral is light. The heavier particles collected 
 in the first chamber are charged into the furnace again ; the 
 lighter dust is unsuitable for this purpose, and must be subjected 
 to special treatment. In the improved Bruckner the diaphragm 
 is discarded, as it is rapidly corroded by the sulphur, and 
 increases the quantity of material carried into the dust chambers. 
 The furnaces are now usually made 7 feet in diameter and 18 
 feet long, a full charge being from 6 to 8 tons. This large size 
 has given greater satisfaction than that formerly employed viz., 
 6 by 12 feet the full charge of which was from 3 to 4 tons of 
 ore. As a charge can be kept in the furnace as long as is 
 
256 THE METALLURGY OF GOLD. 
 
 necessary, the Bruckner is particularly useful when very base 
 ores are to be roasted, or when the ores vary greatly in com- 
 position, so that the duration of roasting is not constant. The- 
 mixing of the ore is rendered more perfect by the conical shape 
 of the ends, which causes the ore to be thrown backwards and 
 forwards, changing its position frequently and exposing new 
 surfaces to the action of the fire. 
 
 Bruckner's cylinder has been found suitable for the dead oxi- 
 dising roasting of pyritic ores, a little charcoal being sometimes 
 added towards the end of the operation to reduce the sulphate of 
 copper formed. It has the disadvantage of being too hot at 
 first, and not hot enough at the finish if the tire is kept uniform, 
 and consequently pyritic ore tends to form into balls. These 
 may contain sulphates, which are not then easy to decompose 
 except with great waste of time and fuel. The consumption of 
 fuel is also rendered greater by the shortness of the furnace, as- 
 a large amount of waste heat passes out into the dust chamber, 
 and is not utilised, a state of things exactly similar to that 
 which occurs in the reverberatory furnace with only one floor. 
 
 An important improvement in the Bruckner furnaces in use 
 at the Portland mill, Dead wood, Dakota, was made in 1895 
 by the Manager, Mr. Hickock. Between the tire-box, which is- 
 mounted on wheels, and the revolving cylinder is placed a short 
 iron cylinder, 18 inches long, rigidly connected with the fire- 
 bridge and closely fitting the throat of the revolving cylinder. 
 In the lower side of the small cylinder is an aperture, from & 
 inches to 10 inches square, closed by a sliding door. This is 
 useful in regulating the draught, without cooling down the fire. 
 By its use a sheet of cold air can be made to pass immediately 
 above the ore and below the products of combustion of the fire. 
 It is stated that a considerable saving of fuel has been effected 
 since the introduction of this simple device, and that a more 
 perfect roast can be obtained in one-third less time than was 
 formerly possible. 
 
 The Hofmann Furnace. This furnace was designed by H. O. 
 Hofmann as an improvement on the Bruckner cylinder, which 
 has the disadvantage of being much hotter at one end than at 
 the other. The result of this is that the ore near the fire is 
 exposed to a higher temperature than that near the flue, and, 
 consequently, it is finished much sooner, so that fuel is wasted 
 while finishing the ore in the colder part of the furnace, and, in the 
 case of antimony ores, if the temperature is high enough to roast 
 the ore next the flue, the portion close to the fire becomes caked 
 and suffers a considerable loss of silver by volatilisation. The 
 Hofmann furnace (Fig. 48) has a fireplace and flue at each end. 
 The flues are between the fireplaces and the cylinder, and open 
 downwards into dust chambers, C, built beneath, which are con- 
 nected with the main stack. The arrangements at each end are 
 
MECHANICAL FURNACES. 
 
 257 
 
 17 
 
258 THE METALLURGY OP GOLD. 
 
 the same, and are such that, by means of dampers, the current of 
 the air and gases can be made to pass through the furnace in 
 either direction. The fire is lighted first at one end, A, and 
 the dampers arranged so that the draught passes through the 
 revolving cylinder and down the flue, B, at the other end. After 
 a few hours the fire is lighted at the other end, in the fireplace, 
 D, and the position of the dampers is reversed. The alternate 
 heating from the two ends is regularly performed until the 
 charge is completely roasted. In this way more uniform heating 
 is obtained, both halves of the charge being raised to the required 
 temperature without any portion being overheated. It is stated 
 that the formation of balls is diminished by this system, and, 
 in particular, the furnace is found to answer well when treating 
 ores which require either a very low roasting temperature, or a 
 very high one. Thus, for example, antimonious ores, which cake 
 readily, are said to be successfully treated by the use of moderate 
 fires ; and ores containing very little sulphur, and so requiring a 
 higher temperature, can readily be made as hot as is necessary 
 by the alternate use of large fires, without using much fuel. 
 When one fire is in operation the other is allowed to go down. 
 By closing one of the large dampers near the main flue, and 
 opening the damper of the corresponding descending flue, and 
 the " plug door " in connection with it, a current of fresh air can 
 be introduced into the furnace beneath the flame from the fire, 
 thus greatly increasing the rate of oxidation of the ore. This 
 arrangement is especially useful when ores carrying a high 
 percentage of sulphur are being treated, their tendency to form 
 balls being in this way greatly diminished. 
 
 It is clear that the use of much larger cylinders than those of 
 the Bruckner furnace is rendered possible by the presence of a 
 fire at each end. Nevertheless, the evil of heating the ore 
 unequally is only palliated ; the ore in the middle of the furnace 
 is never heated so strongly as that at the two ends, whilst in the 
 case of continuous-discharge inclined cylinders, every particle of 
 ore is heated similarly. Moreover, the use of two fires, which 
 are alternately checked and urged, must cause a considerable 
 waste of fuel. The principle is obviously inferior to that of the 
 White-Howell furnace. 
 
 One incidental advantage of this furnace over all other 
 revolving cylinders lies in the absence of loss from the carrying 
 away of dust at the moment of charging-in. In other cylinders 
 the ore falls in a shower into the furnace through a strong 
 draught, by which much of the dust is carried off into the flue. 
 When charging the Hofmann furnace, however, the dampers in 
 both the descending flues are left open, so that the draught 
 passes through both fireplaces direct to the dust chambers, 
 without entering the furnace at all. 
 
 TJie White and White-Howell Roasting Furnaces. The White 
 
MECHANICAL FURNACES. 259 
 
 furnace consists of a long cast-iron revolving cylinder, lined with 
 fire-brick, and inclined towards the fire end. The cylinder 
 is bound by four cast-iron rings which rest on friction wheels, 
 serving as supports, and it is revolved by other friction wheels 
 driven by a shaft and pulleys. Crushed ore is fed into the 
 furnace continuously at the upper end, passes through it by 
 gravity, and is continuously and automatically discharged by 
 falling into a pit through an opening in the floor close to the 
 fire. The time occupied by the ore in passing through the 
 furnace depends on the angle of inclination of the cylinder 
 which can be changed, so that the time of roasting can be 
 shortened or lengthened according to the nature of the ore. 
 Ores containing a high percentage of sulphur require to be 
 subjected to heat for a longer time than those with little 
 sulphur, and the angle of inclination is reduced in such cases. 
 The average inclination is about one in twenty. The cylinder 
 is lined with fire-brick throughout, and projecting bricks raise 
 portions of the ore and drop it through the flames, thus assisting 
 the oxidation. 
 
 The advantages claimed for this furnace are that it is con- 
 tinuous in its operation, discharging its product regularly into 
 the pit at the lower end, and this roasted ore can be allowed to 
 accumulate and be withdrawn as required. The ore is submitted 
 to a gradually increasing temperature, the most favourable con- 
 ditions for dead roasting being thus obtained. The usual size 
 of this furnace is about 4| feet in diameter and 27 feet long, 
 the capacity being usually stated at from 20 to 30 tons per day. 
 When employed for the dead roasting of pyritic ores its capacity 
 is less, arid in some cases the passage through a 27-foot cylinder 
 is insufficient to eliminate the whole of the sulphur. In such 
 cases the ore is made to traverse successively two similar 
 cylinders, the combined length of which is sometimes over 60 
 feet. Great care should be exercised in keeping the ore supplied 
 to the furnaces as uniform as possible, so that when once the 
 rate of feed and the proper angle of inclination for the ore have 
 been determined, no further alterations are needed in order to 
 continue to give a perfectly roasted product. 
 
 In the White- Howell furnace (Fig. 49) only the enlarged part 
 next the fire-box is lined with fire-brick, the remainder being 
 left uiilined. Cast-iron spirally arranged shelves assist in raising 
 and showering the pulp through the flames. To both White 
 and White-Howell furnaces an auxiliary fire is often added for 
 roasting the dust which escapes from the main furnace. The 
 dust, when it has been completely roasted, is shovelled out and 
 mixed with the main bulk of the ore by hand. This arrangement 
 is decidedly inferior to that suggested by John E. Roth well,* 
 who uses a hopper-shaped dust chamber with its bottom con- 
 * Mineral Industry for 1892, p. 234. 
 
260 
 
 THE METALLURGY OF GOLD. 
 
 sisting of an inclined 
 cast-iron plate project- 
 ing about 8 inches into 
 the upper end of the 
 cylinder. The dust car- 
 ried out of the cylinder 
 settles in this chamber, 
 and, as it accumulates, 
 slides down the sides 
 and mixes with the fresh 
 ore. The ore is thus 
 kept more uniform than 
 if re-mixed by hand, and 
 some saving in labour 
 is also effected. Roth- 
 well uses a cylinder 36 
 feet long and 5 feet in 
 diameter, with an in- 
 clination of 14 inches 
 only, rotating once per 
 minute. The lining is 
 of fire-brick, six inches 
 thick. 
 
 Use of Producer 
 Gas in Boasting. The 
 use of producer gas in 
 roasting may be men- 
 tioned, as it bids fair to 
 completely replace solid 
 fuel for the purpose at 
 some future time, great 
 saving of expense being 
 thus effected. It was 
 introduced at the Hoi- 
 den Mill, Aspen, Color- 
 ado, in 1891, and has 
 now completely dis- 
 placed other fuel there 
 for both drying and 
 roasting.* The gas 
 plant consists of two 
 "Taylor" revolving-bot- 
 tom gas producers, one 
 6 feet and the other 7 
 feet in diameter. The 
 
 * W. S. Morse in Trans. 
 Am. Inst. Mng. Eng., Mon- 
 treal meeting, 1893. 
 
CHLORINATION : THE VAT PROCESS. 
 
 261 
 
 coal used is a mixture of one-third "Sunshine" nuts and two- 
 thirds " Newcastle " nuts and pea-coal, the analyses of which are 
 as follows : 
 
 
 Sunshine. 
 
 Newcastle. 
 
 Water 
 Volatile matter, . 
 Fixed carbon, . 
 Ash, . 
 
 2-8 
 36-3 
 37-1 
 23-8 
 
 17 
 37-95 
 48-6 
 11-6 
 
 
 100-0 
 
 99-85 
 
 The drying plant consists of four Stetefeldt double shelf driers, 
 and the roasting furnace is a Stetefeldt shaft furnace. The ore 
 contains 8*10 per cent, of sulphur before roasting and 0*2 per 
 cent, as sulphide afterwards. The coal consumed is 72-22 Ibs. 
 per ton of ore dried and 11744 Ibs. per ton roasted, the cost for 
 the two being 29-37 cents per ton of ore. All the above figures 
 are averages for a run of over a year viz., from November, 1891, 
 to January, 1893. The heat obtained in the Stetefeldt furnace 
 could be made sufficient to fuse the ore ; and there is, therefore, 
 no doubt that similar producers could give heat enough to 
 desulphurise ores in reverberatory furnaces. 
 
 CHAPTER XII. 
 CHLORINATION : THE VAT PROCESS. 
 
 The Vats. The vats used for impregnating the ore with 
 chlorine are usually, in California, about 7 feet in diameter, 
 and are made of staves 3 feet long and 2 inches thick, which 
 consist of the best split sugar-pine. They are coated with a 
 mixture of pitch and tar to protect them from the corrosive 
 action of the chlorine. Before being used for the first time, the 
 vats are thoroughly soaked with water to diminish their absorp- 
 tive action on the chloride of gold solution, but all wood brought 
 in contact with this solution is nevertheless invariably impreg- 
 nated with a certain amount of gold. This may be recovered 
 after the vats, &c., are worn out, by burning them and fusing the 
 ashes with suitable fluxes. The false bottom generally consists 
 of quartz pebbles, the lowest layer being of the size of hazel 
 nuts, and each successive layer consisting of finer material until, 
 at the top, a thin layer of fine sand (passing a 20-mesh sieve, 
 but retained on a 60-mesh sieve) is spread evenly over the 
 
262 
 
 THE METALLURGY OP GOLD. 
 
 surface. The thickness of the filter bed (which is not shown in 
 the figure) is usually from 6 to 12 inches. It is supported on 
 boards (A, Fig. 50), 1 inch thick, in which numerous i-inch 
 auger holes are drilled ; these boards rest on wooden strips (not 
 shown in the figure), 3 inches wide and 1 inch thick, which do not 
 reach the edge of the vat, and so keep a clear space 1 inch deep, 
 just above the true bottom of the vat, in which the solution can 
 accumulate. The solution is drawn off' by a leaden pipe fitted 
 with a stopcock, preferably of stoneware; the pipe should be 
 level with the bottom of the vat, which may with advantage be 
 made with a slight fall towards the outlet to prevent any liquid 
 being left in it. Deetken states * that fine sea-shells (consisting 
 
 of carbonate of lime) have been used instead of quartz pebbles 
 for the filter bed without any prejudicial result. Talcose rocks, 
 and particularly silicates of alumina, must not be used on 
 account of their power of absorbing the chlorine. For the same 
 reason sulphides, magnetic iron oxide, metallic iron, fragments 
 of wood or other organic matter, or, briefly, any substances 
 capable of being acted on by chlorine or of reducing the chloride 
 of gold must be carefully excluded from the filter bed. 
 
 The surface of the filter bed may be covered with boards, not 
 fitted closely together, but made into a framework by cross-pieces 
 and pierced with many auger holes. This cover is useful when 
 
 Mineral Resources of the Rocky Mount tin*, 1873, p. 342. 
 
CHLORINATION ! THE VAT PROCESS. 263 
 
 the tailings are being cleared out, otherwise, in shovelling away 
 the ore, the surface of the filter bed is partly removed also. 
 Messrs. MacArthur & Forrest's devices to avoid this are given 
 on p. 323. Filter cloths of canvas, burlap, or cocoa-nut fibre 
 matting are also frequently used above the filter bed, stretched 
 tightly over a framework of wood which accurately fits the inside 
 of the vat. The space between the canvas and the wall of the 
 vat is packed with hemp or other material closely tamped down. 
 Filter-cloths of every material, except asbestos, are soon rotted 
 and demoralised by the action of the chlorine, and their use is 
 frequently dispensed with. Wool lasts longer than cotton. 
 
 Charging-in the Ore. When the vats are ready to be 
 charged, a layer of dry ore is spread over the false bottom, and 
 time given for the water from the filter bed to be drawn up into 
 this layer by capillary attraction. If attention is not paid to 
 this point, the lowest layer of ore becomes too wet from the 
 combined effect of the water added to it before charging-in and 
 that absorbed from the false bottom. The result is that the 
 passage of the chlorine through the mass is resisted, and there 
 is a great increase in the consumption of the gas. Deetken 
 states that the whole of the usual charge of gas may be thus 
 consumed, not rising more than a few inches above the bottom. 
 The greater part of the charge is damped by sprinkling with 
 water and thorough mixing. The amount of the water added 
 varies with the nature of the ore, but the usual amount is from 
 6 to 12 per cent, for roasted ores. If it is made too wet, dry 
 ore to the required amount is mixed with it. A good rough 
 method of ascertaining when it is of the proper degree of damp- 
 ness is to compress some in the hand ; balls of ore should 
 be readily formed in this way, but should be just dry enough 
 to crumble up again. The reason for the addition of water 
 is that perfectly dry chlorine has scarcely any action on metal- 
 lic gold at ordinary temperatures, and up to a certain point 
 an increase in the amount of water present raises the rate of 
 solubility of the gold. The limit of the amount of water that 
 can be added is, however, determined by physical conditions, as 
 the mass must be of loose porous texture in order to permit the 
 gas to readily permeate through every portion of it. In order to 
 promote this porous texture and uniform dampness, the ore is 
 shovelled upon, and made to pass through, a sieve of four holes to 
 the linear inch. This sieve may be conveniently made to slide 
 on rollers on iron rails placed above the vat, and the ore, shaken 
 through it, falls into the vat in a light shower. Although left 
 undisturbed as far as possible, the charge must be levelled off 
 with a rake occasionally. 
 
 When the vat is tilled to within 6 inches of the top, the 
 surface of the ore is made concave or saucer-shaped, higher at 
 the sides than in the centre. The cover, usually of wood, is 
 
264 THE METALLURGY OF GOLD. 
 
 then lowered on to the vat by means of a chain and pulley, and 
 the rim luted with a mixture of wet clay and sand, or, more 
 usually in former times, with dough. These joints are kept 
 moist during the "gassing" process by wet rags. The gas is 
 introduced through a lead pipe, which is shown en the left-hand 
 side of Fig. 50, passing into the vat below the false bottom. 
 A small hole is left in the cover through which the displaced 
 air may escape, and the issuing gases are tested from time to 
 time by means of a rag tied to a stick and moistened with dilute 
 ammonia. As soon as chlorine is found to be coming off freely, 
 the hole is plugged, but the current of gas is not stopped until 
 after the lapse of one or two hours more, when the charge is 
 supposed to be saturated with the gas, the total time required 
 for the impregnation being usually from five to eight hours. 
 
 Generation of the Chlorine. The chlorine is generated in 
 air-tight vessels of lead fitted with a stirring apparatus passing 
 through the lid and worked from the outside. The gas neces- 
 sary for a 3-ton charge of roasted concentrates may be generated 
 in a leaden vessel of 20 inches in diameter and 12 inches deep. 
 The joints of this generator and of all pipes traversed by the gas 
 or by liquids carrying it in solution must be made by melting 
 the lead, no solder being used except pure lead. The " burning 
 together" of the joints is usually effected by an air-hydrogen or 
 coal gas blowpipe jet. The cover of the chlorine generator is 
 made gas-tight by a water joint, 2 inches deep, as are also the 
 apertures in the lid for the passage of the revolving stirrer 
 (which is usually made of hard wood) and for the delivery 
 tube. Heat is applied by placing the generator on a sand bath 
 standing on a perforated arch over a fire-place. The sand bath 
 is often replaced with advantage by a water bath, as the heat 
 required is not more than 90 P., and sudden heating causes 
 inconveniently tumultuous generation of gas. The charge for 
 3 tons of ore consists of 20 to 24 pounds of rock salt, 15 to 
 20 pounds of manganese dioxide, containing 70 per cent, of 
 available material, and 35 pounds of oil of vitriol of 66 B., 
 diluted with half its weight of water. The cover is usually 
 removed to introduce the solids and the water, but the acid is 
 added, half a gallon at a time, through a siphon. At Hey wood's 
 Works, California, the acid is contained in a lead vessel furnished 
 with a glass stopcock, from which a small continuous stream is 
 allowed to fall into the siphon. 
 
 The outlet tube is of lead, but connections in pipes are often 
 made by short pieces of indiarubber-tubing, well greased on the 
 inside. These resist the action of chlorine fairly well. The gas 
 is passed through the wash-bottle shown in Fig. 51. It is 
 usually a large glass bottle or carboy with its bottom removed, 
 supported in a lead-lined box filled with water. The gas is made 
 to pass through about half an inch of water. The use of the 
 
CHLORINATION : THE VAT PROCESS. 
 
 265 
 
 wash bottle is partly to free the chlorine from hydrochloric acid 
 or other impurities with which it is contaminated, but mainly 
 to give an indication of the rate of flow of the gas ; it is desirable 
 that this should be as uniform as possible, as otherwise leakage 
 is more difficult to prevent. As soon as the current of gas falls 
 off, fresh acid is added and the vessel stirred. The wash-bottle, 
 
 as usually constructed, 
 would be quite inade- 
 quate to free the gas 
 completely from hydro- 
 chloric acid, whilst, even 
 if it were approximately 
 eliminated, some more 
 would be speedily formed 
 by the decomposition of 
 water by the chlorine. 
 It is, therefore, fortunate 
 that this elimination is 
 
 not absolutely necessary, 
 in spite of the customary 
 
 Fig. 51. 
 Scale, 1 in. = 1 ft. 
 
 declaration to the contrary which generally appears in the descrip- 
 tions of the process. Thus it is frequently stated that hydro- 
 chloric acid will act on any sulphides left undecomposed in the 
 ore, generating sulphuretted hydrogen which would precipitate 
 the gold already dissolved in the impregnation tank. This 
 statement would not need any criticism, if it were not for the 
 fact that it has often been made, but has apparently never been 
 contradicted. No doubt matters might be arranged for the 
 above reactions to take place if sufficient care were taken, but 
 in practice they are not to be feared for the following reasons : 
 (1) If undecomposed sulphides were present in the roasted 
 ore, they would be attacked with much greater violence by 
 chlorine than by hydrochloric acid, and before any gold could 
 be dissolved, the whole of the sulphides present would be con- 
 verted into sulphates, according to reactions, the final effect of 
 which is shown in the following equation : 
 
 R 2 S -f 4C1 2 
 
 4H 2 O = R 2 SO 4 
 
 8HC1. 
 
 In some cases, no doubt, chlorides would be formed. (2) If the 
 sulphides were not oxidised by the chlorine, they would be 
 almost equally efficacious with sulphuretted hydrogen in preci- 
 pitating the gold, so that even in this case hydrochloric acid 
 would do no harm. (3) If an excess of chlorine is present, sul- 
 phuretted hydrogen could scarcely be said to be formed at all 
 under any circumstances, as, if it were formed, it would be 
 instantly decomposed by the chlorine. (4) If gold were precipi- 
 tated by sulphuretted hydrogen in the impregnation vat, it would 
 be in such a finely divided state that it would be re-dissolved in 
 
266 THE METALLURGY OF GOLD. 
 
 chlorine in a very short time, provided the gas were in excess. 
 The only disadvantage due to the presence of much hydrochloric 
 acid in the gas lies in the fact that certain metallic oxides (oxides 
 of iron, copper, &c.) are much more readily soluble in the acid 
 than in chlorine, chlorides and water being formed, and the re- 
 sulting solution will be contaminated with these chlorides, so 
 that special precautions are necessitated to prevent the bullion 
 from becoming base. 
 
 Impregnation of the Ore. The ore is allowed to remain 
 impregnated with the gas for from twenty-four to forty-eight 
 hours, the continued presence of a strong excess of gas being 
 ascertained at intervals by removing the plug from the cover 
 and applying the ammonia test. When, as is usually the case, 
 there is a large excess of gas when the impregnation is at an 
 end, it may be disposed of in one of several ways. It may be 
 dissolved by adding water before raising the cover (the usual 
 method of procedure), or it may be withdrawn by aspiration and 
 discharged outside the building, or stored in a gasometer for use 
 in a subsequent charge. If the cover is raised before getting rid 
 of part of the excess of the gas, the atmosphere of the mill is- 
 rendered unbearable for several minutes. 
 
 The time of impregnation varies according to the size of the 
 particles of gold, the fineness of the metal, and the tempera- 
 ture employed. Chlorine has a very slow action on pure gold, 
 the rate increasing gradually with the temperature up to 10l>. 
 In order to obtain some data for the rate of solution of gold by 
 chlorine at different temperatures below 100, the following 
 experiments were conducted by the author in the laboratory of 
 the Royal Mint. Comparisons were made at the same time 
 between chlorine, bromine, and cyanide of potassium, in order 
 to determine their relative efficiency. The gold used consisted 
 of " cornets " weighing about a half gramme each. These cornets 
 offered a large surface to attack, being porous in texture, and con- 
 sisted of gold 9y9'3 parts, and silver 0'7 part. They had all 
 been prepared together by cupellation, parting, and annealing, 
 so that their physical state must have been very similar. The 
 results are given in the table on the next page. 
 
 The results obtained simultaneously by the use of potassium 
 cyanide are given at p. 358. The amount of liquid used was in all 
 cases 30 c.c. These results tend to show that both chlorine and 
 bromine dissolve gold more rapidly at 50 to 60 C. than at ordi- 
 nary temperatures, that bromine is more rapid in its action than 
 chlorine, and that both are considerably more rapid than potas- 
 sium cyanide, particularly at the higher temperatures employed. 
 In all cases an excess of the reagent was present at the end of 
 the operation, but the strength of the solutions obviously fell off 
 during the experiments. The results of these experiments point 
 to the desirability of further investigations on the subject. 
 
CHLORINATION : THE VAT PROCESS. 
 
 267 
 
 Solvent. 
 
 Strength. 
 
 Time of Action. 
 
 Temperature. 
 
 Amount of Gold 
 Dissolved per 
 1,000 Treated. 
 
 Chlorine, . 
 
 Saturated solu- 
 
 1 hour. 
 
 15 C. 
 
 2-3 
 
 
 tion in water, 
 
 
 
 
 Bromine, . 
 
 Pure, 
 
 
 
 ,, 
 
 169-0 
 
 at 
 
 Mixed with 20 
 
 
 
 
 
 83-0 
 
 
 times its weight 
 
 
 
 j 
 
 
 of water, 
 
 
 
 
 n 
 
 > 
 
 > 
 
 50 C. 
 
 157-1 
 
 Chlorine, 
 
 Saturated solu- 
 
 5 hours. 
 
 15C. 
 
 57-6 
 
 
 tion in water, 
 
 
 
 
 Bromine, . 
 
 Aqueous solu- 
 tion containing 
 
 is 
 
 
 
 58-1 
 
 
 0'2 per cent. 
 
 
 
 
 , 
 
 bromine, 
 
 
 
 
 5 J 
 
 > 
 
 > > 
 
 50 C. 
 
 101-0 
 
 Chlorine, . 
 
 Saturated solu- 
 
 Ij hours. 
 
 60 C. 
 
 44-9 
 
 
 tion in water, 
 
 
 
 
 Bromine, 
 
 1 per cent, aque- 
 
 
 
 5> 
 
 64-6 
 
 
 ous solution, 
 
 
 
 
 The fact that the length of time occupied depends on the state 
 of division of the gold needs no demonstration. The influence 
 of the fineness and chemical composition of the particles of gold 
 alloy present in the ore was pointed out by Deetken and Kiistel. 
 Fine gold is acted on more slowly than that of low standard. 
 The alloys containing base metals (copper, <kc.) are dissolved 
 very rapidly, and small quantities even of silver appear to 
 increase the rate of solution, but if the percentage of silver is 
 increased beyond a certain amount, an insoluble coating of 
 chloride of silver is formed over the granule, and further action 
 is chejcked or completely stopped. 
 
 Reactions in the Impregnation Vat. The amount of 
 chlorine to be used depends mainly on the substances present, 
 other than gold, by which chlorine is absorbed. A small 
 amount is invariably converted into hydrochloric acid by the 
 decomposition of the water present, the extent to which this 
 reaction proceeds being increased by light and heat. If any 
 sulphides are present they are oxidised by the chlorine in 
 presence of water, sulphates and hydrochloric acid being 
 formed, as follows : 
 
 4C1 2 
 
 CuS 
 
 4H 2 O = 8HC1 + 20 2 
 
 20 2 = CuS0 4 
 
 These equations do not represent the whole of the reactions 
 that take place. Some oxygen appears to be liberated, but the 
 subject needs investigation. If considerable quantities of hydro- 
 chloric acid are thus formed, certain sulphates are converted 
 
"268 THE METALLURGY OP GOLD. 
 
 into chlorides and sulphuric acid is set free, the reactions being 
 influenced by the mass of the reagents present. Protosulphates 
 or any other protosalts present are converted almost instantan- 
 eously to persalts by the chlorine, as follows : 
 
 6FeS0 4 + 3C1 2 = 2Fe 2 (S0 4 ) 3 + Fe 2 Cl 6 
 
 It is obvious from these reactions that great waste of chlorine 
 in the impregnation vat is caused by imperfect roasting, 1 per 
 cent, of unoxidised sulphur present in pyrites converting 8 '9 per 
 cent, of chlorine (or about 200 Ibs. per ton of ore) into hydro- 
 chloric acid. This simple calculation is sufficient to show the 
 impolicy of neglecting to roast the ore dead and then trying to 
 retrieve the error by increasing the allowance of chlorine. 
 Moreover, it demonstrates the uselessness of eliminating the 
 hydrochloric acid from the chlorine before mixing it with the 
 ore, and expecting in that way to prevent the ill effects pro- 
 duced by sulphides. The fact that many sulphides are almost 
 instantly oxidised by even very dilute solutions of chlorine has 
 been proved by a series of laboratory experiments by the author. 
 These experiments would have been quite unnecessary if it were 
 not that some chemists engaged in chlorination still appear to 
 doubt the rapidity of the reaction. The oxidation of protosalts 
 is almost as rapid, although the percentage waste of chlorine is 
 ^considerably less ; thus 1 per cent, of sulphur present in the ore 
 as ferrous sulphate will convert 1*1 per cent, of chlorine (or 24'6 
 Ibs. per ton of ore) into hydrochloric acid. 
 
 Sulphate of copper (CuS0 4 ) does not appear to be acted on by 
 ohlorine, but, nevertheless, whenever it is present in a roasted 
 ore, chlorination seems to be rendered impracticable. This is 
 possibly due to the fact that some sulphate of iron accompanies 
 it. Whenever sulphates of these metals are left in the roasted 
 ore by accident or design it is necessary to remove them by a 
 preliminary leaching with water before the chlorine is introduced. 
 Of course if the ore is chlorinated in tubs by gas, it must be 
 partially dried and sieved back into the tub before impregnation 
 can be attempted.* 
 
 Organic matter is also oxidised by chlorine, although much 
 more slowly. At ordinary temperatures, pitch and tar are 
 .almost unaifected, and the fibres of matting, canvas, &c., are 
 acted on very gradually. Pieces of decaying wood or dried 
 leaves must not be introduced with the water into the leaching 
 vat, and if surface water is used it should always be carefully 
 strained before being run in. A rough analysis of the water 
 employed will often be serviceable, as it is frequently strongly 
 alkaline in dry countries, and may be softened with advantage. 
 
 The absorption of chlorine by metallic oxides is the most 
 frequent cause of waste, and, in the vat process, there are 
 usually no efforts made to prevent this. Well roasted sesqui- 
 
 * Ore roasted " dead " or " sweet " usually contains from 0'15 to 0'4 per 
 cent, of sulphur, but the lower the amount of sulphur is reduced the better 
 is the extraction of gold and the smaller the amount of chlorine required. 
 
CHLORINATION : THE VAT PROCESS. 269- 
 
 oxide of iron (Fe 2 O 3 ) is scarcely attacked by chlorine, especially 
 if the temperature attained in the furnace has been high. If 
 any magnetic oxide (Fe 3 O 4 ), however, has been formed from 
 over heating, or has been originally present in the ore, the 
 absorption of chlorine is considerably greater, ferric chloride 
 being formed and dissolved. Protoxide of iron is instantly 
 converted into a mixture of chloride and sesquioxide of iron. 
 Hydrochloric acid acts more rapidly than chlorine on all these 
 oxides, but is nevertheless very slow in dissolving the well 
 roasted sesquioxide. Metallic iron, which is sometimes acci- 
 dentally introduced, is dissolved at once by both chlorine and 
 HC1. The oxides of copper and zinc are quickly dissolved by 
 chlorine, and still more readily by HC1. Lime and magnesia 
 also readily absorb chlorine, forming hypochlorites, chlorates 
 and chlorides, but hypochlorites are decomposed by any acid 
 which may be present. 
 
 If any appreciable quantity of oxides capable of absorbing 
 chlorine are present, it is cheaper to dissolve them by adding 
 dilute sulphuric acid to the ore, and then, if possible, to leach out 
 the soluble sulphates formed, before subjecting the ore to the- 
 action of the gas. 
 
 Amount of Chlorine required. The amount of chlorine 
 required varies greatly, both with the nature of the ore and 
 the manner in which it is roasted. In order to roast pyrites- 
 dead, a long time in the furnace terminating at a high tempera- 
 ture is necessary, and the addition of salt may be desirable in 
 order to chloridise oxides which would otherwise absorb the 
 more expensive chlorine in the impregnation vat. These con- 
 ditions in the furnace, however, may cause enormous losses by 
 volatilisation, the endeavour to save a few pounds of chlorine 
 in the vat causing the loss of 30 or 40 per cent, of the gold in 
 the furnace. In ores where the percentage of copper, &c., is not 
 large, and where, in consequence, salt need not be used in the 
 furnace, the roasting may be finished at a high temperature 
 without any disadvantage, and the consumption of chlorine may 
 be thus reduced to a very low point. Thus certain ores from 
 Dakota, containing only 1 or 2 per cent, of sulphur, and con- 
 sisting chiefly of silica, were chlorinated by the author in 
 revolving barrels, using only 3J pounds of chlorine per 2,000 
 pounds of ore. Even in this case, there was a strong excess of 
 chlorine in the ore after the solution was complete, and the 
 amount used could probably have been still further reduced 
 without lowering the percentage extraction of gold. This ore 
 contained 10 dwts. of gold to the ton, and over 80 per cent, was 
 extracted. This was an extreme case, and it is seldom that so 
 little chlorine is sufficient. Mr. Butters states * that at his mill 
 at Kennel, California, where all descriptions of concentrates and 
 * Eng. and Mng. Journ., Dec. 20, 1890. 
 
270 THE METALLURGY OP GOLD. 
 
 pyrites were treated by the vat process, the average consumption 
 of chlorine was 12 pounds per 2,000 pounds of ore. At Deloro, 
 Canada, the amount used in the barrel process was from 12 to 
 18 pounds per 2,000 pounds of ore, but at the Haile Mine, 
 South Carolina, only about 3 or 4 pounds per ton. In giving 
 the amounts of chlorine which are used both in the vat and 
 barrel processes, side by side, the intention is to show that the 
 quantity absorbed depends on the nature of the material treated 
 and not on the process used and the amount of water present, 
 which are immaterial within certain limits as far as this point is 
 concerned. 
 
 Leaching the Charge. When it is judged that the impreg- 
 nation has lasted long enough for all the gold to be dissolved, 
 the excess of chlorine gas is removed, the lid is taken off, and 
 water is added to the charge to wash out the soluble chloride of 
 gold. The water may be added from below, and is then either 
 allowed to overflow at the top, or is subsequently drawn off again 
 at the bottom, the inflow being suspended. It is tar more usual, 
 however, to pour on water at the top, and let it flow out at the 
 bottom. 
 
 The water must be added carefully, as otherwise the ore may 
 pack unevenly, and channels may be formed through the mass, 
 and the leaching thus rendered imperfect. Water is usually run 
 from a tap on to a layer of gunny-sacking placed over the ore, 
 by which it is distributed in a fairly even manner. It has been 
 proposed to attach a coil of lead pipes, pierced with small holes, 
 underneath the cover, and so sprinkle the water all over the 
 ore in fine jets. In any case, water is added until it forms a 
 layer 2 or 3 inches deep above the surface of the ore, and it is 
 then allowed to stand until gas bubbles have ceased to rise 
 through it, which happens in about half an hour. The stopcock 
 below the false bottom is then opened, and the yellow- or blue- 
 coloured solution (coloured by salts of iron, gold and copper), 
 which should have a strong odour of chlorine, is run slowly 
 through a filter, consisting of a canvas bag, into a small barrel 
 about 18 inches in diameter and 2 feet deep, the overflow of 
 which passes, by means of a launder or by rubber hose, to the 
 precipitating tanks. Some of the ore and sand, escaping with 
 the solution, is deposited in the canvas bag and barrel, but, if 
 much slimes are present in the charge, either the canvas bag 
 becomes clogged or the solution still remains turbid when it 
 enters the precipitating vats. Water is supplied on the top of 
 the ore as fast as it is drained away below, care being taken not 
 to let the surface of the ore emerge from the liquid. The leach- 
 ing is continued as long as any trace of gold can be detected in 
 the issuing liquid by protosulphate of iron. As has been ex- 
 plained elsewhere (p. 26), a reaction is visible in clear solutions 
 so long as the liquid contains more than two-thirds of a penny- 
 
CHLORINATIOtf : THE VAT PROCESS. 271 
 
 weight of gold per ton of water, or one part of gold in one 
 million of water. The strongly-coloured turbid solutions usually 
 encountered in mills, however, are not capable of yielding distinct 
 reactions, unless they contain much larger amounts of gold than 
 this. In particular, when large quantities of copper salts are 
 present in the solution, their strong bluish tints mask the 
 slight discoloration due to a precipitate of a small quantity of 
 metallic gold, and, moreover, they appear to interfere with the 
 precipitation itself, in some cases at least preventing it from 
 taking place.* It is always advisable to filter the solution by 
 asbestos or filter paper before testing the liquid. Filter paper 
 may be used, since it is very slow in precipitating gold from 
 dilute solutions, even if they are neutral, and this action is com- 
 pletely stopped if free chlorine is present. It is better to test 
 the clear liquid with stannous chloride under the conditions given 
 at p. 27, since the presence of salts of copper does not seem to 
 interfere with the reaction in this case, and, moreover, the 
 amount of gold present can be determined in very dilute solutions 
 with much greater accuracy than if ferrous sulphate is used. 
 
 Since a few minutes longer time occupied in leaching is of 
 small moment, while the extraction of a few more grains of 
 soluble gold from a ton of ore may be of the utmost importance 
 in the long run, it is advisable to continue to leach at any rate 
 until the water contains less than 1 part of gold in 5,000,000 
 (about 3 grains per ton), a point which can easily be determined 
 by means of stannous chloride properly manipulated. The last 
 charges of wash-water should not be mixed with the strong 
 solution, but stored in other vats and used again for the first 
 washings of other charges. In this way the amount of wash- 
 water does not become excessive, although the tailings are 
 cleaned more effectually than is usually the case. Re-precipi- 
 tation of dissolved gold in storage vats or impregnation vats is 
 not to be feared, so long as there is an excess of chlorine present 
 in the liquid, and this can easily be ensured by adding a small 
 quantity to any solutions not smelling strongly of the gas. 
 
 The amount of water used varies according to the richness of 
 the ore and the method of leaching adopted. It is usually about 
 2 tons of water to 1 ton of ore, but in most cases part of the 
 water is used again in the next charge. 
 
 Precipitation of the Gold. The precipitating vat is of the 
 same materials as the leaching tubs, and may be from 5 to 7 feet 
 in diameter and 3 feet deep. There is no false bottom, and the 
 vat is often made wider at the bottom than at the top to prevent 
 any adherence of the gold to the sides. The wood is protected 
 by a coating of pitch or paraffin-paint, or is left without paint of 
 any kind. The vat receives a smooth finish inside to facilitate 
 perfect cleaning, and is set perfectly level to avoid loss of gold 
 * Letter from Mr. Butters, Eng. and Mng. Journ., Dec. 20, 1890. 
 
272 THE METALLURGY OP GOLD. 
 
 while the waste liquor is being drawn off. The precipitating 
 solution of protosulphate of iron (the reagent which has been 
 chiefly used in practice in the vat process) is usually introduced 
 into the precipitating vat at the beginning of the filtering opera- 
 tion. Care must be taken not to introduce a wasteful quantity, 
 as it is better to make up for any deficiency when the gold- 
 solution has all been run in. This operation is conducted in 
 such a way as to impart a circular motion to the contents of 
 the vat, so that the solutions are mixed without hand-stirring, 
 but the latter is often resorted to in Addition, in order to make 
 the precipitate settle better ; flat wooden staves with round 
 handles are used for the purpose. The contents of the vat may 
 be tested with solutions of gold chloride and of ferrous sulphate 
 to determine whether the precipitation is complete, and the pre- 
 cipitant present in excess. 
 
 If lead or lime is present, dissolved in the solution, it will be 
 precipitated as an insoluble sulphate on the addition of the ferrous 
 sulphate, and thus render the gold-precipitate impure and less 
 easy to treat. The amount of lead in the solution is usually 
 small, unless hot water has been used for leaching, and most of 
 the lead chloride is, in any case, separated by the canvas filter. 
 The usual method of removing the lime is to add sulphuric acid 
 to the gold solution, and to let it stand for a few hours, when 
 calcic sulphate crystallises out, forming a crust on the sides and 
 bottom of the vat. The liquid is then drawn off and transferred 
 to another vat for the precipitation of the gold. Instead of this 
 method, Nelson A. Ferry, E.M., recommends* the addition of 
 molasses to the leach, before the addition of the sulphate of iron.. 
 He dissolves 1 gallon of molasses in 30 or 40 gallons of water, 
 and keeps it for use, determining the quantity to be added 
 by a laboratory experiment. He states that in this way the 
 precipitation of the lime is prevented, but a large excess of the 
 ferrous sulphate should be avoided, and the liquid kept slightly 
 acid. The gold then sometimes comes down fiocculent at first, 
 but soon changes to its normal condition. 
 
 The ferrous sulphate is usually prepared on the mill by dis- 
 solving iron in sulphuric acid. When precipitation is complete 
 the liquid is allowed to remain at rest for some time, in order to> 
 allow the gold to settle to the bottom. The old practice was to 
 leave it " overnight," but the length of time allowed has of late 
 years been greatly extended. Thus Mr. Chas. Butters f states 
 that forty-eight hours is usually sufficient, but that sixty hours 
 is better, and the determination of the extent to which the 
 settling has progressed may be made by tapping the solution at 
 various heights and filtering the liquid thus obtained. When 
 a quart of liquid, drawn from a point 2 inches above the bottom. 
 
 * Eng. and Mng. Journ., Nov. 28, 1885. 
 t Eng. and Mng. Journ., Dec. 20, 1S90. 
 
CHLORINATION : THE VAT PROCESS. 273 
 
 of the vat, gives only a slight dark stain to a No. 7 Swedish 
 filter paper, on being passed through it, the settling may be 
 regarded as complete. Mr. C. H. Aaron quotes instances* 
 where, after forty-eight hours settling, as much gold remained 
 in suspension in the liquid which was drawn off as was equivalent 
 to 50 cents per ton of the ore treated. The conclusion may be 
 drawn from these statements that a certain amount of gold is 
 inevitably lost by being carried away in suspension, but with 
 care and patience the loss may be reduced to a low percentage, 
 and even at the present day, in spite of the introduction of 
 many other precipitants, ferrous sulphate is probably as widely 
 used as it has been at any time in the past. 
 
 When the waste liquid has been drawn off by a floating siphon, 
 more ferrous sulphate and fresh solutions from the leaching vat 
 are poured into the vat, and the process repeated until enough 
 gold has accumulated at the bottom to warrant a clean-up. This 
 may take place at intervals of from a fortnight to three months. 
 The clear liquid is drawn off as closely as possible, and the slime 
 scooped out and filtered through paper, or, by means of a press, 
 through canvas. Finally, the vat is thoroughly cleaned out by 
 rinsing it with water which is run off through a plug-hole, level 
 with the bottom, into a wash-tub. The gold precipitate is then 
 dried carefully and fused in graphite pots, with salt, sand, nitre, 
 borax, &c., as fluxes, according to the requirements of the case. 
 If the precipitate contains any considerable amount of impuri- 
 ties (such as oxides and basic salts of iron), which is usually the 
 case, it may be treated with hydrochloric acid before fusion. The 
 bullion produced varies from 920 to 990 fine, the alloying metals 
 consisting chiefly of iron and lead. 
 
 Cost of Working. The cost of treating concentrates or ores 
 by the Plattner process depends chiefly on the cost of roasting. 
 In 1867, the total cost in California was stated by Kiistel to be 
 $14.55 per ton, but in 1872 it had been reduced to $11, the 
 expense of roasting being in each case about two-thirds of the 
 whole. At the works of the Plymouth Consolidated Mining 
 Company, California, in 1886, f the cost of treating 100 tons per 
 month was $9.40 per ton (the roasting accounting for $4.60 per 
 ton, or nearly one half), and at the Providence Mine, in the same 
 State, it was about $6.30, without including the expenses of 
 general supervision, interest on first cost, and depreciation of 
 plant. 
 
 It was estimated in 1888 J that, generally, throughout Cali- 
 fornia, a plant, capable of treating 6 tons of concentrates daily, 
 cost from $6,000 to $7,000 for its erection, while the cost of 
 extraction was about $10 per ton, and the proportion of the gold 
 
 * Cal. State Mineralogist, 1888, p. 836. 
 t I ram. Am. Inst. Mng. Eng., 1886. 
 t Eighth Report, Cal. State Min., 1888. 
 
 UNIVERSITY 
 
'274: THE METALLURGY OF GOLD. 
 
 extracted was from 90 to 92 per cent. It has, however, been more 
 recently stated by G. F. Deetken that the cost of vat chlorina- 
 tion under favourable conditions need not exceed from $3 to $4 
 per ton, but he gives no details or facts in support of his state- 
 ment, and it seems probable that the conditions required are too 
 favourable to be expected. 
 
 CHAPTER XIII. 
 CHLORINATION : THE BARREL PROCESS. 
 
 THE use of revolving barrels for chlorinating ores was perhaps 
 suggested by the old Freiberg method of barrel amalgamation. 
 It has already been mentioned, p. 224, that Dr. Duflos used a 
 revolving barrel in some of his experiments at Breslau, in 1848, 
 and obtained results almost identical with those given by the 
 vat percolation method. He, therefore, preferred the latter as 
 being cheaper. The next mention of revolving barrels seems 
 to have been contained in a patent taken out by Mr. De Lacy 
 in Victoria, in 1864. The process thus patented appears to 
 have been tried there, probably on a small scale, but certainly 
 never passed into general use and was soon forgotten. It need 
 not be described here.* In 1877, Dr. Howell Hears, of Phila- 
 delphia, patented a process which had some points of resemblance 
 to that described by De Lacy. This process was gradually 
 improved in practice, after having been adopted by several 
 mines in the United States, and in particular the improvements, 
 introduced by Mr. A. Thies, of South Carolina, which were 
 everywhere adopted, caused the name of the Thies process to be 
 applied to the amended method of procedure. In 1887 the 
 barrel process was re-introduced with some modifications into 
 Australia by Prof. J. Cosmo Newbery and Mr. C. J. T. Yautin, 
 who applied it to the ore of the Mount Morgan Mine, where it 
 was worked for some years with conspicuous success. Messrs. 
 Newbery and Yautin obtained patents in many countries for 
 thei' improvements, and the method of treatment of ores by 
 barrel chlorination is now known in Australia as the Newbery- 
 Yautin process. Considerable improvements have been effected 
 in the practice during the last few years, these improvements 
 having been largely due to the attention which was called to 
 the subject by the efforts of Mr. Claude Yautin. 
 
 * For description vide O'Driscoll on Gold Ore*, where is to be found a 
 curious history of the patent literature of chlorination. This work must 
 be read with caution, as it appears to attach too much importance to one 
 of the later patent processes. 
 
CHLORINATION : THE BARREL PROCESS. 275 
 
 The Hears Process. In this process the roasted ore is 
 charged into cylindrical barrels, together with enough water to 
 make an easily flowing pulp ; chlorine is then forced in under 
 pressure, and the barrel, which must be air-tight, is revolved 
 until the gold has been dissolved. The barrel is then opened 
 and discharged by gravity into a leaching vat below, where the 
 soluble gold is washed out and precipitated by any of the known 
 methods. Dr. Hears became aware by an accident of the increase 
 caused by the use of compressed chlorine in the rapidity with 
 which gold is dissolved by the gas. He was experimenting on 
 the action of chlorine on roasted pyrites, when the discharge 
 pipe of his apparatus becoming stopped up, the gas accumulated 
 in the vessel until it was burst by the pressure. He then found 
 that the ore was perfectly chlorinated, although it had been 
 subjected to the action of the gas for a few minutes only. 
 
 The Hears process as practised on a large scale may be de- 
 scribed as follows : Cylindrical iron barrels lined with sheet 
 lead and mounted on hollow trunnions are used, most of the 
 general arrangements being similar to those shown in Fig. 52, 
 p. 277, in which the Thies barrel is depicted. An aperture 
 of from 6 inches to 1 foot in diameter, situated in the middle of 
 the length of the barrel, serves for charging-in and discharging 
 the ore ; the aperture is covered by an iron plate which is kept 
 tightly screwed down while the chlorination is in progress. 
 The ore and water are introduced through this aperture, and 
 the plate then screwed on. The chlorine is generated in vessels 
 similar to those used in the vat process, and stored in a large 
 gas-holder made of lead. From this gas-holder the chlorine is 
 pumped direct into the barrel through a leaden pipe passing 
 through the hollow trunnion, and continued into the barrel as 
 the " gooseneck," which is a pipe passing vertically upwards, 
 close to the end of the barrel, starting from the centre, and 
 terminating in a hook-shaped curve near the top. The gooseneck 
 remains stationary and vertical when the barrel is revolved ; 
 it is made of iron so as to be strong enough to withstand the 
 weight of ore which presses against it on revolving the barrel, 
 and is lined inside and outside with lead to protect the iron 
 from the action of the chlorine. The object of the gooseneck is 
 to prevent the ore and water from flowing into and clogging 
 the pipe intended for the introduction of the chlorine. 
 
 When it has been charged, the barrel is exhausted of air by a 
 steam-jet exhaust pump, and chlorine is then pumped in until a 
 pressure of 40 or 50 pounds to the square inch is attained. 
 Another method of introducing the chlorine is to fill a large 
 receiver with the gas at high pressure, and then to connect it 
 with the pipe leading to the gooseneck, when the gas rushes 
 into the barrel. The effect of the high pressure in the barrel, 
 however it has been obtained, is to make a strong solution of 
 
276 THE METALLURGY OF GOLD. 
 
 chlorine in contact with the ore. The previous exhaustion of 
 the air is of little importance, as the quantity present in the 
 barrel is small, at least three-quarters of the space inside the 
 barrel being usually occupied by the charge of ore and water, 
 while it is doubtful if such air as is present does any harm. 
 Messrs. Newbery and Vautin, indeed, subsequently claimed that 
 the presence of a large quantity of air is actually beneficial. * 
 The barrel is then revolved at the rate of six to ten revolutions 
 per minute by a belt and pulley (the latter being fixed on a 
 sontinuation of the trunnion), or by means of cog-wheels, or best 
 of all by a friction clutch. The ore is kept constantly stirred 
 and tumbled about by tho revolution of the barrel, the advan- 
 tages gained from this course being as follows : 
 
 1. Every particle of ore is exposed equally to the action of 
 the chlorine. 
 
 2. Native gold is usually alloyed with more or less silver. If 
 this metal is present in large excess, the dissolution of the gold 
 is stopped after a certain point has been reached, by the formation 
 over it of an insoluble coating, usually supposed to consist of 
 chloride of silver. Such alloys cannot be dissolved by aqua regia 
 unless this silver chloride is removed at intervals by friction or 
 by solution in ammonia. It is doubtful whether chloride of 
 silver is formed to any extent in the cold by free chlorine, 
 although no doubt the nascent chlorine produced in aqua regia is 
 more potent, but, whether the coating formed consists of silver 
 or of chloride of silver, it must in either case be in a powdery 
 condition, and so is readily removed by the mechanical attrition 
 of the particles of ore against one another, caused by the rotation 
 of the barrel. A clean surface of gold is thus continually offered 
 to the action of the chlorine. Coarse particles of gold may also 
 be covered and protected by a coating of undissolved chloride of 
 gold in the vat process, where so little water is present, but, in 
 the barrel, the larger amount of water present instantly dissolves 
 all such soluble salts. 
 
 The pressure of gas inside the barrel is tested from time ta 
 time by opening a valve to which a pressure-gauge is connected. 
 A pop-valve may also be used for the same purpose. When 
 chlorination is believed to be complete, the excess of gas is 
 drawn off by the exhaust apparatus, and stored up or discharged 
 outside the building. The barrel is filled with water, and the con- 
 tents then discharged into a filtering vat, where the solution is 
 separated from the ore and precipitated as usual. 
 
 This description applies in great part to the other barrel chlori- 
 nation processes which differ little. The process was formerly in 
 use at the Phoenix and Haile Mines, in Carolina, and at Bunker 
 Hill Mine, California. The chief disadvantages in the process 
 were the rapidity with which the gooseneck wore out, and the 
 great strength and corresponding cost of the barrels rendered 
 
CHLORINATION : THE BARREL PROCESS. 
 
 277 
 
 necessary by the high pressure used. It was also found to be 
 difficult to keep the stuffing-boxes in the hollow trunnions from 
 leaking, and the cost of repairs was excessive. Modifications 
 were made by Mr. Adolph Thies, at Bunker Hill, with a view 
 to remove these objections to the process, and his modifications 
 have been everywhere adopted. The amended method of pro- 
 cedure may be called by the name of its inventor. 
 
 The Adolph. Thies Process. Mr. Adolph Thies was in 
 
 Fig. 52. 
 
 bcale. 1 in. = 2 ft. 6 ins 
 
278 
 
 THE METALLURGY OP GOLD. 
 
 charge of the Hears process at the Bunker Hill Mine for nearly 
 four years before he made the improvements, in the year 1881, 
 which have been associated with his name. He found that the 
 hollow trunnions, the gooseneck, and the pressure pumps could 
 all be dispensed with, and the chlorine gas generated inside the 
 barrel itself to any required extent by the use of bleaching 
 powder and sulphuric acid. This method had been mentioned 
 by Mears in one of his earlier patents, but had been abandoned 
 in favour of pumping the gas into the barrel. Thies proved it 
 to be cheaper and better, as all joints liable to leakage are- 
 
 Fig. 53. 
 
 dispensed with, and the machinery is simplified and rendered 
 much cheaper. Although there is no doubt that chlorine 
 under very high pressure acts more efficiently and rapidly than 
 at the ordinary pressure, still Thies showed that very good 
 results could be obtained at a reduced cost by using a moderate 
 pressure of chlorine of only a few pounds per square inch. The 
 barrel used is similar to that already described as suitable 
 to the Mears process, except that the trunnions are solid. It 
 is shown in Fig. 52 in transverse and longitudinal section,, 
 while a complete Thies plant is shown in Figs. 53 and 53a. 
 
CHLORINATION : THE BARREL PROCESS. 
 
 279 
 
280 
 
 THE METALLURGY OP GOLD. 
 
 An enlargement of part of the end of the most modern type of 
 chlorination barrel is shown in A, Fig. 54 ; in B, the construc- 
 tion of an older form of barrel, which is stated to have been in 
 use as late as the year 1891 at the Golden Reward Mill, is 
 shown for purposes of comparison. The shaded parts of A and 
 B are made of iron (the barrel-head and the flange being cast- 
 iron, and the cylinder, boiler-iron) ; the lead lining is not shaded. 
 In A the lead lining is burnt-on to the flanged cylinder and to 
 the head respectively, and the head then firmly bolted to the 
 flange. The lead-joint is made tight by the blows of a 
 hammer on a blunt chisel directed on the surface of the lead, 
 in the direction shown by the arrow. When repairs are needed 
 inside the barrel, the cast-iron head is removed and the work 
 thus easily done. This barrel never suffers by leakage, and the 
 cost for repairs is usually nominal. In the other barrel the 
 lead lining of the end is burnt- n to the lining of the cylinder at 
 D, leaving a hollow space, E. In consequence of the existence 
 
 B 
 
 Fig. 54. 
 
 of this space, the lead-joint is often broken when the pressure 
 in the barrel is great, and leakage then occurs; moreover, 
 repairs must be effected by a workman crouching inside the 
 barrel. 
 
 In charging-in, the water is added first, being run in by a hose 
 through the manhole, until it reaches a certain mark on the 
 inside of the barrel. The amount of water to be used varies 
 with the nature of the ore roasted pyrites absorbing much 
 more water than siliceous ores ; the quantity required is usually 
 from 40 to 60 per cent, of the weight of the ore i.e., from 80 to 
 120 gallons of water per 2,000 Ibs. of ore. This is enough to 
 make an easily flowing pulp. It is much more than that 
 employed in the Plattner process, where it is necessary to keep 
 the mass porous, so as to enable the gas to pass through it. In 
 the barrel, however, this reason for limiting the amount of water 
 does not exist, and the ore is really treated by a solution of 
 chlorine water. If too small a quantity of water is used, so that 
 
CHLORINATION : THE BARREL PROCESS. 281 
 
 the pulp is not quite free-flowing, lumps are formed which are 
 not broken by the revolution of the barrel, and these lumps are 
 not perfectly chlorinated. 
 
 The ore is let fall into the barrel down a shoot from an over- 
 head hopper, which may conveniently be made to contain the 
 exact quantity of ore required for a barrel charge. The ore 
 should be perfectly dry (not cooled after roasting by too much 
 " wetting down "), as otherwise it sticks in the hopper, instead 
 of sliding freely down the shoot. The latter may be conveni- 
 ently made of canvas, so that it can be looped-up out of the way 
 when not in use. The chemicals (bleaching powder and sul- 
 phuric acid) may be added in one of two ways. Either the lime 
 is thrown into the water before the ore is added, and the acid 
 subsequently poured upon the upper dry surface of the latter, 
 just before the manhole is closed ; or the acid is poured into the 
 water, through which it sinks to the bottom without mixing, the 
 ore let fall next, and the lime added on the top. This latter 
 system is perhaps the better, as, if it is used, the generation of 
 chlorine is never begun before the barrel commences to revolve, 
 provided ordinary care is taken. 
 
 The amounts of bleaching powder and sulphuric acid to be 
 added depend on the nature of the ore, and the care with which 
 it has been roasted. If there is nothing in the ore which is open 
 to attack except gold, the amount of chlorine actually absorbed 
 in the course of the chemical reactions is extremely small, 1 oz. 
 of gold requiring only 0*54 oz. of chlorine to convert it into the 
 soluble trichloride. 
 
 The bleaching powder used should be of the finest quality 
 obtainable, to prevent the introduction of an unnecessarily large 
 amount of sulphate of lime into the charge. Bleaching powder 
 usually has assigned to it the formula Ca(OCl)Cl, or Ca(OCl) 2 
 + Ca01 2 , and may be, as the latter formula would indicate, a 
 mixture of hypochlorite and chloride of lime. The reaction 
 with acid is usually expressed thus 
 
 CaCl 2 + Ca(OCl) 2 + 2H 2 S0 4 = 2CaS0 4 + 2H 2 + 2C1 2 
 
 This equation certainly does not accurately represent what 
 happens, as much less chlorine is liberated than is indicated by 
 it.- The amount of "available" chlorine (i.e., that which is 
 liberated by the action of acids) contained in commercial bleach- 
 ing powder varies from 20 to 35 per cent. Bleaching powder is 
 gradually decomposed by the carbonic anhydride in atmospheric 
 air (chlorine being liberated), and must consequently be kept 
 separated from it as completely as possible. Even if preserved 
 in hermetically sealed vessels, however, it suffers a slow change, 
 by which some chlorate of lime is iormed and the amount of 
 available chlorine reduced. Under these circumstances, it is 
 necessary to re-determine the value of the bleaching powder at 
 
282 THE METALLURGY OF GOLD. 
 
 short intervals of a few days. The shortest and best method of 
 effecting this is to grind a sample in a stoneware mortar under 
 water, and add the emulsion to an excess of a solution of potassic 
 iodide. The whole of the available chlorine instantly displaces 
 an equivalent quantity of iodine, which is set free and may be 
 readily estimated by a standard solution of hyposulphite of 
 soda in the usual manner. From this the amount of available 
 chlorine in the sample is calculated. The whole operation can 
 be performed in from ten to fifteen minutes when the standard 
 solution has been prepared. 
 
 In calculating the quantity of lime to be added to a barrel 
 charge, there are several considerations to be taken into account. 
 A saturated solution of chlorine acts more rapidly than non- 
 saturated solutions. At the ordinary temperature and pressure, 
 water dissolves about 2g volumes of the gas, and it is desirable 
 that the amount added to the barrel should be enough for this 
 saturated solution to be formed. Now, suppose the barrel to 
 have a capacity of 40 cubic feet, and that the charge consists 
 of say 125 gallons of water and 2,500 pounds of ore. The 
 volume of water added is 20 cubic feet, and if the particles of 
 ore have a mean density of 3'0, which is approximately true in 
 the case of many samples of roasted pyritic ore, the volume of 
 ore and water will be, together, 33J cubic feet, leaving 6J cubic 
 feet of air space inside the barrel. There must be, therefore, 46|^ 
 cubic feet of chlorine dissolved in the water to make a saturated 
 solution, and G^ cubic feet of chlorine in the air space, in order 
 to keep that amount in solution. The weight of 53 cubic feet of 
 chlorine is about 10*4 pounds, which could be supplied by the 
 decomposition of about 30 pounds of bleaching powder or 24 
 pounds per short ton of ore. This quantity may be taken as 
 one well adapted for the efficient chlorination of ordinary ores, 
 the chlorine acting rapidly at that pressure. Nevertheless, much 
 smaller quantities are in use with complete success on certain 
 ores, and, as has been already stated, no general rule, applicable 
 to all cases, can be laid down. When much oxide of copper or 
 other metallic oxides capable of absorbing chlorine are present, 
 or if the ore is insufficiently roasted, the pressure rapidly falls 
 off, and a further addition of chemicals may become necessary, 
 as at the Phoenix Mine. The total pressure exercised by the air 
 and chlorine, if the solution is saturated by the last-named gas, 
 is equal to two atmospheres i.e., one atmosphere in excess 
 of the normal and this may be taken as the greatest pressure 
 which it is advisable to maintain inside a barrel. If greater 
 pressures are used, either very expensive barrels are required, 
 or else the valves, manhole, &c., soon begin to leak. 
 
 It must not be forgotten that, at higher temperatures, water 
 dissolves less chlorine than the amount mentioned in the last 
 paragraph, and, consequently, a pressure of chlorine equal to 
 
CHLORINATION : THE BARREL PROCESS. 283- 
 
 that of the atmosphere will be obtained with the use of smaller 
 quantities of chemicals. Thus, at 40 0., with the other condi- 
 tions identical with those given above, the amount of chlorine 
 required to give the same pressure will be only 6*5 pounds 
 instead of 104 pounds, and, as the temperature rises above this, 
 the quantity of chlorine necessary falls off rapidly.* 
 
 The relative amounts of bleaching powder and acid used is 
 varied with different ores. Theoretically, according to the equa- 
 tion given above, 7 parts of chloride of lime require 6 parts of 
 sulphuric acid for complete decomposition, but practically a 
 little more sulphuric acid is always added, because it is desirable 
 to maintain an excess of the acid in the charge. This is done 
 in order to prevent lead and lime from getting into solution as 
 chlorides, which would entail a loss of chlorine, and also to assist 
 in dissolving any oxide of copper that may be present, which 
 otherwise would also absorb chlorine. The proportions added 
 are usually 6 parts of bleaching powder to 7 or 8 parts of 
 sulphuric acid of 66 B. 
 
 The time occupied in chlorinating usually varies from three to 
 six hours. The continued presence of an excess of chlorine gas 
 should be tested from time to time by opening a small valve momen- 
 tarily. If the presence of the free gas cannot be detected, the 
 barrel must be opened and further supplies of chemicals added. 
 The amount of pressure in the barrel is not alone sufficient to 
 prove the presence or absence of an excess of chlorine, as other 
 gases may be generated and exert considerable pressure. When 
 the chlorination is finished, the excess of chlorine is discharged 
 by a hose-pipe outside the building, the barrel is filled up with 
 water, again revolved, and the liquid decanted on to large, 
 shallow filter-beds. The barrel is again filled up, revolved, and 
 decanted as before, and, finally, the whole charge is emptied out 
 and another wash-water given to the charge on the filter. In 
 comparing this method of washing by decantation with that 
 of direct filtration, usually adopted, Mr. Thies found that, with 
 similar charges, the amount of water used in the latter case was 
 nearly double, and the time occupied much longer, while the 
 tailings contained 5J dwts. of gold per ton against about 1 dwt. 
 when washed by decantation. He also found that there is little 
 difficulty in filtering through a bed of fine ore from 3 to 4^ inches 
 thick, but if the thickness of the bed is greater, direct leaching 
 becomes very tedious and ineffective, and decantation is much 
 better. 
 
 The leaching vats are usually constructed of wood, which is 
 either lined with lead or coated with tar and pitch. At Bunker 
 
 * For the amount of chlorine dissolved by water at different temperatures, 
 vide Schonfeld, Ann. Chem. Pharm., vol. xciii.,p. 25; vol. xcvi., p. 8. The 
 results are quoted in Roscoe & Schorlemmer's Treatise on Chemistry, vol. i. 
 p. 123, 1881. 
 
284 THE METALLURGY OF GOLD. 
 
 Hill the filter tanks are rectangular, and measure 6 feet by 18 
 feet, and 18 inches deep. They are lined with lead, and incline 
 towards the drain hole, where the bottom is one inch lower than 
 it is at the other side of the tank. The filter-bed, as is usual in 
 California, is made of quartz-pebbles, gravel, and fine sand. In 
 other mills the leaching vats are usually round. 
 
 The Thies barrel process has been greatly altered and improved 
 during the last few years. The modern barrel chlorination pro- 
 cess, as practised in Dakota, differs from it in several essential 
 particulars ; it is described in Chapter xiv., pp. 303-316. 
 
 Mechanical Difficulties of Leaclimg. The difficulties of 
 leaching vary enormously with the character of the ore and the 
 treatment to which it has been subjected. Concentrates are 
 among the best leaching ores, as, even if they were originally 
 in a state of extremely fine division, the oxidising roasting in 
 many cases appears to cause an agglomeration of the particles 
 into porous granules, which do not pack down readily, and do 
 not resist the passage of liquids. A small quantity of red oxide 
 of iron in a very fine state of division is often present in roasted 
 concentrates, and this is carried away by the water, passing into 
 and partly through the filter-bed and appearing in the precipi- 
 tating vat. This material, on settling to the bottom of the gold 
 solution, often appears to carry down with it a large proportion 
 -of the gold, forming a layer of slimes which are extremely rich, 
 and very difficult to treat except by smelting. In some cases, 
 these iron oxide slimes are present in roasted pyrites in such 
 quantities that leaching is greatly interfered with, and made very 
 tedious. Siliceous ores, if properly pulverised, usually present 
 no difficulty in leaching, but aluminous ores are exceedingly 
 troublesome. At Mount Morgan, Queensland, the ore consists 
 chiefly of hydrated oxides of iron, which offer the greatest 
 possible resistance to leaching if treated raw, but, if roasted, the 
 loss of their water of hydration is found to be accompanied by 
 a remarkable agglomeration of the particles of ore, a-nd ordinary 
 gravitation leaching is thus rendered possible. The roasting 
 is in this case performed merely for the purpose of facilitating 
 the leaching, as it is stated that in the raw ore there are no 
 constituents, except gold, which are readily acted on by chlorine. 
 
 In order to quicken the process of leaching various appliances 
 have been suggested. Different forms of vacuum pumps have 
 been used, the amount of air in the space below the filter-bed 
 being reduced by them, and the liquid thus forced through by 
 atmospheric pressure. Such methods have not apparently been 
 attended with any conspicuous degree of success, as the increased 
 packing of the ore tends to neutralise the effects of the pressure 
 on the liquid. In 1889, at the Colorado Gold and Silver Extrac- 
 tion Company's Mill at Denver, the effect of the use of increased 
 pressure of air applied directly on the surface of the liquid was 
 
THE PRECIPITATION OF GOLD. 285 
 
 tried. A special cast-iron vat was constructed capable of sustain- 
 ing an internal pressure of 100 pounds per square inch, and 
 furnished with valves, so that air and water could be simul- 
 taneously pumped into it. It was found that ores, which 
 entirely prevented the passage of water through them, even 
 after a vacuum of 20 inches of mercury had been established 
 beneath the filter-bed, could be leached with great speed under 
 a pressure of from 30 to 50 pounds per square inch. Moreover, 
 when the leaching was complete, the ore could be freed from the 
 water more completely by the passage of a current of air through 
 it than by gravitation alone. This method, which was suggested 
 by Mr. Dennes, the Company's engineer, has since been adopted 
 at the Rapid City chlorination mill, and also, in a modified form, 
 at the Golden Reward Mill, Dakota. The chief objection to it 
 seems to lie in the great additional expense incurred in the 
 construction of the leaching vats and in working the pumps. 
 
 Mr. Riotte has suggested* that the wash-water should be 
 thoroughly mixed with the ore by agitation, and then removed 
 as completely as possible by squeezing in a filter -press. To 
 effect this, the ore, together with the necessary amount of water, 
 is passed successively through two revolving barrels, entering 
 and leaving them by means of hollow trunnions. The mixing is 
 accomplished inside the barrels by means of projecting inter- 
 nal ribs, and the charge . passes continuously through, and is 
 received into large filter-presses. These are set to work as soon 
 as they are full, and squeeze out all the liquid, retaining the 
 tailings, the pressure used being from 25 to 30 pounds per 
 square inch. Mr. Riotte finds that the average amount of 
 moisture retained in the ore after being squeezed is about 6 per 
 cent. As this would cause the retention in the tailings of from 
 3 to 6 per cent, of the soluble gold, according to the amount of 
 wash-water used, the method is obviously inapplicable to rich 
 ores, which would require to be subjected to treatment twice. 
 Centrifugal leaching has also been proposed, and it is stated that 
 the Mount Morgan ore" can be leached in this way without 
 previous roasting. 
 
 THE PBECIPITATION OP GOLD. 
 
 In adopting a system of precipitation suitable for application 
 on a large scale, there are several points to be taken into con- 
 sideration. (1) The precipitant should be capable of decomposing 
 gold chloride rapidly and completely, even when the latter is 
 present in extremely dilute solutions. (2) All other substances 
 likely to be present in the solution should be left unprecipitated. 
 (3) The excess of free chlorine in the solution should be removed 
 by the precipitant by being converted into hydrochloric acid or 
 into chlorides, so as to avoid running any risk of the precipitated 
 * Eng. and Mng. Journ., March 31, 1888. 
 
286 THE METALLURGY OF GOLD. 
 
 gold being partially redissolved. (4) The precipitate of gold 
 must also be in sucli a form that it is easily separable from the 
 liquid. It is possible to obtain it in such a finely divided form 
 that it will not settle in water at all, and will pass through the 
 closest filter. 
 
 The precipitants which have been proposed or used may be 
 conveniently divided into two classes. 
 
 1. Soluble precipitants and gases, by which solid particles of 
 gold are formed, suspended in the liquid. These are either 
 allowed to subside and are separated by decantation, or they 
 are removed by filtration. 
 
 2. Insoluble solid precipitants, on which the gold forms as a 
 deposit, and from which it has to be separated by subsequent 
 operations. 
 
 1. SOLUBLE PRECIPITANTS. 
 
 Ferrous Sulphate. This is the best known precipitant, 
 having been used by Plattner, see p. 226, and being still em- 
 ployed more frequently than any other in small mills, though it 
 has been discarded in favour of sulphuretted hydrogen in the 
 modern large mills in the United States owing to the slowness 
 with which the precipitated gold settles and to the difficulty of 
 collecting it by filtration. Some account of its use has already 
 been given on p. 271. It is made by dissolving iron in sul- 
 phuric acid, with the aid of heat, and the crude solution thus 
 prepared, which is in most cases added direct to the gold solution, 
 always contains some free sulphuric acid. Precipitation takes 
 place according to the equation 
 
 2AuCl 3 + 6FeS0 4 = Au 2 + Fe 2 Cl 6 + 2Fe 2 (SO 4 ) 3 
 
 'The oxidation of the ferrous salt is also effected in other ways, 
 notably by the excess of free chlorine present in the solution, so 
 that much more sulphate of iron is required than is indicated by 
 the equation. The difficulty of collecting and saving the preci- 
 pitated gold has already been dwelt on. The gold settles better 
 if it is well stirred, and Aaron recommends an addition of more 
 sulphuric acid and vigorous stirring, two hours after the 
 precipitation is complete, as a means of assisting the settling. 
 It was proposed by Mr. Yautin to collect the precipitated gold 
 by the use of centrifugal force. For this purpose the liquid, 
 containing finely-divided gold, was placed in a circular vessel, 
 capable of being rotated at great speed. On rotating the 
 cylinder, the liquid in it was also gradually put in motion, and 
 the finely-divided gold then moved outwards by centrifugal force, 
 and was pressed against and adhered to the inner wall of the 
 vessel, while the clear liquid could then be siphoned off. Although 
 stated to be successful on a small scale, this device has not yet 
 been adopted in practice. Besides gold, the only other metals 
 precipitated by ferrous sulphate are those which form insoluble 
 
SOLUBLE PRECIPITANTS. 287 
 
 .sulphates viz., lead, calcium, strontium and barium. The last 
 two of these are rarely present, and the others are dealt with in 
 the manner already described above. Basic iron salts are not 
 precipitated if enough free sulphuric acid is present, and, when 
 precipitated, they may be removed from the gold by treatment 
 with acids, or by slagging them off in the furnace. 
 
 Other proto-salts of iron are equally efficacious, but are never 
 used in practice. 
 
 Organic Substances. Many organic substances, such as 
 oxalic acid, formic acid, ether, &c., also decompose chloride of 
 gold, but are unsuitable as precipitants in practice, owing to 
 their slowness of action and high cost, or to the extremely fine 
 state of division in which the gold is thrown down, so that it 
 is made difficult or impossible to collect. Nevertheless, it is 
 probable that some of the more easily oxidisable organic com- 
 pounds may at some future time be found to be suitable for the 
 precipitation of gold on a large scale. 
 
 Sulphuretted Hydrogen. This was formerly made at Deloro 
 by heating paraffin and sulphur together, and the gas diluted 
 with air was forced through the solution by means of a small air- 
 pump. The use of the air was to keep the solution agitated, and 
 to expel part of the chlorine mechanically, and so economise the 
 sulphuretted hydrogen, which is decomposed by chlorine. The 
 equation for this decomposition is given by Mr. W. Langguth * 
 as follows : 
 
 H 2 S + 4H 2 + 8C1 = H 2 S0 4 + 8HC1 
 
 but it is more likely that the greater part of the change is repre- 
 sented by the well-known equation 
 
 H 2 S + C1 2 = 2HC1 + S 
 
 although a small amount of sulphuric acid may be formed at the 
 same time. The precipitate of sulphur thus formed before and 
 during the precipitation of the gold, which begins before the 
 whole of the chlorine has been destroyed, is to be avoided, and 
 Mr. Langguth has, therefore, suggested the use of sulphur 
 dioxide, generated by burning sulphur to destroy the free 
 chlorine, the reaction being as follows 
 
 C1 2 + S0 2 4- 2H 2 = H 2 S0 4 + 2HC1 
 
 When almost all the chlorine has been thus converted into 
 hydrochloric acid, the point being determined by testing some of 
 the liquid with a few bubbles of sulphuretted hydrogen, the 
 passage of SO 2 is stopped, and sulphuretted hydrogen, now 
 generated by the action of sulphuric acid on iron matte, is 
 forced into the solution, destroying the last traces of chlorine 
 and precipitating the gold. This system was introduced at the 
 Golden Reward Chlorination Works in 1891, and has been sub- 
 sequently adopted at other mills in the United States with 
 conspicuous success. 
 
 * Eng. and Mng. Journ., Feb. 14, 1891. 
 
288 THE METALLURGY OP GOLD. 
 
 The gold is precipitated as a sulphide mixed with more or 
 less sulphur, the reaction being represented by the following 
 equation 
 
 2AuCl 3 + 3H 2 S = Au 2 S 3 + 6HC1 
 
 It is said to take less than an hour at the Golden Reward 
 Works to precipitate the gold from 5,000 gallons of solution 
 (resulting from the lixiviation of from 25 to 50 tons of ore). 
 The liquid is quite cold, but the precipitate is in a collected, 
 voluminous and flocculent form, that settles quickly. It is left 
 undisturbed for two hours, and the liquid is then drawn off to 
 within 4 inches of the bottom of the vat, and passed through a 
 Johnson filter-press, provided with a set of heavy, canton-flannel 
 filter-cloths. The head of liquid used for filtering is 25 feet, and 
 the nitration is said to occupy from three to four hours, accord- 
 ing to the amount of sulphides already contained in the filter. 
 When the latter is full, a small air-pump is connected with it and 
 a current of air passed through it for an hour to dry the mass of 
 sulphides into hard cakes, which are easily handled and removed. 
 The precipitate is then roasted in iron trays, which are placed in 
 cast-iron muffles which are heated only from the top, no stirring 
 being necessary, so that the loss from dusting is small. The 
 filter-cloths are burnt with the precipitate, when they have 
 become either clogged with sulphides, or untrustworthy owing^ 
 to the action of the acid liquors. The precipitate is then melted 
 down with a little borax and nitre, the total loss in handling 
 being very small. The bullion is about 900 to 950 fine in gold, 
 the remainder consisting chiefly of silver, copper, lead and 
 arsenic. The bulk of the precipitate remains at the bottom of 
 the vat. It is allowed to accumulate for a fortnight, and is then 
 filtered and treated as above. The slag resulting from the 
 fusion of the gold is crushed and melted down with litharge, and 
 the reduced lead cupelled (Rothwell). 
 
 It has usually been assumed that all the lead, copper, and 
 silver contained in the liquids are precipitated with the gold, 
 and that if much copper is present, the bullion will be very base, 
 and Langguth suggested the removal of the copper from the pre- 
 cipitate by dilute nitric acid. Rothwell, however, has stated 
 (Mineral Industry for 1896, p. 278) that, under normal condi- 
 tions, the gold can be precipitated, forming a bullion from 820 to 
 960 fine, whilst the copper remains in solution together with 
 barely a trace of gold. 
 
 Sulpliurous Acid, which has been already mentioned as a 
 cheap agent for destroying free chlorine, precipitates gold very 
 completely from dilute solutions, but its action in the cold is 
 slow until the liquid is almost saturated with the gas (water 
 absorbs 39 volumes of the gas at 20), and the gold settles slowly. 
 At higher temperatures the action of the gas is much more rapid 
 
SOLID PRRCIPITANTS. 289 
 
 and satisfactory, and very little is wasted in saturating the liquid. 
 As a precipitant, the gas does not appear to present any advan- 
 tages over ferrous sulphate. 
 
 2. SOLID PREC1PITANTS. 
 
 Charcoal. In 1801, Henry referred to the reducing action 
 of charcoal on chloride of gold in solution, and observed that the 
 gold will be precipitated on the charcoal, if the solution is either 
 exposed to the direct light of the sun or heated to 212 F.* In 
 1869, in Percy's laboratory* some sticks of wood-charcoal were 
 immersed in water, and 32-50 grains of gold in the form of 
 chloride were added on August 7. 85 more grains of -gold, in 
 the form of chloride, were added on November 3, 1869, and 
 the bottle was left to stand. This bottle is in the Percy collec- 
 tion, at South Kensington, and the surface of the charcoal is 
 now coated over with metallic gold which shows on its surface 
 the fibres and vessels of the stem. 
 
 Vegetable charcoal was first employed as a precipitant for 
 solutions of gold chloride on the large scale by Mr. W. M. Davis,, 
 in a chlorination mill in Carolina, in the year 1880. Its use was 
 discontinued, however, and nothing further was heard of it until 
 it was adopted by Messrs. Newbery & Vautin, at the Mount 
 Morgan Mine, Queensland, in 1887. The method adopted there 
 is as follows : The solution is heated to boiling, the free chlorine 
 being thus expelled ; the liquid is then made to run slowly 
 through large shallow tanks filled with pieces of charcoal, varying 
 from the size of a small hazel nut to less than that of a pea. 
 The tanks are about a foot deep, and the overflow from one is 
 allowed to run through others, until the solution is found, on 
 testing, to be free from gold, which is deposited on the surface oi 
 the charcoal. When the latter is coated sufficiently with the 
 precious metal, it is burnt in small furnaces, furnished with 
 large dust chambers or apparatus for condensing the fumes, and 
 the ashes melted with borax, the gold being thus obtained in a 
 remarkable state of purity. Animal charcoal cannot be used 
 owing to the difficulty of burning it afterwards. 
 
 The exact action of the charcoal has not been fully demon- 
 strated. It acts slowly on cold solutions, and its action is not 
 rapid even at boiling point. It is under the disadvantage that 
 it does not destroy free chlorine, which must therefore be 
 expelled by boiling or by passing a current of air through t he 
 liquid before the precipitation of the gold is begun. Mr. Davis 
 states that 240 parts of charcoal are required for the precipitation 
 of 19| parts of gold. The prevailing opinion is that the hydro- 
 gen and hydrocarbons remaining in the charcoal are the active 
 agents in the precipitation, hydrochloric acid and free gold being 
 
 '* An Epitome of Chemistry, by Wm. Henry. London, 1801. P. 95. 
 
 19 
 
290 THE METALLURGY OP GOLD. 
 
 formed. Charcoal is not now used in Carolina, the only place 
 where it is still employed being the Mount Morgan Mine. 
 
 Insoluble Sulphides. The use of sulphide of copper as a 
 precipitant for gold chloride was covered by a British patent, 
 taken out many years ago, but the discovery was not applied in 
 practice, and had been forgotten when Mr. C. H. Aaron described 
 his method of application.* Experiments were subsequently 
 conducted at the various mills engaged in working the Newbery- 
 Vautin process, with the object of finding if the sulphides of 
 copper or of other metals could be used in practice on a large 
 scale. It was found in Denver that the best method of preparing 
 the precipitated sulphide of copper is to add a boiling saturated 
 solution of sulphate of copper to a boiling saturated solution of a 
 mixture of the polysulphides of sodium, prepared by adding 
 sulphur to boiling caustic-soda solution, and to stir the mixture 
 vigorously. The sulphide of copper is precipitated in a granular 
 form, which settles quickly in water, and allows liquids to filter 
 through it readily, but offers a large surface, from the porosity of 
 the granules. It is moreover easily decomposed, and so is 
 especially active in precipitating gold. The reaction that occurs 
 may be expressed thus : 
 
 3CuS + 2AuCl 8 = 3CuCl 3 -f Au 2 S 8 
 
 Some metallic gold may possibly be formed at the same time, 
 according to the equation : 
 
 3CuS + 8AuCl s + 12H 2 O = 8Au + 24HC1 + 3CuS0 4 
 
 The cupric chloride is removed in solution and the sulphide of 
 gold precipitated on the surface of the granules of copper 
 sulphide. The precipitant may conveniently be contained in two 
 or three small vessels, through which the gold-solution passes 
 successively, either flowing in at the top and out at the bottom, 
 or in the reverse direction. The sulphide of copper and the 
 precipitated gold are partly carried off in suspension in the 
 liquid, which must, therefore, be passed through a filter of flannel 
 or other material. In order to quicken both the precipitation 
 and the rate of filtration through the filter-cloths, the liquid may 
 be heated, while a head of 10 or 15 feet makes filtration very 
 rapid. The free chlorine need not be expelled from the liquid, 
 as it is destroyed at once by the sulphide. 
 
 Mr. Blomfield found that the subsulphide of copper, Cu 2 S, 
 prepared by fusing together sulphur and copper, is more easily 
 handled than the precipitated sulphide, besides being cheaper. f 
 The Cu 2 S is prepared for use by crushing it and retaining only 
 that part which remains on a sieve having 100 holes to the 
 linear inch, after having passed through a 60-mesh sieve. 
 The precipitation vessels which he recommends are stout glass 
 
 * Eighth Report Gal. State Min., 1888, p. 839. 
 t Eng. and Mng. Journ., Jan. 24, 189H. 
 
SOLID PRECIPITANTS. 291 
 
 cylinders, 6 inches in diameter by 12 inches long, closed at each 
 end by glass-lined cast-iron caps held together by two bolts. 
 Each vessel has a false bottom to support the filter-cloth, and 
 is charged with 4 to 8 inches of the precipitant. Mr. Blomfield 
 recommends downward filtration of cold solutions, thus saving 
 the expense of heating them. It is stated that this precipitant 
 is now being used on the large scale at a chlorination mill in 
 South America. 
 
 Whether precipitated or fused sulphide of copper is used, it is 
 -advisable to defer the clean-up as long as possible, until the first 
 filter has had the greater part of its copper replaced by gold. It 
 is then cleared out, and the contents of the second filter trans- 
 ferred to the first, and so on, fresh sulphide being added to the 
 lowest filter. The mixture of gold and copper sulphides from 
 the first filter is carefully dried, mixed with borax, and melted 
 in plumbago crucibles. If the bullion produced is too base, some 
 of the copper may be removed by roasting the sulphides and 
 treating them with dilute sulphuric acid before fusion. 
 
 It has been proved by laboratory experiments* that both fused 
 and precipitated sulphide of iron are more rapid in their action, 
 than the corresponding copper compounds. The only apparent 
 disadvantage in their use lies in the fact, that some of the copper 
 contained in the solution is precipitated with the gold. When 
 copper is not present in the ore this drawback would not be felt. 
 
 Metals. In recent experimentst the author has found that, 
 in the laboratory, iron turnings constitute the quickest and most 
 trustworthy precipitant for gold chloride, its superiority over 
 the insoluble sulphides being more marked, however, at 60 or 
 80 C. than at ordinary temperatures. The separation of the 
 gold from the iron may be readily effected by methods similar 
 to those used in the separation from zinc, which has to be done 
 in the MacArthur-Forrest cyanide process, as described below, 
 p. 332. 
 
 The order of the rate of precipitation of gold from moderately 
 -dilute solutions of its chloride (one part in 20,000) appears to be : 
 metallic iron (quickest), sulphide of iron, sulphide of copper, 
 -charcoal, ferrous sulphate, metallic copper (slowest). This order 
 is that noted near the boiling point of water. The relative 
 rate of precipitation by sulphurous acid and sulphuretted 
 hydrogen was difficult to compare closely with that of the above 
 reagents, but they appeared to act more slowly than iron and 
 the sulphides. It does not necessarily follow from the results of 
 these experiments that metallic iron is the best precipitant on a 
 large scale, and it has not been tried in practice as yet. Pos- 
 sibly the order given above would not be preserved unchanged 
 at ordinary temperatures. 
 
 * Mining Journal, March 11, 1893. 
 t Loc* cit. 
 
292 THE METALLURGY OF GOLD. 
 
 MODERN PATENT PROCESSES OP CHLORINAT1ON. 
 
 The Newbery- Vautin Process. In this process an effort 
 is made to combine the advantages of the Hears and Thies pro- 
 cesses. Since a high pressure of chlorine increases the rapidity 
 with which gold is dissolved, it is desirable to use pressure, but 
 Thies had shown that the economy effected by reducing the 
 quantity of chlorine employed outweighed the advantages gained 
 by the high pressure. Messrs. Cosmo Newbery and C. Vautin 
 proposed to obtain the desired pressure by pumping air into the- 
 barrel, keeping the amount of chlorine as low as possible. It 
 was stated by them that the chlorine would be kept in a liquid 
 state or, at any rate, that it would be entirely dissolved in the 
 water if the air in the barrel were at a pressure of 60 to 80 
 pounds to the square inch. This is contradicted by theoretical 
 considerations, and no attempt appears to have been made to 
 prove it in practice. The only effect of the increase of air 
 pressure would be an increase in the amount of air, but not in 
 that of chlorine, dissolved in the water. The latter can only 
 be affected by variation of chlorine - pressure. According to 
 experiments made by the author at Denver, using the Newbery- 
 Vautin barrels, air-pressure does not seem to exercise any 
 influence on the rate of solution of the gold in ores, or upon the 
 percentage of extraction. The other proposal made by Messrs. 
 Newbery & Vautin was to leach by means of a single-acting 
 vacuum pump. Vacuum leaching has already been discussed on 
 p. 284. The particular form of pump proposed by Messrs. 
 Newbery & Vautin does not seem to possess special advantages. 
 Neither of these methods appear likely to pass into permanent 
 use in gold chlorination. 
 
 It has often been asserted that the Newbery- Vautin process 
 is identical with that formerly in use, with conspicuous success, 
 at the Mount Morgan Mine, Queensland. Nevertheless, as Jar 
 as the author is aware, neither air-pressure in the barrel, nor 
 vacuum-leaching by means of a single-acting pump ever formed 
 part of the method of procedure at Mount Morgan. Mills de- 
 signed to use the Newbery- Vautin patents were erected about 
 the year 1888 in the Transvaal, New Zealand, Colorado, and 
 Hungary, but it appears that none of these mills are now at work. 
 
 The Pollok Hydraulic-pressure Process. In this process 
 also, there seems to have been a confusion of thought in the 
 mind of the inventor, who does not clearly distinguish between 
 the pressure of chlorine gas, which increases its chemical activity, 
 and air- or water-pressure, neither of which has any effect on the 
 condition of the chlorine or the rate of solution of gold. Mr. 
 Pollok ensures the solution in water of the whole of the chlorine 
 present in the barrel by filling the latter completely with water. 
 
MODERN PATENT PROCESSES OF CHLORINATION. 293 
 
 When the barrel is closed and all the air has been suffered to 
 -escape, more water is pumped in, until the pressure inside the 
 barrel rises to at least 100 pounds per square inch. The chlorine 
 is generated inside the barrel by the action of bisulphate of soda 
 on chloride of lime, but it is doubtful if there is any economy on 
 any gold-field in the use of bisulphate of soda in place of sulphuric 
 acid. Pollok claims that by this high pressure the solution of 
 chlorine is forced rapidly into the pores of the ore. The process 
 in other respects presents no novel features, and it does not 
 seem likely to come into general use. 
 
 The Swedish Chlorination Process. This method was 
 devised by Mr. Munktell, who seems to have worked it out 
 without having visited either the vat or the barrel chlorination 
 works already established in other parts of the world. The 
 process enjoys the distinction, according to the published accounts, 
 of having been worked at a profit. It was first used at the 
 Fahlun Copper Works, Sweden, and has since been introduced 
 into Hungary, where it is in successful operation at Brade and at 
 Boitzas. The process as worked at Fahlun may be briefly 
 described as follows : The ore is roasted at a low temperature 
 with the view of obtaining the copper in the form of sulphates. 
 If silver is present, salt is added in the roasting furnace. The 
 calcined ore is then placed in false-bottomed, wooden vats and 
 leached with hot water, followed, if necessary, by hot, dilute 
 sulphuric or hydrochloric acids, by which all the copper and 
 much of the iron are removed from the ore. The solutions thus 
 obtained are run through tanks containing scrap-iron, by which 
 the copper and any silver in solution are precipitated. The 
 solution of ferrous sulphate may be saved and used subsequently 
 to precipitate the gold. The residues in the tubs are now in a 
 condition to yield up their gold readily, and are accordingly 
 treated by a solution of - 6 to 0'7 per cent, of chloride of lime 
 (bleaching powder) in water, mixed with an equal volume of 
 dilute hydrochloric acid of specific gravity 1-002 or 1*003, or of 
 dilute sulphuric acid. These solutions are mixed in troughs just 
 before they flow into the ore-vat. Chlorine is slowly generated 
 by the action of the acid on the hypochlorite of lime, and, being 
 partly present in the nascent state, attacks the gold vigorously 
 in spite of the extreme dilution of the solution, while there is 
 very little smell of the gas above the vats. The liquid is passed 
 through the ore until gold ceases to be dissolved, after which the 
 tailings are thrown away. The solution is heated to 1 60 F. by 
 steam, and precipitated by ferrous sulphate. The collection of 
 the gold is expedited and ensured by adding acetate of lead to 
 the solution ; by this device lead sulphate is precipitated with 
 the gold, and, in settling to the bottom, carries the particles of 
 precious metal with it. This process was in continuous operation 
 at the Fahlun Copper Works from 1885 to 1888. In that period, 
 
294 THE METALLURGY OP GOLD. 
 
 1,500 tons of gold ore and the tailings from 29,000 tons of copper 
 ore were subjected to treatment. It is stated that in the year 
 1886, the tailings from 14,000 tons of copper were treated with 
 the following results : 
 
 AVERAGE AMOUNT OF GOLD IN TAILINGS. 
 
 Before treatment, . . , 41 '82 grains per ton. 
 
 After . . 4-04 
 
 COST PEE, TON. 
 Chloride of lime (6 Ibs.), . 5d. 
 
 Sulphuric acid (8 '37 Ibs.), 
 Plumbic acetate and other reagents, 
 Fuel for steam, 
 Labour, 
 
 Id. 
 fd. 
 lid. 
 Id, 
 
 Total, . . 9d. 
 
 The low cost assigned to labour is very remarkable. These 
 tailings had, of course, been previously crushed and roasted 
 before being treated as above. 
 
 In the same year 960 tons of gold ore were treated at a cost of 
 12s. 2|d. per ton. the ore containing 523'62 grains of gold per ton 
 before treatment, and 6*02 grains per ton after treatment, so- 
 that no less than 98 -85 per cent, of the gold was extracted. 
 
 The following details concerning the practice at Brade,. 
 Hungary, may be added.* In addition to gold, the ores contain 
 iron pyrites, barytes, zinc-blende, antimonial minerals and 
 argentiferous galena, and also in some cases calcite and carbonate 
 of magnesia. The concentrates, which are subjected to treatment 
 by the Munktell process, contain 0*86 oz. of gold and 4 ozs. of 
 silver per ton, and 36 to 40 per cent, of sulphur. It is obvious 
 that such concentrates could not be treated advantageously by 
 the modern barrel process as described on pp. 303-315. 
 
 The oxidising roasting takes twenty-eight hours, after which 
 5 per cent, of salt is thrown upon the ore and mixed thoroughly 
 with it, and four hours later the charge is drawn from the 
 furnace and allowed to cool slowly in a covered-in brick pit. 
 The loss by volatilisation is, however, considerable, and it is 
 proposed to build seven-story furnaces. The leaching vats are 
 made of wood, lined with lead, and are 3 metres broad, 5 metres 
 long, and 0'75 metre deep, and hold 10 tons, the charge being 
 0'5 metre deep. There is a false bottom and filter-bed of quartz 
 as usual. The ore is leached with five different solutions in 
 succession, viz. : (1) Warm water, 2,500 gallons being required 
 for the charge : Chlorides of copper and zinc and about 25 per 
 cent, of the silver are removed in solution. (2) A solution 
 containing 2 per cent, of hyposulphite of soda, 1,600 gallons- 
 
 * Proc. Inst. Civil Eng., Session 1891-92. 
 
MODERN PATENT PROCESSES OP CHLORINATION. 295 
 
 being used ; this dissolves the rest of the silver. (3) Dilute 
 sulphuric acid, 2,700 gallons being used to remove the oxides of 
 iron. (4) Weak solutions of bleaching powder and sulphuric 
 acid to dissolve the gold ; and (5) Cold water to wash the ore. 
 The solutions are all preserved separately. The leaching takes 
 in all nine or ten days, the chlorination alone occupying three 
 days. The gold and silver are precipitated from their solution 
 by sodium sulphide, and the precipitate is dried, pressed, roasted, 
 and melted down. 
 
 The expenditure per ton of concentrates is as follows : 
 
 *. d. 
 Salt (110 Ibs.), . 17 
 
 Sulphuric acid (102 Ibs.), 
 Bleaching powder (25 Ibs. ), 
 Hyposulphite of soda (14'7 Ibs.), 
 Fuel, .... 
 Labour, 
 Amortisation, 
 
 6 
 3 7 
 1 10 
 8 
 3 8 
 1 7 
 
 Total, . , , . . 26 3 
 
 The percentage of extraction is not given. 
 
 Cassel's Process. In this process a solution of salt in 
 contact with the ore is decomposed electrolytically, and the 
 chlorine thus set free attacks and dissolves the gold, which is 
 deposited in a hollow iron shaft in the centre of the vessel. The 
 apparatus is complicated and ill-adapted for its purpose, and the 
 process has not been a practical success. It has now been 
 completely abandoned, but is interesting as being the first 
 attempt to generate chlorine by electrolysis for attacking gold. 
 
 The Julian Process. This was devised by Mr. Julian, of 
 Johannesburg, South Africa. After chlorination of the ore in a 
 barrel, mercury is added to the charge and the barrel again 
 revolved. The result of this is to amalgamate the coarse gold in 
 the ore and to reduce the gold chloride, chloride of mercury and 
 metallic gold being formed, the latter amalgamating with the 
 excess of mercury. The charge is then emptied out and passed 
 over a set of amalgamated copper-plate tables, by which the gold 
 amalgam is partly retained. The tailings are then passed 
 through a series of " electrolytic cells," each of which has a bath 
 of mercury lying at the bottom, while an electric current is 
 passed through the water. In these cells, the compounds of gold 
 and silver, soluble in the water, and all floured mercury and 
 amalgam are supposed to be collected. The process has not yet 
 been proved to be a success in practice. The idea of decomposing 
 gold chloride by mercury is not new, and previous experience has 
 shown the action to be very slow and partial. The complicated 
 nature of the whole process is against it. 
 
 Greenwood Process. In this process gold is dissolved in 
 barrels by chlorine produced by the electrolytic decomposition 
 
296 THE METALLURGY OP GOLD. 
 
 of a solution of common salt. The problem of producing chlorine 
 economically by electrolysis is one of enormous importance, as 
 one of the chief causes of expense in chlorination lies in the cost 
 of the chemicals required. If power is alone required, then 
 wherever water power is available, the whole process may be 
 much cheapened. In the Greenwood patents, the chlorine so 
 obtained is not well applied, and the machinery need not be 
 described here.* 
 
 CHAPTER XIV. 
 CHLORINATION: PRACTICE IN PARTICULAR MILLS. 
 
 THE following description of the practice actually employed at 
 the works named will serve to show how far the general descrip- 
 tions given are applicable to particular cases : 
 
 THE VAT PROCESS. 
 
 1. Butters' Ore Milling Works, Kennel, California. f 
 The practice varies with the nature of the ore. The Idaho 
 sulphides from the Grass Valley constitute a case of extreme 
 difficulty. "They contain considerable quantities of lime and 
 magnesia, and some manganese, zinc-blende and chalcopyrite, 
 with 4 ozs. of gold and 8 to 12 ozs. of silver per ton. Roasting 
 with salt is prohibited by the heavy volatilisation losses thereby 
 incurred. The roasting is performed at a black heat, the charge 
 withdrawn, allowed to grow quite cold, and then wetted down 
 and screened immediately, in order to prevent the agglomeration 
 of the ore, as the presence of the anhydrous sulphates makes the 
 mass a natural cement. The ore is then leached in a vat for 
 from 48 to 72 hours with cold water, and the soluble sulphates 
 of copper, iron and zinc thus removed. The solution thus 
 obtained contains small quantities of silver and gold, perhaps 
 dissolved by the agency of the metallic salts. These are 
 recovered in the copper-precipitating tank. The alkaline bases 
 and carbonates are next neutralised with sulphuric acid (60 to 
 100 Ibs. of 66 Baum6 acid being added per ton of ore), and 
 20 Ibs. (diluted with water) of the same acid per ton are then 
 added and circulated for 24 hours to dissolve the oxide of copper, 
 and the charge washed for 48 hours with cold water, shovelled 
 
 * A detailed description of the Greenwood process is given in Eissler's 
 Metallurgy of Gold, pp. 351-369. 
 
 t Eng. and Mng. Journ., Dec. 20, 1890. 
 
THE VAT PROCESS. 297 
 
 -out, partially dried, screened back into the tank and impregnated 
 with chlorine. After leaching, the tailings contain from $4 to 
 $6 in gold and all the silver. Five per cent, of salt are added, 
 and the charge dried and roasted at as high a temperature and as 
 fast as possible, when little loss by volatilisation is experienced. 
 The silver is then removed by leaching with hyposulphite 
 solution." 
 
 It seems difficult to believe that this complicated system 
 (which, it will be observed, bears a strong resemblance to 
 Mr. Munktell's methods in several important particulars), can 
 really be the most economical one that can be devised for this 
 ore. 
 
 2. Plymouth Consolidated Gold Mining Company, 
 Amador County, California,* The ore contains 11 dwts. of 
 gold, mostly in the free state. It is crushed in a stamp battery, 
 no. 8 screens (40-mesh) being used, passed over a length of 
 20 feet of amalgamated plates (the upper one of which is copper 
 ^and the rest silver-plated) and then concentrated on Frue 
 vanners. The concentrates amount to from 1^ to 1^ per cent, 
 of the weight of the ore, and contain from 5 to 10 ozs. of gold 
 per ton. They are treated at the rate of 100 tons per month, 
 being kept damp until charged into the roasting furnace, to 
 prevent the formation of lumps. A " Fortschaufelungsofen " is 
 used for roasting, 80 feet long by 12 feet wide, its hearth 
 consisting of a long continuous plane, holding three charges at 
 one time, which are kept separate. The three stages are called 
 the " drying," " burning," and " cooking " stages. In the 
 (middle) burning stage, the bed of ore is kept thin and occupies 
 double the space of each of the other charges. The furnace is 
 worked by three eight-hour shifts of one man. The charges 
 weigh 2,400 Ibs., including 10 per cent, of moisture, and on an 
 average contain 20 per cent, of sulphur. Just before the sulphur 
 ceases to flame, f per cent. i.e., 18 Ibs. of salt are added. 
 
 The chloridising vat is 9 feet in diameter and 3 feet deep : it 
 holds 4 tons of ore. The filter-bed is 6 inches deep, and con- 
 sists of, (a) at the bottom, wooden strips, J inch wide, placed 
 1 inch apart ; (b) above this, 6-inch boards placed 1 inch 
 apart and laid across the strips ; (c) coarse lumps of quartz, 
 diminishing upwards to fine stuff; and (d) at the top, a cover of 
 6-inch boards placed similarly to the lower layer, but crosswise 
 to them. Chlorine is introduced on both sides of the vat, which 
 is left lu ted-up for two days, and then leached for four or five 
 hours, the tank being kept full of water during the operation. 
 A gunny-sack protects the surface of the ore from the direct 
 impact of the water from the hose. The ore in the impregnation 
 vat contains about 6 per cent, of water (crumbling after it has 
 
 * For a more complete description, see that given in the Eighth Report 
 Col. Stat. Min., 1888, of which the account appended is an abstract. 
 
298 THE METALLURGY OF GOLD. 
 
 been sieved), and is sieved into the vat through a screen of 
 J-inch mesh. 
 
 The gold solution is acidulated with sulphuric acid to pre- 
 cipitate the lead, and ferrous sulphate is then added to it in 
 another tank, after which it is allowed to settle for two days 
 before the supernatant liquor is siphoned off. The gold is left 
 to accumulate for fourteen days and the wet gold is then filtered 
 and fused in graphite pots. The average extraction is from 
 95 to 96 per cent, of the gold, and 7 dwts. per ton are left 
 in the tailings. All wood is protected from the action of the 
 acids by paraffin-paint. The cost of milling is said to be 
 39 cents per ton of ore, and the cost of concentration, roasting, 
 and chlorination, $13 '40 per ton of concentrates. 
 
 3. The Alaska Treadwell Mine.* These works turned out 
 more gold than any other chlorination mill in the United States, 
 until the Golden Reward and other Dakota mills were started. 
 The old-fashioned vat process in use has been adopted after 
 thousands of dollars have been spent in testing newer methods. 
 The works have been in active operation since the year 1884. 
 The material treated consists of the sulphides collected by Frue 
 vanners from a stamp battery, and contains 40 per cent, of 
 sulphur, mostly as iron pyrites, although the percentage of copper 
 is increasing. The gangue is quartzose, containing from 2 to 5 
 per cent, of calcite, which necessitates the addition of salt in the- 
 roasting furnace. 
 
 Roasting. This was first effected in Bruckner cylinders, which 
 were abandoned owing to the large amount of fuel consumed and 
 the enormous losses by dusting and volatilisation (which are said 
 to have amounted to 30 per cent, of the gold). The automatic 
 Spence furnace was then tried and proved to be useless until it 
 was used as a reverberatory. Six were erected, and the cost 
 of roasting was reduced by one-half, but the capacity was small 
 (10 tons per day in the six furnaces) and the consumption of 
 fuel great, and a reverberatory furnace being erected, was found 
 to be more satisfactory. The Spence furnaces were accordingly 
 discarded, and 20 tons of concentrates are now roasted per day 
 in four reverberatory furnaces of 13 by 65 feet, inside measure- 
 ment, at a low cost. The ore is not roasted quite dead, Mr. 
 Burfeind, the superintendent, having decided that when salt is 
 added the best results, with least volatilisation, are obtained by 
 leaving some sulphates undecomposed. He gives it as his 
 opinion that improper roasting is responsible for all losses of gold 
 in the tailings and by volatilisation. 
 
 The roasted ore is spread on the cooling floor, wetted down 
 
 and sifted carefully into vats, each of which holds 4 tons, and 
 
 is then impregnated with the gas. This operation occupies four 
 
 hours, after which the vat is left untouched for thirty hours, 
 
 * Eng. tk Mng. Journ., April 11, 1891. 
 
THE BARREL PROCESS. 
 
 299' 
 
 fresh gas being forced in a few hours before leaching. The 
 leaching usually requires twelve hours. The tailings are sampled 
 and assayed, and, if found sufficiently poor, are sluiced into 
 the sea. 
 
 The solution is run into collecting tanks and thence to the 
 precipitating vats, which already contain the necessary amount 
 of ferrous sulphate in solution. The precipitation is complete 
 when all the solution has been run in, or when the vat is full. 
 It is then stirred briskly for a few minutes and left to settle for 
 from eighteen to twenty-four hours, when the supernatant liquor 
 is siphoned off and passed through a large filter. The super- 
 natant liquor usually contains from 23 to 25 cents of gold per 
 ton of material treated, and this is all saved in the filter. 
 
 The clean-up is made twice a month, the drying and melting 
 of about 600 ozs. of gold being done by one man in one day. The 
 gold is washed into a small tub, allowed to settle, and the 
 supernatant liquor returned to the precipitating vat. The gold 
 is dried in an iron pan without filtering, and melted with a 
 little borax. 
 
 Each one of the chlorination vats, holding 4J tons of ore, cost 
 $50, and lasts for three years without any repairs. The filter in 
 it costs only the price of a few gunny-sacks, and lasts for six 
 months without any attention. The other vats are expected to 
 last a lifetime. The ferrous sulphate is prepared on the works 
 from sulphuric acid and scrap-iron. During the six months 
 ending November 31, 1892, 120,002 tons of ore were mined and 
 milled, the total cost being at the rate of $1'32 per ton. The 
 concentration of the ore yielded 2,703 tons of sulphides, which 
 were chlorinated at a cost of $8'42 per ton. This is very low 
 for the Plattner process, although perhaps high when compared 
 with barrel chlorination. In 1890, the cost was at the rate of 
 about $10 per ton, the yield in gold being over $40 per ton. 
 Later results are as follows : 
 
 YEAR. 
 
 Tons 
 Chlorinated. 
 
 Yield per 
 Ton. 
 
 COST PER TON. 
 
 Labour. 
 
 Supplies. 
 
 Total. 
 
 June, 1890, to May, 1891, 
 
 5869 
 
 2-02 oz. 
 
 $5-03 
 
 $3-99 
 
 $9-02 
 
 June, 1891, to May, 1892, 
 
 6177 
 
 1-52 ,, 
 
 4-80 
 
 2-81 
 
 7-61 
 
 THE BARREL PROCESS. 
 
 1. Mears' Process at Deloro, Canada. The following 
 description of the barrel process adopted at these works ia 
 
300 THE METALLURGY OF GOLD. 
 
 taken from the account given by the manager, Mr. John G. 
 Roth well, M.E. The ore is hand-sorted, crushed in rock- 
 breakers and Cornish rolls so as to pass a 16-mesh screen, then 
 sized and concentrated by jigs. The concentrates are roasted in 
 two White-Howell cylinders, being passed through one, where 
 nearly all the sulphides and arsenides are burnt, and then into 
 the other, where dead-roasting is performed ; the ore falls thence 
 direct on to the cooling floor, from which it is loaded into cars 
 and sent to the barrels. The first cylinder is 30 feet long and 
 -5 feet in diameter, and has eight brick shelves to assist in 
 stirring the ore. The second cylinder is 20 feet long and 4 feet 
 in diameter, with six shelves. The air supplied to the first 
 cylinder is heated by the escaping gases of the second. Both 
 cylinders are jacketted by air-spaces and a covering of mineral- 
 wool and paper. Ten tons of concentrates are roasted in 
 twenty-four hours, yielding over 4 tons of arsenious oxide, 
 which is condensed in brick chambers. These cylinders are 
 specially arranged so as to receive an unusually large supply of 
 .air, by apertures in the fire-bridge and around the fire-box ; they 
 are said to treat more ore with less loss by dusting than ordinary 
 White-Howell furnaces. 
 
 The chlorinators are cylinders of J-inch boiler-iron, having 
 cast-iron heads rivetted or bolted on; each head has a long 
 trunnion turned true and cored-out, with a shafting-box and 
 gland on it. The cylinders are lined with 20-lb. (to the 
 square foot) sheet-lead, the joints being burned. The barrels 
 are of two sizes. The smaller size is 40 inches in diameter and 
 44 inches long (inside measurement) ; it has a capacity of 
 32J cubic feet, and holds 1 ton of ore. The larger size is 
 60 inches in diameter and 72 inches long, has a capacity of 
 118 cubic feet, and takes a charge of 3 tons of ore. 
 
 A manhole in the centre of the shell enables a man to get 
 inside the barrel for repairs ; in the centre of the manhole cover 
 is an aperture for charging and discharging. Inside the cylinder, 
 :and extending from the charging aperture in the direction in 
 which the cylinder turns, is a pocket covered and lined with 
 lead, which is burnt on to the lining of the cylinder. The pocket 
 in the 1-ton barrel holds 100 Ibs. of sulphuric acid, in the 
 3-ton barrel, 300 Ibs. There are several lead-covered shelves 
 bolted to the inside of the cylinder, their use being to stir the 
 pulp thoroughly. Through the trunnion passes the "gooseneck," 
 (viz., an iron pipe covered and lined by lead), which is turned in 
 the direction in which the barrel revolves. The gooseneck is 
 held stationary in its upright position by clamps. 
 
 About 120 gallons (i.e., 60 per cent.) of water are added to the 
 ton (of 2,000 Ibs.) of ore. The proportions of chemicals used are 
 parts of bleaching powder to 7 parts of sulphuric acid, the 
 absolute amounts varying with the quality of the ore and the 
 
THE BARBEL PROCESS. 301 
 
 degree of perfection of tne roasting. The usual amounts are from. 
 36 to 54 Ibs. of bleaching powder, and from 42 to 63 Ibs. of acid 
 per ton. The lime is mixed with the ore and water, and the acid 
 poured into the pocket. The covering plate is then screwed on and 
 the barrel is set in motion, when the acid is immediately emptied 
 out of the pocket and is mixed gradually with the rest of the 
 charge. The cylinder revolves for from one and a-half to three 
 hours, after which the excess of gas is drawn off and the barrel 
 is partly exhausted through the gooseneck. Then the barrel is 
 opened and dumped into the filtering vat, which has a loosely 
 fitting cover placed on it, and an exhaust-pump connected with 
 the space below the filter-bed to quicken the leaching. Another 
 method adopted is to decant the liquid contained in the barrel 
 and conduct it to settling tanks ; then to add more water, revolve 
 for a few minutes and decant again. In this way a charge of 
 three tons can be washed in two or three hours, while by 
 exhaust-leaching, 1 ton takes from six to eight hours. 
 
 The clear solution is precipitated by sulphuretted hydrogen 
 made from paraffin and sulphur, thus The generator is a small 
 cylinder of heavy boiler-iron with cast-iron heads rivetted in, and 
 the seams caulked. In the head is a hole big enough to admit a 
 man's hand for the purpose of cleaning out the vessel. In the 
 shell is a charging hole and an outlet for the gas, consisting of a 
 2-inch gas-pipe. The generator is built in a small brick fire-place 
 with a charging hole or pipe on the top, and a hand-hole outside. 
 A small fire is used, and the charge consists of one part of single- 
 pressed paraffin to two parts of common brimstone. 
 
 The precipitating tanks are fitted with tight covers (in which 
 are well-secured manholes), and the only outlet from them is a 
 wide exhaust-pipe leading out of the building. In front of these 
 tanks are lines of steam- and gas-pipes connected separately with 
 each one. The gas, after being generated, passes into a receiver to 
 condense the oils, &c., and an air-pump forces it from this through 
 a perforated lead pipe, nearly to the bottom of the precipitating 
 vats. An air-hole left in the receiver prevents the formation of a 
 vacuum there, and enables the pump to pass a mixture of sul- 
 phuretted hydrogen and air through the liquid, an arrangement 
 which economises the sulphuretted hydrogen. When precipita- 
 tion is complete, a short time elapses and the liquid is then 
 allowed to flow through several small filters.* Three of these 
 filters will empty a 1,500-gallon tank in from three or four hours. 
 The precipitate is then pressed into cakes, dried, roasted, and 
 melted with borax. 
 
 The ore is an arsenical sulphuret of iron in a gangue of quartz 
 and calcspar. The "mineral" contains about 42 per cent, of 
 arsenic, 20 per cent, of sulphur, and 38 per cent, of iron. The 
 
 * For a description of these filters, see Eng. and Mng. Journ., Mar. 25,. 
 1885. 
 
302 THE METALLURGY OF GOLD. 
 
 roasted concentrates average $78-67 per ton in gold, and the 
 tailings about $3, the extraction being at the rate of 96-16 per 
 cent, of gold. 
 
 2. The Thies Process at the Phoenix Mine, Concord, 
 North Carolina, and the Haile Gold Mine, South Carolina.* 
 The barrel process is used at these mines, the methods of pro- 
 cedure being similar at the two places. Chlorine was formerly 
 forced in through a gooseneck, passing through the trunnion 
 (Mears' process), but it was found that the gooseneck wore out 
 quickly, and that its inner lead lining collapsed when the gas 
 was exhausted. It was, therefore, discarded after two years' 
 experience, and the chlorine is made inside the barrel by means 
 of sulphuric acid and bleaching powder. The ore is roasted in a 
 double-bedded reverberatory furnace, having a capacity of 3 
 tons of concentrates per day. The chlorinator is 42 inches in 
 diameter and 60 inches long, the capacity being 48 cubic feet, and 
 the charging-hole is 6 inches in diameter. The lining consists 
 of 10-lb. to 12-lb. sheet-lead. The weight of the charge is 
 from 2,000 to 2,500 Ibs., to which from 100 to 125 gallons of 
 water is added i.e.. enough to make an easily flowing pulp. At 
 the Phoenix Mine, where the ore contains copper, one-half of the 
 acid and bleaching lime is added first and the barrel turned at 
 .the rate of 15 revolutions per minute for 3 to 4 hours. For the 
 Phoenix ores, this half charge of chemicals consists of 20 Ibs. 
 of lime and 25 Ibs. of acid. After this time has elapsed the 
 barrel is opened, and the other half of the chemicals added, after 
 which the barrel is rotated for from two to three hours longer. 
 The charge is then let fall on to a filter-bed, 6 feet long and 8 feet 
 wide, the depth of the pulp on it being 4 inches, while the 
 filter-bed is 5 inches thick. The first solution is allowed to run 
 through until the surface of the ore begins to be exposed, when 
 it is again covered with water to the depth of 3 to 4 inches, and 
 when the ore surface appears again, the whole space above it, 
 11 inches deep, is filled, which is enough for the Phoenix ores. 
 In all, 2 tons of water are used to leach 1 ton of ore. 
 
 The bottom of the filter-bed consists of perforated glazed tiles 
 or the substance known as mineraline, and stones, gravel and 
 sand are successively piled on these as usual. The tanks for 
 precipitation are made less than 3 feet deep. The precipitation 
 is effected by ferrous sulphate, and care is taken that no soluble 
 chloride of calcium shall remain in the solution, or a bulky 
 precipitate of sulphate of lime will come down on the addition 
 of the precipitant. It is for this reason that an excess of 
 sulphuric acid is added to the charge, and more is added to the 
 solutions if they do not already contain some free acid. The 
 large amount of lime and acid added to this ore is necessitated 
 .by the presence of copper. The solutions are allowed to settle 
 * Eighth Report Col. State Min., 188S, p. 836. 
 
THE BARREL PROCESS. 303 
 
 for three days and then allowed to run through tanks filled with 
 scrap-iron, by which the copper is precipitated. 
 
 At the Haile Mine, where iron pyrites, free from copper, 
 is treated, only 10 Ibs. of lime and 15 Ibs. of acid per ton of 
 ore are used. Ten tons of roasted ore are treated in the large 
 barrels per day of ten hours, and 94 per cent, of the gold is 
 extracted. 
 
 The cost at the Haile Mine is $4-65 per ton, without reckoning 
 the expenses of superintendence and assaying, the cost of roasting 
 being $2-70, and of chlorination, &c., $1'95. The following 
 details are given on the authority of Mr. Adolph Thies, the 
 inventor of the process used and the manager of the mill : 
 
 Fire-assay value of ore, delivered to the stamps, $4*50, of 
 which $1*45, in free gold, is caught on the plates. 
 
 Average assay value of raw concentrates, . . . $30 per ton. 
 
 -Sulphur in raw concentrates, . . . . 40 to 45 per cent. 
 
 Value of roasted concentrates, ..... $50 per ton. 
 Average value of tailings from chlorination, . . $2 per ton. 
 
 COST OF ROASTING. 
 Cord of wood at $1-40, ...... $0'70 
 
 Two labourers at $1 -00, . . . . . .200 
 
 Per ton of roasted ore, . . . . $2 '70 
 
 COST OF CHLORINATING. 
 
 10 Ibs. chloride of lime at 3 cents, . . . . . $0'30 
 
 15 Ibs. commercial sulphuric acid at 2 cents, .... "30 
 Two labourers, J day at 90 cents, . . . "45 
 
 Chlorinator, $ day at $2 '00, 
 Power, .... 
 Sulphuric acid for ferrous sulphate, 
 Wear and tear, . . 
 
 Superintendence, 
 
 50 
 
 12* 
 
 124 
 
 10 
 
 05 
 
 Per ton of roasted ore, . . . . $>1'95 
 
 TOTAL COST. 
 Per ton of roasted ore, . . . . . . . $4 -65 
 
 Per ton of raw concentrates, . . . . . 3 '50 
 
 A lead-lined barrel which had been in use for five years at 
 this mill showed no signs of wear, the abrasion of the lining by 
 the ore being evidently almost nil. 
 
 3. The Modern Barrel Process at the Golden Keward 
 Chlorination Works, Deadwood, Dakota. Mr. John E. 
 Roth well, in exchanging his position at Deloro for that of 
 manager of these works, adopted a system differing considerably 
 from the above, as may be seen from the following description 
 which is based on his writings,* and on accounts furnished 
 
 * Eng. and Mng. Journ., Feb. 7, 1891, and Mineral Industry for 1892, 
 p. 233. 
 
304 THE METALLURGY OF GOLD. 
 
 privately to the author by a late foreman of the mill and by 
 others : 
 
 The ore is crushed in a large Gates crusher and two sets of 
 Krom rolls, roasted partly in four Bruckner furnaces of 3 tons 
 capacity each, partly in a large White-Howell furnace, and 
 chlorinated in barrels of 3 and 4 tons capacity. After being 
 passed through the rock-breaker, which delivers no pieces larger 
 than 1^ inch in diameter, the ore is dried by passing through 
 a revolving cylinder, 18 feet long and 5 feet in diameter, the 
 capacity of which is somewhat increased by being divided into 
 four longitudinal compartments, by the same number of parti- 
 tions, made of plate-iron, which terminate at a distance of 2 feet 
 from each end of the cylinder. The fire-place is arranged to- 
 heat a large quantity of air, which is not entirely admitted 
 through the grate, but partly through channels by its side. 
 The revolving screens are hexagonal in shape, and are made 
 double with a coarse mesh inside the fine mesh, the object being 
 to protect the latter from undue wear. The mesh-frames are 
 slid into grooves in the screen-frame, and are interchangeable 
 and readily replaced when worn. The partially oxidised ores 
 ("red ores"), coming from workings near the surface, are roasted 
 in the White-Howell furnace, while the undecomposed pyritic 
 ores (" blue ores "), derived from deeper levels, and containing 
 much more sulphur than the first-named class, are roasted in 
 Bruckner furnaces. The ore from the roasting furnaces is spread 
 out in a thin, furrowed layer on the cooling floors, and, when 
 partially cooled, it is sprinkled with water and raked over, the 
 water drying out. 
 
 The actual capacity of the barrels used at the Golden Reward 
 Mill is 3J tons of roasted ore per charge. The following dimen- 
 sions apply to a 5-ton barrel. This consists of a ^-inch steel 
 shell, 9 feet long and 5 feet in diameter inside, lined with (J-inch) 
 lead of 24 Ibs. to the square foot. The cast-iron heads are 2J 
 inches thick and heavily ribbed, and are inserted inside the shell 
 and bolted to it through a flange. A 3-inch iron rod, covered 
 with lead, passes through the trunnions and the entire length of 
 the barrel. The trunnions are 12 inches in diameter, and the 
 charging-holes are oval, 11 by 16 inches, the cast-iron covers swing- 
 ing off on levers when required. There are two charging-holes 
 placed in the cover-plate of the manhole. The chlorination 
 barrel is also the washing and leaching vessel; this is arranged 
 by placing a supporting diaphragm, for a filtering medium, to 
 form the chord of an arc of the circle of the barrel. 
 
 The filter or diaphragm, as it is called, is made of asbestos 
 cloth, resting on a framework consisting of oaken planks, each 
 11 inches wide and as long as the barrel, and 2 inches thick. 
 The area of the filter is nearly 30 square feet. In the planks 
 are grooves running longitudinally, each groove being f inch 
 
THE BARREL PROCESS. 305 
 
 wide and deep, and the same distance apart. Transverse grooves, 
 a little deeper than the longitudinal ones, are also cut at 
 intervals of from 4 to 6 inches, and f-inch auger holes are bored 
 through the planks from the bottom of the transverse grooves at 
 short intervals. These planks are supported on wooden cross- 
 pieces placed transversely to the barrel, which rest on longi- 
 tudinal strips bolted to the shell. The liquid passing through 
 the asbestos filter-cloth collects in the grooves, and is drained off 
 by the filter-holes. This oaken filter-plate, which was devised 
 by Mr. D. Dennes, costs little, and is very effective and durable. 
 Before his suggestion was adopted, an iron grating covered with 
 lead, which was burnt-on, had been used to support the filter- 
 cloth. In the grating on one barrel there were no less than 
 28,000 holes, through each of which the lead coating had to be 
 carried and burnt-on separately. The grating thus made with 
 great labour only lasted for a short time, as a faulty joint in the 
 lead in any one of these holes allowed the iron to be corroded, and 
 caused the grating to go to pieces. Boards with auger holes 
 were tried, but the area of the filter was thus restricted to the 
 sum of the areas of the apertures, and the speed of filtration 
 greatly reduced. With the oaken-plates, however, the filter- 
 cloth rests on the sharp ridges between the grooves, the surface 
 being almost entirely available for filtration. Of course, when 
 the pressure is applied, the cloth sags down into the grooves, 
 but this only increases the area available for filtration. On the 
 top of the corrugated plates is placed the filtering medium, an 
 open-woven asbestos cloth. It is nearly as coarse as the ordinary 
 gunny-sack, but the warp and woof are of much heavier thread. 
 Over this is placed an open wooden grating, and the whole is 
 held in place by cross-pieces, the ends of which rest under straps 
 bolted to the inside of the shell ; although the filter, when made 
 in this way, is rigidly held in place, it can be very easily and 
 quickly removed when the changing of the asbestos cloth becomes 
 necessary. The time occupied in changing the cloth is about 
 one hour and a-half, and, under ordinary conditions, it will last 
 for from 15 to 18 charges or from 2J to 3 days. The wear is due 
 to the scouring action of the pulp, from which the filter-cloth is 
 very imperfectly protected by the wooden grating. This is made 
 of 1^-inch wooden pieces, so that its apertures are 1J inches 
 square.* The woodwork of the supporting plates and gratings 
 lasts much longer than the filter-cloths, but requires replacing at 
 somewhat frequent intervals. One carpenter is continually em- 
 ployed in making the woodwork for the filters of the 4 barrels, 
 and in fixing the new frames in position when necessary. The 
 lead lining of the barrels, in which several thousand charges have 
 
 * At the Independence Mill, which stands near the Golden Reward Mill, 
 the asbestos cloth is protected by a secret method, and lasts over three 
 months. In some mills the oak filter-plate has now been replaced by a 
 sheet of lead of 24 Ibs. to the square foot, perforated with noles about 
 | inch in diameter and covered with asbestos cloth. 20 
 
306 THE METALLURGY OF GOLD. 
 
 been treated, show little signs of wear yet. Two valves on each 
 side of the shell of the barrel, above and below the filter, are for 
 the inlet and outlet of the wash-water and solution respectively. 
 
 The barrel is charged by first filling the space under the filter 
 with water, which at the same time is allowed to pass through 
 the filtering medium, and wash it ; then the required quantity of 
 water is put in above the filter. The sulphuric acid is then 
 poured into the water, through which it sinks in a mass to the 
 bottom, without mixing with it; the ore is then charged in, as 
 follows : The hoppers are furnished with 2 shoots, one for each 
 charging-hole. The ore is let fall through these shoots alter- 
 nately, the hole through which ore is not being passed serving 
 as an air-vent. Meanwhile the bleaching powder has been 
 weighed-out and placed in two small kegs. When the ore has 
 all been introduced into the barrel, a workman, stationed at each 
 charging-hole, hollows out a space in the surface of the dry ore 
 with his hands, and, emptying one of the kegs into the barrel, 
 closes up the charging-hole as quickly as possible. If all these 
 operations have been conducted rapidly without a hitch, there is 
 no immediate evolution of chlorine, but, if some time is suffered 
 to elapse after charging-in the ore, the acid liquid, thoroughly 
 stirred-up and mixed by the fall of the ore into it, gradually 
 rises through and wets the charge, and the bleaching powder, 
 falling on ore which has been wetted with acid, gives off copious 
 fumes of chlorine before the cover-plate can be screwed on. 
 
 After the chlorination is complete viz., in about one and a-half 
 to two hours the barrel is stopped, so that the filter assumes a 
 horizontal position ; the hose is attached to on,e of the outlet 
 pipes, and after waiting for a few minutes to allow the pulp to 
 settle, as recommended by Rothwell, the valve is opened and 
 the solution allowed to run out, the pressure of gas left in the 
 barrel being enough to start the leaching. A hose is also 
 attached to the inlet pipe, and water is pumped in under a 
 pressure seldom exceeding 40 Ibs. per square inch, the air in 
 the top part of the barrel being compressed and forming an 
 elastic cushion. By washing in this manner, no chlorine is 
 allowed to escape into the building, as it is all absorbed by 
 the water. If necessary the leaching is suspended at intervals, 
 and the barrel is again revolved for a few minutes, so that 
 its contents are thoroughly mixed-up together again. In this way 
 the formation of permanent channels in the ore is prevented, 
 and perfect leaching is ensured. The wash-water coming from 
 the barrel is tested for gold with sulphuretted hydrogen. 
 
 The length of time required for the leaching varies with the 
 leaching quality of the ore treated charges having been leached 
 in forty minutes with a pressure of from 30 to 40 Ibs. per square 
 inch. With higher pressures the time can be materially shortened. 
 In order to facilitate the leaching of charges carrying an excess 
 
OF THB 
 
 THE BARREL PROCESS. 307 
 
 of dust or slime, the following method, now not in use, is 
 stated by Mr. Rothwell to have been formerly employed at 
 this mill. A valve placed in the casting of the head, on a 
 level with the surface of the pulp, is opened just after the barrel 
 is stopped, and the dust and slimes which remain in suspension 
 are run into an outside washing filter-press, or to the slime-filter 
 where it can be treated separately, and the charge washed in the 
 usual way.* 
 
 The tailings are discharged into a car which will hold the 
 whole charge of ore and water, and then run out of the building ; 
 or, if water is abundant, they are discharged into a sluice, and 
 washed away. The filter-cloth is washed clean by a jet of water 
 under pressure directed successively to all parts of it. This water 
 is discharged by revolving the barrel. 
 
 " For leaching purposes, the amount of water necessary to 
 wash a charge varies very little with the richness of the ore, 
 which goes to show the perfect leaching condition of the ore in 
 the barrel. The amount required is about 120 gallons per ton 
 over and above the quantity used in the barrel for chlorination, 
 which is about 100 gallons per ton. The water from the final 
 washings, which only contain small quantities of gold, is forced 
 up into an overhead tank, and used for the succeeding charge in 
 the barrel ; the quantity of solution to be precipitated is thus 
 reduced to about 120 gallons per ton of ore treated." t 
 
 The solution coming from the barrel, passes through a 2-inch 
 acid-proof hose-pipe into a large settling tank, one of a consider- 
 able number. These are round, wooden vats, lined with lead, 
 and are about 10 feet in diameter and 7J feet deep. When the 
 first tank is full, it overflows into the next one, which is placed 
 at a somewhat lower level, and this arrangement is repeated 
 through the whole series, the lowest vat overflowing into a 
 sump, whence the liquid, now nearly free from suspended 
 particles of ore which have settled to the bottom of the vats, 
 is pumped up into the precipitating vats, situated above the 
 cooling floor near the top of the building. The force-pump 
 used for this purpose is perhaps the weakest point in the whole 
 plant as it is rapidly corroded by the acid liquid, and requires 
 frequent renewal and repairs. 
 
 At the Independence Mill, the slimes are separated by being 
 passed through slime-presses, similar in shape to a Johnson 
 filter-press. J Asbestos cloth is used for the partitions, with the 
 addition of a layer of Canton-flannel if the slimes are very 
 
 *This device was first used by the late Mr. J. T. Blomfield, at the 
 Newbery-Vautin Chlorination Testing Works, in London, in 1889, and was 
 afterwards adopted by Mr. Rothwell in 1890. 
 
 t Eng. and Mng. Journ., Feb. 7, 1891, p. 166. 
 
 For description of Johnson's filter press, vide Stetefeldt's Lixiviation 
 of Silver Ores, p. 126. 
 
308 THE METALLURGY OP GOLD. 
 
 finely divided. The same pressure which effects the leaching in 
 the barrel forces the solution through the slime-presses. When, 
 these are full of slimes, they are washed out by water forced in 
 the opposite direction to that in which the solution had moved. 
 Slime-filters were formerly in use at the Golden Reward Mill 
 also ; they consisted of lead-lined cast-iron cylinders of 18 inches 
 deep and 30 inches in diameter, in which were arranged filter- 
 beds of sand and asbestos cloth. The chief difficulty encountered 
 in the use of these slime-filters is stated to have been that the 
 solution was squirted out in all directions in small jets, and part 
 of the valuable liquid thus lost. At the Independence Mill, the 
 filter-press is placed above a large, wide, shallow, lead tank, while 
 it is completely surrounded by a leaden hood, against which all 
 the jets of liquid strike and then run down into the tank. 
 
 At the Golden Reward Mill there are two precipitating tanks 
 for each barrel. They are 6J feet in diameter and 10 J feet deep,, 
 are fitted with covers, and are constructed of wood and lined 
 with lead of 8 Ibs. to the square foot. The filter press, which is 
 situated on the ground-floor, over 25 feet below the precipitating 
 vats, has twelve chambers of 19 inches square, and has a filtering 
 area of 57 square feet. Two lead pipes connect the precipitating 
 tanks with the press, one leading from the bottom of the tank 
 and the other from a point in the side about 4 inches above the 
 bottom. The precipitating gas generators are cast-iron cylinders 
 of the same size as the slime-filters ; only the sulphuretted hy- 
 drogen generator is lined with lead. 
 
 Before the precipitating tank is full of solution, the sulphur- 
 ous acid gas generator is started, and the gas forced through a. 
 2-inch leaden pipe leading to the bottom of the tank. The gas is 
 made from sulphur burned in the generator, with the aid of a 
 current of air forced in at the bottom and deflected over the sur- 
 face of the burning mass. The mixture of sulphur dioxide and 
 air, which is present in excess, is passed into the solution, where 
 the former is absorbed, while the air, bubbling through, serves to 
 stir the liquid. When the free chlorine has nearly all been con- 
 verted to hydrochloric acid, sulphuretted hydrogen, generated 
 from sulphide of iron and dilute sulphuric acid, is forced in, 
 together with a large excess of air, by which the liquid is kept 
 in a state of agitation, and the free chlorine expelled, thus 
 effecting a saving of SH 2 . The precipitate is flocculent, and 
 settles quickly. 
 
 After the precipitation is complete the "tank is allowed to 
 stand for two or three hours, when it has settled enough to 
 draw the supernatant liquor off through a filter-press by the 
 pipes in the side of the tank. There is little danger of precipi- 
 tating arsenic or antimony that may be in the solution when it 
 is worked cold, as they do not commence to come down till some 
 time after the gold has precipitated and collected. Of course 
 
THE BARREL PROCESS. 309 
 
 any copper or lead in solution will be precipitated with the gold, 
 but if small quantities only are present, they can be removed 
 subsequently. The loss in gold is considerably less if the preci- 
 pitate is allowed to accumulate in the tanks, and a clean-up 
 made after six to ten precipitations, than if it were filtered 
 through a press and collected after every precipitation. Continu- 
 ous filtration causes the filters to be coated and clogged with the 
 sulphides, and retards the operation unless high pressure is used, 
 which is sure to increase the loss, while a similar result attends 
 the handling of a large number of filter-cloths."* Rothwell 
 has stated more recently that gold can be fractionally precipi- 
 tated from solutions containing copper if free acid is present. 
 The sulphide cakes in the presses are compressed by blowing air 
 through them, and are turned out in the form of lumps, which 
 break up into powder when they are touched ; they are then 
 dried, roasted, and fused with borax, nitre, carbonate of soda 
 and sand. 
 
 The amount of precipitants required to precipitate a tank of 
 2,500 gallons of solution are sulphur, 2 Ibs. ; iron sulphide, 
 4 to 5 Ibs.; sulphuric acid, 16 Ibs.; water, 9 gallons. The 
 chemicals required for chlorination are bleaching powder, 10 
 Ibs., and sulphuric acid, 25 Ibs. per ton of roasted ore. The 
 power is furnished by an engine of 125 H.P., consuming six or 
 seven cords of wood, or about 4 tons of bituminous coal. Steam 
 is also required for the air-compressor, for steam-pumps, and for 
 heating the building. No salt is used in roasting, as the ore 
 treated contains little copper, &c., and no attempt is made to 
 extract the silver ; the fuel used (chiefly pine-wood) amounts to 
 a little more than that needed for power. The men employed 
 on the mill number about forty. " One man of ordinary intelli- 
 gence and a helper are able to take care of three barrels that 
 is, to look after the charging, leaching, and discharging. If the 
 tailings are sluiced out they can also attend to that ; but where 
 they have to be trammed out, one more man is necessary." f 
 
 There are four barrels, each of which treats 6 charges in 24 
 hours, or a total of over 90 tons per day. The cost of treatment 
 is given below. When the ore is crushed through an 8-mesh 
 screen, the tailings usually contain about 3J dwts. of gold per 
 ton, or 20 per cent, of the total contents ; but it is stated that 92 
 per cent, was being extracted in 1894. J 
 
 The disadvantages of the process used are due to the construc- 
 tion, but do not seriously interfere with the successful working of 
 the barrel ; they are, among others, the amount of space taken 
 up by the filter and the portion of the barrel under the filter, 
 and the fact that whenever the barrel is charged and running, it 
 is not in a perfectly balanced condition. These disadvantages 
 xmld be minimised, according to Mr. Rothwell, by using a filter 
 
 * Loc. cit. t Ibid. J Eng. and Mng. Journ., Jan. 27, 1894. 
 
310 THE METALLURGY OF GOLD. 
 
 placed close to the shell, and separated from it by a space only 
 just enough to allow of free circulation, and reaching to the same 
 height on the sides as the horizontal filter; then, by using 
 compressed air to displace the solution and wash-water, an 
 equally good result could be obtained. The mechanical diffi- 
 culties in constructing a curved filter-bed, however, prevent 
 its introduction. 
 
 According to Mr. Dennes, it appears that one of the disadvan- 
 tages in the use of the diaphragm is that it is difficult to keep it 
 in place. The ore and water wash to and fro through the filter- 
 cloth, and there is a strong tendency for the latter to become 
 displaced. If one plank in the somewhat complex framework, 
 by which it is held in place, becomes loosened, the whole soon 
 gives way as the barrel revolves, and the mishap is not detected 
 until leaching begins, when ore and water come out together. 
 The asbestos cloth is expensive, and only, lasts for about three 
 days, and all repairs must be done by a man getting inside the 
 barrel, owing to its construction. It has been more recently 
 suggested that the asbestos cloth could be replaced without 
 difficulty by a sand filter. 
 
 The cost of barrel chlorination in 1891, at the Golden Reward 
 Mill, is given by Mr. John E. E/othwell, as follows : * 
 
 July. August. September. 
 
 Amount of ore treated, 1,430 tons. 1,195 tons. 1,513 tons. 
 
 treated per day, 46'1 38'5 ,, 50 '4 
 
 Cost per ton 
 
 Milling, . 
 Roasting, 
 Chlorination, 
 Office, salaries, 
 Construction and 
 repairs, . 0'36 093 0'21 
 
 $1-49 $1-40 $1-44 
 
 1-53 1'35 1-37 
 
 1-76 1-67 1-65 
 
 0'40 0-47 0-37 
 
 $5-54 $5-82 $5-04 
 
 The total cost given above includes all the working expenses, 
 but is exclusive of interest on capital, taxes, insurance and 
 amortisation (i.e., extinction of capital by a sinking fund, which 
 covers the depreciation of plant, &c.) 
 
 The roasting was at that time done in Bruckner cylinders, and 
 its cost was afterwards reduced to about 70 cents per ton by the 
 adoption of a modified White-Howell furnace, and in January, 
 1894, the total cost of treatment per ton was stated to be only 
 $3'77. In 1896 the cost of roasting was less than 50 cents per 
 ton. (See p. 252). 
 
 The amount of chemicals used in the chlorination and precipi- 
 tation departments, in 1891, is given on the next page.f 
 
 *Eng. and Mng. Journ., Mar. 25, 1893, p. 269. t Loc. cit. 
 
THE BARREL PROCESS. 
 
 311 
 
 
 JULY, 1891. 
 
 AUGUST, 1891. 
 
 SEPTEMBER. 1891. 
 
 Lbs. per 
 ton of 
 Ore. 
 
 Cost 
 in 
 Cents. 
 
 Lbs. per 
 ton of 
 Ore. 
 
 Cost 
 in 
 Cents. 
 
 Lbs. per 
 ton of 
 Ore. 
 
 Cost 
 in 
 Cents. 
 
 Sulphuric acid, 
 Chloride of lime, . 
 Crude sulphur, 
 Iron sulphide, 
 
 26 '82 
 11-04 
 0-37 
 0-77 
 
 45-6 
 35-8 
 1-14 
 4-54 
 
 25-14 
 10-17 
 0-31 
 0-79 
 
 44-0 
 33-1 
 0-9 
 4-0 
 
 24-77 
 10-09 
 0-29 
 0-84 
 
 43-3 
 32-8 
 1-2 
 4-2 
 
 Total cost, 
 
 ... 
 
 87-1 
 
 
 82-05 
 
 ... 
 
 81-5 
 
 Method employed at the Bromination Mill, Rapid City, 
 Dakota. The following details were supplied to the author by 
 Mr. D. Dennes, the late foreman of the works. The method of 
 crushing has already been described, p. 231. The ore which was 
 treated in 1892 contained 6 or 7 per cent, of sulphur, a similar 
 amount of arsenic, and some iron, but little or no copper. It 
 was roasted in two White-How ell furnaces, each 30 feet long 
 and 60 inches in diameter inside the iron, lined with firebrick 
 5 inches thick, shaped so as to fit the curve of the cylinder. 
 There are six shelves in each furnace. The inclination of the 
 cylinder was varied with the ore, and the speed of revolution 
 was also varied from one turn in two and a-half minutes to three 
 turns per minute. A large Bruckner cylinder was also tried, 
 24 feet long and 9 feet in diameter, holding a charge of 15 tons 
 of ore, the duration of the loast being about six hours. This 
 gave a better product than the White-Howell furnaces, but the 
 cost of fuel was greater. The ore was cooled by being spread 
 out in a thin layer, and was sprinkled with water when very hot. 
 The roasted ore was then charged into lead-lined revolving 
 barrels of the same dimensions as those used at the Golden 
 Reward Mill viz., 8 feet long and 4$ feet in diameter, and 
 having a capacity of 3J tons. Hot water was introduced into 
 the barrel before charging-in the ore, only 66 gallons per ton or 
 33 per cent, of the weight of the ore being used. This seems too 
 small a quantity for efficient chlorination. The barrels had 
 cast-iron trunnions 12 inches in diameter, and were revolved at 
 the rate of twelve turns per minute by a worm and wheel. 
 They were made of boiler iron, f inch thick, and were lined with 
 thick lead of 24 Ibs. to the square foot on the ends, and 18 Ibs. 
 to the foot on the cylinder. 
 
 After adding the ore and water, the required amount of 
 bromine was poured in, and the barrel closed at once. Little 
 inconvenience is stated to have been caused to the workmen by the 
 bromine fumes, which were evolved owing to this crude method 
 
312 
 
 THE METALLURGY OF GOLD. 
 
 of manipulation; the men's eyes only were slightly affected. 
 From 6 to 13 Ibs. of bromine were added to the charge of 4 tons, 
 the amount being adjusted so that a very slight excess of free 
 bromine remained at the end of the operation. Chlorine 
 generated from sulphuric acid, and bleaching powder was used 
 for several months before the introduction of bromine, but the 
 cost was greater and the tailings only J dwt. poorer in the former 
 case. Moreover, the filter cloths lasted longer, the health of 
 the workmen suffered less injury, and the gold precipitate 
 was purer when bromine was used. The barrel was revolved for 
 
 Fig. 55. 
 
 from thirty minutes to one and a-half hours, and then discharged 
 into the leaching vat below, shown at Fig. 55. This is 7J 
 feet in diameter and 3 feet deep, made of cast-iron lined with 
 lead, and has a strongly ribbed cover, so that it is capable of 
 withstanding an internal pressure of 100 Ibs. to the square inch. 
 Internally, it tapers slightly upwards. The sides of the vat do 
 not rest on the bottom, but are supported by columns direct from 
 the floor. The bottom is not connected with the rest, but is 
 supported by a hydraulic ram, shown in the figure. On this 
 true bottom there is a filter-bed and false bottom. The filter- 
 cloth consisted of gunny-sacking. The vat was worked as 
 follows: When a charge was about to be introduced, the bottom 
 
THE BARREL PROCESS. 313 
 
 was raised by the ram aiid pressed tightly against the sides of 
 the vat, a good joint being made by a thick, rubber ring. The 
 hydraulic pressure was more than enough to overcome the air 
 pressure subsequently applied. The charge was then introduced, 
 the cover replaced and fastened down tightly by screw-bolts, and 
 air was pumped into the vat above the surface of the charge. 
 Water was also introduced by means of a lead pipe extending 
 round the vat at about the surface of the charge. This-pipe was 
 pierced with a number of small holes, by means of which jets of 
 water were thrown upon all parts of the ore surface. The air 
 pressure, which never exceeded 60 Ibs. to the square inch, forced 
 the wash-water through the charge very quickly. The solution 
 was of a strong ruby-red colour, due to the presence of free 
 bromine, and of the bromide of gold, the colour of which in 
 solution is very intense. There was no need to test the issuing 
 liquid, as its colour was found to be a perfect guide to an experienced 
 eye. When it had become colourless, the water supply was shut 
 off and the charge dried by pumping air through it. This was 
 necessary, as the company was not allowed to sluice the tailings 
 into the river. When no more water could be driven out, the 
 air-pressure was let off, and the bottom of the vat lowered until 
 the top of the filter-bed was just clear of the sides of the vat. 
 The bottom was then drawn sideways from below the vat by 
 means of a second hydraulic ram, and the charge fell into a large 
 ore-car below, having a capacity of 4 tons. This ore-car then 
 ran by gravitation to the dump, and was emptied and drawn 
 back by a wire rope winding on a drum actuated by steam 
 power. It is stated that the leaching occupied only twenty 
 minutes, and the vat was used to leach the charges from the two 
 barrels, which discharged into it alternately. The vat was so 
 rapid that it was always waiting for the barrels, and 100 tons 
 per day could be leached by it. It was designed by Mr. D. 
 Dennes, and, though of high initial cost, was found to be of such 
 great efficiency that its introduction was thought to be amply 
 justified. 
 
 The liquid was forced into large, lead-lined wooden precipita- 
 tion vats, each 20 feet x 10 feet x 6i feet, which were placed 
 outside the mill, buried under 2 feet of manure to keep them 
 warm in winter. The precipitation was effected by the successive 
 use of sulphur dioxide and sulphuretted hydrogen, applied in 
 exactly the same manner as at the Golden Reward Mill. The 
 collection of the sulphides was also effected in a similar manner. 
 When dried, the sulphides were put, together with the necessary 
 amount of borax, into a small barrel, revolving by machinery, 
 and were mixed thoroughly. The mixture was then transferred 
 by a scoop to a red-hot clay crucible (size No. 100) in the furnace, 
 additions being made at intervals until the crucible was full 
 of bullion and molten slag. The latter was very rich and 
 
314: THE METALLURGY OP GOLD. 
 
 full of shots of metal. It was stored and eventually sold to the 
 smelters, but owing to the difficulty in sampling it, it would have 
 been better to fuse it with lead and to sell the latter. 
 
 The air-compressor used was 12 inches in diameter, and had 
 an 18-inch stroke ; its maximum velocity was eighty revolutions 
 per minute. A large air receiver, 16 feet long and 4 feet in 
 diameter, was used between the pump and the leaching vat. 
 The maximum air pressure in the receiver was 80 Ibs. per square 
 inch. The dimensions of the bullion smelting furnace were as 
 follows : Depth 3 feet 3 inches. Inside diameter at bottom 2 
 feet 8 inches, at top 2 feet 2 inches. The flue measured 8 inches 
 by 12 inches, and was situated 10 inches below the top of the 
 fire-box. From 60 to 75 per cent, of the gold in the ore was 
 extracted, the tailings usually containing over 5 dwts. of gold. 
 The reason for this unsatisfactory result is said by Dr. L. D. 
 Godshall (Eng. and Mng. Journ., Jan. 6, 1894) to lie in the 
 coarseness of the crushing, a 10-mesh sieve being used. When 
 a 30-mesh sieve was used, only about 2^ dwts. were left in the 
 tailings, and the product was easily leached. It seems possible 
 that inefficient leaching, due to the formation of channels in the 
 ore, may have been another cause of loss, but the chief cause of 
 the failure was doubtless the fact that the amount of bromine 
 used was inadequate, the endeavour to reduce the costs of the 
 process being fatal to success. 
 
 The Cassei-Hinman Bromine Process. In this process the 
 bromine is regenerated, so that a large excess can be used in 
 dissolving the gold. The ore is treated in open vats by nascent 
 bromine generated in the interstices of the ore, and a rapid solu- 
 tion of the gold is thus effected ; the bromine is subsequently 
 recovered and used again. A mixture of bromide and bromate 
 oi soda or lime is prepared by adding bromine to a nearly 
 saturated solution of soda or milk of lime. A solution in water 
 of these salts, which can be obtained drv by evaporation, is cap- 
 able when treated with an acid of liberating all of its bromine, 
 the reactions being expressed thus: 
 
 (1) 6NaOH + 3Br 2 = SNaBr + SNaBrO + 3H 2 O 
 On heating or standing, this becomes 
 
 (2) GNaOH + 6Br = 5NaBr + NaBrO* + 3H 2 
 
 (3) 5NaBr + NaBr0 3 + 3H 2 S0 4 = 3Na 2 S0 4 + 3Br 2 + 3H 2 
 Under similar circumstances only part of the chlorine would 
 
 be liberated. 
 
 The process was introduced at the Nellie Bly Mill, Magnolia, 
 Colorado. The ore contained tellurium and could not be treated 
 either by cyanide or amalgamation. Formerly the richer tellur- 
 ides produced in the region were sold to the smelters, and the 
 mass of the low grade ores was practically worthless. 
 
THE BARREL PROCESS. 315" 
 
 The ore which contained from 10 to 15 dwts. of gold was 
 crushed with a Dodge Crusher, and then reduced by rolls to 
 pass a 15- to 20-mesh. It was then automatically conveyed into 
 a Brown Horseshoe Furnace, which gave excellent results, 
 requiring 6 tons of coal to roast 100 tons of silicious ore, and 
 the services of one man only to attend to the firing (see p. 252). 
 The cost of roasting did not exceed 50 cents per ton. The ore 
 was fed on the hearth at regular intervals in fixed quantities 
 and required eight hours roasting. The ore was not roasted 
 entirely sweet, as it was desired to contain some soluble salts, 
 for reasons which will be readily understood hereafter. From 
 the furnace the ore was conveyed by elevators into large storage 
 bins, from which it was ied into open lead-lined vats, provided 
 with the usual filter bed at bottom. The solution was then run 
 over the ores. It consisted of bromine, to which a dilute solu- 
 tion of caustic soda had been added. In percolating through 
 the ore, the alkali of this solution combined with the acid salts 
 of the ore and set free bromine in the nascent state, which dis- 
 solved the gold. The solution containing the gold, bromine, 
 and soluble bromides was then run into a closed tank, where 
 chlorine was added to liberate whatever bromine had combined 
 with the bases of the ore. 
 
 In the latest form of apparatus designed by the inventors, the 
 bromine thus set free is separated in the form of vapour by 
 allowing the solution to trickle down a scrubbing tower, where 
 it meets an upward current of air. The vapour is carried 
 from the top of this tower to the bottom of another scrubbing 
 tower, and in ascending meets a dilute solution of sodium 
 hydrate <>r milk of lime, by which it is dissolved. The regen- 
 erated solution is thus again ready for use upon a new charge of 
 ore. The loss of bromine is stated to be only about \ Ib. per 
 ton of ore treated, and is estimated by Crookes and Ramsay at 
 about 4 per cent, of the total quantity used. The solution 
 being alkaline emits no fumes, and at all times the bromine, 
 after dissolving the gold, is under perfect control. 
 
 The gold solution, after the bromine has been separated from 
 it by an air current, runs into a precipitating tank, and as no- 
 solvent is present, is at once precipitated by ferrous sulphate^ 
 or sulphuretted hydrogen. 
 
 For the purpose of shipping the bromine in an innocuous 
 state, the inventors make a dry salt, consisting of bromate, 
 bromide, and a chloride of a suitable base. 
 
 The great solvent power of bromine on gold is increased in 
 this process owing to the fact that it acts in the nascent state, 
 so that the extraction is very rapid. 
 
 Moreover, an excess of bromine can be used owing to the fact 
 that most of it is recovered. For the treatment of each ton of 
 ore an average of 7 Ibs. of bromine is used of which about 4 per 
 
316 THE METALLURGY OF GOLD. 
 
 cent, is lost. About 3J Ibs. of caustic soda or lime and 3 Ibs, of 
 sulphuric acid are also consumed. A little chlorine is also used 
 in freeing the bromine which has combined with bases of the 
 ore to form bromides. The inventors of the process estimate 
 the cost of the chemicals at only one shilling per ton of ore, 
 but of course this would vary somewhat with the nature of the 
 nr.aterial treated. 
 
 A large number of ores have been experimentally treated in 
 London, in quantities up to 1 ton at a time. Results obtained 
 by Sir Wm Crookes and Prof. Ramsay showed extractions of 
 from 93 to 98*85 per cent, of the gold from roasted pyritic and 
 telluride ores, with a loss of about 4 per cent, of the bromine. 
 Messrs. Johnson & Sons also treated, with success, a clayey ore 
 from Western Australia, containing 6 ozs. of gold per ton. In 
 this case the coarse gold was removed by amalgamation, and 99 
 per cent, of the gold in all was extracted. The author found 
 that some rich pyritic banket from the Transvaal yielded over 
 96 per cent, of its gold when treated by the process. This ore 
 was also treated by the author with cyanide, but only yielded 
 '25 per cent, of its gold. One of the advantages gained by using 
 bromine instead of chlorine is that solutions are obtained much 
 more free. from base metals which interfere with precipitation 
 and make the bullion base. 
 
 The Future of Chlorination. At the present day chlori- 
 nation is the most favoured method of treating auriferous 
 concentrates obtained from the tailings of stamp batteries, 
 except in districts where smelting works, producing mattes or 
 base bullion, exist- It is probable that, in general, such con- 
 centrates may be more economically dealt with by the cyanide 
 process, hut it seems doubtful it the high percentage of extrac- 
 tion attained by the use of chlorine can be equalled by the 
 newer method without roasting. During the last few years 
 the chief extension given to chlorination has been due to the 
 perfecting of the barrel process ; by its use certain virgin ores, 
 especially those of Mount Morgan, of Colorado, and of South 
 Dakota, have been successfully 'treated. The latest methods 
 introduced in Dakota, especially at the Golden Reward Mill, 
 certainly offer great advantages over the original practice in 
 ^Carolina and Queensland, and improvements in detail will pro- 
 bably continue to be made. For example, the recent increase 
 in size of the barrel in Western America from one containing 
 3 tons of ore to one holding 10 tons is of considerable import- 
 ance. Nevertheless, a doubt may be expressed whether the 
 complicated and costly machinery necessitated by these methods 
 will not stand in the way of their adoption in dther goldfields. 
 
 The cost of a 10-ton lead-lined barrel with accessories is over 
 500, and its weight is about 14 tons, so that its transportation 
 to mountainous mining regions is very difficult, although if 
 
THE BARREL PROCESS. 317 
 
 there were no other objection it could no doubt be made in 
 sections. The framework and foundations for such ponderous 
 barrels, which are usually raised many feet above the ground, 
 are expensive, and the power needed to turn the weight (when 
 charged) of over 20 tons is considerable. The weakest point 
 is, however, the filter bed, the framework of which requires con- 
 stant care and attention. The friction of the ore, and the move- 
 ment of the barrel tend to disarrange the frames which are 
 merely jammed into position. Moreover, the wood-work 
 becomes softened and corroded by the chlorine, and occasion- 
 ally gives way when the pressure is applied. Whenever this 
 happens, or whenever the asbestos cloth gives way, the barrel is 
 thrown out of work, generally for a large part of a day. In con- 
 sequence of this the capacity of such a barrel cannot be safely 
 reckoned as more than a daily average of 30 tons on average 
 ores, and in a mill treating 200 tons per day, the necessary vats 
 would occupy considerably less space than the barrels, so that 
 the cost of the covering building would be lower. Moreover, in 
 consequence of the loss of time caused when the filtering dia- 
 phragm gives way, it is probably unwise to extend the size of 
 the barrel beyond, at the most, 10 tons. 
 
 It is at least possible that American pioneers of chlorination 
 are working in the wrong direction. Dr. Munktell appears to 
 have shown that a weak solution of chlorine in water, acting on, 
 ores not subjected to agitation, is, if sufficient time is allowed, 
 almost as efficacious in dissolving gold as a strong solution com- 
 bined with agitation, or as chlorine gas acting on moistened ore,, 
 and the bromine process in open vats seems even more powerful. 
 If this should prove to be the case with most auriferous pyrites, 
 a plan of operation in chlorination, much simpler than those 
 hitherto in vogue, can be pursued. The roasted ore might be 
 charged into large round tanks of wood or masonry holding 
 hundreds of tons of material, such as those now in use with the 
 cyanide process in the Transvaal. It would then be sufficient 
 to leach with a weak solution of bromine, allowing it to run, 
 through by gravity, and, as it issues at the bottom, continually 
 raising it by means of a pump (preferably of the injector type). 
 The solution could then be poured on the surface of the charge 
 again, strengthened by the addition of chemicals if necessary, 
 and the whole operation repeated until the dissolution of the 
 gold is complete. There can be no doubt as to the cheapness 
 of the plant required in this case, and it is to be remembered 
 that similar methods have been selected as most advantageous 
 in the old hyposulphite, the Russell and the cyanide processes, 
 in all of which the progress that has been made is in the direc- 
 tion of greater simplicity. It is significant that, in spite of the 
 great technical success of the operations in Western America,, 
 the methods used there are not adopted in other countries. 
 
318 THE METALLURGY OF GOLD. 
 
 Moreover, it is reported that in one instance, at least in 
 Nevada, the chlorination process was abandoned in favour of 
 cyaniding in open vats, in spite of a high percentage of extrac- 
 tion, on account of the heavy repairs found necessary. Against 
 this must be set the numerous chlorination mills which have 
 been erected in Colorado and elsewhere. 
 
 As an instance of the direction in which the industry is pro- 
 gressing, the operations in Queensland may be quoted.* Here 
 a few years ago barrels were almost exclusively used in chlorin- 
 ation. At Mount Morgan, the largest chlorination mill in the 
 world, wooden barrels were used in which chlorine was generated 
 by bleaching powder and acid. However, both at Mount 
 Morgan and generally throughout Queensland, a return to the 
 vat process has been made, with a view to reduce the working 
 costs. The vats are, however, made larger, those at Mount 
 Morgan holding 25 tons of ore. The chlorine is now applied 
 there as chlorine water, and it is claimed that the consumption 
 of the solvent when used in this form can be better controlled 
 and kept within the absolute requirements for the extraction 
 of the gold. The gas is generated in stills by means of man- 
 ganese dioxide, salt and sulphuric acid, and the use of chloride 
 of lime has been completely abandoned. The chlorine is ab- 
 sorbed by water in scrubbing towers. The ore contains 1'66 oz. 
 of gold per ton, and 95 per cent, of this is extracted at a cost 
 of about 12s. per ton, the amount treated being about 200 tons 
 of ore per day. It will be observed that all the American 
 methods have now been abandoned at this mill, and the simple 
 method of procedure now adopted seems to have much to recom- 
 mend it. 
 
 Somewhat similar treatment is followed at Bovisa, in N. 
 Italy, where about 25 tons per day of pyritic ore, containing 34 
 per cent, of sulphur, 10 to 12 per cent, of arsenic, and 0'6 to 0-7 
 oz. of gold per ton, are roasted dead (the arsenic being recovered) 
 and chlorinated in lead-lined wooden vats holding 10 tons each. 
 Weak solutions of bleaching powder and sulphuric acid are 
 allowed to pass slowly through the ore, and the total time of 
 treatment is three days. From 85 to 87 per cent, of the gold is 
 .recovered, f 
 
 * See article by E. A. Weinberg in Report for 1894 on Queensland Mines. 
 t F. Clerici in Eng. and Mng. Journ., Aug. 29, 1896. 
 
THE CYANIDE PROCESS. 319 
 
 CHAPTER XV. 
 THE CYANIDE PROCESS. 
 
 The Cyanide Process, like most others, gradually grew up 
 from small beginnings. The power possessed by potassium cyan- 
 ide solutions ot dissolving metallic gold and silver has long been 
 known, as is pointed out, p. 354, but it was formerly believed 
 that the use of an electric current in conjunction was needed to 
 quicken the action, which was otherwise too slow to be of any 
 practical value. The use of cyanide of potassium in the stamp 
 battery has already been referred to, p. 131. It is probable 
 that its effect in dissolving gold was entirely overlooked when it 
 was being used in this way, the effect which it was desired to 
 produce being the removal or prevention of formation of the 
 crusts of carbonate of copper on the plates. In California, prior 
 to the year 1870, gold ores were sometimes digested in tubs 
 with water, to which a lump of cyanide had been added ; after 
 digestion, the ores were amalgamated by stirring with mercury, 
 to which a little sodium -amalgam was sometimes added. It is 
 obvious that, when no sodium was present, only the undissolved 
 gold was amalgamated, the dissolved gold being lost, not being 
 precipitated by mercury. These operations, however, can hardly 
 be said to foreshadow the eyanide leaching process, as they were 
 conducted in ignorance of the principles on which it is based, 
 and partly in defiance of them. 
 
 The first attempt at the direct extraction of gold by the use of 
 cyanide was made by J. H. Rae, who took out a patent in the 
 United States in 1867 (No. 61,866) for a process dependent on 
 the removal of gold and silver from their ores by the combined 
 action of a "current of electricity and of suitable solvents or 
 chemicals," such, for instance, as cyanide of potassium, the gold 
 being simultaneously precipitated on copper plates by the electric 
 current. This patent has now lapsed, and, although other 
 United States patents claiming the use of cyanide in the treat- 
 ment of gold ores were issued in 1880 and 1881, no real progress 
 seems to have been made. 
 
 In 1885, J. W. Simpson, of New Jersey, obtained a patent in 
 the United States (No. 323,222) which was of greater interest. 
 He proposed to treat ores containing gold, silver and copper by 
 a solution containing 3'0 per cent, of potassium cyanide and 
 
320 THE METALLURGY OF GOLD. 
 
 0-19 per cent, of ammonium carbonate. Copper was to be dis- 
 solved at the same time as the gold ; if silver was present also, 
 an addition of common salt was made. The inventor appears 
 to have believed that, by using carbonate of ammonia, the 
 necessity was obviated of employing an electric current, in con- 
 junction with cyanide of potassium, in order to dissolve the gold. 
 Of course the carbonate of ammonia would have no effect on the 
 amount of gold dissolved, whilst its action in inducing the 
 hydrolytic decomposition of cyanide, described in the sequel, is 
 a decided disadvantage. The precipitation of gold was effected 
 by " a piece or plate of zinc." The process does not appear to- 
 have been tried on a large scale, nor did any results follow from 
 the later experiments made on the subject by Endlich and 
 Miihlenberg,* and by Louis Janin, Jr.,f in America. 
 
 In 1886, however, a series of experiments on wet processes 
 of treating gold ores were begun by J. S. Mac Arthur and R. W. 
 and Wm. Forrest in Glasgow, and it was entirely owing to their 
 energy and skill that cyanide of potassium was successfully 
 applied in practice to the treatment of gold ores. Patents were 
 taken out by them in Great Britain in 1887 (No. 14,174), in the 
 United States in 1889 (No. 403,202), and in other parts of the 
 world. Their process consists essentially in attacking gold and' 
 silver ores by dilute solutions containing less than one per cent. 
 of KCy, caustic soda or lime being added to ores rendered acid 
 by the oxidation of pyrites, and then in precipitating the pre- 
 cious metals by zinc shavings. This process has now passed 
 into use in all parts of the world. Its success is complete on 
 many ores, and its extension will probably be very great, partly 
 at the expense of the chlorination process. The chief advantage 
 of the cyanide process over the chlorination process is that 
 roasting is not necessary, even if sulphides are present; this 
 is a most important point in the treatment of low-grade ores, 
 especially in places where fuel and labour are costly. More- 
 over, silver as well as gold is extracted by cyanide solutions. 
 Messrs. MacArthur and Forrest, in the course of their investiga- 
 tions on wet methods, became dissatisfied with chlorine as a 
 solvent, owing to its energetic and preferential action on sul- 
 phides of base metals and other bodies, and its inapplicability 
 to ores containing silver. They desired to find a solvent which 
 would exercise a selective action in favour of the precious 
 metals, instead of other substances. With this object in view, 
 they experimented with a number of solvents (such as ferric 
 bromide, ferric chloride, &c.), and finally decided that potassium 
 cyanide possessed advantages over all other substances. These 
 chemists lay especial stress on the advantages to be obtained by 
 using very dilute solutions, as they believe that, in this way, 
 
 * Ewj. and Mng. Jo/trn., Jan. 17, 1891, p. 86. 
 t Ibid., Dec. 29, 1890. 
 

 THT5 CYANIDE PROCESS. 321 
 
 the waste of chemicals, caused by reactions with the base 
 minerals present, is almost entirely avoided, while the dissolu- 
 tion of gold and silver is not retarded. 
 
 The MacArthur-Forrest Process. The following general 
 description of the working of the MacArthur-Forrest cyanide 
 process is based partly on accounts furnished by Mr. J. S. 
 Mac Arthur, to whom the author is much indebted for a large 
 amount of information, and partly on the actual work being 
 done in South Africa, in America, and elsewhere. The 
 mechanical treatment of the ore viz., crushing, charging into 
 vats, leaching, and discharging the tailings is based on the 
 same principles, and could be effected by the same general 
 methods as are adopted in chlorination or in any other leaching 
 process. Many details will, therefore, be omitted as having 
 been already covered in the course of the descriptions given in 
 Chapters xi. to xiv. 
 
 The process comprises four distinct operations, viz. : (1) Pre- 
 paration of the ore for treatment. (2) Solution of the gold in 
 potassium cyanide. (3) Precipitation of the dissolved gold by 
 means of zinc ; and (4) Conversion of the precipitated gold into 
 bullion by fusion. 
 
 1. Preparation of the Ore. The ore is crushed to a suitable 
 degree of fineness, which depends partly on the condition of the 
 gold and partly on the nature of the gangue. In the Transvaal, 
 where tailings are treated, the ore is crushed in stamp batteries 
 previous to amalgamation, the screens being usually about 25 
 or 30-mesh. In this case, sizing is performed to some extent 
 daring the operation of catching the tailings, the slimes in part 
 passing off suspended in the water used in the stamp-mill. 
 Nevertheless, masses of slimes occasionally accumulate and 
 resist the passage of the cyanide solution or retain the dissolved 
 gold. If mixed with sand, the slimes are less harmful, but 
 they tend to separate from the sand again when the leaching 
 begins. 
 
 At the Mercur Gold Mine, Fairfield, Utah, the ore is a silicious 
 limestone,* carrying magnetic oxide of iron, and in places con- 
 taining much silt, one of the proofs of its aqueous origin. Any 
 attempt at fine crushing results in the sliming of the greater 
 portion, and renders leaching impossible. The course of pro- 
 cedure adopted is consequently as follows : The ore is passed 
 through a Dodge rock -breaker, set to break the ore to a 
 maximum size of 1 inch ; it is then reduced to J inch by a 
 pair of 20-inch corrugated rolls ; thence it is passed over a 
 i-inch grizzly to a similar pair of 12-inch rolls, which reduce it 
 to a maximum size of J inch. After this treatment, 61 per cent, 
 remains on a No. 12 screen, while 26 per cent, passes through a 
 No. 30 screen, this part consisting almost entirely of impalpable 
 
 * Eng. and Mng. Journ., Nov. 5, 1892. 
 
 21 
 
322 THE METALLURGY OF GOLD. 
 
 powder. This material can be leached, although somewhat more 
 slowly than is desirable. 
 
 As a rule, ores are crushed so as to pass through screens 
 varying from sixteen to thirty holes to the linear inch ; in the 
 product thus obtained there is usually from 90 to 95 per cent, 
 of the gold which can be reached by the solution, and, if rolls 
 have been used for crushing, it is not difficult to leach. Stete- 
 feldt gives an instance* in which leaching by hyposulphite 
 solutions was successfully performed on the Bremen tailings, at 
 Silver City, New Mexico, when the average fineness was such 
 that 87 '8 per cent, of the ore would pass through a screen with 
 150 holes to the linear inch. The rate of leaching in this case 
 was about J inch per hour, with a vacuum of 14 inches of 
 mercury, a degree of slowness which would militate strongly 
 against the success of any works. For the separation of slimes 
 from crushed ore, see p. 324. 
 
 2. Solution of the G-old. This is effected in wooden, iron, 
 or cement vats, with false bottoms similar in construction to 
 those used in the Plattner process (see p. 261), or to those 
 employed in the Russell process described by Stetefeldt.f The 
 vats should be round, and may be discharged by sluicing through 
 a door in the side or bottom, or by shovelling out. Those in 
 use in Utah have sheet-iron sides with a wooden bottom, and 
 have a capacity of 14 tons of ore. In South Africa the vats are 
 often of 2,000 cubic feet capacity, and hold 75 tons, but the 
 Langlaagte tanks contain 450 tons each, and the Simmer and 
 Jack vats 600 tons each. 
 
 The use of large vats in preference to small ones is to be 
 strongly recommended, as the attention and superintendence 
 needed by a 10-ton vat costs as much as for a 600-ton vat. 
 Stetefeldt considers | that constructive difficulties limit the 
 diameter of wooden vats to 16 feet, but the Transvaal vats of 
 40 feet diameter give no trouble. The inside depth, Stetefeldt 
 continues, should not be more than 5 feet if the tailings are 
 shovelled out, and 7 feet if they are removed by sluicing. The 
 depth has usually little influence on the rate of leaching, as the 
 greater " head" of liquid tends to compensate for the increase of 
 compactness at the bottom of the vat. At Ousi, in treating 
 silver ores by the Russell process, it was found that the rate of 
 leaching was greater for a charge 6 feet deep than for one which 
 was only 2 feet deep. In South Africa the vats are often 10 feet 
 deep, and are now usually made of steel. 
 
 In calculating the capacity of a leaching vat, the volume of a 
 ton of raw ore may be taken at from 22 to 28 cubic feet when 
 dry, and from 20 to 26 cubic feet when wetted down. 
 
 The false bottom is usually a wooden framework, constructed 
 
 * Lixiviation of Silver Ores, p. 149. t Ibid., pp. 114-119. 
 
 /&*., p. 114. Ibid., p. 149. 
 
THE CYANIDE PROCESS. 323 
 
 of boards pierced with numerous auger holes, or of wooden slats 
 crossing each other and covered by canvas or cocoa-nut matting, 
 which are not rapidly destroyed by the solution, as they are 
 in the chlorination process. Below this a layer of coarse sand 
 and pebbles, through which the solution percolates, is sometimes 
 placed. 
 
 An excellent false bottom recommended by the Cassel Gold 
 Extracting Company, the owners of the MacArthur-Forest 
 patents, consists of angle-irons with the angles uppermost, and 
 the free edges nearly touching. The irons are supported on 
 wooden slats 1 inch thick, which give a space of that depth 
 below the niter-bed, in which the liquid may accumulate. The 
 triangular channels between the irons are filled with sand. 
 When the vat is being cleared out, the workmen's shovels slide 
 along the top of the angles and leave the filter-bed undisturbed. 
 Triangular pieces of wood may be used instead of angle-iron. 
 Thick canvas duck, resting on matting, forms a trustworthy 
 filter-cloth. 
 
 As in the case of chlorination, a layer of perforated glazed 
 porcelain tiles to support the filter-cloth is better than wood, as 
 no absorption of the solution takes place when the former is 
 used. If wood is used it is covered by a coating of paraffin 
 paint, or by a mixture of asphaltum and coal-tar. It was found 
 by Mr. Julian of the Salisbury Cyanide Works, South Africa, 
 that pine wood, lying 34 hours in a solution containing 0'3 per 
 cent, of cyanide reduced the strength to 0-005 per cent., while 
 glass reduced it to 0'2 per cent, and cement to 0*24 per cent. 
 Cement tanks have been used at the Langlaagte Estate since 
 May, 1892. 
 
 An iron pipe communicates with the space below the false 
 bottom, and conveys the liquid to the pumps or to the zinc 
 boxes. The solution does not attack wood or iron : brass and 
 bronzes are attacked and corroded rapidly. 
 
 The crushed ore is conveyed in cars pushed by hand, running 
 on an overhead railway, and emptied into the vats by tipping. 
 The supports of the railway must not touch the sides or any 
 part of the vats or their supports, as the jarring caused by the 
 running of the cars makes the ore settle down, and so renders 
 the leaching more difficult and tedious. 
 
 At the Langlaagte Works the tailings from the stamp batteries 
 are hauled for 600 yards up an incline of 1 in 15 to the tramway 
 above the vats, at a cost of 2d. per ton. At the Salisbury 
 Cyanide plant, a tailings wheel lifts the discharge from the plates 
 into a flume, which carries it to a hydraulic separator, where the 
 slimes are separated from the coarse tailings. The latter are 
 then run directly into the cyanide vats, thus saving handling ; 
 this method of direct filling is now adopted in several works on 
 the Rand, but the separation of the slimes from the sand is not 
 
324 THE METALLURGY OF GOLD. 
 
 complete, and this has a prejudicial effect on the percentage 
 of gold extracted. The separation of the slimes is beneficial in 
 expediting the nitration in the vats. 
 
 A widely-adopted form of slime separator is that known as 
 Butters & Meiii's collecting vat. It consists of a large round 
 vat, 20 to 30 feet in diameter and 20 feet deep. The tailings 
 from the stamp battery are run direct into a revolving distri- 
 butor, which spreads the pulp evenly over the surface of the 
 water with which the vat is filled. The sand settles to the 
 bottom and accumulates there, while the slimes overflow at the 
 top. When the accumulation of sand in the vat reaches to 
 within 5 feet of the top, the water is drained off and the ore 
 discharged by shovelling through an aperture in the bottom into 
 ore-cars below. This sand is now in a suitable condition for 
 leaching, containing only small quantities of slimes. At the 
 Jumpers mine the pulp collected in this way is treated directly 
 without removal. At the May Consolidated and the Croesus 
 mills, where the method of " double treatment " of the tailings 
 is employed, the first solution of cyanide is directly applied to 
 these vats, and, after draining, the pulp, wetted with cyanide 
 solution, is transferred to the second treatment vats (Hatch and 
 Chalmers). As already stated, the usual method is to remove 
 the collected sands to the cyanide vats for treatment. 
 
 The vats are filled to within a few inches of the top, and the 
 charge is levelled by means of hoes. The amount of ore charged- 
 in is such that, after the solutions have been applied, the surface 
 of the charge may stand at about 12 inches below the rim. In 
 levelling the ore, the labourer must not step into the vat or 
 forcibly press down the ore, as irregular filtering is produced by 
 this cause. The shrinkage of the charge on the addition of 
 liquid is from 10 to 18 per cent. The ore should be charged-in 
 as dry as possible, but a few per cent, of moisture makes very 
 little difference to the subsequent leaching. 
 
 The charge is now ready for lixiviation. If soluble salts are 
 present in it, a preliminary washing with water is required, 
 while, if decomposed pyrites are present, treatment with an 
 alkaline solution is also necessary, for reasons to be explained 
 when the chemistry of the process is considered. In each case, 
 from 7 to 9 cubic feet of liquid per ton of ore are added and 
 allowed to remain on the charge for some hours before being 
 displaced. The liquids are run on to the top of the charge 
 through pipes of 2J to 4 inches diameter, and are distributed 
 by a large box with a perforated bottom, by which the force of 
 the discharge is broken. The orifice below the filter-bed is 
 closed before beginning to add the liquids, and not opened until 
 the solution stands about 3 inches above the surface of the 
 ore, when the leaching is begun. After the caustic alkali (if it 
 is used) has been allowed to remain in contact with the ore for 
 
THE CYANIDE PROCESS. 325 
 
 a, sufficient length of time, it is displaced by wash-water, which 
 is added above, while the alkaline solution is allowed to run out 
 at the bottom. A dilute solution of cyanide is then run in, the 
 stopcock below the filter-bed being closed as soon as all the 
 wash-water has been displaced. In leaching, the charge is 
 always kept completely submerged, and the surface of the liquid 
 is never allowed to sink below the surface of the ore for reasons 
 which will be given subsequently. 
 
 The cyanide solution is often allowed to stand in contact with 
 the ore for from 12 to 24 hours undisturbed before the stopcock 
 below the filter-bed is opened and the solution drawn off. It is 
 usually conveyed at once to the zinc boxes, but it was formerly 
 preferred either to raise it and pass it through the charge again 
 (circulation method), or to transfer it to a second or even a third 
 charged vat before precipitating the gold. The advantage of these 
 " circulation" and "transference" methods is that the solutions 
 become much richer in gold than if they are only allowed to per- 
 colate through a single charge of ore, and consequently they give 
 a cleaner deposit on the zinc with much less consumption of 
 cyanide, the volume of solution passing through the precipitation 
 boxes being less. At the Mercur Mine the circulation is kept 
 up for from 24 to 240 hours, according to the speed of leaching, 
 the usual time being about sixty hours. At the Robinson Mine 
 20 tons of solution cover the ore in a 75-ton vat, and are con- 
 tinually pumped backed into the same vat for thirty-six hours, 
 .and then passed to the zinc boxes. 
 
 The " Simmer and Jack " works, South Africa, are shown in 
 Fig. 55a.* Here it was necessary to raise the tailings, and this 
 was done by hauling them up inclined platforms carried on 
 trestles. A double gangway runs over the vats for filling. 
 The leaching vats are the largest on the Rand gold-fields, each 
 holding 600 tons, being 40 feet in diameter with 14-foot staves, 
 the staves, as well as the bottom; being of 9-inch by 3-inch 
 material. There are eight discharge doors, 16 inches in diameter, 
 in the bottom of each vat. The three solution vats are 26 feet 
 in diameter, with 15-foot staves, and the liquid is returned to 
 them from the sumps by three 2-inch centrifugal pumps. 
 
 It has been thought desirable to stir the solution and ore 
 together, in certain cases, but the loss by decomposition of the 
 cyanide thus caused, and the cost of the power, are usually 
 more important than the additional amount of gold recovered. 
 The Cassel Gold Extracting Company recommend the following 
 method of procedure when agitation is used: The solution- vat 
 or agitator (A, Fig. 56) is 4 feet in diameter and 4J feet deep. 
 This is filled to a depth of from 18 to 20 inches with the solution, 
 from the vat C, and the mixer-shaft set going at about sixty 
 
 * Reproduced by permission from the paper by Butters and Smart in 
 the Proceedings of the Inst. of Civil Engineers, vol. cxx., pt. ii., 1895. 
 
326 
 
 THE METALLURGY OF GOLD. 
 
 
 
THE CYANIDE PROCESS. 
 
 327 
 
 revolutions per minute. The ore (li tons, or double the weight 
 of the liquid) is then added gradually from a hopper, B, overhead, 
 in which it has been stored, and agitation is continued until the 
 gold is dissolved. The pulp is then discharged through a 2-inch 
 iron pipe into D, the large filtering vat, which is 12 feet in 
 diameter and 4 feet deep, having a capacity of about 20 tons ; 
 here the leaching is performed as usual. This treatment is 
 recommended for some highly pyritic ores, and in all cases where 
 
 Fig. 56. 
 
 leaching is slow owing to the sliming of the ore, which is 
 especially liable to occur when much galena is present. 
 
 The power required to agitate the pulp is said to be at least 
 one horse-power per ton, and it is also stated that the ore cannot 
 be completely freed from the solution, which does not drain readily 
 from the ore after the agitation is finished, although the super- 
 natant liquid, forming about half of the solution, can of course be 
 removed by decantation. In Fig. 56, E is the vacuum boiler, 
 F the zinc box, and G the sump. 
 
 The rate of leaching may be increased by passing the liquid 
 
328 THE METALLURGY OF GOLD. 
 
 upwards and downwards alternately, or by creating a vacuum 
 below the filter-bed. The latter may be done in several ways, 
 as has already been pointed out in Chapter xiii., p. 284. The 
 use of a large boiler, in which a vacuum is created by a Westing- 
 house or other pump, is perhaps the most advantageous course ; 
 as soon as the pressure in the boiler falls to about half an 
 atmosphere, it is connected with the aperture of the vat below 
 the filter-bed. Korting's injector is recommended by Stetefeldt 
 for the production of a vacuum. The rate of leaching is often 
 doubled by the diminution of the pressure, below the filter-bed, 
 to half an atmosphere, and in some cases it is increased from 
 J or 1 inch to 7 or 8 inches of liquid (in the leaching vat) per 
 hour. In some cases, as already remarked (p. 284), the creation 
 of a high vacuum is rather a cause of delay than otherwise, the 
 ore packing down and making leaching quite impossible. The 
 author has seen an attempt to leach a roasted ore, containing 
 much slimes, by means of a vacuum of 20 inches of mercury, 
 result in the production of a compact mass which could scarcely 
 be penetrated by a pick-axe, while water stood above and could 
 not be drawn through in any way. 
 
 Strength of the Solution. In the Transvaal it is customary 
 to use two different strengths, viz., a " strong" solution, con- 
 taining from 0*25 to 0*35 per cent, of cyanide of potassium, and 
 a "weak" solution, containing from O05 to 0*15 per cent, of 
 cyanide. The strong solution is allowed to filter through during 
 a period of from eight to forty -eight hours, and the weak one is 
 then run through for twenty-four or forty-eight hours. The 
 amount of each solution used is, in general, about J ton per ton 
 of ore, but, as before stated, only about 5J cwts. of solution per 
 ton of ore is used at the Robinson Mine with the " circulation 
 system." At the Mercur Mine the strength of the solution in 
 use is 0'25 per cent. 
 
 The tendency in the Transvaal is in the direction of more 
 dilute solutions, and the average strong solution now certainly 
 contains less than O3 per cent. KCy. In 1895 Caldecott* made 
 a series of experiments on a working scale to test the dissolv- 
 ing powers of solutions of different strengths. Five similar vats 
 were charged with 34-03 tons each of fresh slightly-pyritic tail- 
 ings of uniform composition from the Jumpers' battery, con- 
 taining 100 grs. of gold per ton. The treatment that each vat 
 received was exactly the same except that the strength of 
 solution employed was different in every case. To each charge, 
 15-31 tons of solution were added, and circulated for three days, 
 after which the solution was drained off and displaced by 8 -4 
 tons of wash water. When the charges had drained dry, they 
 were sampled, the results being as follows : 
 
 * Proc. Chem. and Met. Soc. of 8. Africa, vol. L, 1896, p. 293. 
 
THE CYANIDE PROCESS. 
 
 329 
 
 
 I. 
 
 II. 
 
 III. 
 
 IV. 
 
 V. 
 
 Strength of cyanide, p.c. 
 (1) Before treatment, . 
 
 0-041 
 
 0-110 
 
 0-373 
 
 1-023 
 
 3-333 
 
 (2) After 
 
 010 
 
 068 
 
 277 
 
 860 
 
 3-020 
 
 Loss of cyanide (Ibs. per 
 
 
 
 
 
 
 ton) 
 
 34 
 
 48 
 
 1-1 
 
 1-9 
 
 3-6 
 
 Extraction, per cent. , . 
 
 83 
 
 84 
 
 83 
 
 85 
 
 83 
 
 The apparent variation in the extraction was probably due 
 to experimental error in assaying, which, it seems dimcult to 
 believe could have been much less than about 0*01 oz. per ton, 
 or about 5 per cent. The stronger solutions may have dissolved 
 the gold much more rapidly than the weaker ones did although 
 the final results were the same. The conditions under which 
 the experiments were carried out prevented this point from 
 being determined. Since, however, it takes two or three days 
 in practice to wash out soluble salts from Transvaal tailings, 
 there is little or no real advantage to be gained by dissolving 
 the gold quickly. Crosse also found* that almost equally good 
 results were obtained by treating rich Bonanza tailings with 
 solutions containing from O'l to 0'5 per cent. KCy, the time of 
 treatment being twenty-four hours. 
 
 When leaching is complete, so that no more gold is being dis- 
 solved, a point which can be determined by testing the escaping 
 solution with bright zinc shavings, the addition of fresh solution 
 is stopped, and the surface of the ore is laid bare by the sinking of 
 the liquid. There still remains in the vat, however, about 6 or 8 
 cwts. of solution per ton of ore. This might be reduced to 3 or 4 
 cw; s. by draining for some time, but the charge would then shrink 
 and crack and separate from the side walls, so that the succeeding 
 operations would be less effectual. To displace the solution, 
 therefore, water is run on to the ore surface as soon as the latter 
 emerges, and is kept running until the quantity added is equal 
 to that of the solution originally retained by the charge. Up to 
 this point the discharge from below the filter-bed is allowed to 
 run into the " stock " solution, although somewhat more dilute 
 than the main mass, owing to slight admixture with the water. 
 The operation is now stopped and the tailings may be drained 
 and discharged, but, as they still contain a little cyanide, a slight 
 saving is effected by displacing the first wash-water by a second. 
 In this case the first wash-water is run into a separate vat, and 
 used again to displace the stock solution, which is thus reduced 
 in strength less than if fresh water were used. More cyanide is 
 left in the tailings if the solution be drained off as completely 
 as possible before being displaced, as the wash-water finds its 
 way down the cracks, leaving the unbroken masses of ore less 
 efficiently freed from cyanide. 
 
 * Ibid., p. 293. 
 
330 THE METALLURGY OF GOLD. 
 
 By this practice the volume of the solution is kept constant,, 
 but its strength falls off considerably, owing to decomposition in 
 the ore and in the zinc boxes. It is tested in the " sump" to which 
 it flows from the zinc boxes, and enough cyanide added to raise 
 it to the required strength. The cyanide is usually dissolved in 
 a little water before being added to the stock solution, as the 
 amount of KCy present is more easily determined in a strong 
 solution than in any other form. At the Robinson Mine lumps 
 of cyanide are added to the liquid which is circulating through 
 the ore, as the insoluble impurities ("carbides of iron") are in 
 this way left on the surface of the charge, and are dumped with 
 the tailings instead of accumulating in the sumps. 
 
 Messrs. MacArthur & Forrest recommend the concentrated 
 solution of cyanide to be made and kept in a small vat standing 
 at a higher level than the large vats used for the storage of the 
 stock solution. Cyanide of potassium is supplied in zinc-lined 
 boxes holding 190 to 195 Ibs. of cyanide, containing about 78 per 
 cent, of available KCy.* The contents of these boxes are broken 
 into lumps and placed in wire-gauze trays, which are immersed 
 in water contained in the small vat. The soluble salts are soon 
 dissolved, potassium cyanide in particular being very rapidly 
 removed, and the tray is then lifted out, still containing the 
 insoluble material, which is thrown away. The strength of the 
 concentrated solution, which is kept to raise the stock solution 
 to its normal strength, may be from 10 to 25 per cent. It 
 should not be prepared until it is wanted, as it undergoes some- 
 what rapid decomposition. 
 
 Disposal of the Tailings. These are sometimes sluiced out of 
 the tanks by water under pressure. The sluice gates may be of 
 about 1 or 2 square feet in area, and are placed so that their lower 
 edges are level with the surface of the filter. Where the supply 
 of water is not large, or the fall of the ground is insufficient 
 for sluicing out the tailings, they are removed by shovelling out, 
 the cost being about 6d. per ton with European labour. Before 
 the vats are emptied, samples are taken, usually by means of a 
 long iron semicircular probe, shaped like a cheese-taster, which 
 is thrust to the bottom of the vat, then revolved by means of 
 the handle, and withdrawn with the tailings adhering to it. If 
 the results are high, samples from various parts of the vat should 
 be assayed separately, so as to find out if the percolation has 
 been uneven. At the Roodepoort Works, round vats have been 
 constructed, 40 feet in diameter and 8 feet deep, with a capacity 
 of 360 tons. They are placed on firm concrete foundations, 
 raised 6 feet from the ground, and the tailings are discharged 
 
 * High grade cyanide is now largely used containing 98 to 99 per cent, 
 of KCy. Sodium cyanide is also used, and this, calculated as potassium 
 cyanide, contains 125 to 130 per cent, of available KCy. 
 
 
 
THE CYANIDE PROCESS. 331 
 
 through four openings in the bottom communicating with tunnels. 
 It is said that these vats show no signs of leakage, and that they 
 can be discharged in six hours by twenty-eight natives at a cost 
 of IJd. per ton.* 
 
 At the Langlaagte Estate, the five large 450-ton vats, whose- 
 upper edges are level with the surface of the ground, are dis- 
 charged by a steam crane by which truck loads of tailings are 
 elevated and placed on a tram line. The trucks are lowered 
 into the vats and loaded up by Kaffirs, and the total cost ot dis- 
 charge is said to be 2d. per ton, including maintenance. The 
 vats and sumps are excavated in the solid ground, the walla 
 being built of brick and lined with cement. 
 
 3. Precipitation of the Gold. The precipitation of the 
 gold is effected by shavings of zinc freshly turned on a lathe. 
 In South Africa, the zinc linings to the boxes in which the 
 cyanide is imported are worked up for the purpose, a light 
 spongy mass easily traversed by the solution and presenting a. 
 large surface for precipitation being formed. The shavings are 
 about 2-i^ inch in thickness, and of no great width, but are 
 sometimes 2 or 3 feet long. They weigh about 5 Ibs. per 
 cubic foot in the boxes, and are so fine that they can be 
 ignited with a match, burning to zinc oxide. The shavings are 
 placed in wooden troughs, the "zinc-boxes;" they are divided 
 into compartments by partitions, which cause the solution to- 
 flow alternately upward and downward. Each alternate com- 
 partment is empty, in order that, in passing through the shav- 
 ings, the solution may invariably flow upwards. " It is arranged 
 so, because, if the solution were to flow down, the gold slimes 
 would collect on the upper surface of the zinc, and impede fur- 
 ther flow, but by upward flowing the gold slimes are precipitated 
 on the under surface of the zinc, from which they continually 
 drop off, and permit free passage of the solution." 
 
 At the Robinson Mine, the shavings are supported on wire- 
 screen trays, so that the finely-divided, precipitated gold falls 
 through to the bottom of the trough, while the unaltered zinc 
 remains on the sieve. At this mine the boxes are 20 feet long 
 and 2 feet square in cross-section ; there are ten compartments 
 in each, and about 40 Ibs. of shavings in each compartment, 
 except those at the head and foot, which are left empty in order 
 to allow the slimes to settle. About 30 tons of solution pass 
 through each box in a day of nine hours, so that every part of 
 the solution is in contact with the zinc for about twenty-five 
 minutes. The solution contains from 1 to 3 ozs. of gold per ton, 
 and, after passing through the boxes, contains from ^ to 2 dwts. 
 per ton. There are two sets of boxes, one for the "strong" 
 solution, the other for the " weak." In the latter, less zinc is 
 consumed, but the gold-slimes are poorer, less gold being con- 
 * Eng. and Mng. Journ., Mar. 25, 1893, p. 273. 
 
332 THE METALLURGY OF GOLD. 
 
 tained in the " weak " solution. The total amount of zinc 
 -consumed is 100 Ibs. per day, the gold extracted being about 
 120 ozs. At the Mercur Mine, the boxes are 40 feet long and 
 12 inches square in cross-section, with compartments 3 feet long, 
 and contain zinc enough to maintain the contents of the last 
 compartment quite bright. The consumption here is about 1 Ib. 
 of zinc for 1 oz. of gold recovered. 
 
 When the solution comes in contact with bright zinc, the 
 latter turns black at once, owing to the deposition on it of finely 
 divided gold. The zinc is gradually dissolved, and the shavings 
 fall to pieces, those in the first compartment being consumed 
 most rapidly. As the precipitation proceeds, the zinc is trans- 
 ferred from the lower compartments to the upper ones, and fresh 
 zinc is added at the foot of the box. 
 
 Clean-up. The clean-up usually takes place either once a 
 week or twice a month. The screens containing the undissolved 
 shavings are lifted out of the zinc boxes, and the zinc-gold slimes, 
 remaining at the bottom of the boxes, are allowed some time to 
 settle. The clear liquor is then siphoned off and the slimes 
 allowed to run out by withdrawing a plug in the bottom of the 
 box, and drained through a 40-mesh screen, which retains a 
 small part only. After the residue has been rubbed on the 
 screen with sticks tipped with indiarubber, and well washed, it 
 is put back again into the first compartment of the box, on the 
 top of fresli shavings, as it consists mainly of small pieces of 
 unconsumed zinc. The subsequent fineness of the bullion 
 greatly depends on the care with which these operations are 
 conducted. By careful manipulation most of the undissolved 
 .zinc can be separated in this way. The slimes proper are now 
 washed on a filter, and carefully and slowly dried in enamelled 
 iron pans. The richness of the dried precipitate depends on the 
 strength of the cyanide solution, and on the time of contact as 
 well as on the quantities of base metals which are present in 
 the solution. By the prolonged action of the cyanide, the zinc 
 shavings become partly corroded and disintegrated, so that the 
 precipitated gold is mixed with zinc debris. Ordinary commer- 
 cial zinc contains a considerable proportion, generally over 1 per 
 cent., of lead, and a small quantity ol carbon, besides other im- 
 purities, such as arsenic and antimony. All these accumulate 
 ,nd are collected with the gold. The precipitated gold and the 
 carbon from the zinc are invariably in a state of fine division, 
 and, by using a fine sieve, the gold can be very closely separated 
 from the zinc, but, as the moderately fine particles of zinc still 
 retain a considerable proportion of gold entangled, it is not 
 always desirable to use a very fine sieve. If the cyanide solu- 
 tion contains copper, it is found with the gold. 
 
 According to Messrs. Butters arid Clennel,* the pans in use 
 * Eng. and Mng. Joum. y Oct. 15, 1892, p. 365. 
 
THE CYANIDE PROCESS. 
 
 33? 
 
 at the Robinson Works contain about 5 or 6 gallons of dried 
 precipitate (often called " gold slimes ") which may contain a& 
 much as 150 ozs., or as little as 20 ozs., of gold. A little silver is 
 also contained in it, the remainder being chiefly zinc and lead,, 
 with smaller quantities of tin, antimony, organic matter, &c. 
 The average in South Africa is said to be approximately : 
 
 Gold and silver, , , , -, . *. v , . 20 to 50 per cent. 
 
 Zinc,. . . ; Y" . 30 to 60 
 
 Base metals, silica, alumina, oxides 
 
 of iron, &c. , '; i i ; ' ! i '-. * 20 to 40 
 
 A carefully prepared sample of gold slimes which was obtained 
 at one of the African works contains : 
 
 Gold, . . , 
 
 Silver, 
 
 Base metals and carbon, 
 
 60 per cent. 
 10 
 30 
 
 This is by no means average, but is exceptionally good, and 
 must not be taken as showing what is usual. The Hanauer 
 Smelting Company found the precipitate from the Mercur Mines 
 to contain the following substances : * 
 
 Zinc, 
 
 Carbonate of lime, 
 
 Gold, . 
 
 Cyanogen, 
 
 Sulphur, . 
 
 Iron, 
 
 Residue, . 
 
 39 '1 per cent. 
 
 36-7 
 4.4 
 
 3-5 
 2-6 
 2-4 
 6-0 
 
 94-7 
 
 The carbonate of lime present is probably deposited in the boxes 
 chiefly from suspension in the solution, the gangue of the ore 
 mainly consisting of limestone. 
 
 4. Production of Bullion from the Precipitate. In the 
 United States, most mines will, no doubt, find it convenient to- 
 follow the example of the Mercur Mine and sell the dried 
 precipitate to smelting companies, which pay about $20 per oz. 
 for the gold contents. In South Africa this course cannot be 
 pursued, and various methods have been suggested for effecting 
 the elimination of the zinc and other base metals. The usual 
 course is to melt the mass in graphite or clay crucibles, using as 
 fluxes sand, borax, nitre, and carbonate of soda. The slag, which 
 consists of silicates of zinc, soda, &c., corrodes the pots rapidly. 
 Large quantities of zinc oxide are given off as fumes, forming thick 
 crusts in the flues, ai^d doubtless taking some gold with them, and 
 evil-smelling products of decomposition of the cyanides are also 
 evolved. The bullion produced varies in colour from a pale 
 yellow to a brownish linnet-green, and is about 650 fine, but 
 
 * Eng. and Mng. Journ., Nov. 5, 1892, p. 440. 
 
334: 
 
 THE METALLURGY OF GOLD. 
 
 cannot be obtained uniform in composition, so that accurate 
 assays are difficult to obtain. 
 
 The results of analyses made on three ingots of bullion pro- 
 duced by the Mac Arthur-Forrest process in South Africa, and 
 shipped to London, are appended : 
 
 
 I. 
 
 II. 
 
 III. 
 
 Gold, 
 
 60-3 
 
 61-7 
 
 72-6 
 
 Silver, 
 
 7-0 
 
 8-1 
 
 92 
 
 Zinc, 
 
 15-0 
 
 9-5 
 
 7-1 
 
 Lead, 
 
 7-0 
 
 16-4 
 
 4-9 
 
 Copper, 
 
 6-5 
 
 4-0 
 
 4-8 
 
 Iron, 
 
 2-2 
 
 0-3 
 
 1-4 
 
 Nickel, 
 
 2-0 
 
 
 
 
 100-0 
 
 100-0 
 
 100-0 
 
 The slags obtained in this way are always rich in gold, 
 part of which is sometimes in the form of shots, and may be 
 recovered by crushing and panning. The crushed slag is then 
 fused again with the addition of granulated lead, or better still, 
 of litharge, when all the gold is concentrated in the lead. If 
 the lead thus obtained is granulated, it can be used again until 
 rich enough to be worth cupelling. 
 
 If the slimes are roasted in a muffle furnace, some of the zinc is 
 volatilised, and a bullion 800 fine can thus be obtained, according 
 to Butters and Clennel. The separation of the zinc by solution 
 in sulphuric acid is by far the best course to pursue in spite of 
 the difficulty of filtering the slimes, and the tendency of the 
 vessels in which the solution is effected to boil over Irom the 
 copious evolution of hydrocyanic acid. J. S. MacArthur points 
 out that the difficulty of filtering is due to the presence of 
 soluble cyanide salts containing zinc. By the action of acids, 
 cyanide of zinc (ZnCy 2 ) is set free and precipitated, forming a 
 gelatinous mass which cannot be filtered, and is not readily 
 dissolved by solvents other than potassic cyanide. This sub- 
 stance is also very troublesome in the crucible, rendering 
 the slags pasty and rich. If, however, the slimes in their 
 original condition, while still alkaline, are filtered, whether in 
 a filter-press or not, the soluble cyanide salts may readily be 
 separated by washing, and the greater part of the zinc thus 
 removed in solution. The residue is then spread in a thin la^er 
 on iron trays, and dried and heated to a barely-perceptible red 
 heat in the flue of a furnace, The carbon, zinc, arsenic. &c., 
 ignite readily, being in a very fine state of division, and 
 roasting proceeds regularly through the mass to the bottom 
 without any stirring, leaving a completely oxidised porous mass, 
 
THE CYANIDE PROCESS. 
 
 335 
 
 aggregated more or less into granules. This residue contains no 
 carbon and little arsenic, and consists chiefly of oxides of zinc 
 and lead, with metallic gold and silver. It is rapidly acted on 
 by dilute sulphuric acid, whether it is heated or not, which 
 dissolves the zinc, and forms insoluble sulphate of lead. There 
 is less tendency for the mixture to boil over, than is the case 
 when cyanides are present, and the residue is easy to filter, the 
 granular sulphate of lead offering little resistance. When the 
 residue is washed and dried, it contains very little zinc if the 
 operations described above have been properly carried out. It 
 is more usual, however, to treat the washed slimes with acid 
 before roasting. The following analyses of slimes, after treat- 
 ment with acid in this way at the Brodie works, Cripple Creek, 
 Colorado, are given by Mr. W. R. Ingalls.* 
 
 
 I. 
 
 II. 
 
 III. 
 
 Gold, 
 
 49-85 
 
 26-18 
 
 23-60 
 
 Silver 
 
 9-54 
 
 6-84 
 
 6-00 
 
 Silica, 
 
 15-60 
 
 10-00 
 
 6-60 
 
 Lead, 
 
 3'00 
 
 Trace 
 
 Trace 
 
 Copper oxide, 
 Zinc oxide, .... 
 
 5-58 
 014 
 
 8-1) 
 31-80 
 
 6-40 
 41-50 
 
 Lime 
 
 0-51 
 
 5-32 
 
 0-03 
 
 Ferric sesquioxide, \ 
 Alumina, . . / ' 
 
 4-00 
 
 6-80 
 
 9-26 
 
 Sulphur trioxide, . 
 
 11-32 
 
 4-80 
 
 6-45 
 
 
 99-54 
 
 99-85 
 
 99-84 
 
 In Nos. II. and III. the treatment with sulphuric acid was 
 very incomplete. In fusing these slimes, No. 1 would require 
 about 50 per cent, of soda carbonate and 25 per cent, of borax, 
 but silica in addition would be required by the others. The 
 charge is melted in graphite crucibles, and poured into conical 
 moulds. The slag is fluid and glassy and contains from 20 to 
 200 oz. of gold per ton, which is recovered by fusion with 
 litharge. The bullion prepared in this way may be as high as 
 985 fine in gold and silver t and should not be lower than 850 
 fine. It is often brittle from the presence of zinc. The loss in 
 roasting and subsequent handling is estimated by Bettel to be 
 from 0-01 to 0-025 per cent., or less than Jd. per oz. 
 
 The method of cleaning-up in use at the Princess Works, 
 Transvaal, is described by E. H. Johnson, as follows :| The 
 slimes are transferred direct in buckets from the zinc boxes to a 
 filter vat, 6 feet in diameter, and filtered by a vacuum-pump 
 
 * Mineral Industry for 1895, p. 338. t Loc. cit. 
 
 J Chem. and Met. floe, of S. Africa, June 19, 1897. 
 
336 THE METALLURGY OP GOLD. 
 
 through closely- woven canvas, the clear liquid being returned to- 
 the boxes. Water is then added and pumped through until all 
 soluble cyanides have been removed, and the slimes are then 
 baled out (being weighed in the buckets during the process) into 
 a sheet-iron tray. A closed vat, fitted with a stirring apparatus, 
 is then charged with dilute (1 in 10) sulphuric acid, one pound 
 of concentrated acid being added for each pound of moist slimes, 
 the agitator is started, and the slimes added little by little 
 through a hopper, the fumes being carried away by a 3-inch pipe. 
 After all the slimes have been charged in, the agitation is con- 
 tinued for half an hour, and the hopper and sides are then 
 washed down, the agitator washed and removed, and the vat 
 filled with water and allowed to settle. No heat is used, except 
 that caused by the mixing of strong sulphuric acid with water 
 in the vat and the action on the slimes. The clear liquid is 
 syphoned off, and washing by decantation repeated four or live 
 times, the washings containing about 13 grains of gold per ton 
 of solution. The gold residues are then dried in enamelled iron 
 dishes, roasted gently in sheet-iron trays, and then ground, 
 mixed with fluxes, and transferred to the crucible. The charge 
 fuses quietly with little fume, yielding bullion amounting to* 
 50 to 60 per cent, of the weight of the slime. The average fine- 
 ness in 1896 was 821-9, and the slags assay 23 ounces per ton, 
 but only 0-2 per cent, of the gold is locked up in this way.. 
 The cost of reduction in an actual case was as follows : 
 
 Dry vrcight of zinc-gold slimes, ., 
 after acid treatment, . 
 
 672 Ibs. of acidat4d., . 
 66 ,, borax at 37s. 6d. per cwt., 
 9 ,, soda carbonate at 2^d., 
 9 ,, fluor spar at 4d., . 
 
 
 504 Ibs. 
 100 
 
 12 12 
 1 2 1 
 1 10 
 030 
 226 
 
 1 No. 60 crucible, 
 
 
 1 7 6 
 
 17 9 9 
 
 Yield 620 oz. fine gold. Cost 6'7d. per fine oz., without cost of labour. 
 When a filter-press is used, less gold is lost in the washing, 
 and by the use of hot water, the zinc sulphate is more perfectly 
 removed, so that the subsequent smelting is more easily and 
 cheaply executed.* The amount realised by the sale of cyanide 
 gold in the Transvaal is stated to be as follows : f 
 
 Fineness Value per oz. 
 per 1000. fine gold. 
 
 850, . . 84s. 5d. 
 
 800, . . 84s. 2d. 
 
 700, . . 83s. lid. 
 
 600, . . 83s. 8d. 
 
 Fineness Value per oz. 
 
 per 1000. fine gold. 
 
 500, . . 83s. 3d. 
 
 400, . . 83s. Id. 
 
 300, . . 82s. 9d. 
 
 * Trans. Inst. Mng. and Met., vol. v., 1897, p. 147. 
 
 t Chem. and Met. Soc. of S. Africa, Oct. 17, 1896, p. 3L 
 
THE CYANIDE PROCESS. 337 
 
 An average of about one shilling per ounce is consequently 
 deducted from the value of bullion recovered by zinc, whilst 
 gold produced by the Siemens-Halske process is subjected to a 
 deduction of Is. 5|d. per oz. by the smelters who buy the base 
 lead bullion. 
 
 Plant Required. For the treatment of 2,000 tons of tailings 
 per month, Messrs. Mac Arthur & Forrest consider that six cir- 
 cular vats each 19 feet in diameter and 4 feet deep are sufficient. 
 Each vat holds about 40 tons of ore, and must be charged, 
 leached, and emptied in three days. The other requirements of 
 this plant are two large vats, each of 2,000 cubic feet capacity, 
 to hold the strong and weak cyanide solutions, and a similar vat 
 to hold the soda solution. If lime in the solid state is mixed with 
 the ore before it is charged into the leaching vat, this last tank 
 may be dispensed with. Zinc boxes, sumps, pumps, pipes, and 
 launders are also required, and furnaces to fuse the zinc residues. 
 The amount of water required for this plant is about 2,000 to 
 3,000 gallons per day, or 30 to 45 gallons per ton of ore. The 
 labour needed consists of eight men for a shift of twelve hours. 
 A zinc box 14 ieet long, 2 feet deep, and 1 foot wide, divided 
 into six or eight compartments, and holding 100 Ibs. of zinc, 
 should suffice to treat in about nine hours the gold solution 
 obtained each day. 
 
 A plant capable of treating 20 tons of pyritic ore by means of 
 agitation, consists of six agitator- vats (each holding a charge of 
 1 J tons of ore), three large filtering vats, and two vacuum boilers 
 to assist in the filtering. The discharge from the agitators to the 
 filters is by means of pipes which are at least 2 inches in diameter. 
 Pumps are required to raise the liquid from the sump to the stock 
 solution vat, and to create a vacuum in the boiler. The power 
 required for the agitators and pumps is about 6 H.P. The con- 
 sumption of water is from 600 to 1,000 gallons per day, or from 
 30 to 50 gallons per ton of ore. Labour consists of three men 
 for a shift of eight or twelve hours. 
 
 Treatment of Ore Slimes. On the Rand, there have been 
 accumulated hundreds of thousands of tons of slimes which 
 could not be treated until lately, owing to its impermeability, 
 although the gold contained in fresh slimes is readily soluble in 
 the laboratory. The average value of the Robinson slimes is 
 between 7 and 8 dwts. of gold per ton, and the fineness is such 
 that it would pass through a 225-mesh screen. It has been 
 proposed by Bettel to treat this material in Johnson's filter- 
 presses. The slimes were mixed with a very dilute (O'Ol per 
 cent.) cyanide solution and thoroughly agitated, and then filtered, 
 and water forced through under a pressure of 100 Ibs. to the 
 square inch. This pressure leaching is similar to that sug- 
 gested by E. N. Riotte in connection with the hyposulphite 
 lixiviation process (see p. 285). It is stated that about 98 
 
 22 
 
338 THE METALLURGY OF GOLD. 
 
 per cent, of the assay value of the slimes can be extracted in 
 this way, but the method has never been applied on the large 
 scale. 
 
 The following method of treating slimes, devised by Mr. 
 John R. Williams,* has been at work at the Crown Reef Gold 
 Mining Company's Mill since August, 1896, and is working 
 satisfactorily : To the water carrying the slimes from the 
 separating plant, milk of lime is added, and the slimes thus 
 precipitated in a flocculent form. An automatic lime feeder is 
 used, as too much lime is as bad as too little (a large excess of 
 lime preventing the subsequent precipitation of the gold), and 
 regularity of feeding is therefore essential. Three large pointed 
 boxes, two of them 20 feet x 20 feet and 10 feet deep, and the 
 third 40 feet x 40 feet and 10 feet deep, are used for settling the 
 slimes, which are then drawn off at the bottom and pumped into 
 the first two of the eight treatment vats, about 90 per cent, of the 
 water contained in them having been separated. The vats are 
 each 32 feet in diameter and 10 feet deep, having a conical 
 bottom. More water having been separated from the slimes by 
 allowing them to settle, they are sluiced into a pump by a jet of 
 cyanide solution, and transferred to the second series of two 
 vats. These vats are filled with solution, the strength of which 
 is brought up to 0-01 per cent. KCy. About 80 per cent, of the 
 gold is dissolved in the passage through the pumps, but agitation 
 is continued for from one to two hours by withdrawing the solu- 
 tion at the bottom and discharging it in oblique jets at the 
 top and through the sides. The slimes are then allowed to 
 settle, and the clear solution drawn off through side-cocks 
 (which have been advantageously replaced at the Rand Central 
 works by syphon pipes) and passed to the precipitation boxes. 
 The slimes are then pumped successively into the third and 
 fourth series of two vats, where they are further agitated with 
 very dilute solutions of cyanide, and allowed to settle. These 
 solutions are not passed to the precipitation boxes, but are 
 transferred to the preceding series of vats, those going to the 
 second series having cyanide added to it to bring up its 
 strength. 
 
 The " strong " solution from the second series of vats is run 
 into two settling tanks 15 feet in diameter and 5 feet deep, 
 where it is allowed to clarify, the tanks being used alternately 
 (as in the case of each of the other sets of vats), one being filled 
 up, while the liquid from the other is clarifying by settlement 
 and passing to the precipitation plant where eight tons of solu- 
 tion per hour are treated. Electrical precipitation is used, and 
 there are four boxes, each 30 feet x 6 feet and 4 feet 9 inches 
 deep. The iron anodes are set at 6-inch centres, and between 
 * Chem. and Met. Soc. ofS. Africa, July 17, 1897. 
 
THE CYANIDE PROCESS. 
 
 339 
 
 ach pair is placed a double wire frame, carrying on each wire 
 three lead sheets cut into strips. The solution, which contains 
 60 grains of gold per ton, passes upwards through these, and 
 thence to the sump for use over again. The liquid takes about 
 12 to 14 hours to pass through the boxes, and the rate of preci- 
 pitation to be expected may be judged from the following 
 experimental trial : 
 
 No. of Test. 
 
 Time of Precipitation. 
 
 Assay of Liquid. 
 
 
 
 Per Ton. 
 
 1 
 
 Before treatment. 
 
 2 dwt. 12 gr. 
 
 2 
 
 After 3 hours- 
 
 1 
 
 4 
 
 3 
 
 
 6 
 
 
 
 20 
 
 4 
 
 
 9 
 
 
 
 
 12 
 
 5 
 
 
 12 
 
 
 
 
 8 
 
 6 
 
 
 15 
 
 
 
 
 4 
 
 7 
 
 
 18 
 
 
 
 
 2 
 
 8 
 
 
 21 
 
 
 o 
 
 4 ,, 
 
 9 
 
 
 24 
 
 
 Trace. 
 
 About 90 per cent, of the gold is precipitated in practice accord- 
 ing to this table. 
 
 The cost of treatment for the months of April and May, 1897, 
 was as follows : 
 
 Management, assaying, and labour, 
 
 Power, 
 
 Cyanide, . . . . 
 
 Lime, . . . . , . 
 
 Lead foil, V ' . 
 
 Sundries, . . . _ , 
 
 Royalty, . . . . . . 
 
 Maintenance, . 
 
 13'99d. per ton. 
 
 2-75d. 
 
 5'51d. 
 
 6'19d. 
 
 l'75d. 
 
 l'&2d. 
 
 4-77d. 
 8'17d. 
 
 Total, 
 
 3s 9'05d. 
 
 The results obtained were as follows : Tons treated (?), 6643 ; 
 assay value before treatment, 5-32 dwts., after treatment, 0*892 
 dwts. ; theoretical extraction, 83*6 per cent.; fine gold recovered, 
 1067-16 ozs. ; actual extraction, 60-5 per cent. ; profit, 3058, 
 11s. 2d. 
 
 The difference between the theoretical and actual extraction 
 is due to the great difficulty in estimating the tonnage of slimes 
 treated. The clean-up takes place every two months. The 
 amount of lime used is large, and is in proportion to the amount 
 of water and not to that of slimes, so that if, as would be the 
 
340 THE METALLURGY OF GOLD. 
 
 case at some mills, there was double the quantity of slimes in 
 the water, the same amount of lime would precipitate them. 
 
 The Crown Reef Company, by means of its stamp mill, cyanide, 
 and slimes plant, is recovering fully 90 per cent, of the total 
 gold contents of the ore at a cost of under 6s. per ton of original 
 ore. 
 
 The Treatment of Accumulated Slimes.* Although the 
 gold in slimes fresh from the battery is very readily soluble in 
 cyanide solution, it is quite otherwise with slimes which have 
 accumulated in dams and settling pits and have been exposed 
 to the weather for some time. The presence in these materials 
 of finely-divided ferrous sulphide (see p. 366), ferrous hydrate, 
 and other ferrous salts and decomposing organic matter gives 
 rise to the abstraction of oxygen from the cyanide solutions 
 used in their treatment, and the solution of the gold is in this 
 way absolutely prevented. Even in fresh slimes some ferrous 
 sulphide may be present produced during the crushing in the- 
 battery, as Caldecott has shown that this substance is produced 
 by triturating pyrites in a mortar. The obvious remedy for 
 this difficulty is to supply oxygen artificially, either in the form 
 of air delivered from a perforated pipe fixed near the bottom of 
 the agitation vat containing slimes, or in the form of an oxid- 
 ising agent, the best and cheapest of which is found by Caldecott 
 to be potassium permanganate. Aeration of the pulp in the 
 dissolving vats has now been regularly used since the end of 1896 
 at the Rand Central Ore Reduction Co.'s Works, accelerated 
 when this operation takes too long, or when much organic matter 
 is present with | to \ Ib. of permanganate per ton of dry slimes. 
 All slime plants now building will be supplied with large air 
 compressors, delivering up to 1800 cubic feet of air per minute. 
 Vats are made very deep, and the air is delivered in small 
 bubbles to get the greatest efficiency from the air. It has also- 
 been suggested that when aeration is employed, no other form 
 of agitation is necessary. The air is supplied as rapidly as 
 possible (to lessen the time occupied), as the ferrous sulphide 
 is found to require only a certain amount of oxygen and to 
 absorb this very quickly. 
 
 Old accumulated slimes are often highly acid, a consumption 
 of from 8 to 20 Ibs. of lime per ton being not uncommon. 
 Ferrous hydrate, produced by the action of lime on the ferrous 
 sulphate, either existing originally in the slimes or produced by 
 aeration, tends to abstract oxygen from the solution and also to 
 form ferrocyanide with KCy, and hence the ferrous hydrate is 
 carefully oxidised by aeration to form ferric hydrate. After 
 
 *See W. A. Caldecott's paper in Proc. Chem. and Mtt. Soc. of 8. Africa,. 
 July 17, 1897. 
 
THE CYANIDE PROCESS. 341 
 
 this is done the consumption of the cyanide in the dissolving 
 vats is only about Ib. per ton, even with the strongly-acid 
 accumulated slimes, another J Ib. per ton being decomposed in 
 the boxes or thrown away with the residues. 
 
 The method of treatment of the slimes at the Rand Central 
 "Works is as follows : The truck-loads of slimes are dumped 
 into a small constant discharge " pulping " vat containing fast- 
 running paddles. A pipe delivers a continual stream of the 
 weak solution at the bottom of the vat near one side, and at the 
 top on the other side is an overflow launder, through which the 
 slime-pulp continually discharges, as the lumps become disinte- 
 grated, into the dissolving vat. Here it is stirred with paddles 
 and aerated by air delivered from an air-compressor through a 
 perforated pipe fixed near the bottom of the vat, until ferrous 
 sulphide is no longer present in the pulp, the test being the 
 non-blackening of lead acetate paper held above a portion of 
 the pulp after the addition of sulphuric acid. The aeration 
 may take from two to twenty hours, and varies with every 
 charge. After the sulphides are oxidised, an hour or two 
 longer aeration is given to oxidise any ferrous hydrate remain- 
 ing, and cyanide is then added, the solution containing from 
 0-005 to 0-008 per cent. KCy, or about 2 oz. KCy to the ton of 
 water. During the charging and aeration lime is added, raising 
 the alkali strength of the solution to from 0-006 to 0*010 per 
 cent., an amount quite enough to expedite the subsequent 
 settlement of the slimes. Permanganate is also added, especially 
 if organic matter is present, as soon as the vat is full. 
 
 After the addition of cyanide the agitation and aeration is 
 kept up for from two to five hours longer, as experience dictates, 
 and the slimes are then transferred to another vat, diluted and 
 allowed to settle, the subsequent operations being similar to 
 those described in the previous section. 
 
 The percentage of dry slimes in the pulp is determined with 
 fair accuracy from its density x by the formula 
 
 Percentage of dry slimes = 158 ' 8 (a?-1) 
 
 also, 
 
 Slimes : solution = (x-1) : (l-'37a;) 
 
 Testing of Ores. Experiments with small quantities of 
 material in the laboratory will usually determine the maximum 
 extraction that can be looked for. The weight of ore taken may 
 be from 100 to 200 grammes. It should be digested in a beaker, 
 with sufficient solution to form a thin mud, with occasional 
 stirring, or better still, in a funnel or lamp glass, through which 
 the solution slowly passes. The solution is then separated} 
 
342 THE METALLURGY OP GOLD. 
 
 and the residue thoroughly washed by filtration, and assayed 
 to find how much of the precious metals has been removed. 
 The consumption of cyanide may be determined by titrating 
 the solution before and after use. In this way, the maximum 
 extraction of gold and silver from an ore which has been crushed 
 to different degrees of fineness, with solutions at different 
 strengths, acting at various temperatures and for different 
 lengths of time, may be determined. Such data will be of great 
 value in determining the degree of suitability of the process to 
 any given ore, but the results obtained cannot usually be 
 repeated on the large scale. Mechanical difficulties preclude 
 fine crushing beyond a certain point : the relative amount of 
 solution used in practice must be less than the large quan- 
 tities giving the best results on a small scale. Moreover, tho 
 washing and filtering is more perfectly performed in the labora- 
 tory than in the mill, but, on the other hand, the consumption 
 of cyanide may be less in the latter case than in the former. 
 
 Preliminary tests as to the amount of alkali which it is neces- 
 sary to add, and the amount of cyanide decomposed are made as 
 follows : (a) Alkali. 200 grammes of ore are made into a paste 
 with water in a porcelain basin, and tested with litmus paper. 
 If an acid reaction is observed, a titrated solution of caustic soda 
 is run in, little by little, until the mixture is neutral. It is 
 convenient to make the alkaline solution of such a strength that 
 each c.c. corresponds to one pound of caustic soda per ton of the 
 ore which is being tested. (6) Decomposition of Cyanide. 
 100 grammes of the ore is shaken in an 8-oz. stoppered bottle 
 with 50 to 100 c.c. of a solution of KCy for fifteen minutes, then 
 left undisturbed for twelve hours, and finally shaken again for 
 five minutes. The available cyanide in the separated solution is 
 then estimated and compared with that in the original solution. 
 By merely shaking for a few minutes, most of the decomposition 
 of cyanide is made to take place, but not the whole. The 
 strength of solution used may be from 0*1 to 0*5 per cent, of 
 KCy according to the strength required for dissolving the gold. 
 In general, the stronger the solution the greater the amount of 
 cyanide decomposed. If soluble salts are present in the ore, 
 they should be removed by washing with water before applying 
 the cyanide, and the effect of a soda solution in saving cyanide 
 should also be tested. 
 
 The Oassel Gold Extracting Company supply a small testing 
 plant capable of treating 3 cwts. of ore at one time, under 
 conditions closely resembling those obtaining in a good mill. 
 The plant consists of a small revolving barrel for solution, a 
 filtering vat and vacuum boiler, and a zinc box for precipitation. 
 The time of treatment in the barrel varies from six hours 
 upwards. Such a barrel is at the Royal College of Science. 
 
 Direct Treatment of Band Ore. Several attempts have 
 
THE CYANIDE PROCESS. 343 
 
 been made to treat auriferous banket by the cyanide process 
 alone, without first amalgamating it. In general, however, these 
 have been unsuccessful. The coarser particles of gold, usually 
 removed on the copper plates, require far too much time for 
 their complete dissolution to admit of a high percentage of 
 extraction being attained. The pyritic ores have not yet been 
 proved to be amenable to this method, but with certain oxidised 
 ores from near the surface of the ground, all the gold is in a 
 very finely-divided state, and 75 to 80 per cent, is extracted by 
 cyanide in two or three days. At the George and May Mining 
 Company's works, the oxidised banket is very friable and easily 
 disintegrated, and this was crushed dry, and treated direct with 
 complete success in 1895. Here nearly half the ore was con- 
 verted into intractable slimes if it was crushed in the battery, 
 and in the last few months of 1894 only 55 per cent, of the gold 
 had been extracted by amalgamation, followed by treatment 
 with cyanide. In the method now adopted the ore is crushed 
 dry by a Gates crusher, and dumped at once into the vats, where 
 it is treated as follows (Hatch & Chalmers) : The ore is saturated 
 with a solution containing O06 per cent, of KCy, and this is 
 displaced by a "strong solution" containing 0*28 percent.; a 
 "weak" solution containing (HO per cent, of cyanide follows, 
 and is washed out with one containing 0'05 per cent. "About 
 sixty hours are required for the whole treatment. Some lime 
 is mixed with the ore, and on an average 0*89 Ibs. of cyanide of 
 potassium are used per ton of ore treated." The ore is so friable 
 that it is reduced to barren pebbles and loose sand by merely 
 passing through the rock breaker. About 75 per cent, of the 
 gold is extracted, the ore containing about 5 dwts. per ton. 
 The total cost of mining and treatment is about 12s. per ton. 
 
 Use of Cyanide in the Stamp Battery. Efforts have been 
 made in South Africa and the United States to treat the ore 
 direct from the battery instead of first passing it over amalga- 
 mated plates. Ore from the May Consolidated Mine, Johannes- 
 burg, was crushed by the African Gold Recovery Company in 
 the latter part of the year 1892,* with cyanide solution instead 
 of water, and led at once into the filtering tanks. The results 
 are stated to have shown that the coarse gold resisted the attack 
 of the cyanide for so long a time as to render the process 
 uneconomical. No doubt this system will meet with great 
 success in cases where all the gold is in a finely divided 
 condition. 
 
 At the Stewart Mine, Bingham, Utah, a combination of the 
 amalgamation and cyanide processes is said to have been applied 
 successfully to ore containing a mixture of coarse free gold and 
 fine rebellious particles. f The ore is crushed and amalgamated 
 
 * Eng. andMng. Journ., Oct. 8, 18S2, p. 342. 
 April 15, 1893, p. 339. 
 
344 THE METALLURGY OF GOLD. 
 
 in Huntington and Crawford mills successively, the water used 
 being a solution of cyanide of potassium. After leaving the 
 mills the solution and pulp are run into settling tanks and the 
 liquid drawn off. It would appear likely that the decomposition 
 of the cyanide in the mills would be excessive, but no details of 
 the working have been published. 
 
 Method of Treatment at the Robinson Mine. At the 
 Robinson Mine* there is a 60-stamp battery, pulverising 280 
 tons per day through a 40-mesh sieve, the ore containing nearly 
 2 ozs. gold per ton. Of this amount 71*04 per cent, is caught on 
 the amalgamated copper plates, 5-13 per cent, is extracted from the 
 concentrates at the chlorination works, and 18*55 per cent, from 
 the tailings at the cyanide works, while 5 -28 per cent, is lost. 
 The pulp from the battery is concentrated, and the tailings caught 
 in pits and shovelled into cars and conveyed to the cyanide 
 works, while the concentrates go to the chlorination works. 
 There the concentrates are placed in heaps, assayed for sulphur, 
 and mixed so as to make a uniform product for roasting. They 
 are roasted in three furnaces, each 60 feet long, treating 600 to 
 800 tons per month. When roasted, the material is charged 
 into ten circular chlorination vats with bottom discharge and 
 luted rings and lids. The chlorination occupies four days, 6 Ibs. 
 of chlorine being used for each ton of concentrates. The gold is 
 precipitated by a solution of ferrous sulphate, the mass being kept 
 agitated by means of a jet of compressed air. The consumption 
 of ferrous sulphate is from 10 to 15 Ibs. per ton of concentrates. 
 The gold is allowed to settle for some time in a series of large 
 vats, and a complete clean-up takes place only four times a year. 
 
 At the cyanide works there are twelve vats, each holding 75 
 tons. The treatment here also occupies four days, the output 
 being 225 tons per day. The whole of the tailings are subjected 
 to treatment, the coarse gold having been removed on the plates, 
 and the pyrites by concentration, so that almost clean sand, 
 containing from 8 to 10 dwts. of finely divided gold, is left for the 
 action of the cyanide. Of this from 1 to 2 dwts. are left in the 
 tailings, the extraction being over 70 per cent. 
 
 The Cyanide Process at the Sylvia Gold and Silver 
 Mining Company, Thames Valley, New Zealand. f The 
 ore from the deeper levels of the mine contains a high percentage 
 of complex minerals, consisting of galena, zinc-blende, copper, 
 and iron pyrites, and does not, as a rule, show any visible gold. 
 It is crushed in a battery and passed over amalgamated Muntz- 
 metal plates, by which the coarse gold is extracted, the amount 
 thus saved being about 5 dwts. per ton. The ore is then 
 
 * B. H. Brough, Journ. of the Soc. of Arts, vol. xli., 1893, p 173. 
 + This description is taken chiefly from the reports of the manager, 
 Dr. A. Scheidel. 
 
THE CYANIDE PROCESS. 345 
 
 carefully sized and concentrated by a complicated system of jigs, 
 rotary tables, buddies, and sizers. The concentrates obtained are 
 classed as follows : jigger concentrates, first-class slime concen- 
 trates, second-class slime concentrates, and buddle concentrates. 
 The jigger concentrates contain 4 ozs. 5 dwts. of gold and 20 ozs. 
 of silver per ton, and consist chiefly of iron and copper pyrites 
 and zinc blende, with a little galena and quartz. The galena in 
 the ore carries greater amounts of gold and silver than the other 
 minerals. It forms slimes for the most part, and is found in 
 the slmie concentrates. Of these the first-class contain about 
 20 per cent, of galena, and an average of 10 ozs. 6 dwts. of gold 
 and 44 ozs, of silver per ton : the second-class concentrates contain 
 very little lead, and assay 4 ozs. 5 dwts. of gold and 20 ozs. of 
 silver per ton. The buddle concentrates are midway in richness 
 between the two slime-concentrates. About 80 per cent, of the 
 total values in the ore are saved on the tables and in the con- 
 centrates, and the latter can all be treated by the cyanide 
 process. 
 
 It was originally intended to dispose of the concentrates by 
 sale, but the prices realised, after deducting the expenses of 
 bagging, carting, shipping, insurance and treatment, were so 
 small as to render treatment on the spot an imperative necessity. 
 The system ultimately adopted was the MacArthur-Forrest 
 process, after exhaustive trials, and a plant was erected capable 
 of treating 20 tons per day. " It consists of three large agitators, 
 three vacuum filters, a grinding-pan, cyanide solution tank, 
 tanks for gold solution, vacuum and other pumps, and some 
 minor appliances. The whole plant is of local manufacture." 
 The filtration of the concentrates after agitation is difficult to 
 accomplish, and necessitated the introduction of a vacuum filter 
 patented by Dr. Scheidel. " The time of agitation, and the 
 strength of solution applied, vary in accordance with the 
 quality of the material. The quantity of cyanide used for 
 the highest grade of ore amounted to less than 1 per cent., 
 and for low-grade material to considerably under 0'5 per cent. 
 The time of agitation varied between five and twenty-four 
 hours." 
 
 The Sylvia Company acquired the right of using the 
 MacArthur-Forrest process on payment of a royalty of 7J per 
 cent, on the bullion extracted. The patentees did not interfere 
 with the construction of the plant, which was left in Dr. 
 Scheidel's hands. He states that " the results of extraction 
 have varied in accordance with the quality of the material, the 
 slimes generally giving better results than the other products, 
 and the (richer) first-class slimes returned a higher percentage of 
 gold and silver than the lower-grade materials." The extraction 
 was as follows : 
 
346 
 
 THE METALLURGY OF GOLD. 
 
 
 Gold. 
 
 Silver. 
 
 First-class slimes : (1) average extraction, 
 (2) highest 
 Second- ,, average ,, 
 Jigger concentrates : , , ,, 
 Buddie 
 
 86' 11 per cent. 
 96-45 
 85-34 
 80-32 
 77-9 
 
 67 per cent. 
 94-59 ,, 
 68-7 
 50 
 54-45 
 
 The average extraction on the whole was 82-67 per cent, of the 
 assay value ; the tailings are stacked for a possible future course 
 of re-treatment, as they contain large percentages of lead and 
 copper, besides precious metals to the value of about 5 per 
 ton. 
 
 The bullion produced by amalgamation and lixiviation com- 
 bined is thus seen to be about 66 per cent, of the total value of 
 gold and silver contained in the ore. It is obvious that, if 
 smelting works are established in the neighbourhood at any 
 future time, the concentrates will probably be sent to them, as 
 the tailings after lixiviation are still valuable. Nevertheless, it 
 is clear from the results obtained that in places far removed from 
 smelting works, certain complex ores, including those rich in 
 galena, may be treated by the cyanide process successfully. The 
 cost of treatment, consumption of cyanide, &c., are not given by 
 Dr. Scheidel. 
 
 New Primrose Cyanide Works.* At these works the 
 so-called " double treatment " is used, the tailings being treated 
 with cyanide in two vats in succession. The pulp from the 
 battery is passed into hydraulic classifiers (spitzluten), and 
 coarse sands and pyrites, amounting to about 10 per cent, of 
 the whole product of the stamps, are separated from the finer 
 material. The coarser materials, which contain 15 dwts. of 
 gold, are drained and treated for three or four days with a 
 solution containing O3 per cent, of cyanide. They are then 
 transferred to other vats and treated for a further period of 
 three days with weaker solutions, which are finally displaced 
 by water. The extraction amounts to 85 to 93 per cent., the 
 consumption of cyanide being about 1J Ibs. per ton. The fine 
 material, overflowing from the classifiers, is fed into a dis- 
 tributing tank and thence to other tanks in which the sand 
 settles, while the slimes overflow and are carried away. The 
 collected sands are then drained dry, the operation being 
 expedited by the use of vacuum pumps, a vacuum of 15 inches 
 of mercury being employed. A solution containing 0-04 per 
 
 * The account given here is mainly based on the notes supplied by 
 Mr. VV. Bettel, consulting chemist ot the New Primrose Mine, to Messrs. 
 Hatch & Chalmers' excellent work on Gold Mines of the Rand. 
 
THE CYANIDE PROCESS. 347 
 
 cent, of cyanide is then run on, and replaced by a stronger 
 solution, which is allowed to stand for twelve hours, and then, 
 drained off and the sand removed to the ordinary leaching 
 tanks. Here a solution containing 0'2 per cent, of cyanide is 
 added, and allowed to stand for twelve hours ; it is then re- 
 placed by weaker solutions, O08 and 0-04 per cent, being used 
 successively, and finally water, rendered alkaline by milk of 
 lime. An average of 78 per cent, of the gold is extracted, 
 while, before the "double treatment" was introduced, only 55 
 per cent, was obtained. The capacity of the plant is 17,000 
 tons per month, and the working expenses average 4s. Id. per 
 ton. The consumption of chemicals per ton of ore is as 
 follows : cyanide, 0'98 lb., zinc, 0'22 lb., and caustic soda, 
 0*22 lb. The bullion recovered is worth ,3, 5s. per oz. 
 
 Siemens-Halske Process. In this process the gold is de- 
 posited from solution by the passage through the liquid of a 
 current of electricity. Moreover, as the precipitation is equally 
 complete and as readily obtained in extremely dilute cyanide 
 solutions as in those containing O'l per cent, or over, the 
 strength of the solutions used in dissolving the gold from the 
 ores is made less when electrical precipitation is employed than 
 if zinc is the precipitant. Hence the whole method forms an 
 interesting variation of the ordinary Mac Arthur-Forrest process, 
 and bids fair to assume great importance in the future. Gold 
 can be extracted from its ores as completely by a solution 
 containing 0*03 per cent, of cyanide as by one containing 0'3 
 per cent., the only difference being that the time required is 
 considerably longer. On the other hand, the advantage in 
 using the more dilute solution is that the selective action .in 
 favour of the gold is increased, and the amount of cyanide 
 decomposed by " cyanicides " in the ore is diminished. In 
 addition to this, some cyanide solution is invariably left in the 
 ore, and if the "weak" solution used to finish the dissolution 
 of the gold contains only O'Ol per cent, of cyanide, instead 
 of 0*1 per cent., the amount of cyanide lost in this way by 
 mechanical means is also reduced. 
 
 The process was first adopted on a large scale at the Worcester 
 mill in the Transvaal. Here the vats are 20 feet in diameter 
 and have their sides formed of staves 10 feet long; the five vats 
 hold 135 tons each. The battery pulp, after passing over Frue 
 vanners, is classified into four products by hydraulic classifiers. 
 The first series, which consists of spitzluten, removes the coarse 
 sand and pyrites, amounting to 15 per cent, of the pulp and 
 containing 15 dwts. of gold. This product is treated for nine 
 days with solutions of 0-08 per cent, of cyanide, and, after being 
 washed with O'Ol per cent, solutions, gives residues assaying 
 1J to 2 dwts., so that from 87 to 90 per cent, of the gold i& 
 extracted. The second product yielded by the hydraulic 
 
348 THE METALLURGY OF GOLD. 
 
 classifiers comprises 50 per cent, of the pulp, contains 6 dwts. 
 of gold per ton, and, after five days' treatment with solutions 
 ranging from P 05 per cent, downwards, yields residues con- 
 taining from 1 to 1-25 dwts., showing an extraction of 80 to 84 
 per cent. The finest sand is separated from the slimes by 
 pointed boxes ; the slimes amount to 25 per cent, of the whole 
 pulp, and are not treated, being unleachable, though they assay 
 4J dwts. The tine sands, constituting 10 per cent, of the pulp, 
 contain 4J dwts. of gold, and after treatment yield residues 
 assaying 1 dwt. per ton. The consumption of cyanide averages 
 J Ib. per ton of the tailings, of which 3,000 tons are treated per 
 month. 
 
 The precipitation plant consists of four boxes, each 18 feet 
 long, 7 feet wide, and 4 feet deep. Copper wires are fixed along 
 the tops of the sides of the boxes and convey the electric current 
 from the dynamo to the electrodes. The anodes are made of 
 iron, and the cathodes, on which the gold is deposited, of lead. 
 These metals appear at present to be the most suitable for the 
 purpose, but it is possible that other substances may be found 
 equally good. It is necessary to have a very large surface on 
 which to deposit the gold, so that a bath of mercury is out of 
 the question, von Gernet calculating that it would require 80 
 tons of the metal to give the 10,000 square feet of surface 
 necessary to deal with 100 tons of solution in twenty-four hours. 
 Amalgamated copper plates were tried but abandoned, as the 
 mercury penetrated the copper under the influence of the 
 current, and a dry amalgam resulted which did not adhere to 
 the plate. Lead answers all the requirements of the cathode 
 laid down by von Gernet, which are : (1) that the precipitated 
 gold must adhere to it, (2) that it must be capable of being 
 rolled out into very thin sheets to avoid unnecessary expense, 
 (3) that it must be easy to recover the gold from it, and (4) that 
 it must not be electro-positive to the anode, in order to pre- 
 vent return currents being generated when the depositing 
 current is stopped. A fifth requirement might be added, that 
 the gold should be separable from the cathode without de- 
 stroying the latter. This requirement is not fulfilled by lead. 
 
 At the Worcester mill the anodes are iron plates, 7 feet long, 
 3 feet wide, and |- inch thick ; they are supported in a vertical 
 position by wooden strips nailed to the box, and are covered 
 with canvas to retain the small quantity of Prussian-blue pro- 
 duced. The sheet-lead cathodes are stretched on wires fixed 
 in light wooden frames which are suspended between the iron 
 plates. There are in all 3,000 square feet of cathode surface. 
 The solution is made to circulate between the plates passing 
 alternately over and under the edges of the anodes. The boxes 
 are kept locked except when cleaning up is necessary, when the 
 cathodes are removed and replaced with fresh sheets of lead. 
 
THE CYANIDE PROCESS. 349* 
 
 The lead, which contains from 2 to 12 per cent, of gold, is then 
 melted into bars and cupelled. The consumption of lead is- 
 750 Ibs. per month, its cost being equal to l^d. per ton of 
 tailings. The gold is comparatively free from base metals. 
 The total cost of treatment at the Worcester mill is under 3s. 
 per ton of tailings. The electrodes are placed 1J inches apart,, 
 and a current of 4 volts is employed, giving about 0'06 ampere 
 per square foot. In later practice, according to Butters, the 
 boxes are reduced in size, being now 29 feet long, 3^ feet 
 wide, and 2 feet 9 inches deep, the iron anodes are enclosed in 
 canvas sacking, and the space between them is filled loosely 
 with lead turnings, which have replaced the sheet lead formerly 
 used as cathodes. The liquid circulates through the turnings. 
 
 One of the main difficulties in electrical precipitation lies in 
 the great resistance offered to the passage of the current by the- 
 very dilute solutions employed. In order to reduce this as far 
 as possible, the electrodes must be of very large size, and the 
 distance between cathodes and anodes must be small. 
 
 The ions of aurocyanide of potassium are K and AuCy 2 . As- 
 the result of electrolysis, therefore, potassium is set free at the 
 lead cathode, where it attacks the water, forming potash and 
 hydrogen ; at the same time the gold in the double cyanide is 
 displaced and precipitated, both by the potassium and the nascent 
 hydrogen. Such hydrocyanic ncid as may be formed is neutralised 
 by the potash. Meanwhile the anion, AuCy 2 , is set free at the- 
 anode, but is at once split up into AuCy and Cy ; the latter 
 unites with the iron, forming cyanides, which become converted 
 into Prussian blue, and are also oxidised in part, forming ferric 
 oxide, and the cyanide of gold is partly precipitated in thi& 
 substance, although if enough free potassium cyanide is pre- 
 sent it may be kept in solution ; even if it is precipitated, 
 it can afterwards be recovered by treatment with potassium 
 cyanide. 
 
 It is, however, clear that the solutions must be circulated, and 
 also that the more free cyanide is present, the greater is the 
 percentage of gold precipitated on the cathode. It is not usually 
 profitable in the Transvaal to regenerate cyanide of potassium 
 from the Prussian blue. According to Christy, the precipitation 
 is more readily effected in acidified solutions.* As carried out 
 in South Africa, it is stated that the solutions after precipitation 
 still contain from J to 2 dwts. of gold per ton. 
 
 W. Bettel gives the results of work by this process in the 
 autumn of 1896 at two of the Transvaal mines, as follows :f 
 The "Bonanza" mill is designed to treat about 3,000 tons of 
 tailings and concentrates per month. The pulp is roughly 
 
 *See his paper in Proc. Am. Inst. Mng. Eng., 1896. 
 t Proc. Chem. and Met. Soc. of Johannesburg. Meeting held Nov. 21.. 
 1896. 
 
350 THE METALLURGY OF GOLD. 
 
 concentrated in spitzluten, about 20 per cent, of sand (containing 
 the pyrites, &c.) being separated and treated separately from the 
 rest. There are 12 tanks, each holding 100 tons, arranged in 
 two rows, one discharging into the other. Each charge of 
 tailings is treated for 2J days in an upper vat, and 5 days in a 
 lower one. A typical charge of tailings was treated as follows : 
 8 tons of alkaline wash (with 0-008 per cent. KCy) were added 
 at 9 a.m. on October 19, 1896. Between 12 noon and 2.30 p.m., 
 8 tons of weak solution (0-025 to 0-03 per cent. KCy) were run 
 on. From 19th to 21st October, 34 tons strong solution (average 
 0'07 per cent. KCy) were used, and the charge was then drained 
 and transferred to the lower vat, where between 21st and 24th 
 October, 72 tons of 0-07 per cent. KCy solution, 22 tons of 0-025 
 to 0-03 per cent., and 24 tons of 0-008 per cent. KCy ("alkaline 
 solution") were run through, giving a total of 168 tons of solu- 
 tion to 100 tons of tailings. 
 
 A charge of 100 tons of concentrates was treated in the upper 
 vat from 2nd. to 7th October (5 days) with 13 tons of alkaline 
 solution, 10 tons of weak solution, and 100 tons of strong solu- 
 tion. After being drained and transferred to the lower vat, 
 the charge was treated from 7th to 19th October (12 days) with 
 200 tons of strong solution, 15 tons of weak solution, and 30 
 tons of alkaline wash, or 368 tons of solution in all. The con- 
 centrates assay about 2 oz., and the residues 2J dwts., the 
 extraction being about 93 per cent. In 39 charges of tailings, 
 on the other hand, the average gold contents was 16^ dwts. per 
 ton before treatment, and over 3 dwts. after treatment, including 
 nearly J dwt. of dissolved gold, the extraction being 80-9 per 
 .cent. 
 
 The precipitation plant consists of four boxes, each 30 feet x 
 4J feet, and 3 feet deep, divided into two pairs in series con- 
 nected with a dynamo giving 200 amperes with an E.M.F. of 
 8 volts. The resistance of the weak solution is considerable. 
 Each box contains 23,960 sq. feet of cathode, and the rate of 
 flow is 3 tons of strong solution per hour per box, and 1 ton 
 of weak or alkaline solution. The liquid after precipitation 
 usually contains 1 or 2 dwts. of gold per ton, an average of 
 85 to 95 per cent, of the gold being precipitated. In cleaning 
 up, the lead is melted in a reverberatory furnace, and sold to 
 the Rand Central Ore Reduction Company. 
 
 In October, 1896, the total cost was 6s. 8'72d. per ton of ore, 
 but was soon after reduced by the introduction of continuous 
 haulage. The cost of cyanide was nearly Is. lOd. per ton, the 
 consumption being over H Ibs. per ton, the cost of lead was 
 l'2d. pence per ton, and the royalty was Is. 6d. per ton. 
 
 At the May Consolidated works, there are 17 vats of an 
 average capacity of 200 tons each, and the plant is capable of 
 treating 12,000 tons per month. The cost of the plant, which 
 
THE CYANIDE PROCESS. 351 
 
 was completed in April, 1896, was .27,500. The solutions used 
 range in strength from O'l per cent. KCy for strong solutions, 
 and 0-025 per cent. KCy for weak solutions, to O'Ol or 0'02 per 
 cent. KCy for solutions for preliminary treatment. The con- 
 centrates contain from 8 to 12 dwts. per ton, and are treated 
 for 10 days ; the tailings contain from 3| to 4^ dwts. before 
 treatment, and are treated for 4 days. The residues only con- 
 tain from 15 grains to 1 dwt. of gold per ton. The working 
 costs in the three months, August to October, 1896, were an 
 average of 2s. 7'7d. per ton in treating 9,000 tons per month, 
 exclusive of about 4d. per ton for interest and sinking fund for 
 the cost of the vats. 
 
 The Siemens-Halske process has made considerable progress 
 on the Witwatersrand goldfield, although there seems little 
 prospect at present that it will wholly displace the method of 
 precipitation by zinc. Its chief advantage over the latter lies 
 in the possibility of using more dilute solutions for dissolving 
 the gold, with a consequent reduction in the loss of cyanide, of 
 about J Ib. per ton of tailings (Bettel). Where from any cause, 
 solutions containing over O'l per cent. KCy are used, zinc 
 precipitation is preferred as being cheaper, but electrical pre- 
 cipitation is an absolute necessity in treating slimes. The 
 chief disadvantages of electrical precipitation are (1) The length 
 of time during which the gold is locked up, owing to the high 
 cost of the lead and the desirability of precipitating a large 
 amount of gold on it before it is melted down and treated. (2) 
 The greater cost of recovering the gold from the lead as compared 
 with the reduction of zinc slimes. (In the Transvaal, the base 
 lead bullion is sold at 83s. 6d. per oz. fine of gold, and 5 per 
 ton for lead, the silver varying somewhat.) (3) The large size 
 of the plant required as compared with that needed in the 
 MacArthur-Forrest process, owing to the excessive dilution of 
 the solutions used, which necessitate the time of treatment 
 being doubled. Moreover, the cost of the electrical plant is 
 additional, and a royalty is paid, whilst the Mac Arthur-Forrest 
 process is free. On the other hand, as noted above, the con- 
 sumption of cyanide is less under the Siemens-Halske system, 
 averaging about 0*3 Ib. per ton in the Transvaal, as against 0-6 
 to 0-7 Ib. per ton in the MacArthur-Forrest process. The 
 result of these drawbacks and advantages is that there is probably 
 little to choose between the two processes in South Africa, the 
 cost of treatment being about the same, viz., from 2s. to 2s. 6d. 
 per ton in the best modern mills, while the extraction is perhaps 
 higher with the stronger solutions. Most of the work con- 
 tinues to be done by zinc precipitation. 
 
 Results of the Cyanide Process. The success of the pro- 
 cess in South Africa, where a new industry has been created, 
 has been complete. The rapid progress which has been made 
 
 OF THB 
 
 UNIVERSITY 
 
352 
 
 THE METALLURGY OP GOLD. 
 
 there may be judged from the production of gold by cyanide in 
 each year since its introduction in 1890 : 
 
 Year. 
 
 Total Output at 
 Hand. 
 
 Production by 
 Cyanide. 
 
 Percentage of 
 Total Output. 
 
 
 Ozs. 
 
 Ozs. 
 
 Per Cent. 
 
 1890, 
 
 494,817 
 
 300 
 
 0-06 
 
 1891, 
 
 729,382 
 
 35,000 
 
 4-8 
 
 1892, 
 
 1,208,928 
 
 176,231 
 
 14-6 
 
 1893, 
 
 1,476,502 
 
 304,498 
 
 20-6 
 
 1894, 
 
 2,023,198 
 
 587,388 
 
 29-0 
 
 1895, 
 
 2,278,110 
 
 677,435 
 
 29-7 
 
 1896, 
 
 2,2^0,892 
 
 700,000 
 
 30-7 
 
 There were at first large accumulations of tailings to be dealt 
 with, but practically all of these have now been treated, and 
 since the year 1894 little but tailings or concentrates coming 
 direct from the stamps have contributed to swell the output. In 
 the future, cyanide may be expected to continue to yield about 
 30 per cent, of the total output on the Rand. Out of 677,435 ozs. 
 of gold produced in 1895, 38,708 ozs. were from concentrates, 
 and 638,727 ozs. from tailings. The average amount of gold 
 contained in the concentrates was about 2 ozs. per ton, but 
 one parcel of concentrates from the Village Main Reef yielded 
 19 '5 ozs. per ton. In the same period, concentrates, treated by 
 chlorination, yielded 65,833 ozs., the relative amounts of gold 
 from these materials produced by the two processes being about 
 the same as for the previous two or three years. It would 
 therefore appear that there is no striking advantage to be gained 
 by using either system instead of the other in treating concen- 
 trates on this field. 
 
 The cost of treating tailings by cyanide in the Transvaal varies 
 from little above 2s. per ton to 10s. per ton, the average being 
 probably about 4s. The cost of the cyanide consumed is from 
 6d. to 2s. per ton, and with this exception the biggest item is 
 that for handling the pulp, in charging and discharging the 
 vats, which amounts to Is. per ton and upwards. The royalty 
 formerly amounted to about Is. per ton of the tailings treated. 
 The amount of tailings now being treated is about 8,000 tons 
 per day, and the average yield is about 4-8 dwts., which is equal 
 to about 70 per cent, of the gold contained in the material 
 treated. It is evident from these results that the minimum 
 value of tailings capable of being treated at a profit in the 
 Transvaal ranges from about 1 J to 4 dwts. of gold per ton. The 
 cost of the treatment of concentrates is usually over 20s. per ton, 
 while the amount extracted is about 90 per cent. 
 
THE CYANIDE PROCESS. 353 
 
 The importance of the process to the mining industry is not 
 entirely represented by the output given above, as few of the 
 mines could work at a profit if they depended merely on the 
 gold extracted on the plates. The profit of say 15s. per ton 
 derived from the tailings converts the mine from a losing 
 into a paying concern in a large number of cases, and this profit 
 could not be earned without the cyanide process. 
 
 In other parts of the world, the gold-mining industry is less 
 dependent on the cyanide process than in South Africa, but 
 nevertheless it is gradually making its way on every quartz- 
 mining gold-field of any importance. In Australia many mines 
 have adopted it, and a large customs mill is at work in Queens- 
 land where over 14,000 ozs. of gold bullion were produced in 
 1894. In New Zealand the process is being worked at the 
 Waihi Mine, the Crown Mine, and a large number of others. 
 The progress in this colony may be judged from the fact that 
 in 1893, 17,271 ozs. of bullion of the value of 28,657 were 
 produced by cyanide, while in 1895 the amount had increased 
 to 133,162 ozs of gold. The process has only lately been intro- 
 duced into India, where in 1895, 3,614 ozs. of gold were pro- 
 duced, but several works have lately been erected, and the 
 output may be expected to be greatly increased in the near 
 future. In the United States the progress of the Mac Arthur- 
 Forrest process was for some years somewhat slow when com- 
 pared with that in South Africa and New Zealand. It was 
 adopted at the Mercur Mine, Utah, as early as the year 1891, 
 and was a great success from the start. It is now at work in 
 many of the States, the production in Colorado from its use 
 being considerable. In 1895 the production of gold in the 
 whole of the United States by cyanide was valued at over 
 300,000. The process will certainly continue to grow in im- 
 portance for some time to come in all parts of the world, and 
 in 1896 the total output of gold by its use was probably more 
 than 3,000,000. 
 
354 THE METALLURGY OP GOLD. 
 
 CHAPTER XYI. 
 CHEMISTRY OP THE CYANIDE PROCESS. 
 
 Action of Potassium Cyanide on Gold and other Metals. 
 Dr. Wright of Birmingham discovered in 1840 that metallic 
 gold is soluble in potassium cyanide, and Elkington's patent, 
 taken out in the same year, was partly based on this discovery. 
 Bagration, in 1843,* studied the action of cyanide on plates of 
 gold, and announced that they were slowly dissolved. Faraday, 
 in 1857,1 pointed out that gold-leaf is dissolved by a dilute 
 solution of the salt, and also showed that if the gold floats on the 
 surface of the liquid, so that one side of the leaf is in contact 
 with the air, while the other is bathed by the solvent, the action 
 is much more rapid than if the metal is completely submerged. 
 Eisner had previously proved J that the presence of oxygen is 
 required for the solution of the gold. On evaporating the solu- 
 tion, colourless octahedral crystals of auro-potassium cyanide, 
 KAuCy 2 , are formed, which may be viewed as being a double 
 cyanide, produced as follows : 
 
 4Au + 8KCy + O 2 + 2H 2 = 4KA*Cy 2 + 4KOH 
 This equation is exothermic. 
 
 The equation for the solution of silver is similar. 
 
 Theoretically, therefore, 130 parts by weight of KCy in the 
 presence of eight parts of oxygen suffice for the solution of 196-8 
 parts of gold. This has been recently proved by Mr. J. S. 
 Maclaurin to be the case in all carefully conducted experi- 
 ments. The amount of oxygen dissolved in liquids not specially 
 prepared, to say nothing of that contained in a porous mass of 
 pulverised ore, is consequently enough for the solution of great 
 quantities of gold. 
 
 According to G. Bodlander of Clausthal, the chemical action 
 in the dissolution of the gold is as follows : 
 
 2Au + 4KCy + 2H 2 O + O 2 = 2KAuCy 2 + 2KOH + H 2 2 
 and subsequently 
 
 4KCy + H 2 O, + 2Au = 2KAuCy 2 + 2KOH 
 Bettel suggests || the following in place of this : 
 
 (l)6KCy + 2Au + 2 + 2H 2 = KAuCy 2 + KAuCy 4 + 4KOH 
 (2) KAuCy 4 + 2Au + 2KCy = 3KAuCy 2 
 
 * Bull, de VAcad. des Sciences de St. Petersbcurg (1843), vol. ii., p. 136. 
 
 ^Roy. Inst. Proc., vol. ii., p. 308. 
 
 %Erdm. Journ. Prak. Chtm., vol. xxxvii. (1846), pp. 441-446. 
 
 %Journ. Chem. Soc. (1893), vol. Ixiii., p. 724. 
 
 || South African Mining Journal, May 8, 1897. 
 

 CHEMISTRY OP THE CYANIDE PROCESS. 355 
 
 Evidence has been adduced that a substance is formed re- 
 acting like H 2 O 2 , but it does not follow that the actions repre- 
 sented by the equations given above are not limited to an 
 insignificant part of the whole mass. MacArthur has cited 
 experiments* to show that gold in ores can be dissolved by 
 potassium cyanide in the absence of oxygen, and Bettel t found 
 this to be the case if the crushed ore contains basic ferric sul- 
 phate, by which potassium ferric cyanide is formed, the reactions 
 being expressed thus 
 
 (l)Fe 2 (S0 4 ) 3 + 12KCy = 3K 2 S0 4 + K 6 Fe 2 Cy 12 
 
 O) 6 + 
 
 (2)Fe 2 (S0 4 ) 3 + 6KCy = Fe 2 (HO) 6 + 2K 2 SO 4 + 6HCy 
 Bettel and Marais also found in 1894 that gold leaf would 
 not dissolve in a solution of potassium cyanide from which the 
 air had been expelled by the passage of a current of hydrogen, 
 and that the addition, under these circumstances, of either 
 potassium dichromate, chromate, chlorate, perchlorate, nitrate, 
 or nitrite or of ferric hydrate or bleaching powder, did not 
 enable the gold to dissolve. The addition of pyrolusite gave a 
 doubtful result, and lead dioxide caused very slow dissolution 
 of the gold. On the other hand, gold dissolved slowly if 
 ohlorine, iodine dissolved in potassium iodide, or ferric chloride, 
 were added ; rapidly, if bromine were added ; and decidedly, if 
 potassium ferricyanide or permanganate, sodium dioxide, hydro- 
 gen peroxide or barium dioxide, were added. It is clear, there- 
 fore, that certain oxidisers are ineffective, and in practice, when 
 air is not used, potassium permanganate, sodium dioxide and 
 bromine (Sulman-Teed process), have been mainly employed to 
 assist in the dissolution of the gold. In general, artificial 
 oxidation is not resorted to unless there is some reducing agent 
 present in the ore or the water which absorbs the oxygen. 
 It was formerly supposed that any oxygen coming in contact 
 with the solution would be at once taken up by the cyanide, 
 forming cyanates, but it is now well known that oxygen and 
 cyanide remain side by side without rapid union between them 
 taking place. 
 
 It has long been remarked that a small percentage of a 
 soluble sulphide present in the cyanide solution greatly delays 
 the dissolution of gold. Doubtless this is partly owing to the 
 abstraction of oxygen from the solution by the sulphide, for 
 gold sulphide is freely soluble in KOy, so that the surface of the 
 metal is kept free from sulphide. Bettel, however, points out { 
 that silver sulphide is far less soluble than gold sulphide, and 
 that if native gold is alloyed with 20 per cent, of silver, a film 
 almost insoluble in [dilute] cyanide solutions may be formed. 
 It is certain that some specimens of gold leaf dissolve with 
 great difficulty if they have been previously dipped in sulphide 
 
 * Jour Soc. Chem. Ind., 1890, p. 7. 
 
 t S. African Mug. Jour., Loc. cit. J Loc. cit. 
 
356 
 
 THE METALLURGY OF GOLD. 
 
 solutions, or if traces of soluble sulphides or sulpho-cyanides are 
 present in the solution. The difficulty disappears if the sulphides 
 are removed, either by being precipitated with lead salts, or by 
 the action of certain oxidisers. 
 
 The voltaic order of the metals in different solutions ol cyanide 
 of potassium is given in the following table : 
 
 1 part KCy in 8 parts 
 water.* 
 
 1 part KCy in 160 parts 
 water, t 
 
 1 part KCy in 3 parts 
 water, t 
 
 At 50 F. 
 
 At 100 F. 
 
 At 50 F. 
 
 At 100 F. 
 
 + Zn-amalgam. 
 Zn. 
 
 + Al. 
 
 Mg. 
 
 + Mg. 
 Zn. 
 
 + Mg. 
 Zn. 
 
 + Mg. 
 Al. 
 
 Cu. 
 
 Zn. 
 
 Cd. 
 
 Cu. 
 
 Zn. 
 
 Cd. 
 
 Cu. 
 
 Al. 
 
 Al. 
 
 Cu. 
 
 Sn. 
 
 Cd. 
 
 Co. 
 
 Cd. 
 
 Cd. 
 
 Ag. 
 
 Sn. 
 
 Cu. 
 
 Au. 
 
 Sn. 
 
 Ni. 
 
 Co. 
 
 Ni. 
 
 Ag. 
 
 Au. 
 
 Sb. 
 
 Ni. 
 
 Sn. 
 
 Ni. 
 
 Ag. 
 
 Pb. 
 
 Ag. 
 
 Au. 
 
 Sn. 
 
 Ni. 
 
 Hg. 
 
 Au. 
 
 Ag- 
 
 Hg. 
 
 Hg. 
 
 Pd. 
 
 Hg. 
 
 Pb. 
 
 Pb. 
 
 Pb. 
 
 Bi. 
 
 Pb. 
 
 Hg. 
 
 Co. 
 
 Co. 
 
 Fe. 
 
 Fe. 
 
 Sb. 
 
 Sb. 
 
 Sb. 
 
 
 
 
 Te. 
 
 Te. 
 
 Pt. 
 
 Pt. 
 
 Bi. 
 
 Bi. 
 
 Bi. 
 
 Cast iron. 
 
 Sb. 
 
 Fe. 
 
 Fe. 
 
 Fe. 
 
 
 
 Te. 
 
 
 
 - Coke. 
 
 BL 
 
 Pt. 
 
 Pt. 
 
 Pt. 
 
 
 Te. 
 
 
 
 
 ... 
 
 - C. 
 
 - a 
 
 - 6V 
 
 - C. 
 
 The position of the metals in the tables denotes also their 
 relative initial tendency to dissolve under the given conditions, 
 one of which is, of course, that they form part of a complete 
 electric circuit. It may be deduced that mercury is inapplicable 
 as a precipitant for gold from solution, and that aluminium 
 would perhaps act similarly to zinc in cold^ dilute solutions. 
 Moldennauer proposed to apply this in practice in presence of 
 free alkali, when the cyanide is regenerated and alumina formed, 
 thus 
 
 6AuKCy 2 + 6KOH + 2A1 = 6Au + 12KCy + 3H 2 + A1 2 3 
 It is claimed that there is comparatively little waste of alu- 
 minium. 
 
 The following table | shows the actual amounts of metals 
 (stated in grains per square inch of surface per hour (dissolved 
 by a solution containing 1'86 per cent, of KCy : 
 
 * Poggendorf , vide Gore's Electrolytic Sep. of Metals, p. 57. 
 tGore in Proc. Roy. Soc,, vol. xxx., p. 38. 
 JDr. Gore, Proc. Roy. Soc., vol. xxxvii., p. 283. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 
 
 357 
 
 
 At 60 
 
 J, 
 
 At 16( 
 
 ) F. 
 
 
 
 Amounts dis- 
 
 
 Amount dis- 
 
 
 Wt. dissolved. 
 
 solved, Gold 
 being taken 
 
 Wt. dissolved. 
 
 solved, Gold 
 being taken 
 
 
 
 asl. 
 
 
 asl. 
 
 Al, 
 
 0095 
 
 10'5 
 
 4940 
 
 17-5 
 
 Zn, 
 
 0054 
 
 6 
 
 0495 
 
 1-75 
 
 Cu, 
 
 0053 
 
 5-9 
 
 1258 
 
 4-5 
 
 Ag, 
 
 0016 
 
 1-8 
 
 0070 
 
 0-25 
 
 Cd, 
 
 0014 
 
 1-6 
 
 0046 
 
 0-16 
 
 Au, 
 
 0009 
 
 1 
 
 0262 
 
 1-0 
 
 Sn, 
 
 0006 
 
 66 
 
 0042 
 
 0-15 
 
 Ni, 
 
 0005 
 
 55 
 
 0027 
 
 o-io 
 
 Fa, 
 
 Trace. 
 
 ... 
 
 0010 
 
 0-035 
 
 Pt, 
 
 None. 
 
 
 None. 
 
 ... 
 
 Pb, 
 
 
 ... 
 
 0031 
 
 0-11 
 
 It will be observed that the rate of dissolution of gold is increased 
 thirty-one times, and that the "selective action" in favour ot it 
 is also increased (except as regards aluminium) by the increase of 
 temperature. It must be noted, however, that Dr. Gore's table 
 refers to metals in the form of sheets : it must not be assumed 
 that the relative rates of dissolution given above will hold good 
 when ores are being treated. (See also p. 365). The position of 
 .zinc on the table is instructive. Dr. Gore has also shown that 
 the rate of solution of gold is increased by 50 per cent, if it is 
 placed in contact with iron, and that it is then five times more 
 soluble than the iron. 
 
 The heat disengaged in the formation of metallic and other 
 cyanides is shown in the following table, the authorities for which 
 are Berthelot and Thomsen : 
 
 
 
 Heat Disengaged. 
 
 Components. 
 
 Compound Formed. 
 
 
 Compound 
 
 
 
 Compound 
 in Solid State. 
 
 Dissolved 
 in Solution 
 
 
 
 
 of KCy. 
 
 H +Cy 
 
 HCy 
 
 
 + 11-4 
 
 K +Cy 
 
 Na + Cy 
 
 KCy 
 
 NaCy 
 
 + 68-3 
 + 60'4 
 
 + 65-4 
 + 59-9 
 
 *Ca + Cy 
 4Zn + Cy 
 
 4(CaCy 2 ) 
 4(ZnCy 2 ) 
 
 + 26-7 
 
 + 57-7 
 + 31-1 
 
 4Cd + Cy 
 
 4(CdCy 2 ) 
 
 + 17-7 
 
 ... 
 
 4Hg + Cy 
 Ag + Cy 
 
 4(HgCy a ) 
 AgCy 
 
 + 9-5 
 + 4-3 
 
 + 15-3 
 + 10-8 
 
358 
 
 THE METALLURGY OF GOLD. 
 
 The heat disengaged in the union of equivalent quantities of 
 the metals in the solid state (except Hg) with cyanogen is given. 
 
 Messrs. MacArthur and Forrest have stated that no increase 
 in the rate of solubility of gold and silver takes place on 
 increasing the concentration of the cyanide solution. As excep- 
 tion has been taken to this statement by other authorities, the 
 author conducted the following experiments in the laboratory 
 at the Royal Mint, in order to increase the available data on 
 the subject as far as metallic gold is concerned. 
 
 Gold cornets, which had all been subjected to the same 
 treatment, being cupelled, boiled, and annealed together, were 
 submerged in 30 c.c. of KCy solution. Cornets offer a large 
 surface for action, as they are spongy in texture. After some 
 time the cornets were taken out of the solution, well washed,, 
 dried, ignited, and weighed. 
 
 The results obtained were as follows : 
 
 
 
 
 Duration 
 
 GOLD DISSOLVED. 
 
 Weight of Gold taken. 
 
 Solution. 
 
 ture. 
 
 Action 
 of Solvent. 
 
 Weight in 
 Grammes. 
 
 Per 
 
 Cent. 
 
 4997 gramme, 
 
 1 p.c. 
 
 50C. 
 
 11 hours. 
 
 0023 
 
 0-46 
 
 4990 
 
 1 ,, 
 
 15' 
 
 l| 
 
 00095 
 
 0'19 
 
 4991 
 
 5 , 
 
 M 
 
 l| 
 
 0009 
 
 0-18 
 
 4988 
 
 25 , 
 
 
 ll 
 
 0010 
 
 0-20 
 
 5001 
 
 1 , 
 
 
 
 5 
 
 00305 
 
 0-61 
 
 5001 
 
 1 , 
 
 100 
 
 5 
 
 0065 
 
 1-30 
 
 5000 
 
 1 , 
 
 60 
 
 Ik 
 
 00285 
 
 0-57 
 
 These results tend to show that the rate of solution of nearly 
 pure gold is a function of time and temperature, but the degree 
 of concentration of the solution is less important. At 100 the 
 cornet was blackened, and a black precipitate was formed and 
 remained floating in the liquid. Eor the relative rate of action 
 of potassium cyanide, chlorine and bromine, compare the results 
 given above with those on p. 267. 
 
 Mr. J. S. MacArthur remarks on the subject of temperature : 
 " Bagration* states that the solution of metallic gold in cyanide 
 of potassium is facilitated by a rising temperature. We have not 
 found this so in the case of dissolving gold from its ores. Up to- 
 a temperature of about 100 to 110F. (the highest temperature 
 one is likely to encounter in the Tropics) the solution from ores 
 is unaffected, but beyond that the consumption of cyanide is 
 seriously increased and the selective solvent action retarded till,, 
 at the boiling point, it ceases, and probably is to a certain extent 
 reversed, as we have found that gold, which had been dissolved 
 
 * Journal fur Praktische Chemie, Leipzic, 1844, vol. xlvii., p. 367. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 359 
 
 from an ore by a cold solution of cyanide, was re-precipitated on 
 the ore by heating towards the boiling point." 
 
 Mr. J. S. Maclaurin of Auckland University, New Zealand, 
 has recently demonstrated * that the rate of dissolution of pure 
 gold, in the form of plates, in potassium cyanide solutions passes 
 through a maximum when proceeding from dilute to concentrated 
 solutions. The maximum is reached when the solution contains 
 0-25 per cent, of KCy. The solubility of gold is very slight in 
 solutions containing less than 0*005 per cent., but increases 
 rapidly as the strength rises to 0*01 per cent., when the rate 
 of dissolution is ten times as great as in the 0*005 per cent, 
 solution, and about half as great as that in the 0*25 per cent. 
 The rate increases slowly as the strength rises to 0-25 per cent., 
 and thereafter decreases much more slowly, until in 15 per cent, 
 solutions the rate of dissolution is about equal to that in 0-01 per 
 cent, solutions. Higher strengths show a gradual diminution in 
 the rate of dissolution up to saturation point. Silver is also dis- 
 solved at a maximum rate in solutions containing 0*25 per cent, 
 of cyanide, and the changes in the rate are similar to those 
 noted above in the case of gold, the rates for silver being always 
 about two-thirds of the corresponding rate for gold, or, roughly, 
 in the same ratio as the atomic weights of the two metals. In 
 both cases there is hardly any change in the rate of solubility as 
 the strength rises from O'l per cent, to 0-25 per cent. From 
 these results, it is seen that the most active solutions are now 
 used in the ordinary practice of the MacArthur-Forrest process 
 in South Africa, and that the solutions used in the Siemens- 
 Halske practice (0*01 per cent, to 0*1 per cent.) are little inferior 
 to them in activity. It is remarkable that the solubility of 
 oxygen in cyanide solutions undergoes similar changes as the 
 concentration increases, and it is to this fact that Maclaurin is 
 disposed to attribute the variations in the rate of dissolution of 
 the gold. The oxygen which unites with the cyanide, converting 
 it into cyanate, is thereby made inactive, as the presence of 
 cyanate of potassium has no effect on the rate of dissolution 
 of the gold. 
 
 The table on p. 360 gives the results of some experiments 
 made by Mr. Louis Janin, jun.,f on the solubility of metallic 
 silver in potassium cyanide. In this case it appears that a 
 maximum rate of dissolution is reached when the strength of 
 the solution is only about 1 per cent. ; the rate then diminishes 
 gradually as the strength increases from 1 to 3 per cent., and 
 continues to diminish, although much more slowly, until the 
 strength reaches 15 per cent., after which an increase in con- 
 centration has no effect on the rate of dissolution. The influence 
 of the volume of solution seems to be small, and that of time not 
 very great, after the first few hours. 
 
 *Journ. Chem. Soc., vol. Ixiii. (1893), p. 731 ; and vol. Ixvii. (1895), p. 199. 
 . and Mng. Journ., Dec. 29, 1888. 
 
360 
 
 THE METALLURGY OP GOLD. 
 
 Weight of 
 Cement 
 Silver. 
 Mgs. 
 
 Strength of 
 Solution 
 Per Cent. 
 
 Mgs. of 
 Silver 
 Dissolved. 
 
 Percentage 
 Dissolved. 
 
 Volume 
 of Solution 
 in c.c. 
 
 Time of 
 Standing 
 in Hours. 
 
 500 
 
 0-50 
 
 123-8 
 
 24-7 
 
 500 
 
 12 
 
 if 
 
 050 
 
 1265 
 
 25-2 
 
 1000 
 
 12 
 
 9 
 
 1-00 
 
 138-2 
 
 27-6 
 
 500 
 
 12 
 
 9 9 
 
 1-00 
 
 109-8 
 
 21-9 
 
 500 
 
 6 
 
 |f 
 
 1-00 
 
 1795 
 
 35-9 
 
 1000 
 
 96 
 
 | 
 
 2-00 
 
 1124 
 
 22-4 
 
 500 
 
 12 
 
 9 
 
 2-00 
 
 170-0 
 
 340 
 
 1000 
 
 96 
 
 9 
 
 3-00 
 
 100-0 
 
 20-0 
 
 500 
 
 12 
 
 9 
 
 3-00 
 
 145-5 
 
 29-0 
 
 1000 
 
 96 
 
 | 
 
 4-00 
 
 100-0 
 
 20-0 
 
 500 
 
 12 
 
 9 
 
 5-00 
 
 100-0 
 
 20-0 
 
 500 
 
 12 
 
 
 6'00 
 
 921 
 
 18-4 
 
 500 
 
 12 
 
 | 
 
 1000 
 
 76-0 
 
 15-2 
 
 500 
 
 12 
 
 j 
 
 15-00 
 
 88-0 
 
 17-6 
 
 500 
 
 12 
 
 9 
 
 20-00 
 
 88-0 
 
 17-6 
 
 500 
 
 12 
 
 9 
 
 25-00 
 
 88-0 
 
 17-6 
 
 500 
 
 12 
 
 Decomposition of Potassium Cyanide. Hydrocyanic acid 
 is one of the weakest acids known, and is expelled from its salts 
 by all mineral acids and many organic acids. Carbonic acid 
 decomposes potassium cyanide in presence of water thus : 
 
 2KCy + C0 2 + H 2 = 2HCy + K 2 C0 3 
 (Exothermic: + 10'5 + 82-3 + lO'l - 64*7 - 34-5 = + 3'7) 
 
 The srnell of hydrocyanic acid, noticeable whenever KCy or its 
 solutions are exposed to the air, is accounted for by this reaction. 
 In the presence of air, potassium cyanide takes up oxygen, and 
 is converted first to cyanate and then to carbonate 
 
 KCy + O = KCyO( + 69'7cal.) 
 
 2KCyO + 3H 2 = K 2 C0 3 + C0 2 + 2NHs( + 3'7). 
 
 These reactions are much more rapid if heat is applied. Strong 
 solutions turn brown in the air. In dilute solutions, potassium. 
 cyanide appears to be partly changed into HCy and KOH, since 
 the passage of a stream of any neutral gas such as nitrogen 
 through the solution causes an evolution of hydrocyanic acid, 
 while the solution becomes alkaline. The equation is 
 
 KCy + H 2 = HCy + KOH 
 (Endothermic: + 10'5 + 82'3 - 64'7 - 34'5 = - 6'4) 
 
 If the solution is boiled with acids or alkalies, hydrolysis of the 
 cyanide occurs rapidly, ammonia and formic acid being formed, 
 thus 
 
 KCy + 2H 2 O = NH 3 + HC0 2 
 
 This equation does not represent the whole effect, as acetates 
 and other organic substances are also formed. The reactions 
 
CHEMISTRY OF THE CYANIDE PROCESS. 361 
 
 proceed slowly even in the cold, especially if much free alkali is 
 present. It is obvious, from the observed facts of the decom- 
 position of cyanide solutions mentioned above, that (1) the 
 solution must be kept from contact with the air as far as 
 possible, by having closely fitting lids to the storage and leaching 
 vats, the zinc boxes, &c., and by transferring the solution from 
 one to the other by means of iron pipes (not open launders) ; and 
 (2) the solution must be kept free from acids, and this can be 
 effected by the addition of a little alkali. 
 
 Reactions in the Zinc Boxes. It was generally believed 
 at one time that the precipitation of gold and silver by zinc was 
 effected by simple displacement, according to the equation 
 
 2KAuCy 2 + Zn = K 2 ZnCy 4 + 2 Au( + circ. 40 calories). 
 
 Some considerations have been adduced, however, which throw 
 doubt on this view. It is well known that a considerable excess 
 of free cyanide must be present to enable the precipitation to 
 take place quickly and efficiently,* and if the solution con- 
 tains less than about O'Oo per cent. KCy the solution must be 
 in contact with the zinc for an hour. Pure sheet zinc in par- 
 ticular is quite unable to precipitate gold in the absence of 
 free potassium cyanide, although the filiform zinc slowly extracts 
 the gold from the same solution, perhaps owing to the pre- 
 sence of lead in the zinc.t Christy suggests that the cause of 
 this is that hydrogen is set free by the first action of the 
 aurocyanide of potassium on the zinc, thus 
 
 2K AuCy 2 + 2Zn + 2H 2 = 2KOH + H 2 + 2ZnCy 2 + 2Au + (circ. 45 calories). 
 
 This equation practically demands the replacement of the K 
 ions in the aurocyanide by zinc thus 
 
 2(K + AuCy a ) + Zn = (Zn + 2AuCy 2 ) + 2K 
 followed by a re- arrangement of the molecule of zinc-gold 
 cyanide and the replacement of the gold by another atom of 
 zinc. This mechanism of change seems the more probable 
 when it is remembered that the K ions in KCy itself are 
 displaced by zinc in the presence of water, and that the whole 
 reaction is more strongly exothermic than that of simple dis- 
 placement of gold by zinc. 
 
 Assuming, then, that hydrogen is set free as above, it would 
 form an imperceptible layer of hydrogen on the surface of the 
 sheet zinc, which would thereby become "polarised" and pro- 
 tected from further action, so that the reaction would soon 
 cease. Hydrogen would also be set free on the surface of the 
 thread zinc, but in this case the ragged edges of the shavings 
 
 * W. R. Feldtmann in Eng. and Mng. Jour., August 11, 1894. 
 
 t"0n the Solution and Precipitation of Gold," by Prof. S. B. Christy. 
 Am. Inst. Mng. Eng., 1896. "The precipitation of Gold byline-Thread 
 from Dilute and Foul Cyanide-Solutions," by A. James. Ibid., Feb. 1897. 
 
362 THE METALLURGY OF GOLD. 
 
 would assist the gas to form into bubbles and become detached^ 
 so that less " polarisation " would occur. Hydrogen is actually 
 given off after some time has elapsed, but there are other theories 
 to account for its formation, as noted below. It does not 
 appear, moreover, how the presence of free KCy would prevent 
 polarisation by hydrogen, and the prevention of the action may 
 be due to the formation of an invisible protective layer of 
 insoluble ZnCy 2 on the surface of the zinc, which would be at 
 once dissolved if potassium cyanide were present. In the absence 
 of KCy, the reaction proceeds 
 
 4ZnCy 2 + 4KOH = 2ZnCy 2 + ZnCy 2 . 2KCy + Zn(KO)a + 2H 3 O* 
 
 This reaction is, of course, incomplete, but unless the amount 
 of potash in the solution were large, it is clear that some insol- 
 uble zinc cyanide would remain. 
 
 The reaction is completed by the solution of ZnCy 2 in free 
 potassium cyanide. 
 
 ZnCy a + 2KCy = K 2 ZnCy 4 
 
 According to the reactions on p. 361, one part of zinc should 
 precipitate about three parts of gold, but in practice one part 
 of zinc precipitates only from O06 to 0'2 part of gold, the zinc 
 being wasted by direct dissolution in potassium cyanide and 
 by the reactions described below. 
 
 Pure zinc has a slower action on solutions of potassium 
 cyanide than ordinary commercial zinc containing about one per 
 cent, of lead, and the action of the gold-zinc couple, formed by 
 the black deposit of gold (which may be really a compound of 
 gold and zinc), and the unaltered zinc, is still more vigorous. 
 This gold-zinc couple probably develops enough electromotive 
 force to decompose water thus 
 
 Zn + 2H 2 Zn(OH) 2 + H 2 (41'8-34'5 
 
 The positive element, zinc, is thus oxidised, and subsequently 
 dissolved by the cyanide solution. As a matter of fact, Messrs. 
 Butters & Clennel f observed a vigorous evolution of small 
 bubbles, which proved to be mainly hydrogen, in the zinc boxes 
 at the Robinson Mine. The hydrogen thus evolved doubtless 
 carries off mechanically some hydrocyanic acid, and so the odour 
 of the latter, noticeable above the zinc boxes, may be accounted 
 for. The nascent hydrogen also unites directly with hydro- 
 cyanic acid, forming methylamine, thus 
 
 HCN + 2H 2 = CH 2 .NH 2 
 
 The presence of the methylamine may in part account for the 
 ammoniacal odour sometimes occurring above the zinc boxes, 
 
 J. S. C. Wells in Eng. and Mng. Jour., December 21, 1895, p. 585. 
 . and Mng. Jour., 1892, p. 416. 
 
CHEMISTEY OP THE CYANIDE PROCESS. 36$ 
 
 which may, however, be also caused by the hydrolytic decomposi- 
 tion. The hydrate of zinc, formed as shown above, dissolves at 
 once in excess of the cyanide 
 
 Zn(OH) 2 + 4KCy = K 2 ZnCy 4 + 2KOH, 
 
 and the increase in alkalinity of the solution is thus explained. 
 
 Caustic potash, however, acts directly on the zinc, to some 
 extent forming potassium zinc-ate, thus 
 
 Zn + 2KOH = Zn(OK) 2 + H 2 
 
 but the zincate is, according to N. Anderson,* at once dissolved 
 by potassium cyanide, as follows : 
 
 Zn(OK 2 ) + 4KCy + 2H 2 O = ZnCy 2 . 2KCy + 4KOH 
 
 In this way the solution continually tends to become more 
 alkaline, but is neutralised by carbonic acid absorbed from the 
 atmosphere. 
 
 By the reactions in the zinc boxes, not only is the potassium 
 cyanide solution weakened by decomposition, as above, but a. 
 large amount of zinc is dissolved. Goyder has shown f that 
 the presence of iron in contact with the zinc is injurious, check- 
 ing the rate of precipitation of the gold, and increasing the 
 waste of zinc and the tendency for the formation of ZnOy 2 on the 
 zinc. Iron screws should therefore not be used in the precipita- 
 tion boxes. The double cyanide of zinc and potassium is not 
 available for the solution of the precious metals, and as long as- 
 an excess of zinc is present, no gold will be dissolved by a solu- 
 tion of potassium cyanide flowing through the boxes. The 
 presence of large quantities of the double cyanide of zinc and 
 potassium in the solutions is not prejudicial to the solvent 
 action of the simple cyanides. At the Mercur Mine, the stock 
 solution was apparently as efficacious after nine months' use as 
 at the start, although it must have contained large quantities of 
 zinc cyanide. Feldtmann showed that gold in ores can be 
 dissolved by zinc-potassium cyanide, but J. S. C. Wells points 
 out that the double cyanide remains undecomposed by gold so 
 long as any simple cyanide is present. An accumulation of base 
 heavy metals in the solution can be got rid of by the addition of 
 soluble sulphides. 
 
 Ehrmann has shown J by experiments on solutions of various 
 strengths, that about as much gold is precipitated by zinc in 
 24 hours at 20C, and in two hours at 80C. In the latter 
 case, however, there is, according to MacArthur, " an enormous 
 
 * Jour. Colo. Sci. Soc., April 1895. Quoted by Ingalls. Min. Industry 
 for 1895, p. 330. 
 
 t Chem. News., vol. 73 (1896), p. 272. 
 
 Chem. and Met. Soc. S. Africa, April 17, 1897. 
 
364 THE METALLURGY OF GOLD. 
 
 waste of cyanide by the formation of urea, which manifested 
 itself by its strong unpleasant odour.' J 
 
 Action of Potassium Cyanide on Metallic Salts and 
 Minerals Occurring in Ores. The ordinary gangue of most 
 ores (viz., silica and silicates of the alkalies and alkaline earths) 
 -exercises no direct influence on the cyanide solution. The 
 carbonates of the alkaline earths are also probably without 
 influence. The decomposing effects of sulphides of the heavy 
 metals vary with the physical state of the sulphides. 
 
 In 1892, Kedzie* made a large number of experiments, using 
 different samples of pyritic ores and concentrates, and solutions 
 containing from 0-5 to 1-5 per cent, of pure potassium cyanide, 
 and found that the consumption (decomposition) of the cyanide 
 varied from 3 to 50 Ibs. per ton of ore. He found that the 
 consumption increased with the time, the volume of the solution, 
 and the degree of concentration, the effect of the latter being 
 especially marked. An increase of concentration from O5 per 
 cent, to 0-9 per cent, caused the consumption to be doubled, 
 and a iurther increase to 1*5 per cent., doubled the consumption 
 again. Sulphide of iron, he found, acts very little on potassium 
 oyanide, but sulphides of copper and zinc are rapidly dissolved. 
 
 Messrs. MacArthur and Forrest find that dilute cyanide solu- 
 tions exercise a " selective action " in dissolving gold and silver 
 in whatever form they may be present, in preference to sul- 
 phides or other salts of the base metals. There are exceptions 
 to this rule, some of which are noted in the sequel. "For 
 instance, cyanide of potassium solution has a strong tendency 
 to dissolve precipitated sulphide of zinc, but its action on 
 the natural sulphide of zinc, blende, is almost nil. The same 
 holds for compounds of iron, and thus we prove selective action 
 by the average result of a series of experiments on ores. Let us 
 suppose a pyritic ore containing about 7 per cent, of iron and 
 S per cent, of sulphur with about 1 oz. of gold to the ton. After 
 grinding, this ore is treated with a solution containing about 
 1 per cent, of cyanide of potassium. The most of the gold will 
 be dissolved and the rest of the ore left practically untouched. 
 It is obvious that the amount of cyanogen contained in the 
 solution is insufficient to combine with the iron present in the 
 ore, yet, notwithstanding the much greater mass of iron sulphide 
 present and open to attack, it is the gold that is selected for 
 action by the cyanide solution. Taking the average result of 
 our work we find that a higher percentage of gold than of silver 
 is extracted, which justifies us in concluding that the selective 
 action is greater on the former than on the latter. One of the 
 ores on which our early investigations were done was composed 
 as under : 
 
 * Eng. and Mng. Journ., p. 606, 1893. 
 
CHEMISTRY OF THE CYANIDE PROCESS, 
 
 365 
 
 Copper, 
 Arsenic, 
 Antimony, 
 Sulphur, 
 Iron, . 
 Silica, 
 Lead, . 
 Zinc, . 
 Alumina, 
 Gold, per ton 
 Silver, ,, 
 
 0'15 per cent. 
 15-09 
 Traces. 
 
 14-65 per cent. 
 18-77 
 36-20 
 
 2-66 
 
 4-00 
 
 4-20 
 
 2 ozs. 2 dwts. 16 grs. 
 
 2 ozs. 13 dwts. 8 grs. 
 
 In this ore we had an extraction of gold 85 per cent., silver 
 50 per cent., for a consumption of cyanide of about 0-45 percent., 
 and investigations showed that the action was directed in the 
 order, gold, silver, iron, zinc, copper. From the amount of cyanide, 
 consumed it is obvious that the amount of base metals dissolved 
 must, have been very slight. 
 
 " The consumption of cyanide on fresh concentrates varies 
 naturally with the composition of the concentrates. In many 
 cases it is less than 0-2 per cent, of their weight. When the 
 concentrates contain marcasite there is a greater consumption ol 
 cyanide than when the pyrites is entirely of the ordinary yellow 
 cubical description. The presence of compounds of copper, 
 physically soft, also tends to increase the consumption." 
 
 It has been laid down as a general rule that oxides, hydrates, 
 carbonates, sulphates and sulphides of those metals which are 
 electro-positive to gold in cyanide solutions are dissolved more 
 rapidly than the last named metal, whether it is present in the 
 metallic form or contained in its commonly occurring salts. 
 
 This rule certainly applies to the precipitated salts commonly 
 occurring in the laboratory, but J. S. MacArthur has shown that 
 the case may be quite different when the naturally occurring 
 minerals are concerned. Thus, not only is precipitated sulphide 
 of copper rapidly dissolved, but also a sooty form of the same 
 substance occasionally met with as a mineral occurring in ores. 
 On the other hand, fused copper matte is scarcely acted on at all, 
 and in the majority of cases the same may be said of the hard 
 dense sulphides of copper usually found in nature. Sulphide of 
 zinc exhibits the same differences of behaviour : the "black-jack " 
 concentrates of the Ravenswood Mine, Queensland, can be treated 
 with good results, little zinc being dissolved. Again, oxide of 
 copper, if freshly precipitated, is strongly acted on by the cyanide, 
 but if it is heated to dull redness in a muffle it becomes insoluble, 
 and a large excess of this material added to a gold ore makes no 
 difference in the percentage of extraction, while the consumption 
 of cyanide is not increased by its presence. 
 
 The action of cyanide solutions on sulphide of silver is similarly 
 dependent almost entirely on its physical state. Experiments 
 conducted by Fresenius snowed that if a weak solution of silver 
 nitrate is precipitated by a weak solution of sulphide of ammonium, 
 
366 THE METALLURGY OF GOLD. 
 
 the resulting sulphide of silver is soluble in a weak solution of 
 oyanide of potassium. On the other hand, if strong solutions of 
 silver nitrate and ammonium sulphide are mixed together, the pre- 
 cipitated silver sulphide can only be dissolved by concentrated 
 solutions of potassic cyanide, and if this solution is subsequently 
 diluted with water the sulphide of silver is reprecipitated. These 
 results tend to show that sulphide of silver is not decomposed by 
 cyanide of potassium, but is held in solution by it as a hydrate. 
 Similar peculiarities in the behaviour of sulphide of silver are 
 observed in practice when ores are being treated, and in this 
 case an increase in the strength of the solution quickens the 
 action of the potassium cyanide even though dilute solutions 
 may be eventually efficacious if enough time is allowed. 
 
 A number of experiments made by Louis Janin, Jun.,* on 
 various salts of silver point to the following conclusions : 
 Silver chloride is readily soluble in cyanide, and the arseniate 
 is also rapidly dissolved. Silver sulphide and antimonide are 
 less easily acted on, but are not so refractory as metallic (cement) 
 silver, for the solubility of which see the table on p. 338. The 
 presence of copper salts appears to exercise a detrimental action 
 on the solubility of silver sulphide. 
 
 Action of Potassium Cyanide on Oxidised Pyrites. 
 When the pyrites occur in tailings which have been subjected 
 to the action of the weather for some time before treatment, com- 
 pounds are formed which are more prejudicial to the solution than 
 the sulphides. Sulphide of iron, FeS 2 , is oxidised by air and water, 
 ferrous sulphate and free sulphuric acid being formed, thus 
 
 FeS 2 + H 2 + 70 = FeS0 4 + H 2 S0 4 
 
 The protosulphate suffers further oxidation, and normal ferric 
 sulphate (Fe 2 . 3SO 4 ) is produced, which eventually loses acid and 
 becomes a soluble basic sulphate, Fe 2 s . 2SO 3 . Other basic salts 
 of complex and unknown compositions appear to be formed also. 
 
 W. A. Caldecott has investigated the products of decomposi- 
 tion of iron pyrites contained in slimes accumulated in dams or 
 pits in the Transvaal with the following results.! His view is 
 that the main stages in the oxidation of pyrites ormarcasite are 
 .as follows : 
 
 (1) FeS 2 , 
 
 (2) FeS + S, 
 
 (3) FeS0 4 + H 2 S0 4 , Ferrous sulphate and sulphuric acid 
 
 (4) Fe 2 (S0 4 ) 3 
 
 (5) 2Fe 2 8 .S0 3 
 
 (6) Fe 2 8 . 
 
 Iron pyrites. 
 
 Ferrous sulphide and sulphur. 
 
 Ferric sulphate. 
 
 Insoluble basic ferric sulphate. 
 
 Ferric oxide. 
 
 * Eng. and Mng. Journ., Dec. 29, 1888. 
 
 \-Proc. Chem. Soc., April 29, 1897. Proc. Ohem. and Met. Soc., 8. 
 Africa, July 17, 1897. 
 
CHEMISTRY OF THE CYANIDE, PROCESS. 367 
 
 The formation of ferrous sulphide proceeds rapidly, the more 
 pyritic layers of slimes containing, after a few days, as much as 
 0-25 per cent, of FeS, and scarcely a trace of soluble ferrous salts 
 and free acid. It must, however, be remembered that certain 
 ores originally contain ferrous sulphide or Fe 2 S 3 (which equally 
 yield sulphuretted hydrogen when attacked by acids), contained 
 in magnetic iron pyrites and copper pyrites. Caldecott adduces 
 no proof that free sulphur is present, but points out that the 
 slimes formed very compact masses almost impervious to air and 
 water, so that the escape of oxidised sulphur compounds would 
 not be easy. 
 
 In the presence of such oxidised copper and iron pyrites, the 
 following reactions take place : 
 
 (1) The free sulphuric acid liberates hydrocyanic acid. 
 
 H 2 S0 4 + 2KCy = K 2 S0 4 + 2HCy. 
 
 (2) Ferrous sulphate reacts on the cyanide, forming ferrous 
 cyanide, which dissolves in the excess of potassium cyanide, so 
 that it does not appear in the free state. 
 
 FeS0 4 + 2KCy = FeCy 2 + K,SO 4 
 FeCy 2 + 4KCy = K 4 FeCy 6 
 
 The potassic ferrocyanide, if sufficient acid be present, reacts 
 with fresh ferrous sulphate forming a bluish-white precipitate. 
 FeS0 4 + K 4 FeCy e = K 2 Fe 2 Cy 6 + K 2 S0 4 
 
 This precipitate oxidises in the air to Prussian blue if free acid 
 is present * 
 
 4K 2 Fe 2 Cy 6 + O 2 + 2H 2 SO 4 =3FeCy 2 . 2Fe s Cy (Prussian blue) +K 4 FeCy 6 
 + 2K 2 S0 4 
 
 Both these precipitates are decomposed by potash or soda and 
 cannot therefore be formed in their presence. The reactions 
 may be represented as follows : 
 
 K 2 Fe 2 Cy 6 + 2KOH = K 4 FeCy 6 + Fe(OH) 2 
 3FeCy 2 . 2Fe 2 Cy 6 + 12NaOH = 3Na 4 FeCy 6 + 2Fe 2 (OH) 6 
 
 Consequently, if free acid is not present Prussian blue is hardly 
 formed at all, as the solution soon becomes alkaline, and the 
 precipitate is decomposed as fast as it is formed. 
 
 It follows from these reactions that if the blue colour of 
 Prussian blue is visible in the vats or on the surface of the 
 tailings heaps, an enormous waste of cyanide must have taken 
 place, and the matter should be at once investigated. 
 
 (3) Ferric sulphates are decomposed by potassium cyanide, 
 hydrocyanic acid being evolved and ferric hydrate precipitated. 
 * Valentine's Chemical Analysis, p. 42. 
 
368 THE METALLURGY OP GOLD. 
 
 (4) A mixture of ferrous and ferric sulphates produce Prussian 
 blue by reacting with potassium cyanide, ferrocyanide of potas- 
 sium being formed at first as above ; the equation is 
 
 3K 4 FeCy 6 + 2Fe 2 (S0 4 ) 3 = 3FeCy 2 . 2Fe 2 Cy e + 6K 2 S0 4 
 
 (5) Sulphate of copper, CuS0 4 , acts differently from FeSO 4 , 
 cuprous cyanide, Cu 2 Cy 2 , being formed, soluble in excess ot KCy 
 to K 2 Cu 2 Cy 4 , a compound very prone to decomposition. Copper 
 sulphate also gives a precipitate with potassic ferrocyaiiide,, 
 thus 
 
 K 4 FeCy 6 + CuSO 4 = K 2 CuFeCy 6 + K 2 S0 4 
 
 (6) Ferrous hydrate, when formed as above, is instantly 
 dissolved in KCy, thus 
 
 Fe(OH) 2 + 6KCy = K 4 FeCy 6 + 2KOH( + 175'6) 
 
 Ferric hydrate, however formed, does not act on potassium- 
 cyanide, its only action is mechanical, as it collects in a gela- 
 tinous mass on the niters and checks the flow of liquid. 
 
 Copper and zinc in the condition of hydrates or carbonates are 
 quickly dissolved in preference to the precious metals. If 
 sulphates of these metals are formed in an ore containing lime- 
 stone or clay, double decomposition occurs with the production 
 of sulphate of lime or alumina, and oxides or carbonates of the; 
 heavy metals, which are dissolved by the cyanide, thus 
 
 ZnS0 4 + CaC0 8 = ZnC0 3 + CaS0 4 
 ZnC0 3 + 2KCy = ZnCy 2 + K 2 C0 8 
 
 The Soda Solution. Since acidity of the ore causes decom- 
 position of the cyanide, an obvious method of reducing the loss is 
 to add alkali in some form. Before doing this, the free sulphuric 
 acid and soluble salts may be removed by leaching with water, 
 and then a solution of caustic soda or lime is run on to the ore, 
 and after standing for some time is drained off and followed by 
 the cyanide solution. The insoluble basic salts are thus con- 
 verted into ferric hydrate and soluble sulphates 
 
 Fe 2 3 .2S0 3 + 4NaOH + OH 2 = Fe 2 (OH) 6 + 2Na 2 S0 4 
 2Fe 2 3 .S0 8 + 4NaOH + 4.0H 2 = ,2Fe 2 (OH) 6 + 2Na 2 S0 4 
 
 Leaching with water then removes the excess of alkali; but, as 
 this cannot be done completely, except with considerable expendi- 
 ture of time, it is usual at the Robinson Mine to use lime instead 
 of soda. Although the action on the iron salts is slower, an 
 excess of lime is less detrimental than soda to the cyanide 
 solution, and does not attack the zinc. It is found that, even 
 after treatment of the oxidised pyrites by alkalies, the loss of 
 cyanide is much greater than in the case of free milling ores. 
 The reason for this may, in part at least, be attributed to the 
 action of soda on the protosalts (such as sulphates or carbonates) 
 
CHEMISTRY OF THE CYANIDE PROCESS. 369 
 
 of copper, zinc, &c., by which these metals are precipitated as 
 hydrates, readily soluble in KCy. The preliminary washing 
 with water must always be carefully performed, until no colora- 
 tion is obtained with ammonium sulphide, so as to remove the 
 soluble salts as far as possible, but some always remain and are 
 converted into hydrates by the alkali. 
 
 It is now usually regarded as more advantageous to add lime 
 as a dry powder to the ore before it is charged into the vats, 
 instead of an alkaline solution. The necessary amount is added 
 to each truck load of tailings, and is intimately mixed with it by 
 the time it is charged into the vat. 
 
 The amount of alkali to be mixed with a charge of ore is 
 determined by laboratory experiments, adding little by little an 
 alkaline solution of known strength to a given weight of the ore, 
 until the whole is neutral to litmus paper. 
 
 Re-Precipitation of Gold and Silver in the Leaching 
 Vats. If the solution is acid there is said to be some danger of 
 a precipitation of gold previously dissolved, insoluble aurous 
 cyanide being thrown down, according to the equation 
 
 KAuCy 2 + HC1 = KC1 + HCy + AuCy 
 
 This, however, need not be feared as long as there is an excess 
 of KCy, which must all be destroyed by the acid before the 
 aurous cyanide can be precipitated. Christy, however (Am. 
 Inst. Mng. Eng., 1896), denies that aurous cyanide is precipitated 
 by acid, and supposes that aurocyanhydric acid, HAuCy 2 , is 
 formed and not decomposed until all free HCy is expelled from 
 the solution. There is danger in transferring a solution con- 
 taining gold to a vat containing pyritic material. If the latter 
 should contain any soluble salts of the heavy metals, insoluble 
 salts are thrown down, e.g. : 
 
 2KAgCy 2 + ZnS0 4 = K 2 S0 4 + ZnAg 2 Cy 4 
 
 The salt ZnAg 2 Cy 4 is probably a true double salt, but the 
 opinion has been expressed that it is merely a mixture of simple 
 cyanides. 
 
 Testing the Strength of the Solution. The ordinary 
 method of estimating the amount of potassic cyanide present in 
 a liquid is by titration with a standard solution of silver nitrate. 
 Silver cyanide is formed, and re-dissolves in the excess of potassic 
 cyanide, until one-half of the latter has been decomposed. The 
 equations are as follows : 
 
 AgN0 3 + KCy = AgCy + KN0 3 
 
 AgCy + KCy = KAgCy 2 
 
 When one-half of the KCy present has been converted to AgCy, 
 an additional drop of AgNO 3 solution causes the formation of a 
 permanent white precipitate of AgCy. The amount of silver 
 solution added is then read off, and the percentage of cyanide 
 calculated. The equation of the end reaction is 
 
 KAgCy, + AgN0 8 = 2AgCy + KNO S 
 
 24 
 
370 THE METALLURGY OF GOLD. 
 
 A few drops of a solution of potassium iodide are often added 
 to make the end reaction sharper. 
 
 This method is difficult to apply when solutions containing 
 soluble cyanides of zinc and other metals require to be titrated. 
 " A white flocculent precipitate occurs at a certain stage, pro- 
 bably consisting of simple (insoluble) cyanide of zinc, formed by 
 decomposition of the soluble double cyanide 
 
 K 2 ZnCy 4 + AgN0 3 = KAgCy 2 + ZnCy 2 + KN0 3 
 
 This precipitation occurs long before the whole amount of 
 potassium has been converted into the soluble double salt of 
 silver (KAgCy 2 ), for the solution, after the appearance of the 
 flocculent precipitate, still gives the Prussian blue precipitate 
 with acidulated ferrous sulphate."* 
 
 Mr. Bettel has devised the following methods f of testing solu- 
 tions for free cyanide, hydrocyanic acid, and double cyanides 
 respectively. 
 
 (1) Free Cyanide. 50 c.c. of solution is taken and titrated 
 with silver nitrate to faint opalescence, or first indication of a 
 flocculent precipitate. This will indicate (if sufficient ferro- 
 cyanide be present to form a flocculent precipitate of zinc ferro- 
 cyanide) the free cyanide, together with cyanide equal to 7*9 
 per cent, of the potassic zinc cyanide present. J 
 
 (2) Hydrocyanic Acid. To 50 c.c. of the solution add a solu- 
 tion of bicarbonate of potash or soda, free from carbonate or 
 excess of carbonic acid. Titrate as for free cyanide. Deduct 
 the first from the second result, and the percentage of free 
 hydrocyanic acid is obtained. 
 
 (3) Double Cyanides. Add excess of caustic normal soda to 
 50 c.c. of solution, and a few drops of a 10 per cent, solution of 
 KI, and titrate to opalescence with AgN0 3 . The zinc potassic 
 
 * Eng. and Mng. Journ., 1892, p. 417. 
 
 t Ghent. News, 1895, vol. Ixxii., p. 287. See also " Notes on the Analysis 
 of Cyanide Solutions," by J. E. Clenncl. Proc. Chem. and Met. Soc. of 
 S. Africa, 21 Dec., 1895. 
 
 I G. A. Goyder, however (Chem. News., Aug. 16, 1895, p. 80), found 
 that the amount of free cyanide shown by this method varies with the 
 dilution. Thus a sump solution from the Mount Torrens Works, Australia, 
 was found by titration to contain 0'04 per cent. KCy. On adding an equal 
 bulk of water, it appeared to contain 0'05 per cent. KCy (instead of 0"02 
 per cent.), and on diluting with two volumes of water, it showed 0'04 per 
 cent. KCy instead of '01 3 per cent. Goyder also observed that a solution 
 of 1 per cent, of crystallised double cyanide of zinc and potassium appeared 
 to have one-thirtieth of its cyanogen as free cyanide, and on diluting to 
 250 volumes (0'004 per cent. K 2 ZnCy 4 ), the whole of the cyanogen appeared 
 to be in the state of simple cyanide. Bettel's amount, 7 '9 per cent, of the 
 double cyanide, probably, therefore, corresponds only to one definite con- 
 centration. The double cyanide is doubtless dissociated more or less into 
 2KCy + ZnCy 2 in dilute solution, and the potassium cyanide so set free is 
 perhaps available, both to dissolve gold and to keep silver cyanide in solu- 
 tion, the zinc cyanide remaining dissolved in the water, in which it is not 
 quite insoluble. Some experiments of the author in progress will, it is 
 hoped, clear up this point. In any case, it seems quite likely that the free 
 cyanide shown by simple titration represents the cyanide available for the 
 dissolution of gold, so that from the practical point of view, no new methods 
 of assay are necessary. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 371 
 
 oyaiiide is decomposed by the silver nitrate, and the ZnCy 2 thus 
 formed instead of being precipitated is acted on by the soda, 
 sodium zincate being formed and some of the double cyanide 
 regenerated. This method gives the whole of the combined 
 CN present, whence the separate results can now be calculated. 
 
 According to Watts* and Fresenius,f the total amount of 
 cyanogen in a solution, whether present as simple or double 
 cyanides, may be estimated by boiling with an excess of oxide 
 of mercury and water, when all the cyanogen is obtained as 
 cyanide of mercury and the metals pass into oxides. The 
 cyanide of mercury is then precipitated by nitrate of silver, with 
 the precautions recommended by H. Rose and Finkener.J 
 
 The amount of gold in the solution is usually determined by 
 ^evaporating a known bulk to dryness with litharge, reducing the 
 lead by fusion in a crucible with charcoal, and cupelling the lead 
 button. The amount of zinc and other heavy metals present 
 may be determined by concentrating the solution by evaporation, 
 adding an excess of sulphuric acid, and heating until " almost 
 all the sulphuric acid has been expelled. The residual mass is 
 then free from cyanogen. It is dissolved in water, if necessary 
 with the addition of hydrochloric acid, and the oxides deter- 
 mined by the usual methods. This way is not adapted for 
 cyanide of mercury, as a little of the metal would escape with 
 the fumes of the sulphuric acid." 
 
 Estimation of Sulphides in Potassium Cyanide. In view 
 of the marked effect produced by soluble sulphides in checking 
 the solution of the gold, it is necessary to estimate them in 
 samples of cyanide. The former method was to precipitate the 
 sulphur by shaking with excess of lead carbonate, filter off the 
 mixture of carbonate and sulphide of lead, oxidise the precipitate 
 with bromine in presence of an alkali, filter, acidify, boil off the 
 bromine and precipitate the sulphuric acid with barium chloride. 
 
 A more rapid method devised by Feldtmann and Bettel is 
 as follows : 10 grammes of commercial cyanide are dissolved in 
 water and filtered if necessary. The solution is agitated with a 
 small quantity (in excess) of precipitated lead carbonate and 
 filtered. The precipitate is treated in a flask with a few c.cs. of 
 a solution of pure KCy or NaCy free from sulphides, sulpho- 
 cyanides, or ferrocyanides, best prepared by passing HCy in 
 excess into pure hydrate. Hydrogen peroxide, previously puri- 
 fied by shaking with ether, is added in large excess, decoloration 
 of the precipitate being complete, and all the lead sulphide con- 
 verted into alkaline sulphocyanate. Half a gramme of man- 
 ganese peroxide is then added, and the excess of peroxide 
 destroyed by agitation in about two minutes; the solution is 
 then filtered off, acidified with sulphuric acid and titrated with 
 N 
 J^Q potassium permanganate, 1 c.c. of which = 0*000053 grm. 
 
 * Diet. ofChem., 1864, vol. i., p. 202. 
 
 t Quant. Chem. Anal., 7th edition, 1876, vol. i., p. 376. $ Loc. cit. 
 
 % Chem. and Met. Soc., S. Africa, Sept. 19, 189d. 
 
372 
 
 THE METALLURGY OF GOLD. 
 
 sulphur or 0*000182 grm. potassium sulphide. The perman- 
 ganate may be standardised by pure KCNI3, 1 c.c. being equal 
 to 0-0001618 grm. KCNS. 
 
 Methods of determining the amount of available oxygen and 
 of cyanates in cyanide solutions have also been given by 
 Feldtmann and Bettel* but no results obtained in practice 
 have yet been published. 
 
 Strength of Solution Required. Below is a table, given 
 by Mr. J. S. MacArthur, showing the relative effect of weak 
 solutions up to 1 per cent., from which it appears that on 
 certain ores an extremely weak solution does practically the 
 same work in gold extraction as one eight times as strong. 
 While there is a slight tendency to raise the extraction of 
 silver by the increased strength of solution, the greater ten- 
 dency of the stronger solutions to attack the base metals is 
 shown by the fact that, where one of the stronger solutions is 
 used, the amount of cyanide consumed is equal to the whole- 
 amount present in the weaker solutions. 
 
 
 Strength 
 of Cyanide 
 Solution. 
 
 Consumption 
 of Cyanide. 
 
 Extraction. 
 
 
 
 
 
 Au. As. 
 
 
 
 Per cent. 
 
 Per cent. 
 
 Per cent. 
 
 Per cent. 
 
 Decomposed pyritic ore 
 
 
 0-125 
 
 0-090 
 
 93 
 
 70 
 
 containing 
 
 
 0-250 
 
 0-110 
 
 91 
 
 72 
 
 Au. 1 oz. 8 dwts 3 grs. 
 
 
 0-500 
 
 0-150 
 
 93 
 
 73 
 
 Ag. 1 3 3 
 
 
 1-000 
 
 0-230 
 
 91 
 
 78 
 
 Pyritic ore containing 
 Au. 3 ozs. 1 1 dwts. 1 gr. 
 Ag. 1 9 17 grs. 
 
 ! 
 
 0-125 
 0-250 
 0-500 
 1-000 
 
 0-040 
 0-050 
 0-050 
 0-150 
 
 86 
 
 85 
 89 
 
 88 
 
 76 
 74 
 
 82 
 79 
 
 Of late years solutions have been used weaker and weaker, 
 the strength reaching the minimum in treating slimes. Some 
 ores, however, cannot be advantageously treated by solutions 
 weaker than 0-2 per cent., and for very rich ores much stronger 
 solutions are still often used. 
 
 The Consumption of Cyanide varies with the ore. From 
 O'l to 0'2 Ib. per ton of ore is left mechanically mixed with th& 
 tailings unless special means are taken to remove it by washing 
 with water. The decomposition due to exposure to the air, to 
 contact with the ore, and to the reaction in the zinc boxes 
 may not be more than J Ib. to 1 Ib. per ton. In the treatment 
 of pyritic ores it is usually much gt eater, amounting to from 
 3 to 50 Ibs. per ton on Ouray ores.f Agitation of the ore with 
 the solution, by means of mixer shafts, is said, by Mr. Charles 
 
 * Ibid., Oct 17, 1896. 
 
 t Eng. and Mng. Journ., 1892, p. 606. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 373 
 
 Butters, to increase enormously the decomposition of the cyanide, 
 tut Mr. MacArthur declares that this increase is not more than 
 O05 per cent, of the ore, or little more than 1 Ib. per ton. 
 
 Methods of increasing the Speed of Action of Potassium 
 Cyanide on Gold. These methods are described here for 
 convenience, as being more intelligible after the chemistry of 
 the process has been discussed. The necessity of the presence 
 of oxygen has already been dwelt on above. It has, however, 
 been frequently pointed out that in the interior of a mass of ore 
 undergoing treatment the conditions are not favourable for the 
 maintenance of a sufficient quantity of oxygen in a free state. 
 Both the cyanides and the pyrites of the ore tend to unite with 
 it, and further absorption of free oxygen from the air is extremely 
 slow. Hence the time required for the treatment of a charge is 
 many hours, or even days, although under favourable conditions 
 the gold could be dissolved in a few minutes, or at most in 
 two or three hours. To supply the oxygen, various oxidising 
 substances have been tried, such as the passage of a current of 
 air through the solution, and the addition of various materials 
 {see above, p. 355), some of which were tried by Dixon in 1877* 
 in his unsuccessful attempt to find a process for treating refrac- 
 tory ores. These substances hasten the solvent action of 
 cyanides on metallic gold, and are becoming more and more 
 used in practice, especially in the United States, although they 
 act as " cyanicides," destroying some of the solvent by direct 
 or indirect oxidising effects. 
 
 Dr. N. S. Keith suggests f that the action of oxygen is due to 
 its strong electro-negative relation to gold in cyanide solutions, 
 and has experimented with various materials which are electro- 
 negative to gold in solutions of cyanide of potassium. In the 
 list of such materials, given by Dr. Gore, are carbon, iron, lead, 
 and mercury (see p. 356). Keith tried the effect of finely- 
 powdered carbon mixed with the ore, and continually agitated 
 with it in a cyanide solution. He found that the gold was 
 more rapidly dissolved than if no carbon had been present, and 
 supposed that some of the particles of gold came into contact 
 with the carbon and formed galvanic couples, in which the 
 electro-positive element, gold, was quickly acted on. Such 
 contact, however, could not be attained in practice except to a 
 small extent, and the method is therefore useless, whilst the use 
 of lead or iron in this way is a fortiori impossible. On the 
 other hand, mercury can be more readily subdivided and distri- 
 buted through the ore, but, as it amalgamates with the gold, the 
 conditions are changed, and, as a matter of fact, the gold in pasty 
 amalgam is only slightly more rapidly dissolved by cyanide than 
 is pure gold. 
 
 * Chemical News, December 20, 1878, p. 293. 
 \- Engineering, vol. lix. (1895), p. 379. 
 
374 THE METALLURGY OF GOLD. 
 
 From the fact that mercury is electro-negative to gold in 
 cyanide solutions, Skey in 1876 * concluded that metallic gold 
 in contact with a solution of mercury cyanide would rapidly 
 dissolve and mercury be reduced. He found this to be the case 
 alike with gold and silver, which dissolved with almost equal 
 readiness. In 1895, Keith proposed f to add a small quantity 
 (2 ozs. to 12 ozs. per ton of liquid) of potassium mercuric cyanide, 
 HgCy 2 . 2KCy, to ordinary cyanide solutions to quicken their 
 action by enabling the presence of free oxygen to be dispensed 
 with. The gold displaces the mercury from solution, and so is 
 dissolved, whilst the mercury is precipitated on the surface of 
 the particles of gold and amalgamates them. Keith supposed 
 that the whole of the gold would thus be rapidly dissolved, and 
 the precipitated mercury would then be redissolved in the 
 cyanide, and thus be ready to react as before. This view seems 
 to be incorrect. According to the author's experiments, the 
 gold is at first rapidly dissolved, and the mercury precipitated. 
 As the action proceeds, however, the dissolution of the gold 
 becomes slower and slower, the mercury appearing to protect 
 it more and more as the percentage of gold in the amalgam is 
 diminished. If the particles of gold are only moderately fine 
 (e.g., gold precipitated from the solution of the chloride by 
 sulphurous acid) the action becomes extremely slow after about 
 85 per cent, of the gold has been dissolved, the amalgam then 
 consisting of about three parts of mercury to one 'of gold. If r 
 on the other hand, very finely-divided gold is used, such as gold 
 leaf, the action is fairly rapid until about 95 per cent, of the 
 gold is dissolved, and in one case 98 per cent, of such gold was 
 dissolved in four days by a solution containing 1*5 per cent, of 
 HgCy 2 . 2KCy. In the case of gold leaf, however, which contains 
 both silver and copper, the rate of dissolution is higher than it 
 would be for pure gold in a similar state of subdivision, as the 
 presence in the alloy of either silver or copper favours the dis- 
 solution. The retarding effect exercised by metallic mercury 
 when amalgamated with the gold is exemplified by the results 
 of some experiments in which the solutions contained 1*5 per 
 cent, of HgCy 2 . 2KCy, the time of treatment was thirty-six 
 hours, and the weights and state of aggregation of the metal 
 treated were approximately the same, tinder these conditions, 
 about 86 per cent, by weight of some samples of pure gold were 
 dissolved, and only 14 per cent, of the gold contained in amalgams 
 consisting of two parts of mercury and one of gold. Dissolution of 
 gold by solutions containing mercury cyanide is greatly expedited 
 by heat. In the author's experiments on ores, the quickening 
 effect of mercury cyanide on solutions of cyanide was very slight. 
 Later Patent Processes involving the use of Cyanide 
 The Hood Process."^ Careful consideration of the known facts 
 
 * Transactions and Proceedings of the New Zealand Institute, vol. viii.,. 
 p. 334. t tinf/ineering, loc. cit. 
 
 $ This section is inserted by permission of Dr. Hood. 
 
CHEMISTRY OP THE CYANIDE PROCESS. 375 
 
 regarding the dissolution of gold by cyanide solutions, led Dr. 
 J. J. Hood, A.R.S.M., to the conclusion that the generally 
 accepted theories on the subject are only partly true. He 
 suggests that gold can only be dissolved by cyanide when some 
 other metal is present in the solution, which is displaeeable by 
 gold. Thus when gold is digested with solutions containing 
 an alkaline cyanide, together with certain compounds of 
 mercury or lead, it is dissolved, and an equivalent quantity of 
 mercury or lead is precipitated. Dr. Hood does not admit that 
 gold can displace potassium in KCy in the presence of water 
 and oxygen although there seems no reason, according to his 
 theory, for excluding this instance. 
 
 If, for example, gold is digested with an aqueous solution of 
 potassium cyanide and one or other of the chlorides of mercury, 
 the action is represented, so far as weights are concerned, by 
 the following equations : 
 
 2Au + HgCl 2 = Hg + 2AuCl 
 2Au + Hg 2 Cl 2 = Hg., + 2AuCl. 
 
 These equations do not, of course, represent the whole of the 
 interchanges. The presence of the alkaline cyanide evidently 
 assists in dissolving the gold, for a solution of HgCy 2 alone 
 in water has no action on gold. By using an excess of 
 gold, the whole of the mercury can be removed from solution, 
 and equivalents of gold dissolved as represented by the equations 
 given above, 2 x 196 '8 parts of gold being dissolved when 200 
 parts of mercury are added as mercuric chloride or 2 x 200 parts 
 as mercurous chloride. In some experiments made by the 
 author employing 0-5 gramme mercuric chloride and TO gramme 
 potassium cyanide, the amount of gold dissolved by the hot, 
 strong (5 per cent.) solution in thirty minutes was 0-724 
 gramme, whilst theory requires 0*726 gramme. Under similar 
 conditions, the amount of gold dissolved by potassium cyanide 
 alone was only O'OIO gramme. 
 
 Dr. Hood maintains that pure alkaline cyanides could not 
 dissolve gold, and that they act by means of impurities contained 
 in them. Suppose, for example, a trace of a metal were present 
 which could be displaced by gold ; the impurity would be pre- 
 cipitated like copper by zinc, or reduced like ferric chloride 
 dissolving tin or zinc, and an equivalent quantity of gold 
 dissolved. The precipitated metal could not of course be re- 
 dissolved by cyanide while there remained undissolved any 
 portion of the particle of gold on which it had been precipitated. 
 If, however, the impurity were oxidised by contact with air or 
 other means, it might be redissolved, and would then be avail- 
 able for the dissolution of a further quantity of gold. The 
 reduced compound might be similarly regenerated. In this 
 way a small amount of impurity might suffice for the dis- 
 solution of a comparatively large amount of gold. It is to 
 
376 THE METALLURGY OF GOLD. 
 
 such indirect action that Dr. Hood attributes the efficacy of 
 oxygen in promoting the solvent action of potassium cyanide. 
 
 One of the solutions used by Dr. Hood in the treatment of 
 gold ores is obtained by dissolving in water - 03 per cent, of 
 mercuric chloride and 0*06 per cent, of alkaline cyanide. He 
 finds that the mixture in this ratio is so far stable that the 
 decomposition of cyanide due to the presence of acid sulphates in 
 the ore is much less than if potassium cyanide alone were used. 
 He also adds caustic soda or carbonate of soda to the solution. 
 In the Hood process the presence of oxygen is quite unneces- 
 sary, and the treatment of concentrates is thus stated to be 
 more rapid and cheaper than it is by the Mac Arthur- Forrest 
 process. The recovery of the gold from solution is effected by 
 precipitation by means of the copper-zinc couple, which was 
 found by the author on a small scale to be far more rapid in its 
 action than zinc alone. The excess of mercury left in solution 
 is precipitated with the gold, and can be recovered. 
 
 Experiments made upon some of the Australian gold bearing 
 iron oxides and pyritic ores ranging from a few dwts. to 3 ozs., 
 as well as upon some Mexican auriferous sands carrying 700 ozs. 
 of gold to the ton are said to have given a very high percentage 
 of extraction. According, however, to some experiments made 
 by the author, it would appear that the process is inapplicable 
 to ores containing coarse particles of gold. 
 
 It is proposed to prepare the double compounds of mercury 
 from the solutions of the alkaline cyanides obtained in the 
 manufacture of cyanogen compounds through ammonia and 
 carbon bisulphide which is now being worked on a large scale 
 in England. Such double compounds are readily crystallised, 
 very stable, and easy of transport. 
 
 Ehrmann * has also recently tried the copper-zinc couple and 
 found it most efficacious in hot solutions, but even in the cold 
 his results were striking. For example, a cold solution, con- 
 taining 0-017 per cent. KCy and 7 dwts. of gold per ton, had 
 61 per cent, of the latter precipitated by zinc in 24 hours, and 
 under precisely similar conditions, 98 -67 per cent, of the gold 
 was precipitated by the copper zinc couple, the residual liquor 
 containing only 2 grains of gold per ton. 
 
 The Sulman-Teed Process. Among other suggestions for ren- 
 dering the presence of oxygen unnecessary, may be mentioned 
 that made by Messrs. Sulrnan and Teed,f who use cyanogen 
 bromide, CNBr. The addition of this substance to a solution of 
 potassium cyanide makes it three or four times more rapid in 
 dissolving gold They put forward the equation 
 
 CyBr + 3KCy + 2Au = 2K.AuCy 2 + KBr 
 
 * Chem. and Met. Soc. of S. Africa, April 17, 1897. 
 t/Voc. Institute of Mining and Metallurgy, Feb., 1895. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 377 
 
 in explanation of the action of the solvent, which is inopera- 
 tive except in the presence of an alkaline cyanide. Cyanogen 
 chloride and iodide give equally good results as far as rate of 
 solution is concerned, but are not convenient for use on a large 
 scale. In tests on concentrates, and on various complex ores, 
 the results obtained by Sulraan, Teed, and others were remark- 
 ably good, high percentages of extraction being obtained in a 
 few hours from ores which yielded little or no gold to ordinary 
 cyanide solutions in the same time. 
 
 Molloy's Method of Precipitation. Instead of using zinc, 
 Molloy proposes to precipitate the gold by a current of elec- 
 tricity, using a mercury cathode. 
 
 Electro-deposition of the gold is used in the Siemens-Halske 
 process (see p. 347) and the conditions necessary lor success 
 have been worked out and stated by the inventors ot that 
 method. One of the most important provisions is that the 
 layers of liquid between the anodes and cathodes must be very 
 thin in order that the ions may not be compelled to travel tar 
 before they reach the poles. Hence, when great quantities of 
 liquid have to be treated, it is necessary tor the electrodes to 
 have a very large surface. Moreover they should stand in a 
 vertical position so that they may not become coated with slime 
 settling from the liquids, which are usually more or less muddy. 
 It follows that mercury cannot be used in practice for the 
 cathodes. Von Gernet has calculated that 80 tons of quick- 
 silver would be required for the treatment of 100 tons of 
 solution per day, and the initial cost of the mercury would, 
 therefore, be some 20,000. Besides this, about 3,000 ounces 
 of gold, worth nearly 13,000, would be retained by the bath 
 after the whole had been subjected to nitration. Thus the 
 produce of, say, the first four months' work would be perma- 
 nently locked up, until the mercury was re-distilled. 
 
 Messrs. MacArthur & Forrest experimented on the action of 
 sodium amalgam, using a small tower filled with pieces of sodium 
 amalgam, through which the solution of gold in potassium cyanide 
 was allowed to trickle slowly. They came to the conclusion 
 that sodium was not more efficacious than zinc, but was much 
 more expensive. The saving of expense in the production of 
 the bullion from the precipitate was outweighed by the cost of 
 the sodium required to form the precipitate. 
 
 The Pelalan-Clerici Process. In this process, the finely 
 crushed ore is agitated in vats (holding 25 tons each) with a 
 solution of about 04 per cent. KCy. The bottom of the vat is 
 lined with amalgamated copper plates covered with a layer of 
 mercury. A current of electricity is passed from the sheet-iron 
 agitators, forming the anodes, to the mercury bath, which is the 
 cathode, so that gold dissolved by the cyanide begins to be 
 precipitated at once. Several obvious disadvantages incidental 
 
378 THE METALLURGY OP GOLD. 
 
 to the process present themselves. However, it was worked at 
 the De Lamar mill in Idaho in 1897,* using very thick pulp. 
 The practice there seemed to resemble pan amalgamation, with 
 the additions of an electric currentj and some cyanide in the 
 water. From the details, it may be gathered that simple pan 
 amalgamation would have been more profitable. 
 
 The Johnston pro< ess consists in the use of pulverised charcoal 
 arranged in filters to precipitate the gold. Christy's experiments 
 among others have shown the process to be without practical 
 value. 
 
 The Wilde process f is named after its inventor, Prof. P. de 
 "Wilde, who proposed to use a solution of 0*05 per cent. KCy to 
 dissolve gold from its ores, then to acidify the solution with 
 sulphur dioxide, and to precipitate aurous and cuprous cyanides 
 with sulphate of copper, leaving the solution to settle for 
 12 hours. This method involves the destruction of the cyanide, 
 and so, in cases where it is profitable to do so, the solution is 
 acidified, and the cyanide is precipitated as FeCy 2 by means of 
 ferrous sulphate before the gold is thrown down. Prof. Christy 
 proposes to modify this process by acidifying the solution with 
 sulphuric acid, volatilising the free hydrocyanic acid, by passing 
 air or steam through the solution and condensing it in potash 
 or soda solution, which is allowed to trickle down through 
 scrubbing towers. In this way 80 or 90 per cent, of the 
 potassium cyanide is regenerated, and no gold is precipitated. 
 Freshly precipitated sulphide of copper, or any cuprous salt, is 
 then stirred in, when the whole of the aurous cyanide is precipi- 
 tated in a few hours, and the gold may be recovered by treating 
 the precipitate with a solution of sulphide of sodium and de- 
 positing the gold by electrolysis. Although these processes are 
 of great theoretical interest, it does not appear likely that they 
 will ever be of any practical importance. 
 
 Description of Ores suitable to the Cyanide Process. 
 Up to the present the ores on which the most striking success 
 has been obtained on the large scale have been the tailings of the 
 free milling ores of the Witwatersrand Gold Field, in which the 
 gold is finely divided, and the amount of pyrites present is very 
 small. In a large number of cases in South Africa, pyritic ores 
 have been treated with great success as far as the solution of 
 the gold is concerned, over 90 per cent, having been extracted, 
 but the consumption of cyanide is considerable. Ores contain- 
 ing sulphide of silver, mixed with base sulphides, often yield no 
 silver at all. It is, however, no longer doubtful that concen- 
 trates and other highly pyritic materials can often be treated 
 more cheaply by cyanide solutions than by roasting and chlorin- 
 ation. 
 
 * Eng. and Mng. Journ., Aug. 7, 1897, p. 155. 
 
 t Revue Universelle des Mines, de la Metallurgie, <fcc., Oct. 1, 1895. 
 Brussels. 
 
CHEMISTRY OF THE CYANIDE PROCESS. 379 
 
 The presence of decomposing marcasite, especially if some 
 copper is contained in the sulphides, is frequently fatal to the 
 process, as great quantities of cyanide are destroyed by contact 
 with such ores, the amount often exceeding 50 Ibs. per ton, 
 even after careful washing with water and treatment with alkali. 
 However, the concentrates of many mines can be treated suc- 
 cessfully, even if they consist chiefly of sulphides of iron, lead, 
 zinc, &c. The time of treatment is in these cases often as much 
 as three or four weeks, and the charges in the vats are some- 
 times drained and stirred up or transferred to other vats. 
 Nevertheless, high percentages of extraction are obtainable 
 from such materials. 
 
 The process is particularly applicable to low grade ores, con- 
 taining only finely-divided gold and small quantities of base 
 metals. O wing to the necessity for comparatively coarse crushing 
 which exists with all wet processes, and the difficulty of handling 
 great quantities of dilute solutions of the precious metals without 
 loss (which absolutely precludes the use of sufficient solution and 
 wash water to remove the whole of the soluble gold from any 
 ore), the percentage of extraction possible with the majority of 
 ores only amounts to from 70 to 90 per cent. When coarse 
 particles of gold are present, they must be removed by amalga- 
 mation, before treatment with cyanide. Ores containing coarse 
 gold cannot be treated by any wet process, no solution being 
 sufficiently rapid in its action. 
 
380 THE METALLURGY OF GOLD. 
 
 CHAPTER XVII. 
 PYRITIC SMELTING. 
 
 THE best known and most extensively practised smelting 
 processes for the treatment of gold and silver ores viz., lead 
 smelting, copper matte smelting, and smelting for the direct 
 production of copper bottoms, in which the precious metals are 
 concentrated may be best dealt with in the volumes in this 
 series devoted to Lead, Copper, and Silver, and will not be 
 described here. A brief account of iron matte smelting is 
 appended, however, as its main object is the treatment of purely 
 gold ores. This system is said by Eissler * to have originated 
 in Hungary. It consists in fusing auriferous iron pyrites 
 in a blast furnace, with the object of obtaining a regulus of 
 iron, in which the gold is concentrated. The richness of the 
 regulus, under the original system, is increased by repeatedly 
 fusing it with fresh ores, or by alternately roasting and fusing 
 it until the percentage of gold has risen to a certain limit, 
 which varies in practice, but never exceeds 50 ozs. per ton, after 
 which the gold is extracted from this product by some other 
 method, either by roasting and chlorination, or by lead or copper 
 smelting. In the United States, the production of a rich matte 
 in one operation is effected by mixing the ore judiciously, and 
 burning out part of the sulphur in the furnace. It has been 
 proposed by Mr. H. Lang f to restrict the use of the term 
 " pyritic smelting " to the reduction of gold and silver ores in 
 blast furnaces, with the formation of a rich matte in one opera- 
 tion, the distinguishing feature of the work being the use of the 
 sulphur in the ore as a fuel. This description applies to the 
 system invented by Dr. W. L. Austin, of Denver, and now in 
 use in several localities in Western America. The definition, 
 however, is somewhat narrow, and, in this chapter, pyritic 
 smelting is taken as meaning iron matte smelting. The method 
 is especially applicable in districts where no lead ores are to be 
 obtained, where fuel is cheap, and where there are available large 
 quantities of iron pyrites containing a small quantity of gold, 
 with which purely quartzose ores can be mixed if it is desirable. 
 Iron pyrites, as is well known, on being heated with a limited 
 supply of air, may be made to lose about half its sulphur, and is 
 then converted mainly into FeS, which is readily fusible. By 
 
 * Metallurgy of Gold, London, 1891, p. 378. 
 \Eng. and Mng. Journ., Dec. 26, 1891, p. 721. 
 
PYRITIC SMELTING. 381 
 
 increasing the supply of air, the amount of matte produced may 
 be reduced to any required extent, the iron being oxidised and 
 slagged off. If auriferous pyritic ores containing a quartzose 
 matrix are in course of treatment, a flux must be added to slag 
 off the quartz. At the Deadwood and Delaware pyritic smelting 
 works in South Dakota, nearly pure quartzose ores are treated^ 
 and here the flux (in this case pure limestone) is said to cost 
 about 25 cents per ton of ore.* A matte is thus formed beneath 
 the layer of slag, and is drawn off for further treatment. 
 
 On account of the comparatively high temperature at which the 
 matte solidifies, the ordinary blast furnace, with a deep crucible, 
 used in lead smelting, is thought in Hungary to be unsuitable for 
 pyritic smelting, as the matte chills as soon as it sinks below the 
 smelting zone and freezes up the tap-hole, so that tapping is 
 rendered a difficult operation. Consequently, in Hungary, the 
 " Spur-Ofen " has been substituted, in which there is no crucible, 
 the bottom of the furnace sloping to a tap-hole immediately 
 below the tuyers. The tap-hole is kept open continually, and 
 the matte, as fast as it is melted, flows through it by a narrow 
 channel into wells placed outside the furnaces, where it is- 
 separated from the slag by gravity, and is tapped into moulds, 
 while the slag overflows into ordinary slag-pots on wheels. 
 There seems no reason, however, why the methods adopted for 
 keeping metallic copper molten and in a fit condition for tapping 
 at Terrazas, Chihuahua, Mexico, by Mr. H. F. Collins,f should 
 not be equally applicable to pyritic smelting. 
 
 The matte thus obtained is sometimes subjected to partial 
 roasting and then mixed with further quantities of crude auri- 
 ferous pyrites and smelted again, until the product is sufficiently 
 rich. An iron matte can be advantageously made richer than 
 lead without undue losses in the slag, but the limit is reached 
 when it is worth about 200 per ton. The losses in the slag at 
 Mineral City, Idaho, and at Deadwood are said to be about $1 
 per ton in gold and silver together, with mattes of nearly this 
 value. These mattes are produced by the Austin process, which 
 was introduced at Toston, Montana, in 1890, and is now in use 
 at Leadville, Colorado, and at Mineral, Idaho. Details of the 
 work at these places are given below. 
 
 If the sulphides treated contain a high percentage of sulphur, 
 this is of great value as fuel, and the supply of coke may be 
 diminished to a corresponding extent. At the Bimetallic 
 Smelting Company's works, at Leadville, no fuel at all other 
 than sulphur is used, except an occasional charge of coke when 
 some irregularity occurs. J Eaw sulphide ores, direct from the 
 
 * Eng. and Mng. Journ., Jan. 14, 1893, p. 28. 
 
 t Smelting Gold and Silver Ores, Proc. Inst. Civil Eng., vol. cxii. (1893),. 
 part ii. 
 
 %Eng. and Mng. Journ., Feb. 4, 1893, p. 99. 
 
382 THE METALLURGY OF GOLD. 
 
 mine, are fed into the furnace, with the necessary proportions of 
 quartzose ore and limestone, to form slag. The furnace is of 
 peculiar construction, and was designed and built by the Colorado 
 Iron Works, Denver. The following results, however, were 
 obtained by the use of an old blast furnace, which had been 
 previously used for lead smelting. The internal dimensions of 
 this furnace were 36 inches by 80 inches, and it was altered and 
 adapted for the Austin system of pyritic smelting. A hot-blast 
 stove was erected, capable of delivering the required quantity of 
 air heated to 400C., and the other machinery consisted of one 
 100-H.P. Buckeye engine, two 40-H.P. boilers, and heavy line 
 shafting extending through the works conveying power to the 
 blowers, rock-breakers, slag hoist, &c. This machinery is stated 
 to have been enough to satisfy the requirements of six blast- 
 furnaces, but nevertheless it was found that one 40-H.P. boiler 
 was not quite sufficient when one furnace was at work. In a 
 trial run in this furnace, in March, 1892, 1,206 tons of ore 
 were smelted with 216 tons of limestone as flux in twenty-five 
 days. The amount of coke burnt was 6 per cent, of the weight 
 of the ore, its use being mainly to support the fine ores. The 
 hot-blast stove was heated by oil, which was also employed to 
 generate steam. The following is a summary of the cost per 
 ton of ore smelted in this run : 
 
 Coke, at $7 '50 per ton, $0'47 
 
 Oil, at $1-10 per ton, 0'49 
 
 Limestone, at $l - 90 per ton (using 18 per cent.), . 0*34 
 
 Labour and sundries, ...... 2*62 
 
 Total cost per ton, . . . . $3*92 
 
 The ores used were stated to contain about 12 per cent, of 
 sulphur and 1 ton of matte was produced from every 9 tons of 
 ore. The richest matte obtained contained gold 0-33 oz., silver 
 258 ozs., copper 14-2 per cent., the loss being, gold 3'43 per 
 cent., and silver 4*15 per cent, of that contained in the ore. It 
 appears that, as a result of this trial, continuous work on a large 
 scale has been begun and is meeting with considerable success. 
 
 At Mineral, ores containing approximately silica 30, calcite 
 30, iron 10, sulphur 12-5, and arsenic 2-5, with some zinc, are run 
 down in one operation into a very rich matte, without the use 
 of fluxes, and without admixture of other ore, using 7 per cent, 
 of coke. Most of the sulphur and arsenic is burnt off and the 
 corresponding proportion of iron and zinc allowed to go into the 
 slag, thus effecting a desirable concentration of the matte, and at 
 the same time utilising the heat of combustion of the elements 
 named.* Mr. Lang, the manager of these works, also states f 
 that baryta is not disadvantageous in pyritic smelting, and he 
 
 * Eng. and Mng. Journ., March 18, 1893, P- 244. 
 \llid., April 22, 1893. 
 
REFINING. 383 
 
 does not consider that the presence ol even 25 per cent, of heavy 
 spar would render an ore unsuitable, although it would be 
 almost hopeless to attempt to treat such an ore by lead smelting. 
 One of the main reasons for this difference is the fact that 
 sulphates do not increase the percentage of matte formed in 
 pyritic furnaces as they do in lead smelting, their acid being 
 volatilised unchanged in the former case, but reduced by the 
 <joke in lead smelting. Mr. Lang has found in practice, at 
 Mineral, that the best smelting mixture contains approximately 
 silica 30 per cent., sulphur from 10 to 15 per cent, (the larger 
 the percentage of sulphur the less fuel is required), iron 10 per 
 cent., lime, magnesia, baryta, &e., 30 per cent. A few per cent, 
 of zinc, lead, copper, <kc., do no harm, and the lead and copper 
 will be retained in the matte. Arsenic is advantageous as it 
 economises the fuel. The cost of smelting such a mixture, and 
 refining the matte is about $3 per ton in Western America, at 
 points conveniently situated on railroads, within a moderate 
 distance of a coal-field. 
 
 Pyritic smelting, for treating gold ores, is as yet in its infancy, 
 and few details of working have been published by the managers 
 of the various works. It may possibly be found applicable to 
 the deep-level pyritic ores in South Africa and elsewhere. 
 
 CHAPTER XVIII. 
 THE REFINING AND PARTING OF GOLD BULLION. 
 
 General Considerations. By whatever process gold may have 
 been extracted from its ores, it is necessary to melt the crude 
 bullion and cast it into bars so that its value may be ascertained, 
 and that it may be put into a form convenient for transportation 
 and sale. The name " bullion " may be conveniently restricted 
 to the precious metals, refined or unrefined, in bars, ingots, or 
 any other uncoined condition, whether contaminated by admix- 
 ture with base metals or not. It is, however, often applied to 
 coin, and the appellation "base-bullion" is given to the pig-lead 
 or to copper bottoms or pig-copper, which have been obtained in 
 smelting operations, and which may only contain a few parts per 
 thousand of gold and silver, the main portion consisting of base 
 metals. The treatment of base-bullion, however, properly be- 
 longs to the metallurgy of argentiferous lead, and copper, and the 
 descriptions given in this chapter apply only to bullion which 
 consists chiefly of gold and silver. Refining operations which 
 
384 THE METALLURGY OF GOLD. 
 
 involve cupellation on a large scale may also be more con- 
 veniently considered under the heading of the Metallurgy of 
 Silver. 
 
 The operations to which the retorted metal, gold precipitate 
 or bars from the chlorination mills, &c., are subjected may be 
 summarised as follows : 
 
 1. The bullion is melted in crucibles (a rough refining opera- 
 tion being usually effected at the same time) and cast in ingot- 
 moulds. 
 
 2. Assay-pieces are cut from the cast ingots or dipped from 
 the molten metal before pouring, and assays are made on these, 
 by which the value and composition of the bars are ascertained. 
 
 3. The bars are then usually sold to the refineries, where the 
 base metals are eliminated and the gold and silver separated by 
 " parting," and cast into bars separately. Both before and after 
 the parting it is sometimes necessary to subject the bullion to 
 further refining operations. The bars of gold and silver thus 
 obtained, being of a high degree of purity, are in a condition 
 to be used for minting, or for the various industrial purposes ta 
 which they are applied. 
 
 Rough unrefined gold is frequently sold to the refineries 
 attached to the American and Australian mints, in the state of 
 retorted metal, &c., without being previously melted and assayed, 
 the producing mills relying on the good faith of the officials at 
 these establishments. 
 
 REFINING. 
 
 Composition of Bullion. Bullion varies greatly in composi- 
 tion, and gold may be present in any proportion from zero up to 
 nearly 100 per cent. Native gold always contains more or less 
 silver, but silver quite free from gold is not uncommon. The 
 Mount Morgan gold is the finest gold which has yet been found; 
 this is 997 fine in gold, the alloying metals being chiefly copper 
 with a trace of iron. The gold obtained in most chlorination 
 mills is of a high degree of purity and rarely contains much 
 silver. This precipitated gold, however, generally makes brittle 
 bars owing to the presence of a few parts per thousand of lead, 
 bismuth, antimony, and other metals of high atomic volume. 
 Prom some chlorination mills the gold is far from pure, owing to 
 various causes, which include lack of care. If ferrous sulphate 
 is used as the precipitant, the precipitate may contain large 
 quantities of ferric hydrate from which some iron is reduced in 
 the crucible, and if sulphuretted hydrogen is used and the gold 
 precipitated as sulphide, it is contaminated with all the heavy 
 metals contained in the solution, copper, iron, and lead being 
 most often encountered. These may amount to several per cent. 
 
REFINING. 
 
 385 
 
 "Retorted metal is of very different degrees of fineness, according 
 to the nature of the ore and the course of treatment. Placer 
 gold is usually finer than that derived from lodes, containing a 
 smaller percentage of silver, while the nature of the material 
 treated and the methods used in placer operations are not 
 favourable to the contamination of the bullion with base metals, 
 which vary in amount only from to 20 parts per thousand and 
 seldom approach the latter figure. The average fineness of the 
 placer gold obtained in the United States is given in the following 
 table, which contains details concerning the chief producing 
 States : * 
 
 
 AVERAGE COMPOSITION. 
 
 
 Gold. 
 
 Silver. 
 
 Base Metals. 
 
 California, . 
 
 883-6 
 
 112-4 
 
 4-0 
 
 Idaho, 
 
 780-6 
 
 213-4 
 
 6-0 
 
 Montana, . 
 
 895-1 
 
 100-9 
 
 4-0 
 
 Oregon, 
 
 872-7 
 
 123-3 
 
 4-0 
 
 The finest placer gold from California is about 980 fine, viz., that 
 from El Dorado County, and the basest is about 720 fine. 
 
 Australian placer gold is in general finer than Californian, 
 averaging about 950 fine, the remainder being chiefly silver j 
 the silver in gold from batteries has been increasing in amount of 
 late years. Gold from New South Wales contains from 15 to 350 
 parts of silver per 1,000 ; that from Victoria averages 940 fine in 
 gold, containing only 45 parts of silver and 15 of base metals per 
 1,000. The Queensland gold, except that from Mount Morgan, 
 is less fine, containing about 120 parts of silver per 1,000 on an 
 average. Australian gold is generally brittle before refining, 
 owing to the presence of small quantities of lead and antimony. 
 
 Battery retorted gold is usually much less pure than placer 
 gold, the percentage of base metals being in particular much 
 higher, a state of things due in great part to the difference in the 
 method of treatment. In California, the bullion is usually from 
 750 to 800 fine in gold, seldom falling below 600 nor rising above 
 850, though in rare cases it is as much as 920. The impurities 
 are principally silver, copper, and iron, the greater part being 
 copper and iron in many cases. 
 
 The battery gold from the Transvaal is on an average about 
 850 fine in gold, a large part of the remainder being silver. 
 
 The bullion from pan-amalgamation is less fine than battery 
 gold, containing less gold, more silver, and more base metals. 
 
 * Report of the U.S. Census, 1880, vol. xiii., p. 352. 
 
 25 
 
386 THE METALLURGY OF GOLD. 
 
 Some of that produced in California by the Reese Elver and 
 Washoe processes, which are not described in this volume, is 
 only 500 or 600 fine in silver or even lower, with a few parts of 
 gold per 1,000, and the remainder consisting chiefly of lead and 
 antimony, or copper. 
 
 The Furnace. The furnace used for melting the bullion is 
 of simple construction. It is usually square, with walls about 
 12 inches thick, consisting of an outer layer of ordinary brick 
 and an inner layer, at least 4 inches thick, of the best fire-brick. 
 There is often a complete outer casing of iron, which is useful in 
 keeping the furnace from falling to pieces, but radiates more 
 heat than the bricks. The fire-box is about 1 foot square and 
 from 14 to 18 inches deep ; below it is an ashpit, provided 
 with a working iron door, through which the air-supply of the 
 furnace is made to pass, and by which it can be regulated. 
 The fire-bars are movable, and their ends rest loosely on iron 
 supports. The top of the furnace may be made flat or 
 sloping up towards the back at an angle of about 30. In 
 this case a wide flat ledge should be provided at the front, on 
 which crucibles and moulds can rest. The top is always made of 
 a cast-iron flanged plate, with an opening of the same area as 
 the fire-box. This opening is closed by a cast-iron sliding door 
 made in one or two pieces, and preferably lined with fire-brick 
 and running on rollers. The flue is placed at the back of the 
 fire-box near the top ; its cross-section should have an area of 
 about 16 or 18 square inches 6 by 3 inches, and 4 inches square 
 are both convenient sizes. The flue communicates with a 
 chimney, which must be of brick for the distance of 2 or 3 feet 
 from the furnace, but may be of wrought-iron tubing in its upper 
 part. The height of the chimney will depend on the position of 
 the furnace, and should be as great as possible, 60 feet giving 
 better results than any less amount. Some authorities consider 
 that a height of 30 feet is the minimum that can be allowed in 
 order to ensure a good draught, but very satisfactory results can 
 be obtained with a chimney only 16 feet high. The furnace can 
 be built by any bricklayer acting under directions. No mortar 
 is used in its construction ; clay, mixed with an equal bulk of 
 sand, being substituted for it. A sliding damper in the flue at 
 a convenient height above the ground is necessary, so as to 
 regulate the draught. The fuel used in such a furnace may be 
 anthracite, charcoal, or good coke, made in ovens, not in gas- 
 retorts, and broken into pieces of about the size of a hen's egg. 
 If the coke is of high quality, it is the most satisfactory fuel, 
 making a hot fire and lasting for a long time, so that it does not 
 require very frequent replenishing. Neither dust, nor very 
 small, nor very large pieces must be used. Charcoal is preferred 
 in the United States Mints for small charges, and anthracite for 
 large ones. 
 
REFINING. 387 
 
 The Crucibles. The bullion is melted in either graphite or 
 clay crucibles. If nitre is used to refine the metal, the graphite 
 pots are sometimes coated inside with clay. The amount of 
 refining that can conveniently be done in this way is limited 
 by the fact that the molten oxides produced rapidly corrode the 
 crucibles, and may perforate them in course of time, thus causing 
 loss. The Salamander crucibles, manufactured by the Battersea 
 Crucible Company, are the best graphite pots, as they require 
 very little annealing, and will stand frost without being dis- 
 integrated. A No. 20 crucible, of 9 inches in height, and 
 holding 400 to 500 ozs. of bullion, can be used in the furnace 
 described above. The size of the crucibles and the weight of the 
 charges of bullion vary greatly, but in extraction mills, as a 
 general rule, a gold-charge does not exceed 400 ozs., and a silver- 
 charge 1,200 ozs. in weight. In mints and refineries, much larger 
 crucibles are employed, holding different amounts up to 6,000 
 ozs. of metal. 
 
 Melting the Bullion. All crucibles must be thoroughly 
 annealed before being used ; otherwise, the contained moisture 
 being suddenly converted into steam when the crucible is heated 
 rapidly, the pots fly in pieces. The crucible is kept on a shelf 
 near the flue, for as many days or weeks as convenient, before 
 being used. It is then placed on the top of the furnace or in the 
 ashpit for a few hours, when it will probably be safe to hold it 
 over the open furnace by means of the crucible tongs, until it 
 becomes gradually warm. After a few minutes, the crucible 
 being turned round at intervals, it can be lowered rim down- 
 wards upon the burning fuel, and as soon as the rim becomes 
 red-hot, the crucible is quite safe, and may be turned over and 
 placed in position for the reception of the gold. With Sala- 
 mander crucibles, a less degree of care in annealing will suffice, 
 as they are well annealed before being sold. The crucible rests 
 on a fire-brick about 3 inches thick, which is laid on the bars of 
 the grate. If the fire-brick were omitted, the bottom of the pot, 
 resting directly on the fire-bars, would be too cold, while a layer 
 of fuel, if placed below it, would soon be burnt out, and could 
 not readily be replaced, so that the pot would sink down to the 
 bars. The fuel is built up round the pot until it reaches to 
 its rim, and the fire urged slightly until the whole pot is at a full 
 red heat. One or two spoonfuls of borax are then thrown into 
 the crucible by means of a scoop or wrapped in paper ; this 
 flux not only assists the metal to fuse, but slags off the earthy 
 impurities, and makes the metallic oxides more liquid. As soon 
 as the borax is melted, the introduction of the bullion is com- 
 menced. The safest way to do this is to use the shoot shown 
 in fig. 57, which is held in position, its lower edge being 
 inside the crucible, with the left hand, while the metal is 
 transferred to it in a scoop by the right hand. In this way the 
 
388 
 
 THE METALLURGY OF GOLD. 
 
 melter avoids all danger of loss which might be encountered if 
 the metal scrap were wrapped in paper and added by the tongs. 
 Large pieces of metal are added by the crucible tongs. The 
 cover, which must also have been previously well annealed, is 
 kept on the crucible as much as possible, and the fuel pushed 
 down with the poker to avoid scaffolding, and fresh pieces of 
 coke added when required. The crucible is not allowed to 
 become more than two-thirds full at any time, but more metal is 
 added when the first supply has been melted down, and the 
 operation repeated until the pot is sufficiently full of molten 
 material. 
 
 Fig. 57. 
 Scale, 1 in. = 9 ins. 
 
 Refining the Bullion. If the bullion is of a high degree of 
 purity, containing but little dirt or base metals, not much flux 
 is added, a spoonful or so of carbonate of soda and nitre being 
 enough. In this case the slag is not skimmed off but poured 
 with the metal. If the bullion is very base, however, it is usual 
 to refine it partially by adding nitre and borax, a little at a time, 
 and skimming off the slag when all action has ceased. The nitre 
 exercises a powerfully oxidising effect on the base metals in the 
 bullion, and the resulting oxides form a liquid slag with the 
 borax. When graphite crucibles are employed, the nitre must 
 be prevented from coming in contact with the sides, as in that 
 case the carbon would be oxidised and the pot rapidly corroded. 
 On the other hand, clay pots do not withstand the action of 
 molten oxides slagged with borax. For these reasons a favourite 
 practice is to use graphite pots, covering the surface of the molten 
 metal with bone-ash sprinkled on, the layer being thickest round 
 the sides. Holes are made near the centre of this cover with an 
 iron rod, and nitre introduced through them, in small amounts 
 at a time. As the fusible oxides are formed they are absorbed 
 by the bone-ash and prevented to some extent from attacking; 
 
REFINING. 389 
 
 the crucible. When sufficient nitre has been added, a point 
 which is judged by an experienced mel.ter from the appearance 
 of the surface of the molten metal, the slag should be of a moder- 
 ately pasty consistency suitable for skimming. If it is too liquid 
 it is difficult to skim and must be thickened with bone-ash ; if it 
 is too pasty, shots of gold may become entangled in it, and it 
 must be thinned with borax. 
 
 Mr. Hanks, the late State mineralogist, describes the operation 
 of skimming, at the San Francisco Refinery, as follows : * " A 
 skimmer is prepared by bending the end of an iron rod of a 
 J inch diameter into a spiral of about 1J inches in diameter, 
 shaped so that when the skimmer is let down vertically into the 
 crucible, the spiral will lie flat upon the surface of its contents. 
 When the slag and metal are perfectly fluid, the surface of the 
 former is touched by the skimmer, to which some slag adheres. 
 It is then withdrawn, quenched in a bucket of water, and at 
 once replaced in the crucible, thus causing a further small 
 portion of the slag to solidify. This operation is repeated until 
 the greater part of the slag has been removed from the crucible. 
 Care must be taken not to allow the bare iron to come in contact 
 with the molten gold, as in that case some of the latter would 
 adhere to it, and for this reason the slag is left adhering to the 
 skimmer during the latter part of the operation. The wet 
 skimmer must not be plunged below the surface of the molten 
 metal or an explosion would ensue. If too much slag accumu- 
 lates on the skimming tool it is detached by quenching and 
 hammering." 
 
 If the bullion is very base the addition of bone-ash, nitre, and 
 borax, followed by skimming, may have to be repeated two or 
 three times. The method was formerly known as the " Poussee " 
 process. Lead is not readily removed from bullion by the action 
 of nitre, which is best adapted for the oxidation of copper. 
 When much lead is present, alternate additions of sal-ammoniac 
 and nitre are made, by which the lead is rapidly oxidised. It is 
 probable that the sal-ammoniac acts by decomposing basic com- 
 pounds of lead which resist the action of nitre. In obstinate 
 cases, a blast of air is directed upon the surface of the molten 
 metal, and the lead is in this way rapidly oxidised and slagged off. 
 Sometimes sal-ammoniac is sprinkled on, to remove lead, after 
 skimming. 
 
 Other fluxes which are sometimes used are sand, pearlashes, 
 and metallic iron. Sand is added to assist in forming a liquid 
 slag, and to protect the crucible from corrosion by the oxides, 
 especially when iron is present. Pearlashes are sometimes added 
 when the bullion contains tin. 
 
 Bars which contain antimony or arsenic can be rapidly refined 
 by stirring briskly with an iron bar, a little nitre being added to 
 * Calif ornian State Mineralogist's Report, 1884. 
 
390 THE METALLURGY OF GOLD. 
 
 the charge. After three or four minutes' stirring, the greater 
 part of the antimony will have been removed as antimonide of 
 iron. 
 
 The refined gold should now be of a brilliant green colour, and 
 its surface should remain quiet, without showing any iridescent 
 films or other signs of continued oxidation. 
 
 Toughening the Bullion. After melting with nitre, bullion 
 is sometimes toughened before being poured, as small quantities 
 of lead, antimony, arsenic, and bismuth are still retained and 
 render it brittle. Nearly fine gold, which is the product of 
 chlorination mills or of parting operations, is similarly treated. 
 The toughening is usually done in one of three ways, viz. : 
 
 1. Sal-ammoniac and corrosive sublimate are added to the 
 molten metal. 
 
 2. The metal is melted with oxide of copper. 
 
 3. Chlorine is passed through the molten metal. 
 
 The method of procedure in each case is as follows : 
 1. Sal-ammoniac is sprinkled on to eliminate the lead and tin, 
 after which repeated small additions of powdered corrosive sub- 
 limate (mercuric chloride) are made. After each addition the 
 door of the furnace must be at once closed, as dense poisonous 
 fumes arise and must not be breathed by the workers. Volatile 
 chlorides of zinc, copper, antimony, bismuth, &c., are formed and 
 pass off, carrying with them some gold, of which there is an 
 appreciable loss. A little corrosive sublimate sprinkled on the 
 surface of molten gold will completely toughen every part of it 
 without being mixed with it by stirring, even although the 
 crucible contains several hundred ounces of the metal. 
 
 When the metal is supposed to be tough, a small sample 
 is dipped out and made into a thin ingot, which, after it Las 
 been cooled in water, is doubled up by hammering and its 
 degree of toughness thus tested. It is then often remelted with 
 copper to make up the standard alloy of the country, and again 
 cast and hammered or cut in two with a shearing machine. The 
 reason for doing this is that impure gold, although it may be 
 tough when unalloyed with copper, may make brittle standard 
 bars. If the gold is still found to be brittle, the main bulk of 
 it left in the crucible is subjected to a repetition of its former 
 treatment as often as is necessary, and as soon as the toughening 
 is complete, the gold is covered with a layer of charcoal in the 
 form of powder or lumps and thoroughly stirred before being 
 poured. 
 
 The melting under charcoal is sometimes necessary to render 
 silver bars fit for coinage when they have been treated for a long 
 time by nitre. When silver has been raised to a high degree 
 of fineness, it is affected by a peculiar bubbling which may be 
 due to the evolution of oxygen previously absorbed from the 
 nitre. If this continues it is necessary to stir continuously 
 
REFINING. 391 
 
 with a graphite rod, keeping the surface covered with charcoal 
 powder, until the bubbling ceases. If the metal, while still 
 effervescing, is poured into a mould, it sprouts at the surface, 
 and a shower of extremely minute particles are projected, often 
 to some distance from the mould, requiring to be swept up ; 
 if the crucible is covered by a lid, very heavy effervescence 
 ensues when the lid is lifted. In this case the silver ingots 
 formed are not marketable, being brittle, of low density, and 
 covered by heavy efflorescences. 
 
 2. Black oxide of copper is less frequently employed. It is 
 stirred in with the molten metal, and the whole then allowed to 
 remain in the furnace for about half an hour, with occasional 
 stirring. Antimony, arsenic, bismuth, &c., are oxidised at the 
 expense of the oxide of copper, and volatilised or slagged off 
 with borax. It is stated that 2 per cent, of antimony can thus 
 be removed. The process is efficacious, but the pot is rapidly 
 corroded, and the gold is of course contaminated with the reduced 
 copper, a matter which, however, is of little consequence, for the 
 reason that the pure metal is seldom required. 
 
 3. The use of gaseous chlorine is described under the heading 
 of Miller's process of Parting, p. 414. 
 
 The time occupied in refining and toughening a crucible full 
 of gold of course varies greatly, but often occupies from one to 
 three hours after complete fusion has been effected. 
 
 Casting the Ingots. It is necessary to stir the charge 
 thoroughly before pouring, as the bar must be as homogeneous 
 as possible to insure a correct assay. Since segregation may occur 
 on cooling, assay pieces are often dipped out immediately after 
 stirring. The subject of taking assay pieces is further con- 
 sidered in Chapter xx., p. 461. The stirring is usually done 
 with a peculiarly shaped graphite rod, made expressly for the 
 purpose. It is annealed carefully and raised to a full red heat 
 before being introduced into the crucible, and is held firmly by a 
 pair of tongs with special concave curved faces to its jaws, so as 
 to fit the round rod. In the case of very small meltings it is 
 sufficient to lift the crucible out of the furnace with the tongs 
 and to give it a rotary motion just previous to pouring. In 
 doing this, the metal must not be allowed to cool too much or 
 the casting will be defective. It is advisable to close the 
 dampers wholly or in part, so as to check the draught, when 
 stirring is being done. 
 
 Meanwhile the ingot-mould in which the gold is to be cast has 
 been prepared. It is cleaned thoroughly inside by rubbing with 
 emery paper and oil, or with pumice stone, and wiping with 
 an oily rag ; it may also be blackleaded inside, as this prevents 
 contact between the gold and the iron of the mould. It is then 
 warmed by being placed 011 the top of the furnace ; its temper- 
 ature must not be sufficiently high to ignite the oil, but it should 
 
392 THE METALLURGY OF GOLD. 
 
 be too hot to touch with the hand. When the bullion is ready 
 to pour, the mould is placed on a level surface, such as an iron 
 stool, at a height above the floor of about 12 or 18 inches and oil 
 -poured into it. Any cheap non-volatile oil will do, whether 
 animal or vegetable. The mould is tilled to a depth of about a 
 quarter of an inch with oil, which is made to flow over all parts 
 of the interior. 
 
 The crucible is then lifted from the furnace, usually with basket 
 tongs, and when the temperature of the metal has fallen to but 
 little above its melting point, it is poured rapidly but steadily 
 into the mould, the crucible being moved to and fro so that the 
 stream of molten metal is directed to all parts of the mould in 
 succession. The crucible is then held in the inverted position 
 for a short time, and jarred once or twice to cause the last portion 
 of the metal to flow from it. The oil is ignited, and burns on the 
 top of the cast metal, thus keeping it from tarnishing. In small 
 castings, the slag is allowed to flow out and remain on the top of 
 the metal in the mould ; in large castings, the slag is usually 
 skimmed off before pouring. Beads of metal are caught in a 
 large iron tray with raised edges, in the centre of which the 
 mould is placed. If the mould is clean and has been hot enough, 
 and if enough oil has been used, a clean untarnished bar is pro- 
 duced. It is turned out of the mould by inversion of the latter, 
 while still too hot to be handled, and the slag is separated by one 
 or two light taps with a hammer. The bar is then, in many 
 establishments, momentarily dipped into water to assist in the 
 complete removal of the last fragments of the slag, and it is also 
 a favourite practice to dip the bar, first into dilute sulphuric acid, 
 and then into clean water, the bar retaining warmth enough 
 after removal to expel all moisture. This treatment removes all 
 tarnish, and any adherent particles of slag are then chipped off 
 and assay pieces cut from the bar. 
 
 In some refineries large crucibles are used, and 3,000 or 4,000 
 ozs. of metal are refined at once. In this case, several ingot 
 moulds are filled successively from one pot, the weight of the 
 gold bars manufactured being usually either 200 or 400 ozs. each. 
 Silver bars, on the other hand, are made much larger, often 
 weighing 1,000 or 1,200 ozs. Mixed bars of bullion (containing 
 both gold and silver) are seldom cast of a greater weight than 600 
 ozs. in mills. 
 
 The method of refining which has been described above is 
 seldom attempted in extraction mills, and is often unnecessary 
 in refineries, before parting the silver from the gold. The object 
 at a mill is merely to melt down the bullion obtained, so as to 
 bring it to a marketable form with as little loss as possible. 
 With this object in view, no nitre is used, but the metal is kept 
 covered by a layer of charcoal to prevent the formation of oxides, 
 and the crucible is poured as soon as the charge has been fused 
 and stirred. 
 
REFINING. 393 
 
 In conducting all these furnace operations the use of a thick 
 pair of mittens, made of sacking or rubber, is to be recommended 
 for the protection of the hands. 
 
 Losses of Bullion Incurred in Melting. The losses sus- 
 tained in melting vary according to the composition of the 
 metal, but are usually very small. At the New York Assay 
 Office they are found "to vary from 0-5 to 1-5 per 1,000. The 
 losses may be divided into mechanical loss, and loss by volatili- 
 sation. The mechanical loss is reduced by care in the conduct 
 of the operation ; it may be due to a number of causes. The 
 crucible may break and its contents fall into the fire, or be 
 scattered over the floor of the melting house when on the point 
 of being poured. To avoid loss in this way, the ashpit may 
 be constructed of a cast-iron tray, which can be easily scraped out. 
 The floor of the room is made of carefully -laid flagstones, or, 
 better still, of iron plates, in which there are no cracks or 
 crannies capable of hiding metal beads. Projection by spirting out 
 of the crucible may be occasioned if certain impurities, such as 
 tellurium or antimony, are present in the bullion, and if the 
 nitre is added in large quantities at a time, the pot may "boil 
 over," and part be lost. When flux is to be added, the surface of 
 the metal must be at least 6 inches below the rim of the pot. 
 Recovery of any metal lost in the ashes is effected by panning. 
 
 The slags formed in the course of refining frequently contain 
 some small shots of metal, which may be recovered by grinding 
 the slag finely and washing down the product in a pan or on a 
 vanning shovel. At the San Francisco Mint, the residue from the 
 vannings are allowed to accumulate, and at intervals dried and 
 fused with borax in an old graphite crucible at a high tempera- 
 ture. The crucible is left in the furnace over night to cool, and 
 is then broken up, and the shots of metal at the bottom picked 
 out. They are chiefly silver, very little gold being thus 
 recovered. Shots of metal can be recovered with greater 
 certainty from the slags by fusion with lead. 
 
 The crucibles, stirrers, lids, &c., also contain a certain quantity 
 of gold and silver. After each melting they are scraped, and the 
 scrapings panned, or, better still, calcined and fused with lead; 
 but, in spite of this treatment, precious metals accumulate in 
 the pots, and, when worn out, they are ground-up in a Chilian 
 mill or other grinder, and panned, in order to separate the shots 
 of metal. The tailings from this treatment will often pay for 
 fusion with lead, and subsequent cupellation. Certain refineries 
 treat large quantities of such residues, with which may be 
 included sweepings of the floor of the melting house. 
 
 At the Philadelphia Mint it is found worth while to recover 
 the gold from the iron tools used in stirring, dipping, &c. For 
 this purpose they are melted down in a graphite crucible with a 
 little charcoal to make grey-iron, and kept at a white heat for 
 
394 THE METALLURGY OF GOLD. 
 
 some time, after which the charge is allowed to cool slowly. 
 Under this treatment the gold and silver separate out (an alloy 
 containing three or four parts of silver to one part of gold being^ 
 better for the purpose than pure gold), and are found at the 
 bottom of the crucible sharply marked off from the surface of the 
 iron, which is now quite freed from the precious metals.* 
 
 The mechanical loss in melting at the San Francisco mint, 
 where crude bullion is bought and valued, is stated to be only 
 about 0-005 oz. in a melt of 100 ozs. This is only one part in 
 twenty thousand, but the total loss of gold is usually much 
 greater, and, as the crude bullion cannot be accurately valued 
 before it is melted-up, it is always difficult to determine how 
 much loss has actually occurred. 
 
 The loss by volatilisation is often more serious than the 
 mechanical loss. It is increased by the passage of a rapid 
 current of air over the surface of the molten metal; to avoid 
 this, the fire-box is made deep enough to allow the top of the 
 pot to be sunk some distance below the flue, and the exposure 
 to the draught coming from the opening in the top of the furnace 
 is thus diminished. The metal is also sheltered by being kept 
 covered by the crucible lid as much as possible, and covers of 
 charcoal powder, of fragments of charcoal, or of bone-ash are in 
 general use. For the effect of different alloying metals on 
 the volatilisation of gold, see p. 8. Napier found f that much 
 volatilisation took place during pouring, so that if a wet beaker is 
 held inverted over the mould during pouring, gold is condensed in 
 it and can be subsequently detected by the ordinary tests. As for 
 other conditions, it appears from the results on p. 6 that the 
 longer the gold is kept melted, the higher the temperature em- 
 ployed, and the smaller the mass of gold, the greater will be the 
 percentage loss by volatilisation. These results are in accordance 
 with the experience of those engaged in the operations on a large 
 scale. 
 
 The loss by volatilisation when the gold is toughened by the 
 use of corrosive sublimate is particularly high ; the total loss of 
 gold incurred when bars are subjected to this treatment is 
 stated to be on an average about 0-85 per 1,000 or .850 per 
 million sterling, while the ordinary loss on melting, both 
 mechanical and by volatilisation, only amounts to about 0*15 or 
 0*17 per 1,000, part of which is recoverable. 
 
 In order to condense the gold and silver which are carried off 
 by the volatilised copper, Mr. J. Feix, foreman of the refining 
 department of the San Francisco Mint, devised an ingenious 
 flue-dust chamber, which is used there as an adjunct to the 
 melting furnaces, j The chambers are built of brick, and the 
 
 *Egleston's Metallurgy of Silver, Gold, and Mercury, vol. ii., p. 728. 
 
 ^Chem. 8oc. Journ., vol. x., p. 229. 
 
 j Ninth fieport Col. State Min., 1889, p. 66. 
 
REFINING. 395 
 
 horizontal flues, coming from the tire-box, open almost directly 
 into them. The chambers are 12 feet high, 1 foot wide, and of con- 
 siderable length ; the flues from them are placed within 1 foot 
 of the top and communicate with the stack. Iron doors, placed 
 at the end of the chamber, enable the sweepings and dust to be 
 readily withdrawn. It is stated that almost all the precious- 
 metals which have been volatilised are condensed in this way. 
 The sweepings are heated with borax to a high temperature in 
 an old graphite crucible, when the metal accumulates at the 
 bottom as a regulus. It is not stated how much gold and silver 
 is recovered in this way, and whether the amount has paid for 
 the erection of the chambers and the extra labour and fuel 
 required. 
 
 The loss of weight in melting will of course be frequently 
 much greater than the actual loss of gold. Thus mercury, if 
 present, will be expelled by volatilisation, and this metal usually 
 forms from O5 to 1 part per 1,000 of retorted gold. Zinc is not 
 so readily expelled, but, if any is present, some will be volatilised, 
 the amount depending on the temperature and duration of the 
 melting. Moreover, the earthy material and all the base metals 
 which have been oxidised by the nitre will be slagged off by 
 the borax, and the total diminution in weight may thus amount 
 to several per cent. 
 
 Refining by Means of Sulphur is said to be practised in 
 the United States Mints ;* the following is a brief account : It 
 is effected in plumbago crucibles, and has for its main object the 
 elimination from retorted metal of iron, when, as sometimes 
 happens, it is present in large quantities. The metal is kept just 
 above its melting point, the temperature being as low as possible 
 in order to avoid unnecessary waste of sulphur by volatilisation. 
 Sulphur is sprinkled round the edges of the molten mass, and 
 stirred in with a graphite stirrer. If sulphur is added near the- 
 centre, particles of gold are lost by projection. Sulphide of iron 
 is formed with great energy, and sulphide of silver also, but the 
 latter is not produced rapidly until nearly all the iron has been 
 already converted into sulphide. The gold is unaffected by the 
 sulphur and subsides to the bottom. It is not usually cast by 
 pouring, but allowed to solidify in the pot, a better separation 
 between the gold and the matte being thus effected. The pot is 
 turned out as soon as solidification has taken place, and the 
 matte is broken off by a hammer, the gold being remelted and 
 cast into a bar. The small quantity of gold taken up by the 
 matte is separated by melting with metallic iron. 
 
 Osmiridium in Gold Bars. Gold from some districts of 
 
 California and from the Fraser River district of British Columbia 
 
 contains some platinum and palladium, and frequently some 
 
 osmium and iridium. As these metals remain together during, 
 
 * Ninth Report Cal. State Min., 1889, p. 64. 
 
396 THE METALLURGY OF GOLD. 
 
 the treatment, the mixture is commonly called osmiriclium. If 
 this is present in perceptible quantities, a fact \vhich is not 
 usually detected until after the bars have been parted, the gold 
 is remelted in a clean crucible and kept fused at a high tempera- 
 ture for about half an hour, when the osmiridium will settle 
 to the bottom of the crucible. This is due to the fact that 
 osmiridium does not seem to form a true alloy with gold, and, 
 being of high density and very infusible, the particles, unfused 
 or partially fused, settle through the liquid gold. The crucible 
 is then gently lifted out of the furnace, and the greater portion 
 carefully but rapidly poured into a mould ; the remainder, which 
 contains almost all the osmiridium, is allowed to cool in the 
 crucible until it has solidified, and then assayed for osmiridium. 
 An alternative plan is to allow the whole charge to solidify in 
 the crucible, and then to cut off the lowest portion, which is set 
 aside. The osmiridium settles better from an alloy chiefly 
 consisting of silver than from pure gold, and the rich bottoms 
 are consequently melted several times with silver, the lowest 
 part being cut off each time. The gold is thus gradually 
 replaced by silver, which eventually forms by far the greater 
 part of the mass. It is then granulated and parted, and the 
 resulting powder of gold and osmiridium is treated with aqua 
 xegia, by which the gold is dissolved, and the osmiridium 
 separated as a black powder. Such osmiridium is worth from 
 eight to twenty shillings per ounce, and selected grains are used to 
 make the "diamond points" of gold pens. Iridium is, however, 
 not invariably separated from the gold bars in which it is 
 contained, and traces can be observed in a considerable pro- 
 portion of the refined commercial bars met with in London. 
 Platinum and palladium are, in great part, extracted by the 
 nitre, and enter the slag. 
 
 PARTING. 
 
 Parting is the separation of silver from gold. During the 
 -course of the operation the base metals are separated from both, 
 but, as the presence of these base metals is injurious to the 
 successful conduct of the processes which are chiefly in use, a 
 preliminary refining by one of the methods already described is 
 usmally necessary. Only about 10 per cent, of base metals is 
 permissible in the alloys when sulphuric acid is used, although 
 a somewhat larger quantity does no harm if nitric acid is em- 
 ployed. 
 
 The processes of parting may be tabulated as follows : 
 
 1. Cementation. 
 
 2. Melting with sulphide of antimony. 
 
 3. Melting with sulphur, and precipitation of the gold from 
 the regulus by silver, iron, or litharge. 
 
PARTING. 39T 
 
 4. Parting by nitric acid. 
 
 5. Parting by sulphuric acid. 
 
 6. A combination of these last two methods. 
 
 7. The Gutzkow process. 
 
 8. The new Gutzkow process. 
 
 9. Parting by chlorine gas. 
 10. Parting by electrolysis. 
 
 The first two of these methods were Known to the ancients, 
 and the third was described as early as the beginning of the- 
 llth century. 
 
 Cementation. In this ancient and obsolete process, gold was 
 freed from small quantities of silver, copper, &c., contained in it. 
 The method was mentioned by Pliny and described by Geber, 
 who wrote in Arabic, probably in the eighth or ninth century ;. 
 it is possibly still in use in Japan and in some parts of South 
 America. It consists in heating granulations of argentiferous 
 gold mixed with a cement, consisting of two parts of brick-dust, 
 or some similar material, and one of common salt, in pots of 
 porous earthenware. The temperature used is a cherry-red heat,, 
 which is insufficient to melt the granulations. After about 
 thirty-six hours' treatment, the greater part of the silver is- 
 converted into the state of chloride, and this, together with the 
 cement, can be removed from association with the granulations 
 by washing with water. The gold can in this way be raised to 
 a fineness of about 850 or 900. The silver is recovered from the 
 cement by amalgamation with mercury.* 
 
 Parting by Means of Sulphide of Antimony. This pro- 
 cess was also used to purify gold which contained only small 
 quantities of silver. The alloy was repeatedly melted with 
 sulphide of antimony, upon which the gold became alloyed with 
 the antimony and sank to the bottom of the mass, while the 
 silver was converted into sulphide and floated on the top, mixed 
 with the excess of antimony sulphide added. The gold wa& 
 subsequently refined by a blast of air directed upon it, the 
 antimony being thus oxidised and volatilised. The method is 
 now obsolete, bub was in use at the Dresden Mint up to the 
 year 1846, and gold of the fineness 993 was said to be produced 
 in this way. 
 
 Parting by Means of Sulphur. This method was formerly 
 used for the purpose of concentrating the gold contained in 
 auriferous silver in order to obtain a richer alloy. The granu- 
 lated alloy was melted with sulphur and some of the silver 
 was thus converted into a matte. The gold was then precipitated 
 from the matte and collected in a smaller quantity of silver by 
 fusion with pure silver, or with iron, or litharge. No attempt 
 
 * For a full account of this interesting process, as well as of the next 
 succeeding two methods, the student is referred to Percy's Metallurgy of 
 Silver and Gold, pp. 356-402. 
 
"398 THE METALLURGY OP GOLD. 
 
 was made to obtain pure gold in this way, and the enriched 
 alloy of gold and silver was parted by nitric acid. The silver 
 was recovered from the matte by fusion with iron. The method 
 was in use in several refineries in Europe at the beginning of the 
 present century. The employment of sulphur in refining at the 
 United States Mints has been already noticed, p. 395. 
 
 Parting by Nitric Acid. The first clear mention of the use 
 of nitric acid for parting silver from gold is made by Albertus 
 Magnus, who wrote in the thirteenth century, but the process 
 does not appear to have been employed on a large scale until two 
 .centuries later in Venice. Here, according to an old tradition,* 
 some Germans were employed in separating gold from Spanish 
 silver in the fifteenth and sixteenth centuries, the art being kept 
 secret. These refiners were not inaptly named " gold makers " 
 by those who were unacquainted with their methods. The pro- 
 cess was fully described by Biringuccio in his treatise,! published 
 in 1540, and by Agricola J in 1556. It was first used in the 
 Paris Mint about the year 1514, and in London at least as early 
 as 1594, but for a long period the operations were conducted in 
 secret in both countries, and it is supposed that this method of 
 refining was not fully practised in England until about the 
 middle of the eighteenth century. 
 
 Parting by means of nitric acid is conducted on the large scale 
 in the same general manner as in the assaying of gold bullion. 
 It consists of the following operations : 
 
 1. Granulation of the alloys. 
 
 2. Solution of the silver in nitric acid. 
 
 3. Treatment of the gold residues, viz. : Sweetening by washing with 
 water, drying, melting, and casting into bars. 
 
 4. Precipitation of the silver as chloride by salt solution. 
 
 5. Reduction of the silver chloride by zinc and sulphuric acid. 
 
 Granulation of the Alloys. The gold to be parted must be 
 approximately free from base metals, particularly from those 
 which are not soluble in nitric acid, such as tin, arsenic, anti- 
 mony, &c. If these were present they would form insoluble 
 oxides, which would remain with the gold, so that further refining 
 operations would be necessary: they would, moreover, cause a 
 great increase in the consumption of nitric acid, so, if they are 
 present, the gold is freed from them as far as possible by melting 
 with nitre, &c. Copper, lead, and other metals which are 
 readily soluble in nitric acid are less obnoxious, and small 
 percentages of these are allowed to remain, as they are not 
 difficult to separate from the silver when in solution with it, 
 while the presence of copper in particular is advantageous in 
 promoting rapid dissolution of the alloy. If present in large 
 
 *Beckmann's History of Inventions, vol. iv., p. 578. 
 t De la Pirotechnia, Florence, 1540. 
 j De re Metallica. 
 
PARTING. 399 
 
 quantities, however, even these metals would create difficulties 
 and expense, increasing the consumption of acid. 
 
 The bars are melted together to form an alloy which, it was 
 formerly believed, must contain one part of gold to three parts 
 of silver (hence the term " inquartation " applied to this process).* 
 In most refineries, however, the proportion of silver used is less, 
 the minimum being two parts to one of gold. At the Phila- 
 delphia Mint, the proportion is now 2J to l.f Dore" bars con- 
 gaining small quantities of gold are, of course, preferred to bars 
 of tine silver for the purpose of alloying with the argentiferous 
 gold bars. After the "inquarted" alloy has been thoroughly 
 mixed by being stirred while still in the furnace, the crucible 
 containing it is lifted out, and the metal is poured into copper 
 tanks filled with cold water, which is sometimes kept cool by 
 ice, while, in some refineries, a stream of water is kept constantly 
 flowing through the tank. The metal is poured with a circular 
 and wavy motion in a thin stream to prevent the formation of 
 lumps ; leafy granules and small hollow spheres are thus formed. 
 The pouring is done either directly from a crucible or from a 
 dipper, the vessel being held in either case about 3 feet above 
 the surface of the water. In the tank is a perforated copper 
 pan, which is lifted out when the pouring is completed, and the 
 granulations allowed to drain. 
 
 Dissolving the Granulations. The granulated metal is heated 
 with nitric acid in cylindrical vessels of earthenware, porcelain, 
 or platinum. The porcelain vessels in use at the Philadelphia 
 Mint have a capacity of about 33 gallons, and contain 2080 ozs. 
 of granulations : twelve such vessels are placed together on a 
 lattice work of wood, which is laid on the bottom of lead tanks, 
 and are surrounded by water 10 inches deep kept at the boiling 
 temperature by means of steam. The earthenware vessels are 
 covered by a closely-fitting lid provided with a delivery tube to 
 carry away the fumes. Messrs. Johnson & Matthey's platinum 
 vessels, which hold 800 ozs. of metal, are heated by separate 
 furnaces, the fuel being coke or coal gas. 
 
 The strength of the acid used varies from that of spec. gr. 141 
 to that of spec. gr. 1*2. When nitric acid alone is used, about 
 3 Ibs. of acid (spec. gr. 1'2) are used to dissolve each pound 
 of granulations, but of this quantity the amount used in the 
 last boiling (about 20 per cent, of the whole) is available for 
 
 *A smaller proportion of silver, however, was used at least as early 
 as the year 1627 in Paris. Thus Savot observes in his Discours sur ies 
 Medalles Antiques, Paris, 1627, chap, vii., p. 72: "S'il n'y a beaucoup 
 plus d'argent que d'or, 1'eau n'agira aucunement : de sorte qu'il faut qu'il 
 y ayt au moins Ies deux tiers d'argent, et un autre tiers d'or, et encore que 
 1'eau soit tres-bonne : car, si elle est foible, elle n'operara point. " Savot 
 did not seem to regard this proportion of 1 to 2 as of recent introduction. 
 
 ^Report of the Director of the Mint for 1896, p. 135. 
 
400 THE METALLURGY OF GOLD. 
 
 further use ; more acid is required if there is much copper 
 present. The first addition of acid is of about 1J Ibs. to each 
 pound of metal ; it is kept boiling gently for about five or six 
 hours, by which time most of the silver will have been dissolved. 
 The solution is allowed time to settle, and the hot supernatant 
 liquid is siphoned off by a gold or glass siphon, and diluted with 
 water to prevent the formation of crystals on cooling. The 
 second addition consists, in some establishments, of strong acid, 
 (specific gravity 1-414), and in others of acid of the same strength 
 as before. The second boiling is for two or three hours only, 
 and the third boiling for only one or two hours, the liquid being 
 siphoned off after each boiling. 
 
 The vessels are provided with hoods and small chambers in 
 the delivery tubes, in order to effect a partial condensation of 
 the acid, and also to recover the small amount of silver nitrate 
 which is carried over mechanically, owing to the violence of the 
 disengagement of gas bubbles. The fumes are conducted to the 
 melting furnace where they are consumed, giving up their 
 oxygen to the fuel. 
 
 The reactions that occur are partially expressed by the follow- 
 ing equations : 
 
 6Ag + SHN0 3 = 6AgN0 3 + 2NO + 4H 2 
 3Cu + SHN0 3 = 3Cu(NO s )a + 2NO + 4H 2 
 4Zn + 10HNO 3 = 4Zn(N0 8 ) 2 + N 2 + 5H 2 
 
 and similar reactions for other metals. The amount of nitrous 
 oxide evolved increases towards the end of the operation. It is 
 seen that silver decomposes less than its own weight of nitric 
 acid, while copper and zinc destroy nearly three times their 
 weight of the acid. Nitric acid of specific gravity 1'2 contains 
 about 32 per cent, of anhydrous HNO 3 , so that the quantity of 
 acid of this strength theoretically required to dissolve 1 Ib. of 
 silver, copper and zinc is about 2-4 Ibs., 8 -3 Ibs. and 7 -6 Ibs. 
 respectively. 
 
 Treatment of the Gold Residue. The pulverulent gold is 
 " sweetened " by being washed thoroughly in perforated earthen- 
 ware dishes with boiling distilled water, stirring being performed 
 with a spatula of wood, platinum, or porcelain. The gold is thus 
 freed from nitric acid and nitrate of silver, the operation being 
 continued until the washings show no signs of turbidity on the 
 addition of salt. The washings are added to the first silver solu- 
 tions, serving to dilute them, the dilution, as has already been 
 observed, being necessary to prevent crystallisation on cooling. 
 The sweetened gold is generally pressed, dried, melted, and cast 
 into bars, which are now made of a weight of either 200 or 
 400 ozs. The gold thus obtained is usually of a fineness of about 
 997 or 998, the remainder being chiefly silver, which would not 
 pay for extraction, although part of it could be separated with 
 
PARTING. 
 
 a further expenditure of time, fuel, and acid. The gold is 
 pressed by a hydraulic ram, the pressure exerted being about 
 800 Ibs. to the square inch. The cakes of metal are dried at a 
 cherry-red heat, and then broken up for melting. 
 
 Treatment of the Silver Solution. The solution of nitrate of 
 silver is diluted with water, allowed to cool, and then treated 
 with a strong solution of salt which is regulated so as not to be 
 in excess, a constant agitation being kept up by revolving wooden 
 agitators driven by steam power, or by hand paddles. When all 
 the silver has been precipitated as chloride, the whole is allowed 
 to settle overnight, and, in the morning, the clear solution of 
 nitrate of soda, containing most of the base metals originally 
 present in the alloys, is drawn off and filtered. The precipitated 
 chloride is washed several times by decantation and agitation, 
 and finally sweetened in wooden filters by boiling water, which 
 incidentally dissolves out the chloride of lead. The filters are 
 usually lined with linen or some similar material. 
 
 Reduction of the Silver Chloride. The silver chloride is then 
 reduced in lead-lined tanks by means of granulated zinc and 
 water acidulated with sulphuric acid. Thirty-three pounds of 
 commercial granulated zinc are stated to be enough to reduce 100 
 Ibs. of silver from the chloride.* The reactions involved are as 
 follows : 
 
 (1) 2AgCl + Zn = ZnCl 2 + 2Ag 
 
 (2) Zn + H 2 S0 4 = ZnS0 4 + H 2 
 
 (3) H 2 + 2AgCl = 2Ag + 2HC1 
 
 (4) 2HC1 + Zn = ZnCl 2 + H 2 
 
 These reactions explain the fact that, while zinc slowly reduces 
 silver chloride in the presence of water only, the action is quick- 
 ened by the addition of free acid, by which the zinc is attacked 
 and hydrogen evolved. Nascent hydrogen is a powerful reducing 
 agent, and decomposes silver chloride much more rapidly than 
 zinc does, hydrochloric acid being formed and rendered available 
 for the production of more hydrogen. The result of this is that 
 the action, which is at first slow, becomes more and more rapid 
 as hydrochloric acid accumulates in the solution. The sulphuric 
 acid is only needed to start the reaction, but, of course, the more 
 that is added, the more quickly will the operation proceed, and if 
 much is added, the chemical action is more violent at first than 
 afterwards, the amount of free acid present in this case falling 
 off. At the ISan Francisco Mint 1 Ib. of acid of 60B. is added 
 for every 2 Ibs. of silver to be reduced. Hydrogen is evolved 
 copiously and is carried off by a hood and flue. Energetic 
 stirring with wooden paddles is desirable to prevent the forma- 
 tion of lumps of chloride, protected by a layer of silver powder. 
 
 * Ninth Report Gal. State Min., 1889, p. 70. 
 
 26 
 
402 THE METALLURGY OP GOLD. 
 
 The white chloride of silver gradually turns black as the silver is 
 reduced. To test whether the reaction is complete, some of the 
 silver is taken out, washed well with ammonia, filtered, and the 
 clear solution acidified with nitric acid. A white precipitate 
 signifies that undecomposed silver chloride is still present in the 
 vat. 
 
 When the reduction is complete, the vats are allowed to settle, 
 the solution drawn off, and a little sulphuric acid added to dis- 
 solve any residue of zinc that may be present. The dark grey 
 pulverulent silver is then washed by decantation, after which it 
 is removed to a wooden filter, and sweetened by washing with 
 boiling water, and finally pressed, dried, and melted into bars, 
 which are about 998 fine. The zinc and sulphuric acid used in 
 this process are lost, and a considerable quantity of undecom- 
 posed nitric acid is also run to waste, being contained in the 
 solution from which the silver chloride is precipitated. 
 
 The cost of refining and parting by the nitric acid process at 
 the United States Mints in Philadelphia and New York is some- 
 what less than 2 cents per oz. of the parting alloy, and in San 
 Francisco it is nearly 3 cents. The cost for dore silver is 
 considerably lower. In Europe, the cost is less than in the 
 United States. 
 
 Parting by Sulphuric Acid. This process has now, in the 
 majority of refineries, superseded the nitric acid method, which 
 is much more expensive, owing to the higher cost of the acid 
 used and of the plant required. The German chemist, Kunckel, 
 who lived in the seventeenth century, is said to have been the 
 first to employ sulphuric acid in parting, but it was not used on 
 the large scale until the year 1802, when it was introduced into 
 France by 0. D'Arcet, and worked in a refinery built in Paris 
 for the purpose. It was established in London at the Mint 
 Refinery in 1829 by Mr. Mathison, and has been in almost 
 continuous use there ever since, with little change, having been 
 leased to a member of the Rothschild family since 1852. 
 
 The method used varies considerably in different refineries, 
 but essentially consists of the following operations 
 
 1. Mixing and granulating the alloys. 
 
 2. Dissolving the silver from the granulations by sulphuric acid. 
 
 3. Washing and melting the gold residue. 
 
 4. Precipitating the silver from its solution by means of copper. 
 
 5. Recovering the copper sulphate by crystallisation. 
 
 The account given below is a general view of the operations in 
 various refineries, the modifications adopted not being described 
 in most cases. 
 
 Mixing and Granulating. The alloys must be carefully pre- 
 pared so as to be of suitable composition, as otherwise difficulties 
 are encountered. The most suitable proportion of gold in the 
 
PARTING. 403 
 
 alloy is said by Dr. Percy* to be from 18 to 25 per cent., 
 including whatever copper there may be present ; but some 
 American authorities consider the proportion of one part of gold 
 to two and a-half parts of silver to be the most desirable, whilst 
 at a refinery at San Francisco the alloy consists of two parts of 
 gold to three parts of silver. This proportion was instituted 
 when alloying silver was scarce in California and has never been 
 abandoned, but the gold thus separated is only 990 fine, contain- 
 ing ten parts of silver, the maximum allowed by law in the 
 gold coins of the United States. If the ordinary proportion of 
 three parts of silver to one of gold is used, however, the gold can 
 be obtained about 996 fine, and the fineness of the gold can 
 always be increased to about 998 or 999 by fusing it, first with 
 bisulphate of potash and subsequently with nitre. If there are 
 only a few parts of gold per 1,000 of the alloy, it has been stated 
 that the silver left in the gold amounts to as much as 3 or 4 per 
 cent. ; nevertheless, such an alloy, when subjected in the form 
 of bars to the action of the acid, instead of being granulated, 
 yields gold at San Francisco of no less than 996 fine, after one 
 boiling only. 
 
 The amount of base metals present in the alloy must be care- 
 fully regulated, as their sulphates are little soluble in concentrated 
 sulphuric acid, and consequently are precipitated and interfere 
 with the progress of the operation. Bars consisting in great 
 part of copper are often received at the San Francisco works. 
 These are melted with fine bars so as to reduce the proportion, 
 of copper, which must not be more than about 10 or 12 per cent, 
 of the whole \ a small amount of copper facilitates the solution 
 of the silver. A small quantity of lead is said to assist in the 
 solution of the copper, which is somewhat slowly attacked by 
 concentrated sulphuric acid, and a maximum amount of 5 per 
 cent, of lead does not interfere with the operation. From the 
 economy with which this system of parting can be practised, 
 silver containing only 0'5 part of gold per 1,000 can be separated 
 from it at a profit, while the nitric acid process is unremunerative 
 if applied to an auriferous silver alloy containing one part of 
 gold in a thousand. At the Vienna Mint, bars are parted con- 
 taining 0-9 part of gold per 1,000, and at Freiberg bars containing 
 only 0-4 part per 1,000 are profitably treated. 
 
 In England, silver bars are passed through the parting opera- 
 tion, if they contain at least 2 grains of gold per troy pound, or 
 0-35 part per 1,000. 
 
 The parting alloy is usually granulated, but at the San Fran- 
 cisco Refinery the dore silver is not granulated but melted and 
 cast into bars f inch thick, 9 inches wide, and 15 inches long. 
 
 Solution of the Silver. This is usually effected in cast-iron 
 kettles, platinum having been abandoned on account of its 
 * Metallurgy of Silver and Gold, p. 471. 
 
404 THE METALLURGY OF GOLD. 
 
 high cost. The iron used is fine-grained compact white iron, 
 preferably containing 3 or 4 per cent, of phosphorus, which 
 increases the durability, although 2 per cent, only of phosphorus 
 is considered enough by some refiners. The kettle is slowly 
 dissolved by the acid, ferrous sulphate being formed, and, in the 
 course of about two years, the thickness of the vessel is reduced 
 from about 2 inches to from J to \ inch, when it is discarded. 
 The perfect exclusion of air from the interior increases the 
 length of life, and dilute acid must not be allowed to come in 
 contact with the iron, as the latter is freely dissolved by it. 
 The vessels are rectangular or cylindrical, with flat or hemi- 
 spherical bottoms, the latter being preferred in Europe and the 
 former in America. They are covered with cast-iron lids, about 
 \ inch in thickness, which are bolted tightly to the vessels, and 
 have bent leaden pipes fitted to them for carrying off the fumes,, 
 which consist largely of SO 2 . This is sometimes reconverted 
 into sulphuric acid in leaden chambers arranged for the purpose. 
 The cover has also an opening (supplied with a lid made air-tight 
 by a water-joint) through which the alloys and acids are added 
 and the operation watched. The heating is usually done by a. 
 wood fire. (See Fig. 57a). 
 
 The charge for the pots varies from 200 to 1,000 Ibs. of alloy,, 
 and the amount of acid required varies from 2 to 2J times the 
 weight of the alloy, depending on the composition of the latter. 
 About one-half of the acid, which is strong commercial acid of 
 66 Beaume" (specific gravity 1'85) is added at first, and the 
 temperature cautiously raised to boiling point, when the pot is- 
 closely watched, and, if the ebullition becomes too violent, the 
 temperature is lowered by regulating the fire and by adding cold 
 acid a little at a time. The charge is stirred occasionally with 
 an iron tool, particularly towards the end of the operation, when 
 the undissolved granules of metal must be freed from the 
 surrounding sediment, consisting of sulphates of the base metals, 
 and exposed to the action of the acid. The ebullition gradually 
 subsides and action ceases in about five or six hours, the 
 presence of a greater proportion of base metals increasing the 
 length of time required. The reactions are as follows : 
 
 (1) 2H 2 S0 4 + Ag 2 = Ag 2 S0 4 + S0 2 + 2H 2 
 - (2) 2H 2 S0 4 + Cu = CuS0 4 + S0 2 + 2H 2 
 
 and similar reactions with tin and lead. The re-actions with 
 antimony, bismuth, zinc, and iron are more complicated. It is 
 obvious that 63 parts of copper decompose as much sulphuric 
 acid as 216 parts of silver. It is clear, therefore, that an increase 
 in the percentage of copper present necessitates an increase in 
 the amount of sulpKuric acid required. 
 
 One part of sulphate of silver is soluble in J part of boiling 
 concentrated sulphuric acid, but the solubility rapidly falls off as 
 
PARTING. 
 
 405 
 
406 THE METALLURGY OF GOLD. 
 
 the temperature and concentration diminish, so that 180 parts of 
 cold acid of specific gravity 1-08 are required for the same 
 purpose. Sulphate of copper dissolves slightly in the boiling 
 concentrated acid, but is almost all precipitated in the form of 
 the white anhydrous salt on cooling. Tin and zinc behave 
 similarly, and lead makes the solution turbid and milky. The 
 iron would not be so much attacked if it were not for the 
 increasing dilution of the acid during the process, owing to the 
 formation of water, which is, however, in great part boiled off as 
 fast as it forms, or taken up by the anhydrous sulphates. The 
 presence of copper checks the dissolution of the iron. 
 
 When the dissolution is complete, the fire is withdrawn and a 
 few pounds of cold acid of 55 B. are added to the charge, by 
 which the acid is cooled and diluted, and some crystals of silver 
 sulphate are formed. These, falling to the bottom, carry down 
 with them the suspended fine particles of gold, and so clarify the 
 solution. If much copper is present, however, this is not 
 necessary, as the slight cooling of the acid, caused by the 
 withdrawal of the fire, is enough to precipitate some sulphate of 
 copper, which falls to the bottom and adheres to it very firmly, 
 thus clarifying the liquid and enabling it to be poured off or 
 ladled out very closely. The clear silver solution is then ladled 
 out with iron ladles into lead-lined rectangular wooden vats 
 already partly filled with hot water, in which the precipitation 
 is subsequently effected. 
 
 Washing and Melti&g the Gold Residue. The residue in the 
 dissolving pot, if the amount of base metals present is not large, 
 is then boiled twice more with fresh concentrated sulphuric acid 
 added hot, after which the gold residue is hard and heavy and 
 rapidly subsides to the bottom, and the liquors are ladled into- 
 the precipitating vat. The gold is dipped out with an iron- 
 strainer and transferred to a lead-lined filter-box where it is 
 thoroughly washed, first with hot dilute sulphuric acid and 
 subsequently with boiling water, after which it is pressed, dried, 
 and melted. It is almost always brittle, from the occurrence in 
 it of traces of lead or tin which are difficult to separate by sul- 
 phuric acid owing to their insolubility. These metals are 
 eliminated by fusion with nitre or by a blast of air, and the bars 
 thus toughened. 
 
 If the amount of base metals present is very large, the gold 
 residues are ladled into a vessel of hot dilute sulphuric acid and 
 boiled with it by means of steam. In this way, most of the 
 sulphate of silver and the whole of the copper, zinc, iron, &c., 
 remaining with the gold are rapidly dissolved. Care must be 
 taken, however, to add the residues a little at a time, as other- 
 wise the anhydrous sulphate of copper will form lumps, which 
 are only slowly dissolved. The gold is then allowed to settle, 
 and, after the solution has been drawn off, is boiled again with 
 
PARTING. 407 
 
 acid if necessary, or if it is already pure enough it is at once 
 washed, dried, and melted. 
 
 In Europe it is not customary to attempt to obtain pure 
 gold from auriferous silver in one operation, but the gold is con- 
 centrated in a small quantity of silver and then mixed with other 
 alloys rich in gold and parted again. The product of gold thus 
 obtained is purified by heating in a furnace in small iron pots 
 with about half its weight of bisulphate of potash, by which some 
 additional silver is converted into sulphate. The temperature is 
 not raised much above the fusion point of the salt. The fused 
 mass is then boiled in sulphuric acid, and again washed, dried, 
 and melted. In the United States these methods are not used, 
 auriferous silver being cast into slabs and parted merely bv 
 boiling with sulphuric acid; fusion with bisulphate of potash 
 is rarely resorted to, but at the United States Mints, the residues 
 are boiled with fresh acid five times in succession. 
 
 Precipitation of the Silver. On pouring the sulphuric acid 
 solution into water, most of the silver sulphate is precipitated at 
 once in the form of small crystals, consisting of bisulphate, and 
 the liquid must then be raised to boiling, by means of steam, in 
 order to redissolve them. When the original alloys contain 
 much lead this is not redissolved, and it is, therefore, necessary 
 to let the solution settle and transfer the clear liquid to another 
 vessel. Some particles of gold are usually found in the pre- 
 cipitate thus formed. 
 
 The reduction and precipitation of the silver is effected by 
 means of copper, which takes its place in solution. The copper is 
 usually added in the form of scrap while the liquid is being heated 
 up by steam. The precipitation is assisted by constant stirring 
 by means of wooden paddles. In San Francisco, however, the 
 copper is cast into slabs, which are suspended side by side in the 
 solution in a vertical position. The solution should be of about 
 24 B. ; if it is much more concentrated than this, the precipita- 
 tion of the silver is imperfect. The end of the reaction is detected 
 by testing with salt solution, and when complete, the stirring 
 is stopped, the solution allowed to settle for two hours, and the 
 clear liquid tapped into lead-lined vessels, where further settling 
 of the suspended particles of silver takes place. The precipitate 
 of silver is thoroughly washed with boiling water in wooden 
 filters lined with lead, until the reaction for copper can no longer 
 be obtained with ammonia. Care must be taken that no frag- 
 ments of metallic copper remain with the silver. The metal is 
 then pressed, dried, and melted, and is usually from 998 to 999 
 fine, even without fusion with nitre, when the copper plates are 
 used for reduction. 
 
 Crystallisation of the Sulphate of Copper. This is effected by 
 alternate evaporation and crystallisation in lead-lined wooden 
 tanks. The solution, which is still of 24 B., is run from the 
 
408 THE METALLURGY OF GOLD. 
 
 cipitating tank into the evaporating pan and concentrated to 
 40 B. by heating with steam; thence it is transferred to the 
 crystallising tanks, where it is allowed to cool and remain for 
 from ten to twelve days. The mother liquor of 36 B. is then 
 run off and reconcentrated to 45 B., after which it is again 
 allowed to crystallise, reconcentrated to 55 B., and a third crop 
 of crystals obtained, which contain much iron. The clear acid 
 mother liquor can now be used to dilute the solution of sulphate 
 of silver in the dissolving pot as already described. The excess 
 of acid in surplus liquids is neutralised with oxide of copper, 
 more copper sulphate being thus formed. 
 
 The crystals of bluestone are found adhering to the sides and 
 bottom of the tanks. They are detached with copper chisels, 
 redissolved in pure water and recrystallised, the mother 
 liquors being eventually added to the first liquor from the pre- 
 cipitating vats. When the liquors become over-charged with 
 iron, the copper in them is precipitated by means of metallic 
 iron, and they are thrown away or evaporated to get the crystals 
 of sulphate of iron. The bluestone crystals are packed in barrels 
 for the market. One pound of metallic copper with 1-5 Ibs. of 
 sulphuric acid of 66 B. will make 4-5 Ibs. of crystallised sulphate 
 of copper. 
 
 The cost of the process of parting by sulphuric acid in Europe 
 is about one farthing per ounce troy of the parting alloy. 
 
 Combined Process. At the Philadelphia Mint a combined 
 process is used, nitric acid and sulphuric acid being employed 
 in succession. The alloys are granulated and digested with 
 concentrated nitric acid for six hours in the same manner as 
 has already been described ; the solution is then siphoned off, 
 and the gold washed two or three times with distilled water, by 
 decantation, subjected to a second boiling with strong nitric 
 acid, and subsequently sweetened in lead-lined filters with 
 boiling water. The gold is then introduced into cast-iron cylin- 
 drical kettles and boiled for five hours with strong sulphuric 
 acid, the gold being stirred up with an iron rod every ten or 
 fifteen minutes to prevent agglomeration, and the solution is then 
 ladled out and treated as already described, p. 407. For a 
 charge of 190 Ibs. of metal, 175 Ibs. of nitric acid are used in the 
 first boiling, and 50 Ibs. in the second. Some nitre is added to 
 the sulphuric acid. 
 
 The gold is washed thoroughly and sweetened in wooden 
 filters, boiling distilled water being poured through it until the 
 washings will no longer redden blue litmus paper. The silver 
 is precipitated from these washings as chloride by the addition 
 of salt. The gold is then pressed, dried, melted, and cast into 
 bars, which are from 998 to 999 fine. 
 
 The process is much cheaper than the nitric acid process, 
 costing 20 per cent, less for acids, and saving some fuel. The 
 
PARTING. 409 
 
 granulations contain 100 parts of gold in 333 of the alloy. After 
 the boiling in nitric acid very little silver is left with the gold. 
 The cost of refining is a little over one cent per oz. 
 
 The G-utzkow Process. This process of parting by sulphuric 
 acid was invented and patented by Mr. F. Gutzkow in 1867. It 
 has been extensively worked in Germany and in San Francisco, 
 and up to the year 1891 had been instrumental, on the authority 
 of Mr. Gutzkow, in refining one hundred million dollars' worth of 
 silver. It is fully described in Percy's Metallurgy of Silver 
 and Gold, p. 479, and only a brief account will be given here. 
 When the patent had expired, Mr. Gutzkow introduced and 
 patented several improvements on it, which will be described at 
 greater length. This improved process is now successfully at 
 work at the Consolidated Kansas City Smelting and Refining 
 Company's Works at Argentine, Kansas, where it was estab- 
 lished in the spring of the year 1892. 
 
 The original Gutzkow process, as employed at the San Fran- 
 cisco Assaying and Refining Works for many years, may be 
 summarised as follows : The bullion treated is of three kinds, 
 viz., (1) Gold bars from retorted metal, containing about 900 
 parts of gold, 10 to 20 of base metals, and the remainder silver ; 
 (2) Comstock silver bars or dore bars, usually containing 20 to 
 100 parts of gold per 1,000 ; (3) base bars from the Reese River 
 district and from pan-amalgamation of tailings, containing from 
 100 to 800 parts of silver, and the remainder chiefly copper, with 
 sometimes a little gold. The gold bars (1) are alloyed with 
 silver and granulated, but the others are cast into bars, and 
 parted in that form. The dore bars, when prepared for solution 
 in the acid, weigh about 100 Ibs. each, and are 12 inches long, 
 6 inches broad, and 5 inches thick. The base ingots are melted 
 with fine bars to reduce the average copper contents to 12 per 
 cent., and are cast into bars 1 inch thick, the gold from which is 
 only about 992 fine. 
 
 The boiling is done in flat-bottomed thin cast-iron kettles 
 (A, Fig. 58), of which the bottom is only J inch thick when 
 new, and J inch when worn out. The solution can be rapidly 
 heated, owing to the thinness of the iron kettles, and 200 Ibs. of 
 alloy are dissolved in four hours by means of 300 Ibs. of 
 sulphuric acid, which comes from the tank C, and is forced into 
 the kettle through the pipe / by the plunger d. The solution 
 is then siphoned off through the pipe /" into the tank E, 
 and diluted with a large quantity of hot mother liquor from a 
 previous crystallisation, which is mainly sulphuric acid of about 
 58 B. ; some water is also added, and the solution partially 
 cooled, so that some crystals of silver sulphate are enabled to 
 separate out and carry down with them the milky precipitate of 
 lead sulphate and any suspended particles of gold ; green basic 
 sulphate of iron also settles firmly. The clear solution is then 
 
410 
 
 THE METALLURGY OP GOLD. 
 
 siphoned off into H and cooled to 80 F., and almost all the silver 
 sulphate thus crystallised out. If the acid is concentrated, white 
 
PARTING. 411 
 
 soft crystals of bisulphate are formed, which is not desired ; if, 
 however, the acid is only at about 58 B., large hard yellow 
 crystals of monosulphate, free from acid but contaminated with 
 copper, are deposited. The mother liquor is pumped back into 
 the tank E or to the original acid tank, the device employed 
 for this purpose being to exhaust them of air, so that the acid is 
 sucked up without passing through any valves, which would 
 soon wear out. The crystals of sulphate of silver are transferred 
 to the filtering box I by iron shovels, and a hot solution of green 
 vitriol of 25 B. run on to them from G. This is at first mainly 
 occupied in dissolving the sulphate of copper, and the first portion 
 of the solution, after passiog through the filtering box, is run into 
 a storing vat, where the silver, incidentally dissolved, is precipi- 
 tated by copper, and the latter subsequently recovered by means 
 of iron. After a time, the copper being dissolved, the silver 
 begins to be reduced, the green solution of iron turning coffee- 
 brown ; the reaction is as follows : 
 
 2FeS0 4 + Ag 2 S0 4 = Fe 2 8 . 3S0 8 + 2Ag 
 
 The reduction may also be effected by sheets of metallic iron, 
 which is first converted into ferrous sulphate and then into ferric 
 sulphate, the silver being simultaneously reduced to the metallic 
 state. The brown solution of ferric sulphate is boiled with 
 metallic iron in K, in order to regenerate the ferrous salt. The 
 silver is washed, pressed, dried, and melted. The gold from the 
 original dissolving kettle is also washed in a filter, pressed, dried, 
 and melted. 
 
 Such was the original Gutzkow process as employed in 
 treating dor6 bars. Its chief advantage over the ordinary 
 sulphuric acid process was the saving of acid. In the ordinary 
 process, none of the acid used is saved, so that it is reduced in 
 amount as much as possible, but does not fall below twice the 
 weight of the silver dissolved. This reduction in the amount of 
 acid used makes the finishing of the dissolving a difficult and 
 delicate operation. In the Gutzkow process, however, only the 
 acid decomposed by the silver is lost; the weight of this is about 
 equal to that of the metal, the rest of the acid being all recovered 
 and used over again in the boiling. Moreover, the long and 
 tedious crystallisation of copper sulphate is avoided, and the 
 space required for the crystallising vats saved. However, 
 several large lead-lined vessels are required for the storage of 
 the various solutions, and the expense of these, as well as 
 the space required, is greatly reduced by the recent improve- 
 ments described below. 
 
 The New Gutzkow Process.* Mr. Gutzkow has lately 
 
 *This is fully described by Mr. Gutzkow in the Eng. and Mng. Journ., 
 Feb. 28, 1891, p. 257, and May 7, 1892, p. 497, from which this account is 
 summarised. 
 
412 THE METALLURGY OF GOLD. 
 
 pointed out that if a large amount of acid is used for the boiling, 
 not only is the silver more completely dissolved and the operation 
 greatly expedited, but the presence of a high percentage of copper 
 does not hinder the parting, as it is kept in solution by the excess 
 of free acid. Thus, for ordinary dore silver, he uses four parts of 
 acid to one of bullion ; for bars containing 20 per cent, of copper 
 he uses six parts of acid ; for still baser bullion, more acid, and 
 so on, never losing more than one part of acid for one of bullion, 
 and recovering the remainder. 
 
 The charge for a pot 4 feet in diameter and 3 feet in depth 
 is 400 Ibs. of dore silver : the pot is flat-bottomed, with a 
 basin-shaped pocket or well in the centre which is useful for 
 the collection of the gold. The bullion is first attacked by 
 fresh acid of 66 B., run in by gravity from a large tank, and, 
 when most of the silver has been dissolved, mother liquor from 
 a former operation is added, a pitcher-full at a time, until the 
 c'large is completely dissolved, which takes from four to six 
 hours. The lire is then moderated, and the pot tilled with 
 mother liquor to within 1 or 2 inches of the top, when the 
 temperature of the acid will have been so far reduced that only 
 faint fumes are discernible. If no fumes are visible the acid is 
 too cold and some silver sulphate will be precipitated, but other- 
 wise the large excess of acid will keep it in solution. The 
 well-stirred charge is now allowed to settle, which is perfectly 
 accomplished in ten minutes, as the yellowish slowly-subsiding 
 persulphate of iron is transformed to a greenish flocculeiit com- 
 pound by the water in the mother liquor, and this settles quickly 
 and carries all suspended matter to the bottom. More iron is 
 dissolved from the kettle than in the ordinary process, owing to 
 the greater dilution of the acid used in boiling. 
 
 The solution is now siphoned from the kettle by means of a 
 f-inch gas pipe into a large cast-iron vessel, only about 1 foot 
 deep, standing in a larger vessel which can be filled with water 
 for cooling the charge. Steam is blown into the still hot acid 
 solution through a lead nozzle, -J inch in diameter, pointing 
 vertically downwards. This both dilutes and warms the solution, 
 the heating being necessary in order to prevent crystallisation 
 of the silver consequent on the dilution. As soon as the dilu- 
 tion has proceeded sufficiently far to ensure the crystallisation 
 of the hard yellow monosulphate instead of the soft white bisul- 
 phate of silver, a point which is found by dipping out small 
 quantities at intervals, and observing their behaviour on cooling, 
 the steam is shut off and the vat cooled with water and left all 
 night. The silver crystals form a coating of about 1 inch thick, 
 which is contaminated with copper sulphate if the mother liquor, 
 by repeated use. has become saturated with it. The mother 
 liquor is now pumped back into the acid storage tank by the 
 creation of a vacuum, and the crystals of sulphate of silver are 
 
PARTING. 
 
 detached with an iron shovel and thrown into a filtering-box 
 provided with a false bottom. Cold distilled water is sprinkled 
 on the charge, and is allowed to filter through it and flow back 
 into the crystallising vat, until the greater part of the free 
 acid has been removed. The stream is then deflected into a 
 ''silver filter''' where any silver is precipitated that may have been 
 dissolved at the same time as the sulphates of iron and copper. 
 The silver filter is a lead-lined box, partly filled with precipitated 
 copper and provided with a false bottom. The silver separates 
 on the top of the copper as a spongy sheet, a corresponding 
 amount of copper being dissolved. When the crystals of silver 
 sulphate in the first-named filtering box have been completely 
 freed from acid, and from copper and iron sulphates, the stream 
 of water is discontinued. The spongy sheet of silver is then 
 removed from the " silver filter " box and treated with hot 
 water and a few crystals of silver sulphate to dissolve the copper 
 still retained by the sheet. During this whole operation of 
 sweetening the crystals of silver sulphate, only about 3 per cent, 
 of it is dissolved, as it is little soluble in cold water. The copper 
 solution, after passing through the "silver filter," is either run 
 to waste or precipitated by scrap iron in wooden tanks at a 
 nearly boiling temperature. 
 
 The crystals are now dried in an iron pan which is placed 
 above a furnace, and, after being mixed with about 5 per cent. 
 of charcoal, they are at once charged into a hot crucible in a. 
 melting furnace. The silver sulphate is reduced at a low red 
 heat to metallic silver, carbonic and sulphurous acid gases being 
 evolved. By the time the temperature of melting silver is 
 reached, these gases will have all passed away. The silver is, 
 toughened by adding nitre and borax until the so-called 
 "boiling" indicates that the sulphur has all been eliminated^ 
 and the metal is then cast into bars. 
 
 The gold residue in the dissolving kettle contains insoluble 
 sulphates of lead, iron, antimony, mercury, and often some 
 copper and silver. It is ladled out and boiled with water to 
 dissolve out the sulphates of silver, copper, iron, &c., and, after 
 thorough filtering, it is stirred in a dish with hot water, and 
 decanted on to a filter-cloth until the insoluble sulphates of lead, 
 <fec., have all been washed off, and the gold is left bright and 
 clean. The gold is stored until enough is collected to make a 
 200-oz. bar, which is usually brittle. The material collected on 
 the filter-cloth is re-washed once or twice to recover the particles of 
 gold from it, and can then be reduced with charcoal and cupelled. 
 If lead is present in the original alloy, part remains with the gold,, 
 and is dealt with in the manner which has been already described,, 
 but the greater part is carried off with the silver solution, and is 
 deposited both while the steam is being passed in, and also- 
 subsequently during crystallisation of the sulphate of silver, 
 
414 THE METALLURGY OF GOLD. 
 
 which is coated with it. The sulphate of lead is removed from 
 the crystals by stirring them well in a stream of cold water, 
 by which the light insoluble particles of lead and antimony 
 sulphate are carried away ; it can then be collected, reduced, and 
 cupelled. Any silver that may be dissolved in the course of this 
 washing is precipitated by copper as before. 
 
 The process is seen to differ from the original one in three 
 essential particulars : 1. The solution is diluted with steam 
 instead of with mother liquor, the amount of liquid in use, and 
 consequently the number of lead-lined vats required being thus 
 reduced. This is an important item, especially in the United 
 States where lead-burning is expensive, owing to the existence 
 of a powerful union. 2. The weak silver solution is precipitated 
 at once, instead of being stored in tanks to be used again or to be 
 precipitated at leisure. 3. The silver sulphate is reduced directly 
 with charcoal in a crucible in the furnace. This saves the 
 pressing of the silver and, what is of greater importance, avoids 
 the use of the solution of sulphate of iron, which needs to be 
 stored. The reduction in the crucible and subsequent melting 
 requires scarcely more fuel than would be used to melt the 
 pressed silver. One of the minor advantages of the process is 
 said to be that no stirring is required during the boiling, owing 
 to the large amount of acid used. This saves labour and enables 
 the acid fumes to be more easily condensed, as they are not 
 mixed with air, which in the ordinary way would enter through 
 the aperture left for stirring. The exclusion of air also helps to 
 prolong the life of the iron kettles by checking the attack on them 
 by sulphuric acid. 
 
 Titus Ulke, however, states* that the method of reducing 
 sulphate of silver by heating it with charcoal, has been tried by 
 the Consolidated Kansas City Smelting and Refining Company, 
 and abandoned, owing to the great volume of objectionable 
 gases evolved, probably accompanied by heavy losses of silver. 
 Gutzkow's original method of reducing sulphate of silver by a 
 solution of proto-sulphate of iron is now used at the company's 
 works at Argentine, and 36,000 ozs. of dore" (containing on an 
 average 4 per cent, of gold and 95 per cent, of silver) are 
 parted daily at a mean cost (in 1895) of 0'22 cent per oz., not 
 including cost of superintendence and office expenses.! The 
 cost in i892 at this works had been 0-35 cent per oz. The 
 wages are from $2 to $2 per day, and sulphuric acid costs one 
 cent per pound. 
 
 Miller's Chlorine Process. The use of chlorine gas for the 
 purification of molten gold was first proposed by Mr. L. Thompson 
 in 1838, and the results of his investigations were published in 
 the Journal of the Society of Arts J two years later. He stated 
 that "it has long been known to chemists, that not only has 
 gold no affinity to chlorine at red heat, but it actually parts with 
 * Mineral Industry for 1895, p. 348. t Ibid., p. 350. J Vol. liii., part i., p. 17. 
 
PARTING. 415 
 
 it at that temperature, although previously combined. . . . 
 This, however, is not the case with those metals with which it is 
 usually alloyed. It offers, therefore, at once an easy and certain 
 means of separation." 
 
 In 1867, Mr. F. B. Miller, Assayer of the Sydney Mint, 
 applied this property of chlorine to the separation of gold from 
 silver on the large scale, and his process has been in use at 
 Sydney ever since, being particularly suitable for the purpose 
 under the local conditions. Among these conditions may be 
 mentioned the facts that acid is very costly, and that there is a 
 scarcity of silver bullion containing small quantities of gold, 
 while the gold produced in Australia contains but little silver. 
 The result is that the ordinary parting processes would prove 
 very expensive, but the chlorine process can be applied cheaply, 
 as it requires very little acid, and is efficacious in removing 
 small quantities of silver from gold bullion which has not been 
 made up into alloys of definite composition. Before the intro- 
 duction of the chlorine process no attempt was made to extract 
 the silver from any of the native gold of Australia and New 
 Zealand which was coined at the Sydney Mint. Sovereigns 
 were manufactured containing several per cent, of silver, which 
 replaced part of the copper used as the alloying metal. These 
 sovereigns, some of which are still in existence, can be easily 
 recognised by their pale tint, due to the presence of silver. Such 
 sovereigns have not been manufactured since 1867. Besides 
 separating the silver, the chlorine process removes the small 
 quantities of lead, antimony, &c., which render most of the 
 Australian retorted gold brittle, and so in one operation prepares 
 the gold for coinage. Practically the whole of the gold produced 
 in Australia is now deposited in the mints of Sydney and Mel- 
 bourne, and refined by this process, the amount treated in 1892 
 having been 1,673,000 ozs. 
 
 The following description is abridged from that given by the 
 late Mr. Miller,* and from later writings : The furnace used is 
 an ordinary melting furnace, such as has been already described. 
 The tile cover of the furnace has a hole in the centre to allow 
 the chlorine tubes to pass through. French clay crucibles are 
 used, Nos. 17 and 18 being convenient sizes, holding about 600 
 or 700 ozs. of gold ; they are placed inside graphite pots to 
 prevent loss by cracking. They are glazed inside by melting 
 borax in them to prevent them from absorbing molten chloride 
 of silver. Graphite crucibles are said to be unsuitable, as silver 
 chloride appears to be reduced, presumably by the hydrogen 
 contained in them, as fast as it is formed. The crucibles are 
 covered by loosely fitting lids, through which the clay pipe-stems 
 of about T ^-mch bore are passed to the bottom of the crucible for 
 
 * Journ. Chem. Soc., 1868; and Trans. Roy. Soc. of New South Wale*, 
 1869. 
 
416 THE METALLURGY OP GOLD. 
 
 the conveyance of chlorine. The pipe-stem is made red-hot 
 before being introduced into the molten metal, as otherwise it 
 would crack and break off. The chlorine generator consists of a 
 stoneware jar furnished with three necks, and capable of hold- 
 ing from 10 to 15 gallons of liquid. The three openings are 
 fitted with well-secured rubber plugs, through two of which 
 two tubes are passed, viz., the safety tube, which is 8 or 10 feet 
 high, with its open end bent over so as to deliver into a large 
 jar, and the eduction tube, which is closed by a stopcock till it 
 is required. The generator is partly filled with from 70 to 100 
 Ibs. of manganese dioxide in small lumps, an amount which will 
 suffice for many operations ; hydrochloric acid is introduced 
 through the safety tube when the gas is required. The 
 generator is warmed by a steam jacket. 
 
 The chlorine gas is conveyed in leaden pipes to the furnaces. 
 All joints and connections are made by well- wired india-rubber 
 tubes, which must be protected from direct radiation from the 
 furnace. Screw compression clamps on these rubber tubes enable 
 the supply of chlorine to be regulated to a nicety. When the 
 clamps are closed the gas accumulates and forces the acid up the 
 safety tube into the vessel placed overhead, and so the further 
 generation of gas is prevented. Two such generators and three 
 melting furnaces are enough to refine 2,000 ozs. of gold, containing 
 10 per cent, of silver. 
 
 The generators being in readiness, the crucibles are slowly 
 heated to redness, and the full charge of 600 or 700 ozs. of bullion 
 introduced and melted, 2 or 3 ozs. of borax being sprinkled on 
 its surface or poured on in a molten state. The chlorine is now 
 allowed to pass slowly through the clay pipe to prevent " tal 
 from entering it, and the pipe is plunged to the bottom of the 
 molten metal and kept there by means of a weight attached to it. 
 The full stream of chlorine is now turned on and is heard to be 
 bubbling into the molten metal, by which it is completely 
 absorbed, so that no splashing and projection of the metal occurs. 
 A height of 16 to 18 inches in the safety tube corresponds 
 to and balances a height of 1 inch of gold in the refining crucible. 
 The safety tube acts as an index of the pressure in the generator 
 and of the rate of production of the gas : any leakage or the 
 exhaustion of the acid is at once indicated by a fall of the liquid 
 in the tube. Fresh acid is added at intervals as it is required. 
 
 When the chlorine is introduced, dense fumes at once arise 
 from the surface of the metal owing to the formation of volatile 
 chlorides of the base metals, which are the first to be attacked : 
 lead gives especially dense fumes, which can be condensed on a 
 cold object held in them. After a time these fumes cease and 
 silver chloride is formed, very little chlorine escaping from the 
 crucible, even if an extremely rapid current is passed into it ; 
 consequently the operation is expedited by every increase in the 
 
PARTING. 417 
 
 volume of the current. Towards the end of the operation 
 splashing is more noticeable, and dark brownish-yellow fumes 
 appear, consisting chiefly of free chlorine. The completion of 
 the refining, however, is indicated by a peculiar reddish or 
 brownish-yellow stain which is imparted to a piece of white 
 tobacco-pipe when exposed to the 'action of the fumes for a 
 moment. It is suggested by Prof. S. B. Christy that the stain 
 contains gold. This stain appears in about one hour and a-half 
 from the start, when 600 ozs. of gold, containing 10 per cent, of 
 silver, are being subjected to treatment. The current of gas is 
 then at once stopped, and the crucible lifted out of the furnace 
 and allowed to cool sufficiently for the gold to solidify. Pro- 
 bably, if the operation were continued after the appearance of 
 the brown stain, losses of gold by volatilisation would occur. 
 
 The chloride of silver, still molten, and floating on the top of 
 the gold, is then poured off into iron moulds, and the crucible 
 inverted on an iron table, when the red-hot cone of gold falls out. 
 This is now fine, and after any adherent chloride of silver has 
 been detached from it by scraping, it simply requires melting 
 into ingots, 98 per cent, of the gold being thus afc once rendered 
 available for use. The remainder of the gold is contained in the 
 chloride of silver, partly in the form of entangled shots of metal, 
 but chiefly as a double chloride of silver and gold. It was for- 
 merly recovered by melting the chloride with about 10 per cent, 
 of metallic silver, rolled to about J inch in thickness. The gold 
 is reduced by the silver and alloys with the excess, settling to the 
 bottom of the pot where it solidities after ten minutes cooling, 
 so that the chloride of silver can be poured off into large iron 
 moulds, slabs suitable for reduction being thus formed. 
 
 It was found at the Sydney Mint,* that the above method of 
 separating the gold from the silver chloride was subject to several 
 disadvantages. In particular, although on a small scale the 
 amount of gold in the silver could be reduced to from 0'3 to 1*0 
 per 1,000 by careful and continuous stirring with silver foil for 
 a great length of time, nevertheless in practice on a large scale 
 the results varied greatly, and the silver bullion produced usually 
 contained from 10 to 25 parts of gold per 1,000. Several re- 
 ducing agents, such as argol, resin, hydrogen and coal gas were 
 successively tried but were not found to give good results. 
 Finally, the application of soda carbonate, which had been pro- 
 posed by Leibius in 1868, was adopted, the method of procedure 
 being originally as follows : f"The argentic chloride is covered 
 by a layer of fused borax, about J inch thick, and when all is 
 well fused, the powdered soda is sprinkled on the top of the 
 borax, without stirring, as rapidly as the ensuing action will 
 
 * Fourth Annual Report of the Royal Mint, 1873. Report by A. Leibius, 
 p. 63. 
 
418 THE METALLURGY OF GOLD. 
 
 admit. Occasionally the top layer is dipped with a stirrer 
 slightly underneath the molten argentic chloride, with out stirring 
 the latter. When all the necessary soda is added and the action 
 is nearly over, the pot is covered with a lid, and left for about 
 ten to twenty minutes to increased heat, and, when the contents 
 are quite liquid, the pot is lifted out of the fire without previous 
 stirring, and allowed to cool, so as to enable the argentic chloride 
 to be poured off from the gold button at the bottom of the pot. 
 
 " Although in several experiments all but 0-1 of gold per 1,000 
 was eliminated from the silver bullion produced, in no case is 
 every trace of gold removed in one operation. To free the 
 argentic chloride entirely from gold, producing therefore silver 
 bullion free from gold, was, however, accomplished by subjecting 
 the argentic chloride to a second treatment, with a small quantity 
 of soda, in a separate boraxed clay pot, similar to the first opera- 
 tion. 
 
 " A convenient quantity of argentic chloride, to be treated 
 in a No. 18 French clay pot, was found to be 230 ozs. The 
 amount of soda required for 230 ozs. of chloride may range 
 from 16 to 20 ozs. Less than 16 ozs. leaves too much gold in 
 the silver, while more than 20 ozs. produces a very silvery 
 gold button, and yet without completely freeing the argentic 
 chloride from gold. 
 
 "The use of 18 ozs. of soda for 230 ozs. of chloride pro- 
 duces a gold button weighing between 30 and 35 ozs., assaying 
 about 920 to 930, and leaves from 0-5 to 1-0 part of gold in 1,000 
 parts of silver bullion produced. 
 
 "With 20 ozs. of soda the results were : Gold, about 35 ozs., 
 assay 870-880 ; gold left in the silver bullion produced from 0-2 
 to 0-5 per 1,000. 
 
 "With 16 ozs. of soda: Gold from 30 to 33 ozs., assay 
 940-950; gold left in the silver bullion, from 1-0 to 2-0 per 
 1,000, and sometimes as much as 6'0 per 1,000. 
 
 " To free the argentic chloride from gold, a second treatment 
 with 3 ozs. of soda per pot of 200 ozs. chloride, containing but 
 a minute quantity of gold, will always be lound to answer, the 
 only care required being gradual application of the soda and 
 enough heat at the end of the operation." 
 
 The time required for the two operations is about half an 
 hour. 
 
 "The presence of a large proportion of chloride of copper has 
 been found to prolong the operation considerably on account of 
 oxide of copper being formed on addition of soda, as a much 
 greater heat is required in order to fuse the whole mass. The 
 argentic chloride produced from base gold alloys would contain 
 a large proportion of chloride of copper, &c., and it would be 
 better, therefore, to reduce it direct, and dissolve the reduced 
 metals in acid, to separate gold and silver therefrom." 
 
PARTING. 419 
 
 The silver chloride may be assayed for gold by cupellation 
 with lead foil and subsequent parting. 
 
 The method just described was adopted at the Sydney Mint 
 in 1872, and at the Melbourne Mint in the following year, with 
 excellent results. 
 
 The process of reduction of the silver chloride was devised by 
 the late Mr. A. Leibius, fellow-assay er of Mr. Miller at the Sydney 
 Mint, and was described in a paper communicated to the Royal 
 Society of New South Wales.* In this process, 1,400 ozs. of 
 argentic chloride are reduced in 24 hours by the apparatus, of 
 which the following is a brief description : Seven zinc plates, 
 each 14 inches long, 12 inches wide and J inch thick, are sup- 
 ported about 1J inches apart in a vertical position in slots 
 in a wooden frame. Six slabs of argentic chloride, each 12 
 inches long, 10 inches wide and f inch thick, are suspended by 
 loops made of silver bands, in such a way that each slab is placed 
 between two of the zinc plates and separated from them by 
 spaces of about J inch. The silver loops are connected with silver 
 bands on which the zinc plates rest, so that there is metallic 
 connection between the slabs of chloride and the zinc plates. 
 The whole is now plunged into water, to which some of the liquor 
 from a previous operation containing chloride of zinc in solution 
 is added as an exciting agent. Galvanic action soon begins, the 
 liquor gets gradually warmer and a strong current is discernible. 
 The silver chloride is gradually reduced to metallic silver, the 
 slabs undergoing no alteration of form, and the zinc is dissolved. 
 The slabs of silver chloride are generally free from most of the 
 base metals, but copper, if present in the original alloy, is not 
 volatilised in the crucible, and its chloride remains mixed with 
 that of the silver. The two metals are now reduced together. 
 When all action has ceased, the slabs of cupreous silver are lifted 
 out and boiled, first in acidulated water and then in pure water, 
 while still suspended in their silver loops. The porous metal is 
 now ready for melting. As no acid is used the amount of zinc 
 consumed is the theoretical quantity required by the equations 
 
 2AgCl + Zn = ZnCl 2 + 2Ag 
 CuCl 2 + Zn = ZnCl 2 + Gu 
 
 The weight of zinc consumed usually amounts to from 24 to 25 
 per cent, of the weight of the slabs of fused chloride. The zinc 
 plates are used over again until worn too thin for safety, after 
 which they are melted-up and cast into new plates. They suffer 
 no loss if the apparatus is left untouched .for any length of 
 time after the whole of the silver has been reduced. 
 
 At the Melbourne Mint in the year 1889, the zinc plates 
 
 * Trans. Roy. Soc. Sew South Wales, 1869. The paper is given almost 
 at full length in Percy's Metallurgy of Silver and Gold, p. 418. 
 
420 
 
 THE METALLURGY OF GOLD. 
 
 employed as described above were replaced by sheets of iron 
 with satisfactory results. " Upon the reducing bath being- 
 heated with steam, the chloride of copper dissolving, dis- 
 engages itself freely from the slabs of chloride of silver, and 
 coming into contact with the iron is reduced, and the metallic 
 copper falls to the bottom of the bath in large quantities, leav- 
 ing the reduced silver in a much cleaner state than when zine 
 was used. The noxious fumes which were formerly given off on 
 the melting of the reduced silver sponge are also avoided." * 
 
 Fig. 58a. 
 
 The Chlorine Process as now practised at Melbourne. The 
 following description has been kindly supplied by Mr. Francis 
 R. Power, the Assayer at the Royal Mint, Melbourne, by per- 
 mission of the Deputy Master. It gives the exact methods and 
 apparatus in use in the early part of 1896. As will be seen, 
 these differ considerably Irom those described above, and from 
 the practice at Sydney. 
 
 * Twentieth Annual Keport, Royal Mint, 1889, p. 126. 
 
PARTING. 421 
 
 Furnaces. These are thirteen in number. They are built 
 cylindrically (see Fig. 58a, in which one of these furnaces is 
 shown in section, with crucible and pipe-stem in position), being 
 more compact in this form, more easily cleaned from clinker, 
 and more economical in fuel than the square ones. They are 
 12 inches in diameter and 21 inches deep. The five firebars, 
 1^ inches square and 18 inches long, are set in a cast-iron box, 
 D, 12 inches by 2 inches, which passes through the brickwork 
 in front of the furnace, the other ends of the firebars resting on 
 an iron bar set in the brickwork at the back of the furnace. 
 The bars are 6 inches above the floor. The draught is obtained 
 through a grating in the floor, which covers a portion of the ash- 
 pit, over which there slides a cast-iron plate, M, J inch thick, for 
 regulating the admission of air, and pivotted in one corner. The 
 flue, L, is 6 inches square, and communicates with a series of 
 five condensing chambers, 8 feet by 8 feet by 5 feet, running the 
 length of the furnaces (42 feet), all communicating and leading 
 to the stack, 80 feet high, common to refining and melting fur- 
 naces, which are twenty-one in all. There are three furnace 
 covers, two of them 20 inches by 6J inches, the third a little 
 smaller, and all are bound with iron. The middle one is per- 
 forated by a 1-inch hole, through which the chlorine delivery 
 pipe passes. Glenboig arched firebricks, B, 9 inches by 4J inches, 
 and tapering from :2J inches to 2 inches, are used for lining the 
 furnaces, and are set with touching joints in an iron cylinder, A, 
 21 TJ inches in diameter, and at least to |- of an inch thick, 
 which is supported by a cast-iron plate, C, -f of an inch thick, 
 -and 22 inches in diameter, with a 12-inch hole in the centre. 
 This plate is supported by the brickwork which forms the 
 foundation. The ashpit is a cast-iron flanged box, easily 
 cleaned in case of an accident. Round the iron cylinder 
 concrete, N, is rammed, the front iron plate of the furnace 
 being shifted 2 or 3 inches in, until this is set and then moved 
 out, thus providing an air space, E, and keeping the plates 
 cooler. The furnace top is a plate of cast iron and, so as to 
 facilitate repairs, should be in two pieces for each furnace, halved 
 into one another, the hole being slightly bossed at the edge so 
 that the firetiles may run easily on them. One piece has a hole 
 -6 inches in diameter over which the swing ventilating hood, P, 
 is placed by which the pot is covered when removed from the 
 fire. This hood communicates by a passage through the brick- 
 Avork with the flue. The cylindrical furnace is calculated to 
 last for three years, the square ones lasting only eighteen months 
 .and taking three hours to reline, while the cylindrical ones 
 take one hour. 
 
 The Crucibles, &c. The guard pot, placed for safety under the 
 white pot and afterwards used for remelting the refined gold, 
 is a plumbago crucible 8^ inches high, 6 inches inside diameter, 
 
422 
 
 THE METALLURGY OF GOLD. 
 
 J of an inch thick at the top, and f- of an inch at the bottom, 
 which is flat inside and stands on a cylindrical firebrick 5 inches 
 in diameter and 2 inches deep. The white pots, fitting loosely 
 into the guards, are 10| inches high, 5 inches in diameter, and 
 f of an inch thick at the top, tapering from L inch at the bottom. 
 They are covered by a closely-fitting lid, dished at the top to 
 catch any globules spirted out by too rapid a current of gas 
 and perforated by two holes f of an inch in diameter. A. new 
 pattern of lid to be introduced shortly will have a notch in 
 the edge for the pipe-stem to pass through, the advantage of 
 this being the easy removal of the lid without withdrawing 
 the pipe-stem, as is necessary with the old lids. 
 
 Fig. 586. 
 
 The pipe-stem is 24 inches long, tapering from f to \ an inch 
 at the end inserted into the gold, and is wedge-shaped to facilitate 
 the escape of the chlorine when resting on the bottom of the 
 pot. The bore of the pipe-stem is \ of an inch in diameter. 
 The thin end of the pipe-stem is attached to the branch delivery 
 pipe by a piece of |-inch rubber about 2^ inches long, which 
 connects with an ebonite junction, G, 3 inches in length, with 
 a bore of T V of an inch, turned with a ring round the middle, 
 which acts as a rest for the 8-oz. weight, H, used as a sinker 
 for the pipe-stem. One end of the ebonite junction is an inch 
 in diameter, the other J of an inch ; the latter being connected 
 
PARTING. 
 
 423 
 
 by a stout rubber tube 3 or 4 inches long to a 14-inch lead pipe 
 (^ an inch in diameter) which is attached [by a rubber junction] 
 to a glass stopcock, I, from the spigot of which a J-lb. lead 
 weight, J, is suspended to prevent the pressure of gas from 
 blowing it out. The glass stopcocks have replaced the com- 
 pressor clamps, which were not satisfactory owing to the rubber 
 cutting through, and chlorine leaking past. The rubber joints 
 are sufficiently flexible to allow the pipe-stem to bend down 
 into the pot or to be laid horizontally on a rest when not in use. 
 Each furnace is provided with one glass stopcock to control 
 the flow of gas. The cock is far enough away at the back of 
 the furnace, to be unaffected by the heat when the fire tiles 
 
 Fig. 5Sc. 
 
 are removed, and is connected by a J-inch lead pipe with the 
 main pipe running along the wall. Thinner pipe-stems are 
 found to be as serviceable as the above and do not require 
 such careful annealing. The tubes, stopcock, &c., are some- 
 what diagrammatic in Fig. 58a, and can be studied in Fig. 586, 
 which is from a photograph of the furnaces by Mr. R. Law. 
 This also shows the crucibles, guard pots, ventilating hoods, and 
 chloride cakes. 
 
 The generators are shown in Fig. 58c, which is also from a 
 photograph by Mr. R. Law. The pressure regulator and the 
 reduction tank are also shown. The generators, eight in number, 
 
424 THE METALLURGY OF GOLD. 
 
 are three- necked cylindrical stoneware vessels with domed 
 tops, and having a flange round the middle by which they are 
 supported on the stoneware steam jacket 16 inches high, 16 J 
 inches in diameter, and | of an inch thick. The domed vessel 
 is 2 feet high. The three necks have 1 {-, 1|-, and If -inch holes, 
 the first for charging-in the manganese ore and closed by an, 
 indiarubber plug, the second for the pipe leading to the main 
 chlorine pipe, arid the third for a branch acid supply tube, 5 an 
 inch in diameter, fitted with a glass stopcock a foot above the 
 neck, between which and the stopcock another J-inch tube, 
 the overflow, branches. The overflow tube, through which the 
 hot waste of the generators has to pass, is provided with an 
 ebonite stopcock which is turned off' daring refining. Stout 
 combustion tubing is used. The bottom of the generator is 
 covered by 4 inches of quartz pebbles, to prevent choking of 
 the acid delivery pipe, which reaches to within 1 inch of the 
 bottom of the vessel. 56 Ibs. of manganese ore (about 73 per 
 cent, peroxide), broken to to ^ inch square, is placed on the 
 pebbles, and commercial hydrochloric acid, of specific gravity 
 1*16, is added as required through the acid delivery pipe by 
 turning the glass stopcock. The acid pipe is of glass, and leads 
 to the eight storage tanks 20 feet above the floor, which hold 
 320 Ibs. of acid each, and are interconnected by glass tubes luted 
 into the bottoms ; the delivery of acid, however, being from one 
 at a time. The gas delivery pipes from the generators all connect 
 with a 1 inch lead pipe, which leads to a distributing vessel 
 with two necks and partially filled with manganese chloride 
 solution. A pressure guage of 1 inch glass tube and 15 feet 
 high is luted into the bottom of this vessel, and is fixed to the 
 wall by brackets, 10 to 11 feet of the solution being required to 
 overcome the resistance of about 7 inches of metal in the pots. 
 The pressure in refining is equal to 5 Ibs. per square inch. 
 A four-way tube of lead or pottery is passed through the second 
 neck of the vessel, and each arm is connected by thick rubber 
 to glass stopcocks to which |-inch lead pipes are joined, these 
 pipes leading to sets of four, four and five furnaces, so that the 
 supply of gas can be delivered to a few or all the furnaces, as 
 desired, the subdivision being made for safety in case of a 
 leakage or for convenience it only a few furnaces are in use. 
 All the generators are used whether the quantity of gold to 
 be refined be large or small, the same quantity of acid being 
 run into each. When the flow of chlorine through the gold is 
 stopped the acid in the generators is forced back through the 
 overflow pipe by opening the ebonite tap. It is found necessary 
 to have the main pipe in communication with another two- 
 necked earthenware vessel containing such a quantity of water 
 that when the pressure of gas exceeds the working pressure 
 required, the end of a glass tube, passing to the bottom of 
 
PARTING. 425 
 
 the vessel and connected above the neck with an upright 4-inch 
 lead pipe 10 feet high, becomes unsealed, and the gas escapes 
 through the water in large bubbles, escaping through a glass 
 pipe, inclined at an angle at the top of the lead pipe, into the 
 air. When sufficient gas has escaped to reduce the pressure to 
 the working limit the pipe is sealed. Thus the pipe acts auto- 
 matically in keeping the pressure below such an amount as 
 would endanger the apparatus or cause joints to leak. It is 
 found expedient to cover all the rubber junctions in the 
 generating room with calico and then to paint it. Protected 
 in this way it will last until stopped up by the action of the 
 chlorine which tills it with a lemon-yellow incrustation, at the 
 same time reducing its thickness. All junctions are secured with 
 copper wire where practicable. 
 
 Refining Operations. The guard, with the white pot in it 
 containing 2 or 3 ozs. of fused borax, is placed in the furnace, 
 and is heated gradually until the bottom of the white pot is 
 dull red. The ingots (of which the larger are slipper-shaped) 
 to be refined, amounting in all to 650 to 720 ozs. in weight, are 
 then placed loosely in the pot, the furnace filled with fuel, and 
 the dampers opened. As soon as the gold is melted, which 
 generally happens in about one and a-half hours, the boraxing 
 of the pots being also effected at the same time, the perforated 
 lid is put on, and the pipe-stem, previously brought carefully 
 to a red heat to prevent cracking or flaking, is pushed to the 
 bottom of the pot. As the pipe is being inserted, the chlorine 
 is gently turned on to avoid stoppage of the passage through the 
 stem by the solidification of metal in it. The supply of chlorine 
 is controlled by the glass stopcock over the furnace, and the 
 amount is adjusted so that the whole of the gas is absorbed and 
 no globules of metal can be thrown up. This can usually be 
 ascertained by feeling the pulsations of the gas through the 
 indiarubber connections as it escapes in bubbles out of the 
 bottom of the pipe-stem. When the gold contains much silver 
 or base metals, the absorption of the chlorine takes place rapidly 
 but gently, very little motion of the contents of the crucible 
 being apparent, but when the gold to be refined is of high assay 
 and also in "all cases towards the end of refining, the gas is 
 admitted only in a small stream, and requires careful watching 
 to prevent spirting. When base metals are present in large 
 quantities (over 2 per cent.) dense characteristic fumes of the 
 chlorides of these are given off. and the metal or metals present 
 may be generally identified by the fume or incrustation caused 
 by the condensation of the base chlorides on the pipe or lid. 
 
 Gold containing 2 per cent, of silver and 0*5 per cent, of base 
 metal is refined to about 995 fine in one and a-half hours, while 
 that containing 3 -5 per cent, of silver and 1-5 per cent, of base 
 metals takes two hours. When larger percentages of silver or 
 
426 THE METALLURGY OF GOLD. 
 
 base metals are to be dealt with, the time taken is not pro- 
 portionately longer, because, as mentioned above, a much greater 
 stream of gas may with safety be admitted, though, in all cases, 
 at the beginning the chlorine must be introduced gently on 
 account of there being air in the chlorine mains, and, also, at the 
 end of refining, the supply must be greatly reduced. When 
 nearing completion, the " flame " issuing from the holes in the 
 lid becomes altered in appearance, and much smaller; it now 
 contains much chlorine mixed with small quantities of the 
 volatile chlorides. The actual completion of the operation is 
 generally known by the appearance of a very characteristic 
 " flame," which is luminous, with a dark brown fringe. In case 
 of doubt, a piece of clean pipe-stem is used as a test. It is 
 placed, cold, for a few seconds in the issuing flame, and if the 
 refining is finished, a clear reddish-brown stain, tending to 
 yellow, is imparted to the test end. This stain consists of ferric 
 oxide and chloride from the oxidation of ferrous chloride, and 
 contains gold and sometimes chloride of silver, and is probably 
 caused by small quantities of chloride of iron retained by the 
 fused gold and non-volatile chlorides, from which it is freed by 
 the unabsorbed chlorine bubbling through. Traces of copper 
 and iron are always found in the refined gold, the bulk of the 
 alloy being silver. As soon as the stain is found to be of the 
 right colour, the current of gas is reduced, and is allowed to 
 pass for a further fifteen minutes, and the pipe is then with- 
 drawn and the clay pot lifted out of the guard. The pot is 
 allowed to stand under a hood (to carry off the fumes) until the 
 gold is set, which usually takes place in from five to seven 
 minutes, the fact that solidification has taken place being 
 observed by thrusting a piece of red-hot pipe-stem down through 
 the fused chlorides. The chloride is then poured into a mould 
 provided with a ventilating hood, which, in consequence of the 
 high density of the fumes necessitating a sharp draught to 
 remove them, is connected with' the stack. Any borax poured 
 off with the chlorides is allowed to remain, as it is required 
 as a cover for the chlorides in the subsequent fusion for the 
 separation of their gold contents. The pot is then broken, 
 as the cone of gold will not fall out of it soon enough, and the 
 cone of refined gold is remelted in the guard and cast into two 
 flat ingots, 12 inches by 4 inches by 1 J inches, which, when set 
 and still red hot, are placed on a copper lift, dipped in dilute 
 sulphuric acid and then in water, and after removal from the 
 water are still sufficiently hot to dry by their own heat. The 
 broken pots are ground in a small Chilian mill and panned off, 
 and the gold obtained is added to the " end " that is returned at 
 the end of the day. 9,000 ozs., containing up to 10 per cent, of 
 silver and base metals, constitute a day's refining. 
 
 An improvement has recently been introduced, by which a 
 
PARTING. 427 
 
 considerable saving of time and material is made. This is the 
 dipping of the fused chlorides and borax from the pot while it is 
 still in the fire, and without previous solidification of the gold, 
 by a small clay crucible, from which they are poured into a 
 covered mould projecting over the furnace, the drops falling 
 back into the pot. This had previously been the practice when 
 the percentage of silver was large, as the silver doubles its bulk 
 on conversion to chloride, and would have overflowed. The last 
 " dip " always contains some gold, and is poured into a separate 
 mould, in which the metal sets at once. The chloride is thence 
 poured into the larger mould, and the gold returned to the pot. 
 The chloride remaining in the pot is then made into a paste 
 with bone ash, after which the refined gold is stirred and cast 
 into ingots, the pot being at once returned to the furnace to be 
 used a second time. 
 
 The chlorides, which hold from 5 to 10 per cent, of gold in 
 feathery particles, are remelted during the day in plumbago 
 crucibles holding 300 ozs. of chloride. When fused, 7 per cent, 
 of their weight of bicarbonate of soda is added, cautiously and 
 without stirring, which produces a shower of globules of reduced 
 silver, and these falling through the chlorides carry down nearly 
 all the gold. As one addition of bicarbonate of soda does not 
 entirely free the chlorides from gold, a second addition is made, 
 without removing the crucible from the furnace, ten minutes 
 being allowed after each addition. The pot is then lifted out 
 and placed on one side to allow the metal to set, when the 
 chlorides are poured into a mould 12 inches by 10 inches by 
 2 inches, practically free from gold and ready for reduction. 
 The silvery button obtained contains from 40 to 60 per cent, of 
 gold, and is refined on the following day. 
 
 The silvery ingot and the refined gold contain 99 '85 per 
 cent, of the gold issued in the morning for refining, the bulk of 
 the deficiency being in the pots. The amount of gold which 
 goes immediately into work after refining is 9 7 '6 per cent, (an 
 average of thirty days refining). The amount of gold left in the 
 silver after reduction from the chloride reaches a maximum of 
 1 part in 10,000, but is usually from to J of this quantity. 
 
 The cakes of impure chloride, weighing about 250 ozs. each, 
 may be colourless and translucent to brown or chocolate colour 
 and opaque, the colour depending on the amount of copper salt 
 present. They consist of argentic and cuprous chlorides, with 
 traces of other chlorides and 9 per cent, of chloride of sodium 
 from the decomposition of silver chloride by bicarbonate of soda. 
 When cool, each cake is sewn up in a coarse flannel bag to 
 prevent loss of any silver which may become detached during 
 reduction, and they are then boiled with water in a wooden vat 
 for four or five hours. By exposure to air and moisture the cakes 
 become coated with a green deposit, owing to the conversion of the 
 
428 THE METALLURGY OF GOLL>. 
 
 cuprous chloride into cupric oxychloride and hydrated cupric 
 chloride, and the successful removal of a large proportion of the 
 cuprous salt in the vat is due to its solubility in a hot solution 
 of common salt and of hydrated cupric chloride, from which it 
 is redeposited on dilution. 
 
 The cakes for reduction are placed alternately with wrought- 
 iron plates ^ inch thick in a cast-iron tank lined with similar 
 plates. The plates are prevented from touching the bags by 
 laths of wood, otherwise the copper would be reduced in the 
 bags, and would be difficult to separate from the silver. The 
 reduction is slow in starting, unless either some liquor is left 
 from a previous operation or some chloride of iron added. The 
 bath is heated by the direct injection of steam (this is absolutely 
 necessary), and the reduction is complete in from two to four 
 days, though sometimes it takes longer. The time may be 
 lessened to twenty-four hours by putting the chloride cakes in 
 metallic connection with the iron plates. The completion of the 
 reduction may be easily ascertained by feeling the cake, when, if 
 -any chloride be unreduced, it is felt as a hard lump. The 
 reduced silver is taken out of the bags, washed in boiling water 
 for about an hour, and then melted, no fluxes being necessary. 
 The use of the flannel bags makes the reduced silver of high 
 standard, as the reduced copper is thus prevented from adhering 
 to the silver cake, from which it was found very hard to detach 
 it without tearing off silver as well. A small percentage of 
 reduced silver and silver chloride is found at the bottom of the 
 tank, its presence being probably due to the solubility of the 
 chloride in solutions of the chlorides of copper, iron, and sodium. 
 In 1894, the mean standard of the refined silver was 982-2, 
 the lowest ingot being 955 fine, and the highest 998. In 
 1895, the mean standard was 981 '6, the lowest being 936, and 
 the highest 995 fine. 
 
 The mean fineness was only about 930 prior to the year 1889. 
 In 1889, after the introduction of the iron plates in place of zinc, 
 the fineness rose to 948. The subsequent improvement was 
 probably due to the introduction of the flannel bags. 
 
 The silver contained in the gold operated on is distributed in 
 the following manner, the mean results of the last five years 
 (1891-95) being given : 
 
 Silver in ingots, 88 "77 per cent. 
 
 ,, left in refined gold, . . . 7 '62 ,, 
 
 in "sweeps," . . . . . 2'00 
 
 ,, unaccounted for, . . . . 1'60 ,, 
 
 The " sweep " from the condensing chambers amounts to about 
 3 cwts. per annum, and contains an average of 41 ozs. tine 
 gold and 157 ozs. fine silver, which are carried over as globules 
 or volatilised as chloride and condensed. 
 
PARTING. 429' 
 
 The mean amount of gold refined per annum during five years 
 (1891-95) was 949,527 ozs., containing gold 937'7, silver 49-6, 
 and base metals (by difference) 12 -7. The mean assay of the 
 refined gold for the same period was 995 '9, and the mean loss of 
 gold in the refining operations for the same period was 0-175 per 
 thousand. 
 
 The approximate cost of refining per ounce gross weight 
 refined was as follows : 
 
 In 1894. In 1895. 
 
 Material, . 0' 1397 of a penny. 0' 12 15 of a penny. 
 
 Wages, . 0-1485 0'1439 
 
 0-2882 0-2654 
 
 Half the cost for materials was for hydrochloric acid at <20, 15s. 
 per ton. 
 
 The amount of gold refined in 1894 was 1,049,529 ozs., 
 containing in parts per thousand gold, 933-9; silver, 51-4; 
 base metals, 14-7 ; and the gold refined in 1895 amounted to 
 1,083,243 ozs., containing gold, 932-0; silver, 52-3; and base 
 metals, 15-7 parts per thousand. 
 
 In some experiments made by Mr. Barton, who took gold 
 alloyed separately with 4J per cent, of copper, 4J per cent, of 
 lead, 4 per cent, of iron, and 4J per cent, of tin, when the cost 
 of hydrochloric acid was 2Jd. per Ib. and manganese ore (70 per 
 cent, peroxide) Id. per Ib., the following results were obtained, 
 operating on quantities of gold containing 30 ozs. of each of 
 these metals : 
 
 Cost of extracting copper, . . 4 -5 pence per oz. Troy. 
 
 ,, lead, 1'4 ,, 
 
 ,, iron, . . 3'9 
 
 ,, tin, . . 2-5 
 
 The gold in each case was brought up to the usual fineness. 
 These results give approximately the cost of extracting these 
 metals in the quantities in which they ordinarily occur in 
 Australasian gold. 
 
 The Miller Process at the Sydney Mint. Mr. J. M'Cutcheon, 
 the Assayer at the Sydney Mint, writes that the process of 
 freeing the chlorides from gold now used by him is as follows : 
 The chlorides produced during the operation are separated into 
 two classes, termed "balers" and "non-balers." The first is 
 that portion baled, or rather ladled, out during the operation, to 
 prevent overflow; this is re-melted in quantities of 350 ozs., 
 and whilst in the molten state, half a pound of bicarbonate of 
 soda is projected on the surface. This has the effect of reducing 
 some of the chloride, and the metal in sinking to the bottom of 
 the pot carries with it all the gold. The " non-balers," or that 
 portion of chloride which is poured off the refined gold when it 
 
430 THE METALLURGY OF GOLD. 
 
 ;has set, is treated as above, but 7 ozs. of granulated zinc is 
 used instead of the bicarbonate. 
 
 The chlorides are poured into slabs, and are now ready for the 
 reduction process, in which the silver loops formerly used have 
 been abandoned, iron plates being now used instead of zinc ; the 
 water is acidulated with hydrochloric acid. 
 
 The bullion treated at the Sydney Mint during 1895 contained 
 gold, 832-1; silver, 135-5; and base metals, 324 parts per 1,000. 
 
 The average fineness of the gold produced by this method at 
 .Sydney in the period of nine years from 1884 to 1892 was 995*9. 
 The remainder is silver, which apparently cannot be profitably 
 removed by clilorine or by any other method. The average 
 fineness of the silver produced at Sydney was about 970 in 1893, 
 the fineness of individual bars varying from 917 to 987. The 
 .alloying metal in the silver bars is almost all copper. The 
 gold retained by the silver formerly amounted on an average to 
 l'-3 parts per 1,000, but it has since been materially reduced 
 .and is now a mere trace. Analysis of the silver resulting from 
 the refiriage of gold, known originally to have contained, among 
 the base metals in the alloy, copper, lead, antimony, arsenic, 
 .and iron, gave the following results : 
 
 Silver, . . . . 972 '3 
 
 Copper, .... 25-0 
 
 Gold ..... 2-7 
 
 Zinc and iron, . . . Traces. 
 
 1000-0 
 
 The losses of gold in the course of the process are very small, 
 varying from 0-11 to 0*19 per 1,000; this is considerably less 
 than would have been lost by merely toughening the gold with 
 corrosive sublimate without parting it from the silver. The 
 loss of silver at Sydney was about 4-i'5 per cent, in 1895. These 
 losses are reduced if the amounts recovered from the flue-dust 
 .and from the ground-up crucibles are taken into account. The 
 total cost of the process is about Id. per oz. of crude gold at 
 .Sydney, and was about 0'65d. per oz. in Melbourne in 1873, 
 but has since been reduced. The loss of gold by volatilisation 
 is probably prevented from reaching the large amounts which 
 might be expected from the results of Christy and of the author 
 (see p. 22) by the fact that, during the whole time that the 
 chlorine is being passed, silver and base metals are present, 
 and, by absorbing the gas, protect the gold from its action. 
 
 Professor Thomas Price, who treated some of the Californian 
 gold-bullion by this method on a working scale in his laboratory 
 in San Francisco, states * that, with Californian gold, which 
 generally contains more silver than Australian gold, the gold 
 
 * Trans. Am. Inst. Mining Eng. t 1888, vol. xvii., p. 30. 
 
PARTING. 
 
 431 
 
 taken up by the chloride of silver amounted to 5 per cent., 
 and even to 10 per cent, of the total weight of the 
 gold. For this reason, and on account of the large amount of 
 silver bullion in the San Francisco market requiring parting, 
 Professor Price considers that the Miller process, while techni- 
 cally successful with Californian gold, is hardly able to compete 
 commercially with the ordinary sulphuric acid process. He 
 suggests that it might be well adapted for refining the nearly 
 pure brittle gold produced at chlorination works, where chlorine 
 is at hand and other methods of refining are not convenient. 
 The idea of refining brittle gold in this way had previously 
 been acted on by Professor Roberts- Austen in the year 1871 in 
 performing the experiments and operations detailed below. 
 
 Refining Brittle Gold by Chlorine Gas. In spite of the fact 
 that brittle ingots of gold are not accepted for coinage at the 
 Royal Mint, a number of such ingots formerly found their way 
 into work there, their lack of ductility not being observable 
 until after they had been alloyed with the copper requisite to 
 .make standard gold, which is 916-6 fine. In 1871, no less than 
 40,000 ozs. of such brittle standard gold had accumulated at the 
 Mint, having been set aside as unfit for coinage. These bars 
 contained as impurities traces of antimony, lead, bismuth, &c. 
 In order to determine how far such impurities could be elimi- 
 nated by chlorine gas, applied as in Miller's process, the following 
 experiments were conducted by Professor Roberts- Austen at the 
 Royal Mint:* 
 
 1. A bar of standard gold, weighing 241 '20 ozs., which broke 
 with a slight tap from a hammer, was melted in a clay crucible, 
 and chlorine passed for three minutes. The metal was found to 
 be perfectly toughened and to have sustained a loss of weight of 
 only 0'17 oz., which proved to consist entirely of base metal. 
 
 2. Gold bars containing 0*5 per 1,000 of antimony and - 5 per 
 1,000 of arsenic, or a total of IK) per 1,000, were converted by a 
 stream of chlorine into gold of excellent quality in three and 
 .a-half to four minutes. 
 
 3. A bar of standard gold, weighing 301 ozs., in which the 
 impurities noted below were contained, was operated on. The 
 base metals consisted of: 
 
 Antimony, 
 
 Lead, . 
 
 Zinc, 
 
 Iron, 
 
 Tin, 
 
 Arsenic, 
 
 Bismuth, 
 
 3-0 per 1,000. 
 
 20 
 
 20 
 
 20 
 
 2-0 
 
 3-0 
 
 1-0 
 
 150 
 
 * First Annual Report of the Mint, 1870, p. 95. 
 
432 THE METALLURGY OP GOLD. 
 
 This gold was as brittle as loaf-sugar, and could be broken with 
 the lingers. After chlorine had been passed for eighteen 
 minutes, the metal was perfectly ductile. The loss of metal was 
 found to be 
 
 Base metals (as above), . . , 4*51 ozs. 
 
 Copper, . . . . 3-44 ,, 
 
 Gold, . . . .... '15 
 
 Total, . 8-10 
 
 These experiments having been successful, the 40,000 ozs. of 
 brittle gold were treated similarly in graphite crucibles, the 
 charges being about 1,100 ozs. of gold in each. The time of 
 passage of the chlorine varied from five to seven minutes, and the 
 total loss of gold after recovery from the " sweeps," was returned 
 as only ^ oz. With this loss and at a very slight expenditure, 
 the whole of the stock of brittle gold was toughened and rendered 
 fit for coinage. Considering the success of this venture, it is- 
 surprising that the method has not yet found wider application 
 in the treatment of brittle bars. 
 
 Parting by Electrolysis The Moebius Process. This process 
 was patented by Mr. Bernard Moebius, in England, as long ago 
 as the year 1884 (Dec. 16, No. 16,554), and is now in successful 
 operation in several localities in the United States and Germany. 
 It is said* to be specially suitable for refining copper bullion 
 containing large proportions of silver and gold with small 
 quantities of lead, platinum, and other metals, but is chiefly used 
 in parting dore silver containing not more than 20 per 1 ,000 of 
 base metals. 
 
 The apparatus required consists of a number of wooden vats 
 coated inside with graphite paint, and filled with a solution 
 containing 1 per cent, of nitric acid, which constitutes the elec- 
 trolyte. The anodes consist of plates of bullion of about J inch 
 thick, 18 inches long, and 10 inches wide, which are hung in 
 muslin bags destined to catch the insoluble impurities after the 
 silver, copper, &c., have been dissolved. The cathodes consist 
 of plates of pure silver, slightly oiled to prevent adhesion of the 
 deposited metal. These plates are continually scrubbed by a 
 mechanical arrangement of wooden brushes by which their sur- 
 faces are kept free from loose crystals of electro-deposited silver. 
 The loose silver falls on to trays placed below, which are re- 
 moved at intervals, and the silver collected from them. 
 
 The current should have an electromotive force of from one 
 to three volts for each vat. The copper is not deposited unless 
 the solution becomes too weak in silver or too rich in copper, 
 and even if some happens to be deposited, it falls into the cathode 
 trays with the silver and does little harm, since it will be 
 * Gore's Electrolytic Separation of Metals, London, 1890, p. 240. 
 
PARTING. 433 
 
 gradually redissolved if the conditions are corrected. When too 
 much copper has accumulated in the solution the latter must be 
 removed, the method being as follows : The bullion anodes are 
 replaced by carbon ones, and a weak current passed until all the 
 silver is deposited. The silver cathodes are then replaced by 
 copper ones, and a strong current passed so as to deposit the 
 copper as rapidly as possible as a loose powder, which falls into 
 a copper box. This box is connected with the cathode to prevent 
 corrosion by the acid which is set free. The liquid thus regen- 
 erated is used again as the new electrolyte. 
 
 The process is stated to be the cheapest parting process known. 
 If no copper were present in the bullion it is clear that there 
 would be no consumption of acid, and it would never be necessary 
 to change the electrolyte. Under the most favourable conditions, 
 therefore, with water power available and the amount of copper 
 in the bullion very small, the cost of parting dore silver would be 
 merely nominal. 
 
 The following description of the Moebius plant which is in 
 successful operation at the Pinos Altos Mine, Chihuahua, Mexico^ 
 is an abstract of that given by Mr. George Maynard.* The 
 plant is capable of treating from 3,500 to 4,000 ozs. of dore silver 
 in 24 hours, and consists of a tank 12 feet long, 2 feet wide and 
 20 inches deep, divided into seven compartments. Four silver 
 plates (the cathodes) and six dore plates (the anodes) are placed 
 in each compartment in such a way that an anode is opposite to 
 each face of a cathode. The trays to catch the deposited silver 
 have perforated bottoms and are covered with asbestos cloth, 
 and in order to facilitate cleaning-up, the electrodes, frames,, 
 anode bags, trays and brushes can all be lifted up simultaneously 
 by a hoisting arrangement worked with a crank and handle, so- 
 that the exciting liquid alone remains in the cells. 
 
 The bullion is from 800 to 900 fine in silver, and 25 to 50 in gold, 
 the rest being chiefly copper. The exciting liquid, which at first 
 contains 1 per cent, of nitric acid, is gradually converted into a- 
 solution of silver and copper nitrates. The silver nitrate is con- 
 tinually decomposed by deposition of the metal, but the copper 
 nitrate accumulates, fresh acid being added at intervals to prevent 
 the copper from being deposited. The silver dissolves from the 
 anodes and is precipitated on the cathodes in the form of heavy 
 tree-like crystals. The action of the brushes or scrapers is of 
 vital importance to the success of the process, the advantages, 
 derived from their action being summarised as follows : 
 
 1. The liquid is agitated and so kept homogeneous. 
 
 "2. Polarisation is prevented. 
 
 3. The electrodes can be brought very near to one another, 
 and the resistance thus reduced without any fear that short 
 circuiting v/ill ensue by the bridging over of the space between 
 them by crystals of deposited silver. 
 
 * Eng. and Mng. Journ., May 9, 1891, p. 556. 
 
 28 
 
434 THE METALLURGY OP GOLD. 
 
 The lead (as peroxide), platinum metals, antimony and other 
 impurities remain with the gold in the bag surrounding the 
 anodes. When the exciting liquid becomes too highly charged 
 with copper, the solution is used instead of bluestone in the 
 amalgamating pans of the mill, but the copper could of course be 
 recovered as usual if it were desirable. The manual labour is 
 all performed by the assayer and his assistant ; the cathode trays 
 are hinged, and by letting them turn on their hinges the silver is 
 let fall into a movable tank on castors furnished with a false 
 bottom, on which it is washed and dried, and it is then ready to 
 be melted into bars of about 999 fine. The gold slimes are simi- 
 larly washed and filtered ; when they are fused, the lead is 
 almost all slagged off. 
 
 The electric current employed is 170 amperes, the electro- 
 motive force being 8 volts ; this requires an expenditure of 
 2i H.P. to refine from 3,500 to 4,000 ozs. per day. The cost of 
 parting is said to be less than cent per gross oz. of bullion, and 
 the royalty is ^ cent per oz. 
 
 The original cost ot the plant is said to have been about $6,000, 
 including $1,000 for the silver in the cathode plates This sum 
 included the cost ot conveying the plant to the mine, which was 
 very high, as a journey lasting several weeks on mule-back had 
 to be performed. The amount of silver contained at any one 
 time in solution in the bath is about 300 ozs. The weight of 
 the forty-two anode plates is about 4,200 ozs. when they are 
 fresh, and this silver can be melted up and recovered about forty- 
 eight hours after the operation has been started, so that, in 
 this plant, only about 10,000 ozs. of bullion are necessarily 
 locked up continuously in the apparatus. The clean-up takes 
 place once a month. 
 
 The Moebius process has also been in successful operation at 
 the works of the Pennsylvania Lead Company at Pittsburg s nee 
 September, 1886 ; here it is said that from 30,000 to 40,000 ozs. 
 of dore bullion are refined daily. The dore bullion doet> not 
 contain more than 2 per cent, of impurities (lead, copper, 
 bismuth, &c.), and is cast into large anode plates, 1'3 cm. thick, 
 and weighing 13 to 15 kilogrammes, which are less advantageous 
 than the light thin anodes used at Frankfort. Each anode is 
 dissolved in about 2^ days. When the copper rises to more 
 than 4 or 5 per cent, in the solution, part of the liquid is with- 
 drawn, and the silver precipitated as chloride. The silver pro- 
 duced is from 999 to 999-5 fine, and can be raised by repeated 
 washing to 999 -85, but it is better to reduce it to 998 by adding 
 copper, as no excess over that degree of fineness is paid for in 
 the English market.* 
 
 * Mineral Industry for 1895, p. 356. 
 
PARTING. 435 
 
 At the Frankfort refinery, where this process is in use,* the 
 general arrangements are similar to those noted above. The 
 operation is begun with a very dilute solution of nitric acid and 
 with a strong current of as much as 300 amperes per square 
 metre (0*2 ampere per sq. inch). As the operation proceeds, 
 copper accumulates in the liquid (the silver anodes being only 
 about 950 fine), and reaches about 4 per cent., the silver being 
 about 0-5 per cent, and the free nitric acid from O f l to I'O per 
 cent. As the copper accumulates, fresh acid is added, and the 
 current reduced to about 200 amperes per square metre. The 
 E.M.F. required for each cell is 1*4 to 1*5 volts. The anodes are 
 from 6 to 10 mm. thick, and each weighs about 1*5 kilogrammes; 
 they are completely dissolved in about 3G hours, and another 
 plate is at once substituted, the gold slimes being cleared out 
 of the anode bag either once or twice a week. The total capacity 
 of the Frankfort refinery (3200 ozs. per vat in 24 hours), is about 
 30,000 ozs. per day, the area of the plant being only 40 feet x 
 45 feet. 
 
 At the Guggenheim Smelting Company, New Jersey, a con- 
 tinuously moving endless silver belt is used as cathode, the 
 deposited silver being brushed off by fixed brushes, so that it 
 falls into a separate tank. The capacity is 30,000 ozs. per day, 
 and the cost, exclusive of superintendence, office expenses, and 
 royalty, is stated to be about ^ cent, per oz.f 
 
 Electrolytic refining presents many advantages over acid 
 parting, among them being the absence of noxious fumes, except 
 during acid treatment of the gold residues, the small amount of 
 labour and chemicals, the cleanliness and rapidity of working, 
 and the absence of bye-products, so that the loss of metal is 
 reduced to a minimum. In consequence of these advantages, 
 dor6 silver is being more and more treated by this method. 
 
 Parting of Gold from Platinum, Iridium, &c. The 
 method of separating osmiridium, platiniridium, &c., from gold 
 in the United States Mints is described on p. 396. J 
 
 A. Baker & Co., platinum manufacturers of New Jersey, 
 separate platinum from gold by dissolving in aqua regia, driving 
 off acids and precipitating the platinum with ammonium chloride. 
 At Freiberg, platiniferous gold is purified by melting with twice 
 its weight of sodium bisulphate, and subsequently with a little 
 nitre. The slags contain most of the platinum, the resulting 
 gold being 997 to 998 fine. || 
 
 The best and most modern method, however, is an electrolytic 
 
 *Borcher's Klectric Smeltiny and Refining. Translated by W. G. 
 McMillan, London, 1897, p. 243, et seq. 
 ^Mineral Industry for 1895, p. 358. 
 % Report of the Director of the U.S. Mint, 1885. 
 Mineral Industry for 1895, p. 359. 
 II Ibid., p. 361. 
 
436 THE METALLURGY OP GOLD. 
 
 one, and is used at the Norddeutsche Affinerie of Hamburg.*' 
 It may be described as follows : The anode consists of impure 
 gold, 4 mm. thick, and the electrolyte is a solution of gold 
 chloride, containing 25 to 30 grammes gold per litre, and irom 
 20 to 50 c.c. of concentrated hydrochloric acid per litre, according 
 to the strength of the current. The liquid is heated to 60 to- 
 70 C. If the temperature is allowed to fall, or the acid is 
 omitted, chlorine escapes. The cathode consists of pure sheet 
 gold, and the current density is from 400 to 500 amperes per 
 square metre. The gold, platinum, and palladium are dissolved, 
 and the residue consists of iridium, silver chloride, and some 
 finely divided gold derived from the decomposition of AuCl. 
 Nearly pure gold is deposited in an even coherent form on the 
 cathode, and the solution becomes weaker in gold, so that gold 
 chloride must be added from time to time. The operation is. 
 carried on continuously until the amount of platinum in solu- 
 tion is twice that of the gold (or the palladium in solution 
 exceeds 0*5 per cent.), when it is precipitated with ammonium 
 chloride. The anode is reduced to one-tenth of its thickness in 
 24 hours, and is then melted down and recast. The electrodes 
 are 3 cm. apart. Silver, lead, and bismuth are the cause of 
 difficulties necessitating the brushing of the anodes, and the 
 addition of sulphuric acid. 
 
 This method uses very little acid, and offers a cheap way of 
 preparing proof gold. 
 
 * Report on the Precious Metals of the U. 8. , 1896. Quoted by Titus Ulke 
 in the Mineral Industry for 1896, p. 304. 
 
THE ASSAY OF GOLD ORES. 437 
 
 CHAPTER XIX. 
 THE ASSAY OF GOLD ORES. 
 
 THE assay of gold ores is almost universally conducted in the 
 dry way i.e., by furnace methods. Exceptions will be noted 
 later. The plan of operation is to concentrate the precious metal 
 in a button of lead in one of two ways, viz. : (1) By fusion in a 
 crucible; or, more rarely, (2) By scorification. The button of lead 
 obtained by either method is then subjected to cupellation, by 
 which the lead is oxidised and removed, and the resulting bead 
 of precious metal is weighed. Since in these operations silver 
 and the metals of the platinum group remain with the gold, 
 they are subsequently separated by inquartation and parting, and 
 in the case of platinum and its allies by further special methods. 
 
 The exact method to be used in any particular case varies 
 "with the richness of the ore, the nature of its gangue and the 
 presence or absence of compounds of the base metals. As a 
 general rule, poor ores i.e., those containing less than '2 ozs. gold 
 per ton are better assayed by the fusion process so that a com- 
 paratively large quantity of material may be operated on. Rich 
 ores may be assayed either by fusion or scorification, the errors 
 arising from the small amount of material used in the latter 
 process being less important in their case. It is more con- 
 venient to scorify tellurule ores, arsenical and antimonial ores, 
 and ores containing tin, nickel, or cobalt, but in these cases it is 
 often necessary to make the ordinary crucible assay, and then 
 to scorify the lead button obtained. Either method can, how- 
 ever, be used for any ore. 
 
 Assay by means of the Blowpipe* This method, though less 
 exact than that made in a furnace, is of importance, because in 
 prospecting expeditions it is possible by its means not only to 
 detect the gold and silver in any ore, but also to determine its 
 amount quantitatively with fair accuracy. On such expeditions 
 it is impossible to carry the cumbrous apparatus required to 
 make an ordinary assay. The amount of powdered ore taken is 
 usually 100 milligrammes, and this is mixed with borax and 
 about 1 gramme of granulated lead. The whole is wrapped in 
 paper and heated on charcoal in the reducing name of a blow- 
 pipe until the fusion is complete, and then for a short time with 
 the oxidising flame. The lead is then separated from the slag 
 
 * For a full description of this method, see Plattner's Manual of 
 Analysis with the Blowpipe, pp. 360-406. London, 1875. 
 
438 THE METALLURGY OP GOLD. 
 
 and heated on a bone-ash cupel until it is all converted into- 
 litharge. The diameter of the button of silver and gold thus 
 obtained is carefully measured on an ivory scale, which at once 
 gives the percentage amount in the ore. The gold is usually 
 separated from the silver by parting in nitric acid, but Richards 
 has recently stated* that the silver can be distilled off by the- 
 blowpipe, leaving a bead of pure gold, which can be measured. 
 
 Sampling the Ore and Preparation of the Sample for 
 Assay. The value of an assay depends largely on the care with 
 which a sample of the ore is selected. The sampling is done by 
 processes which are as far as possible automatic, and independent 
 of the will or judgment of the assay er. Clarkson's sampling 
 machine is excellent, and gives accurate results. A good method 
 is that of dividing and intercepting a part of a stream of ore 
 falling vertically downwards or sliding down an inclined plane. 
 This is done in many ways in mines and mills, one of which may 
 be described as follows : The mineral, having been passed through. 
 a stone-crusher and one or two pairs of rolls so as to be reduced 
 to the size of coffee beans, falls through a vertical tube, on to the 
 apex of a cone. The surface of the cone is inclined at an angle 
 of 45 to the vertical and has one or more openings or windows 
 bounded by straight lines drawn from the apex to the base. The 
 ore, falling on the cone, slides over its surface into the ore-bin, 
 excepting a portion, usually one-sixteenth, which passes through 
 the openings in the cone and is conveyed down a tube to another 
 receptacle. The operation may be repeated on the sample thus 
 taken, after it has been further crushed. 
 
 In the assay office the sample, however obtained, is further 
 reduced in bulk by the implement known as the sampling tin. 
 This consists of a series of troughs arranged side by side and 
 fastened at equal distances from each other (the width of the 
 spaces being equal to that of the troughs) by strips of metal 
 soldered on to their ends. An even stream of ore being let fall 
 from a shovel on to this sampler, half is retained in the troughs 
 while half passes through. Careful experiments have proved 
 that each half is representative of the whole. A repetition of 
 the process reduces the sample to any required extent. A 
 sampler with troughs 1 inch wide is suitable for treating 
 materials which include lumps of not more than J to J inch in 
 diameter. For finely ground materials a convenient width for 
 the troughs is f inch. 
 
 In taking a sample from a large bulk of ore, or where 
 sampling tins cannot be procured, the method of dividing into- 
 quarters may be employed. The ore is piled into a heap, 
 thoroughly mixed and divided into four by a shovel by two- 
 lines at right angles to each other. The two opposite quarters- 
 are removed, and the remainder again mixed and divided as- 
 
 * Journal of the Franklin Institute, June, 1896. 
 
THE ASSAY OF GOLD ORES. 439 
 
 before. When the sample is reduced to a few pounds in weight 
 it is broken down to the size of coffee beans, reduced in bulk by 
 quartering or the sampling tin, crushed finer, again sampled and 
 so on, until finally, when reduced to a mass of from J Ib. to 1 lb., 
 it is all crushed fine enough to pass through a 60-mesh or 
 80-inesh sieve (i.e., one containing 80 holes per linear inch). 
 Before this fine crushing can be done, it is usually necessary to 
 dry the sample ; the percentage of moisture should always be 
 estimated at the same time by weighing the sample both before 
 and after it is dried. 
 
 The implements employed for crushing samples of ore are 
 various. A small rock-breaker, with reciprocating jaws, similar 
 to those used on the large scale, and worked by hand or steam 
 power, is useful for breaking down large lumps, which otherwise 
 may be broken by a hammer. For finer pulverisation a small 
 pair of steel-faced high-speed rolls may be used if steam power is 
 available. This method is adopted at large smelting or sampling 
 works, where great numbers of samples are crushed daily. In 
 smaller works or offices the buckboard (Fig. 59) is most suitable. 
 
 Fig. 59. 
 Scale, | in. = 1 ft. 
 
 It is a smooth plate of iron about 2 feet square with a 1-inch 
 rim surrounding it on two or three sides. On this a bucking 
 hammer is worked a heavy piece of iron 5 to 10 Ibs. in weight, 
 with a large smooth curved face and a handle 30 inches long. 
 It is moved about on the iron plate (on which the ore is spread) 
 with both hands, one holding the handle, the other pressing the 
 head downwards, the curved face being below, while an oscillatory 
 movement is imparted by the handle. The instrument is very 
 effective if the ore is previously broken down to the size of 
 coarse sand in a mortar. The pestle and mortar are of value in 
 breaking down samples from the size of nuts to that of coarse 
 sand. In grinding down siliceous material, so as to enable it to 
 pass through an 80-mesh sieve, the pestle and mortar is far 
 inferior to the buckboard. 
 
 " Meta/lics" In many ores, both gold and auriferous silver 
 occur native in grains or threads. These " metallics " are not- 
 readily reducible to a fine state of division, and, though a part 
 always passes through the sieve, some of the larger pieces which 
 have resisted abrasion fail to do so. In some assay offices part 
 of the pulverised ore is thrown back into the mortar with the 
 
440 THE METALLURGY OF GOLD. 
 
 inetallics, and grinding is continued until everything passes the 
 sieve. This is a dangerous practice, as it is impossible to ensure 
 the even distribution of metallics through the sample. The 
 smaller pieces which pass through the sieve in the first instance 
 constitute an unavoidable evil which is increased by every piece 
 of metal that follows them. The safer plan is to cupel by 
 themselves the whole of the metallics left on the sieve, and 
 calculate their value per ton of ore independently of the result 
 obtained by the ordinary assay. The total value of the ore is 
 found by adding these two results together. 
 
 The prepared sample may be stored in tin boxes or glass jars, 
 which should be labelled by numbers, none of which are ever 
 repeated. Before weighing out the powdered ore for assay, the 
 whole sample should be turned out into a wide bowl or on to 
 glazed paper, or, better still, rubber cloth (which does not crack 
 and wear out like paper), and thoroughly mixed with a spatula. 
 The sample should never be mixed by shaking, and care should 
 be taken to avoid jarring it after mixing, as metallics in that 
 case tend to settle to the bottom from their superior density, and 
 a fair sample cannot then be easily obtained. For the same 
 reason the part of the mixed sample which is taken for assay 
 should not be hastily shovelled on to the balance pan from the 
 top of the pile, but a vertical slice should be taken, some of the 
 lowest layer being carefully scraped up from the rubber cloth. 
 
 FUSION OR CRUCIBLE PROCESS OF ASSAY. 
 
 This process is divided into three parts, viz. : (1) Fusion; (2) 
 Cupellation; (3) Parting. 
 
 (1) Fusion. The object of this operation is to concentrate 
 the precious metals in a button of lead, while the whole of the 
 remainder of the ore forms a fusible slag with suitable reagents, 
 in which lead sinks. The fusion is made in clay (or rarely iron) 
 crucibles in a wind furnace, the fuel being coke, charcoal or 
 anthracite. The size of the furnace depends on the amount of 
 work to be done and on the fuel employed, charcoal requiring 
 more space than coke. If only one fusion is to be made at once, 
 the internal dimensions of the fire-box may be 8 inches square, 
 while a furnace 14 inches square will hold nine crucibles. The 
 furnace is exactly similar to that used for melting bullion 
 described on p. 387. 
 
 In the United States it is usual to perform fusions in a muffle 
 furnace, similar to that described below under cupellation, p. 480. 
 The temperature required is about the same as that used in 
 scorification. The advantages claimed are greater cleanliness 
 and neatness and more uniformity in the conditions, the tem- 
 perature of a muffle being more easily kept constant and uniform 
 
FUSION OR CRUCIBLE PROCESS OP ASSAY. 441 
 
 than that of an ordinary fusion furnace. Six or eight fusions 
 can be performed at one time in a large muffle. 
 
 There are three shapes of clay crucibles in common use, namely 
 the French, the Battersea, and the Colorado. The French pot 
 differs from the others in the thickness of its walls near the 
 bottom, and consequently it requires very careful annealing. 
 The Battersea pot is of coarser texture than the others, having a 
 larger percentage of sand in its composition. The Colorado 
 crucibles, manufactured by the Battersea Company, are specially 
 made for fusion in the muffle; the "20 gram" size is the most 
 generally useful. 
 
 In weighing the materials for a crucible charge the use of a 
 set of assay-ton weights saves much labour in calculation. The 
 .assay-ton is a weight which contains as many milligrammes as a 
 ton contains ounces. Thus an English or long ton of 2,240 Ibs. 
 {used in England and Australia) contains 32,666 Troy ounces, 
 so that the corresponding assay-ton must weigh .*J,666 m.m.g or 
 32-666 grammes. If the weight of the resulting bead of gold 
 (or silver) from an assay -ton of ore is 1-5 milligrammes, then 
 the ore contains 1'5 ozs. of gold per statute ton. If the value 
 per short ton of 2,000 Ibs. (used in North America and Africa) 
 is required, the weight of the assay-ton is 29*166 grammes, since 
 there are 29,166 Troy ounces in 2,000 Ibs. avoirdupois. If 
 grain weights are preferred, the weight of the assay-ton may 
 conveniently be taken as 326 '66 grains (or 291-66 grains for the 
 short ton); the weight of the resulting bead of gold in hundredths 
 of a grain then gives the value of the ore in ounces per ton. 
 Sets of weights ranging from ^ A.T. (assay-ton) to 4 A.T. can be 
 bought, or they can be made up from an ordinary box of decimal 
 weights. In the following pages this system is used, 1 A.T. 
 being equal to 32-666 grammes : the grammes or grains of the 
 various authors referred to are converted into the approximately 
 corresponding number of assay-tons. 
 
 General Charges. The charge for a fusion assay varies accord- 
 ing to the nature of the ore. The following proportions are 
 supposed by their respective authors to be suitable to all gold 
 and silver ores, and will serve for a rough preliminary assay of 
 .an unknown ore : 
 
 MitcheWs Formula* 
 
 Ore, . . 1A.T. 
 
 Soda carbonate, 
 Litharge, ; , 
 Borax glass, 
 Salt to cover, 
 
 and the amount of argol or of nitre necessary to ensure the 
 formation of a button of lead weighing about 200 grains (13 
 grms.) These quantities can only be determined by a preliminary 
 .assay. 
 
 * Mitchell's Manual of Assaying t 1881, p. 546. 
 
442 
 
 THE METALLURGY OF GOLD. 
 
 Aaron's Formula* 
 
 Ore, . 
 
 Soda carbonate, 
 
 Litharge, 
 
 Borax, . 
 
 Sulphur, 
 
 Flour, . 
 
 Iron, 
 
 Glass. 
 
 Salt to cover. 
 
 1 A.T. 
 3 
 
 1 
 
 A 
 
 & 
 3 nails. 
 
 "Melt and leave in a hot fire about twenty minutes after fusion."" 
 When ores contain only small quantities of base metals the- 
 following formula is recommended by Brown & Griffiths f 
 
 Ore, 
 
 Soda carbonate, . 
 
 Litharge, . 
 
 Borax glass, 
 
 Carbonate of potash, 
 
 Silica, 
 
 Charcoal, . 
 
 Salt to cover. 
 
 1 A.T. 
 H 
 
 H 
 
 I 
 
 *6 gramme. 
 
 Percy's Formula for ores containing only small proportions of base 
 metals 
 
 Ore, 
 
 Red lead, 
 
 Soda carbonate and borax, 
 
 Charcoal, 
 
 i to 1 A.T. 
 
 1 
 
 1 ,, together. 
 13 to 17 grains. 
 (0'8 to 1-1 gramme. > 
 
 The charges given above are varied according to circumstances. 
 One A.T. of the ore is a suitable quantity if the value in gold is 
 from -5 oz. to 10 ozs. or more. With very pror ores 2, 3, or 4 
 A.T. may be taken, and with very rich ores J A.T. may suffice. 
 The fluxes are altered in proportion. Formerly the usual 
 practice of assayers was not to reduce all the red lead or 
 litharge; part was left in the slag, forming easily fusible silicates. 
 Aaron was the first to describe the method of reducing tha 
 whole of the litharge employed; the gangue is in this case slagged 
 off by soda carbonate, borax or silica, and the base metals 
 separated as a matte. The advantages claimed for this method 
 are that the button is never much contaminated by copper, and 
 that the crucible is but little attacked. Brown & Griffiths 
 observe that beginners find Aaron's method less easy than the 
 other ; but it is now perhaps more usual to reduce all the lead 
 than to leave some in the slag, except in rare cases, such as, for 
 instance, when much copper is present. 
 
 Fluxes. The following remarks on the use of the fluxes may 
 serve as an aid in making up charges in most cases : Litharge or 
 
 * Aaron's A ssayincj, 1884, p. 53. 
 
 t Brown and Griffiths' Manual of Assaying, p. 174. 
 
 J Percy's Metallurgy of Silver and Gold, p. 245. 
 
FUSION OR CRUCIBLE PROCESS OF ASSAY. 443 
 
 red lead is added in the proportion of one or two parts to two 
 of ore ; if too much litharge is used the slags are not clean, as a. 
 slag containing lead may mean a loss of silver and gold. What- 
 ever method is used, the amount of lead to be reduced should be 
 from 25 to oO grammes. Raw ores or regulus containing much 
 sulphide of copper may be fused with from 4 to 6 A.T. of litharge 
 to 1 A.T. of ore. In this case the other fluxes, except sand, may 
 be omitted. Some assayers prefer to concentrate the copper as 
 a regulus, and then to treat this over again ; little more than 
 the ordinary quantity of litharge is then used, and not much iron. 
 
 The amount of the charcoal added varies with the reducing 
 power (percentage of ash, &c.) of the particular sample which is 
 employed, as well as the degree of oxidation of the ore. In some 
 highly basic oxidised ores as much as 3 grammes of charcoal 
 powder is required for 1 A.T, of ore, as the oxide must always 
 be completely reduced to FeO, but generally from 1 to 1J 
 grammes is enough. If there is much Fe. 2 O 3 in the ore (e.g., 
 roasted pyrites) the slag is often rich, 3 or 4 per cent, of the 
 gold being lost.* The remedy usually adopted is to increase 
 largely the quantity of soda carbonate, while sand must also be 
 added to prevent the crucible from being perforated by the 
 scouring action of the ferrous oxide. If ores contain much 
 sulphur no charcoal is used, and nitre may even be added to- 
 bum off the excess of sulphur, otherwise, if the old practice of add- 
 ing much litharge is followed, the amount of lead reduced may 
 become too great. The addition of nitre is, however, now seldom 
 made, as the pot is liable to boil over; with very large quantities of 
 sulphides it is better to make a matte and treat the latter again. 
 
 Carbonate of soda is used to flux silica, while borax is valuable 
 in basic ores to prevent corrosion of the crucible, and render the 
 slag more liquid. The relative amounts required are judged from 
 the appearance of the ore in the first place, and afterwards 
 modified acording to the success of the fusion. From 1 to \\ A.T. 
 of soda carbonate, and from J to J A.T. of borax to 1 A.T. of ore 
 are the amounts usually required. Even when the ore is 
 entirely composed of silica, some borax is added. The most 
 convenient form is borax glass. 
 
 Silica is used only for ores full of lime, baryta, compounds 
 of the base metals, &c., or generally whenever the ore does 
 not contain much quartz. It aids fusion in these cases, and 
 protects the crucible from corrosion. From J to J A.T. to 1 A.T. 
 of ore is generally enough. Fluorspar is added to the charge 
 when the ore contains sulphates of barium or calcium, and in 
 the fusion of cupels. Like borax, it increases the fluidity of 
 almost any charge, but it attacks the crucible, and care must be 
 taken to avoid deficiency of silica when it is used. In general, 
 it may be noted that for basic impurities an acid flux is used, 
 and lor an acid gangue a basic flux. 
 
 * C. & J. J. Beringer's Assaying, p. 125. 
 
444 
 
 THE METALLURGY OF GOLD. 
 
 The ore is weighed accurately to '01 gramme unless very poor 
 -when a less exact approximation is sufficient. The charcoal is 
 always weighed with great care ; the litharge is best measured 
 by a ladle or shot measurer ; the fluxes may also be measured 
 out by a ladle. By practice the assayer becomes very rapid in 
 measuring out reagents. The various ingredients are thoroughly 
 mixed together on rubber cloth or in the crucible in which the 
 fusion is to take place. Part of the borax is kept separate and 
 used as a cover, being put on the top of the rest of the charge 
 after transference to the crucible. A cover of common salt is 
 used by German and American assay ers ; for its former em^loy- 
 ment at the Royal College of Science see Dr. Ball's remarks on 
 p. 446. Iron is added to the charge in the shape of large nails 
 or hoop-iron, or even scrap, if sulphur or arsenic is present. 
 Sulphur is thus kept out of the lead button. 
 
 The crucible is carefully annealed in the ashpit of the furnace 
 before using. It is lowered into a hollow in the fuel of the 
 furnace made by piling the coke round an old pot and then care- 
 fully withdrawing the latter. The tongs 
 most used are shown at B, Fig. 60, A 
 being those used when the fusion is per- 
 formed in a muffle. The fire should be at 
 a good red heat on charging-in, and care 
 must be taken that no coke in the furnace 
 is black. If the upper layer of coke is 
 much colder than that below, the top of 
 the charge in the crucible remains un- 
 melted while the bottom begins to effer- 
 vesce, and the crucible may froth over 
 and part of its contents be lost. As the 
 charge melts the fire is urged, and brisk 
 effervescence occurs chiefly from the escape 
 of carbonic acid from the soda carbonate 
 as it unites with silica. After a lapse of 
 from twenty to forty minutes the charge 
 is in a state of tranquil fusion, with the 
 exception perhaps of slight action round 
 the sides or next the iron, if any is used. The crucible is then 
 lifted out of the fire by the tongs, the nails withdrawn and any 
 lead adhering to them shaken off into the pot. The pot may 
 now be tapped on the floor to assist the lead to settle in the 
 slag, allowed to cool and broken by a hammer to extract the 
 lead button, or the charge may be at once poured into an iron 
 mould. The mould must be cleaned, blackleaded and warmed 
 before being used, and, after pouring, it is tapped on the bench 
 to collect the lead at the bottom. 
 
 If the charge does not fuse ompletely, so that the slag is pasty 
 or has lumps in it, it is advisable to recommence the assay, 
 
 Fig. 60. 
 
FUSION OB CRUCIBLE PROCESS OF ASSAY. 445- 
 
 making such alterations in the charge as experience suggests,, 
 having regard to the considerations mentioned above, p. 443. 
 
 When the mould is quite cold its contents are readily separated 
 from it if the precautions mentioned above are taken, and the 
 slag is detached from the lead button with a hammer on an 
 anvil. The slag should be glassy and homogeneous ; if it is- 
 streaked it is probable that the fusion has not been perfect. It 
 is green and transparent if the ore is nearly pure quartz ; black 
 and opaque if much iron is present; but is red if the ore contains^ 
 much copper. 
 
 The lead button should be soft and malleable. If it is hard 
 or brittle it may contain sulphur, arsenic, antimony, copper, &c. 
 Sulphur and arsenic are kept from entering the lead by the 
 addition of iron, and then form with the iron a regulus or speiss, 
 which separates as hard blackish-grey or white layers found 
 just above the lead. They are richer than the slag, and may 
 often yield appreciable quantities of gold on further treatment 
 by scorification or roasting and fusion. Antimony makes the 
 lead hard, white, brittle and sonorous ; it can be removed by 
 adding nitre to the fusion charge, but, according to Rivot,* it is not 
 a source of loss if forming less than 1 per cent, of the lead button 
 (see under Cupellation below). The lead must be completely 
 freed from the slag, very small quantities of the latter interfering 
 seriously with the cupellation, forming a scoria and occasioning 
 loss. 
 
 On the subject of fusing gold ores, Dr. E. J. Ball, who was 
 formerly the Instructor in Assaying at the Eoyal School of 
 Mines, London, wrote as follows in 1893 : 
 
 " I find a very good plan in assaying a gold (or a silver) ore to- 
 be as follows, noting the points : 
 
 "(1) That the main aim at the beginning of the assay is to 
 produce an intimate contact between the molten reduced lead 
 and the gold freed by the original grinding of the ore. 
 
 "(2) That when this has been effected, the particles of ground- 
 up ore must be, as it were, 'atomically' ground-up by solution, in 
 order to liberate enclosed particles of gold. 
 
 "The first of these stages should not be accompanied by a 
 fusion of the charge, because immediately that happens the lead 
 sinks to the bottom of the crucible, and is at once removed from 
 the possibility of contact with the gold floating about in the 
 charge. To bring about this contact in the second stage, the 
 bath must be as thin-fluid as possible, and convection currents 
 should be induced by irregular heating of the crucible, in the 
 hope that, in the course of one of their gyrations, the particles of" 
 gold liberated by the solution of the quartz particles may strike 
 against the surface of the molten lead at the bottom of the- 
 crucible. 
 
 " The maximum quantity of lead which the ordinary cupels in-. 
 * TraiM de Docimasie. 
 
446 THE METALLURGY OF GOLD. 
 
 use at the Royal School of Mines will take is about 450 grains. 
 I therefore recommend (when treating fairly pure quarts, con- 
 _ taining say 1 oz. of gold per ton) that the charge should be made 
 up as follows : 
 
 Ore (passed through an SO-sieve), . 1,000 grains. 
 Red lead, . . . 500 
 
 Charcoal, . -.- . . 25-30 
 
 The charcoal and red lead are first mixed together, and the ore 
 is then carefully incorporated with them. Then 250 grains of 
 sodium carbonate are roughly stirred in, so as to prevent the 
 formation of a sort of sand-bottom, which would not dissolve in 
 stage 2. 
 
 " The charge is then maintained at as high a temperature as 
 possible, actual fusion being avoided for fifteen minutes, when 
 another 1,000 or 1,200 grains of sodium carbonate are charged-in 
 little by little, and the temperature raised to about 950. At 
 this stage, anv necessary addition may be made in order to make 
 the bath fluid, borax for instance, but borax seems to give low- 
 results if added at the beginning of the assay. I always add a 
 piece of hoop iron to help to decompose any lead silicate or 
 sulphide. 
 
 " I never recommend the use of salt, but sometimes when a 
 large excess of sodium carbonate has been added, the 'boil' at 
 the end seems never likely to stop, owing to the action of the 
 acid crucible on the basic charge. In that case, a little salt 
 stirred into the bath seems to volatilise, prevent the contact of 
 the charge with the walls of the crucible for a moment or two, 
 and to quiet the bath, enabling the charge to be poured 
 properly." 
 
 It is worthy of note that Dr. Ball reduces practically all the 
 lead oxide, leaving none in the slag. 
 
 Mr. H. C. Jenkins, the present Instructor in Assaying at the 
 Royal School of Mines, while retaining most of the methods 
 mentioned above, in teaching his students, recommends a larger 
 addition of lead oxide, so that some is left in the slag. He 
 never adds less than 40 grammes ot red lead* to the charge, 
 and aims at reducing 1rom 28 to 30 grammes of lead. He 
 recommends roasting highly pyritic ores (see below) and the use 
 of nitre for aritimonial ores onJy. He also advises his students 
 to cover the crucible after fusion is complete. As giving ex- 
 
 *Mr. Jenkins considers litharge (PbO) to be less suitable than red lead 
 (Pb 3 04). The latter requires more charcoal to reduce it and, accordiug to 
 Prof. Ricketts of the School of Mines. Columbia University, it is objection- 
 able on the grounds that it oxidises silver. (Notes on Assaying, by Prof. 
 Ricketts, p. 39, New York, 1897.) This drawback however, appears to be 
 shared by litharge. In some comparative experiments made by the author, 
 there was no difference in results, so that it would seem that either litharge 
 or red lead can be used with equal advantage. 
 
FUSION OR CRUCIBLE PROCESS OF ASSAY. 
 
 447 
 
 amples of the charges which he employs, the following may be 
 taken as typical : 
 
 
 Quartzose Ore. 
 
 Pyritic Ore. 
 
 Pyrites. 
 
 
 
 ( 
 
 50 grammes. 
 
 Ore, 
 
 50 grammes 
 
 50 grammes < 
 
 (Weighed, then 
 
 
 
 ( 
 
 roasted.) 
 
 Red lead, 
 
 40 
 
 40 ,, and upwards 
 
 40 grammes. 
 
 Charcoal, 
 
 1-5 
 
 
 3 
 
 Soda carbonate, 
 
 50 
 
 20 to 50 grammes 
 
 20 
 
 Borax, 
 
 ... 
 
 10 
 
 10 
 
 Sand, 
 
 ... 
 
 J A little, when ore \ 
 \ charge is basic J 
 
 15 
 
 With hoop iron ad libitum. 
 
 Boasting before Fusion. Ores containing large quantities of 
 sulphur, arsenic or antimony may often be roasted with advan- 
 tage as a preliminary to fusion. Roasting is effected in shallow 
 circular clay dishes, in a muffle, or in the crucibles in which 
 the fusion is afterwards performed. The temperature must be 
 kept low at first and the ore frequently stirred with an iron wire 
 or spatula, to prevent fritting, and to expose fresh surfaces to 
 the air. The roasting takes place in two stages : at first, sulphur 
 dioxide, arsenious oxide (As 2 O 3 ) and antimonious oxide (Sb 2 O 3 ) 
 -are formed and volatilised, the sulphur burning with a blue flame. 
 The formation of lumps is most to be feared during the first few 
 minutes of the operation, and can scarcely be prevented if much 
 sulphide of antimony is present ; in this case an equal weight of 
 pure silver sand is mixed with the crushed ore before charging it 
 into the muffle. 
 
 After a time the blue flame disappears, the odour becomes less 
 strong, and sulphates, arseniates and antimoniates form. By 
 raising the temperature sulphates are decomposed, but arseniates 
 and antimoniates are stable at high temperatures and cause loss 
 of silver in the fusion. To prevent their formation the ore 
 should be roasted in a coke furnace, starting to heat it very 
 gradually and admitting a limited supply of air. In all cases the 
 roasting is nearly complete when the glow caused by stirring is 
 shown only by a few specks of ore; the temperature may then be 
 raised to a strong red heat without danger of fusion. The opera- 
 tion is complete when the ore remains of a uniform colour on 
 stirring. The fusion of roasted ores requires less lead than raw 
 ores and more charcoal powder, the amount of the latter needed 
 being sometimes as much as 3 grammes per A.T. of ore. In 
 general, roasting is to be deprecated, owing to the unavoidable 
 loss by dusting. The results are usually lower than those 
 obtained by fusion with the formation of mattes. 
 
448 THE METALLURGY OP GOLD. 
 
 Cleaning the Slag. The slags of very rich ores may retain 
 enough gold to necessitate further treatment. The slag is roughly 
 crushed and fused with 300 to 500 grains of litharge, 15 to 20' 
 grains of charcoal and a little carbonate of soda, the same crucible 
 being used again. If any regulus or speiss forms during the 
 first fusion it must be preserved with special care and re-fused, 
 charcoal being omitted. The button of lead from the second 
 fusion is cupelled with the first button, or alone ; the slag is 
 almost always poor enough to be thrown away. It is not 
 usually necessary to clean the slag of ores containing less than 
 5 ozs. gold per ton. 
 
 The Treatment of Base Ores. The difficulty of roasting 
 arsenical and antimonial ores may be avoided * by taking ore, 1 
 A.T. ; red lead, 1,000 grains; sodic carbonate, 500 grains; potassic 
 ferrocyanide, 550 grains, with a cover of salt or borax. The 
 button is scorified, together with the matte formed, before cupella- 
 tion. It is perhaps better to scorify these ores as well as those 
 containing much copper if they are rich enough. If poor, copper 
 ores m;ty be treated in three ways,f so that each method may 
 serve as a check on the others, viz. : (a) Fusion with much PbOr 
 the lead button becomes cupriferous, and should be scorified 
 together with the matte ; (6) roasting, followed by fusion and 
 scorification ; (c) treatment with nitric acid, by which all the 
 sulphur and copper are removed. The silver dissolved in the 
 liquid is then precipitated by a solution of common salt, of 
 which a large excess should be avoided. The insoluble residue- 
 is dried, and can now be readily fused and cupelled. By 
 treatment (c) the lead button is kept free from copper, the 
 presence of which in the lead obtained by methods (a) and (6) 
 renders cupellation difficult and unsatisfactory. As already 
 observed, some assay ers prefer to concentrate the copper as a 
 regulus. The regulus first obtained still retains gold, and must 
 be re-treated. 
 
 Brown and Griffiths describe J the following general methods 
 for the treatment of ores consisting mainly of sulphides of copper 
 or iron or both : (1) Roasting followed by fusion with fluxes, 
 the proportions being somewhat similar to those given above by 
 Jenkins. The amounts of silica and of charcoal to be used vary 
 with the colour of the roasted ore, both being increased as the 
 oxides increase, and consequently as the colour becomes a darker 
 red, or if only copper and no iron is present, a deeper black. 
 (2) Oxidation by means of nitre during fusion. (3) Desulphur- 
 isation by means of iron nails, six tenpenny nails being sufficient 
 for 1 A.T. of ore. 
 
 Cupellation. This operation is conducted in a muffle furnace, 
 the construction of which is shown in Figs. 62-64, p. 467. The 
 
 * Rickett's Notes on Assaying, New York, 1887, p. 77. 
 
 t Percy's Metallurgy of Gold and Silver, p. 247. 
 
 I Manual of Assaying, by Brown and Griffiths. London, 1890, pp. 176-187- 
 
FUSION OR CRUCIBLE PROCESS OP ASSAY. 449 
 
 fire is lighted, a little bone-ash is sprinkled on the floor of the 
 muffle to prevent its corrosion by litharge in case of the upsetting 
 of a cupel, and the cupels introduced as soon as a bright red 
 heat is attained. Cupels are little cups made of bone-ash, and 
 are either round or square. In their manufacture the bone-ash is 
 finely powdered so that it will pass a 40-mesh sieve, then slightly 
 moistened with water (to which a little carbonate of potash is 
 sometimes added), put into a mould (Fig. 65) and compressed 
 by the blows of a mallet so as to cohere firmly. Cupels must be 
 dried very carefully and slowly but completely, as otherwise 
 cracks appear when they are heated, and loss is thereby occa- 
 sioned. The cupels at the Royal Mint are made two years 
 before being used, and are dried slowly on shelves at some 
 distance from the furnaces. If too much water is used in 
 mixing the bone-ash, the cupels lose part of their porosity : if 
 too little is used they are too soft and crumble readily. About 
 1 oz. water to each troy pound of bone-ash answers very well. 
 
 The cupels having attained the same temperature as the 
 muffle, the lead buttons obtained as described on p. 445 are 
 charged in by the tongs (A, Fig. 61). The buttons collapse 
 
 r*. . . . ? 
 
 Fig. 61. 
 
 and lose their shape almost instantly if the temperature is 
 sufficiently high, but the molten mass formed is covered by a 
 black crust for several seconds later. The crust then breaks up, 
 and the brilliant surface of a liquid bath is seen. The muffle 
 door should be closed immediately after the charging-in i& 
 completed, and kept closed until all the assays have thus 
 " uncovered." The door is then opened a little way to let in a 
 current of air by which the lead is oxidised, and the litharge, 
 floating to the edge of the bath, is absorbed by the cupel, 
 together with the oxides of other base metals, which are not 
 taken up by bone-ash if they are in a state of purity. TJncom- 
 bined metals are not readily absorbed by the cupel, and only 
 traces of gold and silver are carried into it by the litharge. 
 These quantities depend partly on the nature of the cupels, 
 
450 THE METALLURGY OF GOLD. 
 
 coarse-grained bone-ash absorbing more of the precious metals 
 than fine-grained. By a rise in temperature the absorption of 
 gold is increased. The presence ot other base metals also has 
 some influence on the amounts of precious metals absorbed, oxide 
 of copper in particular carrying with it much gold into the cupel. 
 As the cupellation advances the lead bath is reduced in size 
 by oxidation and absorption ; reddish patches float slowly over 
 its surface, appearing earlier in the richer assays ; it becomes 
 more convex and brighter ; the red spots move more quickly 
 and finally whirl round with great speed and then disappear ; 
 moving iridescent bands take their place for a moment and then 
 disappear likewise, and the bead becomes suddenly mucli duller 
 in appearance, thus indicating that the cupellation is at an end. 
 The temperature must be raised towards the end of the operation 
 to remove the last traces of lead, and the beads left for from 
 three to ten minutes, after all apparent oxidation is at an 
 end. The cupels may then be removed from the muffle, provided 
 the ore is poor or has little silver in it. It must, however, be 
 remembered that silver absorbs oxygen when molten and gives 
 it off suddenly when solidifying, so that if the bead weighs more 
 than O'Ol gramme (Rivot) little fountains of metal are thrown 
 up and some part may be projected out of the cupel. This 
 " sprouting," " spitting," or " vegetation " may take place in 
 argentiferous-gold beads if the gold does not exceed one-third of 
 the silver (Levol). At the Royal Mint it is found that a still 
 larger percentage of gold does not prevent spitting unless a trace 
 of copper is present. Where spitting is to be feared, therefore, 
 either some copper is left in the bead (vide infra, p. 471), a course 
 which is inadmissible if the silver is to be estimated, or else 
 slow cooling is resorted to, the muffle being carefully closed and 
 luted up and the fire withdrawn. The door is not opened again 
 until the beads have solidified; under these circumstances, no 
 sprouting occurs. Slow cooling may also be obtained by cover- 
 ing the cupel containing the silver bead with a red-hot empty 
 cupel. 
 
 When cooled the beads often "flash" i.e., brighten suddenly 
 at the moment of solidification. This is due to the fact that the 
 latent heat of fusion being released raises the temperature of the 
 bead enormously, the metal having been in a state of surfusion 
 at a temperature many degrees below its melting point. The 
 flashing of small beads can rarely be observed. 
 
 The proper temperature for cupellation of gold ores is higher 
 than for that of silver, as loss by volatilisation is less to be 
 feared. The muffle should be at a bright orange-red heat, the 
 cupel red, and the melted lead much more luminous than the 
 cupel ; the fumes should rise slowly to the crown of the muffle. 
 In Western America in both silver and gold assays the heat is 
 kept low enough to enable crystals of litharge to form in a ring 
 
FUSION OR CRUCIBLE PROCESS OF ASSAY. 451 
 
 round the cupel, but the results thus obtained are too high, the 
 lead not being completely eliminated. The whole of the litharge 
 should be completely absorbed. The formation of scales, due to 
 low temperature, is accompanied by a sluggish heavy movement 
 of the fumes, which fall in the muffle. On the other hand, the 
 heat is too great when the cupels are whitish, the fused metal 
 is seen with difficulty, and the scarcely visible fumes rise rapidly 
 in the muffle (Mitchell). If the assays are long in uncovering 
 they may sometimes be started by dropping on them a little 
 charcoal powder wrapped in tissue paper, or still better by 
 placing a piece of charcoal near them. If one freezes before 
 completion it is restarted in the same way, or fresh lead is added 
 or the temperature is raised, but the results are not good. 
 
 The admission of the current of air to the muffle is carefully 
 regulated. Too much air may cause the bath of lead to spirt 
 .and so occasion loss ; too little air delays the operation and 
 causes increased loss by volatilisation and absorption by the 
 cupel. 
 
 The bead thus obtained should be well rounded and bright, 
 loosely adherent to the cupel and slightly crystalline although 
 malleable. If it contains lead it is more globular and brittle and 
 its surface is very brilliant, while it does not adhere at all to the 
 bone-ash. If copper is present the bead adheres firmly to the 
 cupel, and in extreme cases its surface is blackened. Rhodium 
 And iridium occasion black patches at the bottom of the bead ; 
 platinum makes the surface of the bead crystalline and rugose. 
 If the bead cracks on being flattened Van Riemsdijk recommends 
 A second cupellation with some more lead and 10 per cent, of 
 cupric chloride. In this way the metals with volatile chlorides 
 are eliminated. 
 
 Influence of Base Metals on Cupellation Iron. Its oxide is not 
 readily fusible with lead oxide : the button is long in melting, 
 and a brown scoria is left on the cupel, which may entangle lead 
 globules and so contain gold. Care must be taken not to dis- 
 turb the cupel during the operation, as, if the molten lead is 
 moved so as to touch the scoria, part is entangled. The cupel is 
 stained red. 
 
 Zinc burns with a blue flame at first and volatilises, taking 
 gold with it ; it forms a pale yellow scoria, having the same 
 effects as that consisting of oxide of iron. The button is slightly 
 crystalline. 
 
 Tin forms a grey scoria. 
 
 Copper carries gold into the cupel and is usually not wholly 
 removed from the bead. 
 
 Nickel and Cobalt are not so easily oxidised as copper : they 
 may form a dark green scoria and always stain the cupel green. 
 The button is crystalline. 
 
 Antimony does not interfere if less than 1 per cent, is present 
 
452 THE METALLURGY OP GOLD. 
 
 (Rivot). If there is more, in volatilising it takes gold with it> 
 and it forms antimoniate of lead, giving a pale yellow scoria and 
 causing the cupel to crack. 
 
 Arsenic has a similar effect. The last two metals mentioned 
 are removed by scorification if present in large quantities. 
 
 Manganese causes black stains and corrosion of the cupel, and' 
 forms a dark scoria. 
 
 Chromium gives a brick-red stain and scoria on the cupel, and 
 aluminium a grey scoria ; both these metals delay the course of 
 cupellation. 
 
 Cadmium causes a black sooty ring to form inside the cupel 
 near its margin, and gives a brown scoria. 
 
 Tellurium causes loss by volatilisation, and also, in common 
 with some other metals, has the effect of causing subdivision of 
 the cupelled bead. 
 
 Bismuth has less prejudicial effects on cupellation than the 
 metals mentioned above, and may be substituted for lead. How- 
 ever, according to E. A. Smith,* the losses experienced in this 
 case, especially by absorption by the cupel, are much greater 
 than if lead is used. 
 
 The loss during cupellation by volatilisation was proved by 
 Makins. It is never absent, but is insignificant with ores assay- 
 ing below 10 ozs. per ton, especially if highly volatile metals are 
 absent. The absorption by the cupel is more serious (see under 
 Bullion assaying, p. 479). Rivot states that gold is oxidised to 
 some extent at a red heat in the presence of antimonic oxide, 
 litharge or cupric oxide, and that it is the oxidised part which 
 is absorbed by the cupel. This contention is as yet hardly 
 supported by sufficient proof. 
 
 Inquartation and Parting. The bead of silver and gold 
 obtained by cupellation is squeezed between pliers, or flattened 
 by a hammer on a clean anvil, to loosen the bone-ash adhering to 
 its lower surface, and is then cleaned by a brush of wires or stiff 
 bristles. It is then weighed, the silver removed by solution in 
 nitric acid, and the weight of the residual gold taken, when the 
 difference between the two weighings represents the silver. If 
 the bead contains more than one-fourth its weight of gold, 
 enough silver is added to it to make an alloy of about this 
 composition, otherwise some of the silver will remain undis- 
 solved, being protected from the action of the acid by the outer 
 layers of gold. The amount of silver to be added is calculated 
 from the (approximately) known composition of the bead, or 
 guessed from its colour. A pale yellow bead always contains 
 more than 60 per cent, of gold, but a perfectly white bead may 
 not "part" completely. The addition of the silver is effected in 
 the case of small beads by fusion on charcoal by the blowpipe, 
 but large beads are, together with the silver, wrapped in as small 
 * Journ. Chem. Soc., vol. Ixv., p. 624 (1894). 
 
FUSION OR CRUCIBLE PROCESS OF ASSAY. 453 
 
 a piece of lead foil as possible and cupelled. This is called 
 " inquartatian" The resulting bead is cleaned, flattened by the 
 hammer and, if it is large, by passing through the rolls, rolled up 
 into a cornet (vide Bullion assaying) if of convenient size, and 
 dropped into nitric acid. If the gold in the original bead did 
 not exceed one-quarter of the whole, inquartation is omitted. 
 
 The parting is effected in a test tube or a porcelain crucible if 
 the bead is small, in a "parting flask" if large. The strength of 
 acid used for the first boiling varies with the composition of the 
 alloy. Nitric acid of specific gravity I g 25 is used for the 
 inquarted alloy, and more dilute acids for poorer alloys. In 
 .all cases the acids may be previously warmed with advantage, as 
 the gold does not in that case break up into such line particles as 
 if cold acid is used. The acid must be of sufficient strength to 
 .attack the bead instantly, turning it black and giving off nitrous 
 fumes. The temperature is gradually raised to boiling point, 
 and maintained at it for two or three minutes, or until all apparent 
 .action of the acid on the metal has ceased. The acid is then 
 poured off, the residue washed once with boiling water by decan- 
 tation, and if the bead is very large fresh acid added of sp. gr. 1 -32 
 (one part of strong nitric acid to one-half part of water). The 
 boiling is then continued for five to ten minutes longer, when 
 almost all the silver will have been dissolved. A very small amount 
 of silver, weighing from *05 to ! per cent, of the gold, obstinately 
 resists the action of the acid, and remains as a surcharge which 
 may be neglected in almost all ores. The gold may be left as a 
 single piece if it constitutes not less than one part in 10 to 20 
 of the bead : if present in less than this proportion it breaks up, 
 and the finer particles may float and be lost in decantation. 
 Particles floating on the surface may be sunk by touching with 
 a glass rod or by a drop of water let fall on them. Continued 
 heating of the gold in acid makes it agglomerate to some extent, 
 so that it is easier to wash. The second acid is rarely necessary. 
 After decantation of the second acid and washing twice with 
 water, the water is drained off and the porcelain crucibles dried 
 ;at a gentle heat, and then gradually raised to redness. The 
 gold which was previously black and soft, being in a fine state of 
 division, now resumes its usual yellow colour, and hardens so 
 that it can be removed to the pan of a balance and weighed. 
 
 If a parting flask or test tube is used for the boiling, the parted 
 gold is transferred to an unglazed Wedgewood crucible. To 
 effect this the flask is tilled with water, and the crucible placed 
 over its mouth. On inverting both together the gold falls into 
 the crucible, and the flask is removed in such a way as not to 
 disturb the precious metal. The water which has filled the 
 crucible is then poured off, and the crucible heated as before. 
 
 The chief difficulty in parting by the last-named method is 
 encountered in transferring the gold from the glass vessel to the 
 
454 THE METALLURGY OF GOLD. 
 
 "W edgewood crucible ; minute particles of gold may adhere to the 
 glass and are then left behind. The loss of these is the cause of 
 the low results occasionally observed when this method is used. 
 If the glazed porcelain crucibles are used for both boiling and 
 annealing, no transference from one vessel to another of the gold 
 in its soft state is necessary, so that the source of error mentioned 
 above is avoided. On the other hand, the difficulty of boiling 
 the acid in a small porcelain crucible without sustaining loss by 
 projection may prevent the assayer from continuing the boiling 
 for a sufficient length of time, and some silver may thus be left 
 undissolved ; in consequence of this the results obtained are 
 sometimes too high by as much as 2 or 3 per cent, of the weight 
 of the gold. Such errors would, however, be inappreciable in 
 the assay of ores containing less than 1 ounce of gold per ton, and 
 are the less serious for the reason that all the other sources of 
 error (unclean slags, absorption by cupel, &c.), tend to make the 
 result too low. 
 
 By whatever method the parting is performed, very finely 
 divided gold may remain suspended in the liquids, and may thus 
 be lost in the course of decantation. 
 
 To prevent "bumping," one or two small pieces of capillary 
 glass tube or of clay pipe or other porous body are put into the 
 acid together with the alloy to be parted. 
 
 By the operation of parting, silver, palladium and some plati- 
 num are removed in solution, but the greater part of the 
 platinum, and all the rhodium, iridium, &c., remain with the- 
 gold. If the presence of these metals is suspected they must be 
 looked for and removed by special methods (vide infra, p. 486). 
 
 Examination of Assay Materials. The fluxes are usually 
 free from the precious metals, but the litharge, red-lead, and 
 granulated lead contain a small amount of silver and less gold 
 (see p. 30). These are estimated by fusion with charcoal (or in, 
 the case of granulated lead by scorification) and cupellation. 
 
 Examination of the Cupel. When rich ores are assayed, 
 appreciable quantities of gold are carried into the cupel, especially 
 if much copper is present in the lead button. To assay the 
 cupel, all clean bone-ash is detached and thrown away, and the 
 remainder crushed so as to pass through an 80-mesh sieve, and 
 the charge made up as follows : 
 
 Cupel, . 
 Fluorspar, 
 Sand, . 
 Soda carbonate, 
 
 
 
 100 pai 
 75 
 75 
 100 
 50 
 50 
 4 
 
 *ts. 
 
 Litharge, 
 Charcoal, 
 
 
 
 
 
 The fusion is performed in an earthen crucible, and the resulting- 
 button of lead cupelled. The slag of the original fusion may be- 
 
ASSAY BY SCORIFICATION. 
 
 455 
 
 cleaned by adding it to the charge, when less fluxes will be 
 required. 
 
 ASSAY BY SCORIPICATION. 
 
 As stated above, this process is especially applicable to (a) 
 complex ores, (b) very rich ores, (c) ores which are mainly valuable 
 for their silver contents. The losses are small since the slag is 
 highly basic, consisting chiefly of litharge, while the bath of lead 
 is always poor, so that there is little loss of the precious metals 
 by volatilisation. Moreover the operations are easy to conduct, 
 and need not be varied much for different classes of ore. For 
 these reasons the process is generally preferred to the crucible 
 process in Germany and the United States ; if the ore is poor 
 several assays are made and the lead buttons scorified together. 
 The chief disadvantage of the process lies in the small quantity 
 of ore that can be treated, so that the presence of one or two 
 metallic particles of gold may cause the result to be erroneous. 
 
 Scorih'cation is conducted in a muffle at a much higher tem- 
 perature than that required for cupellation. It must be high 
 enough to melt litharge when contaminated by silica and oxides 
 of copper, iron, manganese, &c. A temperature of 1,050 to 
 1,100 C. is usually enough. The charge is placed in a scarifier, 
 a shallow circular fire-clay dish 2 to 3 inches in diameter, which 
 is charged in by the tongs shown at B, Fig. 61. The charge 
 consists of about 50 grains of ore (or ^ A'.T.), 500 to 1,000 
 grains of granulated lead and a few grains of borax glass. The 
 ore is mixed with half the granulated lead, the mixture put in the 
 scorifier and smoothed down, the rest of the lead spread over 
 evenly and the borax put on the top. The amount of lead to be 
 used varies with the nature of the ore. Ricketts and Miller 
 give the following table* as a guide : 
 
 
 For one part Ore take 
 
 Character of Gangue. 
 
 Parts of 
 
 Parts of 
 
 
 Granulated Lead. 
 
 Borax Glass. 
 
 Quartzose, . 
 
 8 to 10 
 
 
 
 Basic, . 
 
 8 to 10 
 
 0-25 to 1-00 
 
 Galena, 
 
 6 
 
 O'lo 
 
 Arsenical, . 
 
 16 
 
 0-10 to 0-50 
 
 Antimonial, 
 
 16 
 
 0-10 to J -00 
 
 Fahlerz, 
 
 12 to 16 
 
 0-10 to 0-15 
 
 Iron pyrites, 
 
 10 to 15 
 
 0-10 to 0-20 
 
 Blende, 
 
 10 to 15 
 
 0-10 to 0-20 
 
 Telluride, 
 
 16 to 18 
 and a cover 
 
 o-io 
 
 of litharge. 
 
 The borax lessens the corrosion of the scorifier and renders the 
 * Notes on Assaying, New York, 1897, p. 96. 
 
456 THE METALLURGY OF GOLD. 
 
 slag more liquid, but its quantity is kept as low as possible to 
 prevent the slag from completely covering the bath of metal too 
 soon. After charging-in, the door of the muffle is closed until 
 fusion takes place. As soon as the lead is melted the door is 
 opened, and a current of air allowed to pass over the bath of 
 metal. Some of the ore is now seen to be floating on the surface 
 of the lead, and is rapidly oxidised, partly by the air and partly 
 by the litharge which immediately begins to form. The sulphur, 
 arsenic, antimony, &c., are thus soon eliminated, while copper, 
 iron and other bases slag oft' with the borax, and the silica, and 
 other acids form fusible compounds with the litharge. Efferves- 
 cence and spirting may occur, especially if the scorifier 
 has not been well dried by warming before it is used. The slag 
 soon forms a ring completely in circling the bath of metal. As 
 oxidation of the lead proceeds, the litharge flows to the sides and 
 increases the quantity of slag until at length the ring closes 
 completely over the metal, leaving a "flat uniform surface" 
 (Percy). This usually happens after from thirty to forty min- 
 utes. The slag should be " cleaned" before withdrawal. This 
 is done by placing 3 grains of charcoal powder wrapped in 
 tissue paper on the surface of the slag, with a pair of cupel 
 tongs, and closing the muffle door. A number of globules of 
 lead are formed by the reduction of the litharge, and these, falling 
 through the slag, extract and carry down with them any gold 
 and silver which it may still contain, and concentrate them in 
 the molten lead below. The fusion being quiet again, the 
 charge is poured into an iron mould by the scorifying tongs, and 
 the lead button cleaned from slag by a hammer. If the button 
 is too large to cupel at once it is re-scorified in the same dish, 
 fresh lead being added if it is not soft and malleable. Less loss 
 of the precious metals is incurred by scorifying lead than by 
 cupelling it, and consequently it is better to reduce any lead 
 button weighing more than 200 to 300 grains by re-scorification 
 before cupellation. 
 
 The slag from the scorifier should always be separately re- 
 treated in the case of very rich ores, but seldom contains much 
 gold. 
 
 The losses incurred in this process are chiefly due to improper 
 temperatures. If the muffle is too cold at first, gold is retained 
 by the slag. This initial low temperature is often indicated by 
 the occurrence of white patches of sulphate of lead on the surface 
 of the slag after pouring. If the slag is pasty, borax is added, 
 but the slag is then rich. It is better to begin again with less 
 ore or more lead. When extremely rich sulphides are assayed, 
 results are often obtained which do not agree well. Stetefeldt* 
 recommends in all such cases that the sulphides be attacked by 
 nitric acid and the residues dried and scorified. 
 
 *Lixiviation of Silver Ores, New York, p. 106. 
 
ASSAY OF SCORIFlCATiON. 457 
 
 The subsequent treatment of the lead button is as already 
 described. Tellurium ores often yield erroneous results, usually 
 ascribed to volatilisation of the gold. Ricketts asserts * that it is 
 not volatile in presence of tellurium (but see page 8), and that 
 no difficulty occurs in either crucible or scorification method if 
 plenty of litharge or lead is used. For rich tellurides 60 to 100 
 parts of lead are required, a large scorifier being used. The 
 tellurium must be driven off before cupellation, as otherwise the 
 gold button will be subdivided into minute particles on the cupel. 
 
 Detection of Gold in Minerals. The detection of gold in 
 loose alluvial ground has been already described, p. 45. A 
 similar method may be employed in the case of auriferous quartz 
 after grinding it. If the concentrates obtained in either case 
 contain sulphides, these are collected, roasted or treated by nitric 
 acid and reground. The light particles of oxide of iron can now 
 be separated from any gold that may be present by washing. 
 " Colour " may often be obtained thus when none could be seen 
 after the first concentration. The washing is made easier by 
 removing from the concentrates the magnetic oxides and iron 
 from the grinding tools by a magnet. All the finest particles of 
 gold are lost in the process of washing, and consequently many 
 auriferous ores cannot be made to " show colour." 
 
 The following method devised by Wm. Skey, analyst to the 
 Geological Survey of New Zealand, is said by him to give good 
 results. The sample of ore is carefully roasted, then digested 
 with an equal volume of an alcoholic solution of iodine for a 
 length of time varying from twenty minutes to twelve hours, the 
 longer time being allowed if the ore is poor. A piece of Swedish 
 filter paper is then saturated with the clear supernatant liquid 
 and afterwards burned to an ash ; if gold is present in the ore 
 the ash is coloured purple, and the colouring matter can be 
 quickly removed by bromine. This method is said to show the 
 presence of as little as 2 dwts. gold per ton in certain ores, but 
 is not uniformly successful. It depends for its success on the 
 solubility of gold in a solution of iodine, and the alcohol must be 
 very pure to prevent reprecipitation of the gold. Iodine, how- 
 ever, is a slow and ineffective solvent for gold,f and the use of 
 bromine is preferable, since it fails less frequently and acts 
 more speedily than iodine. A mixture of 5 to 10 parts of 
 bromine with 100 of water may be used to 100 parts of ore. The 
 ore must be fine enough to pass an 80-mesh sieve and should be 
 reground after roasting. After leaving the mixture to stand for 
 some hours with occasional stirring, the liquid is filtered and the 
 excess of bromine evaporated from the clear solution, which may 
 then be tested by stannous chloride. Occasionally gold ores are 
 met with which give negative results with all these methods, but 
 yield small quantities of precious metal by crucible assay. 
 * Notes on Assaying, p. 79. t Vide p. 11. 
 
458 
 
 THE METALLURGY OF GOLD. 
 
 Estimation of Gold in Dilute Solution. A. Carnot has 
 shown * that the rose colour produced in presence of arseniate of 
 iron is very sensitive, and can be used for colorimetric estimation 
 of small quantities of gold. To attain this end a neutral solution 
 of chloride of gold of known strength is prepared, and some 
 drops of solution of arsenic acid added slowly to it. Then 
 after a time two or three drops of dilute ferric chloride and some 
 hydrochloric acid are added. If the liquid is not acidulated, a 
 flocculent purple precipitate forms ; if it is too acid the reaction 
 fails, and only a faint blue colour is seen. The liquid is made 
 up to 100 c.c. with distilled water, a pinch of zinc dust added, 
 and the mixture shaken in a flask. A colour is produced, 
 varying from rose to purple, according to the amount of gold 
 present. The solution is clear, can be filtered unchanged, and 
 kept without alteration for some time. If more than one milli- 
 gramme of gold is present (1 in 100,000), the colour becomes too 
 intense for small differences to be noticeable ; if less than 
 one-tenth of this amount is present (1 in 1,000,000), the colour 
 becomes too faint. Between these proportions the amount of 
 gold present in a liquid can be determined by comparison with a 
 series of prepared coloured solutions. 
 
 For more dilute solutions, the test described on p. 27, depend- 
 ing on the use of stannous chloride, may be used. A number of 
 precipitates are prepared from solutions of gold containing 
 known amounts, and compared with that given by the solution 
 to be estimated. Suitable volumes of liquid from which the 
 test precipitates are obtained are as follows : 
 
 Strength of Solution. 
 Parts of Water present to each 
 part of Gold. 
 
 Volume of Solution used to 
 obtain a Precipitate. 
 
 1 to 5 millions, 
 5 20 
 20 50 
 50 100 
 
 100 c.c. 
 500 ,, 
 1,000 
 3,000 
 
 If more gold is present than one part per million, the colour of 
 the precipitate is too intense for accurate measurement. 
 
 SPECIAL METHODS OF ASSAY. 
 
 1. Mixed Wet and Dry Method. Complex minerals may 
 be assayed by the method of Cumenge A: Fuchs f as follows : 
 The ore is fused with grey antimonial-copper ore so as to form a 
 
 * Bull, de la Soc. Chimie, 1883. 
 
 t Fr&ny's Encydopcedie. Chimique, T. iii., 16e cahier, p. 131. 
 
SPECIAL METHODS OF ASSAY. 459* 
 
 matte or regulus. All the gold is found concentrated in the 
 matte, which consists of sulphide of antimony, and forms about 
 one-tenth of the weight of the original ore. The matte is fused 
 with litharge and then dissolved in aqua regia, tartaric acid being 
 added, and the solution diluted by boiling water The whole of 
 the silver separates out on cooling as AgCl, which is dried, fused 
 with lead, and cupelled. The gold is precipitated by sulphurous- 
 acid or a little sodic hyposulphite, and is cupelled and weighed. 
 The method is said by its devisers to be accurate, i ut grey 
 antinionial-copper is difficult to obtain free from gold and silver. 
 
 2. Amalgamation Assay. This is useful only in determining 
 the amount of gold present in the ore capable of being extracted 
 by mercury. The sample of ore is pulverised fine enough to pass 
 through a 60 or 80-mesh sieve, made into a paste with a little 
 water, mercury added, and the whole ground in an iron mortar 
 with a pestle for from two to four hours, small additional 
 amounts of water or mercury being added from time to time 
 according to the appearance of the triturated mass. The con- 
 sistency of the mass must be such that the globules of mercury 
 do not sink in it but are broken up into very small particles. 
 A little sodium amalgam dissolved in the mercury prevents it 
 from flouring. The grinding is continued until the particles of 
 ore are all impalpably fine. A machine for the purpose called 
 the "arrastra" mortar has a large pear-shaped muller loosely 
 fitting the inside of the mortar, and capable of being revolved in 
 it by means of a handle. Complete amalgamation is performed 
 in this machine much more rapidly than in an ordinary mortar. 
 When the operator judges that amalgamation is complete, 
 enough water is added to reduce the mass to a thin pulp and 
 stirring is continued for a few minutes to collect the mercury 
 at the bottom. The contents of the mortar are then " washed 
 down " in a pan, the mercury collected and distilled, and the 
 residue, consisting of gold, silver and base metals, cupelled and 
 parted. 
 
 Mr. J. M. Merrick gives a similar method for the assay of 
 pyrites.* He roasts from 8 ozs. to 1 Ib. or more of the finely 
 pulverised pyrites, mixed with marble-dust to prevent caking, 
 and then amalgamates it in a stoneware pot or wooden bucket, 
 collects the mercury and distils it. It is stated that gold may 
 be detected in this way in samples showing none when assayed 
 by fusion. Many auriferous pyritic ores, however, yield but little 
 gold to mercury after roasting, so that the method is uncertain 
 as well as laborious. Such ores, if very poor, are better treated 
 either by chlorination or by Whitehead's method, which will be 
 described subsequently (see p. 461). 
 
 3. Assay by Chlorination. Plattner's method for the assay 
 of roasted pyrites consists in placing the mineral moistened with 
 
 * Mitchell's Manual of Assaying, 1881, p. 694. 
 
460 THE METALLURGY OF GOLD. 
 
 water in a glass cylinder, 200-250 millimetres deep and 20-30 
 millimetres in diameter, and introducing a current of chlorine 
 gas at the bottom. When the odour of chlorine is noticed above 
 the ore, a cover is put on. the stream of chlorine stopped and the 
 whole left for twenty-four hours, after which the reaction is 
 complete if chlorine is still in excess. Boiling water is now run 
 through the ore until all soluble salts have been washed out, 
 and the gold contained in the solution is precipitated by ferrous 
 sulphate, collected, cupelled and parted. According to Balling 
 this method fails if much silver is present, as the chloride of silver 
 formed encrusts the gold and protects it from the action of the 
 chlorine.* He recommends the addition of common salt to 
 dissolve the chloride of silver, but found that the telluride ores 
 of Nagyag yielded only 85 per cent, of their silver and 92 per 
 cent, of their gold when successively treated with chlorine and 
 sodium chloride. 
 
 A better method is to place the completely roasted sample in an 
 ordinary soda-water bottle with enough water to make the whole 
 of the consistency of thin mud. The ore and water should 
 together occupy about two-thirds of the bottle. Bleaching 
 powder and a thin glass bulb filled with dilute sulphuric acid 
 are then added and the bottle securely closed. As cork is 
 attacked by chlorine, glass or vulcanite stoppers are better, and 
 the screw-stoppered bottles are most convenient ; if corks are 
 used they must be wired down. The bottle is then shaken so as 
 to break the sulphuric acid bulb and mix its contents with the 
 bleaching powder, when chlorine is evolved. The bottle is now 
 left for several hours in a warm place, being shaken occasionally 
 by hand to mix its contents. At the end of a period of eight to 
 twelve hours the bottle is opened, and if excess of chlorine is 
 still present the liquid is separated from the ore and the latter 
 washed thoroughly by filtration or decantation. The liquid and 
 washings, whether clear or muddy, are warmed to expel free 
 chlorine and an excess of ferrous sulphate is then added to them. 
 The precipitate is collected, scorified with lead and cupelled. 
 In all cases it is better to keep the first liquid separate from the 
 washings, which should be concentrated by evaporation, since, if 
 this is not done, the precipitate of gold may be too fine to settle 
 and will pass through filter-paper. A better method of precipi- 
 tation is to boil the liquid with iron filings for a few minutes, 
 decant through filter-paper, wash the filings and dissolve them 
 in dilute sulphuric acid when a residue of gold is obtained which 
 is easy to filter. Bromine may be used instead of the materials 
 generating chlorine. The quantities of chemicals required will 
 be such as are sufficient to generate a volume of chlorine equal 
 to twice the capacity of the bottle used, or to make a solution of 
 2 per cent, of bromine in water. Only finely-divided gold is 
 extracted by this method. 
 
 * Balling, L'Art de I'Essayeur, p. 433. 
 
SPECIAL METHODS OF ASSAY. 46T 
 
 4. Whitehead's Method of Assay. A useful method of" 
 determining small quantities of gold and silver in base metals, 
 such as crude copper, or in mattes or base ores, is described by 
 Mr. Whitehead,* as follows : Weigh out from 1 to 4 A.T. of 
 the auriferous material, place it in a beaker of 500 c.c. capacity, 
 and add gradually enough nitric acid to dissolve it completely ;. 
 heat until red fumes cease to come off, dilute with water, and add 
 50 grammes of lead acetate, stir, and when dissolved add 1 c.c.. 
 dilute sulphuric acid and allow the lead sulphate to settle. 
 Filter into a 1,000 c.c. flask and fill to the mark with distilled 
 water. The filter contains the gold which has been collected, 
 and carried down by the sulphate of lead. The filter-paper and, 
 precipitate are dried, the paper burned, and the ash and lead 
 sulphate scorified with test-lead. The button is cupelled and 
 the gold with any trace of silver it may contain is weighed and 
 then parted. 
 
 The solution is divided into two parts, and precipitated by 
 sodium bromide with constant stirring. The precipitates, which 
 consist of a mixture of the bromides of silver and lead, settle 
 quickly, and filter and wash well, only cold water being used. 
 When dry, the precipitate can easily be brushed from the paper, 
 and so the trouble of burning the latter is avoided. The 
 bromides are now mixed with three times their weight of 
 carbonate of soda and a little flour or charcoal, placed in a small 
 crucible, covered with borax, and melted down. The lead thus, 
 obtained should be free from copper, and is easy to cupel. Dupli- 
 cate assays usually agree within two-tenths of an ounce of silver 
 per ton. The use of the lead acetate is to cause the precipitates, 
 of gold and silver to settle quickly, and to enable them to be 
 filtered effectively. Sodium bromide is used instead of the 
 chloride, on account of the greater insolubility of the silver salt. 
 
 5. Assay of Pyrites. Schwartz's Method^ 100 grammes of 
 pyrites is fused with 46'6 grammes of iron turnings under a 
 cover of salt. The protosulphide of iron formed is crushed and 
 dissolved in dilute sulphuric acid; the gold remains undissolved. 
 The liquid is filtered, the residue roasted, fused with borax and 
 granulated lead, cupelled, and parted. It has been suggested 
 that the iron turnings are really unnecessary, and only serve to 
 increase the amount of regulus. Simple fusion with suitable 
 fluxes gives a rich regulus which contains all the gold. 
 
 Stapff's Method.^ The pyrites is fused with sulphur and an 
 alkaline sulphate, and the alkaline aurosulphite thus formed is 
 dissolved in water. The gold is precipitated from the filtered, 
 liquid by acidifying with sulphuric acid, and is cupelled and 
 parted. 
 
 * Chem. News, 1892, vol. Ixvi., p. 19. 
 
 t Dingler's Polyt. Journal, vol. ccxviii. 
 
 Fr^my's Encyclopaedic Chimique, vol. iii., 16e cahier, p 202. 
 
462 THE METALLURGY OF GOLD. 
 
 6. Assay of Purple of Cassius. One part of purple of 
 Oassius is fused with three parts of carbonate of soda, cooled 
 .and dissolved in water. The gold remains undissolved, and is 
 collected on a filter and cupelled after incineration.* 
 
 7. Assay of a Mint Sweep. The assay of sweepings of the 
 floors of jewellers' workshops, refineries and mints, and that of 
 pulverised crucibles, stirrers, &c., which have been used in 
 melting gold bullion, are included in this section. The larger 
 pieces of metal are removed from the sweepings by sieving and 
 washing. The sweep then still contains a large number of 
 metallic particles which will not pass through a fine sieve ; a 
 -considerable percentage of sawdust, charcoal and graphite is also 
 present. The sample must be selected with great care and with 
 due regard to the fact that the lower parts of a heap are richer 
 than the top, owing to subsidence of metallic particles through 
 the mass. The amount of moisture present is first estimated, 
 .and the sample dried and passed through an 80-mesh sieve, after 
 which it is roasted in a muffle and scorified with eight to twelve 
 times its weight of lead and a few grains of borax. The 
 metallics are scorified separately, or passed at once to the cupel. 
 The remaining operations do not differ from those required in 
 the case of a rich ore. Double parting is necessary, as both 
 silver and gold are usually present. From J to | A.T. of the 
 unroasted material is a convenient weight to take for assay, 
 but the metallics must be separated from a much larger quantity 
 .and estimated by themselves. 
 
 CHAPTER XX. 
 THE ASSAY OF GOLD BULLION. 
 
 Assay of Gold Bullion. The assay of gold bullion, as described 
 in this chapter, has for its sole object the estimation of the 
 percentage of gold present in the alloy, all other constituents 
 being disregarded. In the first instance, the simple case of the 
 assay of gold alloys containing appreciable quantities of only 
 copper and silver will be dealt with. Refined gold ingots and 
 the alloys used for coinage, and for almost all jewellery, come 
 under this head. The effect of large quantities of other impurities 
 and the precautions thereby rendered necessary will be discussed 
 later. 
 
 The method universally employed is that of cupellation and 
 .subsequent parting. The gold bullion is cupelled with silver 
 * Ann. de Pharmacie, vol. xxxix. 
 
THE ASSAY OP GOLD BULLION. 463 
 
 and lead, by which the greater part of the base metals present 
 is removed as oxides dissolved in litharge, and an alloy of gold 
 and silver left on the cupel. This is " parted " by nitric acid, 
 which dissolves the silver and leaves the gold unattacked. 
 
 In the following pages the practice at the Royal Mint is 
 described, but the same description would apply, with veiy 
 slight alterations, to the methods used at other mints and assay 
 offices. 
 
 The "parting assay" was first mentioned in a decree of 
 King Philippe of Valois, published in the year 1343.* The 
 methods of procedure in the 17th century have been briefly de- 
 scribed by Savotf and by J. Reynolds, J and more fully in the 
 Compleat Chymist. In 1666 Pepys saw the parting assay being 
 practised at the Mint in the Tower of London, and from his 
 description it is clear that the method then employed bears a sur- 
 prisingly strong resemblance to that of the present day. In 1829 
 a Royal Commission was appointed in France to examine into 
 all questions relating to the methods of assaying gold and silver. 
 The results of the labours of that Commission are to be found in 
 great part in the Appendix to the Report of the Select Committee 
 on the Royal Mint, 1837. The Committee arrived at the con- 
 clusion that the method adopted for assaying gold often over- 
 stated the amount of precious metal by 1 part per 1,000. The 
 Mint Conference held in Vienna in 1857,|| resulted in the almost 
 universal adoption of a more uniform method of manipulation. 
 
 The degree of accuracy now attained in most assay offices 
 reduces the probable error in the report of an assay to 0*1 per 
 1,000, but, to prevent the error from rising above this amount, 
 all weighings must be correct to 0-05 per 1,000, which is not 
 usually the case in ordinary bullion assays. 
 
 The system may be conveniently regarded as comprising six 
 distinct operations, viz. : 
 
 1. Selection of the sample. 
 
 2. Preparation of the assay piece for cupellation. 
 
 3. Cupellation. 
 
 4. Preparation of the assay piece for parting. 
 
 5. Parting and annealing the cornets. 
 
 6. The final weighing and reporting. 
 
 1. Selection of the Sample. At the Royal Mint the con- 
 tents of the pots are poured into moulds, each pot yielding about 
 nine bars, and a piece is cut from the middle of the end of each 
 of the first and last bars formed. Experience proves that these 
 samples are usually representative of the whole mass viz., 1,200 
 
 * First Report of the Royal Mint, 1870, p. 103. 
 
 t Discours sur les Medalles Antiques. Paris, 1627, p. 72. 
 
 A New Touchstone for Gold and Silver Wares, London, 1679, p. 362. 
 
 Report on the Royal Mint, 1837, Appendix B, p. 123. 
 
 || Kunst-und Gewerbeblatt Baiern, 1857, p. 151. 
 
464 THE METALLURGY OP GOLD. 
 
 ozs. In the case of the ingots, weighing 400 ozs., which are 
 received from the Bank of England for coinage, a single sample 
 is cut from the middle of one of the lower edges of the ingot. 
 When the sample is not representative of the whole ingot (a 
 somewhat rare event, judging from results), other base metals 
 besides copper are probably present.* A more certain way of 
 finding the composition of an ingot is to melt it under charcoal,, 
 stir well and dip out samples from the top and bottom of the 
 mass, with an iron ladle. The metal thus obtained is granu- 
 lated by pouring slowly in a thin stream into a porcelain-lined 
 kettle filled with warm water, which must be constantly stirred. 
 The dip-sample may also be taken by a little charcoal crucible 
 fastened to an iron rod, a lid of charcoal being put on directly it 
 emerges from the molten metal. Dip-samples should always be 
 taken when very impure and base ingots are to be assayed. In 
 place of taking a cutting from the edge of an ingot a drilling 
 machine is used in the Paris Mint and in many other offices ; by 
 this instrument portions are taken for assay from any part of the 
 ingot near its surface. 
 
 2. Preparation of the Assay Piece for Cupellation. If the 
 sample to be assayed is a single piece cut from a bar it is " flatted " 
 on a clean anvil by a hammer with a rounded face, weighing 
 about 11 Ibs. A convenient thickness for the sample is about 
 -^2 inch. The piece is then wrapped in paper which is marked dis- 
 tinctively, and an assistant then performs the operation of "bring- 
 ing to weight," or obtaining a piece approximately gramme in 
 weight. This he does by cutting with shears and filing. 
 
 The use of the shears can only be learned by practice, but the 
 following remarks may be of use to a beginner. The metal to be 
 cut is held firmly between the fore-finger and thumb of the left 
 hand. Care must be taken, and considerable force exercised by 
 the fingers if necessary, to keep the plane of the piece of metal 
 to be cut perpendicular to the cutting faces of the shears, other- 
 wise damage is done to the latter. Only clean portions of metal 
 must be used. 
 
 The weights used in gold assaying are the ^ gramme, which is 
 stamped " 1000," and decimal subsidiary weights stamped 900, 
 800, &c., and 90, 80, 70 and so on down to 0'5. These stamped 
 numbers denote the number of J milligrammes ("milliemes ") 
 contained in the weight. Ordinary weights in the gramme 
 system may accordingly be used, each milligramme corresponding 
 to two milliemes in the assay system. The report finally made 
 gives the number of parts (in milliemes and tenths) of pure gold 
 in 1 ,000 parts of the alloy. 
 
 Since the assay is reported to -5^575- part, it is evident that 
 the balance used must clearly indicate a difference in weight 
 of 0-1 per 1,000 or -05 milligramme. It is convenient to have 
 * See under Liquation of Gold Alloys, p. IS, 
 
THE ASSAY OP GOLD BULLION. 465 
 
 the balance adjusted so that one subdivision on the ivory scale 
 traversed by the pointer corresponds to this quantity. In some 
 offices, however, " preparing " balances less delicate than this 
 are used. The weight of alloy must not differ from the 1,000 
 bv more than four subdivisions of the scale. The weighed 
 portion is folded in a small square of paper and tucked inside 
 the paper containing the rest, until it can be checked by the 
 assayer. Weighing becomes a very rapid operation with prac- 
 tice, a skilful workman being able to prepare forty to fifty pieces 
 in an hour. 
 
 The assayer now checks the weighed piece on a more delicate 
 balance. Those used at the Royal Mint have beams only 6 
 inches long, weighing 156 grains, and requiring only ten seconds 
 for a complete double oscillation. Suspension of the pans was 
 formerly effected by means of two hard steel points instead of 
 a knife-edge ; in the year 1892 light agate knife-edges were 
 introduced. Point suspension gave greater lightness, but the 
 wear of the points was rapid, and slight blunting soon impaired 
 the accuracy of the balance, necessitating repairs. These 
 balances are released by a lever working in a vertical direction. 
 The excess or deficiency in weight is recorded on the assay 
 paper in tenths of a millieme, the weight of the piece being 
 adjusted by the assayer if it shows a deviation greater than four 
 of these tenths. The piece is wrapped up in lead foil, which 
 is very pure, and weighs 4 grammes or eight times as much as 
 the assay piece. The lead is twisted into a form resembling 
 that of a conical sugar bag, and the gold together with the silver 
 necessary for parting is dropped in and wrapped up. The lead 
 packets are put in order in the numbered compartments of the 
 wooden tray shown at c, Fig. 67, p. 475, their position being- 
 noted on the assay paper. The corners of the packets are 
 squeezed down so as to fit the cupels by pliers specially designed 
 for the purpose, and the assays are then ready to be charged 
 into the furnace. 
 
 L'he silver added to the assays is in the form of square pieces- 
 cut off by means of gauged shears from a strip rolled to uniform 
 thickness ; the squares weigh about 1^ grammes each or two 
 and a-half times the weight of the alloy. If silver is present 
 in the alloy it must be allowed for. In the case of fine gold 
 (used as proofs} or of any sample of gold of high standard, an 
 alloy of silver and copper, 966'6 fine, is used instead of fine 
 silver. The use of the copper is to make these assay pieces of 
 about the same composition as the standard alloy (916-6 fine) 
 worked with them. Fine silver is added to all alloys containing 
 more than 50 milliemes of copper. All the silver added for 
 purposes of parting must of course be free from gold. 
 
 It was formerly considered necessary for the metals to be pre- 
 sent in the proportion of 1 of gold to 3 of silver, but, as earlv as 
 
 30 
 
466 
 
 THE METALLURGY OF GOLD. 
 
 the year 1627, Savot relates thut the proportion of 1 to 2 was 
 used, and strong acid employed in the boiling, " quand on veut 
 faire quelque essay curieux et exact." * Both Chaudet and 
 Kandelhardt recommend the proportion of 1 to 2J, on the 
 ground that less silver is then retained by the cornet than in 
 any other case. If much more than three parts of silver are 
 present the gold breaks up in the acid. Pettenkoferf found 
 that the proportion of 1 to 1-75 could be employed if the assays 
 were boiled in concentrated nitric acid for some time. 
 
 The amount of lead used varies with the proportion of base 
 metals present. The following table shows the proportions 
 recommended by D'Arcet,J Cumenge & Fuchs, and Kandel- 
 hardt || respectively : 
 
 Gold in 1,000 
 
 Amount of Lead employed for one part of Alloy. 
 
 Metal being Copper. 
 
 D'Arcet. 
 
 Cumenge & Fuchs. 
 
 Kandelhardt. 
 
 1,000 
 
 1 
 
 1 
 
 8 
 
 900 
 
 10 
 
 14 
 
 16 
 
 800 
 
 16 
 
 20 
 
 20 
 
 700 
 
 22 
 
 24 
 
 24 
 
 600 
 
 24 
 
 28 
 
 24 
 
 500 
 
 26 
 
 32-34 
 
 28 
 
 400 
 
 34 
 
 32-34 
 
 28 
 
 300 
 
 34 
 
 32-34 
 
 32 
 
 200 
 
 34 
 
 32-34 
 
 32 
 
 100 
 
 34 
 
 30 
 
 32 
 
 50 
 
 34 
 
 28 
 
 32 
 
 o 
 
 - 
 
 11 
 
 32 
 
 Kandelhardt's table is modified to make it uniform with the 
 others. It must be remembered that, although the above table 
 gives the quantities of lead which will remove the greater part 
 of the copper present in the alloy during cupellation, the last 2 
 or 3 milliemes are obstinately retained by the gold, and cannot 
 be entirely eliminated even by a second cupellation with fresh 
 lead. Instead of attempting to remove all the copper present in 
 an alloy of low standard by one operation, using large quantities 
 of lead as above, it saves time and gives more uniform results if 
 a smaller amount of lead is used in two successive cupellations. 
 The following table shows the proportions of lead employed by 
 Mr. F. W. Bayly at the Royal Mint for gold-copper alloys. The 
 object in view is not to remove all the copper by cupellation, 
 
 * Discours sur les Medcdles Antiques, Paris, 1627, p. 73. 
 
 t Bergwerksfreund, 1849, vol. xii., p. 6. 
 
 I Pelouze et Fremy, Traiti de chimie g6n6rale. 
 
 Encydopadie Chimique, vol. iii., L'or, p. 154. 
 
 I) Oold-Probirverfahren, Berlin, p. 3. 
 
THE ASSAY OF GOLD BULLION. 
 
 467 
 
 i)ut to obtain a well-formed, clean and bright button suitable for 
 parting : in each case silver is added equal to two and a-half 
 times the weight of gold present. 
 
 Amount of Gold per 
 1,000 parts. 
 
 Amount of Lead 
 used for one part of 
 Alloy. 
 
 Proportion of Lead 
 to Copper. 
 
 916*6 
 
 8 
 
 96: 
 
 866 
 
 9-15 
 
 64: 
 
 770 
 
 14-75 * 
 
 64: 
 
 666 
 
 16 * 
 
 48: 
 
 546 
 
 17-5 * 
 
 38: 
 
 333 
 
 18-0 * 
 
 27 : 
 
 * In these cases two-thirds of the lead is charged in at first, 
 and after the cupellation is finished the remainder is added in 
 the form of a bullet. The results are good, and the absorption 
 of gold by the cupel not excessive. 
 
 3. Cupellation A ssay Furnace.^ That in use at the Royal 
 
 Fig. 63. 
 
 s^ 
 
 Fig. 64. 
 
 Fig. 62. Scale, 1 in. = 1 ft. 5 ins. 
 
 t The description of this furnace is, in the main, the same as that given 
 lay Professor Roberts- Austen in Percy's Metallurgy of Silver and Gold. 
 pp. 255-7. 
 
468 THE METALLURGY OF GOLD. 
 
 Mint is shown in elevation in Fig. 62. It consists of an outer 
 casing of wrought-iron plates about J inch thick, united by angle 
 iron. This casing is connected with a chimney 60 feet high by 
 means of a wrought-iron hood (a) and flue which is provided 
 with a damper (6). The lining consists of Stourbridge fire-bricks, 
 and there are five openings in the front of the furnace. Fuel 
 is introduced through the uppermost one of these under the 
 flap, shown partly open, the position of which can be varied by 
 means of two saw-edged inclined planes. The opening (d) com- 
 municates directly with the muffle (e) and may be closed by 
 sliding iron doors, or by the firebrick front (g) in conjunction 
 with a sliding plate, not shown in the figure, covering the upper 
 part of the opening (d). There are also two small openings, 
 closed by sliding doors, at the side of (d), for introducing and 
 withdrawing the fire-bars ; these are seldom used to regulate 
 the draught, for which latter purpose the doors (&) of the ash 
 pit (I) are usually employed. The muffle is formed of fire-clay, 
 and is pierced with four holes at the upper part of the back, by 
 means of which a draught is established ; these slope from within 
 downwards, in order to prevent particles of fuel from finding 
 their way into the muffle. The muffle is shown in front elevation 
 in Fig. 63. 
 
 There are eleven fire-bars, but only the three outer ones on 
 each side are covered with fuel. On the other five bars rest the 
 cast-iron girder-plate (Fig. 64), which is flat on the upper surface, 
 but is strengthened with ribs on the under-surface in order to 
 prevent buckling. Above the girder-plate the muffle is fixed in 
 an inclined position, so that all the cupels may be readily visible 
 from the front. The muffle rests on a bed placed on the girder- 
 plate consisting of pieces of fire-brick plastered over with fire- 
 clay. It is convenient to cover the top of the muffle with a thick 
 luting, composed of fire-clay and a little graphite in order to- 
 check radiation, and to protect the muffle from the coke let fall 
 on it from above, and a fire-brick (m) is placed behind to prevent 
 excessive heating of the back of the muffle. Charcoal, anthracite 
 and coke might all be used in such a furnace, which would be 
 built narrower if either of the first two were employed, but on 
 account of the expense they are now rarely employed, and in the 
 Royal Mint coke is always used. It is broken to about the size 
 of hens' eggs and carefnlly screened. 
 
 Assay furnaces, heated by gas, have long been in use at the 
 Berlin and Utrecht mints, at the Sheffield assay office and at 
 many other places. Their use is becoming more general, especi- 
 ally in America. 
 
 Furnaces to burn soft coal are used in Western America 
 where good coke is very expensive. They differ little from coke 
 furnaces in construction, but have less space between the muffle 
 and the side walls. The flame of the coal is chiefly instru- 
 
THE ASSAY OF GOLD BULLION. 
 
 469 
 
 mental in heating the muffle, a comparatively thin bed of fuel 
 being employed. 
 
 The cupels in use at the Royal Mint are in sets of four, the 
 outer margin being square, as shown in Fig. 65. This form was 
 
 SECTION ON IJNE X, Y. 
 
 Fig. 65. 
 
 THE CUPEL. 
 
 suggested by Mr. Joseph Groves of the Royal Mint, in 1872,* 
 and is found to facilitate charging-in and withdrawal. It will 
 be observed that the cupel mould A B is in three pieces, which 
 are separated to take out the finished cupel. The furnace tools 
 in use are shown in Fig. 66, where a represents the cupel tongs 
 
 Fig. 66. 
 
 Scale, 1 in. = 1 ft. 
 
 and b the tongs used for charging-in the lead packets on to 
 the cupel. The drawing at the left hand of the figure represents 
 a section of the " peel," described on p. 477. 
 
 The furnace is lighted by faggots and charcoal, coke being 
 added at intervals after the first few minutes. The full draught 
 is used until the furnace has nearly attained the proper temper- 
 ature. The floor of the muffle is lightly covered with bone-ash 
 and the cupels introduced at least twenty minutes before charg- 
 ing-in is begun, some pieces of charcoal being laid near the front 
 to assist in the attainment of a uniform temperature throughout 
 the muffle. The furnace is filled with coke, well raked down 
 but not packed too tightly, reaching to a height of about an inch 
 
 * Third Annual Report of the Royal Mint, 1872, p. 30. 
 
470 THE METALLURGY OF GOLD. 
 
 above the top of the muffle. The coke is at a full red heat 
 throughout before the cupellation is begun and no fresh fuel is 
 added during the operation. The muffle is usually ready for 
 work in about an hour from the lighting of the furnace. The 
 charcoal is then removed from the muffle and all dust and ashes 
 blown out by a pair of hand bellows. 
 
 The packets of lead containing the silver and gold are now 
 transferred to the cupels, arranged in rows corresponding to 
 those on the tray, the muffle being at a bright orange-red heat. 
 This charging-in is performed by the tongs, b, in Fig. 66, over 
 the top of the plumbago front (g, Fig. 62), the object of which is 
 to render the temperature of the muffle as equable as possible. 
 The cupels at the back are filled first, and by the time six assay 
 pieces have been introduced, the first one should be melted and 
 " uncovered " by the removal of the black crust which forms at 
 first. At the Royal Mint the full fire of 72 assays takes about 
 five minutes to charge-in. The slider is now put in place over 
 the upper part of the opening (d), and the air supplied to the 
 muffle increased by closing the damper. While the muffle 
 door is open the indraught is diminished by opening thi& 
 damper so as to prevent the cooling action of the current of air 
 from proceeding too rapidly. The air supplied to the muffle 
 and furnace may be entirely regulated by this damper. The 
 furnace operation should be performed rapidly and in such a 
 way that all the cupellations may be completed at as nearly as 
 possible the same time. 
 
 Distinct stages may be noted in the action which npw takes 
 place on the cupel. Almost immediately the surface of the 
 molten metal becomes covered with greasy-looking drops of 
 litharge, which are rapidly absorbed by the porous cupel and 
 replaced by others. They pass over the surface at first slowly,, 
 but as the operation continues move with greater rapidity. In 
 from eight to fifteen minutes the metal suddenly becomes 
 uniformly dull and glowing except for iridescent bands, pro- 
 duced by extremely thin films of fluid litharge, which are seen to 
 pass over it. On the disappearance of these bands a bright 
 liquid globule of a greenish tint is left, but the cupels are not 
 withdrawn from the furnace until the expiration of another 
 fifteen to twenty minutes so that the last traces of lead may 
 be oxidised and absorbed. The completion of cupellation takes 
 place first in the front rows and proceeds regularly backwards. 
 
 The cupels are withdrawn from the furnace while the assay 
 pieces are still fluid, and "flashing" ensues in a few seconds. 
 " Flashing " is most marked in the purer buttons, in which but 
 little copper or lead remains. Slight effervescence may occur in 
 these cases, but the buttons are never sufficiently freed from 
 base metals for " sprouting " to take place. 
 
 Some assayers do not remove the assays from the furnace 
 
THE ASSAY OF GOLD BULLION. 471 
 
 until "brightening" or "flashing" has taken place and the 
 buttons have become solid. The muffle door must be closed 
 and the temperature reduced to effect this. Removal of liquid 
 buttons saves much time and is always performed at the Royal 
 Mint, but has two dangers. A cupel may be upset or " spitting " 
 may take place. Under the conditions given above, however, a 
 trace of copper is always retained by the cupelled button and all 
 fear of spitting thus removed. 
 
 The " flashing " of gold assays was shown by Van Riemsdijk * 
 to be due to solidification after superfusion. The temper- 
 ature of the fluid metal falls until a certain point is reached 
 when it solidifies and the sudden disengagement of the latent 
 heat of fusion reheats the cooling globule to its true melting 
 point viz., 950 and a peculiarly intense light is emitted which 
 rapidly fades as the temperature again falls. A sudden jar at 
 any moment causes the flashing to occur instantly. If the alloy 
 contains a minute quantity of iridium, rhodium, ruthenium, 
 osmium or osmium-iridium (the metals of the platinum group 
 which are non-malleable, refractory to heat and resist the action 
 of acids), the tendency of the cupelled metal to preserve its liquid 
 state below the melting point, and therefore to flash during the 
 final solidification is entirely prevented. These researches point 
 to a simple method for detecting the presence of metals of the 
 platinum group (except platinum and palladium) in ingots of 
 commercial gold ; for, if assays made on them solidify directly 
 they are withdrawn from the muffle without flashing, it will be 
 safe to conclude that the metal contains some of the above- 
 named impurities.! 
 
 The buttons (which are of the form represented at a, Fig. 68, 
 p. 475) are removed from the cupels, after cooling, by a pair of 
 sharp-nosed pliers, cleaned by means of a stiff brush or by immer- 
 sion in warm dilute hydrochloric acid, and are placed in the 
 compartments corresponding to their cupels in the tray (d, Fig. 
 67). If the bone-ash is not completely removed from their lower 
 surface it is of little moment, since bone-ash is readily dissolved 
 by nitric acid on parting. The surface of the cupels must be 
 carefully examined for minute beads of metal due to spirting of 
 the lead bath, which sometimes happens if there is too strong a 
 draught. If any such beads are found in a cupel the fact is 
 noted and the assay repeated. 
 
 If traces of lead remain in the button it is more globular, 
 separates more easily from the bone-ash of the cupel and has a 
 brilliant steely surface. The effect of the presence of other 
 metals is discussed on pp. 484-488. 
 
 Temperature. The exact temperature suitable for cupellation 
 can only be ascertained by practice, and the varying light of 
 
 * Chemical News, vol. xli., 1879, p. 126. 
 
 tProf. Uoberts- Austen in Tenth Annual Mint Report, 1879, p. 43. 
 
472 THE METALLURGY OF GOLD. 
 
 the day may occasion error in judging the degree of heat. The 
 remarks on temperature in the cupellation of buttons from ores 
 (p. 450) apply here. Care should be taken to ensure that the 
 heat, is so high before " charging-in " that the chilling which 
 necessarily takes place during this operation shall not cool the 
 muffle below the requisite temperature. It is of more conse- 
 quence that the muffle should be uniformly hot throughout than 
 that any absolute degree should be attained, as the checks used 
 eliminate uniform errors due to high temperature. 
 
 The measurement of the temperature of the assay muffle has 
 not been frequently attempted. In a paper read before the Royal 
 Society (Phil. Trans., 1828, 79-96), Mr. James Prinsep, Assay 
 Master of the Mint at Benares, gives an account of experiments on 
 the subject by observing differences in the behaviour of a number 
 of silver-gold and gold- platinum alloys when heated. He "made 
 trials in different parts of the same (muffle) furnace. The 
 disparity of heat," he remarks, " is greater than might be 
 supposed, and where, as in assaying the precious metals, so 
 much depends on the temperature at which the operation is 
 performed, it would be useful to know every difference in this 
 respect obtaining in different countries, and its effect upon the 
 quality or standard of bullion." His results were as follows : 
 
 Maximum alloy 
 
 melted. 
 Muffle of an assay furnace : front . . , pure silver. 
 
 : middle (average) 
 
 : behind (average) 
 
 The temperatures at which these alloys melt are 
 
 Silver, . . . 945 (Violle). 
 Silver, 70 \ Qfift0 
 
 Gold, 30 I' 
 
 Silver, 50 1 QQn o 
 
 Gold, 50 /* 
 
 No further experiments were made until the year 1892 y when 
 the temperature of the muffle in the furnace at the Royal Mint, 
 described on p. 467, was measured by the author by means of 
 the Le Chatelier pyrometer.* The arrangements were as 
 follows : The wires, sheathed in clay pipe-stems, were fixed in 
 a porcelain tube, the junction being protected by plates of mica. 
 This tube was introduced into the furnace by sliding through 
 another of wider bore fixed in the brick door of the muffle. The 
 inner end of the narrow tube was plugged by clay and coated 
 over with a fusible mixture of sand, pipeclay and felspar. The 
 galvanometer was shielded from radiation from the furnace, and 
 remained at about a constant temperature, but the zero was 
 frequently determined in order to detect any changes. In 
 
 *Journ. Chem. Soc., vol. Ixiii., 1893, p. 707. 
 
THE ASSAY OF GOLD BULLION. 
 
 473 
 
 calibrating the wire, the effects of heating various lengths of it 
 were noted, and allowance made for the alterations in the 
 deflections of the needle due to this cause. The junction was 
 kept stationary at each point in the muffle until its temperature 
 had become constant, as indicated by the cessation of movement 
 of the spot of light on the scale. 
 
 The muffle is 15 inches long and 6J inches wide, and its full 
 charge consists of seventy-two assays arranged in twelve rows, 
 the distance between the front and back rows being 10^ inches. 
 The position of the junction was about 1 inch above the cupels 
 in all cases. The error in observation in taking the readings 
 was probably not more than O'l mm. on the scale, which was 
 equivalent to O32. 
 
 
 
 Temperature in 
 
 
 
 degrees above 
 
 Place. 
 
 Time. 
 
 m. p. of gold-silver 
 
 
 
 buttons left on the 
 
 
 
 cupel, viz., 952. 
 
 Middle of 1st (front) row of cupels, 
 
 
 10.19a.m. 
 
 114-0 
 
 99 M 
 
 
 10.20 
 
 
 113-0 
 
 3rd row of cupels, 
 
 
 
 10.21 
 
 
 123-0 
 
 4th 
 
 
 
 
 10.22 
 
 
 124-0 
 
 7th 
 
 
 
 
 10.23 
 
 
 144-0 
 
 3rd 
 
 
 
 
 10.25 
 
 
 120-0 
 
 llth 
 
 
 
 
 10.28 
 
 
 140-0 
 
 Right side, llth 
 
 
 
 
 10.29 
 
 
 142-5 
 
 Left llth 
 
 
 
 
 10.30 
 
 
 140-0 
 
 Middle of llth 
 
 
 
 
 10.31 
 
 
 139-0 
 
 3rd 
 
 
 
 
 10.32 
 
 
 112-0 
 
 7th .. 
 
 
 
 
 10.36 
 
 
 128-0 
 
 The assay pieces were charged in between 10.5 and 10.11; cupellation 
 was complete at 10.23, and the charge drawn at 10.47 a.m. 
 
 Place. 
 
 Time. 
 
 Temperature 
 above 952. 
 
 Middle of 3rd row, 
 
 11.3 a.m. 
 
 112-5 
 
 7th 
 
 11.6 
 
 131-5 
 
 llth 
 
 11.8 
 
 143-0 
 
 7th 
 
 11.9 
 
 1230 
 
 3rd 
 
 11.12 
 
 112-5 
 
 The assay pieces were charged in between 10.50 a.m. and 10.55 ; cupella- 
 tion was complete at 11.8, and the charge was drawn at 11.30. The circuit 
 was accidentally broken at 11.15. 
 
474 
 
 THE METALLURGY OF GOLD. 
 U. 
 
 Place. 
 
 Time. 
 
 1 
 Temperature 
 above 952. 
 
 Middle of 3rd row, 
 
 
 
 
 11.56 a.m. 
 
 70 
 
 
 7th 
 
 t 
 
 
 
 
 11.59 ,, 
 
 91 
 
 
 llth 
 
 i 
 
 
 
 
 12.0 noon 
 
 102 
 
 
 , 7th 
 
 
 
 
 
 12.2 p.m. 
 
 97 
 
 
 3rd 
 
 
 
 
 
 12.3 
 
 82 
 
 
 llth 
 
 
 
 
 
 12.8 
 
 105 
 
 
 7th 
 
 
 
 
 
 12.10 
 
 98 
 
 
 3rd 
 
 
 
 
 
 12.15 
 
 88 
 
 
 3rd 
 
 
 
 
 
 12.17 
 
 91 
 
 The assay pieces were charged in between 11.45 and 11.51 ; cupellation 
 was complete at 12.8, and the charge drawn at 12.25. The muffle appeared 
 to the eye to be decidedly cooler than usual. 
 
 These three experiments were conducted on different days. 
 They show that the temperature gradually rises in passing from 
 the front to the back of the muffle, the rate of change falling 
 from 5 per row of cupels at the front to about 2 at the back, 
 whilst along the sides the temperature is 1 or 2 higher than in 
 the middle line. The mean difference in temperature between 
 the 3rd and 7th rows is 15 -7, and between the 7th and llth 
 rows 10*5, whilst the mean temperature of the muffle is 1070. 
 These differences seem to exercise an effect on the absorption of 
 gold by the cupel. The mean "surcharge" (see p. 478) of the 
 check assays in all the charges, 939 in number, worked off at 
 the Royal Mint between January 1, 1880, and March 11, 1892, 
 is as follows : 
 
 Check in middle of 
 
 (a) 3rd row. 
 + 0-1828 
 
 (6) 7th row. 
 + 0-1533 
 
 (c) llth row. 
 + 01347 
 
 Consequently, if, as is probable, the experiments indicate the 
 ordinary conditions of the muffle, a rise of about 5 would cor- 
 respond to a diminution in the surcharge of O'Ol per 1,000. 
 
 4. Preparation of the Assay Buttons for Parting. The 
 operation of " flatting " which now takes place requires some 
 skill, as on it to a great extent depends the accurate and regular 
 formation of the "cornets" finally obtained. A hammer weigh- 
 ing* about 7 Ibs., with a convex face, should be used. The anvil 
 is kept quite bright and clean and used for this purpose only. 
 A heavy blow is first delivered on the centre of the button, the 
 diameter of which is thereby increased to nearly that of a three- 
 penny piece. Two lighter blows are then given on opposite 
 sides of the disc so as to elongate it, giving it the form shown at 
 o, Fig. 68. 
 
 After being annealed in the iron tray, f, Fig. 67, which is- 
 placed in the muffle by the tool, g, and left until it is red hot. 
 
THE ASSAY OF GOLP BULLION. 
 
 475 
 
 the flattened buttons are passed in succession through a pair 
 of jeweller's rolls. The rolls are adjusted so that one passage 
 
 CL 
 
 Fig. 67. 
 
 through them reduces the buttons to the required thickness viz. r 
 about that of an ordinary visiting card. The " fillets " (c, Fig. 68) 
 
 thus obtained should all be 
 
 x >w of uniform size and thickness, 
 
 / \ < with " wire edges," as ragged 
 
 -f \ iSffll edges expose them to loss- 
 
 during the boiling. After 
 being rolled they are replaced 
 in the tray, f^ and annealed 
 at a dull red heat. The object 
 of the first annealing is to- 
 soften the buttons and facili- 
 tate their passage through the 
 rolls, while that of the second 
 is to enable the fillets to be 
 rolled into "cornets''' or 
 spirals, c?, between the finger 
 and thumb. Care is taken in 
 the latter operation to make 
 that which was formerly the 
 lower side of the button form 
 
 Fig. 68. 
 
 Scale, full-size. 
 
 the outer face of the coil, for a reason given below (p. 488). Thi& 
 face is easily recognised, as it is less brilliant than the other and 
 is marked with undetached flattened particles of bone-ash! In 
 the Mints of India, Scandinavia and some other countries, the 
 fillets are rolled tightly round a glass rod and then slipped off 
 the end. This operation occupies a longer time than the coiling 
 between finger and thumb, and leaves the cornet less open to- 
 rapid attack by the acid than if the spiral form is used. 
 
 5. Parting. This was formerly effected by boiling with 
 nitric acid in glass "parting flasks." Platinum boiling trays 
 
476 THE METALLURGY OF GOLD. 
 
 save time, and are now used whenever possible. The silver is 
 dissolved by the acid which must be free from chlorine in any 
 form, sulphuric and sulphurous acids, or sulphide from which 
 sulphuric acid may be formed. A small quantity of silver is 
 kept in solution, in the stock of acid, so that chlorine, if present, 
 would instantly be detected. 
 
 When the flasks are used, 2 ozs. of nitric acid of specific 
 gravity 1'2 are put into each flask and raised to boiling point. 
 A cornet is then introduced and boiling continued for 15 or 20 
 minutes (i.e., for about 10 minutes after nitrous fumes cease to 
 be given off). Hot distilled water is then added, the solution of 
 nitrate of silver decanted off, and the flask washed by filling 
 with hot water, and decanting. Two ounces of hot nitric 
 acid of specific gravity 1-3 is now poured in and the boiling 
 continued for fifteen or twenty minutes, a parched pea or 
 piece of charcoal being added to prevent bumping; after 
 which decantation and washing is twice performed. Another 
 boiling with acid of specific gravity l'-i is recommended by 
 Chaudet with the object of dissolving out the last traces 
 of silver and leaving the gold quite pure. This practice 
 has been adopted by many assayers, but is useless and causes 
 Joss of gold. If any small particles of gold have become detached 
 from the cornet, time must be allowed for them to settle before 
 ^ach decantation. After the last decantation the flask is filled 
 with hot water, the top covered by a small porous crucible, and 
 the whole is carefully inverted ; the pure gold, which is of a 
 dark brown colour and exceedingly fragile, falls through the 
 liquid and rests in the crucible, the water which enters with it 
 being afterwards poured off. The crucible is dried and then 
 annealed at a red heat over gas or in the muffle, when the gold 
 shrinks greatly, though still preserving its shape, hardens and 
 regains its ordinary pale yellow colour. It can now be weighed. 
 
 When a platinum boiler is used the cornets are put on 
 platinum pins, as at the Sydney Mint, or more usually into 
 platinum cups, one of which is shown in Fig. 69. These cups 
 are supported in a platinum tray (which holds 144 
 <jups at the Royal Mint) and the whole lowered by 
 a, platinum hook into a platinum vessel containing 
 about 80 ozs. of hot nitric acid. Great attention 
 must be paid to the temperature of the acid. At 
 the Royal Mint acid of specific gravity 1 -26 (in which, 
 however, a small quantity of silver is already dis- 
 solved) is used, and the temperature at the moment 
 of introduction of the tray is 90 C. If the acid is 
 Bolder than this the cornets tend to break up, some 
 pieces being usually detached in the boiling of 144 
 cornets, even if the temperature is only 1 or 2 lower 
 than the correct point. It the acid is much hotter than 90* 
 
THE ASSAY OF GOLD BULLION. 477 
 
 (93 or more) the action is so energetic that the vessel may- 
 boil over. If the temperature of the acid is as low as 85% 
 the cornets are not immediately attacked ; they turn a coppery 
 hue and then bubbles of gas form slowly on them. In such 
 cases (where the action does not begin within 1 or 2 seconds 
 after entrance into the acid) the only chance of saving the cornets, 
 is to withdraw them instantly and raise the temperature further. 
 Boiling is kept up vigorously for twenty to thirty minutes, 
 and the tray is then withdrawn, drained, washed by dipping 
 vertically in and out of a vessel of hot distilled water, drained 
 again, and placed in a second platinum boiler filled with boiling 
 nitric acid of specific gravity 1 '32 (also containing a little nitrate 
 of silver). In this the cornets remain for a period of thirty to 
 forty minutes, when they are drained and washed as before, and 
 are then ready for annealing. 
 
 The platinum tray may be dried on a hot plate, or may be 
 introduced into the muffle at once while still wet. The iron 
 peel, shown in Fig. 66, is used to support the tray in the muffle. 
 Rotating peels are sometimes used so that the temperature of 
 annealing can be equalised by rotating the tray without with- 
 drawing it from the furnace. A thin platinum plate may be 
 placed between the tray and the peel to prevent scales of oxide 
 of iron from entering the cups. The floor of the muffle should be 
 scraped clean of bone-ash to prevent projection of the latter into 
 the cups owing to the fall of drops of water from the wet tray when 
 it is first introduced, or a clean tile may be put in the muffle to 
 support the tray. The cornets are not injured by the slightly 
 explosive evaporation of small quantities of water contained in 
 their porous substance. Annealing should be conducted at as 
 high a temperature as possible, consistent with the safety of the 
 cornets. They fuse at 1,045 0. (having the same melting point 
 as pure gold), at which temperature the muffle appears orange- 
 red. If annealed at a low temperature, the cornets are rough in 
 texture, dull and fragile, being crushed easily between the finger 
 and thumb. In this condition they adhere to the platinum, and, 
 in detaching them, fragments are often left sticking fast inside 
 the cups. If annealed properly the cornets are smooth, lustrous 
 and hard, showing signs of incipient fusion under a magnifying 
 glass, and only yielding to considerable force exercised by the 
 finger and thumb. Under these conditions they can always be 
 detached from the platinum entire. By being annealed the 
 cornets, which after boiling are very soft and fragile, and dark- 
 brown iri colour, shrink and harden, and regain the ordinary 
 yellow colour of gold (e, Fig. 68). 
 
 Relative advantages of Parting in Flasks and in Platinum 
 Boilers. The use of platinum trays and boilers effects a great 
 saving of time in decanting and washing, as one operation takes 
 the place of as many as 144. If the standard of an alloy is 
 
478 THE METALLURGY OP GOLD. 
 
 unknown, so that it is not certain that the cornet will remain 
 entire in the acid, it must, of course, be boiled separately, as, if 
 one cornet in the tray breaks up, fragments may adhere to a 
 number of others. The manipulation of the platinum tray is 
 easier than that of parting flasks, and, in addition, the treatment 
 of the cornets is more uniform, so that the correction afforded by 
 the use of checks is more trustworthy. On the other hand, if 
 platinum or palladium is present in an individual cornet, it 
 imparts a straw-yellow or orange colour to the acid ; but where 
 :a number of cornets are boiled together, it is obviously impossible 
 to say from which the colour is derived, so that less informa- 
 tion on the subject is obtained. A saving of acid is effected by 
 the platinum vessels, 144 assays being boiled in 80 ozs. of acid 
 at the Royal Mint, while 288 ozs. would be required if parting 
 flasks were used. 
 
 6. Weighing the Cornets. The final weighing is performed 
 
 -on a balance capable of indicating differences of 0'05 per 1,000, 
 or one-fortieth qf a milligramme, with a weight of ^ gramme 
 in each pan. The "checks" or "proofs" (vide infra] are weighed 
 first, and their mean excess or deficiency in weight applied as a 
 correction to all the cornets worked with them. The " weighing - 
 in " correction (p. 465) is also allowed for, and the report is at 
 
 once indicated by the marks on the weights without further 
 
 calculation. 
 
 Surcharge. The gold cornet does not actually contain the 
 whole of the gold present in the original alloy and nothing else. 
 Gold is lost by (a) volatilisation ; (6) absorption by the cupel ; 
 (c) solution in the acid. On the other hand, the cornet always 
 retains (1) some silver; (2) occluded gases. The algebraical sum 
 of these losses and gains is called the "surcharge," since the 
 cornet usually weighs more than the gold originally present in 
 the assay piece ; if the reverse is the case, the work is regarded 
 as less accurate by some assay ers. The various losses and gains 
 
 .are discussed in detail below. 
 
 Losses of Gold in Bullion Assaying. The losses of gold 
 
 can only be incurred in three ways, namely : 
 
 1. Absorption by the cupel. 
 
 2. Volatilisation in the muffle. 
 
 3. Solution in the acid. 
 
 The two latter classes of loss were discussed by the late Mr. G. 
 H. Makins in a paper " On certain Sources of Loss of Precious 
 Metals in some Operations of Assaying," which was read before 
 the Chemical Society and published in their Journal (1860, 13, 
 77). He found gold and silver in the proportion of about 1 to 
 9 in the dust taken from a flue used only in gold and silver 
 cupellation, but did not attempt to ascertain the percentage loss 
 of gold by volatilisation. He also showed that large amounts of 
 gold were dissolved by nitric acid in the course of assaying, and 
 
THE ASSAY OF GOLD BULLION. 
 
 479 
 
 attributed the dissolution of gold to the presence of nitrous acid, 
 but supposed that it would not be dissolved in the weaker acid, 
 where nitrous acid was formed in larger quantities, owing to the 
 protective action exercised by undissolved silver, which formed 
 the positive element in the gold-silver couple. The fact that 
 gold is actually dissolved in nitric acid, presumably as oxide, and 
 remains dissolved even after dilution with water, was proved by 
 Mr. A. H. Allen (Chem. News, vol. xxv., p. 85). 
 
 According to Bruno Kerl,* an alloy which requires from four 
 to eight times its weight of lead for cupellation, sustains a loss 
 which may be as high as 1 millieme of gold ; if sixteen to 
 thirty -two parts of lead are required the loss reaches 2 milliemes. 
 
 According to Rossler the losses are heavier, varying from 1 to 
 3 milliemes of gold when four to five parts of lead are required 
 for cupellation. As the result of a number of experiments con- 
 ducted by the author at the Royal Mint,f it was found that the 
 losses there in the assay of standard gold (916-6 fine) vary from 
 0-4 to 0-8 per 1,000, of which about 82 per cent, is absorbed by 
 the cupel, 8 per cent, is dissolved in the acid, anol the remaining 
 10 per cent., which is unaccounted for, is probably volatilised. 
 These ratios, however, vary considerably, as a hot fire increases 
 the loss by absorption and volatilisation, while by prolonged 
 boiling in acid, and especially by annealing the fillets at a high 
 temperature, the amount of gold dissolved in the acids is 
 increased. An increase in the percentage of copper in the assay 
 piece is accompanied by an increase in the loss under all three 
 heads. This increase of loss has long been known. The follow- 
 ing table of the results of expeiiments on synthetic alloys of gold 
 and copper is given by Pelouze and FreinyJ : 
 
 Actual Standard 
 of the Alloy. 
 900 
 800 
 700 
 600 
 500 
 400 
 300 
 200 
 100 
 
 Difference 
 Surcharge. 
 
 025 
 0-50 
 O'OO 
 000 
 0-50 
 0-50 
 0-50 
 0-50 
 050 
 
 The high surcharge assigned in this table to the 800-alloy is 
 difficult to understand. The usual experience is that an increase 
 of copper in an alloy causes greater loss of gold. The following 
 table gives the relative surcharges obtained in the parting assay 
 of gold of different standards ; it is compiled from a number of 
 results obtained at the Royal Mint by the author : 
 
 * MetaMurflische Probirkunst. , 1880. 
 
 ^Journ. Chem. oc.,lxiii , 1893, p. 710. 
 
 J Trait6 de Chimie., 3ine. Edition, vol. iii., p. 1230 
 
480 
 
 THE METALLURGY OF GOLD. 
 
 Standard of Alloy. 
 
 Surcharge 
 on the 
 Assay Piece. 
 
 Loss of Gold per 1,000 
 of Pure Gold. 
 
 916'6 
 
 + 0-250 
 
 0'63 
 
 900-0 
 
 + 0225 
 
 0-65 
 
 800-0 
 
 - 0-075 
 
 i-oo 
 
 666-6 
 
 - 0-2 
 
 1-20 
 
 5J6-0 
 
 - 0-7 
 
 220 
 
 333-3 
 
 - 2-8 
 
 9-30 
 
 The third column is obtained from the surcharge and the fine- 
 ness in gold of the cornets obtained viz., 999-1. The amount 
 of lead used was eight times the weight of the alloy in the first 
 three cases, and the quantities given in Mr. Bayly's table on 
 p. 467 for the last three. This table, as already explained, only 
 gives relative surcharges and losses of gold ; the absolute 
 amounts vary with the treatment. 
 
 The influence of the temperature of the furnace on the sur- 
 charge is pointed out by Prof. Roberts-Austen.* The following 
 table is compiled from the results of his experiments : 
 
 Temperature of Furnace. 
 
 Composition of Alloy. 
 
 Surcharge. 
 
 (a) Slightly lower than usual, . . 
 (6) Ordinary temperature, . . . 
 (c) Slightly higher than usual, . , 
 
 j Gold 916-7 ) 
 | Copper 83-3 j 
 
 + 0-05 
 
 -o-io 
 
 -0-37 
 
 The loss of gold was found to be 0-645 per 1,000 in series (a), 
 and 0-723 part in series (b). The more exact results of the 
 author are given on p. 474. 
 
 Rb'sslerf has shown that the loss of gold in cupellation 
 increases with the amount of lead used and decreases as the 
 amount of silver is increased. 
 
 Silver Retained in the Cornet. It has been found at the 
 Royal Mint that after boiling in the first acid the average 
 amount of silver retained is about 2-5 parts per 1,000. The 
 amount left undissolved by the stronger acid varies with the 
 length of time of its action. Under normal conditions with a 
 surcharge of 0*2 to 0-3, the silver left in the cornets is from 0-8 
 to 0-9, and this residue is not easily removed. If the acid con- 
 tinues to boil until its constant strength is reached, gold is dissolved 
 to a considerable extent (0'5 per 1,000), while the proportion of 
 silver present suffers little diminution, rarely falling below 0'6. 
 If much gold is dissolved the results are not so uniform. When 
 checks are used perfectly satisfactory results are obtained with a 
 surcharge of + 1*0 per 1,000 or even more.} 
 
 * Percy's Metallurgy of Silver and Gold, p. 275. 
 
 t Ding '1. Polytech. Journ., vol. ccvi., p. 185. 
 
 T. K. Rose, Accuracy of Bullion Assay, Journ. Chem. Soc. (1&93), p. 704. 
 
THE ASSAY OF GOLD BULLION. 481 
 
 The amount of silver retained by the cornets varies with the 
 proportion of silver to gold in the fillet. Thus in an experiment 
 conducted by the author* fillets in which the relative proportions 
 were 247 : 1 and 2-4:1 yielded cornets containing respectively 
 0*6 and 0-7 per 1,000 of silver. This is a common source of 
 error, as the variation of O'l would appear in the result, while 
 such small differences in the proportion of parting silver added 
 to the gold are usually disregarded. Even if the initial weight 
 of silver is constant, the weight in the fillets may vary from 
 unequal losses in the furnace. This loss was found to be about 
 1 per cent, when standard gold is being assayed, 1 -6 per cent, in 
 the case of gold 800 fine, and 3 per cent, if it is 666 fine. 
 
 Occluded Gases. Graham proved f that cornets retained twice 
 their volume of gases (mainly carbonic monoxide) in occlu- 
 sion after annealing. This amounts to two parts by weight in 
 10,000, and is reckoned as silver in the preceding paragraph. 
 According to Varrentrapp, the gas retained varies with the 
 temperature at which annealing takes place. 
 
 Checks or Proofs. Since the losses and gains detailed above 
 are dependent on so many conditions, it is always necessary to 
 subject check-pieces of known composition to the same operations 
 as the alloys under examination. The use of checks in the Royal 
 Mint was prescribed by law as early as the 14th century. % 
 Standard trial plates (916-6 fine) were made and used for this 
 purpose. Since, however, it is impossible to guarantee that a 
 mass of alloyed metal shall have absolutely the same composition 
 throughout, it is better to use pieces of pure gold, a corresponding 
 amount of pure copper being added in order to make the assays 
 absolutely comparable. The correction to be applied to a gold 
 assay is given by the following formula : 
 
 Let 1,000 represent the weight of the alloy originally taken. 
 
 p the weight of the piece of gold finally obtained. 
 
 a: the actual amount of gold in the alloy expressed in thousandths. 
 
 a the weight of pure gold taken as a check, approximately equal to x. 
 
 b the "surcharge," i.e., the loss or gain in weight experienced by a 
 
 during the process of assay expressed in thousandths. 
 Then x ^ p or the corresponding loss or gain experienced by x is equal 
 
 to^^; sothatp = a: (i) or a: = -^ (ii). 
 
 d O, o, -r Q 
 
 Example Let p = 916 '7 thousandths. 
 
 a = 1000-0 
 
 b = 0'3 ,, gain in weight. 
 
 Then & is a positive change, and therefore has the + sign. 
 
 Hence- x - 916 ' 7 x 100 
 
 X ~ 1000 + 0-3 
 = 916-424 
 
 *T. K. Rose, Accuracy of Bullion Assay, Journ. Chem. Soc. (1893),. 
 p. 706. 
 
 iPhil. Trans. Roy. Soc., 1866, p. 433. 
 J Mint Report for 1873, p. 38. 
 
 31 
 
482 THE METALLURGY OP GOLD. 
 
 This result would be reported as 916*4, and, therefore, the following rule 
 is approximately correct ; if there is a gain in weight by the checks, the 
 amount is deducted from the weight of each of the other cornets ; if a loss, 
 it is added. A piece of pure gold weighing 1000 may thus be taken, without 
 appreciable error, as a check on assays of all alloys over 900 tine, provided 
 that the "surcharge" does not exceed + 0'5. This error may be reduced 
 to zero by taking as a check-piece an amount of gold equal in weight to 
 that present in the alloy under examination, an unnecessarily laborious 
 process when assay pieces of various degrees of fineness are present. 
 
 In the above calculation it was assumed that absolutely pure 
 gold was used for proofs. It is doubtful whether it has ever 
 been obtained. The method of preparation of proof gold in use 
 at the Royal Mint is that given at p. 11. Since the assay of 
 an alloy only gives the relative fineness of proof gold and the 
 alloy, it follows that, if the proof gold is not quite pure, the 
 amount found in the alloy will be in excess of the truth. If a 
 sample of proof gold is less pure than the finest yet obtained, an 
 allowance is made. Thus, if it is 999-9 fine, a deduction of 0-1 
 is made from all results of assays checked by it. This deduction 
 is readily proved to be a very close approximation to the correct 
 one. 
 
 Lastly, the " weighing-in " correction is applied. If the 
 original weight taken was say 1000-4 (recorded as + 4), it is 
 sufficient to deduct 0'4 from the final weighing. In this correc- 
 tion 04 of alloy is reckoned as fine gold, but the error is 
 inappreciable, and will remain so if the difference from 1000 is 
 kept less than 0*5. 
 
 Limits of Accuracy in Gold Bullion Assay. Attention 
 may here be drawn to the errors introduced by the lack of 
 delicacy of the very finest assay balances in ordinary use. It 
 has elsewhere been shown by the author * that by weighing in 
 the way indicated above, errors not greater than 0'15 per 1,000 
 may be introduced. It is, therefore, clear that this amount 
 represents the limit of accuracy when such balances are used. 
 By weighing correctly to 0-01 per 1,000, however, and performing 
 all other operations with scrupulous care, then in the determina- 
 tion of gold in high -standard alloys of gold and copper, or of 
 gold, silver and copper, whether pure or contaminated by small 
 quantities of lead, bismuth, zinc, antimony, nickel and some 
 other elements, the error does not exceed 0-02 per 1,000, if 
 the mean of three results is taken. 
 
 Parting by Sulphuric Acid. The use of sulphuric acid of 
 66 B. instead of nitric acid for parting is recommended by some 
 assayers on the ground that the losses of gold by dissolution in 
 nitric acid are variable, while sulphuric acid does not dissolve 
 gold. The inconveniences suffered by the use of sulphuric acid 
 are that (1) lead and platinum are left undissolved by it; (2) 
 violent bumping of the liquid occurs during ebullition ; and (3) 
 * Journ. Chem. Soc., vol. Ixiii., pp. 700-714. 
 
THE ASSAY OP GOLD BULLION. 483 
 
 sulphate of silver is not very soluble in water, and the washing 
 is consequently done with dilute sulphuric acid. However, less 
 silver is left undissolved in the cornets than in the parting by 
 nitric acid if the proportion of gold to silver is between 1 : 2 
 -and 1 : 3. 
 
 Preliminary Assay. If the composition of an alloy is quite 
 unknown, a preliminary assay is necessary in order to determine 
 the right quantities of silver, copper and lead to be added. This 
 determination may be made by the touchstone, by considerations 
 of the colour and hardness of the alloy, or by cupelling 2 grains 
 of it with 6 grains of silver and 30 grains of lead, and parting 
 the button in a flask. Simple cupellation with lead gives satis- 
 factory results if silver is absent or insignificant in quantity ; 
 according to Fremy this method, in which parting is dispensed 
 with, is accurate to 3 milliemes if carefully performed with proofs. 
 
 Assay of Gold by Means of Cadmium. Balling has shown* 
 that cadmium may be substituted for silver in the operation of 
 parting. The J gramme of gold alloy is placed in a porcelain 
 crucible in which a little fragment of potassic cyanide has been 
 previously fused in order to protect the metal from the air. 
 Cadmium is then added in the proportion of 2J to 1 of gold. If 
 silver is present in addition, the combined weight of cadmium 
 and silver must be 2| times that of the gold. The whole is 
 fused and then cooled and plunged into hot water to clean the 
 button, which is then crushed, parted in nitric acid (specific 
 gravity 1'3), dried and weighed. The silver, if any, can be esti- 
 mated by precipitation as chloride from the acid solution. By 
 this method the losses of gold and silver incidental to cupellation 
 are entirely avoided. A similar method, employing zinc in place 
 of cadmium, had previously been recommended by Guptner.f 
 
 Alloys of Gold, Silver and Copper. These may be assayed 
 %y the method just given, the copper being estimated as differ- 
 ence ; or the gold may be estimated as usual, and other assay 
 pieces cupelled with enough lead to remove all the copper. The 
 buttons thus obtained contain silver and gold only, and the pro- 
 portion of silver is found by difference. The method of double 
 cupellabion, by which the button of silver and gold is weighed 
 and then subjected to inquartation and parting is less accurate. 
 
 The cupellation designed to remove the copper is made with 
 less lead than the quantities given on p. 466. If little gold is 
 present, half the amount of lead there given is used, with 
 increasing proportions as the amount of gold present increases. 
 The temperature of cupellation must also be lower than for gold, 
 approximating more to that used for silver. Proofs of similar 
 composition must be used and the operations require much 
 practice before the necessary skill is acquired. If less than 
 
 * Oestr. Zeitsch. fur Berg, und Httmvem., 1879, p. 597. 
 i Zeitsch. Jur Anal. Chem., 1879, p. 104. 
 
484 
 
 THE METALLURGY OF GOLD. 
 
 30 per cent, of gold is present originally, the alloy may be- 
 parted at once without cupellation and the silver estimated by 
 weighing as chloride. 
 
 Effects of the Presence of Other Metals on Gold 
 Bullion Assay. The effects on cupellation are the same as those- 
 given under ore assay, p. 458. In general, if a scoria is formed 
 owing to the presence of large quantities of antimony, arsenic, 
 cobalt, nickel, iron, tin, zinc, or aluminium, there is a loss of 
 gold. The alloy should in that case be scorified with lead as a 
 preliminary to cupellation. If mercury is present gold is vola- 
 tilised (vide p. 486). The presence of tellurium is indicated by 
 the formation of numbers of minute beads of precious metal 
 dispersed over the surface of the cupel. Tellurium compounds- 
 are best analysed in the wet way, see p. 488. 
 
 If members of the platinum group are present they remain 
 undissolved by the parting acid, and hinder the solution of the 
 silver, and the assay is consequently rendered unreliable. The 
 treatment of the alloys is discussed below, p. 486. The effects 
 of the presence of small quantities of various metals on the 
 surcharge in the ordinary parting assay is shown below. The 
 table is the result of experiments made by the author.* The 
 presence of 5 per cent, bismuth does not affect the surcharge. 
 All assay pieces contained 1,000 parts of gold, 2,500 parts of 
 silver, and 91 parts of copper, other metals being added in the 
 proportions indicated. 
 
 Metal added, 
 
 None. 
 
 Antimony, 
 50 parts. 
 
 Zinc, 
 70 parts. 
 
 Tellurium, 
 33 parts. 
 
 Iron, 
 50 parts. 
 
 Nickel, 
 50 parts. 
 
 Surcharge, 
 
 + 0'42 
 0-44 
 0-38 
 
 + 0'28 
 0-28 
 
 + 0-35 
 0-38 
 
 + 0-02 
 0-12 
 
 + 0-22 
 0-32 
 
 + 0-22 
 0-25 
 
 These differences indicate the necessity of employing special 
 checks containing these metals if such be present in the alloys. 
 
 It is often impracticable to apply the ordinary parting assay 
 to the examination of low-standard alloys of gold with other 
 metals. These are then tested by various other methods, of 
 which a summary is given below, the alloys being grouped in 
 four series for convenience : 
 
 A. Alloys requiring scorification. 
 
 B. Amalgams. 
 
 C. Alloys containing members of the platinum group. 
 
 D. Tellurium compounds. 
 
 A. Scorification of Alloys. Alloys of arsenic or antimony 
 are reduced to a fine powder and scorified with thirty parts of 
 
 *Journ. Chem. Soc., vol. Ixiii. (1893), p. 706. 
 
THE ASSAY OF GOLD BULLION. 485 
 
 lead and a half part of borax. If the slag becomes pasty towards 
 the end of the operation more borax is added, a little at a time. 
 If the lead button obtained is hard, a second scorification is 
 necessary, with the addition of more lead. There is considerable 
 loss of gold from volatilisation, and therefore wet methods of 
 analysis are preferable. 
 
 Iron or Manganese Alloys. The operation is tedious and 
 difficult with these alloys, as they are difficult to fuse, having 
 higher melting points than pure gold,* and the oxides of iron 
 do not form easily fusible compounds with the litharge. An 
 extremely high temperature and much borax is required ; ten 
 parts of lead and one of borax usually suffice. 
 
 Cobalt and Nickel. Twenty parts of lead are used, but no borax 
 at first, so that the oxidation of the nickel may not be hindered. 
 A very high temperature and the subsequent addition of two 
 parts of borax are necessary. Several successive scorih cations 
 .are required as nickel and cobalt are difficult to oxidise. 
 
 Zinc. Oxide of zinc does not form a fusible mixture with 
 litharge, and the slag is only rendered pasty by borax, unless it 
 is added in large quantities. Gold is lost by volatilisation, but 
 the loss is minimised by slagging off the zinc as rapidly as 
 possible. Use fifteen to twenty parts of lead and two to three 
 parts of borax added little by little, until the slag is fluid. f 
 
 Tin. Twenty parts of lead are required ; oxide of tin is rapidly 
 formed, but the slag is not easily fusible. Large amounts of 
 borax are necessary, or still better, borax mixed with potash 
 "which forms a fusible stannate with SnO 2 . 
 
 Aluminium. Alloys containing this metal cannot be assayed 
 by scorification and cupellation. As soon as fusion takes place 
 in the muffle, aluminium floats to the top of the bath, being of 
 low density, and is rapidly oxidised, producing alumina which 
 forms an exceedingly infusible scoria not easily removed by 
 litharge. The production of the latter, moreover, is checked by 
 the scum. 
 
 Mr. Edward Matthey observes + that the removal of aluminium 
 by digestion in hydrochloric acid, and collection of the residual 
 gold, does not yield satisfactory results. The process he recom- 
 mends is as follows : Accurately weighed portions of 50 grains 
 each of the alloys are fused with litharge, under a flux of potas- 
 sium carbonate and borax with a small proportion of powdered 
 charcoal, and the resulting slag re-fused with a further small 
 quantity of litharge and powdered charcoal. The lead buttons 
 containing all the gold (the aluminium having combined with 
 the fluxes employed) are cupelled, and the resulting gold 
 cupelled with silver and parted with nitric acid in the usual 
 
 * Fremy, Ency. Chim., T. iii., L'or, p. 147. 
 
 iOp. cit., p. 148. 
 
 SPhil. Trans. Roy. Soc., 1892, p. 643. 
 
486 THE METALLURGY OF GOLD. 
 
 manner. The assays must be worked with checks or standards 
 of fine gold and pure aluminium. 
 
 In the majority of the preceding cases it is better to analyse- 
 the alloys by wet methods, see p. 488. 
 
 B. Amalgams. The alloy is placed in a weighed porcelain, 
 crucible and gradually heated so as to drive off the mercury. 
 After the greater part of the mercury has been driven off the 
 temperature is raised to a full red heat which is maintained for 
 half an hour. About O'l per cent, of mercury still remains in 
 the gold after this treatment and can only be completely removed 
 by cupellation and parting. Checks must be used, as the loss of 
 gold in the operation may amount to 1 per 1,000. 
 
 C. Platinum Group. Cupellation must be performed at a. 
 higher temperature than usual, and the iridescent bands are seen, 
 to remain longer, although they are less numerous. 
 
 (1) Platinum-Gold Alloys. The button obtained by cupellation 
 is dull and crystalline. If the alloy contains as much as 7 or 8- 
 per cent, of platinum, the cupellation proceeds slowly, brighten- 
 ing is only obtained at a very high temperature and the button 
 appears flattened, and has a rough crystalline surface and a grey 
 colour. If more than 10 per cent, is present, brightening does 
 not occur at all, and the other features just mentioned are more 
 strikingly exhibited. On parting, the platinum is dissolved 
 with the silver if it does not form more than 10 per cent, of the 
 original alloy, but the assay-piece must be boiled in acid for a long 
 time, and the parting is less complete than in the simple case of 
 a gold-silver button, from 0'05 to 0'4 per cent, of platinum being 
 left in the cornets. This is completely removed by a second 
 parting. Unless proofs are made up of similar composition the 
 results are not satisfactory if more than 1 or 2 parts of platinum 
 are present per 1,000 of alloy.* 
 
 According to results recently obtained in the laboratory at the 
 Royal Mint by Mr. A. Stansfield, it is impossible to free buttons- 
 containing platinum from lead at the ordinary temperatures 
 attainable in a muffle. It is necessary to employ the oxy- 
 hydrogen blowpipe for the purpose, with the result that the 
 losses are irregular. In the cupellation of pure platinum, 
 buttons weighing from 0*002 to O'Ol gramme retain about 10' 
 per cent, of their weight of lead, and those weighing from OO4 
 to 2 grammes retain about 33 per cent. 
 
 If parting is effected by boiling in concentrated sulphuric acid 
 instead of in nitric acid, almost all the platinum remains undis- 
 solved with the gold. 
 
 Chaudett recommends the following proportions of silver for 
 this purpose: 
 
 * According to Spiller (Proc. Chem. Soc., 1897, No. 180, p. H 8> 
 platinum is but little attacked by nitric acid in the presence of silver,, 
 only about 1 per cent, being dissolved. 
 
 t L'art de I'essayeur, Paris, 1835. 
 
THE ASSAY OF GOLD BULLION. 
 
 487 
 
 Ratio of Platinum to Gold. 
 
 Ratio of Silver to Platinum 
 and Gold together. 
 
 Less than 1 to 10. 
 1 1. 
 More than 10 ,, 1. 
 
 2 to 1 
 1'5 ., 1 
 1-25 1 
 
 The reason for the reduction, shown in the table, in the 
 amount of silver when the platinum is increased is that, according 
 to Chaudet and Haindl, considerable quantities of platinum are 
 dissolved, and the error caused thus is corrected by leaving some 
 silver undissolved in the cornet. The losses of platinum, how- 
 ever, are very irregular, and the method is not a good one. The 
 reason for the variation is probably the lack of homogeneity in 
 the buttons and prills, as gold-platinum alloys show a considerable 
 amount of liquation. 
 
 The platinum is not dissolved by nitric acid if zinc or cadmium 
 is substituted for silver. 
 
 Mr. Edward Matthey recommends the following method for 
 alloys of platinum 900, and gold 100 parts : 50 grains each are 
 taken of the alloys and treated with an excess of nitrohydro- 
 chloric acid which gradually dissolves the whole. The resulting 
 solutions of platinum-gold chloride are then evaporated nearly to 
 dryness to drive off the free acid and diluted with distilled water 
 to about 20 c.c., a degree of strength ascertained by experiment 
 to be the best for the precipitation of the gold. The metallic 
 gold is thrown down by means of crystals of oxalic acid, and is 
 carefully washed, dried a ad weighed. It is necessary to use 
 checks. 
 
 (2) Palladium-Gold Alloys. The palladium is dissolved in 
 parting if the weight of silver is at least three times that of the 
 gold, yielding an orange coloured solution. Matthey recom- 
 mends double parting. Separation may also be effected by fusion 
 with six to eight parts of potassic bisulphate, and dissolving out 
 the dark brown palladious sulphate by boiling water.* A second 
 fusion is usually necessary to render the gold residue quite pure. 
 
 As in the case of platinum, palladium may be separated from 
 gold by dissolving the alloy in aqua regia, evaporating to dry- 
 ness, taking up with water and precipitating the gold by oxalic 
 acid from a very dilute solution : the palladium remains in 
 solution. 
 
 (3) Rhodium- and Iridium-Gold Alloys. Iridium, if present, 
 always sinks to the bottom of the cupelled button as it is very 
 dense (specific gravity = 21 -38), and is not usually fused at the 
 temperature of the muffle, but occurs in the state of fine black 
 crystalline particles. Hence, when the button is rolled into a 
 
 *H. Rose. 
 
488 THE METALLURGY OF GOLD. 
 
 cornet with the lower face outwards (p. 475) iridium occurs as 
 black sooty spots or streaks which are seen by a lens to fill up 
 depressions in the surface of the gold. 
 
 Both rhodium and iridium are almost insoluble in aqua regia. 
 If gold alloys containing both of them are parted in the ordinary 
 way with nitric acid, only a small quantity of rhodium goes into 
 solution with the silver. The residue consisting of the gold, the 
 iridium and most of the rhodium may then be attacked by 
 dilute aqua regia when the gold is dissolved together with only 
 traces of the other metals. These may be separated by evaporat- 
 ing the solution to dry ness and heating to dull redness, when 
 the reduced metals being no longer alloyed may be completely 
 separated by dissolving the gold in aqua regia. 
 
 According to d'Hennin,* iridium may be separated from gold 
 by fusing it with fluxes, the charge being made up as follows : 
 
 Iridic gold, 12*5 grains. 
 
 Sodic arseniate, 3 ,, 
 
 Black flux, 18 
 
 Ordinary flux (consisting of borax, cream 
 
 of tartar, charcoal and litharge), 20 
 
 The iridium forms a speiss with the iron and the arsenic, and 
 the lead button formed at the bottom of the fused mass contains 
 all the gold. 
 
 D. Tellurium Compounds. These must be treated by wet 
 methods. An aqua regia solution containing both gold and 
 tellurium is evaporated with a large excess of hydrochloric 
 acid until no more chlorine is given off, when both gold and 
 tellurium are readily precipitated by a current of sulphur dioxide 
 gas. On attacking the precipitate with nitric acid the tellurium 
 is dissolved in the state of tellurous acid, and the gold residue 
 may be dried and weighed, and its purity ascertained by 
 inquartation and parting. Other methods for separating gold 
 and tellurium will present themselves on studying the properties 
 of the latter. 
 
 Wet Methods of Assay of Gold Alloys, Compounds, 
 &c. Assays or complete analyses of gold bullion, natural 
 minerals, &c., can be made by the ordinary chemical methods 
 given by Fresenius, Crookes and others. From 1 to 5 grammes 
 of bullion are usually enough, but 10 grammes are necessary if 
 the alloy is nearly pure gold. In general the residue left after 
 prolonged action of nitric or sulphuric acid is not sufficiently 
 pure to weigh as gold, and complete solution in aqua regia is 
 usually necessary. From the solution the gold may be precipi- 
 tated by (a) ferrous sulphate, (b) sulphurous acid, (c) oxalic acid, 
 (d) sulphuretted hydrogen, (e) ammonic sulphide, followed by the 
 addition of hydrochloric acid. The following remarks may be 
 
 * Dingl. Polyt. Joum., vol. cxxxvii., p. 443. 
 
THE ASSAY OF GOLD BULLION. 489 
 
 of value in aiding the chemist in his choice of a precipitant in 
 any particular case. 
 
 Nitric acid must always be expelled from the solution by 
 warming with successive additions of hydrochloric acid. The 
 acid solution must not be heated too strongly or loss of gold 
 chloride by volatilisation occurs. Some other chlorides escape 
 more freely. Ferrous sulphate and sulphurous acid act well in 
 strongly acid (HC1) solutions ; oxalic acid, sulphuretted hydrogen 
 and ammonic sulphide best in presence of small quantities of 
 HC1.* The solution should be dilute (say 1 part of gold in 
 -300 of water), so that other metals may not be carried down by 
 the gold. Sulphate of iron gives a very finely divided precipitate 
 which is difficult to wash by decantation without loss ; precipita- 
 tion is slow in cold solutions. Oxalic acid causes plates and 
 .scales to form which are readily washed and are very pure ; it acts 
 best in boiling liquids, but a temperature of 80 for forty-eight 
 hours suffices ; in the cold or in the presence of much hydro- 
 chloric acid or alkaline chlorides the action is very slow and 
 partial \ a large excess of the precipitant must be present. Oxalic 
 acid is used for solutions containing metals of the platinum 
 group, which are not precipitated by it. Alkaline oxalates act 
 better than the free acid. 
 
 Sulphurous acid is an excellent precipitant for solutions from 
 which all other metals but gold have been removed. It acts 
 rapidly and completely in the cold, but unfortunately precipitates 
 many other metals. Sulphuretted hydrogen is used in the 
 absence of all other metals whose sulphides are insoluble in 
 hydrochloric acid. 
 
 In all cases careful consideration must be given to the nature 
 of the base metals present, and the precipitant which will not 
 render any of them insoluble must be selected. 
 
 Other Methods of Bullion Assay. Among methods which 
 have been proposed at various times, and which may still be of 
 service occasionally in particular cases, may be mentioned : 
 {!) The trial by the touchstone (a method formerly more exten- 
 sively used by jewellers than at present) ; the assay by means of 
 considerations as to (2) the colour, and (3) the density of alloys ; 
 {4) spectroscopic assay ; (5) assay by electrolysis, and (6) by the 
 induction balance. A brief description of each of these methods 
 is appended. 
 
 1. Trial by the Touchstone. This is the oldest method 
 of assay. It consists in rubbing the gold bullion to be tested on 
 a hard smooth stone, and comparing the appearance and colour 
 of the streak with those made by carefully prepared needles of 
 known composition. The effect of the action of nitric acid and 
 dilute aqua regia on these streaks is also noted. Touchstones 
 usually consist of Lydianstone or of silicified wood, and black or 
 
 H Rose. 
 
490 THE METALLURGY OF GOLD. 
 
 dark green stones are best. Only alloys of gold and copper or 
 of gold and silver can be thus tested. The trial is more sensitive 
 for alloys below 750 fine than for higher standards. The amount 
 of gold in alloys between 700 and 800 fine can be determined 
 correct to 5 parts per 1,000. 
 
 2. Colour and Hardness of Alloys. These properties 
 form a guide to the composition of copper-gold alloys, an increase 
 of copper corresponding to a heightening of the colour and an 
 increase of the hardness as tested by shears or a knife. On 
 heating the alloy to redness in air, the degree of blackening of 
 the surface is a further indication of the percentage composition, 
 if compared with plates of known fineness. 
 
 3. Density of G-old-copper Alloys. The determination of 
 the fineness of these alloys by taking their densities was investi- 
 gated by Professor Roberts- Austen at the Royal Mint in 1876.* 
 He showed that the densities found by experiment were nearly 
 equal to those obtained by calculation on the assumption that 
 the union of the two metals was accompanied neither by con- 
 traction nor expansion. The alloys examined ranged from 860 
 to 1000 fine, and were made into discs which were all compressed 
 to the same extent. The conclusion arrived at was that the 
 fineness of large masses of gold can be deduced from their 
 densities correct to TO Q 00 part. In the case of individual coins- 
 the results are only approximate. 
 
 4. Assay by means of the Spectroscope. This method 
 of determining the composition of gold-copper alloys was investi- 
 gated by Lockyer <fe Roberts-Austen, f The spectrum of pure 
 gold was shown to be altered by successive additions of copper, 
 and near the English standard (916-6 fine) a difference of two- 
 or three-tenths of a millieme in composition could be readily 
 detected, but the amount of metal volatilised is so small that it 
 cannot be made to represent with certainty the average com- 
 position of the mass, which is never perfectly homogeneous. 
 
 The detection of traces of gold in alloys or ores by means of 
 the spectroscope, though sometimes attempted, is not remarkable 
 for its delicacy or certainty. The method of procedure J is to 
 dissolve the auriferous material in aqua regia, evaporate off the 
 nitric acid, and pass induction sparks through the surface film of 
 liquid, when the spectrum shows some narrow bands and some 
 nebulous bands. The latter only are seen if a drop of the 
 solution is placed in a Bunsen flame. The method may some- 
 times be useful when complex minerals are being examined. 
 Mr. Capel has shown that -f^^ of a milligramme of gold will 
 show a spectrum if the spark be passed through a weak solution 
 
 * Seventh Annual Mint Report, 1876, p. 41. 
 
 -\-Phil. Trans. Roy. Soc., 164, part ii., p. 495, and Mint Report for 1874,. 
 p. 38. 
 
 Frgmy, Ency. Chim., T. in., L'or, p. 134. 
 
THE ASSAY OF GOLD BULLION. 491 
 
 of the pure metal. But when operating on a slip of alloy 
 formed of 
 
 Silver, 708 
 
 Copper, 254 
 
 Gold, 38 
 
 1000 
 
 the spectra of copper and silver alone were visible. In an alloy 
 of gold and copper containing from 200 to 250 parts in the- 
 thousand of the precious metal, the gold spectrum is barely 
 visible. On the other hand, in an alloy of gold and copper 
 containing traces only (-01 per cent.) of the latter, the copper 
 spectrum was distinctly shown. Alloys of gold are rarely so- 
 perfectly homogeneous that the particles of metal volatilised and 
 giving the spectrum represent the whole mass. 
 
 5. Assay by Electrolysis. Solutions of gold are readily 
 decomposed by low tension currents, a difference of potential of 
 one volt sufficing in the case of chloride of gold. Consequently, 
 gold can be separated thus in solutions of chlorides from copper, 
 lead, iron, &c. The process differs little from the electrolytic 
 assay of copper.* The special precautions to be observed are (a) to- 
 replace the positive platinum pole by one of carbon; (b) to cover 
 the platinum cone (negative pole) with a thin deposit of silver or 
 copper, so as to render the gold easily detachable ; (c) to avoid 
 an excess of HC1 or HNO 3 , while having a few drops of free 
 H 2 SO 4 present. The precipitation is effected at 50 with one 
 Bunsen cell ; the gold is washed, detached from the cone by 
 nitric acid, and weighed with the usual precautions. 
 
 6. Assay by the Induction Balance. Full descriptions of 
 the instrument employed may be found in the published papers 
 of Professors Hughes and Roberts- Austen, f but the principle on 
 which it depends may be briefly stated as follows : The balance 
 of two rapidly intermittent induced currents of equal strength, 
 flowing in two coils connected by a wire, is disturbed by the 
 presence of metal within one of the coils, and the fact of this 
 disturbance is made evident by a telephone. The balance can 
 be restored and the telephone silenced by introducing into the 
 opposing coil an identical mass of metal, and it is found that 
 widely different effects are produced by equal volumes of different 
 metals and alloys. In testing discs of the gold-copper alloys, 
 the balance is maintained by a tapered and graduated rod of 
 zinc, which is moved in and out of one coil ; this method, 
 however, is not adapted for examining alloys which differ but- 
 slightly in composition, as is shown by the following table : 
 
 * Crookes's Select Methods. 
 
 tProc. Roy. Soc., vol. xxix. (1879), p. 56; Phil. Mag. [5], vol. viiL 
 (1879), p. 50 ; and Proc. Phys. Soc., vol. iii. (1879), p. 81. 
 
492 
 
 THE METALLURGY OF GOLD. 
 
 Standard by 
 Assay. 
 
 Readings on 
 Zinc Scale. 
 
 Standard by 
 Assay. 
 
 Readings on 
 Zinc Scale. 
 
 1000-0 
 
 200 
 
 9137 
 
 66 
 
 990-9 
 
 138 
 
 911-8 
 
 66 
 
 950-8 
 
 78 
 
 901-3 
 
 65 
 
 925-0 
 
 70 
 
 878-0 
 
 64 
 
 920-3 
 
 69 
 
 859-2 
 
 62 
 
 918-1 
 
 67 
 
 709-2 
 
 60 
 
 916-3 
 
 66 
 
 
 
 By a different arrangement* the instrument may be made 
 more sensitive, but its indications are not trustworthy, as the 
 presence of traces of impurity in. the alloy and even its physical 
 condition have great influence on the readings obtained, which 
 depend on electrical resistance. The instrument is not used in 
 practice although it is capable of distinguishing between most 
 counterfeit and genuine coins. 
 
 Auxiliary Balance for Weighing Cornets. Mr. Robert 
 Law of the Melbourne Mint has invented f an. auxiliary balance 
 which gives the approximate weight of the cornets without the 
 use of any weights. The balance has a pan to hold the cornet 
 .attached to one end of the beam, and the other end is prolonged 
 as a pointer, the position of which (when it comes to rest), with 
 reference to a curved scale, denotes the percentage of gold 
 present. The cornet is then transferred to the pan of an 
 ordinary balance, and the necessary weights, already determined 
 by the auxiliary balance within narrow limits, are placed on the 
 other pan. For weighing large numbers of cornets ot widely 
 differing weights, this balance is a great convenience, saving 
 both time and wear of the balance and weights. It has been in 
 use lor four years at the Melbourne Mint. 
 
 * Tenth Report of the Royal Mint, 1880, p. 47. 
 t Chem. Soc. Journ., May 1896, p. 526. 
 
ECONOMIC CONSIDERATIONS. 49$ 
 
 CHAPTER XXI. 
 ECONOMIC CONSIDERATIONS. 
 
 Management of Gold Mills. The cost of extracting gold 
 from its ores in the mill is chiefly of interest to the producer 
 when it is combined with the cost of mining. Nevertheless, it. 
 is convenient that the two items should be ascertained separately. 
 For this purpose, each truck load of ore delivered from the mine 
 to the mill, and the amount actually treated in the latter, should 
 be weighed carefully. Materials, implements, &c., should not be 
 served out to the mill from the stores without an exactly-worded 
 order. The cost of transport of the ore from mine to mill is 
 often ascertained separately, but if this is not done it is better 
 to include it in the cost of mining. The cost of superintendence 
 must usually be distributed between the various operations, and 
 this course must also be sometimes resorted to in the case of 
 power, lighting, and other items. When the total expenses of 
 milling are ascertained for any period, say for one month, it is 
 necessary to allow for depreciation of the plant each time. It is- 
 also advisable not to neglect the question of interest on the 
 capital sunk in providing the mill, and of a sinking fund, as the 
 ore may come to an end before the machinery is discarded for 
 other reasons. 
 
 Similarly, exactness in ascertaining the percentage of gold 
 extracted is in the highest degree desirable. Each load of ore 
 on its way to the mill should be automatically sampled, and 
 frequent assays made on mixtures of these samples, the value of 
 the tailings being determined with equal care. From the results 
 thus obtained, not only is a watch kept on the relative success 
 of the treatment from day to day, but an additional check is 
 afforded on the amount of bullion produced in the mill. A 
 well-appointed assay office in connection with the works is 
 obviously necessary in order to carry out these tests, and 
 laboratory extraction trials should also be made at frequent 
 intervals in order to determine how far the efficiency of the mill 
 is being maintained. 
 
 Details concerning the cost of treating gold ores by the 
 various processes have already been given separately in the 
 chapters respectively devoted to them. 
 
 Cost of Production of Gold. In view of the fact that the 
 value of money is measured in almost all gold-producing coun- 
 tries by the metal itself, it would be of special interest to- 
 
494 THE METALLURGY OF GOLD. 
 
 -estimate the average cost per ounce of its production. This was 
 done by Prof. Roberts- Austen in the case of silver in the year 
 1887,* and all subsequent computations have served to show 
 that a high degree of accuracy was attained by him. In the 
 case of gold, further difficulties are encountered in the en- 
 deavour to frame a trustworthy estimate, owing to the 
 differences in the mode of treating silver and gold ores. Thus 
 the greater part of the silver produced annually is derived from 
 the output of large mills or smelting establishments, where the 
 total cost of treatment is well known to the managers, even 
 though they withhold it from the public. On the other hand, 
 .a large amount of gold is even now extracted in small mills or 
 by individuals, particularly in the case of placer deposits, and 
 the exact cost is frequently a matter of doubt to the proprietors 
 themselves. Moreover, both silver and gold mining companies 
 are usually reticent as to their costs, although the advantages of 
 this course of action to the proprietors, as distinguished from 
 the managers, are not easy to understand. By certain large 
 companies, systematic accounts are published, and from these 
 the following results are extracted : 
 
 At the Alaska Treadwell Company's mill about 20,000 tons 
 of ore, containing 3 dwts. of gold per ton, are treated monthly 
 by crushing and amalgamation, followed by concentration and 
 chlorination of the sulphides. The cost (including both mining 
 .and milling) per ounce of gold extracted was 2 Is. Id. in 
 1890-9], and 2 4s. 8d. in 1891-92. (The value of 1 ounce of 
 pure gold is 4 4s. lljd.) At the El Callao mine, Venezuela, 
 the cost per ounce was formerly as low as 25s., but rose to 
 3 10s. 9d. in 1891, and 3 Is. 8d. in 1892. The cost to the 
 Montana Company was 3 10s. 6d. per ounce in 1891, and as 
 much as 4 8s. in 1892. In some of the large hydraulicking 
 companies the cost was similar to that of the Alaska Treadwell 
 mine. Thus, at the La Grange Company's workings in 1874-76, 
 the cost was 2 4s. 3d. per fine ounce of gold, and at the New 
 Bloomfield Company's Claim the cost was 3 Os. lOd. in 1875, 
 .and 2 Is. in 1876 and in 1877. The above companies, how- 
 ever, were thriving, and the average cost per ounce in America, 
 even at the mines and mills which pay their way, is probably 
 not less than 3. The average cost of production in California 
 is estimated by A. G. Charleton f at about 2 2s. 6d. per ounce. 
 In South Africa, on the Witwatersrand, the cost of production 
 has been greatly reduced during the last few years. In 1892 
 the cost of production per fine ounce of gold at those mills 
 which were working more or less continuously seems to have 
 been about 3 10s., and in 1895, for the mines situated on the 
 -outcrop of the Main Beef, Hatch and Chalmers estimate it at 
 
 * Nineteenth Annual Report of the Royal Mint, 1888, p. 56. 
 t Trans. Fed. Inst. Mng. Eng., 1893, p. 217. 
 
ECONOMIC CONSIDERATIONS. 495 
 
 about 2 14s. At Barberton, the cost of gold obtained at the 
 Sheba mine is said to be 1 4s. per ounce. In Mysore, where 
 fuel is costly and the climate bad, the cost is said to be about 
 1 18s. lOd. per ounce of gold extracted.* At the placer mine 
 at Saint-Elie in French Guiana, which produces about 1,600 
 ounces of gold per month, the cost of production was 2 Is. 6d. 
 in 1887, and at the Siberian placers the cost is 1 17s. 2d. at 
 Berezovsk,. 1 17s. 8d. at Nijni-Tassnil, and 2 15s. 8d. at 
 Tchernaia-Retchka. 
 
 Judging from the published results generally, it would appear 
 that the average cost of production of gold is somewhat less than 
 3 per ounce. It must be remembered, however, that the interest 
 on the initial cost of discovering and developing the mine, and 
 of the machinery used in both mining and milling is, in many 
 cases, not included in the so-called cost of production. It has, 
 moreover, been frequently pointed out, particularly by Mr. A. 
 Del Mar, that if the working costs of unsuccessful mines, and the 
 money spent in fruitless prospecting, and in the erection of mills 
 unsuitable to the ore to be treated, were to be included in the 
 cost of the production of the gold, it would doubtless appear 
 that a heavy loss is annually experienced by the gold-mining 
 industry. It has been well observed that, no matter what the 
 market prices of gold, silver, copper, &c., may be, mines and ores 
 will be worked which leave no profit, in the never-failing expec- 
 tation that the market will rise again, or that the ore will 
 become more abundant or richer. 
 
 The " market price " of gold at any one time must, of course, 
 be derived from the mean price in gold of all, or at any rate a 
 large number of, other commodities, when compared with this 
 mean price at some other time. 
 
 Annual Production of the Gold Mines of the World. 
 The production of gold in ancient times cannot be closely esti- 
 mated, but, judged from a modern standpoint, it was probably 
 very small. In the middle ages, however, between the fall of 
 Rome and the discovery of America, the production was far 
 smaller than before, and Jacob observes that in this period 
 " the precious metals were sought not by exploring the bowels 
 of the earth, but by the more summary process of conquest, 
 tribute, and plunder." Even after the exploitation of the New 
 World began, the output of gold was for many years much too 
 small to satisfy the cupidity of the conquerors. The develop- 
 ment of the mining industry was prevented by the ruin and 
 destruction of the natives, and by the almost incessant irregular 
 warfare waged against the Spaniards in America in the 16th 
 century, first by the Dutch and later by the English. Fifty years 
 elapsed after Columbus discovered America, before the annual 
 production of gold reached 1,000,000, and even at the end of 
 * Ibid., p. 218. 
 
496 THE METALLURGY OF GOLD. 
 
 the 17th century Soetbeer estimates that it was only 1,500,000, 
 The discovery and working of the rich Brazilian placers during 
 the next half century raised the annual product to over 3,500,000' 
 in the period 1740-1760 (Soetbeer), but as these deposits became 
 exhausted, the output again fell off, and in the period 1810-1820 
 had again sunk to about 1,500,000 per annum. The gradual 
 development of the Siberian placers was the main cause of the 
 subsequent steady increase in production up to an average of 
 7,500,000 per annum in the period 1841-1850, and this was 
 followed by a sudden rise consequent on the discoveries in 
 California and Australia. The maximum output from the rich 
 placers of these countries was reached in 1853, when the world's 
 production of gold is estimated by Sir Hector Hay to have been 
 38,000,000. After tailing to 21,000,000 in 1863 (Hay), the 
 output remained nearly stationary until about the year 1888, 
 when, from various causes mentioned below, the production 
 again began to increase, and in 1897 reached 49,COO,000 the 
 greatest amount on record. 
 
 The doubling ot the world's production which has taken place 
 in the course of the last ten years is due (1) to the discovery 
 and development of new districts, (2) to the progress in the 
 art of metallurgy, and (3) to the increased attention given to gold 
 mining consequent on the fall in the price of silver. This iall 
 administered a severe check to the silver mining industry arid 
 set free a considerable amount of skilled labour and of capital 
 which were largely diverted to gold mining. The most important 
 new districts which have been developed are the Rand, Cripple 
 Creek in Colorado, and Western Australia, The output of 
 gold at the Rand was valued at 81,000 in 1887, and 10,430,000 
 in 1897; Cripple Creek had not been discovered in 1887, and 
 produced about 2,500,000 in 1897, while Western Australia 
 produced 20,000 in gold in 1887, and 2,400,000 in 1897. The 
 total increase in annual output on these three fields was there- 
 fore about 15,250,000 between 1887 and 1897. 
 
 The increase of production due to improved metallurgical 
 methods is difficult to estimate. Most of it must certainly be 
 put down to the account of the cyanide process by which gold 
 valued at about 3,000,000 was produced in 1895, though 
 the greater part of this was produced on the Rand, and is 
 included in the amount already given. The increase of about 
 1,500,000 which took place in the Russian output between 
 1888 and 1897 is partly due to improvements in the methods 
 of washing the gravels, and partly to the extension of the 
 Siberian railway and the greater attention now being paid 
 to vein mining. The third cause of the increase in output 
 of gold has been more felt in the United States and Mexico 
 than elsewhere. Of the increase in the output of the United 
 States from little more than 6,000,000 in 1887 to over 
 
ECONOMIC CONSIDERATIONS. 497 
 
 11,400,000 in 1897, about half is attributable to the output 
 from the new field at Cripple Creek as already noted, and the 
 remainder is mainly due to the changes in the conditions of 
 silver mining and their ultimate effects. The increase of about 
 1,300,000 in Mexico is also to be ascribed to the last-named 
 cause. The output of Australasia (excluding Western Australia) 
 was 2,000,000 more in 1897 than in 1887, and although local 
 causes were largely responsible for this increase, as, for example, 
 the encouragement given by the government of New South 
 Wales in 1894 and 1895 to "fossicking," nevertheless improved 
 methods of extraction were also of great effect. 
 
 Speaking generally, it is clear from the above considerations 
 that of the increase in the annual output of gold of some 
 28,000,000 which took place between 1887 and 1897, about 
 half is due to the discovery of new fields, and the other half 
 to the development of districts already known. The credit for 
 having caused the latter part of the increase may be about 
 equally divided between improvements in metallurgical pro- 
 cesses and the greater attention which has been given to gold 
 mining during the last few years. 
 
 As regards the future production, it is obviously impossible 
 to predict what will be the effect of the opening* up of gold 
 fields as yet undiscovered. Even without these, however, it 
 seems almost certain that the production will continue to rise 
 rapidly for some years to come. The most conservative 
 estimates place the output of the Eand in three years time at 
 from 12,000,000 to 14,000,000 per annum, and when the 
 labour problem has been fully solved, even this huge output 
 might conceivably be doubled soon afterwards. The output 
 in Western Australia may similarly be expected to increase 
 largely when, if ever, the difficulties due to lack of water have 
 been overcome. Siberia, when the transcontinental railway is. 
 completed, may also contribute much more than at present to 
 the world's production, and the improvements in methods of 
 extraction which have been so marked of late years will doubt- 
 less help to swell the total. It is, therefore, no exaggeration to 
 say that present indications point to an annual production of 
 gold of not less than 55,000,000 to 60,000,000 in the year 
 1900. 
 
 The following table gives details of the production of gold in 
 various countries in the period 1891-95, and the year 1896 : 
 
 32 
 
498 
 
 THE METALLURGY OP GOLD. 
 
 of T 
 ion. 
 
 lO iO 
 
 7*^^7*^cogjpi^oi>.r~GiTt<p<p'7-iTrcpr-i.--ii i^OOOpOOO 
 
 OOiCOfMeNOCOtfaoO-^Co' 
 
 '* to <M 
 
 'OO 
 
 i>c^ t-co i^io os co -^ -( 
 
 s 
 
 5[o t^eo^l cot^cofNoot^^t .^^ _ 
 oT co" uT o" oT oo" t-T TJT ^ co co co ci c<f -* r-i 
 
 O OS CO !>!> 
 
 cocoS? 53 
 
 CO ^H 
 C^^^ 
 
 Oi Oi CO CO t^ GO ^* C^ ^ ^H p-^ -^ OO IO ^O 00 O O ^H 
 if5CO^^'!fCOCO^COCOt--COtOii C^^OOr ( 
 
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 2 'i 
 
 .2.S a 
 
 o3 .2 . o- 
 
 = 
 
ECONOMIC CONSIDERATIONS. 
 
 499 
 
 The production of gold in 1897 is estimated by the Engineering 
 -and Mining Journal as follows : 
 
 Name of Country. 
 
 Fine Ozs. 
 
 Value. 
 
 Percentage of 
 Total. 
 
 Africa, .... 
 
 2,712,580 
 
 11,522,800 
 
 23-25 
 
 United States, . . 
 
 2,685,000 
 
 11,405,650 
 
 23-02 
 
 Australasia, . 
 
 2,462,863 
 
 10,454,700 
 
 21-07 
 
 Russia, .... 
 
 1,572,248 
 
 6,720,600 
 
 13-54 
 
 British India, 
 
 346,363 
 
 1,471,300 
 
 2-96 
 
 Mexico, .... 
 
 328,969 
 
 1,400,000 
 
 2-82 
 
 Canada, 
 
 290,259 
 
 1,235,300 
 
 2-49 
 
 Other Countries (mostly 
 
 
 
 
 assumed to be the same 
 
 1,279,643 
 
 5,392,500 
 
 10-85 
 
 as in 1896), 
 
 
 
 
 Total, . 
 
 11,677,925 
 
 49,602,850 
 
 100-00 
 
 The product of each of the Colonies of Australasia was as 
 follows : 
 
 
 Yearly Average for the 
 period 1891-95. 
 
 Year 1896. 
 
 Victoria, 
 Queensland, 
 New South Wales, 
 Western Australia, 
 New Zealand, 
 Tasmania, . 
 South Australia, 
 
 67 1,804 ozs. 
 616,439 
 234,889 
 137,879 
 246,261 
 49,514 
 36,936 
 
 805,087 ozs. 
 638,000 
 296,071 
 281,265 
 263,722 
 62,586 
 29,004 
 
 Total, . 
 
 1,993,722 
 
 2,375,735 
 
 The product for the individual States in the U.S.A. was 
 follows : 
 
 
 Yearly Average for the 
 period 1891-95. 
 
 Year 1896. 
 
 California, . 
 
 630,764 fine ozs. 
 
 737,036 fine ozs. 
 
 Colorado, . 
 
 391,920 
 
 719,264 
 
 South Dakota, 
 
 178,300 
 
 237,978 
 
 Montana, . 
 
 165,573 
 
 209.207 , 
 
 Arizona, . . 
 
 69,536 
 
 124,770 
 
 Nevada, 
 
 70,355 
 
 116,620 
 
 Idaho, 
 
 86,218 
 
 101,263 
 
 Alaska, 
 
 54,569 , 
 
 99,444 , 
 
 Utah, . 
 
 42,625 
 
 91,908 , 
 
 Oregon, 
 
 67,642 
 
 59,313 
 
 Eleven other States, . 
 
 69,024 , 
 
 68,340 
 
 Total, . 
 
 1,826,526 
 
 2,568,132 
 
500 
 
 THE METALLURGY OF GOLD. 
 
 The output in Africa in 1896 may be divided as follows : 
 
 Witwatersrand, 
 De Kaap, 
 Klerksdorp, 
 Zoutpansberg, 
 Lydenberg, 
 
 Other districts in the Transvaal 
 Gold fields not in the Transvaal (csti 
 mated), 
 
 Total, 
 
 1,847,523 fine ozs. 
 
 98,326 
 
 31,443 
 
 4,538 
 
 40,813 
 
 694 
 
 100,000 
 
 2,123,33: 
 
 In 1895, the product of the Witwatersrand field was 2,278,110' 
 ozs., valued at 7,837,779. To obtain this amount, 3,456,575 
 tons of ore were crushed, the average number of stamps at work 
 being about 2,600 ; the yield on the plates was 8-86 dwts. per 
 ton, from the concentrates the yield was 0'61 dwts., and from 
 the tailings by the use of cyanide 3*70 dwts. per ton of the 
 original ore, making a total of 13-17 dwts. of gold per ton. 
 
 In the table given on p. 498, the figures are for the most part 
 taken from the tables given by the Director of the United States 
 Mint, and published in his Annual Reports on the Production of 
 Gold and Silver. The output of Australasia in both years is 
 that given in the government reports of the separate colonies. 
 The African product in each year consists of the returns for the 
 Transvaal made by the Witwatersrand Chamber of Mines, to- 
 gether with a rough estimate of the gold produced in the rest of 
 the Continent. The product of most of the countries in 1896 
 is that given in the Mineral Industry for 1896. The last two 
 columns giving the percentages have been added by the author. 
 
 Without attempting to make an exact estimate of the amount 
 of gold won by each metallurgical method, it is obvious from 
 the fact that almost all the Russian product, about one-fourth 
 of that from Australasia, and probably over one-fourth of that 
 from the United States, is derived from placer washings, that 
 these deposits yield little less than 30 per cent, of the world's 
 product. The yield of gold from quartz crushing is now far 
 more than that obtained from placers, although, up to a few 
 years ago, it had always been less. Gold obtained by smelting 
 (chiefly in the United States) may be about 3 per cent, of 
 the total produce of the world, while that by the wet 
 methods (viz., treatment by chlorine and by cyanide of 
 potassium) is certainly over 10 per cent., the reagent last 
 named being instrumental in obtaining about two-thirds of 
 this amount. It is probable that over nine-tenths of the gold 
 is won by methods best suited to the ores dealt with in each 
 case, taking into account the conditions prevailing in the 
 district. Alterations in these conditions, such as the improve- 
 
ECONOMIC CONSIDERATIONS. 501 
 
 merit of transport or changes in the cost of labour or materials, 
 may render some other method preferable to the one in use. 
 The great problem has long been to find a cheap method of 
 treating large quantities of low-grade pyritic ores which do not 
 readily yield their gold to the action of mercury. Such ores 
 -can usually be treated by chlorination or by smelting, but the 
 expense of either method is in many cases too great to afford a 
 profit. It seems quite possible that the cyanide process will 
 come into general use for these ores. 
 
 Consumption of Gold. Many attempts have been made at 
 various times to estimate the amount of gold used annually 
 for purposes other than that of additional coinage. In 1831, 
 Jacob * estimated the annual amount of gold " converted into 
 utensils and ornaments" at about 4,500,000, while the pro- 
 duction was under 2,000,000, without counting, however, 
 imports from India and China, which were supposed to be 
 considerable. In 1881, Dr. Soetbeer estimated the annual 
 industrial consumption of gold in the world, after making 
 deductions for old material employed, as amounting to 84,000 
 kilogrammes or 11.500,000, and in 1885 he put it at 90,000 
 kilogrammes or 12,300,000, against a production of 21,000,000. 
 In 1891, the same authority gave the industrial consumption of 
 gold added to the amount hoarded and that exported to the 
 East as 120,000 kilogrammes or 16,400,000, against a pro- 
 duction of 24,000,000. In 1894, Ottomar Haupt estimated 
 the industrial consumption at only 280,000,000 francs or 
 11,200,000, but the exports to the East are not included in 
 this amount. In the same year the Director of the United States 
 mint f estimated the industrial consumption at $50,177,300 or 
 about 10,300,000, and for the year 1895 at $58,579,160 or 
 about 12,050,000. Doubtless the industrial consumption of 
 gold is increasing in amount with the increase of population 
 and wealth, but nevertheless it is fair to assume that no such 
 rapid growth has taken place of late years in consumption as 
 has been already pointed out to have taken place in production. 
 
 Taking the annual consumption of gold, including that exported 
 to the East, as being, on the basis of the estimates given above, 
 from 10,000,000 to 15,000,000, it would appear that the 
 amount of gold added to the world's coinage, or retained as 
 reserves in the form of bullion, was from 5,000,000 to 
 10,000,000 per annum previous to the year 1890, and has 
 been gradually increasing ever since, so that in 1895 it was 
 probably over 25,000,000. Now, in 1894, it was estimated 
 that the total available stock of gold in the world was 
 $3,965,900,000 or about 800,000,000, so that the annual 
 
 * History of the Precious Metals, vol., ii., p. 322. 
 
 t Production of the Precious Metals, 1893, p. 53. Report of the Director 
 of the Mint, 1896, p. 58. 
 
502 THE METALLURGY OF GOLD. 
 
 addition to this stock is now probably at the rate of between 
 3 and 4 per cent, per annum, and may be expected to increase 
 gradually for some time to come. With regard to the effect 
 which this increase may have on prices and on commercial 
 prosperity, it is to be observed that in the years succeeding 
 1848, when the available stock of gold is estimated to have been 
 less than 300,000,000, an annual addition (exclusive of the 
 industrial consumption) was made to the available stock of about 
 20,000,000, or 6'7 per cent, of the total amount. This addition 
 is usually supposed to have stimulated trade and inaugurated a 
 long period of commercial prosperity. Jevons estimated that 
 the purchasing power of gold was depreciated by from 9 to 15 
 per cent., and that it was only kept from falling lower by the 
 absorption of vast quantities of gold for use in currency by 
 several countries which had previously mainly employed silver 
 or paper. Similar absorption is now going on at a very rapid 
 rate, the amount of gold held by the Russian and Austrian 
 national banks, for example, being ,40,000,000 more in October 
 1897, than in 1896, while the holdings of the other banks have 
 not fallen. The result is, that in spite of the enormous quantity 
 of gold now being produced, its value has not greatly altered. 
 
BIBLIOGKAPHY. 
 
 IT has been found impracticable to enumerate the articles and paragraphs 
 relating to the metallurgy of gold, which have from time to time appeared 
 in the various periodicals, as very little search results in the accumulation 
 of thousands of such references. In general, therefore, only the names 
 of some of the publications which contain important or interesting matter 
 on the subject are given ; but a few exact references on special points 
 have been added, and many others occur in the footnotes to the text. 
 
 PERIODICAL LITERATURE. 
 
 Anales de la Mineria Mexicana, 6 sea : Revista de Minos. Mexico. From 
 
 1861. 
 
 Analex de Minas. Madrid. From 1841. 
 Annalen der Berg-und Huttenkunde. Salzburg, 1802-5. 
 Annales des Mines. Paris. From 1816. 
 Annuaire du Journal des Mines de Russie. St. Petersburg. First published 
 
 in 1840. 
 Annuaire des Mines et de la Metallurgie Francises. Paris. First published 
 
 in 1876. 
 
 Annual Reports of the Bnllarat School of Mines. From 1882. 
 Annual Reports of the Califomian State Mineralogist. Sacramento. From 
 
 1881. ' 
 Annunf Reports of the Deputy -Master of the Royal Mint. London. From 
 
 1870. 
 
 Annual Reports of the Director of the. United States Mint. Washington. 
 Annual Reports on Gold Mining. Victoria, British Columbia. From 1875. 
 Berg-und Huttenmann inches Jahrbuch. Vienna. From 1866. 
 Berg-und Huttenmanxische Zeitung. Freiberg. From 1842 
 Biennial Reports of the Nevnda State Mineralogist. From 1871. 
 Boletin oficial de minas Madrid, 1844-5. 
 Bulletin de I' Association amicale des anciens elSves de I'tfcole des Mines. 
 
 Paris. From 1869. 
 
 Dingier' s Polytechnisches Journal. From 1815. 
 Engineering and Mining Journal. New York. From 1866. 
 Jahrbuch fur Berg-und Hiittenwesen . Freiberg. From 1827. 
 Journal des Mines de Freiberg. Koehler. Freiberg, 1788-1793. 
 Journal des Mines de Russie. 1832 to 1835. 
 La Mineria. Mexico, 1843. 
 
 Mining and Scientific Press. San Francisco. From 1864. 
 Mining and Smelting Magazine. London, 1862-65. 
 Minmg Journal. London. From 1836. 
 Mining Review. Denver, Colorado, 1873-76. 
 Mining World. London. From 1871. 
 Nouvpau Journal des Mines de Freiberg. Koehler & Hoffmann. Freiberg, 
 
 1795-1804. 
 
504 BIBLIOGRAPHY. 
 
 Oesterreich Zeitschriftfiir Berg-und Huttenwesen. Otto Freihern. Vienna. 
 
 From 1853. 
 Precious Metals of the United States, Annual Reports on. Washington. 
 
 From 1880. 
 Reports of the Mining Commissioner of New Zealand. Wellington. New 
 
 Zealand. From 187 1. 
 
 Revista Minera y Metalurgica. Madrid. From 1850. 
 School of Mines Quarterly. Columbia, (J.S. From 1879. 
 Scientific American. New York. From 1846. 
 JSUliman's American Journal of Science and the Arts. New Haven and New 
 
 York. From 1816. 
 Transactions of the American Institute of Mining Engineers. Philadelphia 
 
 From 1871. 
 Proceedings of the Chemical and Metallurgical Society of South Africa. 
 
 Johannesberg. From 1894. 
 Transactions of the Institute of Mining and Metallurgy. London. From 
 
 1892. 
 
 GENERAL METALLURGY OF GOLD. 
 
 Oeber, the Works of. Translated by R. Russell. London, 1686. 
 
 Biringuccio. De la Pirotechnia. Venice, 1540. French translation, 
 Rouen, 1627. 
 
 Agricola (G-eorgius). De re metallica. Bale, 1556. 
 
 Michaelis (Johannis). De Oro. Leipzig, 1630. 
 
 Barba (Alphonzo). Arte de los metales. Madrid, 1639. French trans- 
 lations, Paris, 1751, and La Haye, 1782. 
 
 Schluter. Principles of Metallurgy and Assaying. Brunswick, 1738. 
 
 Vargas (Perez de). Traite singulier de metallurgie. Translated from 
 the Spanish. Paris, 1743. 
 
 Ijewis (Win.) Commercium Philosophico-Technicum. London, 1763. 
 
 Valerius. Grundriss der Metallurgie. Ulm, 1768. 
 
 Cramer. Principes de metallurgie et docimasie. Blankenburg, 1774. 
 
 Jars. Voyages metallurgiques. Paris, 1774-81. 
 
 Karsten. System der Metallurgie. Breslau, 1818. 
 
 Kiessling. Die Metallurgie. Dresden, 1841. 
 
 Ansted(D. T.) The Gold-Seekers' Manual. London, 1849. 
 
 Landrin (H.) Traite de 1'or. Paris, 1850 and 1863. 
 
 Rammelsberg. Lehrbuch der chemischen Metallurgie. 1850. 
 
 Phillips (J. A.) Gold Mining and Assaying. London, 1852. 
 
 Phillips (J. A.) Encyclopaedia metropolitana. Article on Metallurgy. 
 London, 1854. 
 
 Bivot (L.) Principes generaux du traitement des minerais metalliques. 
 Paris, 1859. 
 
 Crookes (Wm. ) and Rohrig (E.) Treatise on Practical Metallurgy, trans- 
 lated from the German. London, 1860. 
 
 Kerl (B.) Die Rammelsberger Hiittenprozesse. Clausthal, 1861. 
 
 Kustel (Or.) Processes of Gold and Silver Extraction. San Francisco, 
 1863. 
 
 Kerl (B.) Handbuch der metallurgischen Huttenkunde. Clausthal, 1865. 
 
 Phillips (J. A.) The Mining and Metallurgy of Gold and Silver. London, 
 1867. 
 
 Overman (F.) A Treatise on Metallurgy. New York, 1868. 
 
 Blake (W. P.) Report on the Precious Metals. Washington, 1869. 
 
 Baymond (R. W.) Mineral Resources West of the Rocky Mountains. 
 7 vols. Washington, 1869-74. 
 
 Baymond (R. W. ) Mines, Mills, and Furnaces of the Pacific States. New 
 York, 1871. 
 
 Schiern (F. ) Sur 1'origine de la tradition des f ourmis qui ramassent Tor. 
 Copenhagen, 1873. 
 
BIBLIOGRAPHY. 505 
 
 Greenwood (W. H.) Manual of Metallurgy, vol. ii. London, 1875. 
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 p. 1388, 1877. 
 
 Attwood (G.). The Batea, the Milling of Auriferous Veinstones, and tha 
 Mineralisation of Gold. Articles in Alto. California. San Francisco, 
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 Kerl (B. ) Grundriss der Metallhiittenkunde. Leipzig, 1880. 
 Simonin (L.) L'Or et I 1 Argent Paris, 1880. 
 Industrial Progress in Gold Mining. Philadelphia, 1880. 
 Percy (John). Metallurgy of Silver and Gold, vol. i. London, 1880. 
 Ryan (J.) Gold Mining in India. London. 1881. 
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 Egleston (T.) The Progress of the Metallurgy of Gold and Silver in the 
 
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BIBLIOGRAPHY. ^ssss^ssss?** 5Q7 
 
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508 BIBLIOGRAPHY. 
 
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 CHAPTERS IV. AND V. PLACER MINING. 
 
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BIBLIOGRAPHY. 509 
 
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 CHAPTERS VI., VII., AND VIII.-QUARTZ CRUSHING AND 
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 Somieschmied. Traite sur 1'amalgamation. Ronneburg, 1811. 
 Rivot. Nouveau proced6 de traitement des minerals d'or et d'argent (a. 
 
 manuscript in the archives of the l^cole des Mines de Paris). Paris, 
 
 1818. 
 
 Ortmann. Kurze Geschichte der Amalgamation in Sachsen. Freiberg, n.d. 
 Lawson (G.) Improvements in Amalgamation. Trans. Nova Scotia Inst*. 
 
 1866. 
 Hague (J. D.) Gold Mining in Colorado. Report on the Fortieth Parallel,. 
 
 vol. iii. Washington, 1870. 
 Keith (N. S.) Amalgamated Copper Plates. Eng. and Mng. Journ., vol. 
 
 xi., p. 270, 1871. 
 
 Blake (W. P.) Mining Machinery. New Haven, 1871. 
 Fonseca. Memoire sur 1'amalgamation Chilienne. Paris, 1872. 
 Bergmann (E. von). Die Anfange des Geldes in JSgypten. Vienna, 1872.. 
 Thompson (H. A.) Extraction of Gold. Trans. Roy. Soc. Viet., vol. viii., 
 
 pp. 15-26, 1872. 
 Skey (Wm.) Electromotive Power of Certain Metals in Cyanide of 
 
 Potassium with reference to Gold Milling. Trans. N.Z. Inst., vol. 
 
 viii., p. 334, 1876. 
 Attwood (G.) Chile Vein Gold Works, South America. Proc. Inst. C.E.,. 
 
 vollvi., p. 244, 1879. 
 Cumenge and Fuchs. Effect of Antimony and Arsenic on Amalgamation. 
 
 Comptea Rendus, March 17, 1879. 
 
 Sgleston(T.) Californian Stamp Mills. London, 1880. 
 Randall (P. M.) Quartz Operators' Handbook. New York, 1880. 
 Habermann (I.) The Heberle Mill. Oesterr. Ztschr. fur Berg, und 
 
 Htnwesen, 1880. 
 
 Egleston (T.) Treatment of Gold Quartz in California. London, 1881. 
 Egleston (T.) Losses in Amalgamation. Trans. Am. Inst. Mng. Eng., 
 
 vol. ix., p. 633, 1881. 
 Egleston (T.) Causes of Rustiness in Gold. Trans. Am. Inst. Mng. Eng.> 
 
 vol. ix., p. 646. New York, 1881. 
 Yale (Charles G.) Mining Machinery. Prod, of Gold and Silver in the 
 
 . U.S., 1881. 
 Richards (J. ) Quartz Crushing Machinery. Prod, of Gold and Silver in 
 
 U.S. Washington, 1881. 
 Attwood (G.) Milling of Gold Quartz. Prod, of Gold and Silver in U.S., 
 
 1881. 
 Reed (S. A.) Ore Sampling. School of Mines Quarterly, vol. iii., p. 253., 
 
 1881. 
 M'Dermott and Duffield. Gold Amalgamation and Concentration. London*. 
 
 and New York, 1890. 
 
.510 BIBLIOGRAPHY. 
 
 Lock (G. Warnford). Gold Amalgamation. Proc. Inst. Mng. and Met. t 
 
 Session ii. , 3rd meeting. London, Dec. , 1 892. 
 ^Curtis (A. Harper). Gold Quartz Reduction. Proc. Imt. C.E., vol. cviii. 
 
 (1892), partii. 
 Charleton (A. G.) Coarse and Fine Crushing. Trans. Fed. Inst. Mng. 
 
 En.g , 1892-93. Four Papers. 
 
 Louis (H.) Handbook of Gold-Milling. London, 1894. 
 Rickard (T. A.) Variations in Gold Milling. New York, 1895. 
 Rickard (T. A.) The Stamp Milling of Gold Ores. New York and London, 
 
 1897. 
 
 CHAPTER IX. CONCENTRATION OF GOLD ORES. 
 
 Rittinger (P. von). Lehrbuch der Aufbereitungskunde. Berlin, 1867, 
 
 1870, and 1873. 
 .Smyth (Sir W. W.) Dressing, or the Mechanical Preparation of Gold 
 
 Ores. Lectures on Gold. Mng. Journ., 1873. 
 
 Pull descriptions of the different kinds of round buddies are given in the 
 following papers : 
 
 Teague (W., Jun.) "On Dressing Tin Ores." Proc. Mining Inst. 
 
 Cornwall, vol. i., No. 3 (Truro^, 1877). 
 
 Ferguson (H. T.) "On the Mechanical Appliances used for Dress- 
 ing Tin and Copper Ores in Cornwall. " Proc. Inst. Meek. Eng. , 
 1873. 
 'Schmidt (A. W.) Der Schlamfanger auch Kornfanger genannt. Dillen- 
 
 burg, 1877. 
 Reytt (C. von). Comparison of the Hand Buddie and the Rotary and 
 
 Percussion Tables. Berg-u. Hiittenmann. Jahrbuch, vol. xxii., 1877. 
 Habermann (J. ) Comparison of the Salzburg Table, the Rittinger Table, 
 
 and the Hand Buddie. Oesterr. Zeitschr. fur Berg-und Huttenweaen, 
 
 1879, No. 8. 
 
 Cazin (F. M. P.) Dynamical Metallurgy or Mechanical Ore Concentra- 
 tion. Mining Record, 1881-2. 
 Richards (R. H.) A new Hydraulic Separator to prepare Ores for Jigging 
 
 and Table Work. Trans. Am. Imt. Mng. Eng., 1883. 
 Reytt (C. von). Salzburg Percussion Table. Berg-und Hilttenmanisches 
 
 Jahrbuch, xxxiii., p. 3, 1883. 
 Callon (J.) Lectures on Mining. Translated by Le Neve Foster and 
 
 W. Galloway, vol. iii. London, 1886. 
 The Settling of Solid Particles in Liquids. Bulletin No. 36. United States 
 
 Geological Survey. Washington, 1886. 
 Ore Dressing in California. Sixth Report of the Gal. State Min., 1886. 
 
 This is a full account of the machines actually at work. 
 Reytt (C. von). The Linkenbach Table and Hand-Buddie compared. 
 
 Oesterr. Ztch., Oct. 6, 1888. 
 
 Lock (C. G. W.) Mining and Ore Dressing Machinery. London, 1891. 
 Kunhardt (W. B.) The Practice of Ore Dressing in Europe. New York, 
 
 1891. 
 Scheidel (A.) New Process of Dressing the Gold Ores of the Thames 
 
 Valley. New Zealand Mng. Comm, Rent.. 1S91, p. 28. 
 Meier (J. W.) Concentration at Pribram, Bohemia. Eng. and Mng. 
 
 Journ., vol. liv. (1892), p. 665, vol. Iv. (1893), pp. 5, 28, and 52. 
 Commans. Concentration and Sizing of Crushed Ore. Proc. Inst. Civil 
 
 Eng., 1894. 
 Kosales (H. ) Report on the Loss of Gold in the reduction of Auriferous 
 
 Veinstone in Victoria. Melbourne, 1895. 
 Richards (J.) Article in the Mineral Industry for 1896. 
 
BIBLIOGRAPHY. 511 
 
 CHAPTER XL ROASTING OF GOLD ORES. 
 
 Plattner (C. P. ) Die metallurgische Rbstprozesse. Freiberg, 1856. 
 Dixon (Wm.) Methods of extracting Gold, Silver, and other Metals from 
 
 Pyrites. CJiemical News, vol. xxxviii., pp. 281, 293, and 301 ; and 
 
 vol. xxxix., p. 7. 1879. 
 
 Kustel (G.) Roasting of Gold and Silver Ores. San Francisco, 1880. 
 Cumenge (E.) Note relative a 1'emploi de la vapeur d'eau dans certainea 
 
 operations metallurgiques. Annales des mines, vol. i. , p. 1852. Paris, 
 
 1882. 
 Stetefeldt (C. A.) On Salt Roasting. Trans. Am. Inst. Mng. Mug., 
 
 vol. xiii. New Haven, 1885. 
 Christy (S. B.) Losses of Gold in Roasting. Trail*. Am. Inst. Mng. Eng. t 
 
 1888. 
 Producer Gas for Roasting. Tenth Cal. State Min. Report, p. 897. San 
 
 Francisco, 1890. 
 Adams ( W. H. ) Pyrites : Practical Methods for Extraction of Gold, &c. 
 
 New York, 1892. 
 Stetefeldt (C. A.) Taylor's Gas Producer for Roasting. Eng. and Mng. 
 
 Journ., vol. Ivi. (1893), p. 124. 
 
 CHAPTERS XII., XIII., AND XIV. CHLORINATION. 
 
 The earliest researches on the chlorination of gold were made by 
 
 (1) Duflos (Dr.) Die schles. gesell. Uebersicht. Breslau, 1848. 
 
 (2) Richter (Theo.) Journ. fUr Prak. Chem., vol. li. (1849), p. 151. 
 
 (3) Lange (Herr). Karsten's Archiv, vol. xxiv., pp. 396-429. 1852. 
 
 (4) Plattner (C. P.) Probirkunst, p. 570. 1853. 
 
 (5) Percy (J.) Phil. Mag., vol. xxxvi. (1853), pp. 1-8. 
 
 Whelpley and Storer. Method of separating Metals from Sulphurets. 
 
 Boston, 1866. 
 
 Kustsl (G.) Concentration and Chlorination. San Francisco, 1868. 
 Bleasdale (J. J.) On Chlorine as a Solvent for Gold. Trans. Roy. Soc. 
 
 Victoria, vol. vi., pp. 47-52. 1870. 
 
 Pyrites : Report of the Board appointed to report on the methods of treat- 
 ing Pyrites and Pyritous Veinstuff, as practised on the Gold Fields. 
 
 Melbourne, 1874. 
 Attwood (G. M.), and Aaron (C. H.) Prod, of Gold and Silver in the 
 
 United States. Washington, 1881. 
 Prance (Ch. de). Extraction par voiehumide du cuivre, de 1'argent, et de 
 
 Tor. Brussels, 1882. 
 Egleston (T. ) Leaching Gold and Silver Ores in the West. New York, 
 
 1883. 
 
 Egleston (T.) Leaching Gold Ores containing Silver. London, 1886. 
 Stetefeldt (C. A.) Lixiviation of Silver Ores. New York, 1888. This 
 
 work gives many details applicable to all wet processes. 
 O'Driscoll (P.) Notes on the Treatment of Gold Ores. London, 1889. 
 The Pollok Process of Chlorination. Eng. and Mng. Journ., vol. xlix. 
 
 (1890), Feb. 15. 
 Burfeind (J. H.) Chlorination of Gold Ores. Eng. and Mng. Journ., vol. 
 
 li. (1891), p. 446. 
 Precipitation of Gold from Chloride Solution. Letters and Articles by John 
 
 E. Rothwell, W. Langguth, C. H. Aaron, L. D. Godshall, J. T. Blom- 
 
 field, and T. K. Rose in Eng. and Mng. Journ., vol. li. (1891), pp. 74, 
 
 112, 165, 204, 229, 282, 347, 373, 465; vol. lii. (1891), p. 211; and 
 
 Mining Journal, March, 18, 1893. 
 
512 BIBLIOGRAPHY. 
 
 Vautin(C.) Decomposition of Auric Chloride. Proc. List. Mining and 
 
 Met., Session ii., 5th Meeting. London, Feb. 15, 1892. 
 Langguth (W.) Chlorination of Gold. Trans. Am. In*t. Mng. Eng. 
 
 June, 1892. 
 Barrel Chlorination. Eng. and Mng. Journ., vol. Iv. (1893), pp. 244 and 
 
 269. 
 Rothw a ll(J. E.) Recent improvements in Chlorination. Mineral Industry 
 
 for 1892, p. 236. New York, 1893. 
 Godshall (L. D.) Modern Chlorination. Eng. and Mng. Journ., vol. 
 
 Ivii. (1894), Jan. 6 and 13. 
 
 Wilson (E. B. ) The Chlorination Process. London and New York, 1897. 
 Rothwell (John E. ) The present development of the Barrel Culorination, 
 
 Process. Mineral Industry for 1896, pp. 261-284. 
 
 CHAPTERS XV. AND XVI. -CYANIDE PROCESS. 
 
 Scheidel ( A. ) The Cyanide Process. Sacramento, 1894. 
 Janin, Jr. (L.) Article in Mineral Industry for 1892, pp. 239-270. 
 Reunert (T.) Diamonds and Gold in South Africa. London, 1894. 
 Butters (C.) and Smart (E.) Plant for the extraction of Gold by the 
 
 Cyanide Process. Proc. I nut. Civil Eng., vol. cxx. 1895. 
 Eissler (M.) The Cyanide Process. London, 1895. 
 Park (Jas.) The Cyanide Process of GoLl Extraction. Auckland and. 
 
 Melbourne, 1896. 
 Wilson (E. B.) The Cyanide Process. New York and London, 1896. 
 
 CHAPTER XVII. PYRITIC SMELTING. 
 
 Austin (W. L.) Matting Dry Auriferous Sulphides. Trans. Am. Inst~ 
 Mng. Eng., 1889. 
 
 The Austin System of Pyritic Smelting. Denver, Colo., 1892. 
 
 Pyritic Smelting. Articles and letters on the subject have appeared in the 
 Engineering and Mining Journal under the following dates : Vol. 1. 
 (1890), Aug. 2 and Nov. 15; vol. li., p. 229 (Feb. 21, 1891) ; vol. Hi., 
 pp. 174, 471, 721 (Aug. 8, Oct. 24, and Dec. 26, 1891); vol. lv. r 
 pp. 28, 99, 244, 292, 339, and 364 (Jan. 14, Feb. 4, Mar. 18, April 1,. 
 April 15. April 22, 1893). 
 
 Lang (H.) Matte Smelting. New York, 1896. 
 
 CHAPTER XVIII. -REFINING AND PARTING OF GOLD 
 BULLION. 
 
 Goddard (Jonathan). Experiments on Refining Gold with Antimony.. 
 
 Phil. Trans. Roy. Soc., 1676. 
 Leibius (A. ) Separation of Gold from Silver Chloride in Miller's Process. 
 
 Trans. Roy. Soc. of l\/ew South Wales, 1872, p. 67. 
 
 Egleston (T.) Parting Gold and Silver in California. New York, 1877. 
 Egleston (T.) Parting Gold and Silver by means of Iron at Lautenthal. 
 
 New York, 1885. 
 Egleston (T.) The Separation of Silver and Gold from Copper at Oker^ 
 
 Washington, 1885. 
 
 Hugon. Etude sur le raffinage electrolytique du cuivre noir. Paris, 1885. 
 Eglesion (T.) Treatment of Gold and Silver at the United States Mint. 
 
 London, 1886. 
 Skidmore (W.) Parting Gold and Silver in the United States. Ninth 
 
 Report of the Col. State Min., 1889, pp. 67-90. This is a complete 
 
 account of the methods in use in the United States. 
 
 Ulke (Titus). Parting and Refining Gold and Silver. Mineral Industry 
 for 1895, pp. 343-366. 
 
BIBLIOGRAPHY. 513 
 
 CHAPTERS XIX. AND XX. ASSAYING. 
 
 Carranza (A.) El Ainstamieto i Proporcion de las Monedas de Oro, Plata 
 
 i Cobre, i la reduccion distos Metales a su Debida estimaccion. Madrid, 
 
 1629. 
 Bedrock (Wm ) A new Touchstone for Gold and Silver Wares. London, 
 
 1651. 2nd edition, 1679. 
 
 Le Febure (N. R.) Compleat body of chymistry. 1670. 
 Petty (Sir John). Laws of Nature in Assaying Metals. Translated in part 
 
 from the German of L. Erckern. London, 1683. 
 Cramer (J. A.) Elementa artis docimasticse. Lugduni Batavorum 
 
 (Leyden), 1744. 
 
 Symonds (W.) Essay on the Weighing of Gold, &c. London, 1756. 
 Cramer, M D. (J.A.) Elements of the Art of Assaying Metals. London, 
 
 1764. 
 
 Pouchet. Le nouveau titre des matieres d'or et d'argent. Rouen, 1798. 
 Becquerel. Gold and Electricity. Ann. de Chimie et de Physique, vol. 
 
 xxiv. (1823) ; also Article by Oersted, d, vol. xxxix. (1828), p. 274. 
 Chaudet. L'Art de 1'essayeur. Paris, 1835. 
 Boaemann (Th. ) Anleitung zur Berg-und Hiittenmannischen Probirknnst. 
 
 Clausthal, 1845. 
 
 Berthier. Traite des essais par la voie se"che. Paris, 1847. 
 Pettenkofer. Berywerksfreund, vol. xii. (1849). Article on Gold Bullion 
 
 Assaying. 
 
 Watherston (J: H.) The Gold Valuer. London, 1852. 
 Plattner (C. F.) Probirkunst. Freiberg, 1853. 
 
 Bodemann and Kerl. Treatise on Assaying. Translated by W. A. Good- 
 year. New York, 1868. 
 Doraeyko (J. ) Tratado de Ensayes, tanto por la via seca comopor la via 
 
 humeda. Chile, 1873. 
 Foord (G.) Mechanical Assay of Quartz. Trans. Hoy. Soc., Victoria, 
 
 vol. x., pp. 139-147. 1874. 
 Broch (Dr. O.) Assay of Gold by means of its Density. Norwegian Nyt. 
 
 Mag. fur Naturvsk. Christiania, 1876. 
 Ricketts (P. ) Notes on Assaying. New York, 1876. 
 Kerl (B ) Metallurgische Probirkunst. 1880. 
 Attwood(G.) Practical Blowpipe Assaying. London, 1880. 
 Chapman (E. J. ) Assay Notes. Practical instructions for the determin- 
 ation by furnace assay of Gold and Silver in rocks and ores. Toronto, 
 
 1881. 
 Mitchell (W.) Manual of Practical Assaying. Edited by Wm. Crookes. 
 
 London, 1881. 
 
 Balling (C.) L'Art de 1'essayeur. Paris, 1881. 
 Jagnaux (TRaoul). Traite pratique d'analyses chimiques et d'essais 
 
 industriels. Paris, 1884. 
 Rossler (H.) Article on Gold Bullion Assaying. Dinner's Polyt. Journ.. 
 
 vol. ccvi. (1884). 
 
 Black (J. G.) Chemistry of the Gold Fields. Dunedin, N.Z., 1885. 
 Aaron (C. H.) Manual of Assaying. San Francisco, 1885. 
 Hiorns (A. H. ) Practical Metallurgy and Assaying. London, 1888. 
 Fremy. L'or dans le Laboratoire. Cumenge and Fuchs. Ency. Chim., 
 
 vol. iii., c. 16e. Paris, 1888. 
 
 Ross (W. A.) Blowpipe Analysis. London, 1889. 
 Brown and Griffiths. Manual of Assaying of Gold, Silver, &c. London, 
 
 1890. 
 
 33 
 
514 BIBLIOGRAPHY. 
 
 Lieber (O. M.) Assayer's Guide. New York. 
 
 Plattner. Blowpipe Analysis. Enlarged by Richter (Th.) Translated by 
 
 H. B. Cornwall. New York, 1890. 
 Riche(A.) L'Art de 1'essayeur. Paris, 1892. 
 Purman. Practical Assaying. New York, 1894. 
 Ledoux (A. R. ) Assay of Copper and of Copper materials for Gold and 
 
 Silver. Am. fnst. Mng. Eng., October, 1894. 
 Beringer (C. & J. J.) Text-book of Assaying. 5th edition. London, 
 
 1898. 
 
 
515 
 
 INDEX 
 
 Aaron, C. H., 134, 286, 290, 442. 
 Aeration of decomposed slimes, 341. 
 Africa, Gold in, 41, 216, 500. 
 Agitation in cyanide process, 325. 
 Agricola, 2, 91, 95, 96, 398. 
 Alaska Treadwell Mine, 248, 298, 494. 
 Albertus Magnus, 398. 
 Alchemy, 2. 
 Allen, A. H., 479. 
 Allotropic forms of gold, 12. 
 Alloys, Assay of, 484. 
 
 Crystalline, 14, 15. 
 Gold, 13. 
 
 Gold and copper, 17. 
 Gold and silver, 16. 
 Scorification of, 484. 
 Aluminium alloys, Assay of, 485. 
 Amalgam, Cleaning of, 138. 
 ,, Composition of, 140. 
 
 Retorting of, 140. 
 Amalgamated plates, see Plates, 
 
 Amalgamated. 
 Amalgamation, 129. 
 assay, 459. 
 Causes of prevention of, 
 
 151. 
 
 "Designolle process of, 136. 
 Effect of chemicals on, 131. 
 ,, hammering on, 147. 
 , , temperature on, 123. 
 Methods of causing, 148. 
 pans, 165. 
 Amalgams, 15. 
 
 Assay of, 486. 
 Ammonium carbonate used in treat- 
 ing gold ores, 319. 
 Ancient process of milling in Chili, 
 
 91. 
 
 ,, ,, ,, in the Hartz, 97. 
 
 ,, ,, ,, in Norway, 91. 
 
 in the Tyrol, 96. 
 Ancient rivers of California, 64, 68. 
 Anderson, 363. 
 Annealing cornets, 477. 
 
 Annealing crucibles, 387. 
 
 fillets, 475. 
 
 Annual production of gold, 495, 498. 
 Antimony alloys, Assay of, 484. 
 
 ,, in roasting furnace, 240, 
 
 241. 
 
 ,, used for parting, 397. 
 Apron plates, 126. 
 Arborescent gold, 9, 35. 
 Arnold, J. 0., 20. 
 Arrastra, 92. 
 
 for prospecting, 95. 
 Arsenic alloys, Assay of, 484. 
 
 ,, in roasting furnace, 240. 
 Artificial crystals of gold, 10. 
 Asbestos filtering cloth, 304, 307. 
 Assay by amalgamation, 459. 
 blowpipe, 437. 
 bromine, 460. 
 cadmium, 483. 
 chlorination, 459. 
 colour and hardness, 490. 
 crucible method, 440. 
 density, 490. 
 electrolysis, 491. 
 induction balance, 491. 
 scorifieatioii, 455, 484. 
 spectroscope, 490. 
 touchstone, 489. 
 Fluxes used in, 443. 
 furnaces, 440, 467. 
 General charges in, 441. 
 materials, Examination of, 454. 
 of base ores, 448. 
 
 bleaching powder, 282. 
 complex materials, 458. 
 cupel, 454. 
 
 cyanide of potassium, 369. 
 gold bullion, 462. 
 
 ,, Accuracy of, 482. 
 Lead used in, 460. 
 ,, Losses of gold in, 478. 
 , , Use of proofs in, 481. 
 gold ores, 447. 
 
516 
 
 INDEX. 
 
 of gold ores, Cleaning slag 
 
 in, 448. 
 
 ,, Cupellation in, 448. 
 Mint sweep, 462. 
 purple of Cassius, 462. 
 pyrites, 459, 461. 
 rich sulphides, 456. 
 tellurides, 457. 
 Preliminary, 483. 
 Roasting in, 447. 
 Special methods of, 458. 
 Wet methods of, 488. 
 Whitehead's method of, 461. 
 Assay- ton, 441. 
 Attwood, G. M., 153. 
 Atwood's amalgamator, 199. 
 Aurates, 30. 
 Auric bromide, 28. 
 ,, chloride, 21. 
 
 ,, ,, Decomposition of 24. 
 
 ,, ,, Volatilisation of, 22. 
 
 oxide, 29. 
 
 Auricyanide of potassium, 29. 
 Auriferous quartz, Formation of, 33, 
 
 42. 
 
 Aurocyanide of potassium, 28. 
 Aurosilicates, 33. 
 Aurosulphites, 30. 
 Aurous chloride, 21. 
 ,, cyanide, 28. 
 ,, oxide, 29. 
 Austin, W. L., 380. 
 ,, process, 381. 
 Australia, Battery practice in, 208. 
 
 Gold in, 41, 500. 
 Automatic feeding machines, 120. 
 Available chlorine, 281. 
 
 Bagration, 354. 
 
 Balances for bullion assay, 464, 465, 
 
 Ball, E. J., 445. 
 mills, 163, 165. 
 ,, stamp, 158. 
 Balling, C., 460, 483. 
 Banket, 216. 
 Bar mining, Deep, 58. 
 Barba, 91. 
 
 Barrel for cleaning up, 138, 211. 
 process of chlorination, 274, 
 
 299. 
 
 ,, Advantages of, 276. 
 
 ,, Amount of chemicals 
 
 used in, 282. 
 , , , , Amount of water used 
 
 in, 280. 
 
 ,, History of, 274. 
 
 Basalt overlying auriferous deposits, 
 64. 
 
 Base bullion, 383. 
 
 ,, Assay of, 461. 
 
 ,, ores, Assay of, 448. 
 Bassick mine, 38. 
 Batea, 45. 
 
 Battery, Stamp, see Stamp battery. 
 Bayly, F. W., 466. 
 Bazin's centrifugal amalgamator, 136, 
 Beach mining, 62. 
 Beckmann, 90, 95, 398. 
 Becquerel, 4. 
 Berdan pan, 214. 
 Berthelot, 357. 
 Berzelius, 3. 
 Bettel, W., 335, 337, 346, 349, 351, 
 
 354, 355, 370. 
 Bibliography, 503. 
 Biringuccio, 90, 398. 
 Black concentrates, 49. 
 Blackhawk, Stamp mills at, 204. 
 Blake crusher, 101, 231. 
 
 W. P., 10, 36. 
 Blake-Marsden crusher, 102, 217. 
 Blanching, 18. 
 Blanket strakes, 181, 210. 
 Bleaching powder, 281. 
 
 ,, ,, Assay of, 282. 
 
 Blomfield, J. T., 290, 307. 
 Blowpipe assay, 437. 
 Blue gravels, 66. 
 
 ,, lead theory, 66. 
 
 ,, Spur Company, 78, 85. 
 Bodlander, G., 354. 
 Bonanza Mill, Cyanide process at r 
 
 349. 
 Bone -ash for cupels, 449. 
 
 ,, to cover bullion, 388. 
 ,, to protect muffle, 449. 
 Booming, 54. 
 Booth, J., 19. 
 Borax for refining bullion, 388. 
 
 ,, in crucible assay, 443. 
 
 ,, in scorification, 455, 485. 
 Born, Baron, 2, 97. 
 Boss, M. P,, 168. 
 
 ,, system of pan-amalgamation,. 
 
 171. 
 
 Bovisa, Chlorination at, 318. 
 Boxes, Puddling, 48. 
 
 ,, Sluicing, 50. 
 Boyle, 5. 
 
 British Isles, Gold in, 39. 
 Brittle gold, 19, 390, 391, 431. 
 Brodie works, 335. 
 Bromide of gold, 28. 
 Bromination of ores, 311, 314. 
 Bromine, Action of, on gold, 267* 
 
 Assay by, 460. 
 Brough, B. H., 95, 344. 
 
INDEX. 
 
 517 
 
 Brown and Griffiths, 442. 
 Brown, R. E., 68. 
 Bruckner furnace, 242. 
 Biichner, 3. 
 Buckboard, 439. 
 Bucyrus steam shovel, 63. 
 Buddie, Cornish, 182. 
 ,, German, 180. 
 Buisson, 3. 
 Bullion, Gold, Assay of, see Assay of 
 
 gold bullion. 
 Casting of, 391. 
 
 ,, Composition of, 334, 
 
 384, 409, 415, 428, 
 430, 431. 
 
 ,, Definition of, 393. 
 
 ,, Granulation of, 398. 
 
 Losses of, 393. 
 Melting of, 387. 
 Parting of, 396. 
 Price of, 336. 
 Refining of, 384, 388. 
 Skimming of, 389. 
 Toughening of, 390. 
 Bunker' Hill mine, 252, 277. 
 Butters, C., 269, 271, 272, 296, 324, 
 332, 334, 362, 373. 
 
 <3admium, Assay by, 483. 
 Calaverite, 38. 
 Caldecott, 328, 340, 366. 
 Caledonia Mill, 215. 
 California, Battery practice in, 199. 
 ,, Chlorination in, 296. 
 ,, Deep placers in, 64. 
 Cam-pulley, 113. 
 
 -shaft, 111. 
 Cams, 111. 
 Canada, Gold in, 41. 
 Capel, 490. 
 Carnelley, 4. 
 Carnot, A., 458. 
 Carolina, Chlorination in, 302. 
 Cassel process, 295. 
 Cassel-Hinman bromine process, 314. 
 assius, Purple of, 3, 27, 34, 35. 
 
 ,, Assay of, 462. 
 
 Casting ingots, 391. 
 Cazo process, 97. 
 'Cement gravel, 55, 66. 
 
 treated by arrastra, 94. 
 
 mill, 83. 
 
 ,, pan, 84. 
 Cementation, 397. 
 Challenge feeder, 122, 217. 
 Chalmers, see Hatch. 
 Charcoal, Precipitating gold by, 
 
 289. 
 Charging scoop for bullion, 387. 
 
 Charleton, A. G., 494. 
 Chaudet, 466, 476, 486, 487. 
 Checks in bullion assay, 481. 
 Chemicals for stamp battery, 131. 
 Chemistry of cyanide process, 354. 
 
 ,, of oxidising roasting, 238. 
 Chester, 10. 
 Chilian Mill, 91. 
 China, Gold in, 40. 
 Chlor-auric acid, 27. 
 Chloride of gold, 21. 
 
 ,, ,, Decomposition of, 24. 
 
 ,, Melting point of, 24. 
 
 Volatility of, 22, 245. 
 ,, of silver, see Silver, Chloride 
 
 of. 
 Chlorination at Alaska - Treadwell 
 
 Mine, 298. 
 at Deloro, 299. 
 , , at Golden Reward Mill, 303. 
 ,, at Mt. Morgan, 318. 
 at Rapid City, 311. 
 ,, at Robinson Mine, 344. 
 ,, Barrel process, 295. 
 ,, Cassel's process, 295. 
 ,, Greenwood process, 295. 
 
 in California, 296. 
 in Carolina, 302. 
 , , Julian process, 296. 
 ,, MunkteU process, 293. 
 ,, Newbery - Vautin process, 
 
 292. * 
 
 ,, Pollok process, 292. 
 Vat process, 261, 296. 
 Chlorine, Action of, on gold, 266. 
 ,, on gold alloys, 267. 
 
 ,, ,, on organic matter. 
 
 268. 
 ,, ,, on oxides of metals, 
 
 268. 
 
 ,, on sulphates, 268. 
 
 ,, ., on sulphides, 265, 
 
 267. 
 ,, Amount required for ores, 
 
 269. 
 
 ,, Generation of, 264. 
 ,, Parting by, 414. 
 ,, Refining by. 431. 
 Christy, S. B., 244, 245, 247, 349, 
 
 361, 417. 
 
 Chrome-steel, 110. 
 Clarkson, T., 197. 
 Classification of ores, 178. 
 
 ,, of parting processes, 
 
 396. 
 
 Claveus Gasto, 5. 
 Clay crucibles, 387. 
 Clean-up barrel, 138, 211. 
 
 in cyanide process, 332. 
 
518 
 
 INDKX. 
 
 Clean-up in hydraulicking, 80. 
 in sluicing, 53. 
 ,, of stamp battery, 137. 
 
 pan, 139. 
 
 Cleaning slags, 448, 456. 
 Clennel, 332, 334, 362. 
 Cobalt alloys, Assay of, 485. 
 Cohesion of gold, 4. 
 Coinage, Alloys used in, 17. 
 Collins, H. F., 381. 
 Colorado, Battery practice in, 204. 
 Colour of alloys, Assay by, 490. 
 
 of gold, 2. 
 
 Composition of bullion, 384, 409, 415. 
 ,, ,, from chlorination 
 
 process, 384. 
 
 from cyanide process, 334. 
 ,, placers, 385. 
 ,, stamp battery, 385. 
 of gold slimes, 333. 
 of native gold, 39. 
 Compounds of gold, 20. 
 
 Natural, 38. 
 
 Concentrates, Amalgamation of, 172, 
 
 199, 468. 
 Black, 49. 
 ,, Chlorination of, 296, 
 
 298. 
 
 Cyaniding of, 345. 
 
 ,, Grey, 49, 60. 
 
 Concentration, 172, 175. 
 Concentrator, Centrifugal, 183. 
 
 Clarkson & Stanfield's, 198. 
 Duncan, 183. 
 Embrey, 192. 
 Frue vanner, 185. 
 Gilpin County, 184. 
 Golden Gate, 202. 
 Hendy, 183. 
 Llihrig vanner, 193. 
 Raising Gate, 182. 
 Triumph, 193. 
 Conductivity of gold, 5. 
 Consumption of cyanide, 343, 350, 
 
 364, 372. 
 
 ,, of gold, 501. 
 
 Copper amalgamated plates, 97, 98, 
 
 124. 
 
 ,, ,, discoloration of, 130. 
 
 ,, ,, for cement gravel, 
 
 83. 
 
 ,, ,, in sluicing, 53, 62. 
 
 ,, Oxide of, in toughening 
 
 bullion, 391. 
 ,, sulphate, Crystallisation of, 
 
 407. 
 Cornets, 475. 
 
 Silver in, 480. 
 Weighing the, 478. 
 
 Corrosive sublimate, in toughening- 
 
 gold, 390. 
 
 Cost of barrel chlorination, 303, 310. 
 , , cyanide process, 223, 336, 339 r 
 
 343, 349, 350, 352. 
 Munktell process, 294, 295. 
 parting, 402, 403, 408, 409, 
 
 414, 429, 437. 
 Plattner process, 273, 299. 
 production of gold, 493. 
 stamp amalgamation process,. 
 
 ,, working placers, 89. 
 ,, ,, deep placers, 88. 
 
 ,, ,, shallow placers, 86. 
 
 Coyoting, 44. 
 Cradle, 46. 
 Crawford Mill, 163. 
 Creuzbourg, 3. 
 Crinoline hose, 74. 
 Cripple creek, 38, 496. 
 Croesus Mine, 218. 
 Crookes, Sir W., 144, 316, 491. 
 Crosse, A. F., 329. 
 Crown Mine, 223. 
 
 ,, Treatment of slimes at r 
 
 338. 
 Crucibles for assaying, 414. 
 
 ,, ,, melting bullion, 387. 
 
 Size of, 392. 
 Crushing before chlorination, 227 r 
 
 314. 
 
 ,, ,, cyanide process, 321. 
 
 ,, in stamp battery, 89. 
 
 Crystalline alloys of gold, 14, 15. 
 Crystallisation of copper sulphate,. 
 
 407. 
 
 gold, 9, 35. 
 ,, silver sulphate^ 
 
 411, 412. 
 
 Cumenge, 33, 458, 466. 
 Cupellation in assay of bullion, 467. 
 
 ores, 448. 
 Influence of base metals 
 
 on, 451. 
 ,, temperature 
 
 on, 471. 
 Cupels, 449, 469. 
 
 Assay of, 454. 
 Curtis, A. H., 159, 228, 232. 
 Cyanide of gold, 28. 
 
 of mercury, 373, 374. 
 of potassium, action on gold 1 
 and other metals, 
 354. 
 ,, action on salts and 
 
 minerals, 364. 
 , , acti on on sulphides, 
 154, 366. 
 
INDEX. 
 
 519 
 
 Cyanide of potassium, Assay of, 
 
 369. 
 
 Commercial, 330. 
 
 ,, Consumption of, 
 
 343, 350, 364, 372. 
 
 ,, Decomposition of, 
 
 342, 360, 362, 365. 
 
 in stamp battery, 
 
 I'M > I > 
 
 lol, 64:6. 
 Selective action of, 
 
 320, 365. 
 Solubility of gold 
 
 in, 358. 
 Solubility of silver 
 
 in, 360. 
 
 Solubility of vari- 
 ous metals in, 357. 
 of sodium, 330. 
 of zinc, 362. 
 process, 319. 
 ,, ,, Chemistry of, 354. 
 
 ,, Direct treatment in, 
 
 342. 
 ,, ,, Disposal of tailings 
 
 in, 330. 
 ,, ,, Double treatment in, 
 
 324. 
 
 ,, Mac Arthur -Forrest, 
 
 see MacArthur-For- 
 rest process. 
 
 Pelatan-Clerici, 377. 
 
 Results of, 351. 
 
 Siemens-Halske,347. 
 Sulman-Teed, 376. 
 Wilde, 378. 
 Cyanogen bromide, 351. 
 Cyclops mill, 165. 
 
 Daintree, R., 69, 153. 
 
 Dakota, Battery practice in, 215. 
 ,, Bromination in, 311. 
 Chlorination in, 303. 
 ,, Pyritic smelting in, 381. 
 Roasting in, 256. 
 Dams, in river mining, 56. 
 D'Arcet, C., 402, 466. 
 Davis, W. M., 289. 
 De Lacy, 274. 
 De Morveau Guyton, 3. 
 Dead-leaf gold, 16. 
 Debray, 21, 35. 
 
 Decomposition in zinc boxes, 361. 
 of gold chloride, 24. 
 ,, of potassium cyanide, 
 
 342, 360, 362, 365. 
 Deep bar mining, 58. 
 ,, placer deposits, 63. 
 Method of working, 
 
 Deetken, G. F., 202, 206, 228, 262, 
 
 267, 274. 
 Del Mar, A., 43. 
 Deloro, Mears process at, 299. 
 Dendritic gold, 9, 35. 
 Dennes, D., 234, 285, 305, 311, 313. 
 Density, Assay by, 490. 
 
 ,, of alloys of gold, 17. 
 ,, of gold, 4. 
 Designolle process, 136. 
 Desmarest, 3. 
 
 Detection of gold in alluvium, 45. 
 ores, 457. 
 
 ,, solution, 27, 458. 
 
 Deville, 5. 
 D'Hennin, 488. 
 Dibromide of gold, 28. 
 Dies, 110. 
 
 Diffusion of gold, 15. 
 Dingier, 461. 
 Diodorus Siculus, 90. 
 Dip sample, 464. 
 Dissemination of gold, 36. 
 Distribution of gold, 35, 39. 
 
 ,, in gravels, 67. 
 
 Ditches in hydraulicking, 73. 
 Dixon, 373. 
 Dodge crusher, 102. 
 Double discharge in stamp battery, 
 
 119. 
 Double treatment in cyanide process, 
 
 324. 
 
 Dredging for gold, 58. 
 Drift mining, 82. 
 Drop of stamps, 115. 
 Drops in sluice, 52. 
 Dry crushing, 227. 
 
 ,, diggings, 55. 
 Drying ore, 233. 
 Dubois, W. E., 36. 
 Ductility of gold, 3. 
 Duffield, 150, 192. 
 Duflos, 224, 274. 
 
 Egleston, T., 69, 147, 394. 
 Egypt, Ancient workings in, 1, 90. 
 Ehrmann, 363, 376. 
 Eissler, 296, 380. 
 El Callao Mine, 494. 
 Electrical precipitation of gold, 347. 
 Electricity in amalgamation, 135. 
 Electrolysis, Assay by, 491. 
 
 , , of cyanide of potassium, 
 
 349. 
 
 ,, Parting by, 432. 
 
 Electrum, 16. 
 Elephant stamp, 158. 
 Elevator, Hydraulic, 59, 84. 
 Elkington, 354. 
 
520 
 
 INDEX. 
 
 Eisner, 354. 
 
 Embrey concentrator, 192. 
 
 Emmons, 36. 
 
 Endlich, F. M., 320. 
 
 Equal falling particles, 176. 
 
 Essoii, 25. 
 
 Eureka rubber, 199. 
 
 Europe, Gold in, 39. 
 
 Expansion of gold, 5. 
 
 Extra solution, Russell's, 31. 
 
 Faraday, 3, 354. 
 
 Farrar, S. H., 216. 
 
 Feather river, 57, 81. 
 
 Feeding in stamp battery, 120. 
 
 Feix, J., 394. 
 
 Feldtmann, 361. 
 
 Ferry, N. A., 272. 
 
 Figuier, 3. 
 
 Filter beds for leaching, 261, 263, 
 
 304, 307, 323. 
 press, 334, 337. 
 Fineness of bullion produced by 
 
 Miller's process, 430. 
 Finkener, 371. 
 
 Flashing in cupeUation, 449, 470, 471. 
 Float gold, 149, 150. 
 Flouring of Mercury, 143. 
 Flumes for hydraulicking 73. 
 Fluviatile theory of Calif ornian placer 
 
 deposits, 63. 
 
 Fluxes in assaying, 442, 443. 
 ,, in pyritic smelting, 381. 
 ,, in refining bullion, 389. 
 Fly catchers, 56. 
 Foliated tellurium, 38. 
 Formation of auriferous quartz, 33. 
 
 ,, of nuggets, 69. 
 Forms of gold found in nature, 35. 
 Forrest, W., 320. 
 
 R. W., 320. 
 Foster, Le Neve, 123. 
 Foundations of stamp battery, 108. 
 Framework of stamp battery, 108. 
 Frankfort refinery, Parting at, 435. 
 Freezing point of gold, 4, 13. 
 Fremy, 458, 461, 466, 479, 485, 490. 
 Fresenius, 365, 371. 
 Fromm, 14. 
 Frue vanner, 185. 
 
 Riffled belt for, 192. 
 
 Fuchs, 458, 466. 
 Fulminating gold, 30. 
 Furnaces, Assay, 440, 467. 
 
 ,, Brown, 250. 
 
 Bruckner, 255. 
 
 , Chlorination, 421. 
 
 Gas Assay, 468. 
 
 Hofmann, 256. 
 
 Furnaces, Mechanical, 250. 
 
 Melting, 386. 
 
 Muffle, 467. 
 
 O'Hara, 250. 
 
 Pearce Turret, 249. 
 
 Refining, 386. 
 
 Reverberatory, 238. 
 
 Revolving cylindrical, 255. 
 
 Ropp, 252. 
 
 Rotating-bed, 252. 
 
 Shaft, 250. 
 
 Spence, 251. 
 
 White, 258. 
 
 White-Howell, 259. 
 Fusibility of gold, 4. 
 Future production of gold, 497. 
 
 Gas, Chlorine, method of production, 
 
 264. 
 
 ,, ,, used in plattner pro- 
 
 cess, 264, 266. 
 ,, Roasting by, 260. 
 Gates crusher. 103, 217. 
 Geber, 90. 
 
 Geographical distribution of gold, 39. 
 Geological ,. 36. 
 
 George and May mine, 343. 
 Gernet, von, 348. 
 Gilpin County, Battery practice at, 
 
 204. 
 
 ,, ,, concentrator, 184. 
 
 Godshall, L. D., 314. 
 Gold, Action of chlorine on, 267. 
 ,, ,, of potassium cyanide 
 
 on, 358. 
 
 ,, bullion, see Bullion, Gold. 
 ,, Detection of, see Detection of 
 
 gold. 
 
 ,, Losses of, see Loss of gold. 
 ,, ores, Assay of, 437. 
 ,, Properties of, 2. 
 ,, residues, Treatment of, 290, 
 
 400, 410, 413. 
 slimes, 333. 
 
 ' Treatment of, 334, 336. 
 
 Solubility of, 10. 
 Golden Fleece, Legend of, 2. 
 
 Gate concentrator, 202. 
 Reward Mill, 303. 
 Gooseneck in Mears process, 275. 
 Gore, G., 356, 357, 432. 
 Goyder, 363, 370. 
 Grade of plates, 132. 
 
 sluices, 51, 60, 79. 
 Granulation of bullion, 398, 402. 
 Graphic tellurium, 38. 
 Graphite crucibles, 387. 
 Gravel, Cemented, 55. 
 
 Method of washing, 45. 
 
INDEX. 
 
 521 
 
 Gravels, blue and red, 66. 
 
 Gravitation stamps, 107. 
 
 Green gold, 16. 
 
 Greenwood process, 295. 
 
 Grey concentrates, 49, 60. 
 
 Griffiths, H. D., 204. 
 
 Grizzly, 52. 
 
 Grommestetter, 95. 
 
 Ground sluicing, 54. 
 
 Groves, J., 469. 
 
 Gruner, 235. 
 
 Guggenheim Smelting Coy., Parting 
 
 at, 435. 
 
 Guides in stamp battery, 114. 
 Giiptner, 483. 
 
 Gutters in ancient rivers, 67. 
 Gutzkow, F., 409, 411, 414. 
 ,, process, 409. 
 
 New, 411. 
 
 Haile Mine, 136, 302. 
 
 Haindl, 487. 
 
 Hall-marking, 16. 
 
 Halogen compounds of gold, 21. 
 
 Hammond, J. H. , 200. 
 
 Hankey, J., 151. 
 
 Hanks, 389. 
 
 Harcourt, 25. 
 
 Hardness of alloys, 490. 
 
 of gold, 3. 
 Hare, R., 5. 
 Harris, 93. 
 Harscher, C., 95. 
 Hartz jigs, 196. 
 Hatch, 219, 221, 222, 324, 343, 346, 
 
 494. 
 
 Heat of formation of chlorides, 21. 
 ,, ,, of cyanides, 357. 
 
 Hendy's challenge feeder, 122. 
 Heycock and Neville, 13, 14. 
 Hickock, 256. 
 Hinman, see Cassel. 
 Hofmann furnace, 256. 
 Holleman, 12. 
 Homburg, 5. 
 Homestake Mill, 215. 
 Hood, J. J., 375. 
 
 ,, process, 374. 
 Horn spoons, 46. 
 Howe, H M.,241. 
 Howell- White furnace, 258. 
 Huggins, Dr., 4. 
 Hughes, Prof., 491. 
 Huntington Mill, 159. 
 Husband's Stamp, 155. 
 Hydrates of gold, 29. 
 Hydraulic elevator, 59, 84. 
 
 ,, sizing boxes, 179. 
 Hydraulicking, 57, 71. 
 
 Hydrocyanic acid, 370. 
 Hydrogen amalgamator, 173. 
 Hydrolytic decomposition of potas- 
 sium cyanide, 360. 
 Hyposulphites of gold, 31. 
 
 Inch, Miners', 74. 
 
 Independence Mill, 305, 307. 
 
 India, Gold in, 40. 
 
 Induction balance, 491. 
 
 Ingalls, 335. 
 
 Ingots, Casting, 391. 
 
 Inquartation, 399, 452. 
 
 Iodide of gold, 28. 
 
 Iridium alloys, Assay of, 487, 488. 
 
 Iridosmine, 70. 
 
 Iron alloys. Assay of, 485. 
 
 pipes in hydraulicking, 73. 
 
 ,, riffle bars," 78. 
 
 ,, Sulphate of, used to precipitate 
 gold, 271, 286, 298, 299. 
 
 ,, vessels for parting, 401, 409. 
 
 Jacob, 495, 501. 
 
 Janin, Jun., L., 19, 136, 153, 154, 
 
 320, 359. 
 
 Japan, Gold in, 40. 
 Jars, M., 97. 
 Jenkins, H. C., 446. 
 Jevons, 502. 
 Jigs, Hartz, 196. 
 
 ,, Pneumatic, 197. 
 Johnson, E. H., 335. 
 Johnson's filter press, 337. 
 Jordan's amalgamator, 174. 
 Julian, 295, 323. 
 
 ,, process, 292. 
 Jumpers ore, 328. 
 
 Kandelhardt, 466. 
 
 Kasentseff, 15. 
 
 Kedzie, G. E., 364. 
 
 Keith, N. S., 373. 
 
 Kennel, California, Chlorination 
 
 works at, 296. 
 Kerl, B.,225, 479. 
 Kerr, Prof., 36. 
 Knox pan, 139, 199. 
 Kohlrausch, 12. 
 Krafft, 10. 
 Krom, S. K, 197, 228. 
 
 rolls, 228. 
 Kruss, 3, 5, 25, 26, 29, 34. 
 Kumara tail sluice, 55. 
 Kunckel, 402. 
 Kustel, G., 98, 235, 238, 244, 267. 
 
 Lang, H., 380, 382, 383. 
 Lange, 224. 
 
522 
 
 INDEX. 
 
 Langguth, W., 287. 
 
 Langlaagte estate, 323, 381, 
 
 Latent heat of gold, 4. 
 
 Latta, 153. 
 
 Laurie, 5, 14. 
 
 Lava above auriferous gravels, 64. 
 
 Law, R., 423. 
 
 Lazzlo amalgamator, 166. 
 
 Le Chatelier pyrometer, 472. 
 
 Leaching by decantation, 283. 
 
 pressure, 285, 306, 312. 
 vacuum, 284, 292. 
 
 Difficulties of, 284. 
 ,, in chlori nation barrel, 305. 
 cyanide process, 322, 
 
 324. 
 
 ,, ,, with agita- 
 
 tion, 327. 
 
 ,, in Plattner process, 270. 
 vats in chlorination pro- 
 cess, 261. 
 
 ,, cyanide process, 322. 
 
 Lead, amount used in bullion assay, 
 
 466, 467. 
 ,, reduced in ore assay, 
 
 443. 
 lining for chlorination barrel, 
 
 280. 
 
 Leibius, A., 6, 417. 
 Lever or jack in stamp battery, 114. 
 Level, 16, 32, 34, 450. 
 Lime used in cyanide process, 324, 
 
 338, 341. 
 Lindet, 21. 
 
 Liquation of gold alloys, 18. 
 Liversedge, 70. 
 Lockyer, 490. 
 Long Tom, 47. 
 Lossen, 11. 
 
 Loss of gold by volatilisation, 394. 
 ,, in amalgamation, 145. 
 assaying, 478. 
 refining, 393. 
 roasting, 247. 
 mercury, 212. 
 Louis, H. , 4, 90. 
 Lowe, 24. 
 Liihrig vanner, 193, 217. 
 
 MaeArthur, J. S., 320, 321, 330, 
 
 334, 337, 358, 372. 
 MacArthur-Forrest process, 321. 
 
 At the George and May, 343. 
 
 At the New Primrose, 346. 
 
 At the Sylvia mine, 344. 
 
 Composition of bullion from, 334. 
 
 Consumption of cyanide in, 343, 
 350, 364, 372. 
 
 Direct treatment of ore in, 343. 
 
 Leaching in, 324. 
 
 Ores suitable for, 378. 
 
 Plant required in, 337. 
 
 Production of bullion in, 333. 
 
 Results of, 352. 
 
 Strength of solution for, 328, 359. 
 
 Treatment of concentrates by, 344. 
 
 ,, ,, slimes by, 337. 
 
 M'Cutcheon, J. , 140, 429. 
 M'Dermott, 150, 178, 191, 192. 
 M'Dougal, G., 149. 
 Maclaurin, R. C. , 354, 359. 
 Mactear, 152. 
 Magnetism of gold, 4. 
 Makins, G. H., 5, 478. 
 Maldonite, 39. 
 Malleability of gold, 3. 
 MaUet, 5. 
 
 Maltitz, Count von, 95. 
 Management of gold mills, 493. 
 Manganese alloys, Assay of, 485. 
 Marais, 355. 
 Maray, 91. 
 
 Materials for crushing surfaces, 105. 
 Mathison, 402. 
 Matthey, E., 8, 20, 485, 487. 
 Matthiessen, 14. 
 Mattison, Edw., 71. 
 May Consolidated Mill, cyanide pro- 
 cess at, 350. 
 Maynard, G., 433. 
 Mears, Dr. H., 274, 275. 
 process, 275, 299. 
 Mechanical furnaces, 247. 
 Medina, 91. 
 Mein, 324. 
 Meinecke, 120. 
 Melting gold bullion, 387. 
 Mercur Gold Mine, 321, 325, 328, 
 
 332, 333. 
 Mercury, 15, 132. 
 
 , , Chloride of, used to toughen 
 
 gold, 390. 
 ,, cyanide, Use of, in cyanide 
 
 process, 373, 375. 
 ,, Flouring of, in stamp bat- 
 teries, 143. 
 ,, Loss of, in stamp battery, 
 
 144. 
 
 ,, Sickening of, 143. 
 Use of, 15, 80, 90. 
 ,, ,, in sluicing, 53. 
 
 ,, wells in stamp batteries, 
 
 135, 149, 210, 215. 
 Merrick, J. M., 459. 
 Merrifield Mine, 246. 
 Metallics, 439. 
 Methylamine, 362. 
 Mexico, Gold in, 41. 
 
INDEX. 
 
 523^ 
 
 Miers, 91. 
 
 MiU, Site for, 126. 
 
 Miller, F. B., 415. 
 
 ,, process, 414. 
 Minerals in placer deposits, 70. 
 
 ,, occurring with gold, 37, 
 Miner's inch, 74. 
 
 ,, pan, 45. 
 Mint, Method of assaying in, 463. 
 
 ,, sweep, Assay of, 462. 
 Miocene Ditch Company, 73. 
 Mitchell, 441, 451, 459. 
 Modderfontein Mine, 218. 
 Mode of occurrence of gold, 35. 
 Moebius, B , 432. 
 
 ,, process, 432. 
 Moldenhauer, 356. 
 Molloy's amalgamator, 173. 
 
 , , method of precipitating gold 
 
 cyanide, 377. 
 Monitor, 75. 
 
 Monochloride of gold, 21. 
 Morison, D. B., 115. 
 Morison's stamp, 158. 
 Mortar for crushing gold ores, 91. 
 
 ,, stamp battery, 108. 
 
 Morton,- Prof., 154. 
 Mould for ingots, 391. 
 Mt. Morgan Mine, 39, 318. 
 
 ,, ,, Precipitation of gold 
 at, 289. 
 
 ,, Roasting at, 152, 284. 
 Muffle furnace, 467. 
 
 ,, Temperature of, 472. 
 
 Miihlenberg, 320. 
 Miiller, 25, 31. 
 Munktell process, 293. 
 Muntz metal plates, 133. 
 Mylius, 14. 
 
 Nagyagite, 38. 
 
 Napier, 5, 6. 
 
 Native crystals of gold, 9. 
 
 gold, 39. 
 Natural compounds of gold, 38, 39. 
 
 ,, forms of gold, 35. 
 Neville, see Hey cock. 
 New Primrose Mine, Cyanide pro- 
 cess at, 346. 
 
 New South Wales, 'Gold in, 42. 
 New Zealand, Battery practice in, 
 
 208. 
 
 Gold in, 41. 
 Newbery, J. Cosmo, 143, 195, 274, 
 
 292. 
 
 Newbery -Vautin process, 292. 
 Nickel alloys, Assay of, 485. 
 Nickles, 11. 
 Nitric acid, Action of, on metals, 400. 
 
 Nitric acid, Parting by, 398. 
 
 , , Solubility of gold in 
 
 North Bloomfield Mine, 72, 80. 
 Nuggets, 37, 70 
 
 ,, Formation of, 69. 
 
 Occurrence of gold in nature, 35. 
 O'Driscoll, F., 274. 
 O'Hara furnace, 247. 
 Ore bins, 128. 
 Ores, Assay of, 437. 
 Origin of deep placer deposits, 63. 
 ,, gold in gravels, 68. 
 ,, in ores, 42. 
 
 Osmiridium, Separation of, from gold, 
 
 395. 
 
 Oxides of gold, 29. 
 Oxidised pyrites, Action of potassium 
 
 cyanide on, 366. 
 ,, ,, Amalgamation of, 
 
 153. 
 
 Oxidising roasting, 238. 
 Oxygen, Action on dissolution of gold 
 by potassium cyanide, 354, 358, 
 
 373 ! 
 
 O IO. 
 
 Palladium alloys, Assay of, 487. 
 Pan-amalgamation, 168. 
 Pan, Boss, 171. 
 Knox, 139. 
 Patton, 170. 
 Siberian, 61. 
 Washing by the, 43. 
 Paracelsus, 2. 
 Parting, 396. 
 
 assay, 462. 
 
 ,, by sulphuric acid, 
 
 482 
 
 double, 483. 
 History of, 463. 
 by cementation, 397. 
 ,, chlorine, 414. 
 ,, combined process, 408. 
 ,, electrolysis, 432. 
 ,, Gutzkow process, 409. 
 ,, new ,, 411. 
 
 ,, nitric acid, 398. 
 ,, sulphide of antimony, 
 
 397. 
 
 ,, sulphur, 397. 
 ,, sulphuric acid, 402. 
 flask, 453. 
 
 in bullion assay, 475. 
 ,, ore assay, 452. 
 ,, platinum tray, 476. 
 ,, porcelain crucible, 454. 
 test tube, 454. 
 processes, Classification of, 
 396 
 
524 
 
 INDEX. 
 
 Patio process, 92. 
 Pearce, 15, 16, 18. 
 
 ,, furnace, 249. 
 Peel, 477. 
 
 Pelatan-Clerici process, 377. 
 Peligot, 18. 
 Pelouze, 466, 479. 
 Pepys, 463. 
 Permanganate, Use of, in cyanide 
 
 process, 341. 
 Percussion tables, 184. 
 Percy, J., 90, 224, 289, 397, 403, 409, 
 
 429, 456, 480. 
 
 collection, 10, 15, 36, 289. 
 Pestarena Mine, 94. 
 Pettenkofer, 466. 
 Petzite, 38. 
 Philadelphia Mint, Liquation at, 19. 
 
 Parting at, 408. 
 Philippe of Valois, 463. 
 Phoenix Mine, Chlorination at, 302. 
 Pinos Altos, Treatment of bullion at, 
 
 436. 
 Placer deposits, Deep, 63. 
 
 ,, Shallow, 43. 
 gold, 37, 67, 69. 
 ,, mining, Cost of, 42. 
 Plates, Amalgamated, 97, 98, 124. 
 ,, Copper, see Copper 
 
 plates. 
 
 Corrugated, 135. 
 Grade of, 132. 
 in sluicing, 53, 62. 
 in Transvaal, 219. 
 Muntz metal, 133. 
 Preparation of, 124. 
 Shaking, 134. 
 Treatment of, 129, 
 
 140, 220. 
 Platinum alloys, Assay of, 486. 
 
 ,, Parting of, from gold, 435. 
 ,, Tray for parting, 486. 
 Plattner, C. R, 224, 244, 437, 459. 
 
 process, 224. 
 ,, ,, Amount of water used 
 
 in, 263. 
 
 ,, at Reichenstein, 225. 
 
 ,, at various mills, 296, 
 
 299, 344. 
 
 ,, Cost of, 273, 299. 
 
 , , Method of working, 261 
 
 ,, Reactions in, 267. 
 
 Vats used in, 261. 
 
 Pliny, 90, 397. 
 
 Pliocene gravels of California, 64. 
 Plymouth Consolidated Mill, Chlor- 
 ination at, 297. 
 Pneumatic jig, 197. 
 Poggendorf, 356. 
 
 Pointed boxes, 179. 
 
 Pollok chlorination process, 292. 
 
 Poussee process, 389. 
 
 Power, F. R., 420. 
 
 Prat, 21. 
 
 Precipitation of gold chloride, 25, 
 
 285. 
 
 ,, by charcoal, 289. 
 ., by iron sulphate, 271, 286, 
 
 298, 299. 
 
 ,, by metallic sulphides, 290. 
 ,, by metals, 291. 
 ,, by organic substances, 287. 
 ,, by sulphuretted hydrogen, 
 
 287, 288, 301. 
 
 by sulphurous acid, 12, 308. 
 ,, ,, ,, Use of mo- 
 
 lasses in, 272. 
 ,, of gold alloys, 14. 
 ,, of gold cyanide by sodium, 
 
 377. 
 ,, ,, ,, by electricity, 
 
 337. 
 
 by zinc, 331, 361. 
 ,, Influence 
 of temperature 
 on, 363. 
 
 ,, ,, in leaching vats, 369. 
 
 ,, ,, from solution, 25, 69. 
 
 of silver by salt, 401. 
 
 ,, by copper, 409. 
 Pressure leaching, 285, 306, 312. 
 Price, T., 430. 
 Primitive methods of crushing quartz, 
 
 89. 
 ,, ,, of washing river 
 
 sands, 1, 90 
 Princep, James, 472. 
 Production of gold, Causes of increase 
 
 of, 496 
 Cost of, 494. 
 in the world, 495. 
 by cyanide, 352. 
 in the future, 497. 
 
 ,, past, 495. 
 in various 
 
 countries, 498. 
 Proofs in bullion assay, 481. 
 Properties of gold, 2. 
 Proportion o! gold and silver for 
 
 parting, 399. 
 ,, ,, to gangue in ores, 
 
 38. 
 
 Prospecting trough, 46 
 Protobromide of gold, 27- 
 Proust, 3. 
 Puddling tub, 47. 
 Pure gold, Preparation of, 11. 
 Purple alloy, 14. 
 
INDEX. 
 
 525- 
 
 Purple gold, 3. 
 
 of Cassius, 3, 27, 34. 
 
 Assay of, 462. 
 Pyrites, Assay of, 461. 
 
 ,, Conditions of gold in, 152. 
 Pyritic smelting, 380. 
 
 ,, Austin process, 380, 381. 
 Pyrometer, Le Chatelier, 472. 
 
 Quartz crushing in stamp battery, 
 
 89. 
 Quicksilver, see Mercury. 
 
 Rae, J. H.,319. 
 
 Rand Central Works, 340. 
 Randall, P. M., 74, 124. 
 Rapid City Mill, 231, 311. 
 Raymond, 181. 
 
 Reactions in chlorination vat, 267. 
 Recovery of gold lost in refining, 393. 
 Reduction of silver chloride by zinc, 
 401, 419. 
 
 by iron, 420. 
 
 Reduction of silver sulphate by char- 
 coal, 413. 
 
 by copper, 407. 
 
 by iron, 411. 
 
 by sulphate of iron, 411. 
 Refining of gold bullion, 384, 388. 
 
 by borax, 388. 
 
 by chlorine, 414, 431. 
 
 by nitre, 388 
 
 by sulphur, 395. 
 Regnault, 4. 
 Reichenstein, Plattner process at, 
 
 225. 
 
 Reservoirs for hydraulicking opera- 
 tions, 73. 
 Retorting, 140. 
 
 ,, furnace, 141. 
 Reverberatory furnace, 235. 
 Reynolds, J., 463. 
 Rhodio-platinum couple, 472. 
 Rhodium alloys, Assay of, 487. 
 Richards, J., 233. 
 Rickard, T. A., 207, 209. 
 Ricketts, 448, 455, 457. 
 Riffle bars, 50, 60, 78. 
 
 ,, blocks, 77. 
 Riffled belt in Frue vanner, 192. 
 
 ,, sluice, 182. 
 Rim-rock, 67. 
 Riotte, E. N., 285. 
 River mining, 56. 
 Rivot, 42, 450. 
 Roasting at Deloro, 296. 
 
 ,, at Kennel, California, 296. 
 
 at Treadwell Mine, 298. 
 
 by producer gas, 260. 
 
 Roasting, Cost of, 301, 310, 314. 
 
 ,, furnaces, see Furnaces,. 
 Roasting. 
 
 ,, ores, 234. 
 
 ,, ,, for assay, 447. 
 
 , , oxidising, Chemistry of , 238. 
 
 ,, with salt, 242. 
 Roberts- Austen, 4, 11, 14, 15, 18, 235,. 
 
 238, 431, 467, 471, 480, 490, 491. 
 Robinson Mine, 223, 325, 331, 344. 
 Rock-breakers, 101. 
 Rock-cut sluices, 79. 
 Rock pavements for sluices, 77. 
 Rocker for washing gold, 46. 
 Roller feeder, 123. 
 Rolls for crushing, 228. 
 
 ,, compared with stamps, 232. 
 Resales, H., 42, 178. 
 Rose, G. , 4, 39. 
 Rose, H., 371, 487, 489. 
 Rose, T. K., 5, 6, 19, 24, 27, 472,. 
 
 480, 481. 
 
 Rossler, 479, 480. 
 Rotary drying furnace, 234. 
 Rothschild's refinery, 402. 
 Rothwell, J. E., 230, 259, 300, 303,. 
 
 310. 
 
 Royal College of Science, 342, 446. 
 Royal Mint, Bullion assay at, 463. 
 
 Cupels at, 469. 
 Ruby gold, 3. 
 Russell process, 32. 
 Russia, Gold in, 40. 
 
 ,, Placer mining in, 48, 59. 
 Rusty gold, 151. 
 
 St. John del Key Mine, 181. 
 
 Sal ammoniac used to toughen 
 
 bullion, 389, 390. 
 Salamander crucibles, 387. 
 Salisbury works, 322. 
 Salt used in ore assay, 444, 446. 
 
 ,, roasting, 247. 
 Sampling cone, 438. 
 
 ,, of bullion, 463. 
 
 ,, ,, ores 438. 
 
 ,, tin, 438. 
 Sandberger, F. von, 42. 
 Savot, 399, 463, 466. 
 Scheidel, A. , 344. 
 Schemnitz mill, 165. 
 Schnabel, 166. ' 
 Schonfeld, 398. 
 Schranz crusher, 104. 
 Schwartz's assay of pyrites, 469. 
 Scorification, Assay of alloys by, 484. 
 
 ores by, 455. 
 
 Screens, 96, 97, 116. 
 Selective action of cyanide, 320, 365. 
 
526 
 
 INDEX. 
 
 Selwyn, A. C., 69. 
 
 Settling boxes, 177. 
 
 Shaft mining, 82. 
 
 Shaking copper plates, 134. 
 
 Shallow placer deposits. 43. 
 
 Sheba mine, 220 
 
 ,Shovel, Steam, for placer gravels, 
 
 Siberia, Gold in, 40. 
 
 ,, Method of washing in, 48, 
 
 59. 
 
 .Siberian pan, 61. 
 ,, sluice, 60. 
 ,, trommel, 61. 
 trough, 48. 
 Sickening of mercury, 143. 
 Siemens-Halske process, 337, 347. 
 
 ,, Advantages of, 
 
 347, 351. 
 
 Cost of, 350. 
 
 ,, Disadvantages 
 
 of, 351. 
 
 Silicate of gold, 33. 
 .Silver chloride, Reduction of, 401, 419. 
 ,, Compounds of, dissolved by 
 
 potassium cyanide, 365. 
 ,, dissolved by nitric acid, 400. 
 potassium cyanide, 
 
 360. 
 .,, ,, sulphuric acid, 404, 
 
 409, 412. 
 filter, 413. 
 
 for gold bullion assay, 465. 
 Gold separated from chloride 
 
 of, 417. 
 
 in cornets, 480. 
 ,, gold coins, 16. 
 Simmer & Jack, cyanide works, 325. 
 ,, ,, Concentration at, 212. 
 
 Simpson, J. W., 319. 
 Sizing boxes, 179 
 Skey, Wm., 69, 151, 374, 457. 
 Skimming bullion, 389. 
 , Slime presses, 307. 
 
 ,, separator, 324. 
 , Slimed ore, Treatment of, by cyanide, 
 
 337, 340. 
 
 Slimes, Gold, 332. 
 Sluice plates, 126. 
 Sluicing, 50, 77. 
 Smith, E. A., 37, 452. 
 . Soda, Caustic, use in amalgamation, 
 
 132. 
 ,, ,, cyanide process, 
 
 324, 368. 
 , Sodium amalgam, 144. 
 
 carbonate, used to separate 
 gold from silver chloride, 
 417. 
 
 Sodium hyposulphite, Action on gold 
 
 of, 32. 
 used to precipitate gold 
 
 cyanide, 377. 
 Soetbeer, 497, 502. 
 Solubility of gold, 10, 32, 374, 478. 
 ,, in potassium cyanide, 
 
 254, 357. 
 
 of silver chloride, 12 
 ,, sulphate, 404. 
 ,, sulphates, 406. 
 ,, various metals in potas- 
 sium cyanide, 357, 
 358, 360. 
 Solutions of gold, Tests for, 26. 
 Solvents of gold, 10, 32, 478. 
 Sonstadt, 36. 
 South America, Gold in, 41. 
 
 Clunes United Mill, 210. 
 Spanish Mine, 162. 
 Specific gravity of gold, 4. 
 heat 4. 
 
 Spectroscope, Assay by, 490. 
 Spectrum of gold, 4. 
 Spence, furnace, 248. 
 Spitzluten, 346. 
 Splash box, 110. 
 Spring, 11. 
 Stamp battery, 83, 89, 95, 107. 
 
 ,, Chemicals used in, 131. 
 ,, Foundations of, 108. 
 Framework of, 108. 
 ,, General arrangement 
 
 of, 126. 
 
 ,, practice, 199. 
 Calif ornian, 107, 113. 
 Elephant, 158. 
 German, 99. 
 Husband's 155. 
 Morison's, 158. 
 Steam, 156. 
 
 Weight of parts of, 114, 219. 
 Stamps, Order of fall of, 120. 
 
 , , Weight of, in California, 200. 
 Standard gold of various countries, 
 
 17. 
 
 Stanfield, 197. 
 Stanford's self feeder, 121. 
 Stansfield, A., 486. 
 Stapffs method of assaying pyrites, 
 
 461. 
 Steam shovel for placer gravel, 62. 
 
 ,, stamp, 156. 
 Stelzner, A., 42. 
 Stetefeldt, C. A., 173, 234, 244. 
 
 drying kiln, 234, 261. 
 ,, furnace, 247. 
 
 Stewart mine, 343. 
 Stock of gold, 502. 
 
INDEX. 
 
 527 
 
 Stone-breakers, 101. 
 
 Strength of cyanide solution, Method 
 
 of testing, 369. 
 Strength of cyanide solution required 
 
 for ores, 328. 
 Sulman, H. L., 376. 
 
 -Teed process, 376. 
 Sulphate of copper, Crystallisation 
 
 of, 407. 
 
 Sulphide of gold, 31, 32, 33, 34. 
 Sulphides, Action of potassium 
 cyanide on, 154, 366. 
 Amalgamation of, 152. 
 
 Assay of, 456. 
 
 Gold in, 152, 
 
 Oxidation of, 366. 
 
 ,, Roasting of, 236. 
 
 Smelting of, 380. 
 
 used as fuel, 381. 
 
 Sulphites of gold, 30. 
 Sulphur as fuel, 381. 
 ,, in parting, 387. 
 ,, in refining, 385. 
 Sulphuretted hydrogen as precipitant 
 
 for gold, 287, 308. 
 
 Sulphuric acid, Action on metals of, 
 399. 
 
 Assay by, 486. 
 Parting by, 402. 
 
 Sulphurous acid used to destroy 
 
 chloride, 287 
 Surcharge, 478. 
 
 Effects of base metals on, 
 484. 
 
 ,, Influence of temperature 
 on, 474, 480. 
 
 Variations in, 479. 
 Sweating copper plates, 140. 
 Swedish chlorination process, 293. 
 Sweep, Mint, Assay of, 462. 
 Swinging amalgamated plates, 126. 
 Sydney Mint, Parting at, 415. 
 Sylvanite, 38. 
 Sylvia Mine, Cyanide process at, 344, 
 
 Tail race, 54. 
 
 ,, sluices, 55. 
 Tailings, Examination of, 150. 
 
 ,, from hydrauli eking, 81. 
 Tappet, 111. 
 Teed, Dr., 351. 
 Tellurides of gold, 38. 
 
 Assay of, 457, 488. 
 Tenacity of gold, 3. 
 Test for gold in solution, 26. 
 Testing ores by chlorine, 460. 
 
 ,, cyanide, 341. 
 
 ,, solutions of gold, 271. 
 
 toughness of bullion, 390, 
 
 Thames Valley, N.Z., Battery prac-' 
 
 tice in, 212. 
 Theophilus, 90. 
 Theophrastus, 90. 
 Thies, A., 274, 277. 
 
 ,, process, 277, 292. 
 
 Cost of, 303. 
 
 Thiosulphates of gold, 31. 
 Thompson, L., 414. 
 Thomsen, 21, 357. 
 Thorpe, 5. 
 
 Tin alloys, Assay of, 485. 
 Tina process, 92. 
 
 Tools used in assaying, 444, 449, 469. 
 Touchstone, 489. 
 Transvaal, Cyanide process in, 351. 
 
 ,, Production of gold in, 41, 
 
 223, 500. 
 
 ,, Stamp amalgamation in, 
 
 216. 
 
 Trapiche, 91. 
 Tribomide of gold, 28. 
 Trichloride ,, 21. 
 Triumph concentrator, 193. 
 Trommel, Siberian, 61. 
 Trough, Prospecting, 46. 
 
 Siberian, 48. 
 Tub, Puddling, 47. 
 Tulloch-feeder, 123. 
 Tunnels in drift mining, 82, 83. 
 Turret furnace, 249. 
 Tyrol, Ancient process of milling, 96. 
 Tyrolean bowl, 96. 
 mill, 165. 
 
 Undercurrents, 52. 
 
 United States, Gold in, 41. 
 
 Ural Mountains, Gold mining in, 40 
 
 Vacuum pump for dredging, 58. 
 Van Riemsdijk, 451, 471. 
 Vanner, Frue, 185. 
 
 LUhrig, 193. 
 Van't Hoff, 24. 
 
 Vat process of chlorination, 261. 
 Vats for leaching processes, 261, 322. 
 Vautin, C., 274, 286. 
 Victoria, Gold mining in, 42. 
 Vitruvius, 90. 
 Violle, 4, 6. 
 Volatilisation of alloys of gold, 6, 8. 
 
 of gold, 5, 394, 478. 
 
 ,, Effect of base metals 
 on, 6, 8. 
 
 ,, ,, Effect of temperature 
 
 on, 6. 
 
 Volatility of gold chloride, 22, 245. 
 Voltaic order of metals in potassium 
 cyanide, 356. 
 
528 
 
 INDEX. 
 
 Wales, Gold mining in, 39. 
 Washing gravel by the pan, 45. 
 
 ,, ,, in ancient times, 2, 
 
 90. 
 
 ,, ,, in sluices, 51. 
 
 Water supply for hydraulicking, 72. 
 
 used in barrel chlorination, 
 280. 
 
 ,, ,, cyanide process, 337. 
 
 ,, sluicing, 51. 
 
 ,, stamp battery, 123, 
 
 219. 
 
 , , , , vat chlorination, 263. 
 
 Watson-Denny pan, 214. 
 Watts, 371. 
 Wear of screens, 118. 
 
 ,, shoes and dies, 110. 
 Weight of parts of stamps, 114, 219. 
 Weights, Assay- ton, 441. 
 
 ,, for bullion assaying, 464. 
 Weinberg, 318. 
 Wells, J. S. C., 362, 363. 
 Welman dredge, 58. 
 
 Whisk brooms for copper plates, 131. 
 
 White furnace, 258. 
 
 White-Howell furnace, 259, 304. 
 
 Whitney, 64, 70. 
 
 Wiebe, 5. 
 
 Wilde process, 378. 
 
 Williams, J. R., 338. 
 
 Wilm, 13. 
 
 Wing dams, 56. 
 
 Witwatersrand, see Transvaal. 
 
 Worcester mill, 347. 
 
 World's production of gold, 495. 
 
 Wright, Dr., 354. 
 
 Zinc, Action of potassium cyanide 
 
 on, 361. 
 
 ,, alloys, Assay of, 485. 
 Assay by, 483. 
 ,, Precipitation of gold cyanide 
 
 by, 331, 339, 362. 
 T , Reduction of silver chloride 
 
 by, 401, 419. 
 
 BELT. AND BAIN, LIMITED, PRINTERS. GLASGOW. 
 
T X 3 M I 
 
 A CATALOGUE 
 
 OP 
 
 SCIENTIFIC AND TECHNICAL WORKS 
 
 PUBLISHED BY 
 
 CHARLES GRIFFIN A COMPANY, 
 
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 PUBLICATIONS may be obtained through any Bookseller 
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INDEX TO AUTHORS. 
 
 TT 
 
 PAGE 
 
 ALLINGHAM (W.), Navigation, . 5, 21 
 
 Meteorology, 21 
 
 ANGLIN (S.), Design of Structures, . 4 
 BARKER (D. WILSON-), Navigation 
 
 and Seamanship, . . . . 5, 21 
 BERING-ER (J. J. & C.), Assaying, . 8 
 BLACK MORE (E.), British Mercantile 
 
 Marine 6, 21 
 
 BLOUN't(B.)and BLOXAM (A.G.), 
 Chemistry for Engineers and Manu- 
 facturers, .... .7 
 BLYTH (A. Wynter), Foods and Poisons, 9 
 BORCHERS (Dr.), Electric Smelting, 29 
 BROTHERS (A.), Photography, . 
 
 BROUGH (B. H.), Mine-Surveying, 
 BROWNE (W. R.), Fuel and Water, 
 
 Mechanics, 
 
 BUCK (R. C.), Algebra and Trigono- 
 metry, 6, 21 
 
 BUTTERFIELD (W, J. A.), Gas 
 Manufacture, 11 
 
 COLE (Prof.), Practical Geology, . . 12 
 
 Open Air Studies in Geology, . . 12 
 
 CRIMP (W. S.), Sewage Disposal Works, 13 
 VIES (Surgeon-Major), Hygiene, . 14 
 
 DAVIEJ 
 
 DAVIS (Prof.), Biology,. 
 The Flowering Plant, 
 
 - Zoological Pocket- Book, ... 15 
 DONKIN( Bryan), Gas and Oil Engines, 16 
 
 - Efficiency of Steam Boilers, . . 16 
 DUERR (Geo.), Bleaching and Calico- 
 
 Printing ....... 14 
 
 DUPRE&HAKE(ManualofChemistry),16 
 ETHERIDGE (R.), Stratigraphical 
 
 32 
 
 eoogy, ....... 
 
 EW ART (Prof.), Preservation of Fish, . 16 
 FIDLER (Prof.), Bridge- Construction, 17 
 FOSTER (Prof. C. le Neve), Ore and 
 
 Stone Mining 
 
 18 
 
 ...... 
 
 GRIFFIN ( J. J.), Chemical Recreations, 19 
 
 GRIFFIN'S Electrical Price- Book, . 19 
 GRIFFIN'S ENGINEERING PUB- 
 
 LICATIONS ...... 20 
 
 GRIFFIN'S GEOLOGICAL and 
 
 METALLURGICAL SERIES, . 22 
 
 GRIFFIN'S NAUTICAL SERIES, . 21 
 GRIFFIN'S SANITARY PUBLICA- 
 
 TIONS, . ... 22 
 GRIFFIN'S TECHNOLOGICAL 
 
 PUBLICATIONS, . . .23 
 
 GURDEN (R.), Traverse Tables, . 19 
 
 GUTTMANN (O.), Blasting, . 19 
 HARBORD (F. W.), Steel, . 41, 50 
 HUGHES-GIBB (E.), How Plants Live, 17 
 
 HUGHES (H. W.), Coal Mining, . 24 
 
 HURST (Chas.), Valves, . . 23 
 
 HURST (G. H.), Painters' Colours, . 25 
 
 - Garment Dyeing and Cleaning . 25 
 
 - Colour Theory (Practical Applica- 
 
 tions of), ..... 26 
 JAMIESON (Prof.), MANUALS, Ad- 
 vanced and Elementary, . .26 
 
 - Steam Engine, ..... 26 
 
 - Applied Mechanics, .... 28 
 
 - Magnetism and Electricity, . . 26 
 JENKINS (H. C.), Metallurgical 
 
 Machinery, ... . . 41 
 
 JOHNSON (J. C. F.), Getting Gold, . 28 
 
 KN ECHT&RAWSON, Dyeing, . . 27 
 LAFAR (Dr. F.), Micro-Organisms 
 
 (Utilization of) ...... 47 
 
 PAGE 
 
 LAWN (J. G.), Mine Accounts, . . 28 
 MACKENZ IE (T.), Applied Mechanics.21,30 
 M 'MIL LAN (W. G.), Electro - Metal- 
 lurgy, 29 
 
 & BORCHERS, Electric Smelting, 29 
 
 MILLAR (W. J.), Latitude and Longi- 
 tude 21,30 
 
 MUNRO & JAMIESON'S Electrical 
 
 Pocket- Book, ... . 31 
 
 MUNRO (Dr.), Agricultural Chemistry, 31 
 MUNRO (R. D.), Steam- Boilers, . . 30 
 
 Kitchen Boiler Explosions, . . 30 
 
 N YSTROM'S Pocket-Book for Engineers, 31 
 PEARCE (W. J.), Painting and Decor- 
 ating, 33 
 
 PHILLIPS & B AUERM AN, Metallurgy, 34 
 POYNTING (Prof.), Mean Density of 
 
 the Earth, 34 
 
 PRAEGER (R. L.), Open Air Botany,. 33 
 RANKINE'S Applied Mechanics, . 35 
 
 Civil Engineering 35 
 
 Machinery and Millwork, . . 35 
 
 Steam Engine & other Prime Movers, 35 
 
 Useful Rules and Tables. ... 35 
 
 A Mechanical Text-Book. . . 35 
 
 Miscellaneous Scientific Papers, . 36 
 
 REDGRAVE(G. R.), Cements, . . 36 
 RED WOOD (Boverton), Petroleum, . 37 
 REKD (Sir E. J.), Stability of Ships, . 38 
 REID (Geo., M.D.), Practical Sanitation, 39 
 RICHMOND (H. D.), Dairy Chemistry, 38 
 RIDDELL (Dr. Scott), Ambulance, . 40 
 RIDEAL (S., D.Sc.), Disinfection, . 39 
 ROBERTS -AUSTEN (Prof.), Metal- 
 lurgy and Alloys 41 
 
 ROBINSON (Prof.), Hydraulics, . . 42 
 ROSE (T. K.), Gold, Metallurgy of, . 
 ROTH WELL, (C. F. Seymour), 
 
 Textile Printing, 44 
 
 SALTER (Chas.), Micro-Organisms, . 47 
 SEATON (A. E.), Marine Engineering, 45 
 SE ATON & ROUNTH WAITE, Marine 
 
 Engineers' Pocket- Book, . . .45 
 SEELEY (Prof.), Physical Geology, . 32 
 SEXTON (Prof.), Elementary Metal- 
 lurgy 46 
 
 Quantitative & Qualitative Analysis, 46 
 
 SH ELTON (W. V.), Mechanic's Guide, 44 
 SMITH, (Johnson), Shipmaster's Medical 
 
 Help, 21,40 
 
 SMITH (Prof. R. H.), Measurement 
 Conversions, ...... 
 
 Calculus for Engineers, . 
 
 SYKES (Dr. W. J.), Brewing, 
 THOMSON & POYNTING (Profs.) 
 Text-Book of Physics, .... 
 TRAILL (T. W.), Boilers, Land and 
 
 Marine, 
 
 TURNER (Thos.), Iron, Metallurgy of, 
 WALTON (T.), Know Your Own Ship, 
 
 21,51 
 W ATKINSON (Prof. W. H.), Gas and 
 
 Oil Engines, 51 
 
 WELLS (S. H.), Engineering Drawing, 53 
 WIGLEY (T. B.), Goldsmith and 
 
 Jeweller's Art, 53 
 
 WRIGHT (Dr. Alder), The Threshold 
 
 of Science 52 
 
 Oils and Fats, Soaps and Candles, . 52 
 
 YEAMAN (C. H.), Elec. Measurements, 53 
 YEAR-BOOK of Scientific Societies, . 64 
 
 43 
 
INDEX TO SUBJECT 
 
 PAGE 
 AGRICULTURAL CHEMISTRY, . 31 
 
 AIR ENGINES, 16 
 
 ALGEBRA 6 
 
 ALLOYS, 41 
 
 AMBULANCE, 40 
 
 ASSAYING, 8 
 
 BACTERIOLOGY (Practical), . . 47 
 
 BIOLOGY, 15 
 
 BLASTING, 19 
 
 BLEACHING, 14 
 
 BOTANY, . . . .15, 17, 33, 47 
 BRIDGE-CONSTRUCTION, . . 17 
 BOILERS, Construction, ... 49 
 
 Efficiency of, .... 16 
 
 Kitchen, Explosions of, . . 30 
 
 Management, . . . .30 
 
 BREWING 47 
 
 CALCULUS FOR ENGINEERS, . 48 
 CALICO-PRINTING, . .14, 44 
 
 CEMENTS, 36 
 
 CHEMICAL ANALYSIS, Qualitative 
 
 and Quantitative, . . . .46 
 CHEMICAL RECREATIONS, . 19, 52 
 CHEMISTRY FOR ENGINEERS, . 7 
 
 for Manufacturers, ... 7 
 
 Inorganic, . . . . .16 
 
 COAL-MINING, 24 
 
 COLOURS, 25 
 
 COPPKR, Metallurgy of, . . . 41 
 DAIRY CHEMISTRY, ... 38 
 DENSITY OF THE EARTH, . . 34 
 DESIGN (Engineering), ... 53 
 
 of Structures, .... 4 
 
 DISINFECTION and DISINFECTANTS,39 
 DRAWING (Engineering), ... 53 
 DYEING, . . 25, 27 
 
 ELECTRIC SMELTING, ... 29 
 ELECTRICAL MEASUREMENTS, . 53 
 
 Pocket-Book 31 
 
 Price-Book, . . . . . 19 
 
 ELECTRICITY and MAGNETISM, . 26 
 ELECTRO-METALLURGY, . . 29 
 ENGINEERS, Pocket-books for, . 31, 45 
 
 Useful Rules for, ... 20, 35 
 
 ENGINEERING, Civil, ... 35 
 
 Marine, 45 
 
 Drawing and Design, . . .53 
 
 FERMENTATION, . . . . 47 
 FISH, Preservation of, . . . .16 
 FOODS, Analysis of, . . . .9 
 FUEL and WATER, .... 44 
 
 GAS ENGINES 16,51 
 
 GAS MANUFACTURE, ... 11 
 GEOLOGY, Introduction to, . . 12 
 
 Physical, 32 
 
 Practical, 12 
 
 Stratigraphical, . . . 32 
 
 GOLD, Getting, 
 
 Metallurgy of. 
 
 GOLDSMITH'S ART, 
 HYDRAULICS, . 
 HYGIENE, 
 
 . 28 
 . 43 
 . 53 
 . 42 
 13, 14, 39 
 
 PAO 
 
 IRON, Metallurgy of, . . . .50 
 JEWELLER'S ART, .... 53 
 KITCHEN BOILER EXPLOSIONS, . 30 
 LATITUDE and LONGITUDE, 
 
 To Find, . . . .21,30 
 
 MACHINE DESIGN 53 
 
 MACHINERY and MILLWORK, 33, 41 
 
 MAGNETISM, 26 
 
 MARINE BOILERS, .... 49 
 
 Engineering, . ... 45 
 
 MECHANICS, Applied 
 
 Elementary, . . . 21, 26, 30 
 
 Student's, 10 
 
 Advanced, . . . . 26, 35 
 
 MEASUREMENTS, Conversion of, . 48 
 
 Electrical, 53 
 
 MEDICINE and SURGERY, Domestic, 55 
 
 for Shipmasters, . . . .40 
 
 MERCANTILE MARINE, British, . 6 
 METALLURGY, Elementary, . . 46 
 
 Introduction to, . . . .41 
 
 General, 34 
 
 of Gold, 48 
 
 of Iron, ..... 50 
 
 MICRO-ORGANISMS, Utilization of, 
 
 in the Arts and Manufactures, . 47 
 MINING- 
 
 Accounts and Book-keeping, . 28 
 
 Blasting, ... . 19 
 
 Coal, .... .24 
 
 Ore and Stone, . . .18 
 
 Surveying, ... .10 
 
 NAUTICAL SUBJECTS, . . 21 
 
 NAVIGATION, ... .5 
 
 OILS and FATS, . . . .52 
 
 OIL ENGINES, . 16 
 
 ORE MINING, 18 
 
 PAINTING and DECORATING, , 33 
 
 PETROLEUM, 37 
 
 PHOTOGRAPHY, .... 8 
 PHYSICS, .... 26,49,52 
 POISONS, Detection of, . . .9 
 SANITATION, ... 13, 14, 39 
 SCIENCE, Popular Introduction to, . 52 
 
 SEAMANSHIP, 5 
 
 SEWAGE, Disposal of, ... 18 
 SHIPMASTER'S MEDICAL GUIDE, 40 
 SHIPS, Loading, &c., of, . . 51 
 
 Stability of, . . . 38, 51 
 
 SOAPS, Manufacture of, . . 52 
 
 SOCIETIES, Year-book of, . . 54 
 
 STEAM ENGINE, Elementary, . 26 
 
 Advanced. . . . 26, 45 
 
 STEEL, Metallurgy of, . 41, 50 
 
 STRUCTURES, Design of, . .4 
 
 SURVEYING, Mine, . 10 
 
 TEXTILE PRINTING, 14 44 
 
 TRAVERSE TABLES, . ,19 
 
 TRIGONOMETRY, . . 6 
 
 VALVES and VALVE-GEARING, . 23 
 VARNISHES, Manufacture of, . 25 
 
 ZOOLOGY, 15 
 
3 id I 
 
 CHARLES GRIFFIN A CO.'S PUBLICATIONS. 
 
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 With Frontispiece, Twelve Plates (Two in Colours), and Illustrations 
 
 in the Text. 
 
 GENERAL CONTENTS. The Building of a Ship; Parts of Hull, Masts, 
 &c. Ropes, Knots, Splicing, &c. Gear, Lead and Log, &c. Rigging, 
 Anchors Sailmaking The Sails, &c. Handling of Boats under Sail 
 Signals and Signalling Rule of the Road Keeping and Relieving Watch 
 Points of Etiquette Glossary of Sea Terms and Phrases Index. 
 
 %* The volume contains the NEW RULES OF THE ROAD. 
 
 " This ADMIRABLE MANUAL, by OAPT. WiLSON-BAKKER of the u Worcester," seems to us 
 PERFECTLY DESIGNED, and holds its place excellently in ' GRIFFIN'S NAUTICAL SERIES.' . . . 
 Although intended for those who are to become Officers of the Merchant Navy, it will be 
 found useful by ALL YACHTSMEN." A thenaeum. 
 
 " Five shillings will be WELL SPENT on this little book. CAPT. WILSON-BARKER knows 
 from experience what a young man wants at the outset of his career." The Engineer. 
 
 Price 3s. 6d. Post-free. 
 
 NAVIGATION: 
 
 PRACTICAL A1VO THEORETICAL. 
 
 BY DAVID WILSON-BARKER, R.N.R., F.R.S.E., &c., <fcc., 
 
 AND 
 
 WILLIAM ALLINGHAM, 
 
 PIRST-CLASS HONOURS, NAVIGATION, SCIENCE AND ART DEPARTMENT. 
 
 lldftb "Numerous ^lustrations ano ^Examination Questions. 
 
 GENERAL CONTENTS. Definitions Latitude and Longitude Instruments 
 of Navigation Correction of Courses Plane Sailing Traverse Sailing Day's 
 Work Parallel Sailing Middle Latitude Sailing Mercator's Chart 
 M creator Sailing Current Sailing Position by Bearings Great Circle Sailing 
 The Tides Questions Appendix : Compass Error Numerous Useful Hints, 
 &c. Index. 
 
 " PRECISELY the kind of work required for the New Certificates of competency in grades 
 from Second Mate to extra Master. . . . Candidates will find it INVALUABLE." Dundee 
 Advertiser 
 
 "A CAPITAL LITTLE BOOK . . . specially adapted to the New Examinations. The 
 Authors are CAPT. WILSON-BARKER (Captain-Superintendent of the Nautical College, H.M.S. 
 " Worcester," who has had great experience on the highest problems of Navigation), and 
 MR. ALLINGHAM, a well-known writer on the Science of Navigation and Nautical Astronomy." 
 Shipping World. 
 
 VFor complete list of GRIFFIN'S NAUTICAL SERIES see p. 2L 
 
 LONDON: EXETER STREET, STRAND. 
 
6 CHARLES GRIFFIN * CO.'S PUBLICATIONS. 
 
 GRIFFIN'S NAUTICAL SERIES. 
 
 Price 3s, 6d. Post-free. 
 T IHIIE 
 
 British Mercantile Marine. 
 
 BY EDWARD BLACKMORE, 
 
 MASTER MARINER; ASSOCIATE OF THE INSTITUTION OF NAVAL ARCHITECTS; 
 
 FORMERLY A RESIDENT JUSTICE OF THE PEACE FOR THE COUNTY 
 
 OF RENFREW. N.B.; AND A MEMBER OF THE INSTITUTION 
 
 OF ENGINEERS AND SHIPBUILDERS IN SCOTLAND ; 
 
 EDITOR OF GRIFFIN'S "NAUTICAL SERIES." 
 
 GENERAL CONTENTS. HISTORICAL : From Early Times to 1486 Progress 
 under Henry VIII. To Death of Mary During Elizabeth's Reign-Up to 
 the Reign of William III. The 18th and 19th Centuries Institution of 
 Examinations Rise and Progress of Steam Propulsion Development of 
 Free Trade Shipping Legislation, 1862 to 1875 " Locksley Hall" Case- 
 Shipmasters' Societies Loading of Ships Shipping Legislation, 1884 to 1894 
 Statistics of Shipping. THE PERSONNEL : Shipowners Officers Mariners- 
 Duties and Present Position. EDUCATION : A Seaman's Education : what it 
 should be Present Means of Education Hints. DISCIPLINE AND DUTY 
 Postscript The Serious Decrease in the Number of British Seamen, a Matter 
 demanding the Attention of the Nation. 
 
 " INTERESTING and INSTRUCTIVE . . . may be read WITH PROFIT and ENJOYMENT." 
 Glasgow Herald. 
 
 "EVERT BRANCH of the subject is dealt with in a way which shows that the writer 
 'knows the ropes' familiarly." Scotsman. 
 
 "This ADMIBABLE book . . . TEEMS with useful information Should be in the 
 hands of every Sailor." Western Morning News. 
 
 WORKS BY RICHARD C. BUCK, 
 
 of the Thames Nautical Training College, H.M.S. ' Worcester.' 
 
 1. A Manual of Trigonometry: 
 
 With Diagrams, Examples, and Exercises. Price 3s. 6d. 
 
 *+* Mr. Buck's Text-Book has been SPECIALLY PREPARED with a view- 
 to the New Examinations of the Board of Trade, in which Trigonometry 
 is an obligatory subject. 
 
 2. A Manual of Algebra. 
 
 [In Preparation. 
 
 %* These elementary works on ALGEBRA and TRIGONOMETRY are written specially for 
 those who will have little opportunity of consulting a Teacher. They are books for "SELF- 
 HELP." All but the simplest explanations have, therefore, been avoided, and ANSWERS to 
 the Exercises are given. Any person may readily, by careful study, become master of their 
 contents, and thus lay the foundation for a further mathematical course, if desired. It is 
 hoped that to the younger Officers of our Mercantile Marine they will be found decidedly 
 serviceable. The Examples and Exercises are taken from the Examination Papers set for 
 the Cadets of the " Worcester." 
 
 %* For complete list of GRIFFIN'S NAUTICAL SERIES see p. 21. 
 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 7 
 
 CHEMISTRY FOR ENGINEERS 
 AND MANUFACTURERS. 
 
 A PRACTICAL TEXT-BOOK. 
 
 BY 
 
 BERTRAM BLOUNT, AND A. G. BLOXAM, 
 
 F.I.C., F.C.S., F.I.C., F.C.S., 
 
 nsulting Chemist to the Crown Agents for Consulting Chemist, Head of the Chemi 
 
 the Colonies. Department, Goldsmiths' Inst., 
 
 New Cross. 
 
 With Illustrations. In Two Vols., Large 8vo. Sold Separately. 
 
 "The authors have SUCCEEDED beyond all expectation, and have produced a work which 
 should give FRESH POWER to the Engineer and Manufacturer." The Times. 
 
 I. Pvice 10s. 6d. 
 
 CHEMISTRY OF ENGINEERING, BUILDING, AND 
 METALLURGY. 
 
 General Contents. INTRODUCTION Chemistry of the Chief Materials 
 of Construction Sources of Energy Chemistry of Steam -raising Chemis- 
 try of Lubrication and Lubricants Metallurgical Processes used in the 
 Winning and Manufacture of Metals. 
 
 "PRACTICAL THROUGHOUT ... an ADMIRABLE TEXT-BOOK, useful not only to Students, 
 but to ENGINEERS and MANAGERS OF woaxs in PREVENTING WASTE and IMPROVING PROCESSES." 
 Scotsman. 
 
 " EMINENTLY PRACTICAL." Glasgow Herald. 
 
 "A book worthy of HIGH RANK . . . its merit is great . . . treatment of the subject 
 of GASEOUS FUEL particularly good. , . . WATER GAS and its production clearly worked out. 
 . . . Altogether a most creditable production. We WARMLY RECOMMEND IT, and look forward 
 with keen interest to the appearance of Vol. II." Journal of Gas Lighting. 
 
 II. Px*ioe 16s. 
 
 THE CHEMISTRY OF MANUFACTURING 
 PROCESSES. 
 
 General Contents. Sulphuric Acid Manufacture Manufacture of Alkali, 
 &c. Destructive Distillation -Artificial Manure Manufacture Petroleum 
 Lime and Cement Clay Industries and Glass Sugar and Starch Brewing 
 and Distilling Oils, Resins, and Varnishes Soap and Candles Textiles 
 and Bleaching Colouring Matters, Dyeing, and Printing Paper and 
 Pasteboard Pigments and Paints Leather, Glue, and Size Explosives 
 and Matches Minor Chemical Manufactures. 
 
 "Certainly a GOOD and USEFUL BOOK, constituting a PRACTICAL GUIDE for students by 
 affording a clear conception of the numerous processes as a whole." Chemical Trade 
 Journal. 
 
 "We CONFIDENTLY RECOMMEND this volume as a PRACTICAL, and not overloaded 
 TEXT-BOOK, of GREAT VALUE to students." The Builder. 
 
 LONDON : EXETER STREET, STRAND. 
 
8 CBARLBS GRIFFIN * OO.'S PUBLICATIONS. 
 
 ASSAYING (A Text-Book of): 
 
 For the use of Students, Mine Managers, Assayers, dc, 
 BY J. J. BERINGER, F.I.C., F.C.S., 
 
 Public Analyst for, and Lecturer to the Mining Association of, Cornwall. 
 
 AND C. BERINGER, F.C.S., 
 
 Late Chief Assayer to the Rio Tinto Copper Company, London, 
 
 With numerous Tables and Illustrations. Crown 8vo. Cloth, 10/6. 
 FOURTH EDITION; Revised. 
 
 GENERAL CONTENTS. PART I. INTRODUCTORY ; MANIPULATION: Sampling; 
 Drying ; Calculation of Results Laboratory-books and Reports METHODS : Dry Gravi- 
 metric ; Wet Gravimetric Volumetric Assays : Titrometric, Colorimetric, Gasometric 
 Weighing and Measuring Reagents Formulae, Equations, &c. Specific Gravity. 
 
 PART II. METALS : Detection and Assay of Silver, Gold, Platinum, Mercury, Copper, 
 Lead, Thallium, Bismuth, Antimony, Iron, Nickel, Cobalt, Zinc, Cadmium, Tin, Tungsten, 
 Titanium, Manganese, Chromium, &c. Earths, Alkalies. 
 
 PART III. NON-METALS : Oxygen and Oxides; The Halogens Sulphur and Sul- 
 phatesArsenic, Phosphorus, Nitrogen Silicon, Carbon, Boron Useful Tables. 
 
 "A RALLY MERITORIOUS WORK, that may be safely depended upon ekhr for systematic 
 instruction or for reference." Nature. 
 
 " This work is one of the BEST of its kind. . . . Essentially of a practical character. 
 . . . Contains all the information that the Assayer will find necessary in the examination 
 f minerals." Ettfinter. 
 
 New Edition in Preparation. 
 
 PHOTOGRAPHY: 
 
 ITS HISTORY, PROCESSES, APPARATUS, AND MATERIALS. 
 
 Comprising Working Details of all the More Important Methods. 
 
 BY A. BROTHERS, F.R.A.S. 
 
 WITH TWENTY-FOUR FULL PAGE PLATES BY MANY OF THE PRO- 
 CESSES DESCRIBED, AND ILLUSTRATIONS IN THE TEXT. 
 
 In 8vo, Handsome Cloth. 
 
 GENERAL CONTENTS. PART. I. INTRODUCTORY Historical 
 Sketch; Chemistry and Optics of Photography; Artificial Light. 
 PART II. Photographic Processes. PART III. Apparatus. PART IV. 
 Materials. PART V. Applications of Photography ; Practical Hints. 
 
 " Mr. Brothers has had an experience in Photography so large and varied that any work 
 by him cannot fail to be interesting and valuable. ... A MOST COMPREHENSIVE volume, 
 entering with full details into the various processes, and VERY FULLY illustrated. The 
 PRACTICAL HINTS are of GREAT VALUE. . . . Admirably got up." Brit. Jour, of Photography. 
 
 " For the Illustrations alone, the book is most interesting ; but, apart from these, the 
 volume is valuable, brightly and pleasantly written, and MOST ADMIRABLY ARRANGED." 
 Photographic News. 
 
 " Certainly the FINEST ILLUSTRATED HANDBOOK to Photography which has ever been 
 published. Should be on the reference shelves of every Photographic Society." Amattur 
 Photographer. 
 
 "A handbook so far in advance of most others, that the Photographer must not fail 
 to obtain a copy as a reference work." Photographic Work. 
 
 "The COMPLETEST HANDBOOK of the art which has yet been published." Scot. 
 
 LONDON: EXETER STREET, STRAND, 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. g 
 
 WORKS BY A. WYNTER BLYTH, M.R.C.S., F.C.S., 
 
 Barrister-at-Law, Public Analyst for the County of Devon, and Medical Officer of Health for 
 
 St. Marylebone. 
 
 FOODS: 
 
 THEIR COMPOSITION AND ANALYSIS. 
 
 In Demy 8vo, with Elaborate Tables, Diagrams, and Plates. Handsome 
 Cloth. FOURTH EDITION. Price 21s. 
 
 GENERAL CONTENTS. 
 
 History of Adulteration Legislation, Past and Present Apparatus 
 useful to the Food- Analyst "Ash" Sugar Confectionery Honey 
 Treacle Jams and Preserved Fruits Starches Wheaten-Flour Bread 
 Oats Barley Rye Rice Maize Millet Potato Peas Chinese 
 Peas Lentils Beans MILK Cream Butter Oleo- Margarine 
 Butterine Cheese Lard Tea Coffee Cocoa and Chocolate Alcohol 
 Brandy Rum Whisky Gin Arrack Liqueurs Absinthe Principles 
 of Fermentation Yeast Beer Wine Vinegar Lemon and Lime 
 Juice Mustard Pepper Sweet and Bitter Almond Annatto Olive 
 Oil WATER Standard Solutions and Reagents. Appendix : Text of 
 English and American Adulteration Acts. 
 
 PRESS NOTICES OF THE FOURTH EDITION. 
 
 " Simply INDISPENSABLE in the Analyst's laboratory." The Lancet. 
 
 "THE STANDARD WORK on the subject. . . . Every chapter and every page gives 
 abundant rroof of the strict revision to which the work has been subjected. . . . The 
 section on MILK is, we believe, the most exhaustive study of the subject extant. ... An 
 INDISPENSABLE MANUAL for Analysts and Medical Officers of Health." Public Health. 
 
 " A new edition of Mr. Wynter Blyth's Standard work, ENRICHED WITH ALL THE RECBNT 
 DISCOVERIES AND IMPROVEMENTS, will be accepted as a boon." Chemical News. 
 
 POISONS; 
 
 THEIR EFFECTS AND DETECTION. 
 
 THIRD EDITION. In Large 8vo, Cloth, with Tables and Illustrations. 
 
 Price 21s. 
 
 GENERAL CONTENTS. 
 
 I. Historical Introduction. II. Classification Statistics Connection 
 between Toxic Action and Chemical Composition Life Tests General 
 Method of Procedure The Spectroscope Examination of Blood and Blood 
 Stains. III. Poisonous Gases. IV. Acids and Alkalies. V. More 
 or less Volatile Poisonous Substances. VI. Alkaloids and Poisonous 
 Vegetable Principles. VII. Poisons derived from Living or Dead Animal 
 Substances. VIII. The Oxalic Acid Group. IX. Inorganic Poisons. 
 Appendix : Treatment, by Antidotes or otherwise, of Cases of Poisoning. 
 
 " Undoubtedly THB MOST COMPLETE WORK on Toxicology in our language." The Analyst (on 
 the Third Edition). 
 
 "As a PRACTICAL GUIDE, we know NO BETTER work." The Lancet (on the Third Edition}. 
 V In the THIRD EDITION, Enlarged and partly Re-written, NEW ANALYTICAL METHODS have 
 been introduced, and the CADAVERIC ALKALOIDS, or PTOMAINES, bodies playing so great a part in 
 Food-poiioning and in the Manifestations of Disease, have received special attention. 
 
 LONDON : EXETER STREET, STRAND. 
 
io CHARLES QRIFPIN it CO.'S PUBLICATIONS. 
 
 MINE-SURVEYING (A Text-Book of): 
 
 For the use of Managers of Mines and Collieries, Students 
 at the Royal School of Mines, dc. 
 
 BY BENNETT H. B R O U G H, F.G.S., 
 
 Late Instructor of Mine-Surveying, Royal School of Mines. 
 With Diagrams. SIXTH EDITION, Enlarged and Revised. Cloth, 7s. 6d. 
 
 GENERAL CONTENTS. 
 
 General Explanations Measurement of Distances Miner's Dial Variation of 
 the MajEjnetic-Needle Surveying with the Magnetic-Needle in presence of Iron 
 Surveying vrith the Fixed Needle German Dial Theodolite Traversing Under- 
 ground Surface-Surveys with Theodolite Plotting the Survey Calculation of 
 Areas Levelling Connection of Underground- and Surface-Surveys Measuring 
 Distances by Telescope Setting-out Mine-Surveying Problems Mine Plans 
 Applications of Magnetic-Needle in Mining Appendices. 
 
 " Has PROVED itself a VALUABLE Text-book ; the BEST, if not the only one, in the English 
 language on the subject." Mining Journal. 
 
 " No English-speaking Mine Agent r Mining Student will consider his technical library 
 complete without it.'' Nature. 
 
 "A valuable accessory to Surveyors in every department of commercial enterprise. 
 Fnlly deserves to hold its position as a STANDARD.** CUury G^tartUatt. 
 
 O UK S 
 
 BY WALTER R. BROWNE, M.A., M. INST. C.E., 
 
 Late Fellow of Trinity College, Cambridge. 
 
 THE STUDENT'S MECHANICS: 
 
 An Introduction to the Study of Force and Motion. 
 
 With Diagrams. Crown 8vo. Cloth, 43. 6d. 
 
 " Oear in style and practical in method, 'THE STUDENT'S MECHANICS' is cordially to b 
 commended from all points of view. " A then&um. 
 
 FOUNDATIONS OF MECHANICS. 
 
 Papers reprinted from the Engineer. In Crown 8vo, is. 
 
 FUEL AND WATER: 
 
 A Manual for Users of Steam and Water. 
 BY PROF. SCHWACKHOFER AND W. R. BROWNE, M.A. (Sec p. 44.) 
 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. n 
 
 GAS MANUFACTURE 
 
 (THE CHEMISTRY OF). 
 
 A Hand-Booh on the Production, Purification, and Testing 
 of Illuminating Gas, and the Assay of the Bye- 
 Products of Gas Manufacture. For the 
 Use of Students. 
 
 BY 
 
 W. J. ATKINSON BUTTERFIELD, M.A., F.C.S., 
 
 Head Chemist, Gas Works, Beckton, London, E. 
 With Numerous Illustrations. Handsome Cloth. Price 9s. 
 
 " The BEST WORK of its kind which we have ever had the pleasure of re- 
 viewing." Journal of Gas Lighting. 
 
 GENERAL CONTENTS. 
 
 Mann- 
 
 I. Raw Materials for Gas 
 Manufacture. 
 II. Coal Gas. 
 III. Carburetted Water Gas. 
 IV. Oil Gas. 
 V. Enriching by Light Oils. 
 
 VI. 
 
 VII. 
 VIII. 
 IX. 
 X. 
 
 Final Details of M 
 facture. 
 Gas Analysis. 
 Photometry. 
 Applications of Gas. 
 Bye-Products. 
 
 * # * This work deals primarily with the ordinary processes of GAS MANUFACTURE 
 employed in this country, and aims especially at indicating the principles on which 
 they are based. The more modern, but as yet subsidiary, processes are fully treated 
 also. The Chapters on Gas Analysis and Photometry will enable the consumer to 
 grasp the methods by which the quality of the gas he uses is ascertained, and in the 
 Chapter on The Applications of Gas, not only is it discussed as an illuminant, but 
 also as a ready source of heat and power. In the final Chapter, an attempt has been 
 made to trace in a readily-intelligible manner the extraction of the principal derivatives 
 from the crude BYE-PKODUCTS. The work deals incidentally with the most modern 
 developments of the industry, including inter alia the commercial production and 
 uses of acetylene and the application of compressed gas for Street Traction. The needs 
 of the Students in Technical Colleges and Classes have throughout been kept in view. 
 
 LONDON : EXETER STREET, STRAND. 
 
12 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 Works by GRENVILLE A. J. COLE, M.R.I.A., F.G.S, 
 
 Professor of Geology in the Royal College of Science for Ireland. 
 
 PRACTICAL GEOLOGY 
 
 (AIDS IN): 
 
 WITH A SECTION ON PALMONTOLOGY. 
 SECOND EDITION, Revised. With Illustrations. Cloth, IDS. 6d. 
 
 GENERAL CONTENTS. PART I. SAMPLING OF THE EARTH'S 
 CRUST. PART II. EXAMINATION OF MINERALS. PART III. EXAMINA- 
 TION OF ROCKS. PART IV. EXAMINATION OF FOSSILS, 
 
 " Prof. Cole treats of the examination of minerals and rocks in a way that has 
 been attempted before . . . DESERVING OF THB HIGHEST PRAISE. Here indeed are 
 ' Aids ' INNUMERABLE and INVALUABLE. All the directions are given with the utmost clear- 
 ness and precision." Atktna-um, 
 
 "To the younger workers in Geology, Prof. Cole's book will be as INDISPENSABLE as a 
 dictionary to the learners of a language." Saturday Review. 
 
 "That the work deserves its title, that it is full of ' AIDS,' and in the highest degree 
 ' PRACTICAL,' will be the verdict of all who use it." Nature. 
 
 " This EXCELLENT MANUAL . . . will DC A VERY GREAT HELP. . . . The section 
 
 on the Examination of Fossils is probably the BEST of its kind yet published. . . FULL 
 of well-digested information from the newest sources and from personal research." Annals 
 +f Nat. History. 
 
 OPEN-AIR STUDIES: 
 
 An Introduction to Geology Out-of-doors. 
 
 With 12 Full-Page Illustrations from Photographs. Cloth. 8s. 6d. 
 
 GENERAL CONTENTS. The Materials of the Earth A Mountain Hollow 
 Down the Valley Along the Shore Across the Plains Dead Volcanoes 
 A Granite Highland The Annals of the Earth The Surrey Hills The 
 Folds of the Mountains. 
 
 "The FASCINATING ' OPEN-AIR STUDIES' of PROF. COLE give the subject a GLOW OF 
 ANIMATION . . . cannot fail to arouse keen interest in geology." Geological Magazine. 
 
 "EMINENTLY READABLE . . . every small detail in a scene touched with a sym- 
 pathetic kindly pen that reminds one of the lingering brush of a Constable." Nature. 
 
 " The work of Prof. Cole combines ELEGANCE of STYLE with SCIENTIFIC THOROUGHNESS." 
 Petermann's Mittheilunyen. 
 
 " The book is worthy of its title; from cover to cover it is STRONG with bracing freshness 
 of the mountain and the field, while its ACCURACY and THOROUGHNESS show that it is the 
 work of an earnest and conscientious student. . . . Full of picturesque touches which 
 are most welcome " Natural Science. 
 
 " A CHARMING BOOK, beautifully illustrated." A thenssum. 
 
 %* For the Companion-Volume on " OPEN AIR BOTANY " see p. 33. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 
 
 SEWAGE DISPOSAL WORKS; 
 
 A Guide to the Construction of Works for the Prevention of the 
 Pollution by Sewage of Rivers and Estuaries. 
 
 BV 
 
 W. SANTO CRIMP, M.lNST.C.E., F.G.S., 
 
 Late Assistant-Engineer, London County Council 
 With Tables, Illustrations in the Text, and 37 Lithographic Plates. Medium 
 
 8vo. Handsome Cloth. 
 SECOND EDITION, REVISED AND ENLARGED. 
 
 30s. 
 
 PART I. INTRODUCTORY. 
 
 Introduction. 
 
 Details of River Pollutions and Recommenda- 
 tions of Various Commissions. 
 
 Hourly and Daily Flow of Sewage. 
 
 The Pail System as Affecting Sewage. 
 
 The Separation of Rain-water from the Sewage 
 Proper. 
 
 Settling Tanks. 
 Chemical Processes. 
 
 The Disposal of Sewage-sludge. 
 The Preparation of Land for Sewage Dis- 
 posal. 
 Table of Sewage Farn Management. 
 
 PART II. SEWAGE DISPOSAL WORKS IN OPERATION THEIR 
 CONSTRUCTION, MAINTENANCE, AND COST. 
 
 Illustrated by Plates showing the General Plan and Arrangement adopted 
 in each District. 
 
 Map of the LONDON Sewage System. 
 
 Crossness Outfall. 
 
 Barking Outfall. 
 
 Doncaster Irrigation Farm 
 
 Beddington Irrigation Farm, Borough of 
 Croydon. 
 
 Bedford Sewage Farm Irrigation. 
 
 Dewsbury and Hitchin Intermittent Fil- 
 tration. 
 
 Merton, Croydon Rural Sanitary Authority. 
 
 Swanwick, Derbyshire. 
 
 The Ealing Sewage Works. 
 
 Chiswick. 
 
 Kingston-on-Thames, A. B. C Process. 
 
 Salford Sewage Works. 
 
 Bradford, Precipitation. 
 
 New Maiden, Chemical Treatment and 
 
 Small Filters. 
 Friern Barnet. 
 
 Acton, Ferorone and Polarite Process. 
 Ilford, Chadwell, and Dagenham Work* 
 Corentry. 
 Wimbledon. 
 Birmingham. 
 Margate. 
 Portsmouth. 
 
 BERLIN Sewage Farms. 
 Sewage Precipitation Works, Dortmund 
 
 (Germany). 
 Treatment of Sewage by Electrolysis. 
 
 %* From the fact of the Author's having, for some years, had charge of the Main 
 Drainage Works of the Northern Section of the Metropolis, the chapter on LONDON will b 
 found t contain many important details which would not otherwise have been available. 
 
 " All persons interested in Sanitary Science owe a debt of gratitude to Mr. Crimp, . . . 
 His work will be especially useful to SANITARY AUTHORITIES and their advisers . . . 
 KMBMKNTLY PRACTICAL AND USEFUL . . . gives plans and descriptions of MANY or TH 
 MOST IMPORTANT SKWAGH WORKS of England . . . with very valuable information as to 
 the COST of construction and working of each. . . . The carefully-prepared drawings per- 
 mit of an easy comparison between the different systems." Lancet. 
 
 " Probably the MOST COMPLXTK AHD wtrr TRKATISK on the subject which has appeared 
 in our language . . Will prove of the greatest use to all who have the problem of 
 
 Sewage Disposal to &<*."- E &ni>urfh M*dic*l J *rn*l. 
 
 LONDON : EXETER STREET, STRAND. 
 
14 CHARLES GRIFFIN Je CO.'S PUBLICATIONS. 
 
 Griffin's "Pocket-Book" Series. 
 
 Pocket Size. Leather. With Illustrations. 12s. 6d. 
 
 HYGIENE (A Hand-Book of), 
 
 BY 
 
 SURGEON-MAJOR A. M. DAVIES, D.RH.Oamb., 
 
 Late Assistant-Professor of Hygiene, Army Medical School. 
 
 General Contents. 
 
 Air and Ventilation Water and Water Supply Food and Dieting 
 Removal and Disposal of Sewage Habitations Personal Hygiene Soils 
 and Sites Climate and Meteorology Causation and Prevention of Disease 
 Disinfection. 
 
 "This ADMIRABLE HANDBOOK . . . gives FULL infoimation compressed into the smallest 
 possible bulk." Edin. Med. Journal. 
 
 "The elegant dress of the little volume before us is but the outer covering of A TBULT RICH 
 KERNEL, and justly merits the praise it spontaneously calls forth. Attractive to the eye, Surgeon- 
 Major DAVIES' volume is equally attractive to the mind. Students will find that its 690 pages 
 comprise ALL information necessary. COMPACT, HANDY, COMPREHENSIVE, it certainly merits a 
 high place among the text-books of the day." Sanitary Record. 
 
 " We are glad to welcome Surgeon-Major Davies' book ... he has had ample opportunity 
 to make himself A MASTER OP THE SCIENCE, and he has a right to speak. . . . WONDERFULLY 
 WELL UP TO DATE, well and clearly written, pleasant to read.' The Lancet. 
 
 " Really an ADMIRABLE BOOK. . . . A MOST HANDY WORK OF REFERENCE full of information." 
 The Hospital. 
 
 " A singularly compact and elegant volume . . . contains an admirable precis of everything 
 relating to Ilygiene CLEARLY and LOGICALLY ARRANGED and easy of reference. Likely, we think, 
 to be the favourite text-book." Public Health. 
 
 Large 8vo. Handsome Cloth. 12s. 6d. 
 
 BLEACHING & CALICO-PRINTING. 
 
 A Short Manual for Students and 
 
 Practical Men. 
 BY GEORGE DUERR, 
 
 Director of the Bleaching, Dyeing, and Printing Department at the Accrington and Bacup 
 Technical Schools ; Chemist and Colourist at the Irwell Print Works. 
 
 ASSISTED BY WILLIAM TURNBULL 
 
 (of Turnbull & Stockdale, Limited). 
 
 With Illustrations and upwards of One Hundred Dyed and Printed Patterns 
 designed specially to show various Stages of the Processes described. 
 
 " When a READY WAY out of a difficulty is wanted, it is IN BOOKS LIKE THIS that it is found." 
 Textile Recorder. 
 
 "Mr. DUERR'S WORK will be found MOST USEFUL. . . . The information given of GHBAT 
 VALUK. . . . The Recipes THOROUGHLY PRACTICAL," Textile Manufacturer. 
 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 15 
 
 W O H K S 
 
 BY J. R. AINSWORTH DAVIS, B.A., 
 
 PROCESSOR OF BIOLOGY, UNIVERSITY COLLEGE, ABERYSTWYTH; 
 
 EXAMINER IN ZOOLOGY, UNIVERSITY OF ABERDEEN. 
 
 DAVIS (Prof. Ainsworth): BIOLOGY (An Ele- 
 
 mentary Text-Book of). In large Crown 8vo, Cloth. SECOND EDITION. 
 
 PART I. VEGETABLE MORPHOLOGY AND PHYSIOLOGY. With Complete Index- 
 Glossary and 128 Illustrations. Price 8s. 6d. 
 
 PART II. ANIMAL MORPHOLOGY AND PHYSIOLOGY. With Complete Index- 
 Glossary and 108 Illustrations. Price IDS. 6d. 
 
 EACH PART SOLD SEPARATELY. 
 
 %* NOTE The SECOND EDITION has been thoroughly Revised and Enlarged, 
 
 and includes all the leading selected TYPES in the various Organic Groups. 
 
 "Certainly THE BEST 'BIOLOGY* with which we are acquainted. It owes its pre- 
 eminence to the fact that it is an EXCELLENT attempt to present Biology to the Student as a 
 CORRELATED AND COMPLETE SCIENCE. The glossarial Index is a MOST USEFUL addition." 
 British Medical Journal. 
 
 " Furnishes a CLEAR and COMPKKHENSIVK exposition of the subject in a SYSTEMATIC 
 form." Saturday Review. 
 
 " Literally PACKED with information." Glasgow Medical Journal. 
 
 DAVIS (Prof. Ainsworth): THE FLOWERING 
 
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16 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 GAS, OIL, AND AIR 
 
 A Practical Text - Book on Internal Combustion Motors 
 without Boiler. 
 
 BY BRYAN DONKIN, M.lNST.C.E. 
 
 SECOND EDITION, Revised throughout and Enlarged. With numerous 
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 BY THE SAME AUTHOR. [Shortly. 
 
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 Many Experiments on Many Types, showing Results as to Evaporation, 
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 BY T. CLAXTON FIDLER, M.lNST.C.E., 
 
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 2 
 
i8 CHARLES GRIFFIN * CO:S PUBLICATIONS. 
 
 ORE & STONE MINING 
 
 BY 
 
 C. LE NEVE FOSTER, D.Sc., F.R.S., 
 
 PROFESSOR OK MINING. ROYAL COLLEGE 0F SCIENCE H.M. INSPECTOR OF MINES 
 
 SECOND EDITION. With Frontispiece and 716 Illustrations. 345. 
 
 " Dr. Foster's book was expected to be EPOCH-MAKING, and it fully justifies such expec- 
 tation. ... A MOST ADMIRABLE account of the mode of occurrence of practically ALL 
 KNOWN MINERALS. Probably stands UNRIVALLED for completeness." The Mining Journal, 
 
 GENERAL CONTENTS. 
 
 INTRODUCTION. Mode of Occurrence of Minerals: Classification: Tabular 
 Deposits, Masses Examples: Alum, Amber, Antimony, Arsenic, Asbestos. Asphalt, 
 Barytes. Borax, Boric Acid, Carbonic Acid, Clay, Cobalt Ore, Copper Ore, Diamonds, 
 Flint, Freestone, Gold Ore, Graphite, Gypsum, Ice, Iron Ore, Lead Ore, Manganese 
 Ore, Mica, Natural Gas, Nitrate of Soda, Ozokerite, Petroleum, Phosphate of Lime, 
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 Faults. Prospecting: Chance Discoveries Adventitious Finds Geology as a 
 Guide to Minerals Associated Minerals Surface Indications. Boring: Uses ol 
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 iii. By Percussion with Rope. Breaking Ground: Hand Tools Machinery- 
 Transmission of Power Excavating Machinery : i. Steam Diggers, ii. Dredges, 
 iii. Rock Drills, iv. Machines for Cutting Grooves, v. Machines for Tunnelling 
 Modes of using Holes Driving and Sinking Fire-setting Excavating by Water. 
 Supporting Excavations : Timbering Masonry Metallic Supports Watertight 
 Linings-Special Processes. Exploitation: Open Works: Hydraulic Mining- 
 Excavation of Minerals under Water Extraction of Minerals by Wells and Bore- 
 holesUnderground \V orkings Beds Veins Masses. Haulage or Transport : 
 Underground: by Shoots, Pipes, Persons, Sledges, Vehicles, Railways, Machinery. 
 Boats Conveyance above Ground. Hoisting or Winding: Motors, Drums, and 
 Pulley Frames Ropes, Chains, and Attachments Receptacles Other Appliances 
 Safety Appliances Testing Ropes Pneumatic Hoisting. Drainage: Surface Water 
 Dams Drainage Tunnels Siphons Winding Machinery Pumping Engines 
 above ground Pumping Engines below ground Co-operative Pumping. Ventila- 
 tion: Atmosphere of Mines Causes of Pollution of Air Natural Ventilation- 
 Artificial Ventilation : i. Furnace Ventilation, ii. Mechanical Ventilation Testing 
 the Quality of Air Measuring the Quantity and Pressure of the Air Efficiency of 
 Ventilating Appliances Resistance caused by Friction. Lighting : Reflected 
 Daylight Candles Torches Lamps Wells Light Safety Lamps Gas Electric 
 Light. Descent and Ascent : Steps and Slides Ladders Buckets and Cages Man 
 Engine. Dressing: i. Mechanical Processes: Washing, Hand Picking, Breaking 
 Up, Consolidation, Screening ii. Processes depending on Physical Properties: 
 Motion in Water, Motion in Air Desiccation Liquefaction and Distillation- 
 Magnetic Attraction iii. Chemical Processes: Solution, Evaporation and Crystal- 
 lisation, Atmospheric Weathering, Calcination, Cementation, Amalgamation Ap- 
 plication of Processes Loss in Dressing Sampling. Principles of Employment 
 of Mining Labour : Payment by Time, Measure, or Weight By Combination of 
 these By Value of Product. Legislation affecting Mines and Quarries: 
 Ownership Taxation Working Regulations Metalliferous Mines Regulation Acts 
 Coal Mines Regulation Act Other Statutes. Condition of the Miner : Clothing 
 Housing Education Sickness Thrift Recreation. Accidents : Death Rate of 
 Miners from Accidents Relative Accident Mortality Underground and Above- 
 ground Classification of Accidents Ambulance Training. 
 
 "This EPOCH-MAKING work . . . appeals to MEN OF EXPERIENCE no less than to 
 students . . . gives numerous examples from the MINING PRACTICE of EVERY COUNTRY. 
 Many of its chapters are upon subjects not usually dealt with in text books. . . Of 
 great interest. . . . Admirably illustrated." Berg- und Hiittenmannische Zeitung . 
 
 "This SPLENDID WORK." Oesterr. Ztschrft. fur Berg- und Hiittemvesen. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 19 
 
 SECOND EDITION, 8s. 6d. Leather, for the Pocket, 8s. 6d. 
 
 GRIFFIN'S ELECTRICAL PRICE-BOOK. 
 
 For Electrical, Civil, Marine, and Borough Engineers, Local 
 
 Authorities, Architects, Railway Contractors, &e., &e. 
 
 EDITED BY H. J. DOWSING, 
 
 Member of the Institution of Electrical Engineers ; of the Society of Arts; tfthe London 
 Chamber of Commerce, &*c. 
 
 " The ELECTRICAL PRICE-BOOK REMOVES ALL MYSTERY about the cost of Electrical 
 Power. By its aid the EXPENSE that will be entailed by utilising electricity on a large or 
 small scale can be discovered." Architect. 
 
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 jave time and trouble both to the engineer and the business man." Machinery. 
 
 GRIFFIN (John Joseph, F.C.S.) : 
 
 CHEMICAL RECREATIONS: A Popular Manual of Experimental 
 Chemistry. With 540 Engravings of Apparatus. Tenth Edition. Crown 
 8vo. Cloth. Complete in one volume, Parts I. and II. Cloth, gilt 
 top, 12/6. 
 
 Part II. The Chemistry of the Non- Metallic Elements, 10/6. 
 
 GURDEN (Richard Lloyd, Authorised Surveyor 
 
 for the Governments of New South Wales and Victoria) : 
 
 TRAVERSE TABLES : computed to Four Places Decimals for every 
 Minute of Angle up to 100 of Distance. For the use of Surveyors and 
 Engineers. FOU-RTH EDITION. Folio, strongly half-bound, 2i/. 
 %* Published with Concurrence of the Surveyors- General for New Sovtk 
 Wales and Victoria. 
 
 "Those who have experience in exact SURVEY-WORK will best know how to appreciat* 
 the enormous amount of labour represented by this valuable book. The computations 
 enable the user to ascertain the sines and cosines for a distance of twelve miles to within 
 half an inch, and this BY REFERENCE TO BUT ONE TABLE, in place of the usual Fifteen 
 minute computations required. This alone is evidence of the assistance which the Table* 
 ensure to every user, and as every Surveyor in active practice has felt the want of such 
 assistance, few knowing of their publication will remain without them." Enfttuer. 
 
 In Large &vo, with Illustrations and Folding-Plates. los. 6d, 
 
 BLASTING: 
 
 A Handbook for the Use of Engineers and others Engaged in 
 Mining, Tunnelling, Quarrying, &c. 
 
 BY OSCAR GUTTMANN, Assoc. M. INST. C.E. 
 
 Member of the Societies of Civil Engineers and A rchitects of Vienna and Budapest, 
 Corresponding Member of the Imp. Roy. Geological Institution of Austria, &C. 
 
 GENERAL CONTENTS. Historical Sketch Blasting Materials Blasting Pow- 
 der Various Powder-mixtures Gun-cotton Nitro-glycerine and Dynamite 
 Other Nitro-compounds Sprengel's Liquid (acid) Explosives Other Means of 
 Blasting Qualities, Dangers, and Handling of Explosives Choice of Blasting 
 Materials Apparatus for Measuring Force Blasting in Fiery Mines Means cf 
 Igniting Charges Preparation of Blasts Bore-holes Machine-drilling Chamber 
 Mines Charging of Bore-holes Determination of the Charge Blasting in Bore- 
 holes Firing Straw and Fuze Firing Electrical Firing Substitutes for Electrical 
 Firing Results of Working Various Blasting Operations Quarrying Blasting 
 Masonry, Iron and Wooden Structures Blasting in earth, under water, of ice, &c. 
 
 "This ADMIRABLE work." Colliery Guardian. 
 
 " Should prove a vade-mecum to Mining Engineers and all engaged in practical work. 
 Iron and Coal Trades Review. 
 
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20 CHARLES OB1FFIN & CO.'S PUBLICATIONS. 
 
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 Stability Of Ships, . SIR E. J. REED, . 38 
 
 The Steam-Engine, . . RANKINE, JAMIESON, . 35, 26 
 
 Valves and Valve-Gearing, CHAS. HURST, . . 23 
 
 Gas, Oil, and Air-Engines, BRYAN DONKIN, . 16 
 
 Boiler Construction, T. W. TRAILL, . . 49 
 
 Management, . R. D. MUNRO, . . 30 
 
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22 
 
 CHARLES GRIFFIN <k CO.'S PUBLICATIONS. 
 
 Griffin's Geological, Mining, and 
 Metallurgical Publications. 
 
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 (Physical), . 
 (Practical), . 
 ,, (Introduction to), 
 
 Mine Accounts, . 
 
 Mine-Surveying-, 
 
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 Assaying-, .... 
 
 Metallurgy, 
 
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 A. WYNTER BLYTH, . 
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 SANTO CRIMP, 
 
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 39 
 13 
 
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 Catalogue. 
 
 [See Medical 
 
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SCIENTIFIC AND TECHNOLOGICAL WORKS. 23 
 
 Grin's Chemical and Technological Publications. 
 
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 OF ENGINEERS, DRAUGHTSMEN, AND STUDENTS. 
 
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 41 USEFUL and THOROUGHLY PRACTICAL. Will undoubtedly be found of GREAT VALUE to 
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 . . . Should prove both USEFUL and VALUABLE to all Engineers s eking for RELIABLE and 
 CLEAR information on the subject. Its moderate price brings ir within the reach ot all" 
 Industries an > Iron 
 
 " Mr. HURST'S work is ADMIRABLY suited to the needs of the practical mechanic. . . . 
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 The Scientific American. 
 
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24 OHARLSS GRIFFIN A CO. 1 8 PUBLICATIONS. 
 
 COAL-MINING (A Text-Book of): 
 
 FOR THE USE OF COLLIERY MANAGERS AND OTHERS 
 ENGAGED IN COAL-MINING. 
 
 BY 
 
 HERBERT WILLIAM HUGHES, F.G.S., 
 
 Assoc. Royal School of Mines, Certificated Colliery Manager. 
 THIRD EDITION. In Demy &vo, Handsome Cloth. With -very Numerous 
 
 Illustrations, mostly reduced from Working Drawings. i8j. 
 "The details of colliery work have been fully described, on the ground that 
 collieries are more often made REMUNERATIVE by PERFECTION IN SMALL MATTERS 
 than by bold strokes of engineering. ... It frequently happens, in particular 
 localities, that the adoption of a combination of small improvements, any of 
 which viewed separately may be of apparently little value, turns an unprofitable 
 concern into a paying one." Extract from Author s Preface. 
 
 GENERAL CONTENTS. 
 
 Geology : Rocks -Faults Order of Succession Carboniferous System in Britain. 
 Goal : Definition and Formation of Coal Classification and Commercial Value of Coals, 
 Search for Coal : Boring various appliances used Devices employed to meet Difficulties 
 of deep Boring Special methods of Boring Mather & Plait's, American, and Diamond 
 systems Accidents in Boring Cost of Boring Use of Boreholes. Breaking Ground; 
 Tools Transmission of Power: Compressed Air, Electricity Power Machine Drills Coal 
 Cutting by Machinery Cost of Coal Cutting Explosives Blasting in Dry and Dusty 
 Mines Blasting by Electricity Various methods to supersede Blasting. Sinking: 
 Position, Form, and Size of shaft Operation of getting down to " Stono-head" Method of 
 proceeding afterwardv^-Lining shafts Keeping out Water by Tubbing Cost of Tubbing 
 Sinking by Boring Kind - Chaudron, and Lipmann methods Sinking through Quicksands 
 Cost of Sinking. Preliminary Operations : Driving underground Roads Supporting 
 Roof: Timbering, Chocks or Cogs, Iron and Steel Supports and Masonry Arrangement of 
 Inset. Methods of Working : Shaft, Pillar, and Subsidence Bord and Pillar System- 
 Lancashire Method Longwall Method Double Stall Method Working Steep Seams 
 Working Thick Seams Working Seams lying near together Spontaneous Combustion. 
 Haulage: Rails Tubs Haulage by Horses Self-acting Inclines Direct-acting Haulage 
 Main and Tail Rope Endless Chain Endless Rope Comparison. Winding : Pit 
 Frames Pulleys Cages Ropes Guides Engines Drums Brakes Counterbalancing 
 Expansion Condensation Compound Enjjines Prevention of Overwinding Catches at pit 
 top Changing Tubs Tub Controllers Signalling. Pumping: Bucket and Plunger 
 Pumps Supporting Pipes in Shaft Valves Suspended lifts for Sinking Cornish and 
 Bull Engines Davey Differential Engine Worthington Pump Calculations as to size of 
 Pumps Draining Deep Workings Dams. Ventilation: Quantity of air required 
 Gases met with in Mines Coal-dust Laws of Friction Production of Air-currents 
 Natural Ventilation Furnace Ventilation Mechanical Ventilators Efficiency of Fans 
 Comparison of Furnaces and Fans Distribution of the Air-current Measurement of Air- 
 currents. Lighting: Naked Lights - Safety Lamps Modern Lamps Conclusions 
 Locking and Cleaning Lamps Electric Light Underground Delicate Indicators. Works 
 at Surface; Boilers Mechanical Stoking Coal Conveyors Workshops. Preparation 
 of Coal for Market : General Considerations Tipplers Screens Varying the Sizes made 
 by Screens Belts Revolving Tables Loading Shoots Typical Illustrations of the arrange- 
 ment of Various Screening Establishments Coal Washing Dry Coal Cleaning -Briquettes. 
 
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 touches upon every point connected with the actual working of collieries. The illustrations 
 are KXCKLLINT." Athen&um. 
 
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 Colliery Guardian. 
 
 " Mr. HUGHES has had opportunities for study and research which fall to the lot of 
 but few men. If we mistake not, his text-book will soon come to be regarded as the 
 STANDARD WORK of its kind." Birmingham Daily Gazette. 
 
 *+*Note. The first large edition of this work was exhausted within a few months of 
 publication. 
 
 LONDON: EXETER STEERT, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 25 
 
 WORKS BY GEORGE H. HURST, F.C.S., 
 
 Member of the Society of Chemical Industry ; Lecturer on the Technology of Painteri* 
 Colours, Oils, and Varnishes, the Municipal Technical School, Manchester. 
 
 NEW WORK BY MR. HURST. 
 
 COLOUR-THEORY AND ITS PRACTICAL 
 
 APPLICATION TO PAINTING, DYEING, and 
 THE TEXTILE INDUSTRIES. 
 
 In Large Crown 8vo. With Illustrations and Plates (some in Colours). 
 
 HURST'S PAINTERS' COLOURS, OILS, AND 
 
 VARNISHES (A Practical Manual). SECOND EDITION, Revised and 
 Enlarged. With Illustrations. I2s. 6d. 
 
 GENERAL CONTENTS. Introductory THE COMPOSITION, MANUFACTURE, 
 ASSAY, and ANALYSIS of PIGMENTS, White, Red, Yellow and Orange, Green, 
 Blue, Brown, and Black LAKES Colour and Paint Machinery Paint Vehicles 
 (Oils, Turpentine, &c. , &c. ) Driers VARNISHES. 
 
 " This useful book will prove MOST VALUABLE." Chemical News. 
 
 " A practical manual in every respect . . . EXCEEDINGLY INSTRUCTIVE. The 
 section on Varnishes the most reasonable we have met with." Chemist and Druggist. 
 
 " VERY VALUABLE information is given." Plumber and Decorator. 
 
 " A THOROUGHLY PRACTICAL book, . . . the ONLY English work that satisfactorily 
 treats of the manufacture of oils, colours, and pigments." Chemical Trades' Journal. 
 
 " Throughout the work are scattered hints which are INVALUABLE." Invention. 
 
 HURST'S GARMENT DYEING and CLEANING 
 
 (A Practical Book for Practical Men). With Numerous Illustrations. 
 45. 6d. 
 
 GENERAL CONTENTS. Technology of the Textile Fibres Garment Cleaning 
 Dyeing of Textile Fabrics Bleaching Finishing of Dyed and Cleaned Fabrics 
 Scouring and Dyeing of Skin Rugs and Mats Cleaning and Dyeing of Feathers 
 Glove Cleaning and Dyeing Straw Bleaching and Dyeing Glossary of Drugs 
 and Chemicals Useful Tables. 
 
 " An UP-TO-DATE hand book has long been wanted, and Mr. Hurst has done nothing 
 more complete than this. An important work, the more so that several of the branches of 
 the craft here treated upon are almost entirely without English Manuals for the guidance 
 of worker?. The price brings it within the reach of all." Dyer and Calico-Printer. 
 
 " Mr. Hurst's woric DECIDEDLY FILLS A WANT . . . ought to be in the hands of 
 EVERY GARMENT DYER and cleaner in the Kingdom" Textile Mercury. 
 
 LONDON : EXETER STREET, STRAND. 
 
26 CHARLES GRIFFIN A CO.'S PUBLICATIONS. 
 
 WORKS BY 
 
 ANDREW JAMIESON, M.lNST.C.E., M.I.E.E., F.R.S.E.,. 
 
 Professor of Electrical Engineering, The Glasgow and West of Scotland 
 Technical College. 
 
 PROFESSOR JAMIESON'S ADVANCED MANUALS, 
 
 In Large Crown %vo. Fully Illustrated. 
 
 1. STEAM AND STEAM-ENGINES (A Text-Book on). 
 
 For the Use of Students preparing for Competitive Examinations. 
 With over 200 Illustrations, Folding Plates, and Examination Papers. 
 TWELFTH EDITION. Revised and Enlarged, 8/6. 
 
 " Professor Jamieson fascinates the reader by his CLEARNESS OF CONCEPTION ANTJ 
 SIMPLICITY OF EXPRESSION. His treatment recalls the lecturing of Faraday." Athenaeum. 
 " The BEST BOOK yet published for the use of Students." Engineer. 
 "Undoubtedly the MOST VALUABLE AND MOST COMPLETE Hand-book on the subject' 
 that now exists." Marine Engineer. 
 
 2. MAGNETISM AND ELECTRICITY (An Advanced Text- 
 
 Book on). Specially arranged for Advanced and " Honours " Students. 
 
 3. APPLIED MECHANICS (An Advanced Text-Book on). 
 
 Vol. I. Comprising Part I. : The Principle of Work and its applica- 
 tions; Part II.: Gearing. Price 7s. 6d. SECOND EDITION. 
 
 " FULLY MAINTAINS the reputation of the Author more we cannot say." Pract. 
 Engineer. 
 
 Vol. II. Comprising Parts III. to VI. : Motion and Energy; Graphic 
 Statics; Strength of Materials; Hydraulics and Hydraulic Machinery. 
 Price 7s. 6d. [Now ready. 
 
 PROFESSOR JAMIESON'S INTRODUCTORY MANUALS. 
 
 With numerous Illlustrations and Examination Papers. 
 
 1. STEAM AND THE STEAM-ENGINE (Elementary Text- 
 Book on). For First-Year Students. FIFTH EDITION. 3/6. 
 
 " Quite the RIGHT SORT OF BOOK." Engineer. 
 
 "Should be in the hands of EVERY engineering apprentice." Practical Engineer. 
 
 2. MAGNETISM AND ELECTRICITY (Elementary Text- 
 
 Book on). For First- Year Students. FOURTH EDITION. 3/6. 
 
 " A CAPITAL TEXT-BOOK . . . The diagrams are an important feature." Schoolmaster. 
 
 "A THOROUGHLY TRUSTWORTHY Text-book. . . . Arrangement as good as well 
 can be. ... Diagrams are also excellent. . . . The subject throughout treated as an 
 essentially PRACTICAL one, and very clear instructions given." Nature. 
 
 3. APPLIED MECHANICS (Elementary Text-Book on). 
 
 Specially arranged for First-Year Students. SECOND EDITION. 3/6. 
 
 "Nothing is taken for granted. . . . The work has VERY HIGH QUALITIES, which. 
 m*y be condensed into the one word 'CLEAR.'" Science and Art. 
 
 POCKET-BOOK of ELECTRICAL RULES and TABLES, 
 
 FOR THE USE OF ELECTRICIANS AND ENGINEERS. 
 Pocket Size. Leather, 8s. 6d. Twelfth Editinn : revised and enlarged. 
 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 2p 
 
 44 The MOST VALUABLE and USEFUL WORK on Dyeing that has yet appeared in the English 
 language . . . likely to be THE STANDARD WORK OF REFERENCE for years to come." 
 Textile Mercury. 
 
 In Two Large 8vo Volumes, 920 
 pp., with a SUPPLEMENTARY 
 Volume, containing Specimens 
 of Dyed Fabrics. Handsome 
 Cloth, 45s. 
 
 MANUAL OF DYEING: 
 
 FOR THE USE OF PRACTICAL DYERS, MANUFACTURERS, STUDENTS, 
 AND ALL INTERESTED IN THE ART OF DYEING. 
 
 BY 
 
 E. KNECHT, Ph.D., F.I.C., CHR. RAWSON, F.I.C., F.C.S., 
 
 Head of the Chemistry and Dyeing Department of Late Head of the Chemistry and Dyeing Department 
 the Technical School, Manchester; Editor of "The of the Technical College, Bradford; Member of 
 Journal of the Society of Dyers and Colourists ; " Council of the Society of Dyers and Colourists ; 
 
 And RICHARD LOEWENTHAL, Ph.D. 
 
 GENERAL CONTENTS. Chemical Technology of the Textile Fabrics- 
 Water Washing and Bleaching Acids, Alkalies, Mordants Natural 
 Colouring Matters Artificial Organic Colouring Matters Mineral Colours. 
 Machinery used in Dyeing Tinctorial Properties of Colouring Matters 
 Analysis and Valuation of Materials used in Dyeing, &c., &c. 
 
 " This MOST VALUABLE WORK . . . will be widely appreciated." Chemical News. 
 
 44 This authoritative and exhaustive work . . . the MOST COMPLETE we have yet seen 
 on the subject." Textile Manufacturer. 
 
 " The MOST EXHAUSTIVE and COMPLETE WORK on the subject extant." Textile Recorder. 
 
 14 The distinguished authors have placed in the hands of those daily engaged in the dye- 
 house or laboratory a work of EXTREME VALUE and UNDOUBTED UTILITY . . . appeals 
 quickly to the technologist, colour chemist, dyer, and more particularly to the rising dyer 
 of the present generation. A book which it is refreshing to meet with." American Textile 
 Record. 
 
 LONDON: EXETER STREET, STRAND. 
 
28 CHARLES GRIFFIN <k CO.'S PUBLICATIONS. 
 
 GETTING GOLD: 
 
 A GOLD-MINING HANDBOOK FOR PRACTICAL MEN, 
 
 BY 
 
 J. 0. F. JOHNSON, F.G.S., A.I.M.E., 
 
 Life Member Australasian Mine-Managers' Association. 
 SECOND EDITIOV. Crown 8vo, Extra. With Illustrations. Cloth, '3s. Qd. 
 
 " Should prove of the GREATEST VALUE. Almost every page bristles with sugges- 
 tions." Financial News. 
 
 "PRACTICAL from beginning to end . . . deals thoroughly with the Prospecting, 
 Sinking, Crushing, and Extraction of gold." Brit. Australasian. 
 
 "Directors and those interested in the formation of companies would do well to pur- 
 chase Mr. Johnson's book." Mining Journal. 
 
 " The reader, be he miner or novice, will gain from Mr. Johnson's book a GRIP of the 
 gold industry." African Critic. 
 
 "Should be specially commended to all who have any idea of proceeding to the gold 
 fields." Financial Truth. 
 
 " The most striking elements are the numerous ' TIPS ' and USEFUL WRINKLES given." 
 Standard. 
 
 " One is lost in admiration at the wealth of knowledge displayed." Nature. 
 
 " Evidently the well-matured product of a scientist of well-trained and tried experience." 
 South Africa. 
 
 NEW VOLUME OF GRIFFIN'S MINING SERIES. 
 
 Edited by C. LE NEVE FOSTER, D.Sc., F.R.S., 
 
 H.M. Inspector of Mines, Professor of Mining, Royal School of Mines. 
 
 Mine Accounts and Mining Book-keeping, 
 
 A Manual for the Use of Students, Managers of 
 
 Metalliferous Mines and Collieries, and 
 
 others interested in Mining. 
 
 With very Numerous Examples taken from the Actual Practice 
 of leading Mining Companies throughout the world. 
 
 BY 
 JAMES G. LAWN, Assoc.R.S.M., 
 
 Professor of Mining at the South African School of Mines, Capetown, 
 Kimberley, and Johannesburg. 
 
 In Large 8vo. Price 10s. Qd. 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 29 
 
 WALTEE G. M'MILLAN, F.I.C., F.C.S., 
 
 Secretary to the Institution of Electrical Engineers; late Lecturer in Metallurgy 
 at Mason College, Birmingham. 
 
 ELECTRIC SMELTING AND REFINING: 
 
 A PRACTICAL MANUAL OF 
 
 THE EXTRACTION AND TREATMENT OF METALS BY ELECTRICAL METHODS. 
 Being the " ELEKTRO-METALLURGIE " of Dr. W. BORCHERS. 
 
 Translated from the, Second German Edition 
 BY WALTER G. M'MILLAN, F.I.O., F.C.S. 
 
 In large 8vo. With Numerous Illustrations and Three Folding-Plates. 
 
 Price 21s. 
 
 *** THE PUBLISHERS beg to call attention to this valuable work. Dr. BORCHERS' 
 treatise is PRACTICAL throughout. It confines itself to ONE branch of Electro-Chemistry, 
 viz. : ELECTROLYSIS, a subject which is daily becoming of more and more importance to 
 the Practical Metallurgist and Manufacturer. Already in the extraction of Aluminium, 
 the refining of Copper, the treatment of Gold and other metals, electrical processes are 
 fast taking the place of the older methods. Dr. BORCHERS' work is acknowledged as the 
 standard authority on the subject in Germany. 
 
 CONTENTS. 
 
 PART I. ALKALIES AND ALKALINE EARTH METALS : Magnesium, 
 Lithium, Beryllium, Sodium, Potassium, Calcium, Strontium, Barium, 
 the Carbides of the Alkaline Earth Metals. PART II. THE EARTH 
 METALS: Aluminium, Cerium, Lanthanum, Didymium. PART III. THE 
 HEAVY METALS : Copper, Silver, Gold, Zinc and Cadmium, Mercury, Tin, 
 Lead, Bismuth, Antimony, Chromium, Molybdenum, Tungsten, Uranium, 
 Manganese, Iron, Nickel, and Cobalt, the Platinum Group. 
 
 " COMPREHENSIVE and AUTHORITATIVE . . . not only PULL of VALUABLE INFOR- 
 MATION, but gives evidence of a THOROUGH INSIGHT into the technical VALUE and 
 POSSIBILITIES of all the methods discussed." The Electrician. 
 
 ELECTRO-METALLURGY (A Treatise on): 
 
 Embracing the Application of Electrolysis to the Plating, Depositing, 
 Smelting, and Refining of various Metals, and to the Repro- 
 duction of Printing Surfaces and Art- Work, &c. 
 
 BY WALTER G. M'MILLAN, F.I.C., P.C.S. 
 
 With numerous Illustrations. Large Crown 8vo. Cloth 10s. 6d. 
 
 "This excellent treatise, . . . one of the BUST and MOST COMPLETE 
 manuals hitherto published on Electro-Metallurgy." Electrical Review. 
 
 " This work will be a STANDARD. "Jeweller. 
 
 "Any metallurgical process which REDUCES the COST of production 
 must of necessity prove of great commercial importance. . . . We 
 recommend this manual to ALL who are interested in the PRACTICAL 
 APPLICATION of electrolytic processes." Nature. 
 
 LONDON : EXETER STREET, STRAND. 
 
30 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 MACKENZIE (Thos., Master Mariner, F.R.A.S.): 
 
 PRACTICAL MECHANICS : APPLIED to the REQUIREMENTS of 
 
 the SAILOR. Crown 8vo, with numerous Illustrations. Handsome 
 
 Cloth. 35. 6d. {Griffin's Nautical Series. 
 
 GENERAL CONTENTS. Resolution and Composition of Forces Work done 
 
 by Machines and Living Agents The Mechanical Powers : The Lever ; 
 
 Derricks as Bent Levers The Wheel and Axle : Windlass ; Ship's Capstan ; 
 
 Crab Winch Tackles : the "Old Man" The Inclined Plane; the Screw 
 
 The Centre of Gravity of a Ship and Cargo Relative Strength of Rope : 
 
 Steel Wire, Manilla, Hemp, Coir Derricks and Shears Calculation of the 
 
 "Cross -breaking Strain of Fir Spar Centre of Effort of Sails Hydrostatics : 
 
 the Diving-bell ; Stability of Floating Bodies ; the Ship's Pump, &c. 
 
 " THIS EXCELLENT BOOK . . . contains a LARGE AMOUNT of information." 
 Nature. 
 
 " WELL WORTH the money . . . will be found EXCEEDINGLY HELPFUL." 
 Shipping World. 
 
 " No SHIPS' OFFICERS' BOOKCASE will henceforth be complete without 
 CAPTAIN MACKENZIE'S ' PRACTICAL MECHANICS.' Notwithstanding my many 
 years' experience at sea, it has told me how much more there is to acquire." 
 (Letter to the Publishers from a Master Mariner). 
 
 MILLAR (W. J., C.E., late Secretary to the Inst. 
 
 of Engineers and Shipbuilders in Scotland) : 
 
 LATITUDE AND LONGITUDE : How TO FIND THEM. Crown 
 8vo, with Diagrams. 2s. [Griffin's Nautical Series. 
 
 " CONCISELY and CLEARLY WRITTEN . . . cannot but prove an acquisition 
 to those studying Navigation." Marine Engineer. 
 
 " Young Seamen will find it HANDY and USEFUL, SIMPLE and CLEAR." The 
 Engineer. 
 
 SECOND EDITION. Enlarged, and very fully Illustrated. Cloth, 4. 6d, 
 
 STEAM - BOILERS; 
 
 THEIR DEFECTS, MANAGEMENT, AND CONSTRUCTION. 
 
 BY R D. MUNRO, 
 
 Chief Engineer of the Scottish Boiler Insurance and Engine Inspection Company. 
 
 This work contains information of the first importance to every user of 
 Steam-power. It is a PRACTICAL work written for PRACTICAL men, the 
 language and rules being throughout of the simplest nature. 
 
 " A valuable companion for workmen and engineers engaged about Steam 
 Boilers, ought to be carefully studied, and ALWAYS AT HAND." Coll. Guardian. 
 
 " The book is VERY USEFUL, especially to steam users, artisans, and 
 young engineers." Engineer. 
 
 BY THE SAME AUTHOR. 
 
 KITCHEN BOILER EXPLOSIONS: Why 
 
 they Occur, and How to Prevent their Occurrence? A Practical Hand- 
 book based on Actual Experiment. With Diagrams and Coloured Plate, 
 Price 35. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 31 
 
 MUNRO ft JAMIESON'S ELECTRICAL POCKET-BOOK. 
 
 TWELFTH EDITION, Revised and Enlarged. 
 
 A POCKET-BOOK 
 
 ELECTRICAL RULES & TABLES 
 
 FOR THE USE OF ELECTRICIANS AND ENGINEERS. 
 
 BY 
 
 JOHN MUNRO, C.E., & PROF. JAMIESON, M.lNST.C.E., F.R.S.E, 
 With Numerous Diagrams. Pocket Size. Leather, 8s. 6d. 
 
 GENERAL CONTENTS. 
 
 UNITS OF MEASUREMENT. 
 
 MEASURES. 
 
 TESTING. 
 
 CONDUCTORS. 
 
 DIELECTRICS. 
 
 SUBMARINE CABLES. 
 
 TELEGRAPHY. 
 
 ELECTRO-CHEMISTRY. 
 
 ELECTRO-METALLURGY. 
 
 BATTERIES. 
 
 DYNAMOS AND MOTORS. 
 
 TRANSFORMERS. 
 
 ELECTRIC LIGHTING. 
 
 MISCELLANEOUS. 
 
 LOGARITHMS. 
 
 APPENDICES. 
 
 " WONDERFULLY PERFECT. . . . Worthy of the highest commendation we can 
 give it." Electrician, 
 
 The STERLING VALUE of Messrs. MUNRO and JAMIBSON'S POCKET-BOOK." 
 
 MUNRO (J. M. H., D.Sc., Professor of Chemistry, 
 
 Downton College of Agriculture): 
 
 AGRICULTURAL CHEMISTRY AND ANALYSIS : A PRAC- 
 TICAL HAND-BOOK for the Use of Agricultural Students. 
 
 NYSTROM'S POCKET-BOOK OF MECHANICS 
 
 AND ENGINEERING. Revised and Corrected by W. DENNIS MARKS, 
 Ph.B., C.E. (YALE s.s.s.), Whitney Professor of Dynamical Engineering, 
 University of Pennsylvania. Pocket Size. Leather, 155. TWENTIETH 
 EDITION, Revised and greatly enlarged. 
 
 LONDON : EXETER STREET, STRAND. 
 
32 OHARLES GRIFFIN A CO.'S PUBLICATIONS. 
 
 Demy 8vo, Handsome cloth, 18s. 
 
 Physical Geology and Palaeontology, 
 
 OJV TEE BASIS OF PHILLIPS. 
 
 BY 
 
 HARRY GOVIER SKELEY, R R. S., 
 
 PROFESSOR OF GEOGRAPHY IN KING'S COLLBGB, LONDON. 
 
 TKUtb ^frontispiece in Cbromo=lUtbo0rapb, and 3Uu6tratione, 
 
 " It is impossible to praise too highly the research which PROFESSOR SEELEY'S 
 ' PHYSICAL GEOLOGY ' evidences. IT is FAR MORE THAN A TEXT-BOOK it is 
 a DIRECTORY to the Student in prosecuting his researches." Presidential Ad- 
 dress 10 the Geological Society ', 1885, by Rev. Prof. Bonney D.Sc. , LL.D., F.R.S. 
 
 " PROFESSOR SEELEY maintains in his ' PHYSICAL GEOLOGY ' the high 
 reputation he already deservedly bears as a Teacher. " Dr. Henry Wood- 
 ward, F.R.S., in the " Geological Magaxine." 
 
 " PROFESSOR SEELEY'S work includes one of the most satisfactory Treatises 
 on Lithology in the English language. ... So much that is not accessible 
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 aftord to be without it." American Journal of Engineering. 
 
 Demy 8vo, Handsome cloth, 34s. 
 
 StratlgrapMcal Geology & Palaeontology, 
 
 ON THE BASIS OF PHILLIPS. 
 
 BY 
 
 ROBERT ETHERIDGE, F. R. S., 
 
 TM NATURAL HIST. DEPARTMENT, BRITISH MUSEUM, LATE PALAEONTOLOGIST TO 1 
 
 GEOLOGICAL SURVEY OF GREAT BRITAIN, PAST PRESIDENT OF THE 
 
 GEOLOGICAL SOCIETY, ETC. 
 
 TKHftb dfoap, numerous Gables, ano ttbirtB*sir. plates. 
 
 * % * PROSPECTUS of the above important wrk perhaps the MOST ELABORATE of 
 its kind ever written, and one calculated to grve a new strength to the study 
 0f Geology in Bi itain may be had on application to the Publishers. 
 
 " N such compendium of geological knowledge has ever been brought together before " - 
 Werstintnster Review. 
 
 " If PROF. SKELEY'S volume was remarkable for its originality and the breadth of its view*, 
 Mr. ETHBRIDGB fully justifies the assertion made in his preface that his book differs in con- 
 struction and detail from any known manual. . . . Must taJca HIGH RANK AMONG WOKKS 
 OF **raxwcM"At/un4tum. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 35 
 
 Painting and Decorating: 
 
 A Complete Practical Manual for House 
 Painters and Decorators. 
 
 Embracing the Use of Materials, Tools, and Appliances; the 
 
 Practical Processes involved ; and the General Principles 
 
 of Decoration, Colour, and Ornament. 
 
 BY 
 
 WALTER JOHN PEARCE, 
 
 LECTURER AT THE MANCHESTER TECHNICAL SCHOOL FOR HOUSE-PAINTING AND DECORATING. 
 
 In Crown 8vo. extra. With Numerous Illustrations and Plates 
 (some in Colours), including Original Designs. 12s. 6d. 
 
 %* MR. PEARCE'S work is the outcome of many years' practical ex- 
 perience, and will be found invaluable by all interested in the subjects 
 of which it treats. It forms the Companion- Volume to MR. GEO. HURST'S 
 well-known work on " PAINTERS' COLOURS " (see p. 25). 
 
 Open-Air Studies in Botany : 
 
 SKETCHES OF BRITISH WILD FLOWERS IN 
 THEIR HOMES. 
 
 BY 
 
 R. LLOYD PRAEGER, B.A., M.R.I.A. 
 
 Illustrated by Drawings from Nature by S. Rosamond Praeger, 
 and Photographs by R. Welch. 
 
 In Crown 8vo. extra. Handsome Cloth, 7s. 6d. 
 Gilt, for Presentation, 8s. 6d. 
 
 GENERAL CONTENTS. A Daisy-Starred Pasture Under the Hawthorns 
 By the River Along the Shingle A Fragrant Hedgerow A Connemara 
 Bog Where the Samphire grows A Flowery Meadow Among the Corn 
 (a Study in Weeds) In the Home of the Alpines A City Rubbish-Heap 
 Glossary. 
 
 " A FRESH AND STIMULATING book . . . should take a high place . . . The 
 Illustrations are drawn with much skill." The Times. 
 
 "BEAUTIFULLY ILLUSTRATED. . . . One of the MOST ACCURATE as well as 
 INTERESTING books of the kind we have seen." Athenceum. 
 
 "Redolent with the scent of woodland and meadow."! 7 ^ Standard. 
 
 "A Series of STIMULATING and DELIGHTFUL Chapters on Field-Botany." The 
 Scotsman. 
 
 *** Companion-Volume to Prof. Grenville Cole's fascinating " Open-Air Studies in 
 Geology " (see p. 12). 
 
 LONDON: EXETER STREET, STRAND. 
 
34 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 THIRD EDITION. With Folding Plates and Many Illustrations. 
 Large 8vo. Handsome Cloth. 36s. 
 
 ELEMENTS OP METALLURGY: 
 
 A PRACTICAL TREATISE ON THE ART OF EXTRACTING METALS 
 
 FROM THEIR ORES, 
 
 BY J. ARTHUR PHILLIPS, M.lNST.C.E., F.C.S., F.G.S., <fec. 
 AND H. BAUERMAN, V.P.G.S. 
 
 GENERAL CONTENTS. 
 
 Refractory Materials. 
 
 Fire-Clays. 
 
 Fuels, &c. 
 
 Aluminium. 
 
 Copper. 
 
 Tin. 
 
 Antimony. 
 
 Arsenic. 
 
 Zinc. 
 
 Mercury. 
 
 Bismuth. 
 
 Lead. 
 
 Iron. 
 
 Cobalt. 
 
 Nickel 
 
 Silver. 
 
 Gold. 
 
 Platinum. 
 
 ** Many NOTABLE ADDITIONS, dealing with new Processes and Developments, 
 will be found in the Third Edition. 
 
 " Of the THIRD EDITION, we are still able to say that, as a Text-book of 
 Metallurgy, it is THE BEST with which we are acquainted/' Engineer. 
 
 "The value of this work is almost inestimable. There can be no question 
 that the amount of time and labour bestowed on it is enormous. . . . There 
 is certainly no Metallurgical Treatise in the language calculated to prove of 
 guch general utility." Mining Journal. 
 
 " In this most useful and handsome volume is condensed a large amount of 
 valuable practical knowledge. A careful study of the first division of the book, 
 on Fuels, will be found to be of great value to every one in training for the 
 practical applications of our scientific knowledge to any of our metallurgical 
 operations. " A thenceum. 
 
 " A work which is equally valuable to the Student as a Text-book, and to the 
 practical Smelter as a Standard Work of Reference. . . . The Illustration* 
 we admirable examples of Wood Engraving." Chemical News. 
 
 POYNTING (J. H., Sc.D., F.R.S., late Fellow 
 
 of Trinity College, Cambridge; Professor of Physics, Mason College, 
 Birmingham) : 
 
 THE MEAN DENSITY OF THE EARTH: An Essay to 
 which the Adams Prize was adjudged in 1893 in the University of 
 Cambridge. In large 8vo, with Bibliography, Illustrations in the Text, 
 and seven Lithographed Plates. 125. 6d. 
 
 "An account of this subject cannot fail to be of GREAT and GENERAL IHTBRBST to the scientific 
 mind. Especially is this the case when the account is given by one who has contributed BO 
 considerably as has Prof. Poynting to our present state of knowledge with respect to a very 
 difficult subject. . . . RemarKably has Newton's estimate been verified by Prof. Poynting." 
 Athenaeum. 
 
 POYNTING and THOMSON: TEXT-BOOK 
 
 OF PHYSICS. (See under Thomson). 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 35 
 
 WORKS BY 
 
 J, MACQUORN RANKINE, LL.D, F.R.S,, 
 
 Late Regius Professor of Civil Engineering in the University of Glasgow. 
 THOROUGHLY REVISED BY 
 
 W. J. MIL LAB, C.E., 
 
 Late Secretary to the Institute of Engineers and Shipbuilders in Scotland. 
 
 I. A MANUAL OF APPLIED MECHANICS : 
 
 Comprising the Principles of Statics and Cinematics, and Theory of 
 Structures, Mechanism, and Machines. With Numerous Diagrams. 
 Crown Svo, cloth, 12s. 6d. FOURTEENTH EDITION. 
 
 II. A MANUAL OF CIVIL ENGINEERING : 
 
 Comprising Engineering Surveys, Earthwork, Foundations, Masonry, Car- 
 pentry, Metal Work, Roads, Railways, Canals, Rivers, Waterworks, 
 Harbours, &c. With Numerous Tables and Illustrations. Crown Svo, 
 cloth, 16s. TWENTIETH EDITION. 
 
 III. A MANUAL OF MACHINERY AND MILLWORK : 
 
 Comprising the Geometry, Motions, Work, Strength, Construction, and 
 Objects of Machines, &c. Illustrated with nearly 300 Woodcuts. 
 Crown Svo, cloth, 12s. 6d. SEVENTH EDITION. 
 
 IV. A MANUAL OF THE STEAM-ENGINE AND OTHER 
 PRIME MOVERS : 
 
 With a Section on GAS, OIL, and AIR ENGINES. By BRYAN DON KIN, 
 M.Iust.C.E. Crown Svo, cloth, 12s. 6d. FOURTEENTH EDITION. 
 
 V. USEFUL RULES AND TABLES : 
 
 For Architects, Builders, Engineers, Founders, Mechanics, Shipbuilders, 
 Surveyors, &c. With APPENDIX for the use of ELECTRICAL ENGINEERS. 
 By Professor JAMIESON, F.R.S.E. SEVENTH EDITION. 10s. 6d. 
 
 VI. A MECHANICAL TEXT-BOOK : 
 
 A Practical and Simple Introduction to the Study of Mechanics. By 
 Professor RANKINE and E. F. BAMBER, C.E. With Numerous Illus- 
 trations. Crown Svo, cloth, 9s. FOURTH EDITION. 
 
 *.* The " MECHANICAL TEXT-BOOK " was detigned by Professor RANKINS at an [nu>- 
 to the above Series of Manuals. 
 
 LONDON: EXETER STREET, STRAND. 
 
36 CHARLES GRIFFIN <k CO.'S PUBLICATIONS. 
 
 PROF. RANKINK'S WORKS (Continued). 
 
 VII. MISCELLANEOUS SCIENTIFIC PAPERS. 
 
 Royal 8vo. Cloth, 31s. 6d. 
 
 Part I. Papers relating to Temperature, Elasticity, and Expansion of 
 Vapours, Liquids, and Solids. Part II. Papers on Energy and its Trans- 
 formations. Part III. Papers on Wave-Forms, Propulsion of Vessels, &c. 
 
 With Memoir by Professor TAIT, M.A. Edited by W. J. MILLAR, C.E. 
 With fine Portrait on Steel, Plates, and Diagrams. 
 
 " No more enduring Memorial of Professor Rankine could be devised than the publica- 
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 . . . The Volume exceeds in importance any work in the same department published 
 in our time " A rchitect. 
 
 CALCAREOUS CEMENTS: 
 
 THEIR NATURE, PREPARATION, AND USES. 
 
 Witlx some Remarks "upon Gexxa.en/t; Testiii't.gJ. 
 
 BY GILBERT R. REDGRAVE, Assoc. INST. C.E. 
 With Illustrations. 8s. 6d. 
 
 GENERAL CONTENTS. Introduction Historical Review of the Cement 
 Industry The Early Days of Portland Cement Composition of Portland 
 Cement PROCESSES OF MANUFACTURE The Washmill and the Backs 
 Flue and Chamber Drying Processes Calcination of the Cement Mixture 
 Grinding of the Cement Composition of Mortar and Concrete CEMENT 
 TESTING CHEMICAL ANALYSIS of Portland Cement, Lime, and Raw 
 Materials Employment of Slags for Cement Making Scott's Cement, 
 Selenitic Cement, and Cements produced from Sewage Sludge and the 
 Refuse from Alkali Works Plaster Cements Specifications for Portland 
 Cement Appendices (Gases Evolved from Cement Works, Effects of Sea- 
 water on Cement, Cost of Cement Manufacture, &c., &c.) 
 
 " A work calculated to be of GREAT and EXTENDED UTILITY." Chemical News, 
 
 " INVALUABLE to the Student, Architect, and Engineer." Building News. 
 
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 "Will be useful to ALL interested in the MANUFACTURE, USE, and TESTING of Cements." 
 Engineer. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 37 
 
 AND ITS PRODUGTS: 
 
 A PRACTICAL TREATISE. 
 
 BY 
 
 BOYERTON REDWOOD, 
 
 F.R.S.E., F.I.C., Assoc. INST. C.E., 
 
 Hon. Corr. Mem. of the Imperial Russian Technical Society ; Mem. of the American Chemical 
 
 Society ; Consulting Adviser to the Corporation of London under the 
 
 Petroleum Acts, &c., &c. 
 
 ASSISTED BY GEO. T. HOLLOW AY, F.I.C., Assoc. R.C.S., 
 And Numerous Contributors. 
 
 In Two Volumes, Large 8vo. Price 45s. 
 TOitb Numerous dfcaps, fiMates, anfc Illustrations in tbe 
 
 GENERAL CONTENTS. 
 
 I. General Historical Account of 
 the Petroleum Industry. 
 
 II. Geological and Geographical 
 Distribution of Petroleum and 
 Natural Gas. 
 
 HI. Chemical and Physical Pro- 
 perties of Petroleum. 
 
 IV. Origin of Petroleum and Natural 
 Gas. 
 
 V. Production of Petroleum, 
 Natural Gas, and Ozokerite. 
 
 VI. The Refining of Petroleum. 
 
 VIII. Transport, Storage, and Dis- 
 tribution of Petroleum. 
 
 IX. Testing of Petroleum. 
 
 X. Application and Uses of 
 Petroleum. 
 
 XI. Legislation on Petroleum at 
 
 Home and Abroad. 
 XII. Statistics of the Petroleum 
 Production and the Petroleum 
 Trade, obtained from the 
 most trustworthy and official 
 sources. 
 
 VII. The Shale Oil and Allied In- 
 dustries. 
 
 " The MOST COMPREHENSIVE AND CONVENIENT ACCOUNT that has yet appeared 
 of a gigantic industry which has made incalculable additions to the comfort of 
 civilised man. . . . The chapter dealing with the arrangement for STORAGE 
 
 and TRANSPORT Of GREAT PRACTICAL INTEREST. . . . The DIGEST of LEGIS- 
 LATION on the subject cannot but prove of the GREATEST UTILITY." The Times. 
 
 " A SPLENDID CONTRIBUTION to our technical literature." Chemical News. 
 
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 , ". most fully and ably handled . . . could only have been produced 
 by a man in the very exceptional position of the Author. . . . INDISPEN- 
 SABLE to all who have to do with Petroleum, its APPLICATIONS, MANUFACTURE, 
 STORAGE, or TRANSPORT." Mining Journal. 
 
 " We must concede to Mr. Redwood the distinction of having produced a 
 treatise which must be admitted to the rank of THE INDISPENSABLES. It con- 
 tains THE LAST WORD that can be said about Petroleum in any of its SCIENTIFIC, 
 TECHNICAL, and LEGAL aspects. It would be difficult to conceive of a more 
 comprehensive and explicit account of the geological conditions associated with 
 the SUPPLY of Petroleum and the very practical question of its AMOUNT and 
 DURATION." Journ. of Gas Lighting. 
 
 LONDON: EXETER STREET, STRAND. 
 
38 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 Royal 8uo, Handsome Cloth, 25s, 
 
 THE STABILITY OF SHIPS. 
 
 BY 
 
 SIR EDWARD J. REED, K.C.B., F.R.S., M.P., 
 
 KNIGHT OF THB IMPERIAL ORDERS OF ST. STANILAUS OF RUSSIA; FRANCIS JOSEPH OF 
 AUSTRIA ; MEDJIDIE OF TURKEY ; AND RISING SUN OF JAPAN ; VICE- 
 PRESIDENT OF THE INSTITUTION OF NAVAL ARCHITECTS. 
 
 With numerous Illustrations ana Tables. 
 
 THIS work has been written for the purpose of placing in the hands of Naval Constructors, 
 Shipbuilders, Officers of the Royal and Mercantile Marines, and all Students of Naval Science, 
 a complete Treatise upon the Stability of Ships, and is the only work in the English 
 Language dealing exhaustively with the subject. 
 
 In order to render the work complete for the purposes of the Shipbuilder, whether at 
 home or abroad, the Methods of Calculation introduced by Mr. F. K. BARNES, Mr. GRAY, 
 M. REECH, M. DAYMARD, and Mr. BENJAMIN, are all given separately, illustrated by 
 Tables and worked-out examples. The book contains more than 200 Diagrams, and is 
 illustrated by a large number of actual cases, derived from ships of all descriptions, but 
 especially from ships of the Mercantile Marine. 
 
 The work will thus be found to constitute the most comprehensive and exhaustive Treatise 
 hitherto presented to the Profession on the Science of the STABILITY OF SHIPS, 
 
 " Sir EDWARD REED'S ' STABILITY OF SHIPS ' is INVALUABLE. In it the STUDENT, ne\v- 
 to the subject, will find the path prepared for him, and all difficulties explained with the 
 utmost care and accuracy ; the SHIP-DRAUGHTSMAN will find all the methods of calculation at 
 present in use fully explained and illustrated, and accompanied by the Tables and Forms 
 employed ; the SHIPOWNER will find the variations in the Stability of Ships due to differences 
 in forms and dimensions fully discussed, and the devices by which the state of his ships under 
 all conditions may be graphically represented and easily understood ; the NAVAL ARCHITECT 
 will find brought together and ready to his hand, a mass of information which he would other- 
 wise have to seek in an almost endless variety of publications, and some of which he would 
 possibly not be able to obtain at all elsewhere." Steamship. 
 
 "This IMPORTANT AND VALUABLE WORK . . . cannot be too highly recommended to 
 all connected with shipping interests." Iron. 
 
 " This VERY IMPORTANT TREATISE, . . . the MOST INTELLIGIBLE, INSTRUCTIVE, an 
 
 COMPLEX* that has ever appeared." Nature. 
 
 "The volume is an ESSENTIAL ONB for the shipbuilding profession." Westmintttr 
 Review. 
 
 RICHMOND (H. Droop, F.C.S., Chemist to the 
 
 Aylesbury Dairy Company) : 
 
 DAIRY CHEMISTRY FOR DAIRY MANAGERS : A Practical 
 Handbook. ( Griffin's Technological Manuals. ) 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 39 
 
 FOURTH EDITION, REVISED. With Additional Illustrations. Price 6s. 
 
 PRACTICAL SANITATION: 
 
 A HAND-BOOK FOR SANITARY INSPECTORS AND OTHERS 
 INTERESTED IN SANITATION. 
 
 By GEORGE REID, M.D., D.P.H., 
 
 Fellow of the Sanitary Institute of Great Britain, and Medical Officer, 
 Staffordshire County Council. 
 
 TICUtb an Bppen&fj on Sanitary Xaw. 
 
 By HERBERT MAN LEY, M.A., M.B., D.P.H., 
 
 Medical Officer of Health for the County Borough of West Brom-wich. 
 
 GENERAL CONTENTS. Introduction Water Supply: Drinking Water, 
 Pollution of Water Ventilation and Warming Principles of Sewage 
 Removal Details of Drainage ; Refuse Removal and Disposal Sanitary 
 and Insanitary Work and Appliances Details of Plumbers Work House 
 Construction Infection and Disinfection Food, Inspection of ; Chara/:- 
 teristics of Good Meat ; Meat, Milk, Fish, &c., unfit for Human Food- 
 Appendix : Sanitary Law ; Model Bye-Laws, &c. 
 
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 DISINFECTION & DISINFECTANTS 
 
 (AN INTRODUCTION TO THE STUDY OF). 
 
 Together with an Account of the Chemical Substances used 
 as Antiseptics and Preservatives. 
 
 BY SAMUEL RIDEAL, D.SC.LOND., F.I.C., F.C.S., 
 
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 Chemistry, St. George's Hospital Medical School, &c., &c. 
 
 With Folding- Plate and Illustrations of the most Approved Modern 
 Appliances. 
 
 ** " Notwithstanding the rapid development of Sanitary Science in this country, there doe* 
 not exist at the present time in the English language any book which deals exclusively with the 
 composition of DISINFECTANTS. The present volume will, therefore, supply a want which hat 
 been felt not only by the chemist and bacteriologist, but also by those who are concerned with the 
 practical work of disinfection. . . ."EXTRACT FROM AUTHOR'S PREFACE. 
 
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 WORK OF BEFERBNCB." Pharmaceutical Journal. 
 
 "Aw EXHAUSTIVE TREATISE, dealing with the WHOLE RANGE of the subject: Disinfection by 
 Heat, Chemical Disinfectants, Practical Methods, Personal Disinfection, Legal Regulations, and 
 Methods of Analysis ... so very well done and so USEFUL that it wfll be valued by ALL 
 connected with Sanitation and Public Health." Chemist and Druggist. 
 
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 meni Journal. 
 
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40 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
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 THIRD EDITION, REVISED. Large Crown 8vo. Handsome Cloth. 4s. 
 
 A MANUAL OF AMBULANCE. 
 
 BY J. SCOTT RIDDELL, C.M., M.B., M.A., 
 
 Assistant-Surgeon, Aberdeen Royal Infirmary ; Lecturer and Examiner to the Aberdeen 
 
 Ambulance Association ; Examiner to the St. Andrew's Ambulance Association, 
 
 Glasgow, and the St. John Ambulance Association, London. 
 
 With Numerous Illustrations and Full Page Plates. 
 
 General Contents. Outlines of Human Anatomy and Physiology 
 The Triangular Bandage and its Uses The Roller Bandage and its Uses 
 Fractures Dislocations and Sprains Haemorrhage Wounds Insensi- 
 bility and Fits Asphyxia and Drowning Suffocation Poisoning Burns, 
 Frost-bite, and Sunstroke Removal of Foreign Bodies from (a) The Eye ; 
 (b) The Ear; (c) The Nose; (d) The Throat; () The Tissues Ambulance 
 Transport and Stretcher Drill The After-treatment of Ambulance Patients 
 Organisation and Management of Ambulance Classes Appendix : Ex- 
 amination Papers on First Aid. 
 
 "A CAPITAL BOOK. . . . The directions are SHORT and CLEAR, and testify to the 
 hand of an able surgeon." Edin. Med. Journal. 
 
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 us place in EVERY HOUSEHOLD LIBRARY." Daily Chronicle. 
 
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 Colliery Guardian. 
 
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 A MEDICAL AND SURGICAL HELP 
 
 FOR SHIPMASTERS AND OFFICERS 
 IN THE MERCHANT NAVY. 
 
 INCLUDING 
 
 FIRST AID TO THE INJURED. 
 
 BY WM. JOHNSON SMITH, F.R.C.S., 
 
 Principal Medical Officer, Seamen's Hospital, Greenwich. 
 
 With Coloured Plates and Numerous Illustrations. 
 
 *. * The attention of all interested in our Merchant Navy is requested to this exceedingly 
 useful and valuable work. It is needless to say that it is the outcome of many years 
 PRACTICAL EXPERIENCE amongst Seamen. 
 
 " SOUND, JUDICIOUS, REALLY HELPFUL." The Lancet. 
 
 "It would be difficult to find a Medical and Surgical Guide more clear and comprehensive 
 than Mr. JOHNSON SMITH, whose experience at the GREENWICH HOSPITAL eminently qualifies 
 him for the task. ... A MOST ATTRACTIVE WORK. . . . We have read it from cover 
 to cover. ... It gives clearly written advice to Masters and Officers in all medical and 
 surgical matters likely to come before them when remote from the land and without a 
 doctor. . . . We RECOMMEND the work to EVERT Shipmaster and Officer." Liverpool 
 Journal of Commerce. 
 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 41 
 
 d>riffms 
 
 STANDARD WORKS OF REFERENCE 
 
 FOR 
 
 Metallurgists, Mine-Owners, Assayers, Manufacturers, 
 
 and all interested in the development of 
 
 the Metallurgical Industries. 
 
 EDITED BY 
 
 W. C. ROBERTS-AUSTEN, C.B., F.R.S., 
 
 CHEMIST AND ASSAYER TO THE ROYAL MINT; PROFESSOR OF METALLURGY IN 
 
 THE ROYAL COLLEGE OF SCIENCE. 
 In Large &vff, Handsome Cloth. With Illustrations. 
 
 1. INTRODUCTION to the STUDY of METALLURGY. 
 
 By the EDITOR. FOURTH EDITION. Revised and Enlarged, with 
 New Illustrations, Micro-Photographic Plates of Different Varieties of 
 Steel, and Folding Plate. 153. 
 
 " No English text-book at all approaches this in the COMPLETENESS with 
 which the most modern views on the subject are dealt with. Professor Austen's 
 volume will be INVALUABLE, not only to tile student, but also to those whose 
 knowledge of the art is far advanced." Chemical News. 
 
 " INVALUABLE to the student. . . . Rich in matter not to be readily found 
 elsewhere. " Athenceum. 
 
 " This volume amply realises the expectations formed as to the result of the 
 labours of so eminent an authority. It is remarkable for its ORIGINALITY of con- 
 ception and for the large amount of information which it contains. . . . We 
 recommend every one who desires information not only to consult, but to STUDY 
 this work." Engineering. 
 
 " Will at once take FRONT RANK as a text-book." Science and Art. 
 
 " Prof. ROBERTS- AUSTEN'S book marks an epoch in the history of the teaching 
 of metallurgy in this country." Industries. 
 
 2. G-OLD (The Metallurgy of). By THOS. KIRKE ROSE, 
 
 D.Sc., Assoc. R.S.M., F.I.C., of the Royal Mint. SECOND EDITION, 
 2is. (See p. 43). 
 
 3. IRON (The Metallurgy of). By THOS. TURNER, 
 
 Assoc. R.S.M., F.I.C., F.C.S. i6s. (See p. 50). 
 
 Will be Published at Short Intervals. 
 
 4. STEEL (The Metallurgy of). By F. W. HARBORD, 
 
 Assoc R.S.M., F.LC. 
 
 5. COPPER (The Metallurgy of). By THOS. GIBB, Assoc. 
 
 Royal School of Mines. 
 
 6. METALLURGICAL MACHINERY: the Application of 
 
 Engineering to Metallurgical Problems. By HENRY CHARLES JENKINS, 
 Wh.Sc., Assoc.R.S.M., Assoc. M.Inst.C.E., of the Royal Mint. 
 
 7. ALLOYS. By the EDITOR. 
 
 %* Other Volumes in Preparation. 
 
 LONDON: EXETER STREET, STRAND. 
 
i CHARLES GRIFFIN & CO. '8 PUBLICATIONS. 
 
 SECOND EDITION, Eevised and Enlarged. 
 In Large 8vo, Handsome cloth, 34s. 
 
 HYDRAULIC POWER 
 
 AND 
 
 HYDRAULIC MACHINERY. 
 
 BY 
 
 HENRY ROBINSON, M. INST. C.E., F.G.S., 
 
 FELLOW OF KING'S COLLEGE, LONDON; PROF. OF CIVIL ENGINEERING, 
 KING'S COLLEGE, ETC., ETC. 
 
 TKHttb numerous TRUooDcuts, and SfetE*nine plates. 
 
 GENERAL CONTENTS. 
 
 Discharge through Orifices Gauging Water by Weirs Flow of Water 
 through Pipes The Accumulator The Flow of Solids Hydraulic Presses 
 and Lifts Cyclone Hydraulic Baling Press Anderton Hydraulic Lift 
 Hydraulic Hoists (Lifts) The Otis Elevator Mersey Railway Lifts City 
 and South London Railway Lifts North Hudson County Railway Elevator 
 Lifts for Subways Hydraulic Ram Pearsall's Hydraulic Engine Pumping- 
 Engines Three- Cylinder Engines Brotherhood Engine Rigg's Hydraulic 
 Engine Hydraulic Capstans Hydraulic Traversers Movable Jigger Hoist 
 Hydraulic Waggon Drop Hydraulic Jack Duckham's Weighing Machine 
 Shop Tools Tweddell's Hydraulic Rivetter Hydraulic Joggling Press 
 Tweddell's Punching and Shearing Machine Flanging Machine Hydraulic 
 Centre Crane Wrightson's Balance Crane Hydraulic Power at the Forth 
 Bridge Cranes Hydraulic Coal-Discharging Machines Hydraulic Drill 
 Hydraulic Manhole Cutter Hydraulic Drill at St. Gothard Tunnel Motors 
 with Variable Power Hydraulic Machinery on Board Ship Hydraulic Points 
 and Crossings Hydraulic Pile Driver Hydraulic Pile Screwing Apparatus 
 Hydraulic Excavator Ball's Pump Dredger Hydraulic Power applied to 
 Bridges Dock -gate Machinery Hydraulic Brake Hydraulic Power applied 
 to Gunnery Centrifugal Pumps Water Wheels Turbines Jet Propulsion 
 The Gerard- Barre" Hydraulic Railway Greathead's Injector Hydrant Snell's 
 Hydraulic Transport System Greathead's Shield Grain Elevator at Frank- 
 fort Packing Power Co-operation Hull Hydraulic Power Company 
 London Hydraulic Power Company Birmingham Hydraulic Power System 
 Niagara Falls Cost of Hydraulic Power Meters Schtmheyder's Pressure 
 Regulator Deacon's Waste- Water Meter. 
 
 "A Book of great Professional Usefulness." Iron. 
 
 %* The SECOND EDITION of the aboye important work has been thoroughly revised and 
 brought np to date. Many new full-page Plates haye been added the number being 
 increased from 43 in the First Edition to 9 in the present. Full Prospectus, giving a 
 description of the Plates, may be had on application to the Publishers. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 43 
 
 QRIFFIN'S METALLURGICAL SERIES. 
 
 SECOND EDITION. Large 8vo. Handsome Cloth. 21s. 
 
 THE METALLURGY OF GOLD. 
 
 BY 
 
 T. KIRKE ROSE, D.Sc., Assoc.KS.M., F.I.C., 
 
 Assistant Assayer of the Royal Mint. 
 
 Revised and partly Re-written. Including the most recent Improve- 
 ments in the Cyanide Process, and a new Chapter on Economic 
 Considerations (Management, Cost, Output, &c.). With 
 Frontispiece and additional Illustrations. 
 
 LEADING FEATURES. 
 
 1. Adapted for all who are interested in the Gold Mining Industry, being 
 free from technicalities as far as possible ; of special value to those engaged in 
 the industry viz. , mill-managers, reduction-officers, &c. 
 
 2. The whole ground implied by the term " Metallurgy of Gold " has been 
 covered with equal care ; the space is carefully apportioned to the variou* 
 branches of the subject, according to their relative importance. 
 
 3. The MACARTHUR-FORREST CYANIDE PROCESS is fully described for the 
 first time. By this process over 2,000,1)00 of gold per annum (at the rate of) i 
 now being extracted, or nearly one-tenth of the total world's production. The 
 process, introduced in 1887, has only had short newspaper accounts given of it 
 previously. The chapters have been submitted to, and revised by, Mr. 
 MacArthur, and so freed from all possible inaccuracies. 
 
 4. Among other new processes not previously described in a text-book 
 
 (1) The modern barrel chlorination process, practised with great success in 
 Dakota, where the Black Hills district is undergoing rapid development owing 
 to its introduction. (2) New processes for separating gold from silver viz., the 
 new Gutzkow process, and the Electrolytic process ; the cost of separation is 
 reduced by them by one-half. 
 
 5. A new feature is the description of EXACT METHODS employed in particular 
 extraction works Stamp-batteries of South Africa, Australia, New Zealand, 
 California, Colorado, and Dakota ; Chlorination works also, in many parts of 
 the world ; Cyanide works of S. Africa and New Zealand. These accounts are 
 of special value to practical men. 
 
 6. The bibliography is the first made since 1882. 
 
 " Dr. Ross gained his experience in the Western States of America, but he has secured 
 details of gold-working from ALL PARTS of the world, and these should be of GKIAT SERVICE 
 to practical men, . . . The four chapters on Chlorination, written from the point of view 
 alike of the practical man and the chemist, TEEM WITH COHSIDERATIOWB HITHERTO UNBICOG- 
 NISID, and constitute an addition to the literature of Metallurgy, which will prove to be of 
 classical value." Nature. 
 
 "The most complete description of the chlorination process which has yet been published. 
 Mining Journal. 
 
 LONDON: EXETER STREET, STRAND. 
 
44 CHARLES GRIFFIN & CO.'S PUBLICATIONS. 
 
 Companion-Volume to MM. Knecht and Rawson's "Dyeing." 
 
 TEXTILE PRINTING: 
 
 A PRACTICAL MANUAL. 
 
 Including the Processes Used in the Printing of 
 
 COTTON, WOOLLEN, SILK, and HALF-SILK FABRICS. 
 
 Br C. F. SEYMOUR ROTHWELL, F.C.S., 
 
 A/em. Soe. of Chemical Industries ; late Lecturer at the Municipal Technical School, Manchester. 
 In Large 8vo, with Illustrations and Printed Patterns. Price 2 is. 
 
 General Contents. Introduction The Machinery Used in Textile Printing 
 Thickeners and Mordants The Printing of Cotton Goods The Steam Style 
 Colours Produced Directly on the Fibre Dyed Styles Padding Style 
 Resist and Discharge Styles The Printing of Compound Colourings, &c. 
 The Printing ot Woollen Go9ds The Printing of Silk Goods Practical 
 Recipes for Printing Appendix Useful Tables Patterns. 
 
 " BY FAR THE BEST and MOST PRACTICAL BOOK on TEXTILE PRINTING which has yet been 
 brought out, and will long remain the standard work on the subject. It is essentially 
 practical in character." Te ciile Mercury. 
 
 " THE MOST PRACTICAL MANUAL of TEXTILE PRINTING which has yet appeared. We have 
 no hesitation in recommending it." The Textile Manufacturer. 
 
 '' UNDOUBTEDLY the book is THE BEST which has appeared on TEXTILE PRINTING, and 
 worthily forms a Companion- Volume to ' A Manual on Dyeing. 1 "The Dyer and Calico 
 Printer. 
 
 SCHWACKHOFER and BROWNE: 
 
 FUEL AND WATER: A Manual for Users of Steam and Water. 
 By Prof. FRANZ SCHWACKHOFER of Vienna, and WALTER 
 R. BROWNE, M.A., C.E., late Fellow of Trinity College, Cambridge. 
 Demy 8vo, with Numerous Illustrations, g/. 
 
 GENERAL CONTENTS Heat and Combustion Fuel, Varieties of Firing Arrange- 
 ments : Furnace, Flues, Chimney The Boiler, Choice of Varieties Feed-water 
 Heaters Steam Pipes Water : Composition, Purification Prevention of Scale, &c., &c. 
 
 "The Section on Heat is one of the best and most lucid ever written." Engineer. 
 " Cannot fail to be valuable to thousands using steam power." Railway Engineer. 
 
 SHELTON-BEY (W. Vincent, Foreman to the 
 
 Imperial Ottoman Gun Factories, Constantinople) : 
 
 THE MECHANIC'S GUIDE : A Hand- Book for Engineers and 
 Artisans. With Copious Tables and Valuable Recipes for Practical Use. 
 Illustrated. Second Edition. Crown 8vo. Cloth, 7/6. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 45 
 
 Thirteenth Edition. Price 21s. 
 
 Demy 8vo, Cloth. With Numerous Illustrations, reduced from 
 Working Drawings. 
 
 A MANUAL OP 
 
 MARINE ENGINEERING: 
 
 COMPRISING THE DESIGNING, CONSTRUCTION, AND 
 WORKING OF MARINE MACHINERY. 
 
 By A. E. S E AT N, M. Inst. C. E., M. Inst. Meeh. E., 
 M.InstN.A. 
 
 GENERAL CONTENTS. 
 
 Part I. Principles of Marine 
 Propulsion. 
 
 Part XL-Principles of Steam Valves, &e. 
 
 Engineering. 
 
 Part III. Details of Marine 
 Engines: Design and Cal- 
 
 culations for Cylinders, 
 Pistons, Valves, Expansion 
 
 Part IV. Propellers. 
 Part V. Boilers. 
 Part VI. Miscellaneous. 
 
 *** The THIRTEENTH EDITION includes a Chapter on WATEK-TUBE BOILERS, 
 with Illustrations of the leading Types. 
 
 "In the three-fold capacity of enabling a Student to learn how to design, construct, 
 and work a Marine Steam- Engine, Mr. Seatpn's Manual has NO RIVAL." Timts. 
 
 "The important subject of Marine Engineering is here treated with the THOROUGH- 
 NESS that it requires. No department has escaped attention. . . . Gives the 
 results of mucn close study and practical work." Enairteering. 
 
 "By far the BEST MANUAL in existence. . . . Gives a complete account of the 
 methods of solving, with the utmost possible economy, the problems before the Marine 
 Engineer." Athenaeum. 
 
 " The Student, Draughtsman, and Engineer will find this work the MOST VALUABLS 
 HANDBOOK of Reference on the Marine Engine now in existence." Marine EngtJteer. 
 
 FOURTH EDITION. With Diagrams. Pocket-Size, Leather. 8s. 6d. 
 A POCKET-BOOK OF 
 
 MARINE ENGINEERING RULES AND TABLES, 
 
 FOR THE USE OF 
 
 Marine Engineers, Naval Architects, Designers, Draughtsmen, 
 Superintendents and Others. 
 
 BY 
 
 A. E. SEATON, M.I.O.E., M.I.Mech.E., M.I.N.A., 
 
 AND 
 
 H. M. ROUNTHWAITE, M.I.Mech.E., M.I.N.A. 
 
 "ADMIRABLY FULFILS its purpose." Marine Engineer. 
 LONDON: EXETER STREET, STRAND. 
 
46 CHARLES GRIFFIN A CO.'S PUBLICATIONS. 
 
 WORKS BY A. HUMBOLDT SEXTON, F.I.C., F.C.S., 
 
 Professor of Metallurgy in the Glasgow and West of Scotland Technical College. 
 
 In Large Crown 8vo, Handsome Cloth, 6s. 
 
 ELEMENTARY METALLURGY 
 
 (A TEXT-BOOK OF). 
 
 Including the Author's PRACTICAL LABORATORY COURSE. 
 With Numerous Illustrations. 
 
 GENERAL CONTENTS. Introduction Properties of the Metals Combustion 
 Fuels Refractory Materials Furnaces Occurrence of the Metals in Nature Pre- 
 paration of the Ore for the Smelter Metallurgical Processes Iron : Preparation of 
 Pig Iron Malleable Iron Steel Mild Steel Copper Lead Zinc and Tin Silver 
 Gold Mercury Alloys Applications of ELECTRICITY to Metallurgy LABORA- 
 TORY COURSE WITH NUMEROUS PRACTICAL EXERCISES. 
 
 " The volume before us FULLY ENHANCES and confirms PROF. SEXTON'S reputa- 
 tion. . . . Just the kind of work lor Students COMMENCING the study of Metal- 
 lurgy, or for ENGINEERING Students requiring a GENERAL KNOWLEDGE of it, or 
 for ENGINEERS in practice who like a HANDY WORK of REFERENCE. To all three 
 classes we HEARTILY commend the work." Practical Engineer. 
 
 " EXCELLENTLY got-up and WELL-ARRANGED. . . . Iron and copper well 
 explained by EXCELLENT diagrams showing the stages of the process from start to 
 finish. . . . The most NOVEL chapter is that on the many changes wrought 
 in Metallurgical Methods by ELECTRICITY." Chemical Trade Journal. 
 
 " Possesses the GREAT ADVANTAGE of giving a COURSE OF PRACTICAL WORK." 
 Mining Journal. 
 
 Sexton's (Prof.) Outlines of Quantitative Analysis. 
 
 FOR THE USE OF STUDENTS. 
 With Illustrations. FOURTH EDITION. Crown 8vo, Cloth, 3s. 
 
 " A COMPACT LABORATORY GUIDE for beginners was wanted, and the want has 
 been WELL SUPPLIED. ... A good and useful book." Lancet. 
 
 Sexton's (Prof.) Outlines of Qualitative Analysis. 
 
 FOR THE USE OF STUDENTS. 
 With Illustrations. THIRD EDITION. Crown 8vo, Cloth, 3s. 6d. 
 
 " The work of a thoroughly practical chemist." British Medical Journal. 
 " Compiled with great care, and will supply a want." Journal of Education. 
 
 LONDON: EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 47 
 
 In Large 8vo. Handsome Cloth. Price 21s. 
 
 BREWING: 
 
 (THE PRINCIPLES AND PRACTICE OF). 
 
 FOR THE USE OF STUDENTS AND PRACTICAL MEN. 
 
 BY 
 
 WALTER J. SYKES, M.D., D.P.H., F.I.C., 
 
 EDITOR OF "THE ANALYST.'' 
 With Illustrations and Plates. 
 
 %* This work is intended to present a description of the FUNDA- 
 MENTAL PRINCIPLES on which the Art of Brewing is based, and also a bird's 
 eye view of that Art as carried on in accordance with THE BEST AND MOST 
 
 SCIENTIFIC MODERN METHODS. 
 
 In Large 8vo. Handsome Cloth. With numerous Illustrations. 
 
 TECHNICAL MYCOLOGY: 
 
 THE UTILIZATION OF MICRO-ORGANISMS IN THE 
 
 ARTS AND MANUFACTURES. 
 
 A Practical Handbook on Fermentation and Fermentative Pro- 
 cesses for the Use of Brewers and Distillers, Analysts, 
 Technical and Agricultural Chemists, and all 
 interested in the Industries dependent 
 
 on Fermentation. 
 BY DR. FHANZ LAFAR, 
 
 Of the Royal Experimental Station for Industries dependent on Fermentation, 
 Hohenheim, near Stuttgart. 
 
 With an Introduction by DK. EMIL CHR. HANSEN, Principal of the 
 Carlsberg Laboratory, Copenhagen. 
 
 TRANSLATED BY CHARLES T. C. S ALTER. 
 
 In Two Volumes, sold Separately. 
 
 \* It is hoped that the English version of the First Volume of this novel and 
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 The extraordinary rdles played by Micro-organisms in Brewing and Distilling in the 
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 turein Agricultural Industries and the Processes connected with the Dairy : Souring of 
 Cream, Cheese-Ripening, &c., &c. combine to make Dr. Lafar's Text-Book one of great 
 value and interest to a very wide circle of readers. 
 
 LONDON: EXETER STREET, STRAND. 
 
48 CHARLES GRIFFIN Jc CO. 'S PUBLICATIONS. 
 
 WORKS BY PROF. ROBERT H. SMITH, Assoe.M.I.C.E., 
 
 M.I.M.E., M.I.E1.E., M.Fed.I.Mi.E., Whit. Sch., M.Ord.Meiji. 
 
 MEASUREMENT CONVERSIONS 
 
 (English and French) : 
 28 GRAPHIC TABLES OR DIAGRAMS. 
 
 Showing at a glance the MUTUAL CONVERSION of MEASUREMENTS 
 
 in DIFFERENT UNITS 
 Of Lengths, Areas, Volumes, Weights, Stresses, Densities, Quantities 
 
 of Work, Horse Powers, Temperatures, &c. 
 For the use of Engineers, Surveyors, Architects, and Contractors. 
 
 In 4to, Boards. 7s. 6d. 
 
 %* Prof. SMITH'S CONVERSION-TABLES form the most unique and com- 
 prehensive collection ever placed before the profession. By their use much 
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 diminished. It is believed that henceforth no Engineer's Office will be 
 considered complete without them. 
 
 " The work is INVALUABLE." Colliery Guardian. 
 
 " Ought to be in EVERT office where even occasional conversions are required. . . . Prof. 
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 "Prof. Smith deserves the hearty thanks, not only of the ENGINEER, hut of the COMMERCIAL 
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 a subject which is now assuming great importance as a factor in maintaining our HOLD upon 
 doub 
 
 FOREIGN TRADE. There can be no doubt that the antiquated system of Weights and Measures 
 used in this country is doomed to be superseded by the much simpler metho 
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 THE CALCULUS FOR ENGINEERS 
 AND PHYSICISTS, 
 
 Applied to Technical Problems. 
 
 WITH EXTENSIVE 
 
 CLASSIFIED REFERENCE LIST OF INTEGRALS. 
 By PROF. ROBERT H. SMITH. 
 
 ASSISTED BY 
 
 ROBERT FRANKLIN MUIRHEAD, 
 
 M.A., B.Sc. (Glasgow), B.A. (Cambridge), 
 
 Formerly Clark Fellow of Glasgow University, and Lecturer on Mathematics at 
 Mason College. 
 
 In Crown 8vo, extra, with Diagrams and Folding-Plate, 8s. 6d. 
 LONDON : EXETER STREET, STRAND. 
 
SCIENTIFIC AND TECHNOLOGICAL WORKS. 49 
 
 By PROFESSORS J. J. THOMSON & POYNTING. 
 
 In Large 8vo. Fully Illustrated. 
 
 A TEXT-BOOK OF PHYSICS: 
 
 COMPRISING 
 
 PROPERTIES OF MATTER; HEAT; SOUND AND LIGHT; 
 MAGNETISM AND ELECTRICITY. 
 
 BY 
 
 J. H. POYNTING, J. J. THOMSON, 
 
 SC.D., F.B.S., AND M - A - *-B.S.i 
 
 bate Fellow of Trinity College, Cambridge; Fellow of Trinity College, Cambridge; Prof. 
 
 Professor of Physics, Mason College, of Experimental Physics in the University 
 
 Birmingham. of Cambridge. 
 
 THIRD EDITION, Revised and Enlarged. Pocket-Size, Leather, also for Office Use, Cloth, 12s. fid. 
 
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 BY T. W. TRAILL, M. INST. 0. E., F. E. R. N. f 
 
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 4 
 
CHARLES GfilFFIN & CO. 'S PUBLICATIONS. 
 
 GRIFFIN'S METALLURGICAL SERIES. 
 
 THE METALLURGY OF IRON. 
 
 THOMAS TURNER, Assoc.R.S.M, F.I.C., 
 
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 IN LARGE 8vo, HANDSOME CLOTH, WITH NUMEROUS ILLUSTRATIONS 
 (MANY FROM PHOTOGRAPHS). PRICK 16s. 
 
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 Early History of Iron. 
 
 Modern History of Iron. 
 
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 Further Treatment of Wrought 
 
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SCIENTIFIC AND TECHNOLOGICAL WORKS. 51 
 
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 An Introductory Text-book on the Theory, Design, Construction, 
 and Testing of Internal Combustion Engines without Boiler. 
 
 FOR THE USE OF STUDENTS. 
 
 BY 
 
 PROF. W. H. WATKINSON, WHIT. Sen., M.INST.MECH.E., 
 
 Glasgow and West of Scotland Technical College. 
 In Crown 8vo, extra, with Numerous Illustrations. [Shortly. 
 
 LONDON : EXETER STREET, STRAND. 
 
52 CHARLES ORIFFIN <k CO.' 8 PUBLICATIONS. 
 
 WORKS BY DR. ALDER WRIGHT, F.RS. 
 
 FIXED OILS, FATS, BUTTERS, AND WAXES: 
 
 THEIR PREPARATION AND PROPERTIES, 
 
 And the Manufacture therefrom of Candles, Soaps, and 
 Other Products. 
 
 BY 
 
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 Late Lecturer on Chemistry, St. Mary's Hospital School ; Examiner in "Soap" to the 
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 In Large 8vo. Handsome Cloth. With 144 Illustrations. 283. 
 
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SCIENTIFIC AND TECHNOLOGICAL WORKS. 53 
 
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 VOL. II. MACHINE AND ENGINE DRAWING AND DESIGN. 4s. 6d. 
 
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 In Large Grown Svo. With Numerous Illustrations. 
 
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