The Metallurgy of the Common Metals, Gold, Silver, Iron, Copper, Lead, and Zinc by LEONARD S. AUSTIN 1 1 Professor of Metallurgy and Ore Dressing, Michigan College of Mines. Second Edition Revised and Enlarged 1909 Published by the Mining and Scientific Press, San Francisco, and The Mining Magazine, London. COPYRIGHT, 1909 BY DEWEY PUBLISHING COMPANY. ADDENDA Page. 22 line .6, for CO read CO 2 . 29 " 9 for - 62 '^~ read ^A J > I T 12X07 redQ 12 29 " 37, for 9.77 read 9.86. 29 " 39, for 87% read 77%. 30 " 1, for 8.5 Ib. read 7.6 Ib. 30 " 2, for 0.87 read 0.85. 30 " 4, for 'the table' read 'Fig. 4'; for 'the heat values' read 'for 2000 the specific heat'. 30 " 6, for 8.5 read 7.6 ; for 0.251 read 0.281. 54 Omit the sentence, lines 38 and 39. 81 line 19, for 'in the manner seen' read 'in the reaction (6) '. 81 ' 21 and 22, omit reaction (8) and in its place put reaction (9). 81 " 25, omit 'sulphate is formed thus'. 81 " 28, omit 'even more'; also, 'than the former'. 153 last line, omit 'in the same general way'. 155 line 1, for chlorine read chlorate. 155 " 20, for Fe(OH) 4 read Fe'(OH) 2 . 195 Diagram, for 5% washes read 0.05% washes; for washes of 25% and 15% read washes of 0.25% and 0.15% ; for Strong Solution 25% KCN read Strong Solution 0.25% KCN; for Medium Solution 15% KCN read Medium Solution 0.15%; KCN; for Weak Solution 5% KCN read Weak Solution 0.05% KCN. 198 line 14, for 86a read 87a. 223 37, for copper sulphide read copper sulphate. 224 " 29, for HgS read Hg 2 S. 244 " 6, f or = 3Hg read = 3CuHg. 260 " 26, for Na 2 S 2 O 8 Cu 2 O 8 read 2Na 2 S 2 3 , 3Cu 2 S 2 O 3 . 270 ' ' 5, for 30 to 40 read 3000 to 4000. 270 " 18, for 335 read 135. 289 " 8, for Fe read 4Fe. 289 " 14, for - 1400 Cal. read + 1400 Cal. 289 last line, for + 1400 Cal. read + 1600 Cal. 291 line 37, for 1100 read 1112. 297 ' ' 10, for native read 34.4%. 306 ' ' 8, for Fig. 130 read Fig. 132. 320 Insert line 23 between lines 5 and 6. 323 line 30, for 112 Ib. read 1112 Ib. 331 22, for sulphide read sulphate. 332 " 17, for 6CuSO 2 read 6Cu + SO 2 . (OVER) Page. 339 line 33, for Cu 2 S read Cu 2 O. 353 " next to last for 'ferrous' read 'sodium', and omit 'adding'. 353 last line, for 'to' read 'in'. 354 line 1, before 'common salt' insert '200 parts of. 354 " 3, after 'CuO' insert 'and Cu 2 O', omit 'and dissolves'. 354 Omit lines 5, 6, and 7. 354 line 10, for 2Cu read Cu. -^ 354 " 15 Reaction (4) should read 2CuCL, + Cu 2 CL + 3Fe - 4Cu + 3FeCl 2 . 354 " 20, for one read three; for (56)' read (168) ; for two read four; for (126) read .(252). 354 " 24, for 3Fe 2 Cl read 2Fe. 2 Cl (i . 366 " 24, for sulphide read sulphate. 366 " 26, for SO 2 read 2SO 2 . 376 " 12, for 157 read 159. 380 ' ' 21, for Pb + FeO read PbS + FeO. 382 " 36, for C read CO. 368 " 6 and 7, for 'the amount of lime - - for Baryta' read 'the percentage of magnesia is multiplied by 1.4, and of baryta by 0.4'. 387 Charge sheet, last line, for ^ read ^ e 387 line 9, for 'ratains 1% ' read 'retains 0.85% '. 387 ' ; 10, for l~ or 600 Ib. read ^ or 680 Ib. 387 " 13, for 1% of 600 Ib. read 0.85% of 680 Ib. 399 " 15, for 'they are frequently contaminated with impurities' read 'blende as black-jack frequently contains iron.' 399 " 16, omit 'especially iron'. 403 " 39, for 0.95% read 95%. 403 ' ' 40, for retained read recovered. 406 " 21, for retort read condenser. 410 " 35, for 1500C. read 500C. ; for 550 read 85C. ; for boil- ing read melting. 420 " 32 and 33, for ' (!%s place in the gases, and they impart their heat to the cold charge, which is being supplied as fast as the ore sinks below the required level. In this cupola, where the operation is one only of melting, to attain the greatest economy of fuel, the coke should, be dense, the pieces large, and the blast abundant to supply plenty of air. Thus, burning the coke is deferred to the last, less CO is formed, and the combustion, performed largely in zone 1, is more nearly complete, developing the largest possible amount of heat. From equation (1) we find, that to burn one pound of carbon to CO 2 , and thus with the greatest development of heat, there is needed 2.66 Ib. oxygen, or 11.6 Ib. air, since air contains 23% oxygen by weight. At the sea-level 12.4 cu. ft. air weigh one pound. This makes 143.8 cu. ft., or, in round numbers, 150 cu. ft. air per pound 28 THE METALLURGY of carbon. Ordinary coke contains 85% carbon, thus requiring 122 cu. ft. air per pound of such coke. While in theory 12 Ib. air should be sufficient per pound of coal, it has been found that excess is needed for complete combustion. For natural draft, using a thin fire, 18 to 24 Ib. air has given the most satisfactory results, and where air is forced into a closed ash-pit, and through the fire-bed (undergrate blast), then 16 Ib. air, or even less, is sufficient. Figuring from equation (2) in the same way, we find, per pound of carbon 1.33 Ib. of oxygen required, or of air 5.79 Ib., equal to 71.8 cu. ft. Upon the basis of coke containing 85% combustible matter, 61 cu. ft. air are required per pound of fuel when burned to CO. Chimneys or stacks. In a furnace reaction not only is it neces- sary that the reacting elements be present in mutual contact, but that the products of the reaction be removed as fast as formed. Under conditions other than this the reaction ceases. A draft, there- fore, must be provided, to carry away the waste gas, and to expel it into the atmosphere. This draft may be natural or forced. To insure obtaining a sufficient draft in a chimney, the gases must be delivered into the stack while hot. A temperature of 200 C. is ample for this, but since the work of most furnaces is done at a tempera- ture higher than a red-heat, the excess may be utilized to generate steam by conducting the gases through waste-heat boilers before entering the stack. In a reverberatory-furnace, used for smelting, the quantity and intensity of the heat depend upon the amount of coal burned per hour. This varies between 18 and 40 Ib. per square foot of grate area and, to completely burn it, there will be needed 150 cu. ft. air per pound of coal consumed. In these furnaces the gases escape at temperatures between 300 and 1100C., and move with a velocity of 12 to 20 ft. per second. For illustration, take the furnace Fig. 138 and 139, with a grate-area of 56 sq. ft., a consumption of 30 Ib. of coal per square foot of grate-area per hour, needing 252,000 cu. ft. of free air. Allowing a temperature in this instance of 1000C., and a draft velocity of 20 ft. per second at this high temperature, and knowing that these gases expand 1 / 273 of their volume for each de- gree above 0C., we find the volume at 1000C. to be m V 73 273 =4.7 times the volume at 0C. Assuming the temperature of the outside air to be 0C., we shall have as the volume of hot gas per hour 1,184,400 cu. ft. At 20 ft. per second, or 72,000 ft. per hour, this will be an area of stack of ^ o'oo = 15 sq. ft. The actual area is 16 square feet. OF THE COMMON METALS. 29 The total pull, or suction, that a chimney can produce, assuming it to be filled with hot gases, is due simply to the ascensive force of the gas measured by the difference between its weight and the weight of an equal volume of the cold air outside. To maintain the velocity of the gas in the stack, it has been found that a suction, or 'pull', of 0.4 to 0.8 in. of water, as measured by a water-gauge, is needed. Taking a draft of 0.6 in. in the above instance and adding 0.1 in. for friction in the chimney, we have 0.7 in. water equal to i2 6 xfca^ 3 - 647 lb - P er s( l- ft A cubic foot air at C ' wei ^ hs 0.0807 lb. (12.4 cu. ft. per lb.). The gas inside the stack has a spe- cific gravity of 1.03, weight of air being unity, thus making the weight when heated M*vu - \_Wood . 20.00 40.00 1.60 3100 B Peat . . 24.20 45.30 27. 00 3.30 0.20 Gallup, N. M C Lignite 12.14 47.63 32.81 7.42 Hams Fork, Wyo. D Lignite 7.75 50.60 35.10 6.55 Pittsburg, Pa E Bituminous .... 3.00 48.50 38.25 7.50 2.75 7110 Connellsville, Pa. F Bituminous 2.90 52. Ou 33.60 8.20 1.60 7333 El Moro, Colo G Bituminous 0.95 56.41 29.82 12.82 0.41 Pocahontas, W.V. H Bituminous 0.69 73.02 19.96 5.67 0.66 Bennington, Pa... Colorado I Hemi-bitum'ous J Semi-anthracite 1.73 2.27 67.03 78 83 23.89 8.83 6.69 9.39 0.66 0.68 Pennsylvania K Anthracite 2.98 87.13 3.88 5.86 0.65 Rhode Island i. Anthracite 1.18 85.70 3.80 8.52 0.80 M Graphite 99. (0 1.00 Fig. 6. TABLE OF NATURAL SOLID FUELS. The table (Fig. 6), gives the proximate analysis of a variety of fuels, the calorific value of some of these, and also the sulphur con- tained. The latter element is of importance in smelting iron ores, 34 THE METALLURGY since it must be eliminated from the iron which it would render brittle. Bituminous coals. These are distinguished from lignites by their deep-black streak, greater density, and lamellar structure. They contain but little water when first mined. We may distinguish the following different kinds : (1) Non-coking coal with a long flame. These coals closely approach lignites, furnish 55 to 60% pulverulent coke, and give a long smoky flame. The average composition of the dry coal is : Per cent. Fixed carbon 54 Volatile constituent 42 Ash 4 Water 4 The calorific power varies from 8000 to 8500 calories. (2) Coking long-flame gas-coal. These coals give 60 to 68% friable, porous coke. They contain, when air-dried : Per cent. Fixed carbon 63 Volatile constituent 31 Ash 6 Water 6 The calorific power varies from 8500 to 8800 calories. (3) Furnace coal. These coals are black, and never hard. They burn with a long smoky flame, softening the while, and swelling in the fire. They yield 68 to 14% of a swollen coke, and when quite dry are shown by analysis to contain : Per cent. Fixed carbon 69.0 Volatile constituent 26.5 Ash 4.5 The calorific power of the dry coal varies from 8800 to 9300 calories. (4) Coking coal with a short flame (semi-bituminous coal). These yield 74 to 82% of a compact coke and contain, when dry: Per cent. Fixed carbon 73 Volatile constituent 20 Ash 7 The calorific power varies from 9000 to 9600 calories. Anthracite coal. (1) Semi-anthracite. These coals burn with a OF THE COMMON METALS. .35 short flame and yield 82 to 92% of a pulverulent (sometimes fritted) coke. They are products of transition to true anthracite. Their composition on a dry basis may be stated as : Per cent. Fixed carbon 83 Volatile constituent 11 Ash 6 The calorific power varies from 9200 to 9500 calories. The percent- age of ash ranges from 1 sometimes to 30, but seldom exceeds 7. (2) Anthracite proper. This is the final product in the trans- formation of vegetable matter into coal. It is of a jet-black color, of a vitreous lustre, homogeneous structure, and conchoidal fracture. When water-free its composition is: Per cent. Fixed carbon 89 Volatile constituent 4 Ash 6 Anthracite burns almost without flame. The carbonaceous residue after distillation shows no sign of coking. In conjunction with coke for use in the blast-furnace for making pig-iron, Pennsylvania an- thracite is used. Coals of very different properties may appear alike if represented only by proximate analyses. The comparative calorific value may be judged of by Berthier's method. This consists practically in the operations of a lead assay, using an excess of litharge, with a gram of the fuel, and noting the size of lead button reduced. One can also judge a good deal about the character of the coal by coking it in a covered crucible and weighing the coke produced, judging the char- acter by the appearance of the product. The proximate analyses (Fig. 6), showing the different kinds of coal, determine to which class any given kind belongs. Graphite. This is of interest, not as a fuel, but as a refractory material, particularly when combined with clay. The analysis given in the table is that of a pure form of graphite. Ceylon graphite con- tains 79 A% carbon and 15.5% ash and some volatile matter. Petroleum or fuel oil. This is the most concentrated of fuels, and, when the cost justifies, can be used not only for generating steam, but for roasting and melting. It will be found, in burning fuel-oil from various localities, that the calorific power is much the same for the different kinds. Beaumont (Texas) oil has a calorific power of 10,820 calories, and a specific gravity of 0.88 (iy 3 Ib. per 36 THE METALLURGY gallon). Oil can be burned in such a way as to give, not only a high and uniform temperature, but also the oxidizing (roasting), or re- ducing action that may be desired. The air for combustion is best pre-heated as well as the oil, and it will be found advantageous to inject the oil under a high steam-pressure. A mixture of light and heavy oils should not be used. In Russia where oil has been em- ployed in open-hearth steel-furnaces of 10 to 15 tons capacity, oil to the extent of 15 to 20% of the weight of the charge has been used. As regards comparative costs at the Selby Smelting & Lead Works, Vallejo Junction, California, it was found that the saving by using oil was 40 to 60% with oil at $1.71 per bbl. (42 gal.), and coal at $6 per ton. A suitable control of the grade of the matte was possible by the regulation of the flame. Natural gas. In Ohio, Indiana, and Kansas, particularly, there are districts where natural gas has been obtained by boring for it, as for oil. It is the most efficient of natural fuels, having a calorific power of 611 cal. per cu. ft., or 27,861 cal. per pound. The following analysis will give an idea of the composition of Pennsylvania natural gas. It shows that natural gas is composed chiefly of marsh-gas and hydrogen : Per cent, by volume. Carbon-dioxide (CO 2 ) 0.8 Carbon-monoxide (CO) 1.0 Oxygen (0 2 ) 1.1 Ethylene (C 2 H 4 ) 1.0 Ethane (C 2 H 6 ) 3.6 Methane (marsh-gas) (CH 4 ) 72.2 Hydrogen (H 2 ) 20.7 Charcoal. Wood, packed in a kiln, and permitted to partly burn, changes by distillation of the volatile portion by the heat pro- duced from the portion burned into charcoal. The charcoal retains the form of the wood from which it was made, but has a specific gravity of only 0.2. It is of a dull-black color, soils the fingers but slightly if of good quality, but much if poor. It should ring when struck, and should show the annual rings of the wood distinctly. The density of charcoal varies with that of the wood from which it was made, dense woods giving a dense charcoal. A heaped bushel (1.5555 cu. ft.) weighs 14 to 16 Ib. When apparently quite dry, char- coal still contains 10% or more of moisture. Dry charcoal contains 95% carbon, 1.5% ash, and has a calorific power of 7610 pound-calo- OF THE COMMON METALS. 37 ries per pound. Charcoal is used in iron blast-furnaces particularly in localities where wood is abundant ; and it produces a pure, strong iron, free from sulphur, called ' charcoal-iron '. Charcoal has been used also for silver-lead and copper smelting in districts difficult of access. In these cases it has done especially well when coke could be secured to use in conjunction with it. It is, however, a friable fuel, making fine dust sometimes to the extent of 10% ; and this 'fine' is apt to make trouble in the blast-furnace. If under-burned, it is heavier and more dense, and has a brown color. Portions of the wood found imperfectly burned are called 'brands' and are returned for the next burning. Charcoal is generally made in a kiln. One of these is shown in section in Fig 7, showing method of filling. The kiln is set at the foot of a steep bank so that it can be charged conveniently from Fig. 7. SECTION OF CHARCOAL KILN. above. It has two charge-doors A and B. The first of the wood is conveyed through the lower door, and placed. The remainder is brought along the run-way c, and introduced through the upper door B. There are three rows of openings, 3 by 4 in. in size, spaced 2 ft. apart, around the bottom of the kiln. The kiln is lighted at the lower door, and when fairly started, both openings A and B are closed with sheet-iron doors. These are tightly luted with clay, and the air is thus caused to enter by the small holes. When combustion has progressed sufficiently, these openings are tightly closed, and the kiln is permitted to cool slowly. The period of charring or burning is eight days, and the cooling four days additional. Such a kiln holds 25 cords of wood and produces 1125 bushels of charcoal weighing 16 Ib. per bushel, or about 20% of the weight of wood charged. By-product charcoal. An example of the modern method of 38 THE METALLURGY making by-product charcoal for iron blast-furnace use is one at the Pioneer Iron furnace, Marquette, Michigan. Here there are 86 kilns each holding 80 cords. The daily requirement is 20 carloads, of 16 cords each, amounting to 320 cords. The kiln is packed full of wood, the sheet-iron doors put on and closed, and fire is started at a man- hole in the apex of the dome. As soon as combustion gains sufficient headway, this opening is closed, and smoke escapes by way of a flue leading from the base of the kiln to the chimney, continuing thus until most of the aqueous vapor has escaped. At this stage the chimney is closed, and the vapors pass by a smoke-main to the con- densers, the current being aided by an electrically driven fan. The cold surface of the copper tubes of this condenser precipitates the condensible portion of the gas, while the gas itself goes on to the boilers, where it is burned for steam-making. The condensible por- tion, amounting to 41% of the weight of the wood, is called green liquor or pyroligneous acid, and consists mostly of water, but con- tains also alcohol, tar, ammonia compounds, acetone, and acetic acid. The tar is separated in settling tanks, and the liquor passes to the primary still-house. Copper stills here remove the vapors of alcohol, acetic acid, and much water from the liquor. The neutralizing tank receives the product, and into this is mechanically stirred milk-of- lime to neutralize the acid by the formation of acetate of lime. The neutralized liquor is allowed to settle, and the supernatent solution is drawn off and conveyed to the refining-still house. By fractional distillation a crude wood alcohol is obtained here, and a solution of acetate of lime is left behind and recovered by evaporating the solu- tion. The crude alcohol is then purified by further distillation until a clear 95% wood-alcohol is obtained. A cord of wood (4500 lb.), yields 880 lb., or 19.5% dense charcoal of 20 lb. per bushel, 208 gal. of pyroligneous acid, 8 gal. of wood-tar, 64 lb. gray acetate of lime, and 4 gal. wood alcohol. By the sale of the wood-alcohol, acetate of lime and formaldehyde, and by the superior quality and consequently higher price of charcoal-iron, it has been possible to build up this industry, where the supply of wood is abundant, in spite of the seri- ous competition of iron smelted in blast-furnaces using coke. Coke. This is made from coal in kilns, in a way similar to that of making charcoal. Bituminous coal which cokes or fuses at the high temperature of the kiln or oven is used for this purpose. The raw screenings, in the example below, contained much fine passing a lV2-m. bar-screen. From this, the residue left after remov- ing the lump of merchantable coal, coke was made. By washing the fixed carbon was increased and the ash in the coke reduced to OF THE COMMON METALS. 39 14.24%. A part of the sulphur also was removed thereby. The re- fuse was high in ash, and low in fixed carbon, as was to be expected ; but the yield of washed coal was 85% of the raw screenings, arid the coke 70% of the washed coal. When the coal contains slate, 'bone', or pyrite, it is often improved by this process of washing, or separat- ing the waste-matter by concentrating. An example of a semi- bituminous southwestern coal is shown below : Raw scr Washed Coke 'eenings ng g innr. n f^ fryrf Ijft1" 7 rm^l nt n high temperature dag&aiflt4^ftbme^^ if in contatl vulli it in a fmaee. It is used both for basic-lined converters and for basic open- OF THE COMMON METALS. 55 hearth furnaces where the slag contains as little as 15% silica. It is used also as a lining for forehearths in copper matting where it would be in contact with low-grade corrosive matte. The nature of the mineral is shown by the following analysis : Per cent. CaO 1.68 MgO 42.43 SiO 2 0.92 Fe 2 6 3 and A1 2 O 3 4.30 CO 2 and H 2 50.41 99.74 Other refractory materials. A mixture of portland cement 2 parts, clay 1 part, and 'chamotte' or coarsely-ground brick 7 parts, moistened and molded into bricks or blocks, or used for patching furnaces, sets quickly and withstands a white heat without disinte- grating. It is easily made and especially useful for rapid repairs. Only as much is mixed as is to be used at once. While common red bricks are not refractory, the least fusible can be used in that part of the roof of a reverberatory furnace where the temperature is not high or only at a red heat. Such bricks are used for backing fire-brick structures. As a general rule each kind of brick should be laid in a material similar to that of which it is composed. We should expect slagging to take place, for example, at joints made of loam-mortar in fire-brick. Such loam, while cheap, is inferior to fire-clay. An analysis of good loam gives : Per cent. Si0 2 80.99 A1 2 3 9.65 Total impurities 4.91 Ignition loss 4.43 99.98 Here we note that the fluxing bases rise to nearly 5%, while alumina approaches 10%, the ratio of the most fusible compound of alumina and silica. Where the fluxing bases rise above 5%, there is risk of complete melting at high temperatures. 11. SAMPLING. The principles and object of sampling. Sampling consists in ob- taining from a large quantity of ore a small portion for assay. This 56 THE METALLURGY must correctly represent the entire quantity of the ore, whether it be a few hundred pounds or thousands of tons, a wagon-load, or a ship-load. Often we have a lot of ore, in which rich pieces mingle with poorer ones, or even with waste. In sampling we must take this variation into account and represent each part, not only accord- ing to its value, but also to its quantity. Often ore is bought or sold upon the results of sampling. Thousands of dollars are involved and cash is paid for ore before the purchaser has treated it. In other cases ores taken by the reduction works are treated separately, the owner receiving whatever is obtained, a charge being made to cover the cost of treatment and the profit to the reduction works. In this latter case sampling could be omitted. Similarly at a mill and mine, operated in one interest, the sampling may be omitted when consid- ered an unnecessary expense. Efficiency of the work is then de- termined by the assay of the tailing. If a reduction-works is producing lead, copper, or zinc, in a form ready for market, the metals do not necessarily require to be sampled. Whenever the precious metals are also present in such quantity as to pay to separate them, however, the metal is sampled to learn the values contained before selling to the refining-works that is to effect the separation. In blast-furnace treatment, ore and all other con- stituents of the charge are sampled, assayed, and analyzed. From the data thus obtained, the charge can be correctly calculated and proportioned. Not only is it necessary to ascertain the value of ores and of metals that result from metallurgical operations, but as well the value of the portions rejected. The efficiency of the work of the metallurgist depends upon thorough extraction from the parts thrown away. To be assured of this, samples of slag or tailing are taken at frequent intervals. In finding the value of a lot of ore, we first weigh the ore, and base the assay-value upon the dry weight. To do this we must determine the percentage of moisture contained, as shown by a 'moisture sample'. We then sample the ore regularly, and finally assay the regular sample. Thus, suppose we have a lot of ore weighing 10,800 lb., containing 1% moisture and by assay 54% lead worth 3c. per pound. Since the assay is made on the dry weight, we have, after deducting moisture, 10,044 lb. ore containing 5424 lb. lead worth, at 3c. per pound, $162.72. 12. ACTUAL SAMPLING OF ORE. Receiving and weighing. At reduction works that purchase ores (custom works), the ore arrives either loose, or in sacks. Whether OF THE COMMON METALS. 57 received by wagon or by car, the vehicle and ore are weighed together on platform scales, thus finding the 'gross weight'. When the vehicle is emptied, the weight, called the 'tare', is similarly taken. The difference is the 'net weight' or the 'wet weight', and this is recorded. When ore arrives in sacks, the weight of the sacks also is deducted. Often sacked ore may be removed to scales to be weighed, and only the weight of the sacks deducted, the difference being the net or wet weight. Sacks, if of sufficient value, are dried and returned to the owner. Railroads often return them without extra charge. Sometimes the sacked ore, if pulverulent, rich, or frozen, may be charged, sack and all, into the blast-furnace, the sack serving to retain the fine contents until smelted, thus preventing the loss of flue-dust. The moisture sample. In the theory the moisture sample should be taken at the instant of weighing, since the ore may dry and be- come lighter. The sample is taken while the car is being unloaded or immediately afterward. To represent by the sample the ore as contained in the car, holes are dug at average points (setting aside the dry top layer) and small portions are taken of ore that appears to be of average moisture. These portions are put in a covered can, and 50 oz. of the mixture are weighed on a moisture-scale. After cautiously drying on a hot-plate, or preferably over night on steam- coils, the 50-oz. portion is again weighed, and the percentage of moisture determined by the loss in weight. The shipper often sends a representative to witness the sampling of his ore. Such a man should pay attention to this detail, otherwise too high a percentage may be deducted for moisture. Sampling methods may be divided into two class : hand-sampling and machine or automatic-sampling. Any method of sampling in- cludes the starting and finishing operations. Hand-sampling. This includes the methods called 'grab sam- pling' and 'trench sampling', which are imperfect, and the regular methods known as 'coning and quartering', 'fractional selection', and sampling with the 'split shovel'. The grab sample. This imperfect method consists in taking at uniform distances over the pile, proportional amounts, broken from the lumps and taken from the fine. These portions are mixed, then reduced in size by a method of hand-sampling, later to be described. The imperf ectness of the method is due to the fact that only the upper part of the pile is represented in the sample. The method is used only as a quick and inexpensive means of obtaining an approxi- mate idea of the value or composition. 58 THE METALLURGY The trench sample. This is taken from ore in a pile, or at a dump, where an approximate idea of the whole is desired. A trench is dug transversely through the pile, selecting some place for this that is fairly representative of the whole. As the workman advances with the cut he throws the larger part or 'rejected portion' aside, but re- serves an aliquot portion. Thus each tenth, twentieth, or hundredth shovelful is reserved for sample, to be further reduced in size in the usual manner. The lower part of the pile by this means is repre- sented. In place of digging a trench, another method consists in dig- ging trial pits at regular intervals over the pile or dump and reserv- ing an aliquot portion of the excavated material. The small reserved portion of the whole is then subjected to regular sampling methods. The trench-method and the trial-pit method are imperfect, since they represent only a part of the pile. The regular or complete methods of hand-sampling, including coning and quartering, fractional selection, and the split-shovel method, are used where an accurate sample that could be used with confidence in buying or selling is desired. Of these methods coning and quartering is the oldest. Coning and quartering. Before applying this method, and to save unnecessary labor, it is usual to take an aliquot portion of the ore, say every tenth shovelful, as the car or wagon is being unloaded. The sample from a 100-ton lot would then be 10 tons. It is wheeled to a crusher, crushed roughly to l^-hi- size, and as it falls from the crusher, wheeled to a place on the sampling floor and dumped in the form of a circle or ring, leaving a space of 10 ft. diam. inside. The workmen circle around this ring, shoveling the ore to the apex of a cone which they build at the center. Care is taken that each shovel- ful is thrown upon the apex of this cone. The ore is then drawn down with shovels, to the form of a flat disk 8 to 12 in. deep. This is marked by diametrical lines into four equal sections (hence the word quartering) of which two are left on the floor and two wheeled away. The reserved sectors are shoveled again into a ring and made into a cone thus half the size of the first. This is again flattened down, quartered, and two opposite quarters reserved. The process goes on in this way, by coning and quartering the ore, as long as any single rich piece can not appreciably raise the value of the sample if retained, or lower it if rejected. At this point the ore is crushed to half-inch size by means of rolls. Coning and quartering is resumed until it becomes necessary to crush again, this time, say, to wheat size. The ore is then worked down to a sample two pounds in weight, and this quantity is ground to pass through an 80-mesh screen. It is OF THE COMMON METALS. 59 thoroughly mixed by 'rolling' on a sheet of thin rubber cloth; and the mixed product is placed in 4-oz. bottles or manila sample-sacks. These are sealed and marked with name and particulars of the lot of ore. Fractional selection. This differs from the quartering method in that every second or fourth shovelful is reserved for a sample after coarse crushing, but the selected portion is coned for the purpose of mixing. From the cone each second or fourth shovelful is again re- served and coned, and this is continued until it becomes necessary to re-crush. After this, the reduction proceeds in the same way except that a smaller shovel is used toward the end, to accord In capacity with the size of the sample. The split-shovel. By this method a shovel, resembling a fork, is used, the tines of which consist of troughs, each 12 in. long and 2 in. wide. The space between the parallel troughs is 2 in. wide. Operat- ing on ore crushed to half -inch diameter, one half of the ore shoveled upon the sampler drops through the spaces, and half is set aside as a sample. The sample is again divided by the split-shovel, reducing the size again one-half, and so on until, as in other methods, it is reduced to the desired quantity. For finishing a sample a riffle sampler is often used. This consists of a series of parallel troughs like a gridiron, the width of the troughs being at least four times the diameter of the largest pieces of ore that are being handled by means of it. As with the split-shovel, one- half the ore is retained in the troughs, and one-half falls through and is rejected. Each shovelful is evenly scattered upon the riffles, and care is taken not to heap the ore above the troughs. Machine or automatic sampling. It will be seen that the different methods of hand-sampling, especially for large lots of ore by coning and quartering, involve much labor; and it had been sought to over- come this by the use of machinery. A sample, taken from a stream of ore, is called a running sample, and is taken automatically by a sampling-machine. These machines are of two kinds. (1) Those which take part of the stream all the time. (2) Those which take the whole stream at frequent and regular intervals. Since the stream of ore is not homogeneous, the first method is defective and the second generally preferred. Fig. 15, called the pipe-sampler, is an example of the first type. The ore stream is delivered into the hopper at the top, and is split by deflectors, which eject half the stream at the sides of the tube. 60 THE METALLURGY The half retained is caught by the next deflectors below, that repeat the operation, and the division continues as the ore descends. The sampler here shown reduces the quantity of ore to one-sixteenth the original amount. Fig. 15. PIPE ORE-SAMPLER. Fig. 16. VEZIN AUTOMATIC SAMPLER. The Vezin sampler (Fig. 16) is a sampler of the second kind. It consists of a tube carried by a vertical shaft making 30 revolutions per minute. Attached to the side of the tube and opening into it is a scoop. As the shaft revolves ' counter-clock-wise ', it cuts the stream OF THE COMMON METALS. 61 of ore that is falling from the inclined feed-chute. The ore^ thus intercepted, falls through the tube and becomes the sample, while the rejected portion, falling in the main hopper, is delivered by a pipe to Fig. 17. SAMPLING WORKS (PLAN). Fig. 18. SAMPLING WORKS (ELEVATION). 62 THE METALLURGY the bin. In the plan of a sampling works (Fig. 17) three automatic samplers are shown, the two first of which have double scoops. The ore is crushed finer after passing each sampler, and by the time three successive cuts have been made the sample is reduced to one two- hundred-and-fiftieth of the original weight. Fig. 17 and 18 represent in plan and elevation a small sampling mill of a capacity of 10 tons hourly in use at a reduction works. The process in this case is as follows : A car, after weighing, is unloaded into the storage bins of the works, and a sample consisting of one-tenth of the entire carload is retained in the car and the car then sent to the mill. The sample is held upon the floor until similar samples from all the cars containing ore of the same lot are added to it. This sample may weigh 10 tons. It is now put through a Blake crusher, and reduced to l^-in. size. The ore from the crusher then is elevated and passed through a Vezin sampler, while the rejected portion goes to one of the sample- bins. The portion to become the sample, now weighing 4000 Ib. or one-fifth the original weight, is sent to a Dodge crusher to be crushed to %-in. size. It is elevated to a second Vezin sampler, where one- fifth is again cut out, reducing the amount to 800 Ib. It next passes to large rolls that reduce it to pea-size, then by an elevator to the last Vezin sampler, where it is cut to 80 Ib. It is now received upon the floor to be coned and quartered, and when reduced to 20 Ib. in this way, is put through small rolls that crush it to 20-mesh. This is again reduced in quantity to 5 Ib., put through a sample-grinder, mixed, quartered to 2 Ib., ground by hand on a bucking-plate to 80- mesh size, mixed, and put in bottles or sample-sacks as in hand- sampling. The rejected portions of the ore are conveyed by gravity to any bin desired, and retained there until the ore has become the property of the reduction works by purchase. Weight of ore. Value in silver ounces per ton. Highest 300 Average 50 Highest 3000 Average 75 Highest 10,000 Average 500 100 tons to 10 tons Cocoanut Orange Walnut Pea 20-mesh 80-mesh Fist Egg Chestnut Wheat 25-mesh 100-mesh Fist Walnut Chestnut Wheat 50-mesh 120-mesh 10 tons to 1 ton 1 ton to 200 Ib 200 Ib. to 5 Ib 5 Ib. to bottle sample Bottle sample . In the progressive crushing, as described, it is observed that we OF THE COMMON" METALS. 63 make the ore finer as the sample becomes smaller. This is to make sure of a constant ratio between the size of a single rich piece and the whole sample, that it may not produce an appreciable effect upon the assay-value of the sample whether such a piece be present or absent. The richer, and more 'spotty', or varied the ore, the finer it should be crushed therefore before it is cut or quartered. The table above shows how this is arranged in practice. We may conclude that for accurate sampling the principal re- quirements are : (1) The taking of frequent portions to insure an average of the whole stream that is undergoing progressive sampling. (2) Thorough mixing of the ore to insure uniform richness. Sampling of ores containing metallic substances. This is an operation requiring a clear knowledge of the principles of sampling. We come upon these 'metallics' sometimes in the operation of sam- pling. They must be separated, cut smaller, and quartered down separately by a hand-method, and reduced in size, at the same rate as the fine ore. If fine substance is made by cutting up the metallics it can be united with fine ore. Often metallics are brittle, but with diligent work can be broken, cut, and 'quartered down' without serious difficulty. Cost of sampling. The cost of moving the ore cars, unloading into bins, returning the cars to the sampling-mill, and unloading the fractional part, usually one-tenth, retained, may be taken at lOc. per ton. The cost for hand-sampling the tenth part may be taken at 75c. per ton. Hence, for unloading and hand-sampling a 100-ton lot, the total cost would be 17.5c. per ton. At the Metallic Extraction Works, Cyanide, Colorado, ore was unloaded from the car to a feed- shoot, crushed to %-in. size, automatically sampled, and delivered to storage-bins, for lie. per ton. A charge of $1 to $2 per ton has been made for sampling, storing, assaying, and selling ore at cus- tom works or sampling-mills, where the company has acted as sell- ing agent and obtained the best possible price for the shipper. The price for sampling concentrate was 50c. per ton lower. Sampling concentrate tailing, and ore-pulps. Concentrate is sampled easily for it can be thoroughly mixed and sampled by hand. Tailing contains but little value, and needs no close attention. Ore- pulp flowing in a launder is often automatically sampled. When not so sampled, a bucketful, taken each hour from the stream, and all these samples united in one portion after decanting the water, may be used to determine the approximate daily average value. 64 THE METALLURGY 13. SAMPLING METALS. Metals may be sampled either in the solid or molten state. Gold or silver bars or ingots. These are sampled for assay either by granulating a small portion of them or by taking chip-samples from them. In the first case, while the metal is in a molten condi- tion, a small ladelful, weighing an ounce or less, is taken from the crucible immediately after stirring it. This is poured into a bucket- ful of water thereby granulating the metal and forming particles of a variety of sizes, convenient for weighing and assay.. Chip-samples are taken at points diagonally opposite on the edges of the bar, and a cold-chisel, cutting out a small wedge-shaped piece, is used for this purpose. The pieces are annealed and rolled into a ribbon for assay. The average assay-value of the two pieces thus obtained is taken as the true value. Base-bullion. This is lead that comes from silver-lead blast-fur- naces, and it contains commonly 100 to 400 oz. of silver per ton. When poured into molds to be cast in bars, the silver segregates, and the exterior of the bar, that cools first, is richer in silver by several 2602 Z6O.6 264.1 269.0 \ 260.9 \ 264.6 250.0 249.0 249.0 264.6 1 \~ 263.5 2570 24ED 268.0 1 263.6 / \ 26/.O 259.6 26/.6 / Avers ye 258.2oz. stiver per Ton Fig. 19. DISTRIBUTION OF SILVER IN BAR OF BASE-BULLION. ounces, than the central part. This is illustrated in the cross-section of a bar (Fig. 19), in which the center of the bar assays 10 ounces less than the exterior. Base-bullion is sometimes sampled by taking two 'chips' or punchings, one from the top and one from the bottom of each bar. The punch (Fig. 20) , resembling a belt punch, is 8 in. long and removes a cylindrical piece of Vs-in. diam. by about 1%-in. length. From a carload lot of 400 bars, 800 of these chips would be ob- OF THE COMMON METALS. tained. These are melted and the fused metal stirred in a plumbago crucible and cast into a bar. This is a sample of a 400-bar lot of base-bullion, equivalent to 20 tons. A better way of sampling, how- ever, is to re-melt the metal in a large kettle (See Fig. 172), and to Fig. 20. BASE-BULLION SAMPLING PUNCH. skim and re-cast into bars for shipment. While casting the metal, a sample is taken from the molten bath and poured into a bullet- mold of such a size that each bullet weighs approximately a half assay-ton. This is trimmed to the exact weight for the assay. , O.34\ 1 S00. 7\ 36. 6 O.24\ O.22\ O.22 \/3Z./\/3S.2\/34.&/22.0\ 69. / O.26 O.2O . / \ 672 \0.22a230.30\0.260.20 \ "7/.3\ 7BJJ 6S.8\ 6&B Fig. 21. SECTION OF BAR OF INGOT COPPER. Copper ingots or anodes. The segregation of gold and silver in copper ingots is even more marked than in bars of base-bullion. This is shown in Fig. 21, which represents the distribution of gold and silver in an ingot of blister-copper 5 in. deep. In this case, how- ever, the interior is higher both in silver and gold. The usual way to sample such bars is to drill into them and retain the borings for a sample. Manifestly a sample like this is uncertain, and depends upon the selection of the place on the ingot for taking it. To obviate this difficulty, in sampling a lot, say of 100 bars, it is customary to drill into each succeeding bar at a different spot to obtain an average by so doing. It is preferable, however, to take the sample at the time the copper is melted and well mixed in the furnace by poling. As in the case of base-bullion, samples of copper are taken while 66 THE METALLURGY dipping or casting, one at the beginning, one when the charge is half removed, and one toward the end. The average of the three samples is regarded a correct representation. Pig iron. This is sampled and graded by inspecting its fracture and by chemical analysis. When an analysis is to be made, the sample is taken from the drillings of a small bar, molded while the metal is flowing from the furnace. The percentage of silicon deter- mines the grade. 14. CRUSHING AND GRINDING. The method and extent of breaking, crushing, grinding, or pul- verizing ore depends upon the size best suited to subsequent metal- lurgical treatment. Ore comes from the mine already of a size that may need no further breaking if intended for smelting. Or it may come to the reduction-works as concentrate, fine enough for roast- ing or other treatment. Methods of breaking or grinding ore for further treatment, may be classified as follows : (1) Breaking for the blast-furnace. (2) Breaking for stall or heap-roasting. (3) Crushing for reverberatory melting or smelting. (4) Crushing for treatment in roasting-furnaces. (5) Crushing for distillation. (6) Crushing or grinding for amalgamation. (7) Crushing or grinding for leaching. (8) Grinding for sliming. (1) Breaking for the blast-furnace. The blast ascending through the charge carries away fine ore by the strong current and for this reason, and also for the reason that coarse ore makes an open charge that permits the free passage of the blast, coarse ore is preferred. It is desirable, indeed, to smelt ore as coarse as it comes from the mine, breaking only the large lumps with a hammer. When the ore is an oxidized one, and lumpy, the whole of it may be passed through a rock-breaker set for 2y 2 -m. crushing. This is desirable only when the blast-furnace charge is too open, and when reduction, in consequence, is poor. Commonly the case is the reverse of this because of the large amount of fine ore which it is necessary to smelt. With a coarse charge it is desirable to crush also the fluxes, includ- ing iron-ore and limestone. When the charge is fine, or tight, the limestone may be fed in 50-lb. lumps, which though large pieces, become disintegrated in their downward passage by the heat that drives off the CO 2 . Up to this time the lumps do duty in keeping OF THE COMMON METALS. 67 the charge open. To prevent the blast carrying away the finely pul- verized ore, it is a good plan to wet it before feeding. The more effective way is to press the fine material into bricks, using a briquet- ting press (Fig. 164). The bricks become hard on drying and are unaffected by the blast. Briquetting adds about $1 to the expense of treatment. (2) Breaking for stall or heap-roasting. This work may be done either by hand or in a rock-breaker. By the latter method the work is less costly but more fine material is made than by the former. Peters, in his 'Modern Copper Smelting' cites a case of breaking in a jaw-crusher in which the fine resulting amounted to 17.3%, while by spalling or. breaking with hammers, only 9% was made. On the other hand, machine-crushing costs but 9c. per ton, while hand-spalling costs 35c. Peters states that, since 10% fine is suffi- cient for a finishing-covering for roast-heaps or stalls (See Fig. 27 and 28), any excess above this quantity should be avoided. This rule does not apply to fine material roasted separately in a reverberatory furnace (See Fig. 33). (3) Crushing for reverberatory melting or smelting. Fig 138 and 139 are views in elevation and plan of a furnace in which ore is melted or slagged to a liquid. Pieces as large as an egg melt easily in such a furnace. One great advantage that the reverberatory-fur- nace has over the blast-furnace is, that the damper of the furnace may be closed at the time of charging until the dust, arising from the charge when dropped into the furnace, subsides. Thus no appre- ciable loss of fine dusty ore results, whereas in a blast-furnace such a light material would be blown away. (4) Crushing for treatment in roast ing-furnaces. Illustrations and descriptions of furnaces of this kind are given in the chapter on oxidizing-roasting. The ore must be crushed fine so that air can reach it readily to properly roast it. An ore consisting mainly of iron pyrite decrepitates in roasting, so that it is fine enough if crushed to pass a 2 to 3-mesh screen. Many ores and mattes need crush- ing to 4 to 6-mesh size. An ore containing blende or galena is com- pact, and when it must be roasted, especially if to be dead-roasted (that is, until the sulphur is eliminated), had better be ground finer, or to 10-mesh size. When an ore is to be subsequently treated by leaching, for which it should be more finely ground, the grinding can be done before roasting. It is a good plan, however, to grind to a coarse size for roasting and re-grind the product as fine as de- sired for the after-treatment. Ore should be ground no finer than needed for efficient roasting, since more dust is made by so doing. 68 THE METALLURGY An idea of the efficiency of the roast, and the necessary fineness of the particles can be obtained by examining the latter after the roast is finished. If imperfectly roasted, the particles have an un- roasted or raw core or center. In the Stetefeldt furnace, the ore is ground to 20 mesh, since the time of roasting is confined mainly to the few seconds occupied by the fall of the ore particles from the top of the tower of the furnace to the bottom. (5) Crushing for distillation. The purpose of this is to prepare ore for the distillation of zinc or mercury. When zinc ore has been crushed to 8 to 14-mesh size it is fine enough for the preliminary roast and for mixing with coal in the subsequent treatment. Cinna- bar ore is fine enough when 1 in. or less in size. It is generally di- vided into two or more sizes according to fineness, each size being separately treated in the retorts. (6) Crushing or grinding for amalgamation. The ore is milled or crushed in a stamp-battery (See Fig. 44 and 45). When com- minuted to pass the battery-screen, a good deal of it is very fine, or 'slimed' as it is called. Slime is not objectionable when ore is passed over the usual amalgamated plate to collect the gold it con- tains. In silver milling, ore is first crushed in a stamp-battery, then re-ground in pans, thus further comminuting the coarser particles that pass through the battery-screens. Grinding in the pans is con- tinued until the pulp is of 200-mesh size or finer, and feels smooth and free from grit when rubbed between the fingers. (7) Crushing or grinding for leaching. Ore that is to be treat- ed by a leaching method should be ground so fine that the solution has access to the metal in the particles of ore. It must not be ground so fine as to retard the leaching by the slimed material intermingled with the sand. In the operation of leaching we are guided by the activity or nature of the leaching solution. The solution may contain chlorine, bromine, cyanide of potassium, hyposulphite of soda, com- mon salt, or consist merely of water. With an active agent like chlorine the size may be less than with potassium cyanide. Again, ore may be porous, or the precious metal may be in a more soluble condition, or may be rendered so by a preliminary roasting. Thus Cripple Creek ore, crushed to 10-mesh for chlorination, is crushed to 30-mesh size for cyaniding. A point to be observed in preparing ore for leaching is to avoid making slime in crushing. Any considerable portion of finely ground or slimed material hinders percolation greatly. If grinding is performed by rolls a more granular product, containing less slime, is produced than by crushing to the same screen size with stamps. This is shown by the following screen an- OF THE COMMON METALS. 69 alysis. The two products indicated have passed through screens of the sizes specified, and the percentage of the various fine sizes con- tained are represented by weight. TABLE SHOWING THE UNDER-SIZES RESULTING FROM CRUSHING. Wet stamps. Rolls. Crushed through Crushed through 26-mesh screen. 23-mesh screen. Screen sizes. Per cent. Per cent. Through 30, on 40 mesh 11.15 9.30 Through 40, on 60 mesh 28.53 41.85 Through 60, on 90 mesh ' 9.21 15.38 Through 90 51.11 33.47 100.00 100.00 The ore used here is a conglomerate containing a little pyrite. The smaller quantity of slime produced by the rolls should be noticed. If it is desired to obtain the maximum quantity of sand and the minimum of slime, then gradual reduction, or graded crushing, should be adopted. This consists in screening out the fine ore, coarse- ly crushing the residue, screening out its fine and then crushing what is left. By this means the ore, as soon as crushed to the de- sired size, is removed from further action of the crushing machinery and escapes unnecessarily fine comminution. An illustration of how this is done to ore that is to be ground sufficiently fine for roasting and subsequent leaching, is the follow- ing system of dry-crushing. The assumption here made is that the mill is to crush 200 tons per day. The operation is divided into (A) coarse and (B) fine crushing. A. Coarse crushing. A 20 by 12-in. Blake rock-breaker crushes ore as it comes from the mine containing pieces as large as 12 in. diam. at the rate of 25 tons per hour. There are, however, clayey and talcose wet ores containing 25 to 30% moisture that stick to the rock-breaker, and are impossible to crush in the wet state. Such ore is first dried in a cylinderical dryer. The Blake rock-crusher is shown in perspective in Fig. 22 and in longitudinal section in Fig. 23. It consists of a heavy cast-iron frame, marked 1, within which is placed the fixed jaw 5, and the swinging jaw 2, and between them the ore is crushed. A shaft 36, eccentric where it passes through the pitman 3, causes this to rise and fall, producing a corresponding movement of the adjacent ends of the toggles 7, 7. As these rise, the effect is to push the jaw forward to produce the crushing move- ment. As the pitman and the toggles descend the jaw recedes, and 70 THE METALLURGY is pulled back by the spring rod 16, and the rubber spring; 17. Fly- wheels 11, help and steady the movement. The machine is driven by the pulley 12, at 250 revolutions per minute. The movement of the lower end of the jaw is y to % in. For the breaker above speci- fied, the discharge-opening would be 20 by l 1 /^ in- to crush to IVo-in. size. The receiving opening would be 20 by 12 in., and would take pieces as large as 12-in. diameter. The ore, now crushed to l^-in. size or finer, goes to rolls 36-in. diam. by 16-in. face, which supply 25 tons per hour to a revolving Fig. 22. THE BLAKE ORE-CRUSHER. screen or trommel having %-in. holes. A set, or pair, of such rolls of the belt-driven type is shown in Fig. 24. One roll is carried by its shaft in fixed bearings or boxes, and is driven by the large pulley. The other roll, called the movable one, is held to its work by power- ful spiral springs acting on the shaft-boxes with a pressure of 15 tons upon each box. Thus the pressure of the springs is such that if a hard object, such as a hammer-head, were to fall between the rolls, they would open only under a pressure of 30 tons. The smaller pulley on the shaft of the movable roll is intended to keep the roll in motion at the speed of the other, but not for other work, the OF THE COMMON METALS. 71 power being mostly transmitted through the large pulley. The rolls (in Fig. 24), are covered by the housing. The feed-hopper of the rolls is seen immediately above the housing. As the ore drops from the rolls it falls through a chute to the boot of a belt or chain-eleva- tor (See Fig. 189 and 191). The material from the rolls is delivered or discharged through a chute into a Vezin sampling-machine (Fig. 16), making 40 revolu- tions per minute. This takes out one-fifth for a sample, and leaves four-fifths to go to the storage-bins. The sampling is completed by Fig. 23. BLAKE ORE-CRUSHER (SECTION). the methods described under 'machine-sampling'. A single samp- ling-machine can easily handle the 200 tons crushed during the 10- hour day shift. The ore % to % in- and less in size, is drawn as desired from the storage-bins into two-wheeled buggies, and dumped into the hopper of the cylindrical drier. A feeding-shoe or some other kind of au- tomatic feeder, supplies it continuously to the drier, shown also at /, Fig. 110 and 111. This drier is 24 ft. long and of sufficient capacity to dry ore containing 6% moisture to \% or less, at the rate of 10 to 15 tons per hour, heating it at the same time so that it will screen 72 THE METALLURGY readily. When in this condition the ore is 'lively', and will screen without difficulty, but if it were not for the drying, it would stick to the screen and clog it. The cost of this drying approximates 5c. per ton. B. Fine crushing. This work is done continuously. Fig. 25 is a diagram showing the series-crushing system used. The ore-supply from the drier goes to the roughing-rolls a which reduce it from 0.75 to 0.25 inch. The crushed ore is raised by the elevator to a trom- mel, or separating-screen, having screens of Vs and %-in. aperture, respectively. The first two-thirds of the screen takes out all ma- terial less than % in., and the final size is all coarser than y 2 inch. Fig. 24. CRUSHING ROLLS. We thus get three products, an oversize from the coarser screen which goes back to the roughing rolls to be re-crushed ; a screened product or under-size, which goes to the medium rolls &, set at % i n -> there to be crushed and sent back to the separating-screen ; and fin- ally, an under-size through the %-in. screen, fine enough to go to the finishing rolls. Until crushed so fine that it passes the finest mesh, the ore is returned to the trommel. The fine product of the screen is raised by the elevator /' and the ore-stream is equally divided between the finishing screens s" and s"' which are provided with 30-mesh wire cloth. The undersize from these revolving screens or trommels drops into the storage bin m, while the oversize is conveyed to the finishing rolls, after which it goes again by elevator to the finishing screens. Thus nothing en- OF THE COMMON METALS. 73 ters the storage-bin except 30-mesh or finely ground ore (0.02-in. diam.), ready for further treatment. We observe that any 8-mesh or finer product, that has been pro- duced by roughing rolls, does not go to the medium rolls, and that any finished sand, in the 8-mesh product, is not forced upon the Fig. 25-. FLOW-SHEET FOR DRY CRUSHING. finishing rolls, but goes first to the finishing screens. Thus the mo- ment a particle is broken to a finished condition, it passes to the finished-product bin without passing through the crushing machin- ery. One square foot of screen will separate 6 cu. ft. of product per 24 hours. The 200 tons required can be treated in 20 to 22 hours, allowing time for delays and repairs. This gives a capacity for each 74 THE METALLURGY set of finishing rolls of 5 tons per hour. The speed of the rolls varies with the coarseness of the material to be crushed. The coarse rolls, run at 600-ft. the fine ones at 1000 to 1400-ft. peripheral speed, per minute. This system of graded crushing is preferable because the pro- duct contains a minimum of fine or slimed ore ; and being granular, is more easily percolated or leached. As each piece or particle of ore is crushed by a single nip, the fine is separated by the screen and protected from unnecessary breaking with consequent waste of power. The chief costs in this system of dry-crushing, are those of labor, power, supplies, and repairs. These vary with the tonnage. Fig. 26. TROMMEL OR CYLINDRICAL REVOLVING SCREEN. The cost of crushing to 30-mesh size, in preparation for roastirg ;>L* leaching, is as follows : For coarse-crushing and automatic sampling. . .$0.106 For drying and fine-crushing 0.275 For power 0.105 $0.486 In round numbers the cost is 50c. per ton, but to this must be added general expense, including management, office expense, rates, taxes, insurance, cost of water, and improvements. Besides crushing with rolls, after coarse-crushing, ore may be crushed by stamps or Chilean mills. Fig. 44 is a stamp-battery, a favorite and efficient machine; Fig. 100 is a Chilean mill. This will be described later. (8) Grinding for sliming. Ore frequently contains gold and especially silver so finely desseminated in the substance of the ore that giinding finer than 150-mesh is needed to unlock the minerals from the adhering gangue. While rolls or stamps grind ore to a OF THK COMMON METALS. 75 certain fineness, further grinding, after water-separation or classi- fication of sand and slime, is performed upon the granular or sandy portion too coarse to pass uncrushed. Grinding-pans and tube-mills have been used for this purpose, and have proved most satisfactory. The tube-mill (Fig. 96), later to be described, is a favorite means employed for this purpose because of its great capacity, simplicity, and cheapness in repairs. The cylinder or tube revolves 20 times per minute, and half-filled with pebbles the size of the fist rolling over one another, effectually comminutes the ore particles. Grind- ing to any desired size, is done, and the coarser particles which escape grinding, are classified and returned to the machine for re- grinding. PART II. ROASTING PART II. ROASTING. 15. OXIDIZING-ROASTING. This operation is also called * calcining', though the term calcin- ing is applied also to heating carbonates to expel carbon dioxide. We shall proceed to the chemistry of the process. 16. THE CHEMISTRY OF ROASTING. The operation of roasting is for the purpose of burning or expell- ing the sulphur, which the ore contains, by the action of heat with the access of air. This can be done either when ore is in lump-form, in heaps, or when fine, by means of reverberatory furnaces. The various types of these furnaces are to be described and illustrated in the present chapter. For the purpose of illustration, let us consider the action that takes place in a reverberatory furnace such as is shown in Fig. 31, 32, and 33. In this furnace the fire is at the end, and the charge is spread upon the hearth. As roasting proceeds, the roasted ore at the fire-end is withdrawn and the part remaining unroasted is moved from time to time toward the fire, fresh ore being constantly sup- plied at the upper end. After 20 hours, ore fed at the cool end, is in the hottest part of the furnace. Let us take for example an ore with a silicious gangue, containing mixed sulphides such as galena, blende, pyrite, and chalcopyrite. The silica tends to prevent sintering and makes the charge more open or accessible to the air. To do good roasting, certain requirements must be observed. First, the heat must be sufficient to start the roasting. Later, the temperature must be high to drive off the last of the sulphur. Air must be abundant, to oxidize the ore freely, and to carry away readily the products of the re-action. The surface exposed to the heat and air should be ex- tensive. Finally, the ore should be stirred frequently, to bring fresh surfaces into contact with the air. The ore is dropped upon the hearth at the cool end of the furnace, where the temperature (350C.) is sufficient to expel moisture and start the reaction of combustion. In 10 to 15 minutes the charge 80 THE METALLURGY becomes hot enough for the oxidation of pyrite to begin, as evinced by the blue flickering flame that plays over the surface of the charge. Pyrite, under the action of heat, separates thus : (1) FeS 2 + heat + 20 = FeS + S0 2 . The sulphur expelled combines with oxygen from the air. An equiva- lent, or 32 Ib. of sulphur, burning to S0 2 , yields 72,000 calories, or 72,000 -r- 32 = 3220 pound-calories per pound of sulphur burned. The S0 2 is removed by the draft. The remaining FeS and also CuS, the ZnS, and PbS, begin to oxidize, the activity of roasting being in the order here named. While oxidation proceeds with them all, the FeS is most easily roasted, while the ZnS and PbS are the slowest in parting with sulphur. Beginning, therefore, with the FeS we have : (2) FeS + 30 = FeO + SO 2 23,800 66,400 71,000 = 113,600; or in words the iron sulphide becomes oxidized to ferrous oxide, with the evolution of S0 2 , the reaction being accompanied by heat to the extent of 113,600 -r- 32 = 3550 pound-calories per pound sulphur burned, as explained in Part I, Section 3, under ' Thermo-chemistry '. The cupric sulphide of the chalcopyrite acts according to the for- mula : (3) CuS + 30 = CuO + S0 2 10,200 37,200 71,000 = + 98,000 ; or, per pound sulphur, 98,000 -f- 32 = 3030 calories. The blende, un- der the action of the heat and air, is affected in the same way : (4) z n s -f 30 = ZnO + S0 2 43,000 86,400 71,000 = + 104,200 ; or, per pound sulphur present, 3420 calories. Galena also roasts ac- cording to the reaction : (5) PbS + 30 = PbO + S0 2 17,800 51,000 71,000 = + 104,200 ; which gives us 3250 calories per pound sulphur. It will be noticed that the heat evolved per pound sulphur burned is much the same in each case ; and hence the sulphide, containing most sulphur, gives off the most heat in roasting. These re-actions, especially of blende and galena, proceed gradually throughout the roasting operation. The air acts chiefly on the exposed surface, and the activity of the above reactions is accordingly increased by stirring the charge. The FeS that has been formed, as shown in reaction (1), and which lies upon the surface, is exposed to an excess of air. The FeS, in presence of the silica of the ore, acts by catalysis and is a sort of go-between for OF THE COMMON METALS. 81 the FeS 2 and the oxygen of the air. It finally becomes oxidized thus : (6) 3FeS -f HO = 2SO 2 + Fe,0 3 + FeS0 4 3x23,800 2x71,000 199,400 235,600 = + 505.600 ; or, per pound sulphur, 5260 calories. This is observed to be a most energetic exothermic reaction. The products of the reaction are iron sulphate, SO 2 which is carried away by the draft, and Fe 2 O 3 which is acted upon by more FeS when the two are stirred together as follows : (7) FeS + 10Fe 2 3 = 7Fe 3 4 + S0 2 23,800 10x199,400 7x265,800 71,000 = 86,200; If the charge is not stirred, reaction (6) causes a red-colored pro- duct to form on the surface of the charge. This is the compound, Fe 2 O 3 , but on stirring it decomposes into Fe 3 O 4 as shown in equation (7), and the product becomes black. When roasted to excess, the product takes on this red color, undesirable in some roasting opera- tions. As the ore is moved to a hotter part of the furnace, the activity of the above reactions continue, and at an incipient red. (590C.), iron sulphate, which has been formed in the^mfmnfSr tfoffAn/egins to decompose, reacting on the cupric oxide as follows. Thus, as the SO 3 is given off frojnr^the FeS0 4 , and while in afn as- cent condition, it is taken by the CuO, the reaction being an endo- thermic one. Sulphate 'Li funiiGd Hum* (9) FeS0 4 + 2CuO = FeO + CuSO 4 235,600 37,200 66,400 182,600= -23,800. This reaction is t-n un mui'ivendothermic iLan lliu fui'moj,, and hence is only a partial one. There remains uncombined some iron sulphate, which is decomposed by the heat thus : (10) FeSO 4 = FeO + SO 3 235,600 66,400 91,800= -77,400. This also is an endothermic reaction. At a slightly greater heat (655C.), the cupric sulphate, formed but a short time previous according to reaction (8), begins to de- compose, and at a dull red heat (705C.), the corresponding decom- position of the cupro-cupric sulphate begins. These decomposition- reactions of the copper sulphate are complete at 850C., or at a cherry-red heat. To this point, they indicate the action that is going 82 THE METALLURGY on in a sulphatizing roast like that of the Ziervogel process to be later described. At 850C. the zinc and lead oxides, reacting on the copper sul- phate that is being decomposed by the heat, begin to be changed, to sulphates. (11) ZnO + CuS0 4 = ZnS0 4 + CuO 86,400 182,600 230,000 37,200= -1800. (12) PbO + CuS0 4 = PbS0 4 + CuO 51,000 132,600 216,200 37,200= -19,800. In general we may note that reactions (8) to (12), inclusive, are eri- dothermic, and in place of helping along the roasting by heat de- veloped, as shown in the early action of the charge, demand heat and absorb it as the result of the reactions. Fortunately these later re- actions take place only at the hot end of the furnace. As the charge is moved nearer the fire the sulphates completely decompose, the zinc sulphate doing so more readily than the lead sulphate. At 10500. (a dark-orange heat), copper oxide is decom- posed to cuprous oxide, and ferric oxide to Fe 3 4 , oxygen escaping at this high temperature. At this stage the ore begins to fuse, if it contains lead, but when little lead is present, it only slightly agglomerates. If the ore con- tained much leadj even this temperature can not be attained without causing it to begin to slag. In that case the charge, rendered no longer porous or accessible to the air, necessarily ceases to roast, zinc and lead sulphates decompose imperfectly, and the sulphur is not well eliminated. It is hard to roast a leady ore well. On the other hand a zinc ore, free from lead, can and should be brought to a high finishing-heat when it is desired to decompose the zinc sulphate and remove the last portions of the sulphur. Sometimes, with lead-bearing zinciferous ores, to be treated in the silver-lead blast-furnace, after roasting, the roaster is arranged with a fuse-box as shown in Fig. 33. The charge, which has been roasted on the roasting-hearth of the main part of the furnace, is moved from the hearth into the fuse-box, and melted after adding some sil- icious ore. The silica reacts on the zinc and lead sulphates thus: (13) ZnS0 4 + Si0 2 = ZnSiO 3 + SO 3 and (14) PbSO 4 + Si0 2 = PbSi0 3 + S0 3 The sulphur is eliminated as sulphuric anhydride fume, and the re- sulting product is freed from sulphur. Zinc silicate enters slag as such, and is eliminated by that means. Lead silicate is reduced in the blast-furnace with the recovery of the lead. The charge, originally OF THE COMMON METALS. 83 pulverous, is now in the form of a slag, so that no loss of flue-dust results. It has been found that with 2% S0 2 by volume, or 4.4% by weight, in the escaping gas, roasting is active. This corresponds to 23 Ib. air per pound S0 2 or 46 Ib. (570 cu. ft.) air per pound sulphur driven off. Calculating this for a 16-ft. McDougall roasting-furnace, treating 40 tons of ore per 24 hours, and roasting it from 35% down to 7% sulphur in the product, we have an elimination of approxi- mately i/4 Ib. of sulphur per second, which needs 142 cu. ft. free air, equal to 284 cu. ft. of escaping gas at 273 C., a temperature which gives a maximum chimney-discharge as explained in the chapter on combustion. For a velocity in the stack and flues of 20 ft. per sec- ond as a maximum, this would require an area of 14.2 sq. ft., or a di- ameter, for a round stack, of 4 feet 3 inches. When more than 46 cu. ft. of air per pound of sulphur is ad- mitted, though roasting proceeds actively, owing to a good supply of fresh air and to the fact that the products of ihe reactions are re- moved promptly, still the furnace becomes cooled too much. If less air than this is admitted, roasting proceeds more slowly so that with 4% S0 2 by volume the roasting is slow, with 8% very slow, and with 9% it ceases entirely. The various reactions described above need time ; and the larger the body of ore, the longer the time must be to complete the roast. If a few grams of ore are roasted in the muffle, the operation is com- plete in half an hour, but in a hand-reverberatory furnace (like that shown in Fig. 32), roasting 14 tons per day, the operation takes 20 hours. The temperatures at which the reactions described above take place are as follows : At 150 C. the odor, due to the volatilization of some of the first or loosely-held sulphur in pyrite, can be detected. At 350C. the sulphur of the sulphides, particularly of the iron sulphide, begin to burn. At 590 C. the iron sulphate, formed at a lower temperature, be- gins to decompose. At 655 C. copper sulphate (CuSOJ (see equation 8), decomposes. At 705C. cupro-cupric sulphate (CuO,SO 4 ), formed at the time copper sulphate is formed (see equation 9), begins to decompose. At 850C. copper sulphates are entirely decomposed. At 835 to 850 C. the maximum amount of soluble silver sulphate, (AgSOJ (when silver is present in the ore) is formed. At 1050C. copper oxide (CuO) is decomposed to Cu 2 O. 84 THE METALLURGY At 1100C. ferric oxide (Fe,0 3 ) is decomposed to the next lower oxide Fe 3 O 4 . Reactions in blast-roasting copper-bearing sulphides. When a mixture of iron and copper sulphide is pot-roasted, desulphurization proceeds rapidly if the ore be wet and silica added ; otherwise it pro- ceeds slowly. For these reactions we have : (15) 3FeS + 4H 2 O = Fe 3 4 + 3H 2 S + 2H, (16) 2Fe,O 3 -f 7H 2 S = 4FeS + 3S0 2 + 14H. When air is blown into the charge both hydrogen and H 2 S burn. In a well-burned charge some Fe 2 3 can be seen at the sides and top where cooled by radiation, but Fe 3 4 reacting on FeS gives FeO as follows : (17) FeS + 3Fe 3 4 == lOFeO + SO 2 . This reaction is exothermic and at a high temperature with silica would form ferrous silicate, again producing heat. Indeed, in action the. formation of this, with the consequent sintering, can be seen spreading as the burning proceeds. 17. ROASTING ORBS IN LUMP FORM. Heap-roasting. Roasting in heaps sulphide ores containing cop- per is an operation that must be performed with knowledge and care, in order to obtain satisfactory results. Before selecting this method a careful study of the environment must be made. It can not be adopted in a settled country where the fumes would be a nuis- ance, or where it would be likely to injure live-stock, or vegetation, or damage crops. In the arid and scarcely settled Rocky Mountain region, of the United States and Mexico, it may be employed to ad- vantage. Around an installation of moderate size, where no more than 25 tons of sulphur escape into the atmosphere daily, the area affected may be not more than four miles in extent ; and the site for the roasting-plant may be chosen with this in view. Often the pre- vailing winds blow from a single quarter, and permit placing the roast-piles where they give little offence. The location should be chosen with regard to the needs of the plant itself, so that smoke and fume will seldom be driven into the buildings. A roast-yard should be of ample size, approximately level, and so drained that the surface water shall not flow over it. At Jerome, Arizona, a track follows the contour of the hillside, and the roast- heaps are ranged close beside it so that the ore is trammed conveni- ently to the pile, unloade ft. inside diameter, and 75 ft. high. To charge a stall, the floor is prepared with large irregular pieces to form a rough flue or passage from front to back and two trans- verse ones, by which air can enter at the bottom. In the passages kindling wood is laid. The remainder of the floor is covered with a thin layer of long thin sticks of wood split from logs and poles (See Fig. 30). The stall is now filled with coarse ore and ragging OF THE COMMON METALS. 89 (1 to y-m. size), distributed through the mass. While partly filled, single small sticks of wood are placed at the back and sides as well as occasional sticks at the front. The front wall is also carried up with the larger pieces of ore. The filling completed, a single car- load of ragging is added above the ore, then a 3-in. layer of shav- ings, bark, and chips, then a layer of fine ore that can be roasted with care, and finally a coating of well-roasted ore. A sheet-iron cover on this, luted round the edges to the walls, is of great benefit. The air enters at numerous places. It enters beneath the ore by the rough channels, above described, through the interstices of the temporary 90 THE METALLURGY front wall, and through openings, a Fig. 29, in the side walls. The smoke leaves by the openings b into the main culvert. The wood of the stall having been ignited at the front near the bottom, the roasting proceeds rapidly, and, by the end of the fourth day, the heap is burning throughout. Successful burning is indi- cated by the swelling of the contents and the rising of the surface, sometimes to the extent of a foot. Because of this swelling, the front or temporary wall should be braced with wooden braces to oppose the outward thrust. When, on the other hand, the burning pro- ceeds too rapidly, no such swelling occurs, but instead, the surface subsides. Subsidence is due to the melting of ore owing to the great heat, and is an indication that roasting is imperfect. The heat can, however, be regulated by the use of fine ore to stop cracks, and by closing the passages to the draft openings a. If the ore were left to burn and cool slowly, it would require 15 days to do this. In order to hasten matters, the front portion of the ore as it cools may be removed, taking care not to penetrate beyond the cold portion. Beginning at about the fourth day it is possible to take ore away, so that in seven or eight days the stall is again empty. The ore is brought to the stalls by a track above the culvert or mid-flue. While a stall is being filled, a turn-plate, as shown in Fig. 30, is laid down with a branch-track, so that the ore can be conveyed above the stall and dumped into it. Tracks are also provided at the floor-level of the stall, and by these the roasted ore is conveyed to the blast-furnaces. Cost of heap-roasting. We find the cost of roasting at Duck- town, Tennessee, to be 42c. per ton. Peters gives as an average cost for fuel, labor, and supplies, 48. 5c. per ton, with common labor com- puted at $1.50 per day. Heap-roasting often can be done by con- tract to advantage. At the United Verde, Jerome, Arizona, 75c. per ton was the contract price. Cost of roasting-stalls. Peters gives for the cost of building 56 stalls a total of $3303.80, or about $60 per stall. To the total should be added $400 for the cost of the stack. Cost of roasting in stalls. This may be estimated at 50c. per ton, with common labor at $1.50 per 10-hour day. Relative advantage of heap and of stall-roasting. Heap-roasting has the advantage that it requires only the necessary site, and needs no investment for plant. The method is a simple one and the result is satisfactory. On a small scale, primitive methods of handling ma- terials are sufficient, but for a large scale we must not forget the cost OF THE COMMON METALS. 91 of grading, of trestles, of tracks, etc. Stall-roasting saves much time, requiring 10 days as against 70 days for heap-roasting. In large plants, where from 10,000 to 50,000 tons are in process of treat- ment, heap-roasting may cause the locking up of several hundred thousand dollars in the heaps. By reducing this to one-seventh by stall-roasting, an important saving is effected. In stall-roasting the stack removes the fume, and the entire contents of the stall becomes roasted, including the fine, which is more thoroughly roasted than in heaps. The elimination of sulphur is perhaps a little less thorough in stalls than in heaps. In stalls, rain and snow have little effect on the process, and in a moist climate the consequent leaching causes no trouble. For stall-roasting one-fifth cord of wood per stall is enough for charging, or 1% the weight of the ore roasted, while in heap-roasting average practice calls for 2% per cent. 18. ROASTING OF ORES IN PULVERIZED CONDITION. General. This work is performed in furnaces, generally of the reverberatory type, the ore being exposed upon a hearth to the ac- tion of the flame and air. The reactions by which the ore is roasted are given in section 13, on 'Chemistry of roasting'. The ore if not already fine, is crushed to the size designated under 'Crushing for treatment in roasting furnaces'. In these various furnaces advan- tage is taken of the heat developed by the oxidation of the sulphides. If the percentage of sulphur is high, this is often enough to supply the required heat (after combustion has once started) without the aid of extraneous fuel. Thus, in the McDougall roaster, after the furnace has been heated, an ore containing 25 to 30% sulphur con- tinues to roast by its own heat. The various mechanical furnaces roast ore cheaply, but for ores containing lead, which agglomerate, roasting-furnaces that are mechanically stirred do not give such satisfaction as hand-rever- beratory roasters. With a slight accession of heat above the normal, caused by the lack of care in firing, the ore is liable to agglomerate, and eventually to stick to and collect upon the hearth. In the hand- reverberatory roaster the hearth is accessible, and when this oc- curs the furnace can be used until the matter becomes serious, then the accumulation can be removed by 'cutter-bars'. On the other hand, such an accumulation soon stops the movement of a mechanical roaster. To remove it a flat bar of iron may be attached to one of the rabble-arms in the place of a rabble-blade. This is made stout enough to plow up the accretions and by setting it in different posi- tions on the arm, the hearth is finally cleared. This device has not, 92 THE METALLURGY however, proved to be altogether successful. The hand-reverbera- tory works well on ores that need a high finishing heat to break up the sulphates. Zinc ores are of this kind. Such temperatures are destructive to any kind of mechanical iron stirrer. We may classify roasting-furnaces into: (A) hand, and (B) mechanical-roasters. 19. HAND-OPERATED ROASTERS. The long-hearth reverberatory-roaster or calciner. In this fur- nace the charge is dropped and removed at intervals. The essential features (See Fig. 31 and 32), are a floor, or hearth upon 'which lies the ore spread over the entire surface of large area, a 'fire-box' at one end, a space below the fire-grate called an 'ash-pit', and a wall saparating the fire-box from the hearth called a 'fire-bridge' (or simply a 'bridge'). The whole furnace is covered by a flat arch, or 'roof, against which the heat of the flame is reverberated or thrown down, upon the charge. Thus the flame imparts heat during its en- tire passage over the ore to the outlet-flue at the opposite end of the hearth, where it escapes by a flue to the stack. These furnaces are distinguished from the reverberatory smelting-furnaces shown in Fig. 139, by the relatively small grate-area, and the flat hearth at the level of the door-sills. The figures represent, in sectional plan and in longitudinal sec- tional elevation, a long-bedded reverberatory hand-roaster. Disre- garding the fuse-box, Fig. 33 gives a good idea of its appearance. The width inside for convenience in stirring and moving the charge, should be 14 ft., while the length is designed according to the char- acter of the ore it is to treat. The length of the hearth must accord with the quantity of sulphide the ore contains, and consequently the heat the ore developes in roasting. Without the oxidation of sul- phur the fire would not maintain enough heat to roast the ore more than 32 ft. distant from the fire-bridge. The heat-generating power of the ore depends upon the percentage of sulphur contained but is greater when the sulphur is in the form of the loosely-held first equivalent. An ore containing only 10% sulphur would be roasted to good advantage in a short furnace. The length of the first hearth of the furnace shown in Fig. 31 is 16 ft. Where 15% sulphur is pres- ent it is proper to add another hearth, making a furnace 32 ft. long. A 20% ore would work rapidly on a three-hearth furnace ; and ore containing 25% sulphur or more, a four-hearth furnace, as shown in the figure, making a total length of hearth of 64 ft., which is suf- OF THE COMMON METALS. 93 ficient for any ordinary ore. Hearths of greater length have been tried, but have not been found satisfactory. The stack, proposed by Peters to furnish draft for two roasters, \ . x. > s <. is 42 in. square inside and 65 ft. high. The cost, together with that of the connecting flues, he gives as $728, and the cost of each rever- beratory roasting-furnace, $2713. ' A two-furnace roasting-plant 94 THE METALLURGY with building and accessories would cost approximately $10,000. This type of furnace has several advantages in the roasting of ore. The operation is started at a low temperature at which there is but little tendency for ore to cohere or agglomerate. The pulver- ous condition causes a thorough contact with the air, and makes it easy to rabble the ore. There is a saving in fuel, for this length of furnace reduces the temperature of the escaping products of com- bustion to 270 C. There is thorough stirring and turning, resulting from the movement of the charge toward the fire-end of the hearth. The firing is uniform and there is economy in repairs due to the uni- form and moderate heat of the furnace at the rear where red brick can be used. Moreover the heat near the bridge can be readily in- creased to decompose sulphates and to agglomerate if fusible, which improves it for treatment in the blast-furnace. Construction of furnace. The cast-iron side-door frames are set 6 ft. apart and opposite (See Fig. 33). The doors are made from plates of sheet-iron and are removable by means of a 'lifter'. The floor of the furnace is divided into hearths or divisions, separated by steps with a 2-in. drop, so that successive charges, kept on sepa- rate hearths, do not mix with one another to the detriment of the roast. Sometimes these steps are omitted, but in that case the charges must still be kept separate. The furnace is strongly stayed by 'buckstaves' which are tied across by tie-rods to resist the expansion due to the heating of the brickwork, and to take the thrust of the arched roof. Through the walls of the fire-box, openings 2% by 4 in. are often left for the admission of air above the level of the fire. The bridge has a passage through it to cool it, with side openings 2% by 4 in. by which air can be introduced beneath the flame and in contact with the ore. These openings furnish air, which, to- gether with that which" enters through the fire, produces an oxidiz- ing atmosphere at the hearth. The fire-box, bridge, and the first 16 ft. of the hearth should be of fire-brick. Beyond this, red brick of the better quality may be used, both for the roof and for the pave- ment of the hearth. All brick must be laid in clay, not in lime-mor- tar. A charge of ore is kept in the hopper at the flue-end of the furnace ready to be dropped upon the hearth when needed. The finished charge is removed through the discharge-opening into a car or wheelbarrow standing below. Operation of furnace. Assuming that the furnace is in regular operation, the charge on the first hearth is withdrawn by pushing aside a square cast-iron plate that covers the discharge-opening; and with the aid of a long-handled paddle and rabble, the roasted OF THE COMMON METALS. 95 charge on the first hearth is raked into the tram-car beneath. The paddle used for the purpose has a handle made of 1^4-in. pipe 16 ft. long. It has a blade 6 in. wide by 18 in. long. The rabble, really a jams 96 THE METALLURGY hoe, has a handle of the same length and a blade 6 in. wide by 10 in. long. The hearth being thus cleared, the charge on the second hearth is transferred to it by means of the paddle. In the same way the content of the third hearth is transferred to the second, and the fourth to the third, thus leaving the last hearth empty. The charge is dropped in from the hopper, and by means of the paddle, spread out on the fourth hearth. This moving down of the charge occurs every 4 hours, so that if we charge 2 tons, we roast 12 tons daily. From the 12 tons daily charged into the furnace we may obtain 10.2 tons of roasted ore, the difference in weight being due to the loss of sulphur, etc. Besides the movement necessary to advance the ore through the furnace the charge must be stirred or raked at least every 20 minutes. The proper fuel for the furnace is a free-burning semi-bituminous coal, which should be burned in a shallow bed, 6 to 8 in. thick upon the grate, and it should be added every 15 minutes. A roaster consumes 5000 to 6000 Ib. coal per 24 hours. With an out- put of 12 tons daily this is a consumption of 21 to 25% of the ore charged. The cost of roasting a copper sulphide ore in a long- bedded reverberatory furnace is $1.81 per ton ore charged. If the ore contains lead (which makes it slow and more difficult to roast), the cost may rise to $2.25 per ton. Reverberatory furnace with fuse-box. This is an ordinary long- bedded roaster to which has been added a slagging-hearth or fuse- box. Fig. 33 shows in perspective such a furnace, 75 ft. in total length. Its roasting-hearth is 57 ft. long by 15 ft. wide inside. Next comes the slagging-hearth 26 in. low r er, inside dimensions 11 by 13 ft. Separated from the slagging-hearth or fuse-box by the fire- bridge is the fire-box with grate-dimensions of 8 ft. by 2 ft. 10 in. or 23 sq. ft. area. In the fuse-box a high temperature, sufficient for melting or slagging the roasted ore, is produced, so that the fire- bridge must be furnished with a water-jacket, or coil of water-cooled pipes inserted within the brick-work of the bridge, to protect it from the corrosive or scouring action of the slag. Ore, in the hopper at the fuel-end, is dropped upon the bed of the furnace, and roasted as in the ordinary roaster. When roasted to this extent, it is discharged into the fuse-box where the tempera- ture is high enough to melt it down. When melted it is skimmed or withdrawn by means of a rabble. Certain ores containing blende, that in roasting produces sulphate difficult to decompose, are roasted and then slagged. Silica, which may be in the ore, or may be added in the fuse-box, reacts according to reaction (13), or when lead sul- phate is present by reaction (14), thus eliminating sulphur. The OF THIS COMMON METALS. 97 chief objection to such treatment is that there is loss of silver and lead as mentioned under 'Chemistry of roasting'. 20. MECHANICAL ROASTERS. These may be divided into : (1) Revolving cylinders, set nearly or quite horizontal, and re- volving on the long axis with (a) Continuous discharge, as the White-Howell and the Argall. (b) Intermittent discharge, as the Bruckner. (2) Automatic reverberatory roasters or calciners with continu- ous discharge having : (a) Straight-hearths, as the Brown-O 'Harra, the Wethey, the Edwards, the Merton. (b) Curved or circular-hearth, as the Brown-horseshoe, the Pearce-turret, and the McDougall. Besides these may be mentioned the Holthoff and the Raymond, the hearth of which itself revolves, and also the Stetefeldt shaft- furnace in which the ore is showered down a shaft. The White-Howell furnace. Fig. 34 is a longitudinal elevation of this furnace. It consists of a cylinder, 50 in. inside diameter by 34 ft. long, set at an inclination of 21/2%, supported on friction-rollers carried on the driving shaft. At one end is the fire-box, at the other a dust-chamber which connects by a flue to the stack. The hotter end of the cylinder, near the fire-box, is of larger diameter, to per- mit of its being lined with brick, thus leaving the cylinder of uni- form interior diameter throughout. Projecting, longitudinal, fire- brick ledges, set spirally, raise the ore and shower it back through the flame as the cylinder revolves, so as to roast it more rapidly. The unlined part for the same reason is furnished with longitudinal, cast- iron, projecting shelves. Ore is fed at the flue-end, by means of a screw-feed (See Fig. 194), and when dropped into the revolving cylinder, travels along, discharging at the fire-box end. Just before it reaches the fire-box it passes out from the cylinder to a brick chamber below, and is withdrawn from that, when cool. The furn- ace makes much flue-dust. It is used chiefly for chloridizing-roast- ing, upon ores containing but little sulphur, and has a capacity of 50 tons per 24 hours for low-sulphur ores. The Bruckner roasting-furnace. This consists of a brick-lined cylinder, supported by and revolving on four rollers or carrier- wheels. As in the White-Howell furnace, there is a fire-box at one end, and a flue at the other. The flame from the fire-box is. drawn 98 THE METALLURGY directly through the end-openings of the cylinder to the flue. The furnace treats the ore in charges that require 24 to 48 hours. The OF THE COMMON METALS. 99 cylinder is provided with man-holes for charging and discharging the ore, the ore being first put into a double hopper from which it is quickly drawn when needed. Fig. 35 shows the end and side of the furnace, indicating founda- tions in section. In the end-view at the right, the fire-box is re- moved. The cylinder, 8% ft. diam. by IS 1 /^ ft. long, has end-open- ings 3 ft. diam. for the entrance and exit of the flame. It is driven by a worm-gear, the motion being communicated to two of the car- rier-wheels on the common shaft and transmitted to the cylinder by the friction between the bearing rings and the carrier-wheels. The fire-box is a movable one that can be set aside when it is desired to reach the interior of the cylinder. The present practice is to revolve the cylinder slowly, say once an hour, since the contents, even at this speed, is constantly shift- ing and representing new surfaces to the air. The old way was to revolve it once in four minutes. At the time of discharging, the speed should be at least one revolution in 2 to 4 minutes to discharge the ore quickly. This adjustment is made by throwing in a clutch by means of a quick-working mechanism. All the man-holes are opened, and it takes but few revolutions to discharge the cylinder. The ore falls into a pit below, whence it can be withdrawn to tram- cars for use. Charges that contain lead are liable to agglomerate, even with cautious firing, and to 'hang up', that is, adhere in a layer to the brick-lining. Should this occur, the movable fire-box can be pushed aside and the layer removed by long chisel-ended slice-bars. The attachment to the brick-work is so slight that, when the layer is cut from end to end, even at one place, the key or con- tinuity is broken and the mass falls. Thus it is easily dislodged and by further rolling breaks up and is ready for continued roasting. A moderate cohesion of the fine particles does not present a good roast. Operation. The charge of 20 tons having been dropped into the cylinder from the hopper, the man-holes are closed and vigorous firing begun to start the ore to burning by its own oxidation, which begins at a barely visible red. This takes about six hours, the nec- essary temperature being attained first at the fire-box end and ex- tending then to the flue-end. The charge thus started burns by its own heat 12 hours longer and the fire meanwhile is withdrawn from the fire-box. Indeed the reactions are so vigorous that the charge must be closely watched and the air-supply regulated lest agglom- eration occur if the ore be a leady one. As the heat toward the end of this stage diminishes, firing is resumed, increasing the heat gradually, 6 to 8 hours more, to the finish. In this way sulphates, 100 THE METALLURGY formed early in the roast, are decomposed, and a good result is ob- tained. The whole cycle of operations takes 24 to 36 hours for cop- per ores and 48 hours or more for leady ores. For copper ores it is sufficient to reduce the quantity of sulphur to 7 to 8%, and for leady ores to 3 to 5%. The charge is removed by revolving the furnace rapidly, and after 10 to 15 minutes little is left to mix with the next charge. To charge the cylinder again, all but two man-holes are closed, and the cylinder is revolved to bring these directly under the hopper-spouts. The hopper-slides are withdrawn and the ore is quickly run into the cylinder and is ready for firing on closing the man-holes. These operations of charging and discharging need not take more than 20 minutes. For copper sulphides, roasted in 24 hours, the furnace capacity accordingly is at least 20 tons daily, and for leady sulphides roasted in 48 hours, 10 tons. The cost of roasting leady ores is 85c. per ton, and for copper sulphides roasted to 7 to 8%, 42c. per ton. The cost of one of these roasters, installed, may be estimated at $3000. The Wethey roasting-furnace. Fig. 36 is a perspective view, and Fig. 37 a cross-section of this furnace which consists of two, super- imposed, straight hearths, heated by fire-boxes placed at the sides of the upper hearth. The roasting is done on the upper hearth ; and the lower hearth is for cooling the ore after it has been roasted. The upper one of the two hearths, each of which is 121 by 12 ft: in size, is held between heavy transverse I-beams above and below, as shown in Fig. 37, which tie the upright buckstaves and support the hearth. The roof of this hearth is so suspended from the upper I- beam as to leave slits the entire length of the hearth; and through these project the ends of the rabble-arms which rest on carriages. The stirring blades, or rabbles, within the furnace, are set diagonally upon the rabble-arms (see the detail of such an arrangement for a Pearce-turret furnace shown in Fig. 42). The ends of the rabble- arms are attached to endless chains, which drag them along the upper hearth, stirring and gradually moving the ore forward. They return by the lower hearth upon which the ore, from the finishing end of the upper hearth, falls through an opening. Here it is again moved along, stirred, cooled, and finally discharged into a hopper, thence to be drawn into two-wheeled buggies and transferred to the leaching-vats. The endless chains pass around sheaves at the ends of the furnace, and around sprocket-wheels at the driving end which impel them. The rabble-arms are so arranged that they can be re- moved readily without disturbing the carriage connections. There are two rabble-arms, and these pass along the hearths at the rate of OF THE COMMON METALS w . 100 ft. per minute. The blades of one rabble are set at the opposite angle to those on the next, so that the ore tends toward neither side. To hasten the cooling of the ore upon the lower hearth, that it may be ready promptly, for further treatment, water-cooled pipes are Fig'. ?,6. VIEW OF WETHEY ROASTING FURNACE. * * # Fig. 37. CROSS-SECTION OF WETHEY ROASTING FURNACE. laid the length of the hearth in grooves in the brick pavement flush with the top surface. The ore is regularly fed to the furnace from the hoppers by automatic feeders at the driving end (shown at the right in the perspective view, Fig. 36), and is heated by two sets of fire-boxes, one of which is at the feed end, the other half way along THE METALLURGY the hearth. The flame from the fire-boxes descends to the hearth by a flue crossing the roof (See Fig. 36), then moves horizontally to the left, where by a flue at the end it escapes to the stack. Thus the ore and flame move in the same direction over the roasting-hearth. At each end of this hearth flap-doors of sheet-iron are hinged. These hang by the top edge of the sheet, and close the ends of the furnace at all times except when the rabbles enter or pass out of the furnace. The doors are lifted by the rabble, but drop again as soon as it passes. It might be thought that the slits, each 120 ft. long, would admit too much air and injure the draft, but the furnace is found to work well Fig. 38. PLAN AND ELEVATION OF EDWARDS ROASTING FURNACE. notwithstanding this feature. The rabbles are in the open air more than half the time, and have a chance to cool after each passage through the furnace. We estimate the capacity of this furnace, roasting ore containing Vi2 to 3y^% sulphur that needs 13 to 15 sq. ft. hearth-area, at 100 tons per 24 hours. The Edwards roasting-furnace. This is a single-hearth rever- beratory furnace with hearth dimensions 57 ft. long by 6 ft. wide. Fig. 38, in plan, shows a portion at the fire-box end, the feeding mechanism and the cooling floor in section. The elevation shows the side, constructed like a plate-iron beam, the stirring mechanism and the conveyor for transferring the roasted ore to the cooling-pit. F1g )9 is a transverse section of tho hearth showing the details of OF TIIK COMMON METALS. 103 the stirring mechanism. The slope of the furnace can be changed a little by tilting. This regulates the rate of travel of the ore through the furnace ; but for a given kind of ore, this slope, once determined,- is not again changed. The furnace has a slope of 2 in. per foot to- ward the discharge or fire-box end. The stirring and propulsion of the charge is affected by means of rabbles fixed to vertical shafts, as shown in the elevation of Fig. 39, and in the plan of Fig. 38. The rabbles at the fire-box end are water-cooled, and this is found especi- ally necessary where a high finishing heat is needed. The blades or plows of the rabbles can be easily replaced through the doors adja- cent to them. The figure indicates the hearth as broken away, at the discharge end, to show two of the rabbles in plan. The last Fig. 39. CROSS-SECTION OF EDWARDS ROASTING FURNACE. rabble sweeps the roasted ore into the discharge shoot, and the push- conveyor then moves the ore to the cooling-pit. The bottom of the conveyor-trough is furnished with slides, by means of which the ore can be dropped at any desired point on the cooling-floor. The ore is fed to the furnace from the feed-hopper, by an endless-screw con- veyor which discharges into a feed-opening in the roof of the fur- nace. The smoke is carried off by a flue. The furnace takes 1 hp. to operate, and has a daily capacity of 25 tons on sulphide ore of 30 to 35% sulphur. The roasted ore contains 3 to 8% of sulphur. The moving parts are durable, and the furnace has proved efficient in practice. Large installations, of the duplex type with a double in- stead of a single row of rabbles, and of hearth-dimensions 120 by 12 ft., have been built for a daily capacity of 60 tons. These furnaces do not have the tilting hearth. 104 TTTIO METALLURGY The Pearce-turret roasting-furnace. The word 'turret' has ref- erence to the circular form of the furnace. Three types have been developed, namely, the one-deck, the two-deck, and the six-deck or Fig. 40. TWO-DECK PEARCE-TURRET ROASTING FURNACE (PL.AN). six-hearth furnace. The greater the number the hearths, the more economical the furnace is as regards fuel and output. On the other hand, the multiplicity of hearths makes it more complicated, and 106 TIIK METALLURGY causes more flue-dust by the repeated dropping of the ore from hearth to hearth. We shall now consider the two-deck type shown in Fig. 40 and 41. Fig 40 is a plan, with opposite portions broken away to expose the hoppers and the rabble. Fig. 41 is a central sectional elevation through a fire-box and the hearths. To understand these views will require careful study. There are two superimposed annular hearths with space between for the 15-in. I-beam supporting the upper hearth and a flue-chamber below. Referring to the plan, the ore-hopper, partly broken away, is shown at the near' side. Next, at the left, is the outlet-flue with its barred top, for the escaping gases ; and at the right, where the furnace is broken down to the lower hearth, is shown the discharge-hopper. The ore is fed continuously from the ore-hopper (See Fig. 41), through the roof to the upper hearth until, having made its circuit, it falls through a transverse slit to the lower hearth. Here, as before, it is stirred by rabbles and moved forward until, having again made the circuit, it is received into the discharge- hopper of the lower hearth. There are two rabble-arms for each hearth. Of these, the upper two are shown in the plan. The lower two are at right angles to the upper ones and are shown in the elevation. Each pair of rabble- arms is screwed into a central hub, that revolves on a fixed central column. A gear-wheel on the rabble-arms, in connection with a driving mechanism, moves the rabble at the rate of about one revolu- tion in two minutes. As will be noticed in the plan, Fig. 40, and in the detail, Fig. 42, the rabble-blades are set at an angle to their line of travel. Nearly touching the hearth they pass through the ore, stirring it and moving it slightly forward in pushing it aside. The blades of the opposite rabble-arm incline in the opposite direc- tion and stir the ridges thus formed. Thus every minute fresh sur- face is exposed to the action of the fire. An objection has been urged against this, as against all the mechanically stirred roasters (except the Edwards), that a part of the imperfectly roasted ore, adhering to the blades, creeps forward and mixes with the advanced portions of the charge, raising the percentage of sulphur in a way that need never occur in hand-roasting. Air from a fan-blower is forced into the central hollow column, thence out through the rab- ble-arms, thus cooling them, finally through nipples screwed into the underside of the arm between the blades, as shown in Fig. 42. In addition to this forced air from the fan, air enters through open- ings at the sides to oxidize the ore. The fire-boxes, one of which is well shown in Fig. 41, are sup- OF THE COMMON METALS. 107 plied by air from the fan-blower under a slight pressure. By means of the under-grate blast the flame is caused to enter the furnace under a slight pressure, and thus oppose the inward draft of air through the side-doors. Suction increases as we approach the exit- flue, as indeed it must in order to take away the products of com- bustion. The fire-box is furnished with a step-grate suitable for burning slack coal. The ash-pit is closed by tight sheet-iron doors, so that the under-grate blast can be sustained. Occasionally these are opened so that the firemen can clean the grate. The ashes and clinkers are dislodged from between the step-grates, falling into the rta. 9. ,' tNVCPTtO JftMt. Fig. 42. DETAIL, OF RABBLES. ash-pit arid being thence removed. The fire-box is supplied with coal by means of a hopper kept constantly full. To prevent an intense action of the heat where the flame would naturally strike the charge, a curtain-arch, lower, and having less curvature, as seen in Fig. 41, deflects and distributes the flame. There are three fire-boxes, all of which are upon the upper hearth ; and the movement of the flame is in a direction contrary to that of the rabbles and the ore. The flame from the fire-box nearest the feed moving toward the outlet- flue heats the ore and sets it afire. The second and third fire-boxes raise temperature of the ore to the full heat ; but the final action on the lower hearth is- performed by the heat from the oxidation of the ore. The outlet-flue leads from the roof of the upper hearth, down the side of the furnace, to the dust-chamber below. An under- 108 THE METALLURGY ground-flue from the dust-chamber leads to the chimney. The hearths are bound by heavy bands. The upper hearth and the mechanism of the furnaces are sustained by the I-beam construction that serves also to provide for the slit through the inner wall of the hearth through which the rabble-arms must enter. The slit is cov- ered by a sheet-metal band, which is attached to and revolves with the rabble-arms. The question might be asked, whether the flame from the third fire-box would not go directly to the outlet-flue in- stead of following round the hearth. To prevent this, the upper hearth is broken or interrupted for a space of 5 feet along the hearth, the rabbles being visible for inspection as they pass through this open section. The ends of the hearth are closed by swinging sheet- iron doors, like those of the Wethey furnace. As the rabble enters the open space the doors are lifted by the rabble-arm, and drop by gravity when it passes on. Though complicated in design, this is a furnace that has worked well in practice. It is 33 ft. interior diame- ter and has a 6 ft. hearth, this giving for the hearths an area of 1020 sq. ft. Its daily capacity is 35 tons of ore of 35% sulphur, which it roasts to 6 to 1% sulphur with a consumption of 9.1% fuel. The labor needed is but little more for a double-deck furnace than for the single one, while the fuel-cost per ton of ore is reduced one- half. Thus a single-deck furnace consumes 18% of fuel. More flue- dust is made in the double-deck furnace. To operate the two-deck furnace requires 3 hp., and the cost of roasting is 98c. per ton. Such a roaster costs $8000 installed. In a multiple, or six-deck furnace, the fuel has been reduced to 1.4% of the ore tonnage, but the flue- dust is increased to 4 per cent. The McDougall roasting-furnace. There are several kinds of furnaces of this type. Among these are the Herreshoff, the Wedge, and the Evans-Klepetko. The latter, as manufactured by the Allis- Chalmers Co., is shown in sectional elevation in Fig. 43. It is a vertical, cylindrical furnace, 16 ft. diam., with six arched hearths, over which travel rabbles which stir and move the ore gradually toward the central drop-opening through the floor of each hearth, situated alternately at the center and at the periphery. A centra I shaft is provided, carrying six radial rabble-arms (three of these arc hidden by the shaft in the illustration), provided with nibble-blades set at an angle on the arm as shown in Fig. 42. The rabble-blades on the even-numbered hearths are so set as to push the ore toward the periphery ; the odd-numbered ones toward the center. The ore, fed continuously into the furnace from a cylin- drical hopper shown above and at the right, Fig. 43, drops upon the OF THE COMMON METALS. 109 upper hearth near its outer edge. The rabble-blades of that hearth stir and move the ore gradually toward the central drop-opening where it falls to hearth No. 2. The rabbles of this hearth again stir and move it to the outer drop-openings, through which it falls to hearth No. 3. The ore advances by this means until it reaches the lower hearth, where an opening at the periphery gives it exit to a Fig. 43. McDOUGALL ROASTING FURNACE (ELEVATION). receiving-hopper, shown beneath the hearth, from which it is drawn into a car as is required. A high-sulphide ore roasts by its own heat when the furnace is in full operation. The ore fills the hearth to the level of the blades, and is spread out evenly by them. On the upper hearth, as the ore moves toward the central opening, it becomes dry and hot, and when dropped upon hearth No. 2, begins roasting. On hearth No. 3, the 110 THE METALLURGY ore roasts freely emitting sparks and forming sulphates. On hearth No. 4 no sparks are seen, and the ore has attained its highest tem- perature. On hearth No. 5 the ore looks less bright; and on No. 6, especially at the discharge, it has become cooler. The air for oxidation is admitted by side-doors, mostly those of the lower hearths. The gas, and dust, passing up through the drop- openings, are drawn through the horizontal main flue. In starting, the furnace is heated to the kindling temperature of the ore which, if rich in sulphur, burns by its own heat, without the aid of fuel. If the sulphur content is low, additional heat is supplied by one or more external fire-places, near the bottom of the furnace. To protect the rabble-arms from the intense heat they, and like- wise the central shaft, are water-cooled. The cooling-water is forced down the 9-in. hollow, central shaft in a 3-in. pipe to a point near the bottom, and out to the ends of the arms in 1-in. pipes. It then returns up the annular space between the 3-in. pipe and the hollow shaft, and discharges at the top through two spouts into a launder. The furnace is IS 1 ^ ft. high by 16 ft. diam., and has a total hearth- area of nearly 1000 sq. ft. The structure is supported on columns to give room below for the hopper and the car into which the roasted ore is discharged. The shell made of %-in. plate-steel, is lined with 9 in. of brick-work. The rabble-arms consume I 1 /!' to 2 hp., and make one revolution in 1% minutes. A furnace treats, in 24 hours, 40 tons of sulphide ore of 35% sulphur, reducing it to 7 per cent. About 4% flue-dust is made ; and the ore itself contains more ferric oxide, and is lighter and more porous than if treated in a hand-reverberatory roaster. The cost of roasting such ore is approximately 35c. per ton, which is the lowest figure thus far known for any furnace. The compact form of the furnace reduces radiation to a minimum and enables roasting with little or no fuel. Taking capacity into consideration, the furnace is one of moderate price, and one that costs little to keep in repair. Of the two revolving-hearth furnaces the Holthoff and the Ray- mond, the latter has some popularity for the preliminary roasting of ores for lime or pot-roasting, the powdered ore being showered down a vertical shaft or tower and coming in contact with an upward flame from a fire-box. An objection to its use is the flue-dust that is made. 21. ROASTING OF MATTE. Copper-bearing matte is not difficult to roast, but must be crushed at least to 4-mesh size. OF THE COMMON METALS. Ill The term 'roasting' is applied also to a method of treating cop- per matte in a reverberatory furnace in large pieces, upon which an oxidizing flame is allowed to play. Such masses slowly melt and are acted on by the air, whereby a part of the material becomes oxidized or roasted sufficiently for the next operation. As com- pared with ordinary roasting this is slow, and the method is one but little used. Lead-bearing matte from the silver-lead blast-furn- ace, to be roasted in a reverberatory furnace of the kind shown in Fig. 32, needs a different treatment from that given to ore. This kind of matte contains but 20% sulphur, and does not take fire like pyrite ore, but must have a high finishing heat to expel the sulphur. Such matte is considered well roasted when it contains 4% sulphur. Ores low in lead can easily be roasted to 2 to 3% sulphur, while galena, when roasted, still contains 5 to 6% when drawn from the furnace. Like matte, galena starts burning slowly, and must be roasted slowly, for rapid heating at once causes it to sinter and thus stops further roasting. Typical leady matte contains metals and sulphur as follows : Raw, low-grade. Eoasted, low-grade. Raw, shipping. Per cent. Per cent. Per cent. Pb 10.66 10.49 9.06 Cu 4.62 4.12 42.30 Fe 53.11 52.41 20.00 S 26.87 6.13 17.89 95.26 73.15 89.25 The roasted, low-grade matte tabulated above contains 23% oxygen. This explains why it does not lose weight in roasting. Pyrite ores of 20 to 30% sulphur, on the contrary, easily lose 15% in weight. 22. LOSSES IN ROASTING. Such loss depends upon the extreme to which the roasting is car- ried as well as upon the nature of the ore. When ore is so roasted that it is not sintered at the final high temperature, the lead lost averages 2.5% but no loss of silver occurs. When the temperature is carried higher, and the ore is agglomerated, the loss is slightly higher. When fused it may reach 15 to 20% of the lead and 2 to 5% of the silver. Of the gold little is lost in oxidizing roasting. 23. CAPACITY OF FURNACES AND COST OF ROASTING. These depend upon the surface exposed to the oxidizing influ- 112 THE METALLURGY ences and upon the quantity of sulphur contained in the ore. Sili- cious ore, containing !/> to 3 l / 2 % sulphur, requires 13 to 15 sq. ft. of hearth-area per ton ore roasted per 24 hours. Matte containing 20 to 25% sulphur, when it is required to reduce the sulphur content to 4%, needs 45 sq. ft. hearth-area; copper sulphide ore, roasted to 1% in preparation for smelting, requires 33 to 35 sq. ft. For roast- ing iron-sulphide concentrate, which carries 35 to 45% sulphur, down to 3 to 10% sulphur, 55 to 60 sq. ft. hearth-area is needed. Ore-roasting in heaps, at Jerome, Arizona, costs 80c. per ton, in- cluding general expense. Ore-roasting in stalls costs 50c. per ton. For reverberatory roasting, in long, hand-rabbled furnaces the low- est price attainable on copper ores was $1.50, with an average of $1.81 per ton. For roasting lead-bearing ores, $1.75 is a moderate cost, and from this the cost, when all items are included, may rise to $2.25 per ton. The Allen-O 'Harra automatic furnace, having two straight hearths each 94 by 9 ft., and resembling the Wethey fur- nace, roasts 45 to 50 tons daily at a cost of 78c. per ton. The Wethey furnace, of the type having four hearths, each 65 by 10 ft., the roast- ing proceeding on all the hearths, roasts 90 tons daily to 5 to 6% sulphur, at a cost of 98c. per ton. The 16-ft. McDougall furnace (Herreshoff type), having five hearths, 14% ft. diam., and a total area of 830 sq. ft., roasts 33 to 35 tons daily to 7% sulphur at a cost of 50c. per ton. The Bruckner roasting cylinder, 8Vi> ft. diam. by 22 ft. long, takes a charge of 20 tons (10 tons daily), and in 48 hours roasts it to 4% sulphur at a cost of 80c. per ton. It will be noticed that the low cost of roasting in some of these furnaces is due to their needing no fuel after coming into full op- eration. To obtain this effect such furnaces have several hearths, and are compact. On account of this compactness they lose but little heat by radiation. 24. BLAST OR POT-ROASTING OF ORES. Both lead and copper ores are treated by blast or pot-roasting, though the method was at first intended for lead-bearing ores, es- pecially for galena. We have alread.y mentioned the difficulty of roasting galena by the old method, in the reverberatory furnace ; but by pot-roasting, it can be so treated as to remove most of its sulphur, with less loss by volatilization. Treatment of galena. By the Huntington-Heberlein process, called also the ' II and H process', the galena-bearing ore is given an incomplete, rather rapid roast, to reduce the amount of sulphur to 12 to 14%. The product from the roaster is mixed with a certain OF THE COMMON METALS. 113 proportion of limestone and silicious ore, wet down, and charged into a hemispherical cast-iron pot 8 1 /? ft. diam. by 4 ft. deep, having a capacity of 8 to 10 tons. Within the pot, and forming- a false-bottom, is placed a circular arched plate perforated with %-in. holes to admit air to the charge under pressure. Upon the false-bottom is scattered a wheelbarrow-load of ashes, then a carload (one ton) of hot ore from the roaster. On this is dumped 8 tons of wet charge. Air, under the pressure of a few ounces, is admitted beneath the false- bottom, and coming up through the hot ore, it produces a burning- temperature and starts the combustion of the charge. The heat gradually ascending to the top, the charge becomes red-hot, and S0 2 and SO 3 escape. At the end of the roasting, which lasts sometimes 16 hours, desulphurization is complete and there remains only 3 to 5% sulphur if the charge is properly burned. The pot is now in- verted to discharge the contents, and this falls out in an agglome- rated, red-hot mass. It is broken to a size suited to subsequent treatment in the blast-furnace. There are three patented variations of this treatment of galena ores. In the first, the Huntington-Heberlein process (above de- scribed}, the ore is mixed with limestone, partly roasted, wet down, charged into the pots, and blown, to the point of agglomeration. In the second, the Savelsburg process, the ore is mixed with limestone and charged directly into the pot or 'converter', the preliminary roasting being omitted. In the third, the Carmichael-Bradford pro- cess, the ore is mixed with gypsum and charged into the pot without preliminary roasting. Treatment of copper sulphides. This treatment, like the Huiit- ington-Heberlein, consists in blowing the partly-roasted ore or matte in pots, but in this case no lime is used. The charge is first moist- ened so that it easily coheres. The minimum quantity of water to procure the best result is 3 to 4% for the low-grade matte, 4 to 6% for high-grade matte, and 6 to 9% for ore. In the case of ore it is found that, unless the water is in excess, ferric oxide is produced, and this, forming round the particles, prevents proper slagging; whereas with sufficient water, ferrous silicate is produced. The charge that works best consists of about one-third of pieces 1 to l 1 ^ in. diam., and two-thirds of fine concentrate. It should contain 15 to 35% Si0 2 and 15 to 25% sulphur. In treating a charge the false-bottom is covered with lumps of roasted ore, after this, a small fire of chips and sawdust is started, and urged by a light blast. Ore is charged next, keeping it deeper at the sides, until the pot is half filled. Holes are punched into the 114 THE METALLURGY mass with a half-inch pointed rod, and through these appear sul- phur vapor and sulphur dioxide, forced by the blast. In about an hour a ring of fire begins to show, the remainder of the charge is then put on, the pot is covered with a hood, and the blast gradually increased to 13 oz. per sq. in. After some hours the evolution of SO 2 slackens and the charge gradually becomes red-hot throughout. The blast is then stopped, the blast-pipe disconnected, and the con- verter inverted to discharge the contents. As arranged in some cases, the pot with its load is picked up by an overhead traveling- crane and dumped near a large crusher, and the roasted material crushed into lumps. The time required to treat a charge varies from 8 to 12 hours. The product consists of a porous sintered mass of ferrous silicate containing shots of matte and free silica. It is well suited to blast- furnace smelting. . A large quantity of the product in one instance contained 5.65% sulphur, while as little as 5% has been obtained in the treatment of ore. The process is particularly suited to the treat- ment of matte, to which 15 to 25% of silicious ore is added, to unite with the iron, forming ferrous silicate. In ores containing pyrite, the loosely held equivalent of sulphur is first driven off. The reac- tions should proceed rapidly so that a high temperature can be main- tained to permit the formation of ferrous silicate, otherwise ferric oxide will be formed, diminishing the amount of active oxygen near it. Wherever this occurs, the temperature drops, and a patch of unsintered material is formed. PART III. GOLD PART III. GOLD. 25. GOLD ORES. Gold occurs in nature, both in the native state and combined with tellurium. Native gold occurs in vein-matter disseminated in grains or par- ticles of various size, and it is found not only in quartz veins but in veins or lodes containing hematite, iron-pyrite, arsenical-pyrite, blende, and galena. In pyrite it occurs not only in the substance of the crystals but as films on the surface of these crystals. It is fre- quently accompanied by silver. When gold-bearing veins have be- come disintegrated and swept away into alluvial deposits, the par- ticles of gold, where released, are found in the sand and gravel of the beds, the pebbles and boulders themselves (which have come from the country rock), being in general barren of gold. Gold occuring in this way is called alluvial gold, and is recovered by methods of hydraulic mining or dredging, which belong to mining engineering rather than to metallurgy. We shall here consider, therefore, the treatment of gold ore. Gold tellurides. In South Dakota, at Cripple Creek, Colorado, in Western Australia and elsewhere is to be found gold combined with tellurium as calaverite, AuTe 2 (containing 41.4% Au and 57.3% Te) ; also gold and silver combined with tellurium as sylvanite (AuAg) Te 2 , and as petzite (Ag 2 Te, Au 2 Te). Classification on treatment basis. Gold, regarded from the stand- point of milling and amalgamation, occurs in the following condi- tions : (1) In ordinary amalgamable form, generally called 'free-mill- ing'. (2) In some form of intimate physical admixture, with other minerals, or in chemical combination with other elements. (3) As rusty gold, sometimes metallic in appearance and of usual golden color, but often brown and lusterless. In this supposed allotropic form, it resists the action of mercury in the process of amalgamation. 118 THE METALLURGY 26. STAMP-MILL AMALGAMATION. This consists in crushing the gold-bearing ore, generally in a stamp-mill, to a size of 20-mesh, or finer (using 6 to 8 tons of water to one ton of ore), and running the ore-pulp over plates 4% ft. wide by 6 ft. long to which the gold adheres by amalgamating with the mercury with which the plate is coated. Occasionally the battery is stopped for a short time, and the gold-bearing amalgam is scraped from the plate and treated to obtain the gold. This process is called 'outside amalgamation'. Sometimes amalgamated plates are placed inside the mortar, and the particles of gold contained in the ore are driven against the plates by the movement of the pulp and adhere. The gold is recovered in the same way as on the outside plates. Some of the gold particles and mercury fall into the crevices be- tween the dies, at the bottom of the mortar. This gold is recovered at the time of the monthly clean-up. From time to time a little mercury is placed in the mortar (about 1.5 oz. per ounce of gold in the ore), as the crushing proceeds. The tailing, or residue after obtaining the gold, may be run to waste. If, however, ore contains pyrite, or other heavy sulphide minerals within which a part of the gold is locked, only a part of the gold can be recovered by amalgamation. Pyrite, being heavy, can be caught on concentrating tables, and the concentrate so recovered can be sent away to be smelted or be treated by leaching methods. At Treadwell Island, Alaska, only half the gold in the ore is obtained by amalgamation, and the concentrate, 2^/2% of the weight of the ore, carries most of the remainder. The tailing from the concentrat- ing tables commonly contains a small amount of gold. The mechanical operation of pulverizing the ore to a fineness such as to liberate the gold particles from the enclosing gangue consists first in coarse-crushing the ore in a Blake type crusher (See Fig. 22), which discharges into an inclined-bottom bin, and withdrawing it thence to be crushed in a stamp-battery. The battery. Fig. 44 is a perspective view, Fig. 45 a side eleva- tion, and Fig. 46 a front elevation of a 10-stamp battery such as is used in gold milling. Fig. 45 is a good example of a dimensional drawing. Referring to Fig. 44, the parts will be found designated as fol- lows : A, mortar-block; B, mud-sills; C, cross-sills; I), side-posts; E, platform ; F, G, buck-staves ; H, lower guide-timbers ; 7, upper guide- timbers ; J, mortars ; K, screen ; L, die ; M, shoe ; N, boss or head ; 0, stem; P, tappet; R, cam-shaft; S, collars; T, cam-shaft boxes; U, OF THE COMMON METALS. 119 cams ; V, cam-shaft pulley ; W, line-shaft ; X, tightening-pulley ; Y, water-pipes ; Z, automatic feeder. On two of the cross-sills C, is to be seen the frame that supports the apron-plate. The mortar-blocks, A,A, are made of timbers set on end, as shown also in Fig. 45 and 46. They are set in a pit that extends down to Fig. 44. PERSPECTIVE VIEW OF TEN-STAMP BATTERY. solid rock or to a concrete foundation, and upon the solidity de- pends the durability of the battery. They are bolted together hori- zontally, and are held in position by a tamping of sand, rock, or con- crete, completely filling the pit around them. Instead of wooden mortar-blocks, concrete ones have, of late, come successfully into use. These are more solid, more durable, and cheaper. The mortar J is firmly held to the mortar-blocks by long bolts 120 THE METALLURGY (See Fig. 46), three thicknesses of blanket of a piece of rubber-sheet- ing being interposed to give an even bearing. Fig. 45. SIDE ELEVATION OF TEN-STAMP BATTERY. OF THE COMMON METALS. 121 The stamp-frames consisting of the mud-sills B, the cross-sills (7, the side-posts D, the buck-staves F and G, and the guide-timbers H and I are of the dimensions shown in Fig. 45 and 46. The mud-sills, supporting the cross-sills (7, run the length of the mill, under all the batteries, and carry the line-shaft W. A tightening pulley just above -4- 25'- Fig. 46. FRONT ELEVATION OF TEN-STAMP BATTERY. the lower pulley, serves to tighten the belt and set the battery in mo- tion. To the guide-timbers H and / are bolted the guides, often made of two planks each 4 by 14 in. bored vertically with 3-in. holes for the wrought-iron or steel stems. Otherwise cast-iron individual guides are employed. A mortar, as seen in Fig. 47 and 48, is a cast-iron box having at 122 THE METALLURGY one side a feed-opening through which the ore enters, screen-open- ings at one or both sides for the screens, through which the stamped ore or pulp discharges, a heavy base on which rests the die, and sides enclosing the whole, the mortar being a single heavy casting weigh- ing 2 to 3 tons. Fig. 47 is a mortar of the single-discharge type, used in gold-milling; and Fig. 48 a double-discharge mortar, used in sil- \ I Fig. 47. SINGLE-DISCHARGE MORTAR. ver milling. With the single-discharge mortar the output is less, and the pulp is retained a longer time in the mortar. In consequence it is more finely ground, and more intimately brought in contact with the inside plates when used. The double-discharge mortar on the other hand on account of the greater total opening, discharges freely, and pulverizes more ore. It is not a mortar intended for amalgama- tion, but for crushing only. Upon either mortar will be noticed a OF THE COMMON METALS. 123 deflecting-lip at the bottom of the feed-opening. This is provided to discharge the ore nearer the center of the mortar. Beneath it, and protected from the wear of the entering ore, an inside plate is some- times placed. Screen-openings are provided at the front, and in the double-dis- charge mortar also at the back of the mortar. The screen K, Fig. 44. Fig. 48. DOUBLE-DISCHARGE MORTAR. tacked to a wooden frame, is made either of wire-cloth or punched- plate. Of the two kinds, the punched-plate screen has the advant- age in strength and first cost. The wire-cloth, on the other hand, if of copper or brass, is durable and gives a greater discharge area. Thus in the case of No. 7 punched-plate, Fig. 49, we have effective in discharge openings but 10% of the total area ; while in the case of the wire-screen 27% is open and in consequence, there is less sliming 124 THE METALLURGY or excessive powdering of the ore because of its prompt escape from the mortar. To increase the height of discharge, when the pulp is to be re- tained a longer time in the battery for finer crushing a wooden chuck-block is placed in the discharge-opening beneath the screen- frame. The height of discharge may be defined as the vertical dis- tance from the top of the die to the bottom of the opening of the screen-frame. Where inside amalgamation is practised, an inside amalgamated plate covering the chuck-block is used in addition to other plates, and on these plates as much of the gold as possible is collected. Referring to Fig. 47 and 48, we see a foot-plate that cov- ers the bottom of the mortar, upon which rest the five cylindrical dies 8 to 9 in. diam. by 7 in. high when new. Resting on the die we Fig. 49. PUNCHED SCREEN. Fig. 50. CANDA CAM. notice the stamp-shoe fitting into the head, which in turn is secured to the stamp-stem. The shank of the shoe is made a little smaller than the corresponding recess in the head, and in inserting it, wooden shims are placed around the shank, and the stamp is dropped upon it, thus wedging the shoe into the head. Dies and shoes are made of chilled cast-iron, manganese steel, or chrome-steel, chilled cast- iron being common. The wear upon shoes and dies amounts to 14 to 114 lb. per ton of ore crushed, depending upon the toughness of the ore. On an average these parts last three months. In the case of chilled cast-iron, % lb. at 6c. per lb. or 4.5c. per ton of ore treated is an average cost for wear. The tappets P, Fig. 44, of cast-iron are keyed to the stems. The stamps are lifted 8 to 16 in. by the cams H, and are adjustable to give them the drop desired. At the same time, they cause partial OF THE COMMON METALS. 125 rotation of the stem on its axis and insure an even wear of the shoe and die. As the toe of the cam passes on, the whole 'stamp' (shoe, head, stem, and tappet), drops with the impact of 1000 lb., falling freely upon the ore in the mortar. The replacing of a cam of the ordinary type, requiring the removal of the key that holds it, is a tedious operation, and to overcome this, a self-tightening cam like the Canda, Fig. 50, has been devised. This consists of a curved tap- ering key that fits an eccentric recess in the cam. Into the shaft is set a slightly projecting pin which engages the key in a groove or recess so that when cam and key are set upon the shaft over the pin and turned, the key wedges and ^tightens the cam to the shaft. The cams are set on the cam-shaft at various angles so that the stamps drop at regular intervals and in a pre-determined order. The stamps being numbered consecutively from left to right, in a 5-stamp bat- tery, a favorite order of drop is No. 1, 4 7 2, 5, 3. The cam-shaft is carried by boxes secured to the side-posts, and is driven by a large pulley made of wood in preference to iron, in order to withstand the shock of operation. When it is desired to stop the battery, the stamps are 'hung-up' by inserting under the tappets the 'fingers', one of which is seen in the elevation, Fig. 45. Ore is fed to the battery automatically by a feeder marked Z, in Fig. 44 and 45, that takes its supply from the sloping-bottom stor- age-bins above. As seen in Fig. 47 and 48, there is a ledge or lip on the mortar over which the pulp flows to the apron-plate. In the double-discharge mortar, the pulp-flow at the back of the mortar joins that at the front by passing through a hole or passage in the base of the mortar. 27. OPERATION OF THE STAMP-BATTERY. From the storage-bin, indicated behind the battery in Fig. 44, and shown in section in Fig. 45, the ore, crushed to l^-in. size, passes a chute to the hopper of the automatic-feeder. A feeder of the Chal- lenge type is shown in Fig. 51. At each drop of one of the stamps, a collar on the stem strikes the end of the projecting horizontal lever, and this lever actuates the feed-plate, revolving it slowly against a fixed scraper, and causing ore to fall into the feed-opening of the mortar. As the ore accumu- lates under the shoes, the stroke shortens, and the thrust of the hori- zontal lever being less the feed corresponding is lessened. As the ore supply lessens the stroke lengthens. Thus the action, according to the needs of the battery, is automatic. Water is introduced through a pipe at the feed-opening, and mixing with the pulverized ore, 126 THE METALLURGY splashes out through the screen at each fall of the stamps. The feed is regulated so as to cause the stamp to strike with a sharp, hard blow, but with little of the rebound that would occur with too thin a layer of ore. Mercury fed to the battery. This varies from 1 to 6 oz. per ounce of gold caught, the average being 1.5 oz. Added a little at a time in- side the mortar, it works out in part upon the apron-plates. As to the Fig. 51. AUTOMATIC FEEDER. amount to use, a safe guide is the appearance of the plates. If they are hard the indication is insufficient mercury ; if mercury is dis- tinctly visible on them, either in patches or in streaks, too much is being added. The mercury should be free from base metals that cause it to 'sicken', or break into coated globules. Such globules refuse either to coalesce or adhere to the amalgamated surface, and are swept away with the pulp. Mercury is best that already con- tains gold and silver. OF THE COMMON METALS. 127 Dressing the plates. The outside or apron-plates are dressed three or four times daily, the operation taking about 15 minutes. To do this, the feeding is stopped to permit the ore to work out of the mortar, the stamps are 'hung-up', and the apron-plate washed clean with a stream of water. A rubber-edged scraper, resembling, but heavier than a window-cleaner, is used to scrape the plate. The amalgam, perhaps half a pint in quantity, is scraped together with this, gathered up, and placed in an enameled cup. If the surface is too hard for the scraper, the amalgam is softened by sprinkling a little mercury upon it. Dressing being completed, the stamps are started, and feeding is resumed. Plates are liable to become tarnished with salts of copper that form a coating like verdigris upon the surface. Since the tarnished Fig. 52. MERCURY TRAP. spot collects no gold, the stains must be removed. To do this a solu- tion of sal-ammoniac is applied to the stained parts with a scrubbing brush when the battery is stopped. In a few minutes this chemical is washed off, and potassium cyanide and then mercury are rubbed on, and the plate immediately washed clean. Apron-plates have a grade of 0.5 to 1.75 in. per foot, ore containing sulphides requiring the steepest grade. When the pulp is flowing over the plate in a proper manner it travels down in the form of series of ripples or waves, bringing the particles of gold in contact with it. To save escaping particles of amalgam or globules of mercury that fail to adhere to the plate, a mercury trap, Fig. 52, is provided. This is especially important where no concentration of the tailing is attempted. In shape it is an inverted frustum of a pyramid. The pulp enters by the vertical pipe, and escapes over the wooden block shown belted to the side of the trap. The pulp overflows through the pipe at the side. Some heavy sulphides accumulate in this, but 128 THE METALLURGY the amalgam works down into the bottom. Occasionally the accumu- lation at the bottom of the trap is emptied through the plug-hole in the bottom, and is panned to recover the amalgam. The clean-up. In operating a 40-stamp mill, it is customary to remove the entire accumulation of amalgam from the mortar every Fig. 53. CLEAN-UP PAN. two to four weeks, and to thoroughly dress and scrape the plates. To do this, two batteries (10 stamps), at a time are hung up. The screens, inside plates, and dies, are taken out, and the contents of the mortar, perhaps two or three bucketfuls, is carefully scraped out and fed to the next batteries. The plates are then dressed, the dies and OF THE COMMON METALS. 129 screens returned to place, and the two batteries again started. The next two are taken in the same way, and so on. Finally the last batteries are hung up, the contents removed and put into a clean-up pan, Fig. 53, with the amalgam from the well-scraped plates. Three men can 'clean up' a 40-stamp mill in five to seven hours. The clean-up pan, Fig. 53, 3 ft. diam. by 3 ft. deep, making 12 to 15 r.p.m., is used for grinding the sand, pyrite, fragments of iron, and other substances, collected with the amalgam in the battery clean-up. The charge, of 300 lb., mixed with water, partly fills the pan. It is ground to a fine mud in 3 to 4 hours, during which time 50 Ib. mercury is added, and the mixing continued a few hours longer. The pulp is then diluted with water, the muddy portion de- canted by taking out a plug that is a little above the bottom. The residual mercury and amalgam, with some of the mud, is withdrawn through the lowest plug-hole, panned, treated in an enameled-ware bowl with nitric acid, and well washed until clean. The residual mercury and amalgam is strained through chamois skin, or through canvas, to remove the excess of mercury. Gold amalgam, when well squeezed through cloth, contains 35 to 45% gold. The mercury that has been removed by filtration still retains 0.5 per cent. Retorting. In the smaller gold-mills, amalgam is retorted in a pot-shaped retort, Fig. 54. In larger mills a horizontal cylindrical retort is used, though the latter is chiefly for silver-mills where a large quantity of amalgam is made. The retort, Fig. 54, is filled two- thirds full of amalgam, placed in a wind-furnace and supported by perforated cast-iron thimble that rests on the grate-bars. The cover is luted and tightly clamped. A pipe through which the mercury vapor passes leads out from the cover and turns downward. The outer leg of the pipe is water-cooled and the end dips into a tub of water. A fire having been started on the grate-bars, the retort gradually heats until mercury vapor begins to come over. It con- denses in drops in the cool pipe and collects in the tub below. When a distilling heat is obtained, the fire is checked and the retort kept at an even temperature for one or two hours, after which it is heated to redness to expel the last of the mercury. The mercury, collected by condensation in the tub under water, is used again. This accounts for the small loss, which averages in California practice but 0.5 oz. per ton of ore treated. Mercury is lost by 'flouring' and by 'sickening'. The first of these losses is in- dicated by a white appearance, and is caused by excessive agitation in the air, which breaks it into globules or particles so fine as not again to unite or at least, not without great difficulty. Sickened 130 THE METALLURGY mercury is black, and owes its appearance to the presence of base metals, as already explained. The retorted residue, still containing 0.5 to \% mercury, is porous, and consists of gold from 500 to 900 fine. It is melted in a plumbago crucible in a wind-furnace with soda and borax, and when it contains base metal, with a little nitre which serves to toughen it. The melt is poured into an ingot mold ; and the bar, cleaned from adhering slag, is shipped to the mint. 28. GENERAL ARRANGEMENT OF A GOLD MILL. Fig. 55 is a plan, and Fig. 56 an elevation of a 20-stamp mill, where ore is amalgamated and the tailing concentrated. OF THE COMMON METALS. 131 The ore enters the mill on a high-level track in tram-cars and is dumped into a gyratory crusher and broken to 1 to l^-in. size. From the crusher it goes by a shoot to either of two storage-bins for the two batteries of 10 stamps each. The crushing is done during the 10-hour day-shift, and the bins are large enough to hold a day's supply. From these bins the ore discharges into automatic feeders (not shown) , and from there, in constant supply to the stamp bat- Fig. 55. PLAN OF TWENTY-STAMP MILL. teries. The batteries are driven by a direct-coupled driving shaft supported upon the mud-sills of the batteries, as shown by IF, Fig. 44. The pulp, splashing through the battery screens, flows over a set of short apron-plates. The tailing from the plates unites in a launder, and finally falls into a distributing box that commands the tables. A distribution is made here and one-fourth is supplied by a launder, to each vanner. The tailing from the vanners is wasted. The concentrate is collected, and shipped for smelting. The method 132 THE METALLURGY of driving the machines is indicated in Fig. 56. Above the battery runs an over-head track, carrying a trolley and heavy chain tackle, by means of which a stamp or any heavy part can be removed readily or replaced. When ore contains heavy sulphides, such as pyrite, the tailing from the plates is concentrated on the Frue vanner, or the Wilfley concentrating-table. Thus a valuable product is obtained in a small bulk, and any escaping particles of amalgam are caught with the Fig. 56. ELEVATION OF TWENTY-STAMP MILL. concentrate. When ore contains gold, both coarsely and finely dis- seminated, the usual method has been to recover the coarse gold by amalgamation, and treat the residue by cyanidation to extract the gold not recovered by amalgamation. Another practice, dispensing with amalgamation, has been to crush with stamps, separate the slimed material, and re-crush the coarser portion (the sand), thereby finely grinding the auriferous particles so that they may be cyanided in a reasonable time, not possible with ore consisting of coarse auri- ferous particles. OF THE COMMON METALS. 133 29. CALIFORNIA AND COLORADO PRACTICE IN GOLD- MILLING. For free-milling ores, where the gold is coarse, and where in con- sequence fine grinding is not needed to release the gold from its matrix, California practice, with rapid drop and low discharge is preferred, since in this way large tonnage is secured. On the other hand in Gilpin county, Colorado, the ore contains 10% gold-bearing pyrite in which the gold is finely disseminated. To release the gold, the ore must be longer retained within the mortar and finely crushed. While retained there the fine float-gold, set free, has time to come in contact with the inside amalgamating plates. These recover about 75% of the gold. In California practice, the stamps weigh 900 to 1100 lb., drop 80 to 110 times per minute, and have a fall of 5 to 9 in. In Colorado practice the stamps, weighing 600 to 800 lb. r drop only 25 to 30 times per minute, but through a height of 18 to 20 in. In general, the greater number of drops per minute the greater is the tonnage. To pulverize the ore finer and retain it in the battery a longer time for inside amalgamation, a high discharge is needed, and this is ob- tained by using a deeper chuck-block, bringing the height in the Colorado type of battery to 13 to 18 in., while in California practice it is but 5 in. The objection to a high discharge is that more of the ore is slimed and is difficult to concentrate, causing a high loss of gold-bearing pyrite in the tailing. In California practice 4 tons per head may be crushed; in the Colorado practice it is but 1 to 1.5 tons. We thus have the following practices for average ore : Drops per minute California. 95 Colorado. 28 Height of drop (inches) 7 18 Height of discharge (inches) Weight of stamps (pounds) 5 1000 15 700 Actual horse-power per stamp Capacitv in 24 hours (tons) . 2.02 4 1.07 1.25 The theoretical horse-power of a single stamp is calculated by multiplying the weight in pounds by the distance lifted per minute, in feet, and dividing by 33,000. The actual horse-power may be computed as 1.2 times the theoretical. The harder and tougher the ore, the slower will be the crushing; while the softer, and more fri- able the ore and the coarser the product, the larger will be the ton- nage. 134 THE METALLURGY 30. COST OF GOLD-MILLING. In South Africa, where amalgamation has been followed by cy- anidation of the tailing, the cost of milling and amalgamation has been $0.72 to $1.20 per ton and $0.96 to $1.44 per ton additional for cyaniding. These costs are based upon an output of 4% to 5 tons per stamp. In California in 1896 at a 30-stamp mill there were crushed and concentrated 33,512 tons of ore at the following itemized cost per ton: Shoes and dies $0.029 Screens 0.003 Mercury 0.007 Hardware, belting, and firewood 0.021 Water for power 0.095 Freight, cyanide, oil, and grease 0.006 Lumber 0.008 Miscellaneous 0.007 Assay and office supplies 0.008 Silver-plated plates 0.007 Water-pipes and connections 0.021 Hauling sulphides 0.020 Express on bullion 0.006 Taxes and insurance 0.010 Superintendence and labor 0.160 $0.408 A summary shows that of this cost, $0.153 was for repairs, $0.16 for labor, and $0.095 for power. At a certain 30-stamp mill in Idaho, the cost was $0.41 per ton ; at a 40-stamp mill in California, $0.50, and at another $0.49 per ton. At a 40-stamp gold-mill in Gilpin county, Colorado, the cost was $0.84 and at another $1.47 per ton. Olcott gives the cost of milling at sev- eral California mills as varying from $0.20 to $0.75 per ton, and in a Gilpin county, Colorado, mill $0.95 per ton. 31. THE HYDROMETALLURGY OF GOLD. At the present time there are two methods by which gold is dis- solved from its ore by chemical solvents. In either process the first step is to obtain the gold in aqueous solution, then to precipitate it from the clear filtrate, and finally to get it in the form of a bar or ingot. OF THE COMMON METALS. 135 These two processes are: (1) The chlorination or Plattner process, by which the gold is obtained in solution as a chloride by the action of an aqueous solution of chlorine gas. (2) The cyanide or Mac Arthur-Forrest process in which the solution of the gold is effected by a weak cyanide solution, the dissolved gold then being present as potassium auro-cyanide. With certain refractory ores, the activity of the solution is greatly increased by the use of bromine or bromo-cyanogen in addition to the potassium cyanide. Extraction of gold by means of a solvent in aqueous solution is also practised where gold cannot be completely extracted by amal- gamation. This often is the case with pyrite ores; and extraction can be practised to advantage not only where amalgamation is un- suitable but where smelting is expensive. Gold in ore occurs in particles of various sizes, both as grains readily seen, and in particles of microscopic size. When the particles are visible, or when the ore shows 'colors' upon panning, the gold is called coarse, and such particles generally can be recovered by amalgamation. Gold often occurs in finely disseminated, microscopic particles, not visible to the eye, and in films on the surface of pyrite crystals. If the ore can be ground so fine as to unlock the crystals, or if it is permeable to solutions, gold can be dissolved in aqueous sol- vents, such as chlorine or potassium cyanide. Advantage is taken of the solubility of the released gold particles, and leaching or per- colation methods, in tanks or vats, are practised with this in view. The solution soaks through the ore 4 comes in contact with gold par- ticles, and dissolves them, or by another process, the finely-ground ore or slime is agitated with the solution, and the pulp is filtered and washed in filter-presses. The clear filtrate, in any case, is treated by a suitable precipitant to obtain the gold in small bulk, and the precipitated gold is melted and cast in the form of a bar or ingot for sale. There are thus three stages in any method of extracting gold by aqueous solvents: (1) The ore is finely ground, and when refrac- tory roasted, to convert the gold into a soluble form, and render it accessible to the solution. (2) The gold is extracted from the ore by means of a dilute solvent, using a tank with a filter-bottom, or agitating the ore, pulverized to a thin pulp, using a filter-press for the separation of the solution. (3) The gold in the solution is pre- cipitated (a), in chlorination by hydrogen sulphide or other precipi- tating agent or (b), in cyanidation by the use of zinc-shaving or dust. The precipitate is collected, dried, and melted into an ingot. The cyanidation has proved a remarkably cheap and efficient 136 THE METALLURGY method of extraction, but it has limitations, not only in respect to the solubility of the gold, but because of the interference of com- pounds that sometimes are present, notably those of copper, that interfere with extraction in various ways. The process has the advantage over the chlorination method in that silver, as well as gold, can be extracted. Under favorable conditions the extraction is high, and modern methods have reduced the cost of treatment to a low figure. Pyrite ore, exposed to the weather, becomes acid in reaction, and if treated by cyanide, decomposes and destroys the potassium cyan- ide. To correct this, ore is first treated by a wash of dilute caustic soda, or if acid, mixed with caustic lime in sufficient quantity to overcome acidity. When ore is refractory and requires preliminary roasting, the cost of roasting adds much to the cost of treatment. In chlorination, roasting is always necessary, and in any case it improves the condi- tion of the ore and makes it porous and permeable when leached or filter-pressed. 32. CHLORINATION OF GOLD ORES. This consists in attacking the gold of the ore with chlorine to form the soluble gold chloride, and dissolving* out the gold chloride in water. The complete process consists of the following parts: (1) Pre- paration. Crushing and roasting the ore. (2) Extraction. Bring- ing gold into the form of gold chloride, which is soluble in water, and leaching this out, obtaining a clear filtrate that contains the gold chloride. .(3) Segregation. Precipitating the gold from the fil- trate in metallic form, or as a sulphide and collecting and refining the precipitate to obtain the gold in the form of an ingot. Ores suited to chlorination. An ideal ore for chlorination is one in which the gold is present in a state of division, in which bases are absent that would be attacked by chlorine, and silver if present is in such a condition as not to coat the particles of gold with insoluble silver chloride. While the cyanide process is better for the treat- ment of low-grade ores, many refractory high-grade ores have given better results by chlorination. Ores, in which the gangue consists of hydrated iron-oxide, are extremely difficult to amalgamate. Not only is the gold finely di- vided, but the ore is slimy and forms a coating on the amalgamat- ing-plates. Such ores give satisfactory results by barrel-chlorina- tion. Silver is not recovered by chlorination since it becomes an in- OF THE COMMON METALS. 137 soluble silver chloride. If, however, sufficient silver be present to pay the increased cost, salt may be used in roasting and the silver extracted by means of sodium hyposulphite. A recovery of 60% of the silver is possible in this way. Below are analyses of ores that have been successfully treated by chlorination : Fig. 57. TABLE OF GOLD ORES. (1) (2) (3) (4) Delano mine, Eureka and Portland mine, Boulder Idaho mines, Cananea, Cripple Creek, county, Grass Valley, Mexico. Colo. Colo. Cal. Cu 0.10 0.85 Zn 0.78 Pb 0.78 Mn and Fe 3.40 4.15 6.00 40.65 S 0.80 2.49 2.20 32.80 SiO a 83.50 54.91 89.50 12.64 A1,0 S 3.20 17.80 0.19 MgO and CaO 2.40 0.25 5.83 Alkalis 12.00 Ag (oz. per ton) 1.35 0.50 1.25 2.00 All (oz. per ton) 1.23 1.00 0.65 8.00 Roasting. Oxidized ore at Mount Morgan, Western Australia, containing but a trace of sulphide, is crushed in Krupp ball-mills and quickly roasted (flash-roasted), in a cylindrical roaster like the White-Howell, Fig. 40, to dehydrate it and make it porous. Ore, containing sulphur, arsenic, and antimony, is crushed to 10 to 30-mesh size, and is roasted to expel these elements, to oxidize the bases, to leave the gold in such form as to be attacked by chlorine, and to make the ore porous, accessible to chlorine, and more easily leached. In attempting to chlorinate ore (1) of the table unroasted, the sulphur (0.8%), consumed chlorine, and an extraction of only 25% resulted. After dead-roasting, 98.4% of the gold was extracted. Cripple Creek ore (2) containing 2 to 3.5% sulphur was roasted to 0.08 to 0.10%, then cooled, and chlorinated. The extraction of gold was 92 to 95% of the amount present. The loss of gold in roast- ing, due to volatilization and dusting, is commonly 3 per cent. Ore (3) containing gold telluride, is broken by graded crushing (See Fig. 25), to 20-mesh size, and is roasted in a Pearce-turret roaster. (See Fig. 40 and 41). The ore, discharged from the furn- ace, passes to an automatic cooling-device, consisting of vertical tubes surrounded by water. The ore passes down the tubes, and is gradually removed in cooled condition at the bottom. Ore (4) is a concentrate from gold-milling, containing much sul- phide, and typical of California ore to which chlorination is applied. 138 THE METALLURGY The coarse gold has been removed, by milling and amalgamation, and the concentrate, generally 1.5 to 2% the weight of the ore milled, contains gold in fine particles. The concentrate is roasted generally in long-bedded reverberatory furnaces (Fig. 32), 60 ft. long, and 3 tons capacity per 24 hours. This ore contains copper, lead, lime, and magnesia, all of which consume chlorine, and form chlorides. To prevent this, it has been customary to add salt, to the extent of 0.75 to 1.5% of the charge, at or near the completion of the roast. If roasting has been thorough up to this time, copper is present as CuO, lead as PbS0 4 , lime as CaO, and magnesia as MgO. Were the copper present at the end as CuS0 4 it would react with the salt forming a chloride of copper. The common salt also reacts upon the gold and forms gold chloride. Both these chlorides are volatile, and the CuCl, in volatilizing, promotes the volatilization of the gold. Professor S. B. Christy, who conducted muffle roasting- tests on pyrite mixed with 5% of salt, found at a dull-red heat, 12% of the gold, and at a cherry-red 21%, to be lost. The losses in silver were somewhat less, being 7% and 17% under these conditions. As long as sulphur is present it protects the gold from attack, but when sulphates have been formed, and are causing the abundant evolution of chlorine by reaction with the salt, the escaping gas carries gold chloride, the gold being unprotected by sulphur at the time of chloridization. Lead sulphate similarly reacts with salt, forming lead chloride, which does not consume chlorine. When much lead is present, how- ever, it is removed by leaching with hot water before treating with chlorine. Lime and magnesia are converted by the salt into chlorides, and in this form consume no chlorine. The process of roasting is therefore conducted as follows : The ore is thoroughly roasted at a low-red heat. The tempera- ture is then a bright-red (850C.), to decompose copper sulphate. The salt is then added and thoroughly incorporated, and the tem- perature reduced to prevent volatilization of the gold. The quantity of salt to be added, the time needed for roasting, and the tempera- ture compatible with the minimum loss of the gold, should be de- termined experimentally for each kind of ore. 33. EXTRACTION OF THE GOLD BY CHLORINATION. There are two methods of chlorinating ore. These are (1) the vat, or Plattner process. (2) The barrel, or Theiss process. In the vat method the ore, after moistening, is charged into the vat and subjected to the action of chlorine gas conducted into it from a sepa- OF THE COMMON METALS. 139 rate vessel. No motive power is required, and, aside from the cost of the roasting-furnace, but small investment is involved in the plant. The process is suited to the treatment of a small daily supply of concentrate. In barrel chlorination, ore is charged into lead-lined steel barrels or cylinders, and there exposed to the action of chlorine generated from chemicals within the barrel itself. The barrel is rotated, bringing the gold intimately into contact with the chlorine, which acts powerfully upon the gold while in the nascent condition. Barrel chlorination requires a large initial investment in machinery and apparatus. It is, however, suited to a large tonnage, and is a less costly process and gives a high and rapid extraction. 34. THE VAT OR PLATTNER PROCESS OF CHLORINATION. The ore, which as above stated, has been subjected to an oxidiz- ing roast to free the gold and render the ore porous, is moistened and charged into a tank or vat 8 to 9 ft. diam. by 3 to 3.5 ft. deep. Fig. 58. CHLORINATION LEACHING VAT. This vat, Fig. 58, has a false-bottom of perforated 1-in. boards sup- ported on 1-in. strips above the bottom of the vat. Upon the perforated bottom is spread a layer of quartz first in pieces the size of eggs, then smaller toward the top. Above this is a 2-in. layer of sand, and on this is laid a canvas filter-cloth or an open 140 THE METALLURGY grating of boards, to protect the filter and give a surface on which to shovel. Chlorine gas is admitted through the pipe n, and the solu- tion is discharged through the hose b, into the launder c, which leads to the settling vats. When not in use the hose is turned upward in the position shown. The 4-ton charge of ore is carefully moistened with a suitable amount of water (6%). If too dry, the gold is not well acted upon by the chlorine ; if too w r et, the gas does not penetrate the ore in a suitable manner. A layer of dry ore is first scattered over the bot- tom to soak up the water left in the filter, that the gas may pass upward freely. The charge is then thrown into the vat through a %-in. screen, which breaks the lumps, and causes the ore to scatter loosely. When the vat has been filled a foot deep, the gas is intro- Fig. 59. CHLORINE GENERATOR. duced from below and begins to rise through the ore. Charging is continued and the vat is filled within 3 in. of the top, and burlap- sacking is spread over the surface. The cast-iron cover d, is then brought by an overhead crane to the vat, and lowered upon it ; and the joint between the cover and the top of the vat is made tight with clay-mortar and a strip of cloth. Chlorine gas, generated in a sepa- rate vessel, is allowed to enter the vat for 5 to 12 hours, according to the fineness of the gold. The finer the gold the faster is it chlor- inated. The charge is known to be sufficiently saturated with chlor- ine when, upon opening the stopper e, on the cover, fumes of escaping chlorine gas can be detected by the odor. After this the covered vat is allowed to stand 24 to 40 hours to chlorinate. OF THE COMMON METALS. 141 In presence of moisture the soluble tri-chloride of gold is formed as follows : H 2 O + Au + 3C1 = AuCl 3 H,O. The chlorine is produced in a generating vat (Fig. 59). This is of cast-iron, lead-lined, and has a heavy, tight cylindrical cover c. For a 4-ton charge of ore the vessel should be 24 in. diam. by 12 in. deep. It is charged by lifting the cover and putting in the solid chemicals, or adding these through the plug li. The chemicals con- sist of 20 to 27 Ib. dioxide of manganese, 27 to 32 Ib. common salt, and 40 to 60 Ib. sulphuric acid of 66 B. The acid is added through the funnel-tube i, and is followed by 24 to 33 Ib. water. The cover c, has a water seal ;', which prevents the escape of the gas. The generator stands upon a sand-bath q, or preferably on a steam-bath or coil, by which it is heated to 60C., the best temperature for gen- erating the chlorine. The reaction is as follows : 2NaCl + Mn0 2 + 2H 2 SO 4 = 2C1 + Na 2 S0 4 + MnS0 4 + 2H 2 0. The stirrer e, is turned by the handle g, from time to time. This revolves on a projecting pivot at the bottom and has an inverted cup /, soldered to the shaft that forms an air-tight water- joint with the corresponding socket filled with water like that of the cover at /. The plug d, serves to discharge the exhausted contents of the gen- erator. Through the delivery-pipe A 1 , and the horizontal tube I, the gas passes to the wash-bottle o, where hydrochloric acid, if present, is absorbed by the water contained in the barrel p. The gas, thus washed, passes on by the tube ra and n to the chlorinating tank, Fig. 58. The charge, having remained in the vat an average of 48 hours it is ready for leaching. For leaching, the cover is removed, and water from a hose is run in, and evenly distributed over the ore. As soon as the ore becomes saturated and covered with water, the solution is allowed to escape through the hose 6, which is lowered for that purpose. The level of the water is maintained at the top by a further supply until the escaping solution, on being tested, gives no reaction for gold. Two tons of water per ton of ore is used. The tailing or leached ore often contains silver, and this when in suffi- cient quantity, can be recovered later by hyposulphite-lixiviation. Thus at the Sierra Buttes mine, California, this tailing is leached 48 hours in vats 3.5 ft. diam. by 5 ft. high, with 3% solution of sodi- um-hyposulphite. The silver, precipitated from the filtrate by means of sodium-sulphide, is collected and sold. The gold solution from leaching is conducted through a wooden 142 THE METALLURGY launder, either to a settling vat from which, after settling for sev- eral hours, it passes to the precipitating vat, or through a filter-bag, and then at once to the precipitating vat. When obtained from an 8-oz. ore, it may contain in solution, 1% of various base-metal sul- phates, 0.9% metal-chlorides, 0.02% gold chloride, and 0.2% free chlorine. The precipitating tank is 6 ft. diam. by 3 ft. high, painted (as also are the other tanks), with hot asphalt, to which has been added Portland cement. The precipitant for gold is a solution of ferrous sulphate, prepared at the works by dissolving scrap-iron in sulphuric acid. It precipitates the gold as follows : 2AuCl 3 + 6FeS0 4 = 2Au + Fe 2 Cl ( . + 2Pe 2 (S0 4 ) 3 . To save time, the precipitant is added with the first of the gold solution entering the tank and enough more is stirred into the solu- tion to complete the precipitation. The vat is then covered, and the gold is allowed to settle 12 hours, or much longer. The clear super- natant solution is then run off, preferably through a filter-press (See Fig. 78), and the filtrate received into a sawdust filter to recover any particles of gold that escaped the press. Precipitation with ferrous sulphate has the disadvantage of being slow. The purple color, due to the presence of a trace of gold, can be detected in the solution days after precipitation. It is inferior to hydrogen sulphide in this respect. The residue in the tank is allowed to remain, and fresh solution in turn is run in and precipitated. The precipitate thus gradually accumulates. Finally, the precipitating tank is drained completely through the filter-press, and the residue is washed with dilute sulphuric acid to remove the ferric salt remaining. The moist product from the press is mixed with soda, borax, and nitre, and melted in graphite crucibles. The molten gold is poured into molds, and after removing the slag, is re-melted with a little borax to obtain a uniform bar. The gold thus obtained is 920 to 990 fine, the impurities being lead and iron. The extraction or recovery of the gold is 90 to 92 per cent. Cost of plant and treatment. The cost of erecting a plant in California, capable of treating 6 tons daily, is $6000 to $7000. In 1886 the cost of chlorination at the Providence mine, exclusive of supervision, interest, and depreciation, was $6.30 per ton. About one-half of this ($3), was the cost of roasting. This cost has since been considerably reduced by the use of oil fuel and by the em- ployment of mechanical furnaces of the Edwards type (See Fig. 39). OF THE COMMON METALS. 143 35. BARREL CHLORINATION. An example of an ore generally treated by chlorination is that of Cripple Creek: the ore (2) of the table. The ore contains gold tel- Fig. 60. SECTION THROUGH CRUSHING MILL AND ROASTER. luride, and must be roasted to release the gold from combination with tellurium, and to expel all sulphur above 0.1%. The complete process of barrel-ohlorination is as follows : 144 THE METALLURGY The ore is coarsely crushed in breakers and rolls to %-in. size or less and placed in storage-bins. The crushing is done during the day-shift. Thus an accident to either the coarse or the fine-crushing system need not stop the operation of the mill. The ore from any of the storage-bins is drawn off as needed to the feed-hopper of the dryer. It is dried and fine-crushed precisely as described in section 14, except that it need not be crushed finer than 10 to 16-mesh. By avoiding fine crushing, less dust is made, and after roasting, the product is pervious to the attack of the chlorine gas, and is more readily leached. Cripple Creek ore is roasted in a mechanical furnace, such as Fig.61. CHLORINATION BARREL. the Wethey, Fig. 36 and 37, or the Edwards, Fig. 38 and 39. The finishing temperature should not be higher than necessary to break up the sulphates formed in roasting. In operating the Wethey furn- ace, the lower hearth is used to cool the roasted ore. The cooling is effected in the Edwards furnace by surrounding the troughs of the discharge conveyor by water pipes. Ore thus cooled is raised by an elevator to storage-bins, whence it is drawn as needed to the chlorina- tion barrels. The ehlorination barrel. These barrels, as shown in perspective in Fig. 61, and in transverse and longitudinal section in Fig. 62, are rotated on a horizontal axis. They are supported on trunnions, and driven at 12 rev. per min. The shell d, is of sheet-steel with heavy 146 THE METALLURGY cast-iron ends c, and provided with two charging 1 doors or man- holes 6,6. The cylinder is lined with sheet lead % in. thick, bolted to the shell. It contains, as shown, a filter-frame, or diaphragm //, of hard wood, intended for filtering the clear solution after treat- ment with chlorine, leaving the exhausted ore behind in the barrel. The filter is made of blocks / that sustain a floor of lead plates % in., thick, perforated with %-in. holes. The plate itself is corrugated to allow circulation of the filtrate over it. Instead of these plates, a perforated 2-in. plank floor has been used. Upon the plate (or floor), rests a lead sheet of 4 Ib. per sq. ft. This sheet is perforated with 0.05-in. holes, % in. between centers. To hold down the filter- sheet, a wooden frame or grating is placed upon it. This is held by blocks //, and heavy strips i, securely bolted to the barrel. The wood frames beneath last three months; those above, but two or three weeks. This wood-work, if immersed in boiling tar or asphalt until thoroughly impregnated, lasts longer and absorbs but little solution. In place of lead filters, a woven asbestos cloth may be used. The cloth needs renewal after 50 to 60 charges have been treated. Bar- rels have been made 6 ft. diam. by 16 ft. long, with a capacity of 18 tons. The common size, however, is 6 ft. diam. by 12 ft. long, and the capacity is 8.5 tons. Charging the ore. Crushed, roasted, and cooled, the ore, in weighed charges, is conveyed from the storage-bins in two-wheeled buggies, and placed in the charging hoppers that belong to each cylinder (See the charging hoppers of the Bruckner roasting-f urn- ace, Fig. 35). Into the cylinder is first run 80 to 140 gal. water per ton of ore, enough to make with the ore, an easily flowing pulp. Next, a measured quantity of sulphuric acid is added, and then the charge of ore. Finally a weighed amount of 'bleach' or bleaching powder (CaCl 2 O) is added. The quantity of chemical to be added to the charge depends on the nature of the ore, and is determined by experiment. On roasted- Cripple Creek ore 12 to 15 Ib. bleaching powder of 34 to 36% available chlorine, and 24 to 30 Ib. sulphuric acid of 66B. is used per ton of ore. The charge-openings of the barrel are now closed. The barrel is started to slowly revolving (12 rev. per min.), for a period of 1 to 4 hours. In the case of Cripple Creek ore, for 3 hours. The chemicals enter into thorough contact with one another and with the ore, and react as follows : (1) 2CaOCl 2 + 2H 2 S0 4 == 4C1 + 2CaSO 4 + 2H 2 0. Chlorine, in nascent condition, acts with greater energy upon the OF THE COMMON METALS. 147 gold then chlorine formed outside the barrel. With the gold it forms a soluble gold chloride thus : (2) Au + 3C1 + H 2 O == AuCl,,.H 2 0. To determine when the ore is thoroughly saturated with chlo- rine, a stop-cock j is opened and gas that issues is tested for chlorine with ammonia which produces a white fume of NH 4 C1 with chlorine. If free chlorine is not detected, the barrel is stopped, opened, and more bleach and acid added. In place of using chloride of lime and sulphuric acid, as above described, chlorine (about 1 Ib. per ton of ore), has been introduced into the barrel in liquid form. Chlorine can be obtained in this form, in strong steel cylinders or drums. After saturation with chlorine the barrel is revolved an hour, then stopped in position for filtering with the filtering diaphragm down and level. The outlet pipe K, is connected by a hose to the settling tank and opened ; and water is pumped into the barrel above the charge through the valve /. The solution is now drained off, and the water above the charge forced rapidly through by means of compressed air introduced through' the valve j. The excess of chlo- rine is absorbed by the wash-water, and does not enter the building. The operation of filtering is suspended, connections are broken, valves closed, and the barrel revolved a few times to mix the con- tents again, and to break up channels that may have formed during the leaching. The barrel is then stopped, water run in, compressed air admitted, and the washing resumed. This is repeated until no gold is found in the escaping filtrate when tested. The compressed air admitted is under a pressure of 40 Ib. per sq. in. The time of filtering and washing on an average is 2% hours. The water used is 50% the weight of the ore. All connections are finally broken, valves closed, and man-holes opened, and the cylinder is revolved several times to discharge the contents. It is then washed out with a hose to prepare it for another charge. The solution from the barrels runs into lead-lined or asphalt- coated settling-tanks 10 ft. diam. by 7% ft. high. Here any sediment which escaped the filter is settled in about 8 hours. The clear sup- ernatant solution is withdrawn at a point 10 in. above the bottom of the settling-tank to avoid disturbing the sediment, and run into the gold-solution tanks on a lower level, where it is stored. The clear solution from the gold-solution tank is pumped through the opening .4, Fig. 63, into the precipitation tank X, which is 10 ft. diam. by 12 ft. high. The free chlorine is removed by passing sulphur dioxide through the solution from the SO 2 generator //. This generator 148 THE METALLURGY is a cast-iron receptacle containing a pan F in which sulphur is burned. Compressed air, admitted from a pipe W, supplies oxygen to burn the sulphur to S0 2 and pressure to drive the fumes into the solution along the pipes o,v and the lead pipe r, through numerous v- Fig. 63. PRECIPITATION APPARATUS OF BARREL CHLORINATION. small holes where the horizontal part of the pipe crosses the bottom of the tank. The sulphur dioxide acts upon the chlorine as follows: (3) C1 2 + SO 2 + 2H 2 O == H,SO 4 + 2HC1. Sulphuric and hydrochloric acids are formed by the action of sulphurous acid upon chlorine, and both remain in aqueous solution. The operation is completed in 15 to 20 minutes as indicated by test- ing the solution with H 2 S, when a permanent precipitate forms. The chlorine having been removed, we are ready to precipitate the gold by passing in H 2 S. To do this, the pipe v is connected to the lead-lined generator G which contains lumps of iron sulphide resting OF THE COMMON METALS. 149 on a perforated lead false-bottom. Dilute sulphuric acid admitted below the false-bottom comes in contact with the iron sulphide and abundantly generates H 2 S according to the equation : (4) FeS + H 2 SO 4 =* FeS0 4 + H 2 S. Compressed air, entering by the pipe w and the valve c, drives the gas through the pipe v and the lead pipe r into the solution in the tank, and precipitates the gold as follows : (5) 2AuCl 3 + 3H 2 S = Au 2 S 3 + 6HC1. The gold is thus thrown down as an auric sulphide in a solution containing both sulphuric and hydrochloric acids. The precipitation is rapid, taking about 10 minutes, and, it is possible by selective pre- cipitation and careful working to leave copper in solution while pre- cipitating the gold. Where the chemicals in the generator become exhausted, the waste liquid is discharged into the waste-launder p. The chemicals consumed per ton of roasted ore are, 1 Ib. iron sulph- ide, ^4 MX sulphur, and 2% Ib. sulphuric acid. More recent treatment does not include the use of S0 2 gas, H 2 S gas alone being used. At first, H 2 S is oxidized by the chlorine, thus: (6) H 2 S + 8C1 + 4H 2 = H 2 S0 4 + 8HC1. Sulphuric and hydrochloric acids are formed by this reaction after which auric sulphide is precipitated as in equation (4). After being precipitated, Au 2 S 3 is allowed to settle two hours. The clear solution then is drawn off at C, 10 in. above the bottom of the tank, through the pipe n into the filter-press, where it is filtered under its own hydrostatic head of 25 ft. This is done to recover any possible flakes of gold sulphide that failed to settle in the tank x. In three or four hours after precipitation, the tank can receive a fresh charge of gold-bearing solution. The precipitate collects upon the bottom of the tank, and after several charges have been treated, the united precipitate is drawn off at D and delivered through the man-hole L into the pressure-tank 2, by the hose y. The precipitating tank is then washed clean with the aid of a hose. The pressure-tank is 4 ft. diam. by 4% ft. high. When charged, the cover L is clamped in place, and compressed air, under a pressure of 40 Ib. per sq. in., is admitted through t. At the same time connection is made to the filter-press T through the pipe u, and the precipitate collects in the press under the above pressure. This filter-press is more clearly illustrated in Fig. 78. A set of filter- frames outlasts the treatment of 6000 to 8000 tons of ore. The filt- rate from the press passes over a sawdust filter-bed, as a safeguard, before it is run to waste. The sawdust is collected occassionally, and 150 THE METALLURGY burned, to recover the small amount of gold which it may have caught. The precipitate of gold sulphide also contains sulphur, and sul- phides of arsenic, antimony, copper, and silver, forming a 'sulphide cake'. The press is next opened and the cake is withdrawn. While still moist it is placed in trays 44 in. long, 24 in. wide, 4 in. deep, and mixed with borax, nitre, and soda. The trays are then placed in cast-iron muffles that are heated by coal, and connected to a flue- chamber, where any mechanically escaping flue-dust is caught. The precipitate dries here, and after drying, the temperature is raised to decompose and expel the sulphides, the operation taking an hour. The roasted material has a light-brown color, and contains 70 to 80% of gold. It is carefully transferred to a crucible and melted in a wind-furnace. The content of the crucible, including the slag, is poured into a conical mold where the gold collects at the bottom. Upon cooling, the gold ingot, 900 to 950 fine, is separated from the slag, remelted, and cast into a bar suitable to be forwarded to the United States mint. After deducting a charge of 2c. per ounce for melting and assaying, the gold at the mint should net $20.65 per ounce of contained gold. The slag, from the melting operation, may be melted again with one-seventh its weight of litharge with a reducing agent, and a lead button obtained. This lead can be scorified several times, until it is reduced to a small button, and finally cupelled to recover the small amount of additional gold. The cost per ton of treating Cripple Creek ores, on a large scale, by barrel chlorination, is as follows : Labor, including salaries $1.34 Chemicals and supplies 0.72 Fuel, roasting and power 0.70 Renewals and repairs 0.45 General expense 0.32 Total cost per ton $3.53 The cost of sulphuric acid, 66B., is 0.9 to l.lc. per Ib. ; bleaching powder 1.8c. (New York) ; sulphide of iron 3c. (this can be made at the works from scrap wrought-iron and sulphur) ; sulphur 2 cents. The water needed is approximately two tons per ton of ore treated, which includes that used for power. This quantity can be reduced if settling tanks are employed and the water used again. OF THE COMMON METALS. 151 Following is a flow-sheet indicating the procedure in the barrel- chlorination process above described. MlNEORE I CRUSHING PLANT vSft FIG 2) CRUSHED RAW ORE ^ | STORAGE BINS [ | BELT ELEVATOR) 3 EXTRACTION - ^ (FEED HOPPER | X [ROASTING FURNACE | RO/\ST!D CRE ^ [COOLING HEARTH& FLOOR] X ROASTED Cooueo ORE [STORAGE SINS FLUX e INGOTS -To MARKET -SOLUTION Fig. 64. FLOW-SHEET OF BARREL CHLORINATION. 36. CYANIDATION OF GOLD (AND SILVER) ORES. The process of cyanidation consists in attacking the gold or silver contained in ores with dilute solutions containing 0.5 to 0.7% of potassium cyanide. If the ore contains pyrite it may be acid as the result of weathering. Before applying potassium cyanide this acid- ity must be corrected by adding caustic soda or quick-lime. The gold-bearing cyanide solution is filtered off, and the precious metals contained are precipitated upon zinc shavings, zinc dust, or by elec- trolysis. The precipitate is collected, treated to purify it, melted, and cast into bars or ingots of metal. The solution, free from gold, called 'barren solution', is used again. The cyanide precess is a success with many ores. The field of its usefulness is extending, not only in the variety of ores, but ores con- taining smaller and smaller amounts of metal are being treated. An important advantage of cyanidation over chlorination is that roast- ing by no means is essential though sulphides be present. Sulphides containing silver, however, must be small in quantity, if silver is to be extracted in a period of time compatible with economy. Dilute cyanide solution has a selective action in extraction and can dissolve gold and silver from an ore, at the same time attacking but little the base-metals that are present. If, however, copper is present in soluble form it consumes cyanide, though in modern prac- tice a small quantity is not considered a serious obstacle. Ores suited to cyanidation. With reference chiefly to gold, we have the following classes of ores : 152 THE METALLURGY (1) Free-milling ore. Those in which gold occurs in a fine or microscopic form, and in which panning reveals to the naked eye few or no visible colors of gold. Where gold particles are coarse the time required to dissolve them is long, and cyanidation becomes im- practicable. It is possible, however, to remove coarse gold by am- algamation,, and after so doing the residue can be treated by cyanida- tion. (2) Telluride ores. Before these ores are cyanided they must be treated to liberate the gold from its combination with tellurium. When rich and containing telluride in spots, roasting develops shots or particles of gold, and these must be removed by amalgamation or concentration before cyaniding. (3) Pyrite ores. The gold in such ores is supposed to occur in the form of films on the surface of the pyrite crystals. When these are crushed fine to expose fresh faces, the gold can be dissolved by cyanide. Some pyrite ores, however, contain the gold within the substance of the crystals, and these ores must be roasted before cyaniding. Eoasting improves ore by rendering it more porous, and hence more accessible to the solution. It destroys the colloid of slime, and makes a material more easily filtered. (4) Talcose or clayey ores. These, when crushed for cyaniding, produce much slime that is impenetrable by solution, and can not be leached. Slime is treated in a way that adds to the cost of extrac- tion. Roasting, as stated, would accomplish this end, but the cost would be prohibitory in some cases. 37. DEVELOPMENT OF THE CYANIDE PROCESS. This process, originally called the MacArthur-Forrest process, was patented and vigorously advanced by the inventors. It was first put in practice at the Robinson mine on the Rand, South Africa, and applied to the recovery of gold in large tailing-dumps, that had ac- cumulated in stamp-milling and had been impounded behind crib- work, or low dams. This tailing was leached in filter-bottom vats with cyanide solutions. The gold-bearing filtrate was passed through zinc-boxes, containing zinc shavings, to precipitate the gold, and the precipitate was purified and the gold collected and cast in ingots. So long as these dumps lasted the simple process was sufficient, but when it became necessary to treat the tailing immediately from the mill, another kind of practice was begun. This 'second method' in outline is as follows: The ore is crushed and amalgamated in the customary way. The tailing from the amalgamated plate is classi- OF THE COMMON METALS. 153 fied into two products, one of which consists of coarse sand, the other of slime and fine sand. The coarse sand is more readily leached when freed from slime. To correct the acidity, milk of lime is added to the finer portion as it passes through the launder to the leaching-vats. The sand here settles and the slime passes on, suspended in the escap- ing water. When the tank is filled it is drained, and the contained water is replaced by weak cyanide solution. Discharge-valves or doors (See Fig. 68), are opened and the contents of the vat are shoveled into a vat directly beneath. A solution of 0.15 to 0.25% cyanide is allowed to percolate some days through the sand, and then is displaced by a weaker solution, and the weaker one by water. The gold in the gold-bearing filtrate is precipitated on zinc shaving, and the solution, after removing the gold, is used again. This is called 'double treatment', since the extraction of gold, begun in one vat. is completed in a second one. While a little more expensive in equip- ment and in handling, it is more thorough. The coarse sand requires much longer time to leach than the fine sand, and hence is treated separately. The slime is caught in large tanks and allowed to settle. After settling, the clear supernatant water is drawn off. A 0.01 to 0.02% solution of cyanide is added to the thickened slime remaining in the tank, and the whole is agitated by a mechanical stirrer for some hours. It is then settled, and the clear gold-bearing solution is drawn into a gold-solution tank, from which it is supplied in a regulated stream to the zinc-boxes for precipitation. The residual slime is treated again with weak solution, settled, and decanted. The spent slime then is washed out to the dump. This method of treatment is slow, and requires a large number oi tanks ; and gold is carried away in the solution that remains in the slime. To overcome these difficulties filter-pressing has been tried. The slime, after agitating as above described, with cyanide solution, is pumped through a filter-press (Fig. 78), where it is separated and washed, and the filtrate is collected in the gold-solution tank, while the exhausted slime removed from the press is thrown away. The trouble with filter-pressing, as ordinarily practiced, is that it is ex- pensive and slow. A number of costly presses are required for a large plant and these consume power and require labor to operate them. In the United States this has led to the adoption of suction- filters of the Moore type. In South Africa, Western Australia, and elsewhere, the Butters and Ridgway filters have given good service, while pressure-tank filters of the Blaisdell type, that use filter-leaves in the name general way* have advocates. 154 THE METALLURGY Western Australian ores require fine grinding, to liberate the gold. While fine material cannot be leached in vats, it can be filter- pressed. This has led to the use of the grinding pan as in silver milling, and to the tube-mill in particular, which has proved particu- larly well adopted to fine grinding. The outline of the process as practised in Western Australia is as follows : The ore is coarsely crushed in rock-breakers, then dry-crushed either in a Griffin mill, a ball mill, or rolls. Western Australian ore containing tellurides must be roasted to decompose and expel the tellurium, and liberate the gold for attack by cyanide solution. By roasting, the ore becomes porous, and colloidal compounds are de- stroyed, yet it is necessary to grind fine if a high extraction is to be obtained. The pulverizing has been done in grinding-pans or tube- mills, in the presence of cyanide solution. The product is sent to a settler to remove large particles of gold, if present, and all the pulp is then filter-pressed, and washed in the press, the process giving an extraction of 93 per cent. The trend of modern practice is toward fine grinding, and treat- ment of the whole product as slime, under the idea that the fine re- grinding of the product insures pulverizing the coarse gold, so that it can be cyanided in reasonable time. The alternative process to this is to finely grind the product, classify it into sand and slime, and treat the sand by percolation and the slime by pressure of suction filtration. 38. THE CHEMISTRY OF THE CYANIDE PROCESS. When a solution containing 0.5% potassum cyanide is brought into contact with finely-ground ore containing gold in microscopic particles, the gold gradually goes into solution. The reaction that occurs was first shown by Eisner, and is called Eisner's equation. It is as follows : 2Au + 4KCN + O + H 2 = 2AuK(CN) 2 + 2KOH The gold is dissolved by the action of potassium cyanide in the pres- ence of oxygen and water; the compounds formed are auro-potas- sium cyanide and caustic potash. Silver, but more slowly, dissolves according to a similar reaction, thus : 2Ag + 4KCN + + H 2 = 2AgK(CN) 2 + 2KOH. . It is noticed that oxygen is needed to fulfil the requirements of the reaction, and consequently ore, or solution acting on ore, must be in some manner aerated. When oxygen of the dissolved air is con- sumed, action ceases, but it proceeds again with access of fresh air. OF THE COMMON METALS. 155 Oxidizing agents like potassium^ chloride- and permanganate, and the peroxides of lead, manganese, and sodium may be used to furnish oxygen, in place of air, but are too expensive for practical use. When an ore containing pyrite is exposed to the weather, air and moisture slowly act upon the pyrite, according to the reaction : 3FeS 2 + 2H 2 O + 22O = FeS0 4 + Fe 2 (SO 4 ) 3 + 2H 2 SO 4 Ferrous and ferric sulphates and sulphuric acid are thus formed. The two first named would tend to precipitate gold. Ferric sul- phate is acid in its reaction, and with the sulphuric acid, if un- neutralizecl it would decompose and cause a serious loss of potassium cyanide. Such compounds are called 'cyanicides'. Ore, therefore, that contains pyrite, and that has lain exposed to the weather, needs caustic soda or quick-lime Ca(OH) 2 to neutralize the acidity, and an excess to provide for the acid resulting from further decomposition. This excess is the 'protective alkalinity' as it is called. To remove the soluble ferrous sulphate and sulphuric acid, a water-wash before treatment would be sufficient. This would require much time and in place of it quick-lime is added. Quick-lime acts upon the products of the above reaction as follows : FeSO 4 + Ca(OH) 2 = Fe(OH)^+ CaS0 4 Fe 2 (SO 4 ) 3 + 3Ca(OH) 2 = 2Fe(OH) 3 + 3CaS0 4 H 2 S0 4 + Ca(OH) 2 = 2H 2 + CaS0 4 The reaction causes the formation of a harmless iron hydroxide and calcium sulphate. The reaction that occurs when gold-bearing solution comes in contact with zinc shaving in the precipitating boxes, or when 'zinc dust' is stirred into such a solution, is generally assumed to be a simple replacement of zinc by gold, thus : 2KAu(CN) 2 + Zn = K 2 Zn(CN) 4 + 2Au The gold falls as a brown or black precipitate, while the zinc potas- sium cyanide remains in solution. The barren cyanide solution from which gold has been removed is used repeatedly, and accumulates impurity from the ore, and from the zinc with which it was in contact in the zinc-boxes. In conse- quence it becomes gradually less efficient than fresh solution. It has been found that the addition of quick-lime increases the solvent ac- tion of cyanide solution upon pure quartz ore, but it is without effect on ore containing sulphide. Such solution, if treated with sodium sulphide to the point of exact neutrality, and with a small excess of lead acetate, and given time to permit the resultant sulphide to pre- cipitate, is improved in extractive power thus : 156 THE METALLURGY K 2 Zn(CN) 4 -f Na 2 S = K 2 Na 2 (CN) 4 + ZnS The cyanide is here regenerated, while the zinc sulphide separates. This is a means of overcoming the accumulation of zinc in solution which is one of the drawbacks to the use of zinc for precipitation compared with electrical deposition. We are not, however, altogether dependent upon chemicals to dis- pose of the zinc. We have shown by Eisner's equation that caustic potash is set free. Caustic potash reacts upon sulphide contained in the ore and causes the formation of a soluble sulphide which in turn reacts like the sodium sulphide in the reaction above, and precipi- tates zinc sulphide upon the ore. The following minerals and chemical compounds destroy or com- bine with cyanide, and render it incapable of dissolving gold : Copper in the form of sulphate, carbonate, copper glance, erubes- cite, or copper pyrite. The sulph-antimonites of copper are without action. Manganese as 'wad' (impure hydrous oxide), but not the carbonate or oxide. Zinc as calamine, but not blende or zinc silicate. A distinct loss is traceable to the presence of organic matter like grass-roots, decayed wood, etc. To increase the activity of zinc shaving in precipitating of gold it may be dipped in a 10% solution of lead acetate just before it is placed in the zinc-boxes. Lead is precipitated upon the surface of the shavings, forming zinc-lead couples, which electrically react upon the gold in solution. 39. CLASSIFICATION OF CYANIDATION METHODS. First method. For oxidized ore which forms but a moderate amount of slime when crushed and can be leached, and in which the contained gold is fine and free. Example : Mercur, Utah. Second method. For ore that crushes with the formation of slime and contains both coarse and fine gold (coarse gold can be recovered by amalgamation), and pyrite and other sulphides, that can be re- moved by concentration. The tailing is separated into sand and slime. The sand is treated by leaching. The slime is either rejected or treated by: (a) Decantation. Best example: South Africa, (b) Filter-pressing. Best example : Homestead mine, South Dakota. (c) Vacuum-filtration. Best examples: Terry, South Dakota, Lib- erty Bell mine, Colorado ; Goldfield, and Virginia City, Nevada. Third method. For ore that contains both coarse and fine gold. The coarse gold is recovered by amalgamation, and a large part or the whole of the tailing from the plates, is re-ground or slimed in pans or tube-mills. The ground product is agitated and treated by OF THE COMMON METALS. 157 decantation, filter-pressing, or vacuum-filtration. Concentration is omitted. Best example : El Oro, Mexico. Fourth method. For ore that contains fine gold only. The crushed ore is classified into sand and slime, the former being leached. Concentration is omitted. A feature of the process is that the ore is crushed in weak cyanide solution. Best example : Mait- land mill, South Dakota. Fifth method. For ore that contains gold in combination* with tellurium. The ore receives an oxidizing roast after which it can be treated by the 'first method'. As the result of roasting some of the gold may take the form of shot, beads, or coarse particles of gold that settle and must be caught by a classifier, riffle, sluice, concentrat- ing table, or amalgamated plate. If ore is re-crushed in tube-mills, the shot or coarse gold becomes ground so fine it all may dissolve in the cyanide solution. Best examples : Cripple Creek and Western Australia. 40. FIRST METHOD OF CYANIDATION. This was the first method put in practice for the treatment of gold ores. It is applicable particularly to free-milling gold ores, in which the gold is in a free or uncombined state. Also, since the cyanide solution acts but slowly upon gold, it is necessary when using this method, that the gold particles be extremely fine; otherwise it re-- quires a long time to leach the ore. When coarse gold is present it it first caught on amalgamated plates. The ore is crushed preferably by rolls to avoid the formation of slime, which would interfere with the leaching. The crushed ore is leached in vats with weak cyanide solution, and the gold is precipitated from the filtered solution by passing it through boxes, treated, and obtained in the form of gold ingots. Method of crushing. A comparison between stamps and rolls shows the former to be durable, less expensive to install, and capable of finely crushing ore from a rock-breaker. Where stamps are used the ore is wet and four to eight tons of water is needed per ton of ore. They make a greater proportion of extremely fine material or slime in crushing than rolls. Where ore contains talc or clay, that would cause it to slime, rolls give a more granular product, and one that is more easily percolated or leached by the cyanide solution. With good rolls, an efficient system of graded crushing can be used and dry crushing is preferred for the following reasons : (1) The ore after crushing is delivered dry to the leaching- 158 THE METALLURGY tanks, and there is no dilution of solution as when treating a wet- crushed ore which contains 10 to 20% moisture. (2) In dry-crushing water is not needed. This is an important consideration in a dry country where water is scarce. (3) A uniform bed of easily percolated ore can be put in the OF THE COMMON METALS. 159 leaching-tank, the ore is better aerated, and the oxygen present as- sists in the solution of the gold, as shown in Eisner's reaction. On the other hand the cost of dry crushing beyond 40-mesh is excessive. Leaching tailing. Fig. 65 and 66 represent, in sectional elevation 160 THE METALLURGY and plan, a 75-ton plant designed to treat impounded tailing. which has accumulated from a mill treating gold ore of a kind suited to this method. The tailing contains so little slime that only 37% passes an 80-mesh sieve; and the gold can be extracted by four days leaching. Referring to plan. Fig. 66, &, b, &, &, are the leaching vats (some- times called percolators) the most important units of the plant, and the ones to which the others are accessory. They are 20 ft. diam. Fig. 67. WOODEN LEACHING VAT SHOWING FALSE BOTTOM. by 6% ft. deep, having 75 tons capacity above the filter-bottom. The four together allow four days treatment for each charge. They are constructed with filter bottoms, as shown in Fig. 67 and in perspective Fig. 68. They may be made of wood or steel. In warm countries, like Australia, South Africa, or Mexico, the steel vat is to be preferred ; but in cold countries, where it is necessary to house the plant, the wooden vat gives satisfaction. The latter is cheaper in first cost and easier to set up, though the wood absorbs gold solution. The steel vat on the other hand is less liable to leak, but costs more to maintain in requiring a OF THE COMMON METALS. 161 periodical coat of protective asphalt paint to preserve the steel from the action of the cyanide solution. A wooJcn tank, a!so 7 \\h_i requiring it, should be painted. In Western America wooden vats predominate. They are carefully built by companies who make this their business. Steel vats are also used. Each kind has advocates, but in general the steel vat is preferred. Fig. 68 is a perspective view of a wooden vat with the filter- cloth omitted. This shows the false-bottom of slats and the hinged bottom-discharge opening through which the exhausted tailing is shoveled or sluiced out. A wooden ring, 2 in. high and 2 1 /-> thick, Fig. 68. PERSPECTIVE VIEW OF WOODEN LEACHING VAT. is nailed to the bottom of the vat, leaving a space of % in. between it and the side. Parallel strips, one inch high, are nailed a foot apart upon the bottom, and across these 1 by 4 in. strips or slats are laid with 1-in. space between. Upon this false-bottom cocoa- matting is spread, and over it 8-oz. canvas filter-cloth cut 12 in. larger in diameter than the vat. The edges of the cloth are held down by a rope laid upon the canvas and driven into the %-in. space between the staves and the wooden ring. This way of securing the canvas is shown in further detail in Fig. 71. Fig. 69 gives, in plan and section, the construction of another form of filter-bottom for a wooden leaching-vat. Wooden rings of graded heights are nailed to the bottom of the vat and upon these are nailed radial pieces bored with %-in. holes. This arrangement of a sloping bottom facilitates the discharge of tailing when sluiced 162 THE METALLURGY or hosed out of the vat. The wooden ring, in Fig. 67 is omitted and instead, the joint is made by nailing a wooden strip around the tank so as to hold the canvas tight against the staves. Fig. 70 and 71 represent, in plan and in elevation respectively, the construction of a steel vat having a perforated board bottom. Fig. 69. PLAN AND SECTION OF WOODEN LEACHING VAT. A ring of flat iron % by 2 l / 2 in. is rivetted to the side of the tank, with space-thimbles to hold it % in. from the side. The cleats that sustain the false-bottom are 2 in. high by I 1 /) in. wide. The 1-in. bottom-boards are bored with %-in. holes and screwed to the cleats. As in the case of the wooden vats, the thick stiff cocoa-matting is laid upon the false-bottom, and covered with a filter-cloth of 8 to OF THE COMMON METALS. 163 10-oz. canvas. The edges are calked with a %-in. rope into the %-in. space, as shown at g in the sectional view, Fig. 71. (b) Fig. 70. PLAN OF STEEL LEACHING VAT. Leaching vats vary in size from 16 to 50 ft. in diam. and 4 to 9 ft. in depth. The shallow ones are for the more finely-ground sand. The tank is generally made to hold one day's supply of ore to insure Fig. 71. SECTION OF STEEL LEACHING VAT. uniform work in the mill. Hence the number of tanks indicates the number of days of treatment. The exhausted sand of the vat may be shoveled out through side or bottom openings, or, where water is abundant, may be washed by 164 THE METALLURGY a hose into a launder as shown in s Fig. 65, and thus conveyed to the dump. Fig. 72 represents a side-discharge door which is bolted to the outside of the vat, and which can be quickly opened or closed. The joint between the door and frame is made tight with a rubber gasket. Fig. 73 is a bottom-discharge valve. Opening upward, it is conveniently operated from the platform above the tank when the charge is to be sluiced out. It is securely bolted to the bottom, and is self-sustaining. For a large tank four such valves may be used. The opening in each is 10 in. diam. In Fig. 68 is shown a drop-bottom door that opens from below. At Z Fig. 65, is the charging platform or bridge (not shown in Fig. 66), grated where it is over the vats, with 2-in. openings. The Fig. 72. SIDE DISCHARGE DOOR. tailing is brought by a wagon having movable bottom-planks, and dumped upon the grating from which it falls lightly into the vat. When full the surface is leveled with a hoe. At y is a platform set 41/2 ft. below the top of the vat for convenience in leveling the ore and in adjusting the supply-valves. At a and a' Fig. 66 are the strong and weak-solution stock-tanks, one of which contains 5 Ib. potassium cyanide solution, the other a weak solution, 0.1%, or 2 Ib. per ton. These tanks deliver solution to the leaching-vats by pipe c for the strong and by c/ for the weak solution. The filtrate from the vats goes by a double launder, d for the strong and d' for the weak solution. These deliver by the cross-launder j and /' to the strong gold-solution tank h and the weak gold-solution tank li' '. Charging the vat. Where tailing has been accumulated in dumps, it is coarser at the point of discharge from the mill and OF THE COMMON METALS. 165 finer toward the further limit, and to treat the accumulation, the two must be mixed. They may be mixed in the reservoir by plowing and breaking up the lumps with a disk-harrow. When the material has become pulverized and dry it is scraped into heaps. The coarse and fine are together shoveled into a wagon, and the mixing is completed when the tailing is dropped from the wagon, through the grating, into the leaching vat. The material should be loaded evenly into the vat and should fill it, so that when settled by the wash-water, Fig. 73. BOTTOM DISCHARGE VALVE. the level will be 10 in. below the top. This gives room for the wash-water and cyanide solutions. Since the tailing is acid and the acid would destroy potassium cyanide, it is necessary to add a certain proportion of slaked quick-lime when charging into the vat. Leaching the ore. A drop-pipe from the strong-solution pipe c extends just below the filter bottom; and by this a 0.25% KCN solution is introduced. It rises through the ore and is shut off when the ore is covered 2 or 3 in. deep. The charge is allowed to remain in this condition a fixed time after which the discharge- valve at the bottom of the tank is opened and the solution, at first weakened by the water it has taken from the charge, is run to the 166 THE METALLURGY weak-solution gold-tank. The strength steadily increases and when above 0.1% (the grade of weak solution) the solution is turned into the strong-solution gold-tank. Percolation now proceeds, solution being added above the surface of the charge in a succession of washes between each of which the OF THE COMMON METALS. 167 solution is allowed to sink beneath the surface, to draw air into the charge. Weak solution is now admitted above the charge to displace strong, and as the effluent weakens it is turned into the weak-solution tank taking care to prevent a needless accumulation of Aveak-solution, by using as little as possible. Displacement of one solution of the charge by another does not take place uniformly ; the compactness of the charge is irregular and the more permeable parts allow the quicker advance of the solution last added, causing it to mix more or less with that preceed- ing. It has been found, in fact, that of weak solution l 1 /^ to 2 times the theoretical quantity is needed to displace the strong. Different charges percolate at different rates, and no hard-and-fast rule can be established for handling tailing. Each charge should Fig. In. PERSPECTIVE VIEW OF ZINC BOXES. be studied by itself, and frequent samples of the out-going solution must be chiefly relied on to make sure of proper extraction. The cycle of treatment of a 75-ton charge is as follows : Hours. Charging, or filling the vat with tailing 9 Leveling charge and saturating with strong solution 3 Percolation until effluent becomes 0.18% KCN. 20 Continued percolation with strong solution ... 20 Displacing strong solution with weak 29 Displacing weak solution with water 13 Sluicing out the exhausted tailing 3 Total time (4 days 1 hour) 97 The solution from the gold-solution tanks enters the zinc or 168 THE METALLURGY extractor-boxes of which three receive the strong and two the weak solution. Fig. 74 gives a plan and elevation of a zinc-box, and Fig. 75 a perspective view. As shown, each set of boxes contains seven compartments, each 12 by 15 by 24 inches in size. The compartments have perforated false-bottoms of sheet-iron or wire-cloth that sustain the zinc-shaving with which they are filled. The partitions are set alternately up and down, to compel an upward flow of the gold-bearing solution through the shaving, and to bring it intimately in contact with the zinc to insure the precipitation of the gold upon the surface of the zinc. At a in the side elevation (Fig. 74) the solution enters the box through a pipe indicated at the left, passes through all the compartments, flows over the last partition, and discharges through a down-turned pipe into the sump-tank. The boxes are set at a grade of % in. to the foot. Precipitation. The strong and weak solutions, leaving the leaching vats, are gathered in the respective gold-tanks. These Fig. 76. FLOATING HOSE FOR GOLD SOLUTION TANK. tanks accumulate the gold-bearing solutions from the leaching-vats, and provide a uniform discharge into the zinc-boxes. Since the solutions contain a little flocculent precipitate, which if allowed to enter the zinc-boxes would seriously interfere with the precipita- tion of the gold on the zinc, they are allowed to settle while the clear supernatant solution is drawn from the tank by a 2-in. floating- hose, as shown in Fig. 76. This hose is attached, at its free end, to a float /, which may be a 5-gal. oil-can painted with asphalt paint. The high end of the hose is thus sustained slightly below the surface. The solution enters behind a partition &, so that it deposits its sediment at the bottom of the tank. The accumulated sediment is occasionally drained from the tank through the plug-opening s close to the bottom. Solutions pass through the zinc-boxes in a regulated flow into the sumps, the gold being precipitated on the zinc-shaving, which permits the free passage of the solution. Zinc shaving is either bought prepared, or preferably is made at OF THE COMMON METALS. 169 the works. When freshly made, it is more efficient. To make it, a sheet of zinc is wound around a mandrel in a lathe, and the edge of the sheet is soldered down. A side-cutting tool is used to cut the shavings which are V 120 o i n - thick, 1 / 32 in. wide, and several feet long. All the compartments of the zinc-box, except the last, are compactly but uniformly filled with the shaving, the last compart- ment serving as a settling box. Strong solution flows through the three strong-solution zinc-boxes at the rate of one ton per hour. Fig. 77. ZINC LATHE. The gold precipitates chiefly in the compartments first traversed, so that, by the time the solution reaches the seventh one, no discolor- ation of the shaving is to be noticed, and the barren-solution can go to the sump-tank /. The deposit on the zinc has a brownish, then a grayish-black hue. As it increases in quantity, the shaving in the first compartment becomes soft and stringy, and both precipitate and small pieces of the zinc settle through the screen to the bottom of the compartment. As the shaving in the first compartment settles down it is usual to replenish it with zinc from the last compartments, which in turn are filled with fresh shaving. Before putting the shaving into the boxes it is usual to dip it 170 THE METALLURGY in a solution of lead acetate. Lead deposits as a film on the shaving, forming a zinc-lead couple that is more active than the zinc. Where copper is present in the solution selective precipitation is practised. This consists in running into it, just before it enters the zinc-boxes, fresh cyanide solution to raise its grade. In this way copper is held in solution, less of it precipitating upon the shaving. From 95 to 99% of the gold precipitates on the zinc. Since the barren solution is pumped back to the stock-tanks, the residual gold is not lost. The weak solution, passing through the two boxes k' A:' Fig. 66 in series, deposits the gold with more difficulty and less completely than the strong. The solution goes through the first box, then the second, or through 14 compartments in all, to precipitate as much of the gold as possible. It then flows into the tank I' . The clean-up. This is made monthly or bi-monthly, according to the bulk of the precipitate to be treated, and the need of realizing values for operating expenses. At the time of the clean-up, only one zinc-box is taken care of at a time, the flow continuing in the other boxes. The flow of gold solution to this box is stopped, and water is run in to displace the solution contained. Beginning in the first compartment, the shaving is agitated in the water for five minutes with the hands protected by rubber gloves. This is not done roughly, for the brittle shaving would be unnecessarily broken and the water would be black with the floating precipitate. The plug in the side (see k in the cross-section Fig. 74) is gradually withdrawn, and the accumulated slime and water allowed to flow into the launder //. The double line at the side of the zinc boxes, in Fig. 66, shows this launder. It connects by a cross-launder n and a main one m to the acid tank o, 6 ft. diam. The plug is replaced and the compartment is again filled with water. The zinc again is rinsed and rubbed, and the loosened precipitate once more drawn off. About three such washes free the shaving from precipitate and short-zinc. The compartments are thus successively cleaned up, and the shaving from each compartment successively is moved toward the head, and in the last compartment, where needed, replaced by fresh shaving. Finally, the launder is cleaned with a hose, and everything washed into the acid-tank o. When all the boxes have been cleaned, the precipitate is allowed to settle a short time in the acid-tank and the supernatant liquid is syphoned into the 8-ft. settling tank p. In this larger tank the particles of precipitate have opportunity to settle, and to be recovered subsequently in the filter-press. The acid-tank is stirred 172 THE METALLURGY by hand with a wooden hoe, or preferably a power-driven agitator in constant motion. This insures a thorough agitation of the sludge and precipitated in the acid-treatment now to be described. The acid treatment. Upon the watery slime about 30 Ib. of sulphuric acid is poured. This acts upon the short-zinc and produces a violent effervescence. After subsidence the whole is stirred. When again the action abates, 15 Ib. hot water and the same amount of acid is added with occasional stirring. This is repeated until further addition of acid produces little effervescence. Then the mixture is allowed to stand two hours, and a portion is tested with more acid to see that decomposition is complete. The total time for the operation is four to six hours. Filter-pressing. The black mixture, containing zinc sulphate in solution, is diluted with hot water to within a few inches of the top of the tank. The whole content is stirred and then pumped through the filter-press r Fig. 66. The tank is washed and the washings also are pumped through the press. Finally the residue in the press is washed with hot water to entirely remove the zinc sulphate. Fig. 78 represents the filter-press used. It consists of a series of flat cast-iron frames 18 in. square. The frames are recessed, and between them are held canvas filter-cloths when in position for filtering. At the left-hand corner of the illustration, a dotted line on the cross-section of the frame indicates how the filter-cloth rests against the filter-frame. The filter-cloths, for use, are covered with filter-paper, so that the precipitate does not touch the cloth. This paper is burned after use, and the ash is mixed with the precipitate. The slime from the pulp enters the press under pressure through a pipe and valve, as shown at the left. The filtrate escapes by a row of bronze cocks, one on each frame, into a launder, and thence flows to the settling tank p Fig. 66. Leaning against the launder are to be seen two of the filter-frames. These show the grooved surfaces along which the liquid, after passing through the filter-cloth, reaches the outlet hole drilled parallel with the frame and leading to the bronze cocks. In case a filter-cloth breaks, so that pulp begins to issue from any cock, this can be shut off. The mixture enters through the centers of the plates and distributes itself sideways into the recesses between the filter- cloths. The entire precipitate, having been transferred to the press, while in this position is washed with water under pressure. The water is followed by compressed air to dry the precipitate, which, OF THE COMMON METALS. 173 after this, is ready to discharge. To discharge the press, the tightening-screw at the right is slackened ; the follower is drawn back, the frames are successively separated. The grayish-black residue in the recesses of the frame containing 20% or less of water, drops into a drying-pan placed beneath the press to receive it. Drying and final treatment of the precipitate. The product, still damp, is transferred to pans 44 by 24 in. by 4 in. deep. A pan is slid into a cast-iron muffle and heated until the precipitate is dry and then heated to an incipient red. It is then removed, allowed to cool, and the weighed contents cautiously mixed with half its weight of a flux composed of 4 parts borax, 2 parts soda, and 1 part sand. At the Standard plant, Bodie, where these Fig. 79. CENTRIFUGAL PUMP. proportions are used, the precipitate contains 6% silver and 9.5% gold. A clean and fusible slag is obtained by the use of the flux. Since the product is light and dusty, care must be taken in handling it, and for fusing, it must be put cautiously into the No. 60 plumbago (3-gal.) crucible. The melting is performed in the furnace shown at the right in Fig. 106. The crucibles are placed in a wind-furnace and packed in coke. The melting is done as described under silver milling. The molten metal is stirred in the crucible, then poured into conical molds ; and on cooling, the slag is removed and the gold remelted into an ingot. At v Fig 66 is a vacuum pump by which air and solution are withdrawn from beneath the filter-bottoms of the leaching-tanks 174 THE METALLURGY through the pipes, e, e, and discharged by the pipe u to the stock- tanks a and a'. The solution that has lost strength by mixing with weak solution or water is increased to standard strength in the stock-tank. The position of the engine which drives the overhead line- shaft is indicated by g in Fig 66, while w Fig. 65 is the position of the centrifugal pump driven from the same shaft by which solution from the sumps / and I' is returned to the stock-tanks a and a'. Fig. 79 is a view of such a pump. The solution enters through the elbow at the left and delivers from the discharge at the top. The pump has no inside valves to get out of order, and can be run at a velocity of 500 or more revolutions per minute. A complete plant. Fig. 80 is an elevation and Fig. 81 a view of a plant for the treatment of ore by the first method directly from the mine. This has been designed by the Allis-Chalmers Co., Milwaukee, Wisconsin. As shown in Fig. 81, the buildings include a storage-building for the ore-bin, a coarse-crushing and drying-room, and a fine-crushing building, all of which are for the crushing plant and the preparation of the ore. The building used for the cyanide plant is the largest of all, and adjoining it, shown in the foreground, is the power house. At the right, and in front of the cyanide plant, the tailing-dump is indicated. The ore from the mine enters at a high level and is discharged into the ore-bin a Fig. 80. From the bin it is withdrawn through a gate to a pile on the floor near the mouth of the Gates crusher, size 2. The entire ore, including fine and coarse, is regularly shoveled into the crusher, and broken to 1-in. size, the operation being confined to the 10-hour day-shift. Since we desire to crush ore into a product as granular as possible that it all may be leached, dry crushing is here practised. The finer crushing is preceded by drying the ore in order that it may be screened during the crushing. It is accordingly passed through a cylindrical dryer c 18 ft. long with ends 36 and 44 in. diam. respectively. From the dryer, the ore goes to a special fine crusher H, where it is reduced to y-in. size. This discharges into an elevator d, which carries it to an inclined double shaking-screen, so arranged that the ore-stream may be diverted to either screen, thus guarding against loss of time in case of accident or repairs. This screen is simple in action, and low in first cost, and it easily treats the desired quantity (5 tons per hour) screening it to 8-mesh size. The over-size of the screen slides back to the special fine crusher, 176 THE METALLURGY while the under-size goes to crushing rolls set for fine crushing. The product from the rolls is raised by the elevator / to another double 30-mesh shaking-screen, the reject from which passes back to the rolls, while the under-size goes to a storage bin k. It will be noticed that in this system there is no automatic feeding of ore, and no storage after coarse-crushing. Simplicity is obtained by the use of a fine crusher in place of rolls, and by the use of the inclined screen. On the other hand, the system requires the more costly side-hill site, and at the same time, does not escape the need of elevators. The product when crushed is drawn from the bin into scoop- cars standing on platform scales; and the content of each car is weighed. A suspended tract /, with turn-tables and tracks at right angles, is constructed over the leaching or percolating-vats, I, ?, by means of which the cars may be dumped directly into the Fig. 81. VIEW OF FIFTY-TON CYANIDE PLANT. vats, the ore being leveled with a hoe. There are six leaching-vats, each of a capacity of 50 tons, for a six-day-cycle treatment each vat being 24 ft. diam. by 3 ft. deep. Instead of dumping the ore directly into the vat from a height, which would tend to pack it in spots, the better way proposed is to have a hopper centrally placed over each tank just below the OF THE COMMON METALS. 177 track-level. An adjustable spout leads from this. By it the ore from the hopper may be more gently directed to any part of the vat, and kept in a looser, open, and uniform condition for leaching. The strong and weak-solution stock-tanks are situated at the top of the building, and supply solution to the leaching vats for ore treatment. There are two vacuum-tanks used in connection with the vacuum-pump. Upon connecting them to the leaching-vats by a pipe that enters below the false-bottom, the operation of percolation is hastened, especially toward the last, when the ore becomes settled and compacted after the washes. When the vacuum- tank is filled, the solution is stopped and the tank is discharged into Fig. 82. THREE-COMPARTMENT SPITZKASTEN. the gold-solution tanks m or m'. From these the solution flows through the zinc-boxes to the respective strong and weak-sumps // and n". From the sumps the solutions are pumped back to the storage or stock-tanks above. The tailing, still retaining some 16 to 18% water, is shoveled through bottom-discharge doors into tram-cars and run to the tailing dump. The treatment of the zinc-box precipite has been described. 41. CLASSIFIERS AND SETTLING TANKS. Sand and slime, mixed and in suspension in water as in ore pulp, may be separated under free-settling conditions by devices called classifiers. The pulp passes through the classifiers, from the entrance to the discharge in a current, and the heavier and 178 THE METALLURGY larger particles (the sand) separate by gravity, while the finer portion (the slime), still in suspension, passes out through the over- flow with the water. The products are therefore termed the 'over- flow' and the 'under-flow, ' or spigot-discharge. Some of these classifiers have a supply of water, 'hydraulic water', rising from below, just sufficient to keep the slime in suspension until it escapes in the overflow, while the sand settles against the rising current and escapes in the spigot-discharge. Of the classifiers that do not require hydraulic water we here enumerate two kinds: spitzkasten, and classifying cones. Fig. 82 shows, in plan and sectional elevation, a three- FIG. 83. CALLOW CLASSIFYING CONE. compartment spitzkasten. The current enters by a launder at a, and traverses the three compartments successively, the overflow escaping by the launder I. The current widens, and as it enters each larger compartment, flows more slowly, and drops successively the finer sand. The sand, as it settles, escapes by the bent pipes s called goosenecks. Raising thus the end of the pipe diminishes the velocity of the outflow, and less water escapes with the sand. We get three sizes or grades of sand from the apparatus shown : the coarse from compartment 1, medium from compartment 2, and the fine from compartment 3. In cyaniding where it is sought to OF THE COMMON METALS. 179 obtain only two products, a single compartment in a classifier suffices. Fig. 83 is a perspective view of a Callow classifying-cone, operating on the principle of the spitzkasten. The flowing pulp enters by launder to the center of the cone where it is received into a short, vertical, 12-in. cylinder. This deflects the flow vertically downward. Rising again, it overflows evenly around the rim, and being caught in the circular launder at the rim, flows away by the spout shown at the front. The gooseneck coming from the Fig. 84. DOUBLE-CONE CLASSIFIER. apex of the cone is extended into a rubber hose. The end can be adjusted to alter the pressure of the outflowing sand-pulp. The pipe is provided with a side valve, through which water can be forced and an outlet at the apex, to clear it when choked. Of the classifiers using hydraulic water, Fig. 84 is an example. It is a double sheet-metal cone with a cast-iron sorting-chamber K against an upward current of water admitted at E, and escaping by the spigot at H, The slime and fine sand, flowing downward to M are deflected by the adjustable cone L, caught by the rising water-current, and carried upward between the cones, escaping over the edge of the outer cone at the level C. The overflow is 180 THE METALLURGY caught in the circular launder and escapes by the spout D. The cast-iron cone L is adjusted by the hand- wheel 0. Sometimes the inner cone is omitted. The apparatus then becomes a settling-cone with a rising current of water. These are not really satisfactory in their action. Hydraulic water, rising through a sorting-column that enlarges rapidly in sectional area decreases in velocity. If the current enters the classifier in a horizontal stream, the heavier grains settle out and light ones overflow. Heavy grains escape in the sorting column, on the steep sides of the cone and from a bank there. Occasionally the bank falls in a mass into the spigot-discharge where it does not belong, or remains and hinders the action of the apparatus. To obviate this defect the sides of the cone should incline 55 at least. If a pointed box is used instead of a cone, the bank is apt to form at the corners. The cone, being made of metal, has smooth sides and no corners for a bank to begin to form. In a plain cone, using hydraulic water, it is better to use a pulsating current. This descends to the bottom, breaks up the bank, and agitates the pulp, but the expedient has the drawback that slime enters the spigot- discharge with the sand. 42. SECOND METHOD OF CYANIDATION. Where the ore carries gold and pyrite, a stamp-mill, such as shown in Fig. 44, is employed. The coarse gold is recovered by passing the crushed pulp from the battery over amalgamated plates, and the heavy sulphide (principally pyrite) is removed by concentration on Frue vanners or other concentrating tables. The tailing from the mill is classified into sand and slime. The sand is treated by leaching as in the first method, and the slime by one of three methods: (a) Decantation. (b) Filter-pressing, (c) Suction- filtration. The plant, for the cyanidation of tailing is called the cyanide plant; and this may be divided into a sand-plant and a slime plant. When a stamp-mill is provided with an added building for treating tailing by cyaniding, such an addition is called a cyanide-annex. Plant using decantation. The practice in South Africa today has been brought to a high state of development. At first, ores containing $9 gold per ton were milled by wet-stamping and amalgamating. The free coarse gold was recovered, but the tailing, still containing $3.50 per ton, was accumulated in extensive deposits and retained behind dams. Cyanidation was first applied on the accumulated tailing which was shoveled into cars and hauled up OF THE COMMON METALS. 181 inclined tramways to large leaching tanks. This practice, so long as the impounded tailing lasted, was comparatively simple ; but, when the reserve was exhausted it was necessary to devise a method for treating tailing as it came from the mill. Fig. 85 is a diagram illustrating how this is done. The tailing from the stamps, having a value of $9.60 per ton, is neutralized with milk of lime at the launder a. and enters the receiving tank /;, where it is kept agitated by a 4-arm stirrer mounted on a vertical shaft. A centrifugal pump forces it by the pipe c, to the pointed-box classifier d, which takes out a mixture containing about 10% coarse sand and $25 per ton concentrate, this being delivered to the tank g by a distributor /. This is shown more clearly in the Fig. 85. CYANIDE PLANT FOR DOUBLE TREATMENT. illustration of another plant, Fig. 86, where, in the foreground, is a settling tank arranged with a Butters distributor. This is a horizontally revolving device supported by a vertical shaft on which the distributor revolves. The pulp from the classifier is carried in a launder, shown at the right, to the central hopper of the distributor from which radial pipes discharge the pulp into the vat. Each pipe is bent at the end so that the reaction of the escaping pulp sets the distributor in revolution. Owing to the varying lengths, the pipes evenly distribute the sand in the vat. Water and slime, floating above the sand, are withdrawn through four vertical gratings set equidistant on the sides of the tank. One of these gratings can be seen at the edge of the first tank in Fig. 86. As the tank fills with sand, a roller, carrying a narrow canvas curtain, is unrolled to the level of the deposited material. Another method of removing slime and water consists in entirely filling the 182 THE METALLURGY tank with water, and letting the sand drop through it, the surplus of slime and water overflowing into the encircling launder. The overflow from d, Fig. 85, containing $7.90 per ton, goes to the classifier e, which removes a portion, amounting to 65% of the whole, that has an average value of $9 per ton. This clean and finer sand is distributed in the tank g' ', while the overflow from the classifier goes by the pipe A; to classifier /. The overflow from settling tanks g and -g' also flows into the classifier /. The object of the classifier is to take out the little sand that escapes settling Fig. 86. TANKS AND BUTTERS DISTRIBUTOR. previously. The amount is small and it is returned by the pipe m which leads back to the receiving tank 6, and caused once more to go through the system. The overflow from /, consisting only of suspended slime, 25% the weight of the ore crushed, and having a value of $5 per ton, is treated with additional milk of lime from an automatic feeder n. The addition of the lime is for the purpose of hastening the settling of the slime that flows through q to settling tanks, while the clarified supernatant water escapes at p, and is returned to the battery. Assuming that the tank g } which holds 114 tons, has been OF THE COMMON METALS. 183 filled with the mixture of coarse sand and concentrate, and has been drained, it receives 10 tons of weak solution (0.03% KCN). After this the ore is drained and transferred to the lower tank, this part of the treatment occupying five days. The advantages of making the transfer are two-fold: first, that the imperfectly packed portion becomes mixed with the other sand, and second, that the material, thus moistened with cyanide solution, is exposed to the oxygen of the air during transfer. In the lower vat the charge receives 50 tons of strong solution (0.2% KCN), followed by 75 tons of medium, and 65 tons of weak solution. It is then drained and discharged. This takes twelve days more, or seventeen days for the whole treatment. The fine sand in / requires less time for treatment than the coarse. It receives 10 tons of weak solution and 20 tons of medium solution. It is then treated with 50 tons of strong solution, 20 of medium, and 20 of weak. The time is three days for the upper, and five days for the lower vat. The Blaisdell excavator. In place of shoveling the contents of the upper vat into the lower one, an operation which by no means breaks up the lumpy ore, the Blaisdell system for transferring and distributing has been devised. Fig. 87a represents the excavator. It consists of a trussed steel-bridge, supporting at its center a vertical steel shaft with four horizontal arms below, called excavator beams. This bridge travels on rails and may be set over any vat. The central vertical portion of the bridge carries guides and screws for raising or lowering the shaft and arms. The arms carry steel harrow-disks. In operation, the central door or trap at the bottom of the vat is opened and a hole is quickly dug down to it through which the sand may fall. The excavator is now placed in position and the shaft and arms, having been put in motion by an independent motor, the steel disks pass over the surface of the sand, cut it, and roll it toward the center where it falls down through the trap. The shaft is gradually lowered as the cutting proceeds, much as an auger would advance in boring a hole, until the vat is emptied. The arms are then raised and the excavator moved to another vat. The excavated sand from the settling tank falls upon a belt- conveyor which carries it to the leaching vat. When it reaches the vat a distributing machine or distributor, shown in Fig. 87b, that consists of a movable steel bridge supporting a conveyor- Fig. 87a. BLAISDELL EXCAVATOR. Fig. 87b. SAND DISTRIBUTOR, BLAISUELL, SYSTEM. OF THE COMMON METALS. 185 belt, takes the sand from a tripper and drops it into a hopper at the center of the bridge. It falls from the hopper upon a rapidly revolving steel plate which breaks lumps and distributes the sand lightly and evenly in the vat in an ideal condition for rapid and uniform leaching. The large circular band or ring, suspended from the bridge, is to prevent throwing the sand too far. (a) Decantation. The slime coming from the pipe q, Fig. 85 is collected in large settling tanks. Each tank when filled is cut out from the flow, and allowed to settle. After settling, the supernatant water is drawn off by the decanting device. Fig. 76. This removes the surface water continuously. A weak solution (0.1% KCN) is then added to the slime in the settling tank, and agitation effected with mechanical stirrers, or by transferring the pulp from one tank to another with a centrifugal pump. After several hours the solution is allowed to settle, then drawn off by decantation. More cyanide is then added, the whole stirred, allowed to settle, and half the solution again withdrawn. The operation is repeated several times, the final wash being of water. In this way 75% of the gold in the slime is extracted. The various cyanide solutions mentioned above are run through the zinc precipitating boxes, the zinc-shaving having first been prepared by dipping it in a solution of lead acetate. The zinc becomes covered with a film of lead forming a zinc-lead couple, and is more active in depositing gold than zinc alone would be. The objections to this method of slime treatment, are : the large and awkward plant required, and the loss due to the impossi- bility of saving the last traces of dissolved gold. These were soon recognized by metallurgists in Western Australia, and the following method was devised : (b) The filter-press method of slime treatment. This consists in agitating the cyanide containing slime-pulp (in a settling tank) and removing this solution containing the gold, by forcing the material under a high pressure into a filter-press, Fig. 78. The solution retained in the slime-cake is displaced by forcing dilute solution, or water, through the press. The comparatively dry compressed cake of slime containing 20 to 25% water, is dropped into a car beneath and trammed to the dump. The filtered solution goes to the gold-solution tank, and thence through the zinc boxes, to the proper sump. An estimate of the cost of this kind of filter- pressing is 38c. per ton. (c) Suction filtration. With the exception of the system used 186 THE METALLURGY at the Homestake slime plant, the comparatively high cost of filter- pressing and the heavy cost of installation has led to the adoption, in the United States, .of various systems of vacuum filter-pressing of which the Butters filter, Fig. 88, is an example. The Butters solution-filter. The white rectangular object, at the center, in the background of the illustration, is a cluster of four filter-leaves suspended from an over-head crawl. These have just been hoisted out of the tank below. A single leaf is 10 ft. long by 5 ft. high. The upper part of the frame of the leaf is Fig. 88. BUTTERS FILTER TANKS. a wooden bar, sufficiently long to span the tank and rest upon the sides. The remaining three edges of the frame are of iron pipe. The filter-leaf consists of a piece of cocoa-matting cut to exactly fit the frame, and this is covered on each side by a sheet of canvas. These three sheets of material are sewed together, and the whole fitted round and fastened to the frame. Five vertical strips bolted on the outside serve to stiffen the frame. Fig. 88 gives a view of two 60-frame filter-tanks which, in Fig. 89 and 90 are marked E, E. Fig. 89 is an elevation and Fig. 90 a plan of a 120-frame Butters vacuum-filter plant. The slime is in cyanide solution several hours, or until the gold is dissolved. For convenience the pulp is collected in a tank A, called the pulp-storage tank, and from there is run, as desired, into one of the filter-tanks E, E, filling it to the tops of the frames. It will be noticed in Fig. 88 that the pipes of the OF THE COMMON METALS. 187 frames are connected to a common header. This header leads to a vacuum pump N. Upon starting the pump, the solution is sucked through the filter-frames, discharging into a gold-solution tank outside the building. Pulp is run into the box at intervals to keep Fig. 89. GENERAL ELEVATION OF BUTTERS FILTER PLANT. Fig. 90. GROUND PLAN OF BUTTERS FILTER PLANT. the filter-leaves submerged. The slime remains as a cake on the outside of the filter-leaves, and forms a deposit sometimes nearly an inch in thickness. When the cake is formed the supply of pulp is shut off, and the surplus pulp run to the tank C and pumped back to A. A weak 188 THE METALLURGY cyanide solution from tank B is next admitted to the filter-tank, and is sucked through the cake to displace any gold solution remaining, the level being maintained as before. The surplus wash- solution is pumped back to B. In the same manner the cake is washed with clear water to displace the last of the cyanide solution. Both washes go to a weak gold-solution tank near (}. While the tank is being emptied and filled and the cake is exposed to the air, the vacuum must be reduced to 5 in. to hold the cake firmly in place. The vacuum pump is now shut off, and air under a slight pressure is forced into the leaves, displacing or throwing off the cake, which drops to the bottom of the tank. The 12-in. quick- opening discharge gates, 6% are opened, and the slime is washed out into a waste-launder. The gates are again closed, and the filter is ready for a new charge. The formation of the slime-cake adjusts itself as the composition of the material varies at any point, the coarser forming a thicker layer than the finer, thus maintaining the even permeability. The time of washing is 15 minutes. The operation is conducted by one man. A precipitate of calcium carbonate gradually forms upon the leaves, decreasing their permeability, and every two to nine months this is dissolved by dipping the leaves in tanks J which contain a weak solution of hydrochloric acid. In Fig. 88, a set of leaves is shown in the process of transferring to the washing-tank. The operating cost for filtering, including labor, power, and repairs, in a large installation, amounts to 8c. per ton of slime treated. 43. CYANIDATION AT THE HOMESTAKE MILL, SOUTH DAKOTA. The ore is a garnetiferous hornblende schist containing 7 to 8% pyrite and pyrrhotite. It is crushed with eight to ten times its weight of water, and amalgamated, using inside plates for the mortar, as well as outside amalgamated apron plates. The pulp from the mill, some 1300 tons daily, is carried by launder to a cone-house, Fig. 91, to a set of twelve settling cones that are 10 ft. diam. with sides having an angle of 50 degrees. The stream is distributed evenly to all, entering each at the center. Half the water and the finest slime is removed by the overno\v. Some fine sand is carried over with slime, and this is separated by a series of settling tanks, the overflow running to a pond. Here the water has an excellent chance to settle, and when nearly clear, it is pumped back to be used again in the mill. The sediment accumulates in this pond and finally is washed out with a hose and OF THE COMMON METALS. 189 run to waste. From the bottom of the cones is withdrawn a thickened pulp containing the sand and some slime. This is carried by a 12-in. pipe on a 2 l /2% grade to the sand-plant, Fig. 92 and 93. Classification at the plant is effected by means of six gravity settling cones 6, 7 ft. diam. and with sides of a 50 slope. The underflow from each of these goes to four classifying cones c, 24 in all, which are 4 ft. diam., and with steep sides of 70 degrees. These are provided with hydraulic water, as in the lower cone of Fig. 99, by means of which the sand settles nearly free from slime. The slime from these cones, containing 90c. to $1 per ton, and amounting to 40% of the total tailing, was formerly wasted, but more recently has been treated by filter-pressing, at the slime- plant to be described later. The sand-plant. The prepared sand contains 40.5% coarse par- ticles that remain on a 100-mesh screen; 30.8% middles, between 100 and 200 mesh, and 28.7% fine passing 200 mesh. This leaches at the rate of 3 or 4 in. per hour. Before the sand enters the Elevation ofCo-neJfouse PJan *f Gynp-ffcvff Fig. 91. CONE CLASSIFIERS, HOMESTAKE MILL. leaching vats it receives a stream of milk of lime which has been prepared by being stamped in a five-stamp battery reserved for the purpose. From 3 to 5 Ib. of the lime are added per ton of sand. The classified pulp and lime thus mixed, pass to a Butters distributor, see Fig. 86, which can be transferred from one vat to another by an overhead trolley. There are 14 leaching-vats, each 44 ft. diam., 9 ft. deep, and capable of holding 610 tons. The tank is filled with w r ater and the sand runs in. It takes 11 hours to charge the tank, and treatment lasts five days. When the tank is filled the ore is drained and a series of washes of the stronger of the stock solutions (containing 0.14% KCN) is run in, allowing each wash to drain below the top of the ore to draw in air. The effluent, its strength reduced to 0.10%, is run to the two weak- 190 THE METALLURGY solution precipitation tanks / /, Fig. 93, each 26 ft. diam. by 19 ft. deep. After this, the weak solution is brought upon the charge and retained two days more. The solution escaping during this period is run to the two strong-solution collecting tanks, e e. . This is followed by a water wash which finally reduces the unextracted gold to 5 to 7c. per ton. The charge is now ready for sluicing out. This is done by two men in four hours, four side gates and one bottom gate being OF THE COMMON METALS. 191 used for the purpose. The 8-oz. duck filter-cloth underlaid with another of cocoa matting is washed clean. The vat is then filled with water and is ready for the next charging. Precipitation. The solution, resulting from the leaching with strong solution, run into one of the weak-solution tanks / /, contains $2 gold per ton. When the tank is filled the stream is turned into the second tank, and the first is ready for precipitation. The tank holds 300 tons of solution that is agitated by compressed air admitted from a pipe across the bottom of the tank pierced with numerous small holes for the escape of the air. Sixty pounds of zinc powder in the form of an emulsion is sprayed in, during the agitation, and in 20 minutes the precipitation of the gold is complete. A duplex pump is now started and the mixture in the tank is pumped up to two large filter-presses, Ti and i, each containing 24 frames 36 in. square. The filter-cloths of the presses soon become coated with gold precipitate and zinc-powder, and every drop of solution passing through, comes into molecular contact with the zinc dust. The value of the solution is reduced from $2 to 5 or lOc. per ton (a precipitation of 97.5 to 95.0%) and is then termed barren solution and passes to .the weak-solution storage tanks A*. The weak-solution wash above referred to fills the strong- solution collecting-vats e e and is strengthened to 0.14% KCN and pumped directly, without prior precipitation, to the strong- solution storage tank /, from which it is drawn for the first treat- ment of the sand. It is seen that the strong solution of one day becomes the weak solution of the next, and that all the gold accumulates in the weak- solution precipitation tanks. The strong solution has an approx- imately constant value. One-half the total effluent solution is pre- cipitated, the other half having a nearly constant value of 30 to 50c. per ton. The precipitate, containing gold and the fine zinc, accumulates in the filter-presses, that are run a month without opening. In this way about a ton of precipitate, worth $50,000, is obtained. The presses are opened, the precipitate removed and sent to a lead-lined mixing tank that is equipped with a mechanical agitator, a hood, and an exhaust fan for removing acid fumes. Here it is treated with dilute hydrochloric acid to dissolve the zinc, agitated, and settled. The supernatant solution is then drawn off, by means a montejus or pressure-tank to another smaller filter-press. Sulphuric acid is next added in the mixing-tank and the mixture 192 THE METALLURGY is agitated and heated. Upon settling, the supernatant acid solution follows the first one through the press. Wash-water is next run into the tank and without further settling is run to the press. Finally the precipitate in the press is washed with clean water. The acid solutions and the wash-water go to a large settling tank that acts as a guard to recover any escaping particles. The acid-treated precipitate is now transferred from the press to a large steam-dryer, where a part, but not all, of the moisture is removed. It is then mixed with litharge, borax, silica, and powdered coke and sprinkled with lead acetate solution, and briquetted in a press under a pressure of 2 to 3 tons per square inch. The briquettes are dried, charged into an English cupelling-furnace, see Fig. 179, and quietly fused, suffering no loss. The lead, as in assaying, absorbs the gold and sinks to the bottom, while the slag flows away as it forms. The 'test' gradually fills with lead, which is oxidized to litharge in cupellation, leaving a metal 975 to 985 fine in gold and silver, which is run into ingots. The litharge produced is reserved to be added to the precipitate at the next clean-up. The cupel-slag, cupel-bottoms, sweepings, or other metal- bearing cleanings are accumulated, and occasionally run through a small silver-lead blast-furnace. There is produced a slag assaying less than $5 per ton, and base-bullion which is returned to the cupel-furnace at the next clean-up. The method of treating the precipitate, it is claimed, is practised with a loss of only 0.1 per cent. The cost of treatment per ton of sand is 26c. This has been brought into comparison with South African costs that, before the Boer war, varied from 55 to 72c. per ton. The slime-plant. The slime pulp, amounting to 1600 tons daily, has an average value of 91c. per ton. It contains three tons of water to one ton of solid, and is carried two miles by a 12-in. pipe at a grade of 1.5% to the slime-plant. Here, two small vats are provided for slaking lime. The content is drawn to a screen- bottom box where the undissolved lumps separate. The box over- flows into an agitator from which the milk of lime continuously runs into the main slime-stream at the rate of 5 Ib. of lime per ton of dry slime. Two storage vats, 26 ft. diam. and 24 ft. deep, having conical bottoms with 47 sides, receive the stream. From the bottom of these storage vats the slime-pulp is drawn continuously through a 10-in. pipe to large filter-presses 65 ft. below to obtain a pressure of 30 Ib. per sq. in. The 10-in. main extends the whole length of the press-building. Between each OF THE COMMON METALS. 193 pair of presses, the main branches into 10-in. pipes, which in turn send two 4-in. branches to each press. The smaller branches connect to a 4-in. passage or channel that extends along the center of the top of the filter-press frames. From the channel the slime- pulp flows into the press. There are 92 frames each 4 ft. by 6 ft. and 4-in. distance-frames to form slime-cakes 4 in. thick. The press having been filled, a 0.1% solution of KCN is run in, entering at the lower corners of the press by two channels each 2.5. in. diam. Compressed air is admitted at the upper corners of the press by two channels each 2.5. in. diam. This is followed by a 0.04% KCN wash, another treatment with air, and again a water-wash, until the exhausted slime contains no more than 9 cents gold per ton. Along the center of the bottom of the frames is a continuous 6-in. channel, and within it a 3-in. pipe extending the length of the press. This pipe is provided with 92 nozzles 1 in. long and r> / 32 i n - diam. By a special mechanism a revolving motion through 200 degrees is given to the pipe, so that under a water-pressure of 50 Ib. per sq. in., the nozzles play against the compact slime-cakes, removing the cake completely from the compartments in 45 minutes, the 6-in. channel (generally closed) being then opened for the exit of the slimes. The operation, exclusive of the time of filling and emptying, occupies 6 hours. It will be noticed that the solution of the gold, and the extraction of the slime, is done entirely at one operation and within the press. The effluent solutions now flow to the strong and weak gold precipitation-vats where the gold is precipitated by means of zinc dust. Zinc-dust is a fine powder produced in retorting zinc ores and contains 90% Zn. An emulsion of this zinc in water is added to the solution in a conical-bottom tank. With an air-lift pump a thorough mixture of the zinc with the solution is obtained. As at the sand-vats, the solutions are pumped to the storage solution-tanks and to the precipitate-presses, the treatment at this stage being the same as that already described as taking place at the sand-plant. The estimated cost of this plant is half a million dollars, and the cost of treatment, per ton of slime, is 25 cents including all items. The recovery is 90 per cent. 44. CYANIDATION AT EL ORO, MEXICO. (THIRD METHOD.) The third method of cyaniding is well exemplified in the practice at El Oro, Mexico. The oxidized ore, a hard compact quartz, contains from $6 to $15 gold and 3 to 5 oz. silver per ton. The 194 THE METALLURGY gold is distributed so finely through the quartz that seldom can a ' color' or visible speck be discovered. The silver is partly metallic, and partly in the form of sulphide, arsenide, and antimonide. Both metals are considered to be deposited between the faces of the elementary quartz crystals. The necessity of fine grinding to unlock the gold and silver particles is well illustrated by Fig 94. The mine-ore is dumped from skips at the shaft into ore-bins, coarsely crushed, and conveyed to the mill by cars. It is stamped through a 25-mesh screen, then goes over amalgamated plates which catch from 13 to 18% of the gold and 1 to 5% of the silver. The 100% Fig. 94. RELATION BETWEEN MESH AND EXTRACTION. pulp from the plates is re-ground through tube-mills, then flows to the cyanide plant where it is separated into sand and slime, these being treated separately. The daily duty per stamp is 4.7 tons, and the ratio of water to ore is as nine or ten to one. . Fig. 95 is a flow-sheet showing the method of treatment. The ore pulp from the plates passes to two cones 4.5 ft. diam. beneath which, in series, are two 2-ft. cones called 'pulp thickeners.' The overflow from all the cones gives a product containing 81% slime, while the underflow or spigot discharge, a sand still retaining 9% slime, passes to the tube-mill No. 5, where half of it is ground to a slime. OF THE COMMON METALS. 195 Fig. 96 represents a view and Fig. 97 a longitudinal section of a tube-mill. It consists of a steel shell, like a boiler-shell, 5 ft. diam. by 23 ft. long, making 27 revolutions per minute, and having a daily capacity of 130 to 150 tons. It is lined with silex blocks OF THE COMMON METALS. 197 (a compact quartz) about the size of ordinary brick, and in operating is filled with hard quartz pebbles the size of the fist. As the barrel revolves, these pebbles, dropping back upon one another, finely grind any material fed to the mill. The ore is fed in from a hopper placed at one end. The ground pulp escapes through the perforated false head at the other end, and overflows through the other trunnion. 198 THE METALLURGY Referring to Fig. 95, the discharge joins the overflow of the cones and goes to spitzkasten No. 1, which takes out the insufficiently ground sand. This is raised by the elevator wheel and sent to return cone No. 5 (the purpose of which is to remove the unground sand), then goes to tube-mills No. 3 and 4 for re-grinding and finally back to spitzkasten No. 1, while the fine-sand overflow of return cone No. 5 passes to one of the nine 'sand-receivers' over a Butters distributor. The overflow of these sand-receivers goes on to spitzkasten No. 2, which takes out a little fine sand which returns to the elevator wheel, and an overflow that joins the over- flow from spitzkasten No. 1 and passes on to the 'slime tanks.' Each sand receiver, when filled, is drained 24 hours, and with- out further preliminary treatment, is discharged by a Blaisdell excavator, see Fig. 8jTa, upon a belt-conveyor that takes it to one of the 12 treatment tanks to be charged by a Blaisdell distributor, see Fig. 87. In these tanks the sand is regularly leached. The treatment consists in leaching the sand with a certain number of strong and medium solutions or washes, given alternately, and followed by a number of weak-solution washes. By these, the gold in the tailing is reduced to 20c. per ton, and the total time of leaching is 90 hours. The use of the sand-receivers to do the first part of the leaching has been discontinued because of the loss of time and cyanide solution, and because the tank becomes tightly packed and uneven. The aeration effected by excavating and redistributing the sand, has a great influence on the rate of extraction. This is helped by the speed at which the washes follow one another. The gold is more speedily extracted than the silver. Cyanide consumption is also rapid at first. Upon the completion of the treatment the tailing is removed by another Blaisdell excavator and by a troughed conveying-belt that takes it to the dump. The slime-pulp, overflowing from spitzkasten No. 2 is composed of 1 part of slime to 12 of water. To cause the slime to settle readily, caustic lime is added to the mill-pulp (just before it enters the tube-mills) at the rate of 12 Ib. lime per ton of ore. The slime-pulp flows continuously into one of the slime-tanks, 75 tons of slime settling in the bottom of the tank, while the clear supernatant portion overflows to the 'mill-water sumps,' whence it is pumped back for use in the mill. When a charge has been completed the slime is allowed to settle 8 hours, and the clear water decanted. The slime is now stirred by means of the mechanical agitator of the slime-tank. In ten minutes thereafter a OF THE COMMON METALS. 199 centrifugal pump is started by which the pulp is elevated and thrown back into the tank. The pulp having been thoroughly mixed, a specified amount of 0.05% cyanide solution as well as lead acetate (0.4 Ib. per ton of slime) is added, and the whole agitated 7 hours. The tank is filled with fresh cyanide solution, the stirring is stopped, the content allowed to settle, and the solution decanted as closely as possible. Four additional washes are given as rapidly in succession as the settling of the slime will allow. The residual slime is then pumped into a slime-settler, is gradually settled, then decanted, and the discharge slime run to waste. It will be seen that there are two of the gold-solution tanks, the 'strong' and the 'medium,' which take their supply from the sand-tanks, while a group of three (the weak-solution tanks) take their supply, not only from these, but from the slime-tanks as well. The final wash, which is decanted from the slime-settlers, contains but little gold and need not go through the zinc-boxes but instead, to the weak-solution sump-tanks. To aid precipitation, at the strong and medium zinc-boxes, a regulated flow of 2.5% KCN is added to the inflowing solution from the sand-plant. After this has passed the boxes, it is raised to its full strength by the addition of a concentrated solution of KCN contained in a 'dissolution tank' near the sumps. When pumped to the slime-tanks it is there strengthened by letting it run over a sack containing solid KCN salt. At the sides of both the slime tanks and the slime-settlers are swinging decanting pipes, which withdraw the upper portion of the clear supernatant solution. 45. CYANIDATION IN SOUTH DAKOTA (FOURTH METHOD). An example of this method is found at the Maitland mill, Maitland, South Dakota. The ore is close-grained, hard, and in the main oxidized, through carrying a little pyrrhotite and pyrite. It averages 75% SiO 2 , 11% Fe, 1.2% S, and contains $9 gold and 0.5 oz. silver per ton. The gold occurs evenly distributed and finely divided, so that cyanidation is necessary to recovery. Fig. 98 is a plan of the mill and indicates the operation. The ore from the mine is dumped into a flat-bottom bin, the idea being to permit the ore to form its own slope, thus avoiding wear upon the bottom of the bin. It is coarsely crushed by a large 24 by 13 in. Blake crusher to 1.5-in. size, and is elevated by a belt elevator to the battery bin, a sample amounting to 2% of the ore being taken at the elevator head. The ore from the battery storage-bins is fed by automatic feeders as shown in Fig. 45, to the 40-stamp mill 200 THE METALLURGY of 910-lb. stamps, where it is wet-crushed in cyanide solution through a 26-mesh screen at the rate of 3 tons daily per head or 120 tons in all. The battery solution is kept at a strength of 0.06% SAMPLE ROOM r Sliiue Vyfsiime ' to Jistril.utor over vata 1_/V_L Cle.in sand to distributor over VMS i 1 ._ . J_qbld_ '^solution _ C L . LE , C J IN \ \ /BATTERY) "I ( GOLD V Cyanide tot V fiOL / \ TA \t I VACUUM!/ .wash water net to FII >L^I J Gold sol. Fig, 98. PLAN OF OPERATION AT THE MAITLAND MILL OF THE COMMON METALS. 201 cyanide, and has 0.05% protective alkalinity. Four to five tons of solution are used per ton of ore. In crushing, 4 to 6 Ib. quick- lime is added in the battery to assist later in the precipitation of the slime. The proper separation of sand from slime is important. It is performed in the following way : The battery-discharge is elevated by Frenier sand-pumps to the distributor-box of the cone system as shown in Fig. 99. The feed in this box is divided evenly between discharge Distributing box fo sand tanks Fig. 99. SINGLE-CONE CLASSIFIERS. two simple settling cones 50 in. diam. using no hydraulic water, so that as the sand settles and discharges at the bottom of the cones, it still carries slime. This spigot discharge is received in a launder which delivers it to a 50-in. hydraulic cone-classifier having a hydraulic supply, not of water but of battery solution. Thus a clean sand is separated at the spigot discharge that contains no more than 1 to 5% slime, while the slime with much of the solution joins the overflow from the first cones. Treatment of the sand, The clean sand from the spigot discharge of the lower cone, amounting to half the total weight of 202 THE METALLURGY ore, and containing 2 x /2 to 3 parts solution to one of sand, Hows by launder to the Butters distributor and fills the tanks evenly. The six ' sand- vats, ' holding 140 tons each are 30 ft. diam. by 6 ft. deep, with lattice bottoms as shown in Fig. 68. The lattice is covered with cocoa-matting and the matting with 8-oz. duck. The filter lasts 10 months. A tank is filled in 60 hours. A system of dry filling is used, the vat not being filled first with water. All solution coming in with the sand is allowed to drain through the filter-bottom. This gives a more porous and easily leached product, since the slime is evenly distributed through the sand as it could not be in a vat first filled with solution. The average weight of a cubic foot of the sand is 93 pounds. The vat, having been filled, is leveled with a stream of solution from a hose under low head. Battery solution is then run on for a period of 10 days. A small amount of slime in the solution forms a slight coating on the sand, and to insure satisfactory leaching, this occasionally is lightly raked to keep it pulverous. The battery- solution is followed by barren solution for six days more. This is allowed to drain and is followed by a water-wash, amounting to 10% by weight of the sand. To treat a charge of 140 tons of sand requires 900 tons of battery solution (exclusive of that which has passed through the filter filling) and 450 tons of barren solution. This large quantity of solution (ten parts of water to one of sand) and prolonged treatment (16 days) is by no means excessive, since experience shows that much weak solution must be constantly percolating through the charge. The exhausted sand is sluiced through a launder of 8% grade with 100 to 150 tons of water. The tank is then ready for another charge. Treatment of the slime. Referring to Fig. 98, we note that the overflow from the upper and lower cones is united into a single flow, which goes by launder to one of the slime vats, No. 4 or 5, called 'loading-vats.' Each loading-vat has a central partition extending to 30 in. from the bottom. When the vat is filled the stream of slime-pulp continues to enter the vat quietly on one side, the solid portion settling out, while the nearly clear solution escapes over the rim of the tank at the opposite side. The vats are 24 ft. diam. by 12 ft. deep and hold 150 tons, and receive during 12 hours 300 tons of slime-pulp containing 12 tons of solution to one of slime. Thus 150 tons of slime is decanted and 30 tons is retained in the vat. It is allowed to settle and is decanted as closely as possible, then pumped by a centrifugal pump to slime OF THE COMMON METALS. 203 vat No. 1, barren solution being also added. Meanwhile the other loading-vat is filled, settled, and transferred to slime vat No. 2 with barren solution. Both vats, No. 1 and 2, are now decanted and the contents combined and transferred to vat No. 3, making a fuil charge of 60 tons of slime. Two more transfers, with barren solution added each time, are made to slime-vats No. 6 and Y, and finally "a transfer to slime-vat No. 8 where the charge receives a final water-wash. After each transfer and dilution, several hours of agitation are given by pumping from the bottom of the vat and discharging into the top. After decanting the wash-water, the slime contains 55 to 60% moisture. The layer of thin slime is drawn back on the next charge, leaving a dry slime with but 47% moisture. This is sluiced out and run to waste. Course of the solutions. The rich gold-bearing solutions from the sand-vats, the only ones that run through zinc-boxes, are received in a collecting-box and pass to the gold tank where they are united and brought to standard strength by the addition of fresh KCN. This promotes precipitation at the zinc-boxes. The poor solutions from the sand-vats, and the decanted and overflow solutions from the slime, still called battery solutions, flow to another collecting box and thence to the battery sump. The solution in the sump is pumped back to the two battery-solution stock-tanks for use again. It is not yet sufficiently rich in gold to justify passing it through the zinc-boxes. It is thus seen that the cyanide solution makes a closed circuit, the only part escaping being that in the exhausted sand and slime. The consumption of potassium cyanide is but 0.84 Ib. per ton of ore treated. Precipitation. From the gold-tank the solution passes in a regulated flow r to the five zinc-boxes each of 224 cu. ft. capacity, having eight compartments. Here the gold is precipitated. The barren solution passing on, enters the barren-solution sump whence it is pumped back to the barren-solution stock-tank at the highest part of the mill. Of this solution, amounting to some 400 tons daily, 25% is used for the treatment of sand and the remainder for the slime. The consumption of zinc averages 1.33 Ib. per ton of ore treated. Cleaning up. A clean-up, varying somewhat in detail from that described under the * first method,' is made twice monthly. To clean up a box, the flow from the gold-solution tank is shut off and water run in 15 minutes to displace the solution in the shaving. The zinc in the first compartment is shaken up and down to remove 204 THE METALLURGY adhering precipitate and set aside. The water is bailed close to the precipitate in the compartment into the next compartment. The plug at the bottom of the compartment is then withdrawn and the remaining water, with some precipitate, Hows through a launder to the acid-tank. The screen-bottom is taken out, and the remaining precipitate is removed, placed in a tub, and carried to the acid- tank. The compartment is washed with a hose, the water going to the acid-tank. The plug and screen is replaced, and the zinc shaving put back, with shaving from the second compartment, to completely fill the first. The compartments are cleaned in succession, the zinc- shaving being moved toward the head until all compartments HIV filled. In filling the first compartment a slight flow of solution is started to avoid undue exposure to the air, and the water from the last compartment is poured into an adjoining zinc-box. The acid- tank thus receives all the precipitate and some of the wash-water. One man cleans up the five boxes in 12 hours. The precipitate is allowed to settle in the acid-tank, then the clear supernatant solution is closely syphoned into a waste tank that holds 25 tons of water, and shown in Fig. 98. Concentrated sulphuric acid is next added to the acid tank and the mixture is stirred in by hand to avoid boiling over. The treatment lasts an hour. To avoid fume a well ventilated portion of the precipitating room is chosen for the acid-tank. Acid treatment being now complete, the tank is partly filled with water, the vacuum filter started, and the entire content of the acid-tank filtered without further washing. The filtrate goes to the waste tank. The precipitate is blown from the filter-frame, collected in the hopper-shaped bottom of the filter vat, withdrawn in pans, and taken to the melting-room and dried in a muffle furnace. To 10 parts of the dry precipitate are added and mixed 4 parts soda, 1 part borax, 1.5 parts sand and 0.2 parts fluorspar. The mixture is melted quickly in a No. 200 graphite crucible, in a wind-furnace having a forced or under-grate blast. The content of the crucible with the slag is poured into molds and the slag removed upon cooling. This slag contains matte, carrying as much as $10 gold per pound. It is accordingly remelted with sand, flux, and 10% by weight of scrap-iron. The fusion gives a coppery bullion 700 fine in silver and 80 fine in gold, and a slag that is shipped to a smelting works. The ingot, weighing 200 ounces, is treated with hot concen- trated nitric acid in a porcelain-lined kettle and gives a residue containing 50% Au and 25% Ag. This residue is added to the zinc-box precipitate of the next clean-up, while the acid solution, OF THE COMMON METALS. 205 containing most of the silver, is treated with soda which precipitates the silver carbonate. This is roasted at a low red heat, and yields silver which is easily fluxed and melted in a crucible and cast in the form of a bar. The united solutions from the acid tank and the suction filter are retained in the waste sump. They contain $2 gold per ton. The solution is treated w r ith fine zinc to precipitate the gold; sulphuric acid is added and the mixture well stirred to remove the excess of zinc. Some 90% of the gold is thus precipitated. The sweepings around the zinc-boxes are thrown into the sump and their content recovered once in six months. The extraction at the Maitland mill in 1901 amounted to 75.5% of the gold and 44.3% of the silver. The total cost, including general expense, but not depreciation of the plant was $1.61 per ton of ore treated. 46. CYANIDATION OF CRIPPLE GREEK ORE (FIFTH METHOD). The ores of the Cripple Creek district, Colorado, are phonolites and are of two kinds: first, the weathered surface ore, having little or no tellurium and containing free, finely disseminated gold; and second, the deeper-lying ores containing gold combined with tellurium as syivanite and calaverite, with some nearly barren pyrite. The surface ores are treated raw. The telluride ores receive an oxidizing roast before cyanide treatment. The treatment of the telluride ores we are now to consider. Metallic Extraction Co., Cyanide, Colorado. The ore, coarsely crushed to %-in. size, is stored in bins, then supplied to a revolving dryer, and after drying crushed to 30-mesh size, as described in connection with Fig. 25. The product is elevated to storage hoppers from which a regulated feed is delivered to the roasting-furnaces. The roasting not only breaks the combination of gold and tellurium, but makes the ore porous, and accessible to the solution, and causes it to be more readily leached. The roasting is done in one of the mechanical roasters of the Argall or the Ropp straight-line type. The cooling apparatus, which receives the hot ore from the roasting furnace, is a flat-bottomed trough, water-cooled by pipes beneath. The ore (first mixed with 10 Ib. quicklime per ton) is moved along by scrapers which deliver it to a storage pit at the end of the trough. At the discharge end, the ore is sprinkled, before leaving the trough, with the cyanide solution. This cools it, prevents dusting, and starts dissolution. 206 THE METALLURGY The drying and roasting-furnaces, as well as the housing of the dry crushing machines, are connected by flues to two fan blowers, each capable of delivering 10,000 cu. ft. air per minute. These fans deliver the dust to a bag-house, described under the 'Bartlett Process.' Dust, thus recovered, is about twice the value of the ore from which it arises. It is briquetted in a White briquetting-press, Fig. 164, and smelted. The leaching tanks are 50 ft. diam. by 6 ft. deep, constructed as shown in Fig. 70 and 71, each holding 400 tons of ore. In operation, the tank is filled with ore from two-wheeled buggies, and 0.7% KCN solution is admitted below under pressure, so that the ore and solution rising at the same rate fill the tank in 50 hours. The solution stands on the ore 24 hours, then more is added and drawn off. (This practice is inferior to that of actually percolating the ore.) The operation, repeated several times, reduces the gold content to one third of the original in five days. A weak solution of 0.25% KCN is used in the same way for three days. The charge is then given several water washes, drained, and finally washed out with water under high pressure from a hose into a launder, the washing and discharging taking two days more. At Cyanide, where the plant is situated on a level site, the tailing is elevated by a centrifugal pump to a launder on a high trestle, discharged to a dump, and impounded behind a low dam. In the launder is a series of transverse riffles behind which any coarse particles of gold, produced by roasting, are caught. These particles are likely to be found in roasted telluride ore, in consequence of the condition of the gold when first released from combination with tellurium that causes it to form drops, shot, or other coarse particles. As an alternative to washing out the tailing, it may be shoveled through side-discharge doors, Fig. 72, and trammed away, a more expensive method. The cycle of treatment as above detailed was 12% days. The value of the original ore, charged to the vat, may be assumed to be 0.90 oz. Au and 0.5 oz. Ag per ton, while the tailing retains 0.06 oz. Au and 0.2 oz. Ag per ton. The extraction accordingly is 93.33% for the gold and 60% for the silver while the loss in cyanide is 1.75 Ib. per ton of ore treated. The assay of the tailing shows as good extraction from the grains of 40-mesh size as from those passing a 200-mesh screen, and indicates how effective roasting is in preparation for subsequent leaching. Each kind of solution passes to a separate gold-solution tank and is run in regulated flow through separate zinc boxes to its own OF THE COMMON METALS. 207 sump to be strengthened for re-use. The zinc consumed amounts to 0.9 Ib. per ton of ore treated. The precipitation and the details of subsequent operation are described under the * first method.' The cost of treatment of Cripple Creek ore is $2 to $2.50 per ton including roasting, when performed on the large scale indicated. The Golden Cycle Mill, Colorado City, Colorado. The ore from Cripple Creek, treated at the mill, is coarsely crushed to 2-in. size Fig. 100. MONADNOCK (CHILEAN) MILL. in rock-breakers and passed through coarse-crushing rolls which reduce it to one inch after which a sample is removed automatically. The ore is taken by a belt-conveyor to storage bins holding 8000 tons. This collects the ore into large lots and insures greater uniformity of treatment. The ore is removed from the storage bins on a troughed belt-conveyor and is delivered to intermediate rolls able to crush without preliminiary drying* to ^-in. size. This rather coarse ore is delivered to Edwards roasters, see Fig. 38 and 39. It has been found that roasting can be thoroughly done with 208 THE METALLURGY ore no finer than this. The telluride of gold, where segregated in pieces, forms particles of gold in roasting that are too large to be dissolved in the potassium cyanide. The roasted ore, after cooling, is ground to 60 mesh in cyanide solution in Chilean mills, see Fig. 100. The bed of the mill carries a shallow pool of mercury, and large particles of gold are amalgamated and retained here. It has been found, that the extraction is quite as good when the ore is ground to this mesh as when ground to 200 mesh. The mill in the illustration is shown with the surrounding screens removed to expose the rollers. The rollers travel upon a ring-die, the crushing being done between the die and the roller. The ore is fed into the hopper near the top and enters through pipes in front of the rollers. A scraper in front of each roller deflects the pulp in front, but is so set as to escape the mercury pool inside the ring- die. The ore, when ground sufficiently fine, splashes out through the screens, and is conducted by the circular cast-iron launder that surrounds the mill to the point of discharge. The resulting pulp is elevated to classifiers, and separated as usual into sand and slime, the sand being conducted directly to the leaching tank, while the separated slime, after four hours agitation in a tall conical-bottom tank, is drawn to Argall vacuum filter-presses of the Moore type. 47. ACTION OF COPPER IN CYANIDING. Copper in ore is generally understood to be a serious detriment to successful cyanidation, since it passes into solution and consumes cyanide. The difficulty can be overcome, \vhere little copper is present, by running strong cyanide solution into the stream of gold- solution as it enters the zinc-box. This strengthens the entering solution and prevents the precipitation of copper. The copper that enters the solution consumes potassium cyanide. The Hunt ammonium cyanide process. This is a simple process dependent on the fact that ammonium chloride added to potassium cyanide solution has a protective influence upon the cyanide and a solvent action upon copper. At Dale, San Bernardino county, California, occurs a copper bearing gold ore containing the copper in the form of silicate. The silicate of copper, being soluble in cyanide, causes a loss of 7 to 8 Ib. cyanide per ton of ore by the ordinary treatment. By Hunt 's method 8 Ib. quick-lime is added per ton of ore. The ore in the vat is then leached with a 0.15% solution of cyanide to which has been added 6 Ib. ammonium chloride per ton of solution. The solution after OF THE COMMON METALS. 209 contact with the ore is drained, and the ore given six days' continuous contact and washing. The gold-bearing solution can be precipitated in zinc-boxes, but electrolytic precipitation is preferred. In the precipitation, lead anodes and aluminum cathodes are desired. The anodes are peroxidized by dipping into a solution of potassium permanganate before use. A current density of 3 amperes per square foot is employed. The gold, silver, and copper do not precipitate as an adhering coating upon the aluminum cathodes, but fall from them as sludge to the bottom of the tank. The method has been successfully applied to the treatment of copper-bearing mill-tailing dumps exposed for years to the weather. Potassium or sodium cyanide costs 18 to 20c. per pound. Ammonium chloride, commercially called muriate of ammonia, costs as an estimate 5 to 6c. per pound. 48. CYANIDATION OF CONCENTRATE. Two cases arise in respect to the treatment of the sulphide concentrate obtained from gold or silver ores. In the first, the valuable metals enter the concentrate, while the gangue is comparatively barren. In this case only a small percentage of the total ore need be cyanided. In the second, all parts of the ore have value. Here we have the choice between finely grinding all the ore for extraction or on the other hand treating a large part of the ore by ordinary methods of extraction, but giving the concentrate, which is of higher grade, and in smaller quantity, a more careful treatment. The removal of the concentrate from the ore frees the tailing from acid sulphates and base metals which would interfere with cyanidation. Gold and silver-bearing concentrate can be cyanided with a high extraction in some cases. The gold must be fine and the silver present as silver chloride or sulphide. When silver is the principal metal to be extracted, the treatment must be prolonged. Thus with ore re-ground to 100-mesh size, an extraction of 93.5% of the silver and 96% of the gold was obtained in one instance on ore of 130 oz. Ag and 1 oz. Au per ton. This was accomplished by an eight-day treatment by agitation and decantation, which is equivalent to leaching 30 days. Ror high-grade impure concentrate the best method is to give a chloridizing roast, following with a careful and prolonged leaching with cyanide solution. The silver compounds are thus converted into a readily-soluble form, and the ore is rendered porous and penetrable by the solution. Treatment at the Standard plant, Bodie, California. 210 THE METALLURGY concentrate, consisting of iron oxide with a little pyrite is charged in one-ton lots into an ordinary 5-ft. pan (see Fig. 103), lime and water first having been added to dilute the pulp to 45% solid matter. The charge is ground 48 hours, then cyanide is added in quantity sufficient to strengthen the solution to 1.2%. Grinding is continued 24 hours with additions at intervals of quick-lime and cyanide. The lime insures alkalinity, the cyanide maintains the strength of the solution at a fixed percentage. During the grinding an oxidizing action occurs and the pulp changes in color from a greenish shade to a brownish red. It is drawn into a settler, maintained in motion for 24 hours, and is diluted with weak cyanide solution, filling the settler. The whole content is then run to the filter press. The filtrate from the press, containing the gold in solution, is precipi- tated with zinc-shaving. The extraction of metals, 96.8% of the gold and 84.1% of the silver, is high. The cost also, $7.91 per ton, is high, and 18 to 20 Ib. cyanide is consumed per ton of concentrate treated. Treatment at Harrisburg, Arizona. Seventy tons of ore daily were concentrated, yielding two tons of concentrate, while the mill- tailing being of little value was thrown away. The raw concentrate is exposed to the air for a while to partly oxidize it, since it is found that when thus treated it grinds more easily, and the extraction is better. It is then charged to a 5-ft. pan, the charge consisting of 1.5 tons of concentrate and 1 ton of solution containing 6 Ib. quick-lime and 6 Ib. potassium cyanide. The grinding, which takes 8 hp., was continued 8 hours after adding 2 Ib. more cyanide, the temperature is increased some 40. The content of the pan, not too finely ground, is now run into a 15-ton leaching tank. When the slime is partly settled in the vat, dry tailing is sprinkled in. This coarse material facilitates percolation, and we thus have a number of layers of charge interstratified with middling. The tank having been filled, the charge is continuously leached with cyanide solution while the other tank is being filled. Thus each tank is 7 days filling, followed by 4 days leaching, and 3 days washing, before it is discharged. The average extraction is 94%, and the consumption of cyanide is 8 Ib. per ton. The cost of treatment of the concentrate is estimated $5 per ton. 49. THE BROMO-CYANOGEN PROCESS. This method, called also the 'Diehl process,' is practised upon the sulpho-telluride ores of Western Australia. A typical ore contains 50% Si0 2 ; 10% Fe; 7 to 22% CaO and MgO, 3 to 7% S; OF THE COMMON METALS. 211 0.03 to 0.10% Te, and 2 to 3 oz. gold per ton. The ore contains iron, calcium, and magnesium carbonates which give it cement-like properties after roasting, disposing it to cake or set when subjected to the action of cyanide solution in the tank. Since the Diehl process is applied to the raw ore, this trouble is avoided, and the process has became a successful method of treating a certain kind of ore. Fine grinding and amalgamation was tried upon the ore, but an extraction of only 20% resulted. Upon roasting and amal- gamating in pans, the extraction was 44%. Concentration gave poor results owing to the loss of the finely ground almost inpalpable tellurides escaping with the tailings. Ordinary cyanidation gave extractions of 60 to 77%, but with preliminary roasting, it yielded a total of 93 to 94 per cent. It is found that the ore must be finely ground, or slimed as it may be termed, to obtain a good extraction. At the same time we must distinguish between a product that contains many fine hard particles of quartz and one consisting of real slime composed of colloid substances, in which no grit exists. Moreover, the above tests show that only a part of the gold of the ore is in a metallic state, the rest being present as telluride and locked within the crystals of pyrite. Owing to the brittleness of the telluride, the finer the slime the richer is the content of gold. The process. The ore is coarsely crushed and fed to the stamps. The pulp from the stamps is classified, the overflow going to the agitation vats, while the underflow, of coarse sand, goes to the tube-mills for re-grinding. In the tube-mills the sand is crushed so that most of it passes a 200-mesh screen. The product, still containing 5% sandy particles, is again classified, the overflow as before going to the agitation vats, while the sand is returned to the tube-mill for re-crushing. Thus the agitation vats recover only a slimed product, fine enough to pass a 200-mesh screen. The agitation vat (See Fig. 101) is 25 ft. diam. by 8 ft. high. When it has been filled, potassium cyanide is added to produce a 0.22% KCN solution. This is used in the proportion of two of solution to one of the dry slime, and the whole is stirred for a period varying between 16 and 24 hours. This is followed by the addition of bromo-cyanogen to increase the strength of the solution by 0.055% salt used per ton of dry slime. Agitation is continued until the gold is in solution, after which the whole is filter-pressed. Since, by this method of final treatment, but little water is required, the 212 THE METALLURGY process has an important advantage in the country mentioned where water is worth 50c. per 1000 gallons. The pulp is pumped to presses of the Dehne type that have a capacity of three to five tons of dry slime per charge, the thickness of the cake in the recesses varying from one to three inches. The distance-frames in which the cakes form are 40 in. square, there being 30 to 50 cakes in a press. Of the solution 70% is recovered at once ; then a weak solution or water is forced through, and complete washing accomplished by the use of a half ton of solution per ton of slime. Compressed air, under 90 Ib. pressure is next Fig. 101. AGITATING VAT. introduced. This soon expels the water from the cakes leaving only 11 to 15% moisture. The press being opened, the cakes are discharged into cars below and trammed to the dump. It takes two men 30 to 45 minutes to discharge a press, clean the frames, and close it, ready for another charge. The time of filling, using a pressure of 60 to 100 Ib. per sq. in., is 20 to 30 minutes. Wash- water is used 15 to 25 minutes, and compressed air for 10 minutes. The full cycle of operations can be completed therefore in two hours. Two men, using two presses, handle seven charges per 8-hour shift, or 40 tons dry slime per press in 24 hours. The extraction by the use of the press is 90 per cent. Bromo-cyanide is made by dissolving the commercial crystals in water, or, when produced at the works, by adding bromine-water OF THE COMMON METALS. 213 to cyanide solution. The cyanide reacts with potassium bromide as follows : KCN + BrCN = KBr + C 2 N 2 Cyanogen (C 2 N 2 ) in nascent condition acts energetically on the gold as follows : 2Au + C 2 N 2 + 2KCN = 2KAu(CN) 2 Bromo-cyanide does not require the prior aeration necessary in the use of the ordinary solution. It is expensive, however, and has been applied only to sulpho-tellurides that are unattacked, or are attacked but slowly by plain cyanide. 50. SMELTING OF GOLD ORES. Gold may be recovered from its ore by the processes of silver- lead, or of copper-matte smelting. It often is found in copper or lead-bearing ores and when in excess of 0.05 oz. per ton, is paid for by the smelting works at the rate of $19 to $19.50 per ounce. Practically all the gold is recovered in smelting, and this would be the best method of treatment were it not for the high cost of freight and of treatment. If smelted near the mine in a works operated by the mining company, the cost of freight is eliminated. The charge for dry, fairly silicious ores, from Cripple Creek and from Boulder county, Colorado, is from $4 to $10 per ton, according to grade. The low-grade ores are subject to a low treatment rate. On the other hand ore treated by milling and amalgamation, or by cyanidation, while the extraction is less, often yields higher net returns. An example is found in the case of the dry silicious gold ore from Boulder count} 7 , Colorado, containing 0.5 oz. Ag per ton, giving 70% extraction by milling and amalgamation, or of 90% by cyanidation. In comparing the costs we have : SMELTING. 100% of 0.5 oz. Au at $19.00 $9.50 Mining 2.00 Freight 1.50 Treatment 4.00 7.50 Net returns $2.00 MILLING AND AMALGAMATION. 70% of 0.5 oz. Au at $20.50 $7.17 Mining 2.00 Milling 1.00 3.00 Net returns $4.17 214 THE METALLURGY CYANIDATION. 90% of 0.5 oz. Au at $20.50 $9.23 Mining 2.00 Cyaniding 1.65 3.65 Net returns $5.58 From the above comparison it is seen that cyaniding is the most profitable method of treatment for this grade of ore, and at this place. PART IV. SILVER /* PART IV. SILVER. 51. SILVER ORE. The silver minerals of importance in treatment are as follows : Native silver sometimes occurs in the form of flakes or leaves, and as wire-silver and metallic silver adherent to native copper. Native silver can be readily amalgamated, but when present in particles of visible size it is so slowly soluble in cyanide, that practically no extraction can be obtained. Cerargyrite (horn-silver, silver chloride), AgCl, is widely distributed. At mines it is found in the upper oxidized zones. It is probable that much of the so-called chloride ore is really a chloro-bromide (embolite). The ore is readily amalgamated and is free-milling. The silver also is readily soluble in cyanide and in sodium hyposulphite solutions. Argentite, Ag 2 S, is one of the common silver ores. By using chemicals (bluestone and salt) it can be amalgamated in pans, and the silver extracted thus from the ore. It is soluble in potassium cyanide solution. Stephanite, 5Ag 2 S,Sb 2 S 3 ; pyrargyrite, 3Ag 2 S,Sb 2 S 3 ; proustite, 3Ag 2 S,As 2 S 3 ; dycroasite, Ag 3 Sb are silver sulph-arsenides or sulph-antimonides, refractory in amalgamation, even with chemicals, sparingly soluble in cyanide solution, but readily soluble in a solution of mercurous potassic cyanide. Finally we have those silver sulphides that contain also copper. These are polybasite, 9(Ag 2 Cu)S(SbAs) 2 S 3 , and tetrahedrite (gray copper ore, fahlerz), 4CuFeAg 2 (HgZn)S,(SbAs)S 3 the most complex of all, in which the silver varies from 0.06 to 31%, being higher in the arsenical and lower in the antimonial varieties. These sulphides are refractory to any amalgamation method, and because of their copper content are precluded from treatment by cyanide, even when roasted. This does not interfere with treatment by hyposulphite lixiviation after roasting. A number of rare minerals containing silver could also be enumerated, but for the metallurgist the minerals above named are the important ones. 218 THE METALLURGY Silver ores in general contain but a small percentage of precious metal. They are composed mostly of gangue (waste matter of the ore), and many are treated that contain less than 0.1 to 0.2% silver. Thus we have at the Comstock Lode, Nevada, silver in native form and as sulphide, but oxides of iron and manganese with the associated sulphides, pyrite, blende, galena, and chalcopyrite. At the Ontario mine, Park City, Utah, the silver occurs as argentite and tetrahedrite in a gangue of quartz and clay associated with a little of the heavy minerals blende and galena. These sulphides carry silver which is recovered with the concentrate in case of concentration. 52. THE EXTRACTION OF SILVER FROM ORES. Silver is extracted from its ores by amalgamation in pans, by hydro-metallurgical methods, and by smelting. Amalgamation is now used less in America than formerly, smelting having taken the place, not only because of better extraction and lower treatment rates, but because ores are more easily shipped to smelting works in result of the extension of railroad facilities. Certain ores contain the silver in a form suitable for cyaniding, and the hydro- metallurgical method in consequence is coming forward. The other methods have partly dropped out of use. No reason appears why hyposulphite lixiviation should not revive under the stimulus of the recent methods of agitation and filter-pressing. The patio process, formerly much practised in Mexico where conditions favored, has been superceded by cyaniding in many cases, on account of the lower cost of operating the latter process, but in the past, large quantities of silver have been extracted by the patio process. 53. SILVER MILLING AND AMALGAMATION. The silver ores suitable to treat by milling and amalgamation are those that contain the metal in such form as to be acted upon by mercury when assisted by agitation, heat, and certain chemicals. The ore is first crushed fine by stamps, as in gold milling, then treated for several hours in pans, the reactions being slow compared with those of the amalgamation of gold. In gold milling, the greater part of the gold can be arrested on an apron-plate during the few seconds in which the ore is passing; while in silver milling, the ore-pulp has several hours contact with mercury aided by heat and chemicals, and is but slowly amalgamated. In gold milling, ore containing 0.5 oz. Au per ton can be profitably milled. In OF THE COMMON METALS. 219 silver milling, ore of equivalent value would contain 20 oz. Ag per ton, or 40 times as much metal. Thus is seen why so much time is allowed in silver milling, and why so many precautions must be taken to be sure that all metal possible is recovered. Several ounces of silver per ton often remain in the tailing. The silver metals suited to pan amalgamation are cerargyrite (horn silver, silver chloride), native silver in flakes, wire, or other forms, and certain silver sulphides, notably argentite (Ag 2 S). When the ore is refractory, containing arsenical and antimonial sulphides, and especially containing tetrahedrite, galena, or blende, it is necessary to roast with salt, setting free the silver or converting it into the form of a chloride, which becomes susceptible to amalgamation. There is no sharp line of demarkation between free- milling and roasting milling ores. Often the upper part of a vein is free-milling. In depth base metals and sulphides begin to come in, and it finally becomes necessary to roast the ore. The best extraction therefore is obtained from decomposed or oxidized ore, in which the silver minerals occur in a form that renders possible the action of the mercury. There are few deposits of oxidized ores containing silver chloride and native silver that as a whole are suitable for free silver milling. Such ore, so far as silver chloride is concerned, can also be treated by cyanidation, but the latter method would not recover native silver. Arsenic and antimony compounds interfere with amalgamation by fouling the quicksilver, checking the reactions of the chemicals added to promote amalgamation, and by carrying off silver, which is incapable of being amalgamated with them. We may divide the methods of silver-milling into (1) Wet silver-milling, or the Washoe process, which includes (a), the tank mill, and (b) the Boss process. (2) Combination process, combining the gold-mill and the silver-mill. (3) Dry silver-milling, or the Reese River process. 54. WET SILVER-MILLING WITH TANK SETTLING. This is also known as the Washoe process, receiving the name from the place where it was perfected for the treatment of ores from the Comstock Lode, Nevada. The process is applicable to the so-called free-milling ores, in which the silver occurs native, as chloride or in small amount as argentite. The ore should be free from lead and from any tough clayey gangue. In wet silver-milling, when settling tanks are used, the process OF THE COMMON METALS. 221 consists in coarse-crushing the ore, stamping it fine, and collecting it in settling tanks. The crushed sand is ground in amalgamating pans using mercury to collect the silver. The sand is separated from the silver-bearing mercury in settling pans and is rejected. The amalgam is strained from the mercury, retorted, and the retort- residue melted into silver ingots. Gold present in the ore is recovered as well as the silver. The process resembles gold milling except that amalgamation and the removal of the amalgam is effected in pans. Fig. 102 is a sectional elevation of a wet-crushing tank-mill for the treatment of free-milling silver ores. The ore from the mine, as in gold milling, is dumped from the tram-car over a grizzly or bar-screen a, the bars being set 1*4 in. apart. The fine falls into the sloping bottom storage bin c, while the lump ore is received upon the feed-floor at the mouth of the Blake crusher 6, and after being coarsely crushed, joins the fine in the bin. The coarse crushing is done during the 10-hour day-shift. The ore passing to the battery is fed by an automatic feeder d such as is used in gold milling. The flow of ore to the automatic feeder is controlled by gates at the inclined outlet chutes and the feeder is kept full to supply the stamps e uniformly as required. Water (6 to 8 tons per ton of ore) is at the same time supplied in the mortar, and the suspended ore forms a pulp, which is splashed through the 30-mesh screens by the motion of the stamps. A double- discharge mortar of the type suited to silver milling is seen at Fig. 48. The large screen-opening possible with a double discharge favors a more rapid pulverization than would be possible with a single-discharge mortar. (For further particulars of gravity stamps see the detailed description under 'Gold Milling.') The pulp flows by launders / into the settling boxes or tank g 1 ft. square by 3 ft. deep. There is a double row of these tanks, 20 in a row, occupying the length of the mill in front of the stamps. The flow of the pulp is from box to box until it goes by launder to a settling-pond outside the mill. Most of the solids settle in the first boxes, a further portion dropping in the succeeding ones, and the turbid water passing to the pond. Here it has its final chance to settle before running to waste, or it may be again used in the mill if water is so scarce that it pays to do this. The settled slime is dug from the pond at a later time and treated like the rest of the crushed ore. A variation of this method, shown in Fig. 102, consists in conducting the flow from the last box g' by an inclined elevator to 222 THE METALLURGY a tank h situated in front of and above the battery, the dirty water being again used for stamping. When the first settling box is full the flow of pulp is by-passed into the next one. The contents of the full box is shoveled upon the floor adjoining, and thence taken as needed to the amalgamating-pans q. The emptied box has the flow of the last one turned into it, thus making it the last in the series, and the launders are so arranged that this can be done. The ore, thrown out upon the floor, is fed directly into the pan, Fig. 103. COMBINATION AMALGAMATING PAN. or loaded into the tram-car seen in Fig. 102, and conveyed to the pan. Fig. 103 represents two of these pans. One is shown in section, the other in elevation. The pan shown in section is 5 ft. diam. by 30 in. deep, and has a cast-iron bottom and the sides made of wooden staves. It is furnished with a central sleeve or cone through which rises a shaft carrying a cylindrical casting called a spider, which becomes bell-shaped and broadens into feet below. The spider carries, bolted to the feet, a flat cast-iron ring called a muller, and to the underside of the muller is attached six shoes OF THE COMMON METALS. 223 or plates of chilled cast-iron 2 1 /2 in- thick. The spider, muller, and shoes are raised or lowered as desired, by means of a hand-wheel and screw at the top of the shaft, which is driven by bevel gearing from the horizontal shaft and pulley below. Upon the bottom of the pan rest chilled cast-iron plates or dies that furnish the lower or fixed grinding surface. The shoes attached to the muller revolve 60 rev. per min., and rubbing upon the dies, grind the ore. In working the pans, the shoes are raised % i n - from the dies and set in motion, the pan is partly filled with water, and 3000 Ib. of the damp pulverized ore are shoveled in. The ore and water nearly fill the pan and the mixture is stirred until it is of the consistence of honey. The motion establishes a movement or current of pulp beneath the muller toward the periphery. At the periphery it rises, flows toward the center, sinks, and passes again under the shoes. To assist the action, the rising pulp is deflected inward by cast-iron wing-plates. After thorough mixing in the pan the shoes are lowered until they touch the dies, and grinding goes on for 1% hours, the content of the pan being meanwhile heated nearly to boiling by steam under pressure from a pipe that dips beneath the surface of the charge, the pan being covered. Sometimes the pan is provided with a double bottom into which exhaust steam from the engine is introduced. After grinding, the shoes are raised and 300 Ib. mercury (10% the weight of the ore) is added, by sprinkling it through a fine strainer. The mixing is then continued four hours. The mercury takes up silver most rapidly at first, but the action afterward slackens. At the end of the first hour, in one instance, 74.7% was amalgamated, of the second hour 76.3%, of the third 77.7%, and at the end of the fourth, 81.0% the silver contained in the ore. After the fourth hour no further silver amalgamated. The globules of mercury suspended in the pulp take the silver as they come in contact with it. Care is taken to have the pulp of the right consistence so that mercury will not settle out. This condition is shown when a wooden stick, dipped in the pulp and withdrawn, is found to be covered with a thick mud in which are disseminated minute globules of mercury. If the ore is refractory, salt and copper-sulphS^ are advantageously added at the beginning of grinding to accelerate the reactions, promote amalgamation, and increase the yield of silver. The time required for grinding varies from four to six hours. In mills treating chloride ore, the charges are finished in four hours without grinding. In treating refractory ores, the total 224 THE METALLURGY time for a charge becomes six to eight hours, of which time four hours are required for grinding. The charge above treated having been amalgamated, the pan is ready to empty into the settler r. About 15 minutes before the discharging, the speed of the muller is reduced to 40 rev. per min. and the pan filled to the top with water. A plug closing the discharge opening at the bottom of the pan, seen at the left in section, Fig. 103, is pulled out, and the entire content run by launder to an 8-ft. settler, at a lower level, shown at the right of the amalgamating pans in Fig. 102. Emptying the pan and washing it with a hose takes half an hour, after which time the plug is replaced, and the pan is ready for another charge. Thus the total time for the cycle of operations described is six hours, making it possible to treat four charges daily. The reactions that take place in the pan are as follows : Native silver in threads, films, flakes, or grains readily combines with the mercury and forms an amalgam which contains a large excess of mercury. Silver chloride in contact with the mercury decomposes as follows : (1) 2AgCl + 2Hg = Hg 2 Cl 2 + 2Ag The metallic silver liberated amalgamates with additional mercury. The particles of iron, abraided from the stamps and the bottom of the pan, decompose the mercury salt and liberate the mercury as follows : (2) Hg 2 Cl 2 + Fe = FeCl 2 + 2Hg In many so-called free-milling silver ores argentite is contained which in part is decomposed by mercury as follows : (3) Ag 2 S + 2Hg = Ag 2 + HgS The sulphide of mercury thus formed is lost. We have already stated that chemicals, notably copper sulphate and common salt, are added to promote the decomposition of the silver sulphide. There is added in the amalgamating pan from 6 to 18 Ib. salt and from 3 to 9 Ib. copper sulphate per ton of ore treated. The reactions as generally given are the following: (4) CuS0 4 + 2NaCl = Na 2 S0 4 + CuCl 2 The chloride of copper acting on the silver sulphide decomposes it : (5) Ag 2 S + CuCl 2 = CuS + 2AgCl The silver chloride amalgamates as shown by reaction (1). OF THE COMMON METALS. 225 The complete separation of the mercury with the silver-amalgam is effected in the settler, there being one settler provided for two amalgamating pans. The settler is a cylindrical pan 8 ft. diam. by 3 ft. deep, 3 times the capacity of the amalgamating pan, but of similar construction, as shown in Fig. 104. It is required to do no grinding, but to gently agitate the pulp with the wooden shoes with which it is provided. The shoes nearly touch the bottom of Fig. 104. EIGHT-FOOT SETTLER. the settler, the exact height being adjustable. The grooved border at the bottom just within the sides of the settler has a slight grade to the outlet and mercury-well at the left. The mercury settles from the pulp, flows to the lowest point and stands at a height that balances the hydrostatic head of the content of the pan. Since the specific gravity of mercury is 14 and the content of the settler approximately 1.5 the height of the mercury is a little less than 4 inches. The bottom outlet-hole of the well is plugged. At 226 THE METALLURGY different heights in the side of the pan there are provided openings that are kept closed by plugs. When the plugs are withdrawn the tailing and water, free from mercury, passes out of the pan. The shoes of the settler having been set in motion, at the rate of 15 rev. per min., and raised 8 in. above the bottom, the contents of the two pans are run in, as has been described. Water is then added to within 6 in. of the top, greatly thinning the pulp, and filling the settler. After half an hour the shoes are gradually lowered until, at the end of two hours, they nearly touch the bottom. The purpose of the agitation is to keep the lighter portion of the ore (now called the tailing) in suspension, while the silver- bearing mercury, the heavier particles of sulphide, and the particles of iron from the stamps collect at the bottom. The stirring is continued 3% hours, after which the highest plug in the side of the settler is removed, and the turbid water containing tailing is allowed to escape by launder, a stream of clear water being mean- while allowed to flow through. The plugs are then withdrawn one by one until the settler is emptied of all the content except the heavy portion containing sulphide, iron particles, and the mercury. Emptying takes half an hour, and the cycle of operations becomes six hours as in the case of the amalgamating pan. Since escaping tailing contains sulphide, it may be run over riffles, or blanket- lined launders, before running to waste. In Fig. 109 is see a system of pans and settlers. The silver-bearing mercury or diluted amalgam, a mixture of silver-amalgam and mercury, collecting in the mercury well, over- flows by an escape-opening indicated in Fig. 104. From the opening it passes by a half-inch pipe to the amalgam safe shown at the right of the settler, Fig. 102. The safe, arranged to prevent theft of the amalgam, is shown on a larger scale in Fig. 105. The amalgam and mercury enter a conical canvas sack or filter. The mercury oozes through the pores of the canvas while the amalgam containing as little as 14% silver is retained. Occasionally, after amalgam has accumulated, the sack is squeezed between the hands to remove the surplus mercury, and the compressed amalgam containing 20 to 28% silver, is reserved for retorting. The mercury flows out at the bottom through an outlet provided, as seen in Fig. 105, and is collected at a lower level in the boot w, Fig. 102, of the mercury elevator, shown at the right. The elevator discharges to a mercury tank s commanding the amalgamating pans, to which it is delivered as needed through pipe shown in the figure. Over the stamps and the pans are seen the overhead tracks that carry OF THE COMMON METALS. 227 crawls by which the heavy parts of the machines are lifted or transferred. This facilitates the work of repairs and replacements. The loss of mercury is commonly 1 to 1.5 Ib. per ton of ore treated. A part is lost in handling, but the principal cause of loss is the flouring which causes the mercury to escape in the tailings. The loss is greater with talcose or clayey ores, and in those carrying cerussite, chalcopyrite, or galena. Loss is caused by grease coating the particles of mercury, in case this enters the ore from the machinery. Treatment of the amalgam. Since the weight of metal recovered in silver milling is much greater than in. gold milling, the retorting Fig. 105. AMALGAM SAFE. of amalgam must be performed on a larger scale. Fig. 106 shows a sectional elevation and a plan of a combined retorting and melting furnace with the overhead crawl and chain-blocks by which the large melting crucibles are lifted from the fire in the melting furnace and transferred for pouring. At the left is shown in the elevation a cross-section of the cast-iron cylindrical retort which is 10 in. diam. inside by 28 in. long, resting upon arched cast-iron supports. There is a horizontal pipe, and, not shown in the illustration, a vertical water-cooled pipe in which the mercury .condenses and from which it falls into a tub of water below. As seen in the plan, the front end of the retort is provided with a cover which can be securely clamped in position. 228 THE METALLURGY The charge of amalgam, containing 20% mercury, should weigh 500 Ib. and only half fill the retort. After filling, the cover is clamped on, first luting the joint with flour paste. A wood fire is started on the grate under the retort. The temperature is kept Fig. 106. HORIZONTAL RETORT AND MELTING FURNACE FOR SILVER MILL. low at first, increasing to a red-heat at the end, % to % cords of wood being used. The operation lasts 10 to 14 hours, care being taken not to heat the retort rapidly, nor, for fear of blistering it, to raise the temperature too high. The fire is then allowed OF THE COMMON METALS. 229 to burn down, and the retort to cool. The lid is taken off and the silver residue removed. The melting of the residue is performed in plumbago crucibles, adding soda and borax as fluxes. The crucible is set in the wind- furnace (shown in Fig. 106) at the right and supported upon a brick resting on the grate-bars. It is surrounded with coke which burns by natural draft, the smoke escaping by the flue at the back to the stack. When the charge is melted the crucible and contents are removed from the fire with basket-tongs, which fit the crucible and clasp it firmly so that it can be lifted by the chain- hoist, transferred by the crawl to the ingot, and poured. These molds, 11 in. long by 4% in. wide and deep, hold 1000 oz., or 70 lb., silver. The settler tailing contains heavy unaltered ore which it may pay to concentrate on a Frue vanner of some similar table. At some places, sluices are used, two or three in parallel. These may be 2 in. high, 20 in. broad, and 1800 ft. long. The bottoms of the sluices are covered with coarse blanketing which can be easily removed and washed to recover the concentrate or particles of amalgam that settle into the meshes of the blankets. Costs. The cost of pan-amalgamation with tank-settling ( Washoe process) per ton of ore treated is : Power $0.087 Labor 0.361 Chemicals (salt, acid, bluestone) 0.465 Loss of mercury 0.750 Wear of pans 0.200 Wear of dies and shoes 0.400 Oil, interest, and superintendence 0.100 Total cost per ton $2.363 One notes in particular the larger cost of supplies (chemicals, mercury, and castings) compared with like items in gold milling. 55. THE BOSS PROCESS OF SILVER MILLING. This system, originated by M. P. Boss, a California engineer, differs from the Washoe process in being continuous and generally requiring less labor. However, the Allis-Chalmers Co., have designed for the Washoe process a wet-crushing mill in which the settling-boxes have sloping bottoms, so arranged that the content is transferred to the pans with but little labor. This takes away 230 THE METALLURGY the advantage urged in favor of the Boss system. It may be added that the settling of the pulp in large tanks, combined with a mechanical system of excavating the content as in the cyanide process, ought to be efficient and labor-saving. The Boss system may be applied to free-milling ores and to rebellious ores that need to be first roasted. Fig. 107 is a sectional elevation and Fig. 108 a plan of a 30-stamp mill arranged according to the Boss system. Fig. 108 shows that the ore, coming in by two tram-tracks, is dumped over grizzlies, the fine falling into the storage bins, while the coarse ore falls upon a Cross Section Fig. 107 BOSS-PROCESS SILVER MILL (ELEVATION). feed-floor between them. Here the coarse is crushed in a Blake crusher, discharging to the storage bin below. From the bins the ore is fed automatically to the stamps, and wet-crushed. The screen-discharge from each ten stamps flows through pipes to a pair of 4-ft. grinding-pans in series or six pans in all (shown in the plan, Fig. 108) where the ore is finely ground. It overflows from the top of the second pan into a common launder and is carried by pipe to the next series of pans at a lower level, flowing continuously through the line of pans and settlers there situated. There are ten amalgamating pans and four settlers. They are connected one to another near the top by 4-in. nipples, which permit the flow of a pulp that is thinner than that in the tank system. 232 THE METALLURGY The amalgamating pans are provided with double bottoms for the admission of exhaust steam from the engine, to heat the pulp as in the tank system. The chemicals are fed continuously to the first two pans in the series by two 'chemical-feeders,' and the mercury also is supplied continuously to all the pans by pipes leading from the mercury distributing tank. Whatever mercury settles in any of the amalgamating pans, or in the settlers, is collected in the mercury well of each, and the overflow from all the wells flows through suitably arranged pipes to the strainer. The strained mercury is elevated to the distributing tank to be again used. The continuous flow from the last settler goes to three settling cones in series. The spigot-discharge from each cone goes to a separate table, and the overflow to a fourth one. The fifth table is used for re-treating the heads of the first four. Steam siphons are provided for cleaning out the pans, or for carrying the pulp past any pan when necessary to cut out the pan for repairs. The main line shaft runs directly under the pans and s.ettlers, each of which is driven by a friction clutch. Thus any machine may be stopped in case of an accident, or for cleaning, without having to stop the whole line. All the battery-flow, including the slime, must pass through the system, and no ore passes untreated as in the tank system. The loss of mercury is small and it is possible to correctly sample the tailing, which are advantages over the older system. In spite of the thin pulp used it is claimed that an ore adopted to pan-treatment can be more successfully worked by the continuous system. 56. THE COMBINATION PROCESS OF SILVER MILLING. This process is used on ores carrying silver, gold, and sulphides of the heavy metals, such as galena, blende, and pyrite, and sulphides which contain silver and gold. It is necessary that the silver, not in the sulphides, be amalgamable, as is silver chloride, argentite, or native silver. The process consists in wet stamping the ore, running the pulp over apron plates as in gold milling, concentrating the sulphides, and, as in the Washoe process, pan-amalgamating the tailing and treating the amalgam to recover the silver and gold. Compared with either wet or dry silver-milling the process has much to commend it. The ore being refractory, the wet process would recover little value. The tonnage stamped by the dry-method with roasting would be low compared with wet-stamping which is one-and-one-half to twice as rapid. It is true that by dry-stamping 234 THE METALLURGY and roasting we are able to extract at least 10% more metal than can be obtained by raw amalgamation, but this is offset by the cost of treatment and the loss of precious metal in roasting. The combination process also saves lead and removes galena, sulph- arsenides, and sulph-antimonides, all of which tend to foul and cause the loss of mercury. Such minerals are not amenable to amalgamation, and by removing them for separate treatment there results a cleaner or higher-grade bullion. Manganese minerals that consume chemicals in the pan are also removed by concentration. Because of the advantages, it is probable that the field of the combination process will extend. Certainly it will be used in preference to roasting and amalgamation, a method rapidly losing ground ; and in some instances it may take the place of lixiviation, so far as respects the pan-amalgamation feature of the treatment. The combination mill. Fig. 109 is a perspective view of a 10-stamp combination mill. It shows at the left a tram-car on the upper stage which is about to be emptied over the grizzly. The fine falls through the grizzly while the lumps fall upon the feed-floor and are shoveled into a Blake crusher, crushed to 1-in. size, and united with the fine in the inclined bottom storage-bin below. Let us suppose we are to treat an ore, in part oxidized, but containing the heavy minerals of lead and copper, with p3^rite, arsenides, and manganese minerals. The ore contains the precious metals, a gangue of quartz, calcite, and a little clay, and disseminated through it gold and the amalgamable silver minerals cerargyrite, argentite, and native silver. The purpose is to save the precious metals by plate and pan-amalgamation, and the heavy minerals with silver and gold by concentration. Some of the silver and gold escapes recovery and is lost in the tailing. Since sulph- arsenides and manganese minerals are mostly removed, they do not interfere with subsequent pan-amalgamation where arsenic would sicken the mercury and manganese consume chemicals. The ore and water are fed automatically to a ten-stamp battery, each stamp crushing 4 tons per 24 hours to pass a 30-mesh screen. The pulp issuing from the mortar flows over two apron-plates (one for each 5-stamp mortar) and a part of the gold and silver is recovered. The flow is distributed evenly to four Frue vanners at a lower level, the concentrate (10% of the whole) being separated to ship to smelting works, while the tailing is carried to the ten settling boxes in a double row. These are seen at the left of the pans. The distribution is into a double launder between the OF THE COMMON METALS. 235 two rows. By drawing the plugs in the bottom of the launder, the flow can be directed into any box desired. From this point on, the operation is conducted as described for the Washoe process. There are four amalgamating pans and two settlers. Bluestone and salt are used to decompose the argentite. Mercury or amalgam escaping the apron-plates finds its way into the settling boxes and thence to the pans, and is more thoroughly recovered than would be possible if the recovery depended upon obtaining it in the vanner concentrate as in gold milling. The four 5-ft. combination grinding and amalgamating pans each treat 3000 Ib. per charge, and with a 4-hour treatment, this equals 36 tons daily, which with the 4 tons of concentrate already mentioned is a 40-ton output of the mill. Some ores, not so readily treated, take 6 to 8 hours, and lessen the capacity of the mill accordingly. The recovery of precious metals in a certain ore, containing 0.40 oz. Au and 9 oz. Ag per ton, was as follows, stated in percentages : Gold Silver Recovered on the apron plates 22 3 Recovered on the Frue vanners 28 32 Recovered in amalgamating pans. . . ,32 35 -Lost in the tailing 18 30 100 100 The total recovery of the gold was accordingly 82%, and of the silver 70%. Of the lead and copper 85% was saved in the concentrate. The tailing contained 0.1 oz. Au and 3 oz. Ag per ton. Concentrating adds but little to the cost of the pan amalgamation, and $3 per ton may be taken as a fair estimate of the cost of combination milling. 57. CHLORIDIZING ROASTING OF SILVER ORES. Silver ore containing sulph-arsenides, sulph-antimonides, or tetrahedrite, cannot be treated directly by amalgamation nor by any of the lixiviation methods. Such ore must be subjected to a roast with salt to convert the silver into a chloride, before it can be successfully treated by any of these methods. The above minerals are often accompanied by pyrite, blende, chalcopyrite, and galena. Preliminary to roasting, such ore is dry-crushed, either by rolls or by stamps. Ores containing galena and blende are preferably crushed to 40-mesh size, those having pyrite to 8 to 10 mesh. The roasting is done in a reverberatory furnace, and requires 236 THE METALLURGY the use of salt. There must also be 3 to 8% pyrite present to furnish sulphur for the reactions, and if the ore does not contain this, it must be added. If more than 8% sulphur is present, the percentage is reduced to that point by roasting afterward, before the salt is added. The amount of salt required varies according to the quantity of copper and iron sulphides present which consume the evolving chlorine. The roasting operation is at first an oxidizing one conducted at the temperatures specified in the chapter on roasting. The action is chiefly upon the heavy metals, converting them into either oxides or sulphates. It may be divided into three stages (1) the kindling, (2) the desulphurization, and (3) the chlorination of the ore. In the first or kindling stage we find the loosely held sulphur being driven off, and the ore taking fire, producing a blue flame. In the second stage, the air oxidizes the sulphides, and particularly the newly formed iron sulphide. Reacting upon the sulph-antimonides and arsenides, it volatilizes them and removes them from the ore. Copper and iron sulphates are also formed, the later according to the following reaction: 3FeS + 110 = 2SO 2 + Fe,0 3 + FeS0 4 In the third stage, at 590 C., the suiphate formed, reacts upon the salt, thus : 2FeS0 4 + 2NaCl = Na 2 S0 4 + 2FeO + S0 2 + 2C1 The odor of sulphur dioxide and chlorine is pronounced. The chlorine thus liberated acts at once upon the silver compounds and converts them into chlorides. Zinc blende becomes oxide and zinc sulphate, while sulphur dioxide escapes. Galena and zinc sulphate remain inactive and fail to decompose the salt. They roast slowly, while pyrite, in presence of salt, decomposes quickly, and generates chlorine at a period in the roasting when neither the blende nor galena are sufficiently oxidized to expose silver to the action of the chlorine. If therefore the salt is mixed with the ore at the battery, the chlorine generated by the reaction of the ferrous sulphate and salt is lost, an imperfect chlorination results no matter how long roasting is continued, nor how much salt is added. Hence in roasting an ore containing blende and galena it is of the greatest importance to add the salt later and not at the battery. On the other hand if the roasting continues until the sulphides are well oxidized, the iron sulphate decomposes and no chlorine is generated OF THE COMMON METALS. 237 and again we have a badly chloridized ore. The desirable time to add the salt is after continued roasting at a low heat that does not break up the iron sulphate. This is shown when the black color of the ore changes to brown, but shows still the presence of black particles. A distinct odor of chlorine is then to be noticed, due to the decomposition of the salt. The best results could be obtained by adding a mixture of green vitriol (ferrous sulphate) and salt; but the ore would hardly justify the expense. For moderate quantities of ore, the chloridizing roast is performed in reverberatory furnaces. For large quantities, roasting is cheaper when conducted in the mechanical roasters. The salt is added to the dry ore at the time of charging, if the percentage of sulphur is suitable, or later if the excess of sulphur must be first removed by roasting. The temperature is increased only gradually to kindle or start the ore to burning and to begin oxidation. As the temperature rises oxidation and the formation of sulphates occur, and at the necessary high temperature these act upon and decompose the salt and chloridize the ore. It is not considered necessary to continue the roasting to convert all possible silver into chloride, but to withdraw the charge while hot before this stage is reached. During the gradual cooling (12 to 30 hours) further chloridizing proceeds, due to the action of the free chlorine, with which the ore is saturated, acting on the undecomposed silver sulphide. This may increase the cloridiza- tion 10 to 40 per cent. Upon completion of the operation of 'heap chlorination', as it is called, and with ores containing copper chloride, a wetting down or sprinkling causes an additional chlorination of 3 to 6%. Thus at the Lexington mill, Butte, Montana, the ore, after roasting in a Stetefeldt furnace was chloridized to 65% after two hours in the heap to 75 or 80%, and at the end of 36 hours to 92% of the silver content. The loss in silver by volatilization, when the ore has been properly and carefully roasted, should not exceed 8% except in presence of volatile elements like arsenic, antimony, selenium, or tellurium. If, however, the roasting is completed at a high temper- ature the loss may rise to 18 per cent. The most difficult, and at the same time the most important process for the treatment of silver ores by wet-methods, is undoubtedly chloridizing roasting. It is always the safest plan for the operator to roast as thoroughly as possible. If the ore is well chloridized sodium hyposulphite extracts all the silver chloride. 238 THE METALLURGY A high chloridization does not necessarily involve a high loss by volatilization. 58. DRY SILVER-MILLING (REESE RIVER PROCESS). This process for the treatment of rebellious silver ores, in which the metal is so locked up as to require roasting before it can be amalgamated, was developed at Reese river, near the Comstock Lode at Virginia City, Nevada. The ore contains silver sulphide, particularly the antimonial sulphides, and the sulphide of the base metals such as copper, iron, zinc, and lead. Galena, however, if present exceeding 5 to 10%, renders the ore unsuitable for chloridization. The treatment in brief consists in dry-crushing and roasting the ore, then amalgamating in pans to recover the silver and gold. The dry-crushing is done either with rolls or stamps. Crushing with rolls is described in the chapter on 'Crushing.' If dry stamping is employed the work is done in the dry-crushing silver mill, Fig. 110 and 111, here described. Fig. 110 is a sectional elevation, and Fig. Ill a plan of a 20-stamp mill. The ore is dumped from the tram-car over the grizzly 6, which separates the fine from the lumps which roll to the mouth of the crusher c and are crushed to l^-in. size. The crushed product falls with the fine into the storage bin d. From this the ore is fed continuously throughout the 24 hours by means of an automatic feeder e, which supplies a revolving drier / that is 18 ft. long. The drier is provided with a fire-box g, the smoke passing through a flue to a separate stack. Flue-dust is retained in the flue. The ore from the drier containing less than 1% moisture, slides by an inclined launder or feed-chute to the automatic feeders i of the stamps. The fine particles drive through the screens, while the coarse part drops back upon the die to be further crushed. Naturally the process produces dust, and to prevent escape into the mill, the mortar is housed or boxed. There is also an exhaust-fan connected with the housing by which the dust is drawn to a dust- chamber and settled. The finely ground ore passing through the screens, drops to the screw-conveyors, one at the front of the other at the back of the mortars, and is conveyed by them to the hopper or boot of the belt-elevator fc. It is customary to feed into the last battery the rock salt required by the ore. Being finely crushed it can be intimately mixed with the ore. The elevator raises the mixed ore and salt to the conveyor m. Thence the screw conveyor n delivers it into the 240 THE METALLURGY hopper of the roaster. (In Fig. 194 is shown the manner in which the transfer from ra to n is effected, the delivering conveyor being shown at the right, and the receiving one at the left.) From the feed-hopper the ore slides into the White-Howell roasting furnace o (See also Fig. 34), where it receives the chloridizing roast. In the plan of the mill, Fig. Ill, is to be seen the flue-chamber, and a stack like that for the ore-dryer. In the chamber the dust produced in roasting settles and is recovered. The roasted ore, discharged from the roaster, is received into one of the hoppers r. When either hopper is filled, the content is drawn off, moistened with a little water to prevent dusting, and stored on the cooling floor s. The level of the floor s is slightly above that of the pan- floor where the man in Fig. 110 is seen standing. The tracks on the pan-floor-level are arranged so that a tram-car can take the ore from the cooling floor to any of the ten amalgamating pans t. The amalgamation of roasted ore is performed in pans having wooden sides (See Fig. 103.) as in the Washoe process. Water is run into the pan first, and while in motion, the 3000-lb. charge of ore is shoveled in until the pulp is of the consistence of honey. If the ore is imperfectly roasted, salt and bluestone are added to decompose the silver sulphide. The free chlorine in the ore would be consumed by contact with the iron surfaces of the pan, and hence the wooden sides are provided. After a time, the iron particles that have come from the stamps are thus removed and prevented from acting upon the mercury and converting it to mercury chloride in a way that would be detrimental to the process. When the ore has been ground two hours the shoes are raised about half an inch from the dies, and 300 Ib. mercury is added. If mercury were added during the grinding it would flour seriously. The shoes and muller intimately mix the pulp and mercury. Amalgamation proceeds rapidly at first, but more slowly toward the last, the silver chloride being reduced to metal as in the Washoe process. The iron present reduces the higher chlorides of copper and iron, and precipitates mercury from the mercurous chloride. The mixing is of six hours duration. Water is added and the entire content of the pan is run into the settlers u Fig. 111. Here it is diluted with water and the mercury and amalgam settled. The tailing is decanted, and the amalgam is recovered and treated as described under the Washoe process. The recovery of silver is thorough, being in some instances 97%. The loss of quicksilver is slight, being less than i/ Ib. per ton of ore. At the Lexington mine, ore containing 28.5 oz. Ag and Fig. 111. DRY-CRUSHING SILVER MILL (PLAN). 242 THE METALLURGY 0.58 oz. Au per ton yielded 93.3% of the silver and 60% of the gold after roasting. The loss of silver in roasting, however, was 1%, and that of gold 20%. This gives in actual recovery, 86.8% of the silver, but only 48% of the gold. The cost of dry stamping followed by chloridizing roasting may be considered $6.48 per ton of ore treated, this being an average of results in three mills that vary but little from one another, situated in different parts of the Rocky Mountain country. Upon comparing the recovery at the Lexington mine with one of 60% of the silver and gold as obtains by direct amalgamation of the raw ore, we have, with the price of silver 60c. and of gold $20.65 per ounce, the following result: Raw Roasted Treatment $2.46 $6.48 Loss of silver 6.84 2.14 Loss of gold 4.79 4.79 $14.09 $13.41 It is seen in the extraction represented, that there is but little advantage in roasting over the raw treatment of ore of this grade, to say nothing of the extra cost of installing a roasting plant, and the large mill needed for the tonnage. Of course, in case ore is so refractory that only a low extraction of metal is possible, a chloridiz- ing roast may be the alternative. While rebellious ores have been worked by dry silver-milling, it pays in many cases, where accessible to smelting-works, to ship ore rather than to mill it. If the heavy part of the ore carries most of the value, it is better to concentrate it and submit to the loss in the tailing to gain the advantages of treating a concentrated ore. Such problems are solved on a trial scale before undertaking the erection of a plant. Comparing dry silver-milling with smelting, where the smelter pays for 95% of silver and for the gold at $19.50 per ounce, and makes a freight rate of $5 and a treatment rate of $8 per ton we have the following condition : Smelting Dry Silver-Milling Treatment $8.00 $6.48 Freight 5.00 Loss in silver 0.86 2.14 Loss in gold 0.67 4.79 $14.53 $13.41 OF THE COMMON METALS. 243 This indicates a gain of $1.12 in favor of roasting-amalgamation over smelting. On the other hand the cost of erecting an expensive plant is saved by shipping the ore to the smelting-works as is today generally done. 59. SILVER MILLING AT BLACK PINE, NEVADA. The ore of the Combination Mining & Milling Co., at Black Pine, Nevada, contains silver in malachite and tetrahedrite, with a quartz gangue, and no lead or zinc minerals. The process consists in milling and concentrating the ore to separate the refractory tetrahedrite from the free-milling malachite and roasting and pan-amalgamating the concentrate. The ore is partly sorted underground, and upon hoisting, dumped over grizzlies in a rock-house, the fine falling into a hopper-bottom bin, to be thence hauled by team to a sloping bottom storage-bin near the mill in which a good supply can be carried. From this bin the ore is trammed to the mill, and dumped over a grizzly with one-inch openings to separate the fine. The coarse ore falls to the mouth of a 10 by 20-in. Blake crusher running 250 rev. per minute. The crushed product joins the fine in the mill storage-bin as described in connection with Fig. 102. All the foregoing operations occur in the day shift. From the mill storage-bin the ore is fed by automatic feeder to a 20-stamp battery, having 1100-lb. stamps dropping 6 to 9 in. 100 times per minute. It is crushed through 30-mesh brass-wire screens at the rate of 3.75 tons per stamp in 24 hours. The battery- pulp is carried by launder to a belt elevator which raises it above the level of the Frue vanners. The pulp is discharged into a distributing box and thence to the twelve vanners of 6*4 tons daily capacity each of 75 tons total capacity. With a concentration of four into one, there is obtained nearly 20 tons of concentrate containing the refractory tetrahedrite needing chloridizing roasting, and 55 tons of free-milling tailing containing malachite. The tailing is settled as in the Washoe process (See Fig. 102 and 109) in a series of 18 tanks, each 9.25 ft. long, 6.5 ft. wide, and 2.75 ft. deep. The settled pulp is shoveled from the tanks to the floor adjoining. It is fed as needed into 14 wooden-sided, 6-ft. amalgamating pans, the charge consisting of 3000 Ib. tailing and 1000 Ib. concentrate that has received a chloridizing roast to be described. To the charge is added 300 Ib. mercury to amalgamate 244 THE METALLURGY the silver, 17 Ib. salt to react with the copper salts present and decompose silver sulphide, 8 to 16 Ib. iron filings to decompose chlorides, and 1 Ib. lye to correct any acidity of the ore. The. iron filings react as follows: GAgCl + 2Fe + 3Hg = 3Ag 2 Hg + Fe 2 Cl 6 3CuCl 2 + 2Fe + 3Hg = 3Hg + Fe 2 Cl n A Without the iron, mercury chloride (HgCL) would be formed and lost in the pan-tailing. Silver and copper chloride are accordingly decomposed and the metals amalgamated. The amalgamation is completed in eight hours, then the content of a pan is run into an 8-ft. settler 39 in. deep, of which there are seven. The settler is run at 20 rev. per min. and the charge, now diluted with water, is worked off in four hours, one settler treating the output of two pans. At the side of the settler near the top is a rectangular overflow notch 3 in. deep, and at distances of 8, 14, 21, and 25 in. down are 3-in. plugs. The pan, having been for an hour in motion, decanting separated pulp and water, the first plug-hole from the top is opened. After another hour the second plug is withdrawn and decantation continued two hours longer, the sand in the settler below the second plug remaining until the periodical cleaning out. The plug-holes are now closed, and the pan is ready for another charge. The exhausted tailing is conveyed to a settling pond and impounded behind a crib-dam thrown across the canyon below the mill. The mercury. From the mercury-well at the side of a settler (See Fig. 104), the mercury overflows to an amalgam safe (See Fig. 105). After straining through the canvas filter-sack inside the safe, it drains to the foot of a 4-in. belt elevator provided with cast-iron cup-buckets. This raises and delivers it to a mercury storage-tank where it is reserved until again used. This arrange- ment of settler, safe, elevator, and tank, is well shown in Fig. 102. To remove mercury chloride which causes mercury to flour and fail to coalesce, one pound of sodium is stirred into 16 Ib. mercury and added to the mercury supply every two days. The sodium reacts as follows: 2Na + HgCL = 2NaCl + Hg. The NaCl then dissolves in the water. Eight pounds of potassium cyanide also is added weekly, the cyanide reacting upon the mercury chloride thus : HgCL, + 2KCN = Hg(CN) 2 -f 2KC1. OF THE COMMON METALS. 245 The KC1 passes into solution. The mercury cyanide is soluble in the mercury and adds to its activity. Treatment of amalgam. The amalgam removed from the canvas strainers is often dirty and must be sent to a clean-up pan to remove the waste-matter before retorting. It is placed in the pan for cleaning, ground, settled, and the sand and dirty water decanted as from a settler. The amalgam thus formed is again put through a strainer, and the cleaned amalgam sent to the retorts. Retorting. From 75 tons of 25-oz. ore there comes 1600 oz. silver or 12,800 oz. coppery bullion 125 fine. This corresponds to 64,000 oz. or 4400 Ib. amalgam of 20% Ag. There are four retorts, each capable of holding 1600 oz. amalgam. The amalgam is charged to the retorts, Fig. 106, the covers luted on, and the retorting conducted in 8 to 10 hours. A jacket (See Fig. 54) that surrounds the condensing tube, is supplied with cold water under pressure. The mercury-fumes thus condensed are caught in a tub containing water. Melting the residue. This is done in an efficient way in a small reverberatory furnace, which is first heated to a bright-red heat. The furnace is charged with 4400 Ib. retort residue, upon which when melted is thrown in 20 Ib. each of borax and nitre to act as a flux. The bath is then thoroughly stirred and two samples for assay taken with a long-handled spoon, and granulated by pouring into water. Brass molds, holding 100 Ib. each, placed close together on a carriage, and painted with a mixture of lamp- black and benzine, are brought beneath the furnace spout and the bullion tapped into them, the carriage being moved along as the molds are successively filled. Roasting the concentrate. The concentrate amounts to 20 tons daily. It is trammed to the roaster-room and emptied on the floor above the Bruckner roaster. The charge is shoveled into the cylindrical roaster, which is revolved 15 minutes for the ore to dry. Then 10% salt is added, and if there is not sufficient sulphide in the ore sulphur is added to increase this to 5%. The charge is now fired and the temperature gradually raised, dense grayish- white arsenical and antimonial fumes being evolved. Next, with increasing heat, S0 3 escapes, while the presence of CuSO 4 is indicated by the blue copper flame. When this grows faint the chloridizing reactions begin, and a bright-red heat is maintained to the end. Both roasting and amalgamation are here simplified by 248 THE METALLURGY (2) Ag 2 S + CuCl 2 = 2AgCl + CuS The above reaction is slow, and it is probable that the cupric chloride acts directly on the mercury as follows : (3) 2CuCl 2 + 2Hg = Cu 2 Cl 2 + Hg 2 Cl 2 The mercurous chloride is lost. While Cu 2 Cl 2 is insoluble in water it is soluble in salt solution, and according to Laur can act directly on silver sulphide with the production of metallic silver : (4) Ag 2 S + Cu 2 Cl 2 == CuS + CuCl 2 + 2Ag 61. THE HYDRO-METALLURGY OF SILVER. A wet-process for the recovery from the ore consists in dissolving the metal by means of a solvent and precipitating from the solution in a convenient form. The silver compounds which can be obtained readily in solution are the sulphate and the chloride. In cyanide solution argentite is readily soluble, while ruby silver, freislebenite, and stephanite, are sparingly so, though readily soluble in mercurous potassic cyanide. Silver sulphate is soluble in hot water, while silver potassic chloride is dissolved by brine solution or by sodium hyposulphite (thiosulphate). From the aqueous solution of the sulphate silver is precipitated by metallic copper; from the brine solution of its chloride by copper, or when in dilute solution, by zinc iodide ; from the hyposulphite solution by sodium sulphide ; and from the cyanide solution by metallic zinc. There are four well known wet-processes for the extraction of silver : (1) The Augustin process, based upon the solubility of silver chloride in brine. (2) The Ziervogel process dependent on the solubility of silver sulphate in hot water. (3) The Patera process in which silver chloride dissolves in a solution of sodium hyposulphite. (4) The cyanide process in which the silver minerals above enumerated dissolve in dilute potassium cyanide solution. 62. THE AUGUSTIN PROCESS. This has been used for the extraction of silver from ore and from copper-bearing matte, obtained as a product of smelting. At Kosaka, Japan, ore consisting of one-half heavy spar and containing 10.5 oz. silver per ton is thus treated. The ore is crushed and roasted with salt in a furnace 6, Fig. 112, and after drawing OF THE COMMON METALS. 249 from the furnace and moistening on the cooling floor, contains 80% of silver in the form of chloride. It is leached with a hot 18% salt solution in regular leaching vats r. The leaching is continued until a polished plate of copper shows no precipitate of silver when held in the flowing filtrate. It requires 0.66 tons of brine to leach a ton of the ore. The sand is washed with hot water, and the tailing rejected. The brine solution from the vat c is received in a precipitating tank d. The tank, like a leaching-vat, has a false bottom. Upon the false bottom is spread a 2-in. bed of bean-copper, and on this rest plates of copper 6 by 8 in. by 1 in. thick. The silver, Fig. 112. FLOW-SHEET OF AUGUSTIN PROCESS. precipitated in crystalline form upon the copper, is called 'cement- silver.' At a lower level is a tank e containing scrap-iron where the copper in solution is precipitated, while the barren brine goes to the brine-sump. It is there brought up to the full strength and pumped back to be again used. The cement-silver, 150 to 750 fine, is removed from the plates, squeezed in a screw-press into 'cheeses' 12 in. diam. by 3% in. thick, dried, and refined in an English cupelling-furnace in charges of 150 Ib. with 300 Ib. of lead added to each charge. The refined silver, the result of the operation, is melted in crucibles and cast in bars of 1000 oz. each, 985 fine. Treatment of matte by the Augustin process, As indicated in the diagram (See Fig. 112) the matte after a preliminary roasting is smelted to a higher-grade matte in furnace a. It is then crushed, 248 THE METALLURGY (2) Ag 2 S + CuCl 2 = 2AgCl + CuS The above reaction is slow, and it is probable that the cnpric chloride acts directly on the mercury as follows : (3) 2CuCl 2 + 2Hg = Cu 2 Cl 2 + Hg 2 Cl 2 The mercurous chloride is lost. While Cu 2 Cl 2 is insoluble in water it is soluble in salt solution, and according to Laur can act directly on silver sulphide with the production of metallic silver : (4) Ag 2 S + Cu 2 Cl 2 = CuS + CuCl 2 + 2Ag 61. THE HYDRO-METALLURGY OF SILVER. A wet-process for the recovery from the ore consists in dissolving the metal by means of a solvent and precipitating from the solution in a convenient form. The silver compounds which can be obtained readily in solution are the sulphate and the chloride. In cyanide solution argentite is readily soluble, while ruby silver, freislebenite, and stephanite, are sparingly so, though readily soluble in mercurous potassic cyanide. Silver sulphate is soluble in hot water, while silver potassic chloride is dissolved by brine solution or by sodium hyposulphite (thiosulphate). From the aqueous solution of the sulphate silver is precipitated by metallic copper; from the brine solution of its chloride by copper, or when in dilute solution, by zinc iodide ; from the hyposulphite solution by sodium sulphide ; and from the cyanide solution by metallic zinc. There are four well known wet-processes for the extraction of silver : (1) The Augustin process, based upon the solubility of silver chloride in brine. (2) The Ziervogel process dependent on the solubility of silver sulphate in hot water. (3) The Patera process in which silver chloride dissolves in a solution of sodium hyposulphite. (4) The cyanide process in which the silver minerals above enumerated dissolve in dilute potassium cyanide solution. 62. THE AUGUSTIN PROCESS. This has been used for the extraction of silver from ore and from copper-bearing matte, obtained as a product of smelting. At Kosaka, Japan, ore consisting of one-half heavy spar and containing 10.5 oz. silver per ton is thus treated. The ore is crushed and roasted with salt in a furnace &, Fig. 112, and after drawing OF THE COMMON METALS. 249 from the furnace and moistening on the cooling floor, contains 80% of silver in the form of chloride. It is leached with a hot 18% salt solution in regular leaching vats c. The leaching is continued until a polished plate of copper shows no precipitate of silver when held in the flowing filtrate. It requires 0.66 tons of brine to leach a ton of the ore. The sand is washed with hot water, and the tailing rejected. The brine solution from the vat c is received in a precipitating tank d. The tank, like a leaching-vat, has a false bottom. Upon the false bottom is spread a 2-in. bed of bean-copper, and on this rest plates of copper 6 by 8 in. by 1 in. thick. The silver, \ Fig. 112. FLOW-SHEET OF AUGUSTIN PROCESS. precipitated in crystalline form upon the copper, is called 'cement- silver.' At a lower level is a tank e containing scrap-iron where the copper in solution is precipitated, while the barren brine goes to the brine-sump. It is there brought up to the full strength and pumped back to be again used. The cement-silver, 150 to 750 fine, is removed from the plates, squeezed in a screw-press into 'cheeses' 12 in. diam. by 3 l / 2 in. thick, dried, and refined in an English cupelling-furnace in charges of 150 Ib. with 300 Ib. of lead added to each charge. The refined silver, the result of the operation, is melted in crucibles and cast in bars of 1000 oz. each, 985 fine. Treatment of matte by the Augnstin process, As indicated in the diagram (See Fig. 112) the matte after a preliminary roasting is smelted to a higher-grade matte in furnace a. It is then crushed, 250 THE METALLURGY and roasted with salt in the furnace 6, and the roasted product leached in the tank c, as already described for ore. The residue from c cannot be rejected, as was the case with the ore, for it contains copper and iron oxides, and must be further treated b,y the Welsh method as described in the chapter on the metallurgy of copper. It is smelted in furnaces / with silicious ore and copper sulphates, producing a copper matte. The matte, in coarse pieces, is charged into another furnace from the vat e, and slowly melted by an oxidizing flame. The smelting is completed at a high temperature producing blister copper. This is cast in ingots or pigs, and refined in a furnace to produce market copper. 63. THE ZIERVOGEL PROCESS. This process, practised at Mansfeldt, Germany, and at the Boston and Colorado smelting works at Argo, Colorado, is adopted to the treatment of rich copper matte containing little or no arsenic, COPPER ORE-S & GOLD ORES ORES FINAL SOLUTION MARKET COPPER To WASTE Fig. 113. FLOW-SHEET OF ZIERVOGEL PROCESS. antimony, or bismuth, any of which would form insoluble compounds with silver. The method may be divided into three parts : the roasting for silver sulphate, the leaching, and the precipitation of the silver. OF THE COMMON METALS. 251 The process. Referring to the flow-sheet of the process (See Fig. 113) we have in furnaces a, the operation of producing the matte or regulus from gold and silver-bearing copper ores. The details of the process are described in the chapter on the metallurgy of copper, under the head of ' Reverberatory Matte Smelting.' The composition of the matte is Cu, 47.3; Pb, 8.1; Zn, 2.7; Fe, 17.7; S, 21.6%, with 400 oz. silver and 15 oz. gold per ton. Preparation of the matte. The matte is crushed and passed through rolls at 1) to reduce it to 6-mesh size, and sent to a Pearce- turret furnace c (See also Fig. 40 and 41), where it receives preliminary roasting. The roasting reduces the sulphur to 6.3%, and converts the iron and copper sulphides to the corresponding oxides and sulphates, as described in the chapter on the chemistry of oxidizing roasting. This partly roasted product then goes to a Chilian mill d (See also Fig. 100), where it is finely ground to 60-mesh. Sulphatizing roasting. The partly roasted matte is next treated in charges of 1600 Ib. by a sulphatizing roast in small single-hearth reverberatory roasters at e. In the process the iron and copper remaining in the form of sulphides are converted into sulphates which react on the silver sulphide at a slightly higher temperature as follows : Ag 2 S + 30 + CuS0 4 = Ag 2 S0 4 + CuO + S0 2 It has been found that the addition of 2% sodium sulphate (salt cake) facilitates the change. The roasting takes place in four stages as shown below. During the first stage, of one and one-half hours, the draft is checked, the side doors kept open, and the charge held at a low temperature. The charge becomes evenly heated throughout, and glows from the oxidation of Cu 2 S to Cu 2 0. During the second stage, of one and one-half hours, the heat is increased and the charge constantly rabbled. Iron sulphate is decomposed with the consequent formation of copper sulphate. The charge swells and becomes spongy by the formation of this salt. In the third stage, of a like duration, the temperature is increased for an hour until tests show that the silver is 'out,' that is, in the form of sulphate. The following reaction occurs. CuSO 4 + Ag 2 O = Ag 2 SO 4 + CuO During the fourth stage the temperature is kept constant. The charge is gathered and pressed down with a heavy, long-handled, iron paddle to break the lumps, and then vigorously stirred to 252 THE METALLURGY oxidize the remaining Cu 2 to CuO, and decompose copper sulphate. The temperature is not further increased, since it would decompose silver sulphate forming silver oxide rendering the silver again insoluble. The progress of the roast is tested by dropping small samples from time to time into hot water. Soluble sulphates dissolve in the hot water; and in the tests made early the solution becomes deep blue. Later, as the silver sulphate begins to form, it is immediately reduced to silver spangles by the cuprous oxide present. As the roasting advances during this stage, the copper sulphate decomposes, and the solution becomes less blue in the test and the silver spangles increase and afterward diminish. During the fourth stage the Cu 2 is changed to CuO and the spangles no longer show. A light blue color of the solution remains, due to the presence of a little copper sulphate, which indicates that the silver sulphate is not itself becoming decomposed. A sample thus roasted, showed by analysis 2.5% FeS0 4 and ZnS0 4 ; 0.6% CuS0 4 , and 1.73% Ag 2 SO 4 (348 oz. silver per ton), so that there was left in the matte (there being no loss of weight in roasting matte) 52 oz. per ton or 13% of the silver in insoluble form. Leaching. This is performed in tanks having filter bottoms and holding 1000 Ib. of matte. The roasted matte charged into the tanks /' is leached with hot water to dissolve the sulphate above described. The filtrate goes to a series of boxes, h, containing copper plates upon which the silver precipitates in the form of white shining crystals. The silver-free solution, containing in addition to the original copper sulphate that which it has taken from the copper plates, goes to tanks i where the copper is precipitated upon scrap-iron employed to recover the copper. The final solution is rejected. The cement-silver from the precipitating boxes is transferred to a tank and dilute sulphuric acid is added. It is boiled by forcing in a mixture of air and steam from an injector. The treatment oxidizes and dissolves the traces of copper still retained by the silver crystals, and keeps them in agitation at the boiling temperature of the acid mixture. The copper sulphate solution is run off and the residue repeatedly washed by decantation with hot water to free it entirely from copper. It is transferred to a long pan over a coal fire for drying and is then melted down in crucibles in a wind-furnace and obtained in ingots 999 to 999.5 fine. Residue from the leaching tanks. The extracted matte remaining OF THE COMMON METALS. 253 in the tanks /', still retaining 52 oz. silver per ton as above stated, freed from sulphates, and composed mainly of iron and copper oxides, is sent to a reverberatory furnace k, to form copper matte. The slag produced in the treatment goes back to the ore-smelting furnace a, while the matte tapped into sand molds, is sent to the reverberatory furnace I to be treated by the English process of making 'best-selected copper.' Here the matte, in large lumps, is piled up in the furnace near the bridge, and exposed to a flame made oxidizing by an excess of air admitted through the fire and through openings in the bridge and roof of the furnace. The effect is to 'roast' the matte as the lumps slowly melt and the drops of liquefying matte come in contact with the air. Finally the whole charge becomes melted, and the copper oxide which has been formed, acts on the unoxidized copper sulphide of the matte as follows : 2Cu 2 O + Cu,S = 6Cu-+ S0 2 The aim is to extend the roasting only so far as to obtain in the form of metallic copper one-fifteenth of the total matte. When the charge is tapped from the furnace into the sand molds, the copper is found in the form of plates or bottoms in the first of the molds beneath the lighter matte. The bottoms absorb the impurities such as arsenic, antimony, lead, and bismuth, practically all the gold (100 to 200 oz. per ton), and some of the silver. On the other hand, the matte has risen to the grade of white-metal of 75% copper, and carries 90 to 100 oz. silver but not more than 0.2 oz. gold per ton. To prepare it for the extraction of the silver, the matte is again given a sulphatizing roast, but in a different furnace from the one used for the first matte. The residue after this second treatment, principally a copper oxide containing 10 oz. silver per ton is sold to the oil refiners. At the Argo works the bottoms, formerly treated by a secret process for the extraction of the precious metals, is now more satisfactorily treated by the processes of electrolytic refining. (See 'The Electrolytic Refining of Copper'.) 64. THE HYPOSULPHITE LIXIVIATION OF SILVER ORE. (PATERA PROCESS.) Hyposulphite lixiviation can be practised upon ore containing simple or compound sulphides of silver that have undergone a preliminary chloridizing roasting. The silver sulphides, in the roasting, become converted into silver chloride. The process also applies to silver ore already containing the silver as chloride. Free- 254 THE METALLURGY milling ore, such as oxidized ore containing the silver in the native state, or as chloride, or to some extent as argentite, are preferably treated by milling and amalgamation. Native silver and silver sulphide in a favorable form can be recovered by milling and -amalgamation, whereas by hyposulphite extraction, they would remain insoluble. The process is not suited to the treatment of gold ore. The extraction of gold is slow; usually less than 50 to 70%. One of the most useful applications is to the treatment of 40 Fig. 114. Section onB-3 PLAN AND ELEVATION OF LIXIVIATION PLANT FOR SILVER ORES. argentiferous blende that has been hand-picked or concentrated from galena, and may still contain lead up to eight per cent. The hyposulphite process is based upon the fact that silver chloride readily dissolves in dilute solutions of sodium hyposulphite. The chloridizing roasting is unquestionably the most important part of the process, and the chief attention and study is to be given it. The process consists in crushing the ore, roasting it, and treat- ing the roasted ore in filter-bottom vats, first with water to remove the soluble chlorides and sulphates of the heavy metals (base-metal OF THE COMMON METALS. 255 leaching), then leaching with a dilute solution of sodium hypo- sulphite to dissolve the silver chloride. Silver sulphide is pre- cipitated from the filtrate with sodium sulphide, dried, and roasted to remove the sulphur, and the residue is sent to the smelting works, or treated in an English cupelling furnace. The treatment of the ore to the point where it is stored on the cooling floor, has been described under dry silver-milling, which see. Fig. 114 and 115 show the plan and elevation of a plant for the treatment of roasted ore by hyposulphite lixiviation. At the high level there are six leaching tanks, each 20 ft. diam. by 6 ft. Fig. 115. SECTIONAL ELEVATION OF PLANT FOR THE LIXIVIATION OF SILVER ORES. deep and each holding 60 tons. These are commanded by a double track over which the ore is brought to them. At the next lower level are the six precipitating tanks 13 ft. diam. by 10 ft. deep, and at a still lower level are four sump-tanks, serving also as storage tanks, whence the barren solution is returned by centrifugal pump to be again used. The tanks rest upon trestles, so as to be accessible for inspection and repair. The precipitate is treated in a building not shown. Base-metal leaching. The cooled roasted ore is sprinkled with water from a hose, and piled at the side of the track awaiting to be trammed to the leaching vats. Fig. 116 represents, in plan and sectional elevation, a leaching-vat having a false-bottom of wooden strips covered with cocoa matting and canvas. Into the vat is dumped the ore, which is leveled a foot below the top of the tank. If the wash-water were admitted above and allowed to leach downward, it would dissolve the silver chloride, carrying it away 256 THE METALLURGY in the wash-water. To avoid this, water is admitted below the false-bottom and made to ascend through the charge until the vat is full. The wash-water is then allowed to flow out at the bottom Fig. 116. VAT FOR HYPOSULPHITE LIXIVIATION. of the vat, and fresh water is admitted on top. By this means concentration of the silver in the solution at the lower part of the vat is avoided, and that in the upper part is soon counteracted by the dilution. The precaution does not prevent the dissolution OF THE COMMON METALS. 257 of a small amount of silver chloride in the salty water, and the first of the discharge is run to a special precipitation-vat at the end of the row. The flow is through the outlet d, Fig. 116, and the hose ?/, opened for the purpose, and thence by launder g (marked d in the plan Fig. 114) to the vat. Here the silver is precipitated by sodium sulphide. When the quantity of copper in the ore warrants, the clear solution, decanted from the silver precipitate, is passed through launders containing scrap-iron to recover the copper. In any case, the washing of the ore in the leaching vat continues two or three days, the hose being turned into the next compartment of the launder g (e of Fig. 114) and the last of the wash-water run to waste. The completion of the washing is so thorough as to entirely remove the soluble base-metal salts. Silver leaching. A cold dilute solution of sodium hyposulphite ('hypo solution') is admitted to the leaching vat above the ore. It enters from the main distributing launder c, by a short rubber-hose, the end of which can be raised to stop the flow. The leaching proceeds rapidly, the outlet d being kept open. The hypo-solution dissolves silver chloride, reacting as follows: 2Na 2 S 2 O 3 + 2AgCl = AgS 2 3 Na 2 S 2 O s + 2NaCl Not only the silver chloride, but also silver, silver oxide, silver arsenate, silver antimonate, and gold pass into solution. Copper chloride dissolves much like silver. Lead sulphate and calcium sulphate dissolve, but the solvent power of the hypo-solution is diminished by the presence of lead sulphate or sodium sulphate, and particularly of caustic alkalis and alkaline earths such as quicklime, presence 'of the latter being due to the roasting of the limestone in the ore. The lower part of the ore in the roasting vat, after washing, contains 15 to 20% moisture, causing the first of the hypo-solution to be greatly diluted. To avoid weakening the whole stock of solvent the portion first traversing the ore is rejected. As soon as the effluent, tested with sodium sulphide, shows traces of precipitation, the hose is transferred to launder g and the solution admitted to the special precipitation-vat and treated as already described. When samples from the discharge yield distinct precipitate, the liquid is run to one of the regular precipitating vats. Leaching proceeds until tests show that the silver is extracted. The flow of hypo is then stopped, and a water-wash run on. This removes the hypo-solution, and by changing the hose, the flow can be directed to one of the four sumps, or stock-tanks, at the lowest 258 THE METALLURGY level, until the solution becomes weak and does not pay to save. To avoid excessive solvent, and to provide adequate circulation the outlet d, Fig. 116, of the tank is connected through t with the Koerting injector k, which produces a partial vacuum below the filter-bottom, and forces the solution into the launder g (c in Fig. 114), whence it is admitted again to the charge. The final draining is through the hose u to the launder g. When lead or copper is contained in the ore it is advisable to use dilute hypo-solution (1.0 to 1.2%) circulating in large volume. The dilute solution acts upon the silver salts in preference to the lead salts (selective affinity) and the former can be extracted without serious dissolution of the latter. The greater part of the solution is extracted during the first 8 to 12 hours of the silver leaching, and is known as the sweet solution, because of the sweet taste of silver chloride dissolved in the hypo- sulphite. The metal precipitated from this solution will be 80% silver ; on the other hand, the precipitate from solution flowing from base ore, during the remainder of the period, is poor in silver, hence the advantage of using dilute solutions, and of leaching rapidly. Under these conditions, with ores containing much lead and copper, we may expect to obtain a solution that will yield 1.5 to 2.4 oz. silver per ton when 30-oz. ore has been successfully chloridized. Precipitation. A strong solution of sodium sulphide is added to the silver-bearing solution in the precipitation tank and well stirred by hand or mechanically. The quantity of sodium sulphide should be such as to leave a slight excess of silver solution rather than an excess of sodium sulphide. The proportion is determined by testing a sample with a drop of sodium sulphide. The silver precipitates as sulphide as indicated in the following reaction : Ag 2 S 2 3 Na 2 S 2 3 + Na 2 S = Ag 2 S + 2Na 2 S 2 8 Sodium hyposulphite is regenerated : Gold, copper, zinc, and lead are thrown down. When the proper time has arrived, the precipitate is allowed to settle, and the supernatant solution, drawn down to 1% to 2 ft. above the bottom of the precipitating tank, is run to one of the sump or stock-tanks below, making use of the floating hose, see Fig. 76. The effect of adding the sulphide is to regenerate the hyposulphite, making it again suitable for use. If necessary, it may be further strengthened by the addition of the dry salt. Sodium sulphide solution may be made at the works by dissolving caustic OF THE COMMON METALS. 259 soda in its own weight of water at 80 C. in a 3-ft. diam. iron kettle, adding gradually powdered sulphur to the solution. The addition of sulphur causes the liquid to increase to two or three times the original bulk, and in the beginning the pot should be only one quarter full. The sulphur, 60% the weight of the caustic soda, dissolves in a few minutes. The sodium sulphide solution resulting is poured into a mold in which it solidifies. For use it is dissolved in water. The sulphur needed varies from 4 to 9 Ib. per ton of ore treated, and the sodium hyposulphite 2 to 4 Ib. per ton. Treatment of the precipitate. The residue in the precipitating tank is discharged through the pipe /, Fig. 114 and 115, to a collect- ing tank in a building not shown. From the collecting tank it is pumped to a filter press, and after pressing, the damp precipitate is put in a reverberatory furnace 16 by 6 l /2 ft. hearth dimensions hav- ing a small grate to give a moderate heat. It is heated gradually in the furnace until quite dry, after which the temperature is increased to burn the sulphur. The dried precipitate is 14 to 35% silver and 15 to 27% copper. Where ore is comparatively free from copper and lead, the sil- ver may rise to 45 or 55% in the precipitate. In the following table we give an analysis of sulphides from the Marsac mill, Park City, Utah. Per cent. Ag (10,124 oz. per ton) 34.78 Au (11.7 oz. per ton) 0.04 Cu 21.60 Pb 0.50 Fe 0.75 Sb 0.18 A1 2 3 0.25 Si0 2 0.25 S 20.74 The dried and roasted sulphides are generally sold, and go to smelting works where they are smelted directly, being fed into the blast-furnace in sacks, or treated in an English cupelling-furnace as follows : Thrown upon the molten lead-bath in the hot furnace, a few shovelfuls at a time, the roasted precipitate melts. The silver enters the lead. The base metals, undergoing a scorifying action, enter the litharge. The slag continuously forming flows from the furnace, and being a litharge slag, and containing silver, is sent to the silver- 260 THE METALLURGY lead blast-furnace. When all the sulphide has been treated, the lead of the molten bath is cupelled. The silver is left behind and is tapped from the furnace into molds to form bars or ingots. Cost of treatment. At Sombrerete the ore contains 9 to 10% lead as galena, blende, chalcopyrite, silicious gangue, and silver sulphide. The ore assays 41.9 oz. silver per ton. It loses by roasting 4.8% of this. The extraction, figured on the raw ore, is 82.5%. The treatment cost is as follows: Crushing $1.36 Roasting (including the use of 6% salt) .... 2.68 Labor at the leaching plant 0.27 Chemicals 0.30 Superintendence 1.02 Heating, lighting, pumping, and repairs. . . . 0.08 Cost per ton $5.71 Clemes has given the cost in Mexico, at works treating 40 to 50 tons per day, as being 1*8 to W in the coast districts, and f*ll to W2 in the mountain districts. In small works the ratio of the cost of superintendence to the total cost is high. The small works are r\V usually directed by the owners. W 65. THE RUSSELL PROCESS. This is a modification of the Patera process, principally by the use of another solution, in addition to the hypo-solution, for the extraction of the silver. By mixing in solution two parts of the hypo-salt with one of copper sulphate we obtain a double salt Na 2 S 2 O 3 3Cu20 3 , called the extra-solution. It has a solvent power nine times as A great as that of the ordinary hypo-solution for native silver, silver sulphides, silver arsenides, and silver antimonides. In the case of an imperfectly roasted ore the use of the extra- solution insures the extraction of more silver from the compounds mentioned than could be obtained by the use of ordinary hypo- solution. The practice at the Marsac mill, Park City, Utah (See Fig. 117), is as follows: The silicious ore, containing 5.3% copper, 0.39% lead, 37.3 oz. silver, and 0.05 oz. gold per ton, is dried and dry- milled, using a 30-mesh screen, and obtaining an output of 2% tons per stamp in 24 hours. It is then mixed with 8.9% salt and roasted in a chloridizing furnace, then delivered to the point marked 'cooling floor', Fig. 117. OF THE COMMON METALS. 261 The ore, when cooled and chloridized, is charged to the ore- lixiviation vat 17 ft. diam. by 9 ft. deep, holding 72 tons: It is here treated by a water-wash for 19 hours, using 0.56 ton water per ton ore, the water percolating the charge at the rate of 4 in. per hour. Dissolving the base-metal salts and a part of the silver BORDER OF OP C*A T/CW3 - ist. Preparation of the Mater/mis xtraction 4tt,. Treatment ef ft-oefucrs Fig. 117. FLOW-SHEET OF RUSSELL PROCESS. chloride the percolating wash-water passes to the silver and base- metal precipitation tank. Here, by the addition of sodium sulphide, a precipitate is obtained containing on an average 4.2% lead, 3.9% copper, and 1238 oz. silver per ton. This is separated from the contained water in a Johnson filter-press, removed from the press, dried and shipped to smelting-works. The ore is next treated with a 1.5% hyposulphite stock-solution by which the greater part of the silver is extracted. The first of the flow goes to the base-metal precipitation tank, but, as soon as 262 THE METALLURGY the hypo-solution begins to appear, the solution is diverted to the lead precipitation tank. The silver leaching lasts 81 hours. Now the extra-solution containing 0.75% bluestone and 1.5% hypo is applied to the ore, and further silver extracted from undecomposed sulphides not attacked by the hypo-solution. This operation takes 27 hours, the filtrate being also run to the lead- precipitation tank. The cycle of leaching takes 130 to 150 hours. In the lead precipitation tank lead carbonate, which is insoluble in hypo-solution, is precipitated by the addition of 5 Ib. sodium carbonate per ton of solution, precipitation occuring in result of the following reaction : Na 2 C0 3 + PbCl 2 = 2NaCl + PbC0 3 The precipitate contains, on the average, 32% lead and 526 oz. silver per ton. After the settling, the supernatant solution is transferred to the tank marked 'silver-gold-copper-precipitation,' while the wet precipitate is filter-pressed, dried, and sold to be smelted. The solution in the silver-precipitating tank is now treated as in the Patera process with just enough sodium sulphide to bring down the precipitate of silver sulphide containing 35% silver, 20% copper, and 20% sulphur. The clear solution is pumped back to the hyposulphite stock-solution tank to be again used, while the precipitate is pumped to the filter-press, the solution removed, and the resultant moist precipitate dried and roasted at a low heat to remove sulphur. Thus the filter-press and dryer treat the base-metal precipitate, the precipitated lead carbonate, and the silver precipitate, one after another. The products are sacked, being sold to smelters, who either smelt them directly or treat them in an English cupelling furnace as described in the Patera process. The extraction by the Kussell process is as follows: Per cent. In base-metal sulphides 6.2 In lead carbonates 2.6 In silver precipitates 75.7 In sweepings of the mill. . 0.5 Total recovery 85.0 The Russell process has proved successful in exceptional cases only. At Sombrerete and at Cusihuiriachic, Mexico, it has been displaced by the Patera process, while failures have occured else- where. Compared with the Patera process, the cost of chemicals is greater (92c. against 42c.), the plant is more complicated, and OF THE COMMON METALS. 263 greater skill is needed to work it successfully. It can be applied to oxidized ores, or to those that have been subjected to an oxidizing roast, and though the yield of silver is greater where the extra solution is used, yet this is offset by the consumption of copper sulphate. In presence of much galena or blende the extra-solution extracts silver no better than the ordinary hypo-solution. In presence of much lime, silver is but slowly extracted and copper sulphate is consumed. Cost. The cost by the process, based upon an output of 100 tons daily, is as follows: Crushing and roasting $4.62 Labor in leaching 0.83 Tools, lighting, pumping, and heating 0.12 Chemicals 0.92 Kepairs and superintendence 0.18 Assaying 0.08 Treatment of products 0.10 Total $6.85 66. CYANIDATION OF SILVER ORES Silver ores carrying gold, and in which the silver occurs as chloride, argentite, or stephanite, have been sucessfully cyanided. Thus at Chloride Point, Utah, where silver occurs as the chloride, the extraction is 71% ; at the Palmarejo mine, Chihuahua, Mexico, the extraction is 54% silver and 96% gold; at El Salvador, 85 to 90% silver and 90 to 92% gold; while at Guanajuato, a silver ore, containing silver sulphide and a little gold, is treated with an extraction of 87% of the total metal of value. Of the silver minerals, native silver, in particles so large as to be visible, is insoluble in potassium cyanide in any reasonable time. Silver chloride, bromide, and argentite, are readily soluble. Ruby silver, stephanite, and frieslebenite are sparingly soluble in potassium cyanide but readily soluble in murcurous potassium cyanide solution. Important matters in cyaniding silver ore are the following: (a) A long time (10 to 25 days) in the case of sand treatment is needed for leaching. For slime treatment 96 hours would suffice for a complete cycle in which time a higher percentage of extraction would be obtained than by a 14-day treatment of the 264 THE METALLURGY corresponding sand. The silver compounds are more difficultly soluble than gold, and a larger amount must be dissolved. (b) Thorough oxygenation is necessary, not only because of the large amount of silver present, but because silver compounds need at least initial oxidation to become properly soluble in cyanide solution. Hence an advantage is secured by the double treatment of sand, as described under the 'second method' for the treatment of gold ore. Also during leaching, if the solution be allowed to sink several inches below the top of the charge, before another wash-solution is run on, air is drawn and penetrates the ore, and the solution following forces the air downward through the ore. In the treatment of the slime the pulp may receive through aeration by agitating with air. (c) Stronger solution is used than for the treatment of gold ore. Thus the first or strong solution may be 0.1%, the weak one 0.25%, while for gold ore, a 0.5% solution would be called strong and 0.05% weak. (d) The consumption of potassium cyanide is higher than in the treatment of gold ores. It varies from 2.5 to 4 Ib. per ton as compared with 0.4 to 0.8 Ib. consumed in the treatment of gold ores. (e) The precipitation of silver from cyanide solution by zinc shaving presents no difficulties, and is practically complete. Despite the fact that a relatively great amount of silver has to be precipated, as compared with gold, no more zinc is consumed. The separation and leaching of the sand should be discontinued in many cases, since leaching by cyanide solution only slowly dissolves the fine mineral particles that are enveloped in the quartz grains, and the time that economically can be allowed for extraction is inadequate. When the ore is tough and hard and the added expense of fine grinding offsets the additional extraction, the above practice would not apply. In most ores a part of the silver sulphide is so finely disseminated in the gangue that it becomes necessary to grind the ore to pass a 150-mesh screen before the sulphides are liberated sufficiently from the enclosing quartz to permit a satisfactory extraction of the silver by cyanidatioii. While mechanical agitation of slime in tanks provided with stirring-arms is common, agitation can be more economically and effectively executed by the use of Brown 's agitating-tank, Fig 118. This is provided with an air-lift consisting of a 12-in. pipe extending nearly to the bottom of a tall conical-bottom tank 13 ft. diam. by 55 ft. high. A central pipe c supplies air under pressure. The air OF THE COMMON METALS. 265 bubbling through the pulp causes a flow upward, and creates a circulation. By this method a large proportion of fine sand can be suspended in the slime pulp as would not be possible by the 7;cN OF BROWN AGITATOR Fig. 118. SECTIONAL ELEVATION OF BROWN AGITATOR. usual stirring and pumping method of agitation. Upon the dissolution of the gold, the slime is filter-pressed. The tendency in advanced practice is to classify and finely grind the sand to pass a 150-mesh screen, to add this to the slime, and 266 THE METALLURGY to agitate and filter-press the united product. The fine sand thus present with the slime makes it more easily filtered. 67. CYANIDATION OF SILVER ORES AT GUANAJUATO. The ore consists of pyrite associated with a quartz gangue with, disseminated, the silver sulphides argentite, stephanite, and pyrargyrite. It carries 14 oz. silver and 0.08 oz. gold per ton. The treatment is that described in outline in the 'second method' of cyaniding gold ores, except that here the ore is crushed in barren cyanide solution. The ore is fed from storage bins to Blake breakers and is crushed to 2-in. size. It is elevated by a bucket elevator to the sorting-room, and passed over a grizzly to separate the fine while the lump-ore falls upon a sorting belt. From 10,000 tons of ore, the monthly output, ore sorters pick out the waste, leaving 8000 tons which is hauled in cars drawn by industrial locomotives to the storage- bins of the stamp-mill, the total cost being 8c. per ton. Thus the grade of the ore is increased at the start and the quantity for treatment diminished. As the ore (260 tons daily) is dumped into the storage-bins, \% quicklime is added to neutralize the free sulphuric acid in the ore and furnish protective alkalinity to the cyanide solution. The ore is fed automatically to the stamps, where 7.2 tons of barren solution (0.025% cyanide) per ton of ore is used. Thus the silver begins to dissolve even while the ore is being crushed. From the battery the ore passes through a 50-mesh screen at the rate of 3.3 tons per stamp daily. The smaller tonnage compared with the output elsewhere is due to the fine screen used. No portion of this pulp is subsequently re-ground and about 60% will pass 200-mesh. The rich parts of the ore are the more friable, and consequently enter the finer sizes. The pulp from each five-stamp mortar flows through spitzkasten each having two compartments 22 in. square by 22 in. deep. The overflow of slime goes to the slime-tanks, while the spigot-discharge, or underflow, unites and passes to the Wilfley concentrating table. The concentrate from the table goes to storage-tanks in the concentrate room, the slime-overflow from the last five feet of the "Wilfley table goes to the slime-tanks, and the remainder of the tailing to a Frue vanner for re-treatment. The vanner concentrate goes to the concentrate-room and is mixed with the Wilfley concentrate. The concentrate amounts to 2.5% ore and contains 50% total value. It carries 40% Fe and 6% SiO 2 , and is sent to the smelter. Mean- OF THE COMMON METALS. 267 while the vanner tailing, containing all the sand, flows to classifying cones 3 l /2 ft. diam. provided with hydraulic water. The overflow goes to the slime-tanks. The sandy spigot-discharge, containing slime, passes by a launder to the sand-elevator and discharges into the small cone- separators 18 in. diam. provided with hydraulic water. Here the overflow removes further slime which goes to the slime-vats. The underflow, or spigot-discharge, of the cone-separator is discharged into the collecting vat near the side. Here it accumulates, sloping away gradually to the opposite side of the tank. A grating, such as was described under the 'second method,' and shown in Fig. 86, permits the escape of slime-bearing water. The grating is covered as high as the level of the top of the sand with a canvas roll- curtain to retain the sand. Should the sand still contain slime, a boy with a hose can wash the slime from the surface of the depositing sand through the grating to the slime-tanks. The sand- vats, seven in number, have filter bottoms and bottom-discharge valves or doors. They are 26 ft. diam. by 5 ft. deep, and hold 90 tons of classified sand. Preliminary treatment of sand. The collected sand receives the first treatment in the vat in which it settles, thus making the vat one of preliminary treatment. To fill the vat requires 15 hours, and to drain 15 hours more. The sand retains 18% moisture. Upon the charge is now run 20 tons of solution containing 0.3% KCN to which has been added 20 grams of lead acetate per ton. This is allowed to soak 12 hours. While adding the solution, a basket, containing 50 Ib. dry cyanide salt, is placed under the stream and the salt slowly dissolved and carried into the charge. The addition is sufficient to re-enforce the solution to the strength of 0.3% KCN in spite of the dilution by the moisture in the charge. Percolation and drainage now follow. Drainage causes aeration of the charge. The preliminary treatment is completed by a second saturation, percolation and drainage of the same tonnage and strength of solution Of the first. The recovery of 35 to 40% of the original value of the ore is the result. Second treatment of sand. The first treatment being complete, four discharging-doors in the bottom of the tank are opened, and the sand containing 14% moisture is shoveled into double side- dumping tram-cars running on tracks below the tanks, and above the second treatment or sand-tanks. There are 15 sand-tanks, each 26 ft. diam. by 6 ft. deep, being made more capacious to provide for the sand which does not pack in them as in the collecting tanks. 268 THE METALLURGY While a tank is being charged, 1000 Ib. quicklime is distributed uniformly through the charge. The sand is next saturated with 30 tons of strong solution to which is added 50 Ib. dry cyanide, and allowed to remain in the tank one day. The outlet valve is opened, and the solution is allowed to drain through the charge. The sand is saturated a second time with the regular strong working-solution of the same strength and tonnage as before. The solution valve is opened again and the admission of strong solution, and the percolation simultaneously proceed 6% days. The sand is next washed with weak solution (0.15% KCN) two days, and then with water 18 hours. The two bottom doors of the tank are opened and the sand sluiced out and carried by launder to the dump. The total time of the double treatment is 16 days, during which 9 Ib. lead acetate is added for a 90-ton charge. The progress of the double treatment of a charge of sand is shown in Fig. 119. Slime treatment. There are 24 masonry slime-tanks, each 8 by 9 ft., in two rows, beneath the concentration-mill floor. The stream of slime, separated from the pulp as already described, flows through one row of the tanks until filled. The stream is then diverted to the other row. The slime in the first row is permitted to settle quietly, and the clear supernatant water is decanted into sump-tanks and again used at the mill. The residual slime in the tank contains three parts water to one of dry slime, and amounts to one-half the original ore. Prom the settling tanks the thickened slime goes to 14 slime- treatment tanks, each 30 ft. diam. by 10 ft. deep, provided with stirring arms for mechanical agitation. The charge weighs 140 tons and contains 35 tons of slime, dry weight. As the stream of slime enters the tank it flows upon a basket containing potassium cyanide to strengthen the solution to 0.1% KCN. Then 200 Ib. quicklime and 18 Ib. lead acetate are added. The lime corrects acidity, gives protective alkalinity, and assists settling; the lead acetate decom- poses alkaline sulphides that would precipitate silver from the cyanide solution. The charge fills the tank to a depth of 8 ft., and the remaining 2 ft. is filled with 0.1% KCN solution from the stock- tanks. The solution dilutes the slime five to one, a proportion found necessary for the successful treatment of the charge. Agitation now proceeds uninterruptedly 24 hours, after which the charge is allowed to settle six hours. By this time three feet of the clear solution can be decanted to the adjacent 'clarifying' or clear- solution tanks. The slime-treatment tank is again filled with 0.1% KCN solution, adding the same charge of lime and lead acetate as OF THE COMMON METALS. 269 before, and agitation is resumed and continued half an hour. The pulp is allowed to settle ; the clear solution to the depth of 18 in. is decanted, and the tank filled with weak solution after the requisite addition of lime. Six washes like this follow, making eight in all, followed then by two water-w T ashes. The settled slime is now pumped to one or two settling tanks set at a high level, where it mixes with weak solution previously pumped into the tank to the depth of 6 ft. The mixture is allowed to settle as long a time as the tank can be spared for the purpose. The clear solution is decanted into the clarifying tanks, each 6 ft. diam. by 12 ft. high, with the solution decanted from the slime-treatment tanks. The pulp remain- 4 5% Transfer? *d fo First Tf 9at/nent Ksn r a 2 4 6 DAYS 10 /,? 14 Fig. 119. SAND-TREATMENT CHART AND RECORD. ing in the high-level settling-tank is then run to waste. The various decantations entering the first clarifying tank, flow across it to the second, passing through three curtain-screens of cocoa-matting cov- ered with burlap which form partitions through which the solution is strained. Treatment of the silver and gold-bearing solution. The strong solution from the leaching tanks is received in the strong gold- solution tanks. The weak solution and water-wash of these tanks, as well as the solution from the clarifying tanks, goes to the weak gold-solution tank, and after passing through the zinc boxes to the strong-solution and weak-solution sump-tanks respectively. From the sumps the barren solution is pumped to the stock-tanks to be 270 THE METALLURGY re-enforced to working strength by the addition of dry cyanide and again used. Some of the weaker solution (0.025% KCN) is pumped to the reservoir storage-tanks for use in the batteries. The total weight of solution used is 15.5 tons per ton of ore, or for the daily tonnage of 260 tons, 3(1 to 40 tons of solution passes through the zinc- i o Q boxes. The clean-up. The zinc-boxes are cleaned up three times a month as described under the first method of cyaniding, the precipitate being run to a clean-up sump-tank, and without acid treatment, pumped through a filter-press. The cake of precipitate from the press contains 30% moisture, 52% silver, and 3% gold. The 15% remaining consists of zinc, lead, copper, sulphur, and silver. The precipitate is mixed with fluxes in a large tray, using 100 parts precipitate, 30 of borax, and 18 soda. The mixture is transferred to shallow sheet-iron trays, dried by steam, then charged into No. 300 graphite crucibles and melted. The molten metal is cast into bars weighing 1000 oz. each, 860 fine in silver and 5 fine in gold, the remaining ^35 parts being zinc and lead. The total extraction is 87% of the contained gold and silver. It will be noticed that, whatever the grade of the ore, the assay value of the tailing remains uniform (See Fig. 94). The recovery not only applies to ore, but to the concentrate, which yields 98% when treated by cyanidation. In the practice at the Guanajuato plant the low extraction is due to decantation of the slime and could be increased by filter-pressing. An increase also would result if the sand were more finely ground. The itemized cost per ton of ore is as follows: Cyaniding : Labor $0.205 Supplies, including cyanide, etc. . . 1.030 Electric power 0.130 Assaying 0.060 General office expense 0.070 Plant expense, management, etc ... 0.120 1.615 Crushing and sorting ore 0.082 Milling 0.550 $2.247 PART V. IRON PART V. IRON. 68. IRON ORES. The oxides of iron occur with earthy materials and never in a pure state. Only those are called iron ores that contain sufficient iron to make the recovery of the metal profitable. Iron ore is also used as a flux in the smelting of copper and lead ores, in which case the iron enters the slag and is wasted. The three kinds of iron ore used in making pig iron are : the hematites, the magnetites, and the carbonates. Hematite (Fe 2 3 ) is the best known of the iron ores. It contains, when pure, 70% iron. When the proportion of contained water is low it is called red or brown hematite, while the hydrous varieties are called soft hematites or limonites, although the name limonite is more properly applied to bog-iron ore, containing 20% water. The rich varieties (Lake Superior ores) contain 58 to 64% iron, while the limonites run as low as 50%. By roasting the ore the water can be mostly driven off and the ore improved in grade and made better for smelting. Oolite is a variety that exists in the form of grains or nodules and contains silica or lime. When silicious, as in places in Alabama, the ore is well-nigh worthless, but when limey, as in the Minette district of Germany, the ore is self-fluxing, that is, the lime will flux the silica that the ore contains. It then becomes unnecessary to add flux. The ore of these districts runs as low as 30 to 35% iron, yet it can be smelted at a profit because of the self-fluxing property. Red hematite iron ore is the most desir- able. Most of the Lake Superior deposits are of this variety, and they are there divided into two classes : the hard or lumpy ore of the Michigan and Wisconsin ranges, and the soft ores of those ranges, but especially of the Mesabi range in Minnesota. Among the ores of the United States we should also mention the Alabama beds, and those of Colorado and Wyoming in the West. Magnetite (Fe 3 4 ) as a pure mineral contains 72.4% iron. It is so named because it is strongly attracted by the magnet, whereas other varieties of iron ore are but little affected. Magnetite occurs in large deposits in Sweden, and in various parts of the United 274 THE METALLURGY States. While some of the beds are rich, many contain no more than 40% iron and carry so much silica that the fluxing and smelting is not profitable. Much work has been done in the concentration of these ores, both in Sweden and the United States. In New Jersey extensive beds occur that have been utilized by Edison for the pro- duction of a high-grade ore on a commercial scale. He has mined the deposit, crushed and concentrated it, and made it into briquettes that contain as little as 3.3% silica and 0.04% phosphorus, and as high as 67% iron. The enterprise, however, could not continue at a profit in competition with foreign ores from Cuba and Spain. The extensive plant, so ingeniously devised and constructed by him, is not now in operation. Some of the New York beds near Lake Champlain have been considered valueless on account of the pres- ence of titanium, it having been asserted that this element produces an infusible sticky slag. This, however, has been proved to be un- founded, and it should not prevent their use as a source of iron. It is to be noted that the famous Iron Mountain, Missouri, is a deposit containing 31% iron and 6% titanium oxide, but the deposit is not now worked. In Pennsylvania the Cornwall beds are the most important, and yield a pig-iron carrying not more than 0.04% phosphorus. The ore runs 2.5% sulphur, and about half of this is removed by kiln-roasting before smelting. It contains also copper, which will be found in the pig to the extent of 0.5 to 0.75%. This does not matter in the finished product, but, if the pig-iron is made into steel, the copper causes 'hot-shortness' or brittleness when hot,* thus causing imperfections when rolled into shapes. The average of ore mined is 40 to 42% iron and 20% silica. Carbonate ore, siderite (FeCO 3 ), as a pure mineral contains 48.3% iron. The varieties are spathic, black band, clay-band, or clay iron- stone. It is often roasted to expel the moisture and carbon dioxide before going to the blast-furnaces. In England it forms the well known clay iron-stone of the Cleveland district, but in the United States, though widely distributed, it is too low in grade to be used in competition with the abundant rich ores. 69. SMELTING FOR PIG-IRON. Outline of the process. The operation is conducted in a furnace, often 100 ft. high, filled with a mixture of coke, iron ore, and lime- stone. Superheated air is blown in at the bottom. The coke is burned to maintain a high temperature in the furnace and to reduce the iron in the ore to the metallic form as pig-iron. The pig-iron collects at the hearth or bottom of the furnace, and is removed from OF THE COMMON METALS. 275 time to time. The gangue, or silicious part of the ore, is fluxed with limestone, and produces a worthless slag, or cinder, which is also removed (tapped) as it accumulates in the furnace. 70. BLAST-FURNACE PLANT. Fig. 120 is a view of an iron blast-furnace plant for the manu- facture of pig-iron from iron ores. In the foreground is a cylindrical Fig. 120. IRON BLAST-FURNACE. 276 THE METALLURGY furnace-stack 100 ft. high, immediately in front of which is the forked 'down-comer' (See 39, Fig. 122), a large pipe that conveys the smoke from the stack near the top downward to the flue-system Fig. 121. IRON BLAST-FURNACE WITH AUTOMATIC CHARGING. that carries it away. In front of the down-comer we see the inclined hoist for the 'stock' or the materials that are put into the furnace. At the middle of the illustration are the four cylindrical 'stoves', as high as the furnace, used for pre-heating the air blown into the furn- OF THE COMMON METALS. 277 ace, while the highest stack behind them draws away the gas from the stoves. In front of the four stoves is the blast-main, a pipe 5 ft. diam., by which the air is conducted to the furnace. At the left of the stoves is the building (not shown), that contains the vertical blowing engines by which air, under 15-lb. pressure, is delivered through the stoves to the blast-furnaces. Fig. 125 represents a verti- cal blowing engine. The general arrangement about the furnace is understood from the sectional elevation Fig. 121. The blast-furnace 100 ft. high is at the right. It is served by an inclined hoist, one skip of which is in position in the pit ready for loading, the other in discharging posi- tion at the top. Within the stock-house, at the left, are three stan- dard-gauge tracks on three levels. The upper one at the left is for the hopper-cars that deliver iron ore and fluxes to sloping-bottom bins beneath, shown in section. The second one leads to similar bins (not shown in section), and the third to the floor on the ground-level. On this level is a charge- car, electrically driven, with a weighing attachment that can be brought to any bin to receive a weighed amount of stock. The load is then transferred to and discharged into the skip. In case of accident to the charge-car, or any trouble at bins, the furnace can be supplied by the use of hand-barrows or buggies, taking the stock from the piles that have been made beneath the third track. Hoisting is done by a hoisting engine set well out of the way at the top of the stock-house. Just beyond and at the right of the furnace is the cast-house where the molten iron is molded into pigs when a cast is made. 71. IRON BLAST-FURNACE. Fig. 122 is a furnace, shown in part section, and part elevation. It is circular in cross-section. Beginning at the bottom we have a heavy foundation of concrete and fire-brick upon which rests the hearth 15 and columns 4 which support the upper brick work that constitutes the shaft of the furnace. The hearth or crucible (14% ft. diam. by O 1 /^ ft. deep), that contains the molten iron and slag, extends from the foundation to a height slightly above the tuyeres 22. The bottom and walls are of fire-brick, and near the bottom is the iron-tap 27 through which the molten pig-iron is withdrawn when a quantity has accumulated. At 24 is the cinder-notch or tap by which the slag or cinder is drawn off. The crucible is surrounded by a hearth-jacket of steel plates, cooled on the outside by sprays of water that play against 278 THE METALLURGY it, cooling and protecting it and the brick-work lining from the corrosive action of the molten slag inside. Air^ under a pressure Fig. 122. IRON BLAST-FURNACE, DETAILED SECTION. of 5 to 14 lb. to the square inch enters through the tuyeres 21, which have projecting nozzles 22. Care is taken to withdraw the slag before it reaches the level of the tuyeres, for it would enter the OF THE COMMON METALS. 279 openings and close them. Of these tuyeres there are six. The air is supplied through the tuyere-stocks 33 from the bustle-pipe 13, which encircles the furnace, and connects with the blast-main supplying air at the temperature of a red heat from the stoves. The bustle-pipe, tuyere, and tuyere-stock are shown in the section Fig. 121. The bosh, or that part of the furnace that widens from 14 ft. 6 in. at the hearth to 22 ft. in a vertical distance of 13 ft. is also shown. It is in the region of the bosh that the formation of the slag occurs, and the brick-work of the bosh is subject to a slagging and scouring action that tends to attack and destroy it. To prevent this, hollow water-cooled bosh-plates 14 are laid in the brick-work of the bosh, making rings around the furnace at nearly every two feet vertically. The slag cuts into the brick-work nearly as far as the inner ends of these plates, but the circulation of water within them protects the adjacent brick-work from deeper corrosive action. The shaft, or main brick-work structure of the furnace, is carried by the cast-iron mantel 5 resting upon the columns 4. It extends from the top of the bosh to the throat at 9. The upper part of the furnace is closed by a bell 47 (Fig. 121 shows a double bell), and the gas escapes at the side through the down-comer 39. The in-wall 69 is of fire-brick, while the main portion is of common brick and is sheathed with a shell 46 of steel plates. When in operation the furnace is kept full to a level just below the outlet to the down-comer. This level is known as the stock-line, and the furnace at this point is 15 ft. diam. As the stock smelts and sinks, charges are introduced and the stock-line is maintained at this level. Ordinarily the top of the furnace is kept closed by the conical bell 47 which is suspended from the ends of the counterweighted beams 55. The bell closes the bottom of a circular hopper 48, into which the charge in this particular furnace is supplied by buggies brought up by elevator to the upper or charge-floor of the furnace or 'tunnel-head,' as it is called. To drop a charge into the furnace, the outer end of the lever is raised by the piston-rod and piston of the air-cylinder 60. The bell thus lowered permits the charge to slide into the furnace, after which it is immediately raised to close the opening and stop the outward rush of smoke and gas that mainly escape through the hood 61. The gas, containing dust from the charge, passes off by the down-comer 39 to the dust-catcher 40 (where a part of the dust settles), and by the goose-neck pipe 41 to an underground flue that leads to the stoves and boilers where the gas is burned. Rising from the down-comer is the bleeder 37 ? 280 THE METALLURGY that is used when it is desired to relieve the top pressure of the gas rising from the charge. It is occasionally used. At many furnaces the stock is raised in hand-barrows or charge-buggies to the furnace-top or tunnel-head 51 by means of a platform hoist. In Fig. 120 is seen at the extreme right, the frame of the hoist. In Fig. 121 is shown the more recent method of charging with the inclined hoist. A double bell is used to prevent the escape of the gas. The charge is dropped from the hoist into the upper hopper, where it is retained until the lower hopper is empty. The smaller upper bell is then lowered and the charge slides from the upper into the lower hopper, while the upper bell is closed. The hopper is then ready to take another charge. The charge in the lower hopper, when needed is dropped into the furnace by lowering the lower bell. It slides outwardly to the walls, forming a ring or ridge, the stock in the middle being a little lower than at the sides. Some coarse lumps roll toward the middle, so that part of the charge is more open than at the walls. Just beyond and below the lower bell is noticed the oval outlet to the down-comer. The stock-line must be kept below this. The skips of the hoist run in balance and are charged as follows: The charge-car on the ground level is run to the chute of an iron-ore bin to receive the required weight of ore. It is moved to the limestone-bin beyond to get the needed quantity of limestone, and then to the skip standing below in the charge-pit where it is discharged. The skip is next hoisted and dumped, while the empty one is in position to take the load of coke. After elevating the fuel, a charge of ore and flux next goes. These charges alternate in the furnace and form layer upon layer. The dimensions of a blast-furnace are limited. The considerations are as follows: The hearth should be not more than 15 ft. diam. lest the blast fail to properly penetrate to the center and maintain intense combustion there. The slope or angle of the bosh-wall must be such as to give proper support to the charge, which rests upon it, and yet allow the solid coke to slip down. The height is limited to the height of the smelting zone. These conditions limit the diameter of the bosh to 22 ft. From the top of the bosh the stack-wall must decrease in diameter to the throat to give room for the descending charge to swell by reactions that occur in the down- ward progress. This leaves, at the throat, a diameter suitable for the proper distribution of charge. Furnaces have been built higher than 100 ft., but such height has been found to be excessive, OF THE COMMON METALS. 281 especially for fine ores; and the best practice calls for 90 ft. or less. 72. THE STOVES. The efficient operation of an iron blast-furnace requires that the air entering at the tuyeres be heated, generally to the temperature of a red heat (500 to 750C). To do this the furnace is equipped with three or four (as in, Fig. 120) regenerative fire- Fig. 123. COWPER HOT-BLAST STOVE (ELEVATION). brick stoves 80 ft. high and 14 ft. diam. The Cowper stove (of Fig. 123 and 124), for example, consists of a tight shell, like a boiler shell of steel plates, lined with fire-brick, and containing a checker- work of bricks of special shape laid in open order so as to have numerous openings or passages from the top to the bottom of the stove. The gas from the furnace, containing 24% CO, which in 282 THE METALLURGY burning supplies the heat, flowing along the underground flue from the goose-neck before mentioned, enters the stove at nr> ^RJnpd i thn mrnnnp nf "til IT'S i (2) 3Cu 2 + 2PeCl 2 = 2Cu 2 Cl 2 + Pe 2 8 + 2Cu But the copper with cupric chloride reacts as follows : (3) $Cu + CuCl 2 = Cu 2 01 2 The cuprous chloride is soluble in the salt-liquor. The copper-containing nitrate from the filter-vats passes to precipitation tanks or to launders containing scrap or pig-iron, and the copper precipitates as follows: JL Cu Ci* ^" Cts^Cl 2. Ferrous chloride is regenerated, and the final liquor, after passing through the boxes, is used again. The process has an advantage over the precipitation of copper sulphate by iron in that less iron is reo.uired. In reaction (4), ee- equivalents of iron {j) precipitates mt of copper (Ss&J. The objection urged against the process is that the reaction of air upon the ferrous chloride solution is to decompose it, forming an oxychloride thus: (5) 6FeCl 2 + 3O =SFe 2 Cl e + Fe 2 O 3 The ferric oxide formed in reaction (1) and (2) also tends to clog the filter. To avoid the difficulties the process was altered to the following one. 114. THE NEW HUNT AND DOUGLAS PROCESS. The ore is crushed and roasted as described for the old process, then treated with a dilute solution of sulphuric acid to dissolve the copper oxide, and give a filtrate of copper sulphate containing ferrous and ferric sulphates. The solution is delivered into a precipitating-tank, and a solution of ferrous chloride is added which transforms a part of the copper sulphate into cupric chloride (CuCl 2 ). Sulphur dioxide, obtained by the roasting of ore, is now forced into the liquid. It precipitates the copper as an insoluble cuprous chloride, as follows : (1) CuCl 2 + CuS0 4 + SO 2 + 2H,O = Cu 2 01 2 + 2H 2 SO 4 The cuprous chloride is filtered off and treated with milk-of-lime, or with iron according to the reaction (4) of the old process. The OF THE COMMON METALS. 355 consequent FeCl 2 is used in the treatment of the next charge, while the copper precipitate is shipped away. The sulphuric acid solution of reaction (1), filtered from the cuprous chloride, is freed from the excess of S0 2 by blowing into it hot air from an injector, and the recuperated solution is used again to dissolve ore. The method has the advantage that no ferric hydrate is formed to clog the filter, and that but little iron is needed to precipitate the copper, and that precipitated copper is pure. It has been used for the extraction of copper from the matte rather than from the ore. 115. EXTRACTION OF COPPER FROM OXIDIZED ORES (NEILL PROCESS). This process depends upon the use of a sulphur dioxide solution for dissolving the copper. The copper is subsequently precipitated by heating the filtered solution and expelling the S0 2 gas. It is used preferably upon oxidized ore, such as native carbonate and oxide, which is soluble in an aqueous solution of sulphur dioxide, but not in water. The process can be used also in the treatment of roasted copper-bearing ore. Lime and magnesia consume sulphur dioxide and are objectionable. For oxidized ore, or carbonate, the crushing is done with rolls, which reduce the size to 20-mesh. The ore is then charged into a leaching barrel, like that used in chlorination, and water is added. A stream of sulphur dioxide is forced into the barrel by means of an air-compressor until the pulp is saturated with the gas. The saturation is maintained some hours, until the copper compounds have dissolved. The sulphur dioxide is produced by the roasting of iron sulphide in a pyrite-roaster. From the barrel the ore passes to filter-presses where the solution is removed, the residual tailing being rejected. The solution passes to the precipitation tanks where it is heated by steam and boiled until the SO 2 gas is expelled. The liquid, now freed from S0 2 , no longer retains the copper in solution. The copper comes down as a cupro-cupric sulphite (CuSO 3 ,Cu 2 S0 3 + H 2 0), a heavy crystalline compound of a dark- red color, containing 49.1% copper. The supernatant solution from the precipitating tank runs to waste through launders containing, as a precaution, scrap-iron to insure removing the last of the copper. The precipitated sulphite readily settles from the solution. It is washed by decantation, dried, and reduced and melted in a reverberatory furnace giving metallic copper. 356 THE METALLURGY The process has the advantage that a unit of copper, converted into cuprous sulphite, needs but half the sulphur that would be required to convert it to cupric sulphate. Cuprous sulphite is here precipitated from the solution without the use of scrap-iron. This is an advantage in remote districts where the cost of transportation is high. Sulphur dioxide has little action upon other metals, and thus a pure copper is furnished by this process.. 116. THE HENDERSON PROCESS. The well roasted residue, or cinder, resulting from the pyrite used in making sulphuric acid, contains 2 to 4% copper, with silver and gold. All these metals can be extracted by a chloridizing roast followed by leaching with weak liquor from a previous operation, and containing water and dilute hydrochloric acid. The copper in the clear filtrate is precipitated upon scrap-iron. Operation of plant. Fig. 149 shows the plan and a transverse sectional elevation of a 200-ton plant of the Pennsylvania Salt Manufacturing Co., Natrona, Pennsylvania. The cinder (red-roasted or burned pyrite) that is brought from the various sulphuric acid plants throughout the country is ground dry to 20-mesh in a pan-mill, Fig. 145, and mixed during the grind- ing with 12% of the weight of salt. The mixture is raised by a belt-elevator to storage bins (not shown) commanding the charge floor k. It is weighed in 5-ton charges and put into the charge-tubes of the muffle roasters shown in the longitudinal sectional elevation, Fig. 150. There are four charge-tubes, each 20 in. diam. for each furnace. The gases from the fire pass along the 14-in. space above the 8 by 35-ft. muffle- hearth containing the ore, thence by a flue downward to the space below the muffle, and finally by a main underground-flue to the stack. The gases from the roasting ore pass by an 18-in. pipe to condensing towers a, filled with lump coke that is wet by a spray of water above. The water, coming in contact with the ascending gas, absorbs the chlorine and hydrochloric acid. Raw pyrite is charged with the ground cinder to make the sulphur content I 1 /-? times that of copper. The charge is heated to a visible red heat (525 C.) and well stirred during 8 hours. When finished, it is drawn out upon the floor, allowed to cool, shoveled into charge-cars, raised by platform-elevator to the charge-floor level, and put into the leaching tanks d, each of which is 12 by 14 ft. in size. The ore is first lixiviated with a weak liquor from a previous OF THE COMMON METALS. 357 operation to remove most of the copper. The solution becomes a strong solution. The ore is then treated with water, to remove the remaining copper and the solution becomes the weak solution of the succeeding operation. Finally, the weak solution of hydrochloric acid from the towers a is applied, dissolving the cupric oxide and cuprous chloride, hitherto insoluble. The residue, called 'purple ii HI riin ini nininiriinin i n i a JUUUUi LJ J/yacs for Mi/fs, Sfy/ns, Bo/'/erj, C'naer and -Saff JL 149. HENDERSON-PROCESS PLANT. ore' is shoveled from the vats to the floor c and thence discharged into the railroad cars below. The weak solution is sent to the lixiviation tanks. The strong solution when the specific gravity reaches 18 B. is drawn to tanks 12 by 12 by 6 ft. where the copper is precipitated upon scrap-iron. The tanks have false bottoms of slats 2 ft. above the bottom. Live steam, directed into the solution, agitates it. The copper precipitating upon the iron works down between the slats to the bottom of the tanks and is removed to tanks g, 10 by 10 by 5 ft. The solution from this tank is drawn into launders containing scrap-iron as a guard, and to retain any remaining particles of 358 THE METALLURGY precipitate. The precipitate is 90% copper, 35 oz. silver, and 0.15 oz. gold per ton. It is sold to the blue-vitriol makers who pay 95% of the silver and the full value of the copper and gold. The cost of treatment by the process, with common labor at $1.50 per day, is $1.87 per ton of cinder treated. PART VII. LEAD PART VII. LEAD. 117. THE LEAD ORES. The lead ores are those in which lead is the principal constituent. The term is applied also to mineral aggregates consisting of more than 10% lead. The lead ores may be divided into two classes: the sulphide and the oxidized. The terms are used only according to the constituent that is in excess, in many lead ores both sulphides and oxides are found. Ore containing no lead is called dry, and when carrying lead, leady. The latter term is the opposite of dry, but we do not term a leady ore a wet one. Galena. Pure galena contains 86.6% lead and 13.4% sulphur. In nature it occurs with gangue or vein-matter. When there is much of the latter it can readily be concentrated. The following table gives an idea of the lead-content of ore, before and after dressing : GALENA ORES. Raw ore. Concentrate. Pb, Pb, Ag, oz. Locality. % % per ton. Minnie Moore, Wood River, Idaho 62.0 80.0 Rockville, Wisconsin ... 0.3 St. Joseph, Missouri 7.0 70.0 Kellogg, Idaho 11.0 60.0 30.0 Col. Sellers, Leadville, Colorado 10.0 55.0 19.8 Galena from the Mississippi Valley contains little silver, but from the Rocky Mountain region it is not only argentiferous, but may contain gold. The precious metals as well as the lead determine the value. Metallic sulphides, such as pyrite and blende, are often associated with galena, and with the gangue may carry so much of the gold and silver that concentrating leads to a serious loss of the metals and is omitted. If by hand-picking ore can be brought to contain 30 to 40% lead, it is a desirable ore for the smelter. When of this tenor in lead, and free from other sulphides, it carries but 5% sulphur and needs no preliminary roasting, and is smelted directly. Oxidized lead ores. Little lead oxide is found in nature. The ores classed here under oxidized ores are the result of the alteration 362 THE METALLURGY of galena. They include the carbonate (cerussite) and the sulphate (anglesite) of lead. The minerals are mixed with metallic oxides and vein matter or gangue in nature, and when sandy or earthy, the ore is called sand or soft carbonate, and when hard and stony, hard carbonate. In many deposits we find ore, that originally was galena, profoundly altered to cerussite or anglesite. The subjoined table gives the composition of some of the so-called carbonates. CARBONATE ORES. Locality. Pb, % 72 SiO L >, % Fe, % CaO, % s, % Ag. oz. per ton. 38 25 Leadville Colorado . 21.0 22 5 18.2 2.4 0.9 65.0 Red Mountain Colorado 18 4 41.6 11.4 1.7 1.8 128.0 33.2 3.0 24.1 1.1 2.0 27.5 Bingham Utah . 51.5 12.5 2.6 3.2 6.0 21.1 Horn Silver mine, Frisco, Utah.. . 50.0 15.2 3.4 0.5 8.3 78.3 Of the ores of the table, that from Eureka, Nevada, contains 4.2% of arsenic, which forms an arsenical speiss when smelted. The Horn Silver ore, apparently oxidized, has the lead in the form of anglesite (PbSOJ, and matte is formed from it in smelting. In oxidized ores the silver is apt to occur as a chloride, the gold probably is native. There are many lead minerals but those not mentioned occur in small quantity and are not considered among the commercial lead ores. 118. RECEIVING, SAMPLING, AND BEDDING ORES. Where ore is treated in a small way for the recovery of the lead, as in Missouri, no particular provision is made for storage. In various smelting works in the Rocky Mountain region, where lead ores are treated with others by methods of silver-lead smelting, and where ores are bought outright for treatment, the handling becomes complicated. Such plants are called custom works. A plant treating ore from a single mine is called a mine's works, and here less attention is given the sampling and storing of the ore. In customs works, therefore, ores of many kinds are received, some containing lead, some having little lead but carrying silver and gold. The ore is received in lots of a few tons up to those of several carloads. Each lot is separately weighed, sampled as fully described in the chapter on sampling, assayed, and purchased. The company then is free to treat the ore as it pleases. If different kinds of ore were smelted separately the process would involve endless OF THE COMMON METALS. 363 change and labor, and so it has become the custom to 'bed' the ore in large bins holding several hundred tons. When so bedded the mixture is treated as a single ore. The different kinds of ore are unloaded separately into the bin and each kind is spread out in an even layer before the succeeding one is added. This is indicated in Fig. 151, and it is seen that we thus have a series of Ia3 r ers in a bin. When the ore is to be used, shoveling is done at the floor and all parts above fall down and mix, since a steep face of ore is constantly maintained. Thus a uniform mixture of the different ores is obtained for smelting, and so long as we are using t : Fig". 151. ORE-BED. ore from this bin, the quality remains constant. The supply remains practically unchanged in quality several days, and the content of the bed is treated in the books of the company as a single ore. In the laboratory the aggregate analysis of the bed is obtained as follows: A list of the ores and the dry weights is prepared. The chemist weighs out, on his balance, from the reserved samples of each of the ores, an amount proportionate to the weight of ore in the bin. The total portion, amounting to one or two ounces, is thoroughly mixed, and from it portions are taken for the determination of Si0 2 , Fe, CaO, and S. The mixture may also be assayed for silver, gold, and lead, as a check on the calculated 364 THE METALLURGY content obtained from the weights and calculated assay of the individual ores. The determinations thus made are used in com- puting the charge. Besides ore bedded in this way, lots that would fill a large bin remain unmixed, and the ore is treated as a separate item of the charge. Small lots, of which a moderate amount is to be used upon the charge, also may be kept separate, and such are called 'side ores.' Crushing and bedding ores for roasting. Sulphide ore that is to be roasted also may be bedded. When thus made uniform, it becomes known and is handled better at the roaster, and it makes a uniform product for the blast-furnace. Such beds are made to contain the proportions of pyrite, silica, and galena that work best in the roaster. Crushing the sulphide is often performed in two stages. The coarse crushing of lumpy ore is done with rock-breakers, either of the jaw or the gyratory type, the ore being reduced to %-in. size. It is then passed through rolls, 36 in. diam. by 14 in. face, where it is crushed to pass a 3 to 10-mesh screen according to the nature of the ore. A pyritiferous ore need be crushed no finer than 3-mesh. while ore carrying galena and blende roasts better when crushed to 10-mesh. Sulphide ore is preferably bedded before roasting, because the men then soon learn how best to roast it, whereas with constant changes, they fail in this. Ores of different composition roast to advantage when judiciously mixed with a silicious pyrite ore. A pyrite ore, which readily starts to roast, assists the slow galena or blende. By combining different kinds we obtain a mixture that agglomerates only when the roasting is completed and the roasted material is ready to be withdrawn from the furnace. A bed formed of 10 to 15% SiO 2 , 20 to 28% Fe, and 20 to 28% Pb, roasts well. Mixtures containing less lead and more pyrite than this roast readily, but ore that is pulverulent, when roasted, tends to make more flue-dust in the blast-furnace, while with the proportion of lead above specified it tends to sinter and make a desirable lumpy product for the blast-furnace. 119. THE SMELTING OF LEAD ORES. When lead-bearing ores are to be smelted only for the lead content, as is done in parts of the Mississippi Valley, a simple plant with a reverberatory furnace, or the American ore-hearth, is sufficient. In the Rocky Mountain region the lead ore is not smelted OF THE COMMON METALS. 365 to recover only the lead. The lead of the ore is employed as a collector of the gold and silver of other ores that are smelted at the same time. In the first case, a large part of the lead is recovered cheaply and simply, but a part is lost in the resultant slag. In silver-lead smelting it is essential that the slag be comparatively free from lead and the consequent silver. 120. REVERBERATORY LEAD SMELTING. The treatment of lead ore in reverberatory furnaces has not made much headway in the United States. There are two reasons for this : In the silver-lead districts, the ore has not been of sufficient grade in lead to warrant the treatment, and lead ore has been in great demand as a collector to mix with other ores. Secondly, in the Mississippi Valley, where silver-free high-grade lead ores occur, the question of skilled labor for reverberatory- furnace work has had an influence. The reverberatory lead furnace. Fig. 152 and 153 represent one of the large recent furnaces, 16 by 9-ft. hearth dimensions, HORIZONTAL SECTION ON LINE G, H Fig. 152. LEAD-SMELTING REVERBERATORY FURNACE (PLAN). having a fire-box 8 ft. by 20 in. or of 14 sq. ft. grate-area. The floor or bottom of the hearth slopes from the fire-bridge to the corner near the external well or basin / at the cool end of the furnace. The flame passes to the stack (not shown) by a flue at a. The charge is dropped into the furnace, as needed, through a 12-in. hole in the middle of the roof. There are four working- 366 THE METALLURGY doors on each side, so that the interior is easily reached to spread, rake, or withdraw the charge. The lead, as it forms, drains to the basin /, from which it is dipped from time to time as it accumulates, and molded into bars or ingots. Operation of the furnace. The operation is divided into two stages: the first that of roasting, or oxidation; the second, reduction. Oxidation. Four tons of ore, crushed to 5-mesh size, is dropped into the furnace from the hopper and spread over the hearth in a layer 3 in. deep. An oxidizing fire is maintained in the fire-box to raise the temperature of the charge to visible red (500 to 600C.)- The roasting is kept up three or four hours, and continued only until Firebrick LONGITUDINAL SECTION ON LINE A, B. Fig. 153. LEAD-SMELTING REVERBERATORY FURNACE (ELEVATION). an incomplete but definite degree of oxidation has resulted. The reaction is as follows : (1) 2PbS + 70 = PbO + PbS0 4 + SO, The galena is converted in part into oxide, in part into sulphate. A part remains as sulphide. The temperature is kept at the required degree and the charge is frequently raked to expose new surface to the action of the air, and to prevent the agglomeration of the charge. Reduction. The fire-box is filled with a thick. bed of coal to give a neutral flame, and the temperature is raised to the point at which^the charge begins to soften, but not to melt. The oxide and of lead react upon the unchanged galena thus : (2) PbS + 2PbO === 3Pb -f S0 2 (3) PbS + PbS0 4 = 2Pb -\3$6. and they are reckoned as having equal values for fluxing. OF THE COMMON METALS. 383 Manganese is found in some of the Leadville iron ores to the extent of 10 to 15%, and since by introducing another element, it adds to the complexity of the slag, it also adds to the fusibility. The alkaline earths. Lime, magnesia, and baryta act in inverse ratio to the atomic weights in fluxing silica. Hence to obtain the equivalent in lime, the amount of lim^ needed, c^prconcd in- per ^ is multiplied by 1.4 for magnooia, and by 0.4 for baryta. A slag, high in lime and consequently low in iron like the last three in the table, is of low specific gravity. Its use thus results in a better separation of slag from the heavier matte. Lime being a stronger base by one half than iron, and generally a cheaper flux, the tendency is to choose the limey slags. It is noticed that the higher the silica content of the slags of the table the higher is the lime, and that high silica calls for high lime. Dolomite, having a high content in magnesia, generally is avoided in silver-lead smelting, for it tends to make slag pasty and streaky, and the unfavorable effect is aggravated when zinc also is present. Two analyses of limestone and of dolomite are given below to show conditions typical of actual practice. Canyon City limestone: CaO, 49.8%; MgO, 3.0%; SiO 2 , 3.1%; Fe, 0.8%. Iron county, Missouri, dolomite: CaO, 26.6% ; MgO, 17.6% ; SiOo, 5.1%; Fe, 3.3%. Flourspar. This has no unfavorable, but rather a favorable effect upon the quality of the slag. The fluorine, however, uses CaO, and hence the slag must analyse higher in CaO than the type requires, or it will not be clean. Alumina. It is uncertain whether alumina acts as an acid or a base. It is sufficient for the purpose of silver-lead smelting to regard it as a neutral constituent that dissolves in slag and acts in neither way. Zinc. Either blende or zinc oxide causes difficulties in the blast-furnace, the blende being the more objectionable. Blende is in part decomposed in the presence of iron to zinc oxide, but the zinc in any form tends to make a stiff, pasty, difficultly fusible slag. It may be regarded, like alumina, as being dissolved in the slag. It goes into both the slag and the matte, and diminishing the specific gravity of the latter it causes a less perfect separation of the two. Where much zinc is in the charge, it is customary to modify the type-slag by calculating the zinc oxide as replacing one half the percentage of lime. Take for example the half-slag J of the table. 384 THE METALLURGY Without With Recalculated Zinc Zinc Zinc Per cent. Per cent. Per cent. SiO, 31 31 29.5 FeO 38 38 36.0 CaO . 21 17 16.0 ZnO 8 7.5 90 94 90.0 In the first column we write the slag as the type requires. In the second column we add the 8% Zn and reduce the lime by 4% by which the total becomes 94. Since the constituents should amount to but 90%, all are reduced proportionally in the third column so as to give 90% as the sum. Copper. Copper present in the charge enters the matte when, as generally is the case, sulphur is present with which it can com- bine. In smelting carbonate or oxidized ores, which furnish no sulphur, the copper becomes reduced, and enters the base-bullion, giving a lead so drossy sometimes as to clog the lead-well, and accumulate and solidify in the crucible. The remedy is to supply sulphide to form matte into which the copper can enter. Antimony. Either as an oxide or a sulphide, antimony is reduced like lead. It alloys with the base-bullion, making it hard, and is removed and recovered later in refining the base-bullion. Arsenic. This frequently is encountered in silver-lead smelting. When present in small quantity it is volatilized, but in large quantity it forms a speiss. Where it is intended to produce a speiss, iron is provided with which the arsenic unites. In the fire-assay of arsenic-bearing lead ores, a bead of speiss is found attached to the lead button. From the percentage of this we can compute the weight of the speiss that will be formed: and we may assume that 70% of it is Fe. Where a direct determination of arsenic is made we can compute the weight, and multiply this by 2.3 to express the quantity of Fe to be provided on the charge for the purpose. 129. FUEL IN SILVER-LEAD SMELTING. The fuels used in silver-lead smelting are coke, charcoal, or a mixture of the two. Wood and hard coal have been used experi- mentally, the former in certain cases of scarcity of fuel. Coke. Coke is the kind of fuel commonly used. The ash varies from 10 to 22%, and the fixed carbon from 89 to 77%. In coke of OF THE COMMON" METALS. 385 high ash, not only is the ash to be smelted, but the carbon is correspondingly low, so that the coke is less efficient. A great difficulty with high-ash coke is that it is often friable, making accretions or scaffolds. Analyses of two typical samples of coke give the following results : Connellsville coke contains fixed carbon 87.5%, ash 11.3%, and sulphur 0.7%. El Moro coke, fixed carbon 77.0%, ash 22.0%, and sulphur (when the coke is made from unwashed coal) 0.9 per cent. In computing a charge the coke-ash is taken into account, analyses being as follows : Ash of Connellsville coke : Si0 2 , 44.6% ; Fe, 15.9%; CaO, 7.0%; MgO, 1.9%; ash of El Moro coke: Si0 2 , 84.5%, and Fe, 5.0%. Charcoal. This fuel is used in districts far from railroads, where the cost of coke is high. It is a good fuel for oxidized ores, but is friable and makes undesirable fine which may form accretions or scaffolds in the furnace. It renders a charge more open than coke, and contains less than 2% ash. Coke weighs 25 Ib. and charcoal 10 Ib. per cu. ft. when loose, the weight of a bushel of charcoal being 14 to 16 Ib. Even where charcoal is cheap it is desirable in operating the furnace to use part coke which, fed to the walls, burns more slowly than charcoal and makes the tuyere- zone hotter and gives a more liquid slag. Quantity of fuel. This varies according to the nature of the charge, and generally is from 12 to 15%. Charges that contain sulphur and make matte need less fuel than oxidized ores. Only sufficient is used to give adequate reduction and a hot slag; and the metallurgist is guided by these requirements in adding the fuel. 130. CALCULATION OF A CHARGE. When sulphur-bearing, oxidized or silicious ore is used, we have to consider not only the sulphur, silica, and other constituents of the ore, but also the products of the furnace that remove, the constituents. Ore (galena for example) containing less than 10 to 12% sulphur generally is smelted without roasting. It is cheaper to do this, for by roasting, the sulphur of the ore is reduced to but 3 to 4%. Many ores within the above limit are leady ores, and difficult to roast because of the fusible nature, but the matte that they produce is easy to roast for the elimination of sulphur. Ore intended for roasting may be simple, consisting of iron sulphide, or complex as shown by the following analysis of a roasted ore; Si0 2 , 10 % Fe and Mn, 27%; CaO, MgQ, and BaO, 2%; 386 THE METALLURGY Zn, 8.8% ; Cu, 0.4%, S, 6% ; Pb, 35%, and Ag, 50 oz. per ton. The base in the roasted ore was present as sulphide in the raw ore. The so-called oxidized ores consist of the carbonate of lead with a gangue of iron oxide, limestone, dolomite, and silica. Such ores though called oxidized, often contain a little sulphur, as sulphide (galena or pyrite) or as sulphate. Silicious ores are added to charges, in spite of the large excess of silica, because the gold and silver are present in quantity to pay to recover. The lead of the charge takes the gold and silver contained in such ore, while the silicious gangue is fluxed into a barren slag and sent to waste. Both iron ore and limestone are added to the charge for fluxing the silica, making a slag of a predetermined composition or type. If the fluxes contain gold or silver the metals can be recovered, since they go into the base-bullion or work-lead. Without gold or silver they are called barren or 'dead' fluxes. Ore carrying an excess of iron or lime over silica (called iron or lime excess) is in the same category, since the excess is useful for fluxing, and is credited in purchasing ores. Thus, ore containing 10% SiO 2 and 40% Fe is said to carry 30% iron excess. , Not all the slag that issues from the furnace is clean. At the spout where the matte flows, and in the shell lining the cavity of the fore-hearth, slag, containing drops of lead and matte, is found. When a slag-pot is emptied at the edge of the dump, there remains a shell or coating of solidified slag. This shell, half an inch thick, is found to contain drops of matte that did not entirely settle in the fore-hearth. This is particularly true when the fore-hearth has formed a thick lining and soon must be replaced by another. All this slag having value, and called 'foul slag,' is an acceptable addition to the charge because of the fusibility, and the coarse condition, permitting free passage of the air of the blast. Computation of the charge. To determine the amount of the fluxes (iron ore and limestone) to add to the charge, to give a slag of a desired composition, it is necessary to know the weight of the ores to be used, and the results of analysis of the ore, fluxes and fuel, and also the composition of the slag and matte that are to be produced. First example. Following is an example of a charge-sheet for the calculation of a charge containing no roasted ore. The charge is to weight approximately 1000 lb. ? a quantity sufficient to fill a charge-buggy. OF THE COMMON METALS. 387 CHARGE SHEET. Name of Ore Weigut Pb SiO 2 Feaud Mn CaO and MgO S ^g H 2 Wet Dry % Wt % Wt % Wt. % Wt. % Wt Oz Chrysolite No. 28 3.0 515 500 21.0 105 32.0 160 15.0 75 100 50 4.4 22.0 75 18.7 Ontario Iron Ore 5.0 2.0 52 20 1 200 150 50 200 200 105 60.0 30 4.0 1 30 6 4.0 6 203 20 57.0 3.5 2.6 1 114 7 4 2.0 3.0 54.0 1.5 1 6 108 2_ 167 0.5 0.7 1.1 23.1 75 1.9 20.6 Limestone Coke 950 201 Fe for Matte = _S_j U OS for Slag. etc. Fe for Slag 168 12.0 S for Matte V 2 Slag SiO, = 30.0 = 1.00 factor FeO = 40.0 = 31.1 Fe = 1.04 CaO = 20.0 = 0.67 Matte S = 20.0 Fe = 55.0 Cu = 5.0 Pb = 15.0 90.0 Fe* 55 1 2.75 factor S 20 The items and the analyses of the charge are written in. the proper columns. Lead ore, 'Chrysolite, lot No. 28' is used, to give approximately 10% or 100 lb. of lead. To this is added silicious 'Ontario ore' with the fluxes, to make up the remaining 500 lb. as experience shows necessary. If upon calculation we find that too little silicious ore, and hence too little flux has been taken to make up this 500 lb., we can increase it as needed. t Of the sulphur, we assume that 20% is volatilized, and that the yslag retains fffo ^s weight of sulphur. From the silica we compute the slag to be equal to ||^ or 6gD lb. Neglecting fractions we then have : 0. >-J7 Pounds. Sulphur in slag>$f of 600 lb 6 volatilized 20% of 23.1 lb 5 remaining for matte 12 23 Multiplying the 12 lb. sulphur in the matte by the factor 2.75, which expresses the ratio of sulphur to iron, we find 33 lb. to be the amount of iron needed for matte. This leaves 168 lb. iron for the slag. We have chosen a half-slag for this particular case, and find from it a factor of ( ?i = 1.04) expressing the ratio of silica to 388 THE METALLURGY -.. iron. Multiplying the weight of the silica by this factor gives 211 Ib. iron needed. We have, however, but 171 lb., and the difference can be made up by increasing the iron by 40 lb., or the iron ore by 80 lb., since the iron ore is approximately one-half iron. Again, multiplying the factor expressing the ratio of silica to lime (0.67) by the weight of silica, we find 136 lb. needed, so that we have 31 lb. too much. Now, since the limestone is approximately one-half CaO, we diminish the limestone 60 lb. Erasing the old values and substituting the new ones we carry the calculation through again. The results should be correct within a few pounds, and may be corrected again if not so. Fractions of pounds are neglected, and the weight of the fluxes need not be written nearer than 10 pounds, since variation in the charge, due to variation in the ore and weighing, easily may be greater than fractions of 10 pounds. The slag actually produced may vary from the desired composi- tion. When a new charge has been put on, and after several hours the slag has come down, that is, has begun to flow from the furnace, it is analyzed and the result known in two or three hours. The charge is then altered to give a correct slag. Before making any changes we must be sure that the slag is hot and well reduced. The total weight of the charge can be varied conveniently by varying the amount of silicious ore, retaining unaltered the weight of the lead. Second example. For a charge that is to contain roasted ore or other sulphur-bearing material, the following sheet gives a good illustrative example. We assume that 15% of fuel is to be used, and that the weight of the charge is to be 1000 lb. Let us say that we have ores coming in proportions to permit using them in the ratio indicated by the first two items of the charge-sheet (300 lb. roasted ore and 200 lb. lead ore). We can make up the remaining 500 lb. with silicious ore and flux, in the proportion that experience suggests, being assured that the lead-bearing ore furnishes about 100 lb. lead, or 10% of the charge. We find, including the lead in the silicious ore, that we have 104 lb. total lead which meets the requirement. Next enter the 150 lb. coke, estimating the percentage composi- tion from the following data (for example of Connellsville coke) with 11% ash of Si0 2 , 44%; Fe 2 3 , 22.7%; CaO, 7.3%; MgO, 1.9%. Since the iron has been reported in ferric form, we have to multiply the 22.7% Fe 2 O 3 by 0.7 to obtain the equivalent iron. Magnesia being the stronger base, we multiply the percentage (1.9%) by 1.4 to obtain the lime-equivalent of the magnesia, adding OF THE COMMON METALS. 389 it then to the lime, making in all 11%. Estimating this for the coke, by the percent of the percent, we have Si0 2 , 4.9% ; Fe, 1.7% ; Ca Oand MgO 1.1 per cent. CHARGE SHEET. Name of Ore Weight Pb S1O 2 Fe and Mn CaO and MgO S H 2 O Wet Dry % Wt % Wt % Wt. % Wt. % Wt Roasted Ore Lead Ore. 50 300 200 200 100 200 1000 12.0 25.8 8.0 36 52 16 104 12.0 22.6 63.3 5.0 3.5 4.9 48 45 127 10 3 7 240 32.0 18.1 16.0 540 1.7 96 36 32 51 3_ 221 2.0 5.4 52.0 1.1 6 11 101 2_ 123 5.0 2.0 30 0.7 15 4 6 1 26 Silicious Ore Iron Ore Limestone Coke Fefor Matte - = 37 S volatil- ized 5.0 FeforSlag=l4 S in Slag 6.0 11 15 For Matte Slag. SiOo =33% FeO =33% CaO = 24% Other bases = 10% Fe = 25.7 = 0.77 X SiO 2 = 0.73 X Si0 2 Matte. S = 20% 20% S X 2.5 = Fe = 50% Cu . 5% Pb = 157o 100% Slag = 730 Ib. S in slag = (730 X 0.8%) = 6 Ib. Write in the constituents in percentage and pounds, footing up the columns to obtain the total for each constituent. All fractions are neglected. At the lower left corner of the charge-sheet write the type of slag chosen from the list of type-slags, section 127. The type chosen depends upon the kind found commercially to be the most profitable. This again depends upon the cost of flux and upon the ores that give the most profit in treatment. Where it pays best to treat silicious ore, one uses a silicious slag, but where there is more profit in an irony ore, then a basic slag is preferred. For the present calculation we choose the three-quarter slag, F, containing Si0 2 , 33% ; FeO and MnO, 33%, and CaO and MgO, 24%. The other bases such as ZnO, A1 2 O 3 , alkalis and sulphur make up the remaining 10 per cent. The 33% FeO is calculated to Fe, making 25.7%. Finding the ratio of Si0 2 to Fe, we get the factor 0.77. For CaO we obtain, in the same way, the factor 0.73. On the other side of the sheet 390 THE METALLURGY tabulate a matte analysis, and calculate the factor expressing the ratio of S to Fe, or 2.5. First consider the sulphur. Experience shows that, of the amount of sulphur present, we can estimate that 20% or 5 Ib. will be volatilized, and also that the slag will contain 0.8% S. The weight of the slag is found by dividing the weight (240 Ib.) by the percentage expressed decimally (0.33). This equals 730 Ib. Hence there remains 15 Ib. of sulphur to enter the matte, and this multiplied by the factor 2.5 gives 37 Ib. Fe. This subtracted from the total Fe leaves 183 Ib. for the slag. Multiplying the 240 Ib. of silica by the factor 0.77, we find 185 Ib. Fe, so that we are close to the calculated iron needed. Again, calculating the lime needed (240 multiplied by 0.73 or 175 Ib.) we find we are short (175 minus 123) or 52 Ib., equal to 100 Ib. of limestone of 50% CaO. This increases the total of the charge 100 Ib., and we may let it go at that, submitting to a little less than 10% lead on the charge ; or we may decrease the silicious ore, which permits us to lessen the quantity of flux. Write the new figures 170 of silicious ore, and 270 of iron, and 260 of limestone, making in all 500 Ib. for the three constituents, and, after neatly erasing the old figures, try again. The amended calculation will be nearly correct, or if not so, make a new correction with the figures erased and new ones inserted. The final sheet should be a fair copy. This second example is a simple one. Frequently the metallurgist adds other items to the charge, and methods, other than the trial one here given, are too complicated. Since the atomic weights of Fe and Mn are nearly identical (56 to 55), we add the percentage of each metal to obtain the equivalent Fe. For MgO multiply by 1.4, and for BaO by 0.4 obtain the equiva- lent CaO. 131. SAMPLING AND HANDLING BASE-BULLION. The practice at large silver-lead smelting-works is now to re-melt all base-bullion from the blast-furnace. Sometimes the lead is taken in molten condition to the re-melting kettle from the blast-furnace. .When the lead is melted in the re-melting kettle it is carefully skimmed, and as in lead refining, the cleaned lead is molded into bars. The skimming or dross, containing copper and other impurity, is returned to the blast-furnace. The copper there enters the matte and the lead again goes to the base-bullion. While the lead is being molded samples are taken from the kettle at intervals, and from the samples the assay-results are obtained. The bars for a 40-ton car-load, 800 in number, are stamped with the number of OF THE COMMON METALS. 391 the lot, and are carefully weighed, 20 at a time. Careful assays are made of each lot, both by the shipper and by the refiner. At smaller plants the punch sample is taken as described in tKe chapter on sampling. The results are exact. 132. FLUE-DUST. A blast-furnace 42 by 120 in. takes 5000 cu. ft. of air per min. when in full operation. The escaping gas has an average temperature of 150C., and we calculate that the velocity of the gas rising through the charge is at least 5 ft. per sec. Additional air enters the feed-doors, especially at the time of charging, and particles of 20-mesh size may be carried into the down-take and to the flue that leads to the tall stack producing the draft. The flue, which is common to all the blast-furnaces, in some cases is made hundreds of feet long and of large cross-section, for the purpose of settling and collecting the particles called 'flue-dust.' Flue-dust, after suitable preparation, as for example by making it into briquettes, can be re-melted. It amounts to 0.8 to 15% the weight of the charge, but in good practice a near approach to the former figure is possible. A little of the lead and silver, and much of the zinc and arsenic of the charge are volatilized. The volatilized substance in part adheres to the cool surface of the flue. Eventually it flakes off and falls to the bottom of the flue and is there recovered. Flue-dust, therefore, is composed of: (1) dust, carried along by the draft, and (2) lead fume, condensed on the cool surface of the flue. All this material is not recovered ; the finest part may escape. An analysis of flue-dust at Pueblo, Colorado, shows PbO, 37.6%; ZnO, 5.3% ; Fe,0 3 , 25% ; A1 2 O S , 1.3% ; CaO (from the limestone) 5.3%; SiO 2 , 8.6%; S, 2.5%; S0 3 , 1.6%; H 2 0, CO 2 , and C (from the coke) 11.2%. When arsenic is present in the charge a part condenses in the flue. Where in outlying regions, the lead has a low value, there is a point at which, in order to eliminate the arsenic, it is better to permit the loss of some of the lead. There is a difference in the quantity of flue-dust made by a furnace according to the way it runs. A furnace having accretions or scaffolds in the shaft, or one driven with a large volume of blast, or one fed carelessly so that the gas fails to ascend evenly through the charge, increases the production of flue-dust. The carrying power varies as the square of the velocity, or directly as the draft-pressure, therefore, to arrest as much flue-dust 392 THE METALLURGY as possible in the main dust-flue, it is made large in sectional area thus reducing the velocity of the gas. Flues have been made of sheet-metal, but metal corrodes because of the sulphuric acid that is found in the gas or the sulphates that are in the flue-dust. Reinforced concrete has been used, but it also is somewhat attacked. Brick remains the favorite material for such construction. A flue built of reinforced concrete on the Monier system consists of an arched frame-work of angle-iron tied by longitudinal bars and covered by wire netting, or expanded- metal, the whole being plastered inside and out with a coating of cement-concrete. The bottom of brick flues is frequently made of sheet-steel hoppers. Since these are continually covered by the flue-dust, they are protected from corrosion and last a long time. A flue is rectangular in cross section and has brick walls, a low arched top, and a hopper-bottom, as above described, at such a height as to leave room beneath for a car running on a track at the ground level. The car can be set under any hopper and the content drawn into the car through a canvas sleeve fitted over the outlet spout to avoid dust. Other methods, such as spraying water upon the dust-laden gas ; the use of plates, called Freudenberg plates, hung in the flue parallel to the length; Rosing wires, a multitude of wires hung from rods parallel with the flue, all have been tried with a degree of success, but abandoned in favor of plain flues free from baffle-walls or other obstructions. The two important considerations in the construction of flues are, first, a slow and general movement of the gas, in order to settle out the dust; and second, a large cooling surface for the condensation of the lead-fume. 133. BAG-HOUSE. Fig. 163 is a transverse section of the bag-house which is coming much into use for the recovery of the flue-dust and lead fume by filtering the gas through bags that leave the resultant gas colorless. It is a building 40 ft. wide, 200 ft. or more long and 45 ft. high, divided into two stories. The bags, of which there are several hundred, 20 in. diam. by 35 ft. long, are suspended from the beams of the roof by cords that, at the same time, tie the mouths of the upper ends. The second floor, 14 ft. from the ground, is pierced with openings 18 in. diam. with thimbles over which the lower ends of the bags are slipped and tied. The flue-gas is drawn in by a suction- fan and delivered into the lower story, beneath the second floor. OF THE COMMON METALS. 393 The gas, distending the bags, becomes filtered and passes into the second story, escaping at the ventilator in the roof. The building is divided by several cross-walls, so that it is possible to cut out a section when it is desired to enter it to make repairs, or shake the bags one by one as is done once or twice daily. The dust falls into the lower chamber and accumulates there. When one or two feet deep it is ignited. It burns of itself, becoming agglomerated and hingeddoor Bag suspended from screw-eyes tytfl/8 (or -better Copper) w/ns Fig. 163. BAG-HOUSE, GLOBE SMELTING WORKS. in a condition favorable for feeding back to the blast-furnace. Analysis of the burned dust shows it to contain oxide and sulphate of the metals as follows: Pb, 75%; Zn, 3% Fe, 0.5%; As, 1.3%, and Ag, 4 oz. per ton. The bag-house has been introduced because of the complaint that fumes injure vegetation, and also because of the actual saving it effects. 134. BRIQUETTING FLUE-DUST. Flue-dust can be wet down and fed back to the blast-furnace. If fed a little at a time, it is simply carried again into the flue, but while wet, in occasional large charges, it may be fed so that most 394 THE METALLURGY of it is carried down and smelted. The effective way is to make it into briquettes with milk of lime as a binder. Fig. 164 represents a plant containing a White briquetting-press for making briquettes composed of flue-dust and milk of lime to which is added fine roasted ore. At the right in the figure, shown to be on a high plat- form, is a pile of quicklime. This is fed, together with water, into the lime-mixer, a trough divided transversely by a partition. One compartment is shown as containing the lime being mixed to a thin paste, while the other is now empty. The paste is drawn from either compartment to a horizontal double-shaft pug-mill* Each Fig. 164. WHITE BRIQUETTiNG PRESS. shaft is provided with mixing blades. Flue-dust from the pile at the front of the lower platform is shoveled into the pug-mill and thoroughly mixed with the milk of lime by the revolving blades of the pug-mill which, being set at an angle, propel it to the discharge- opening immediately over a troughed conveying-belt. It drops into the hopper of a six-mold briquetting-press where it is made into briquettes that drop upon a flat conveying-belt, delivering them to a pile. The briquettes may be used at the blast-furnace freshly made, but the usual plan is to dry them, as clay-bricks are dried. At times the briquetting is omitted, and the pug-mill mixture is wheeled to a drying-floor, or is distributed evenly upon one of the ore-beds. By the time the bed is used the mixture has set and become a hard mass capable of withstanding handling without being broken. The lead-smelting charge generally contains copper, and the copper accumulates in the matte. Since matte is roasted and OF THE COMMON METALS. 395 135. LEAD-COPPER MATTE. returned to the blast-furnace, the content in copper gradually increases. When increased to 12%, the copper matte is again roasted and treated in a separate blast-furnace, with silicious ore and oxidized copper ore, to produce a matte of 40% copper, called 'shipping matte' because it often is shipped to a copper works to be treated for the copper. A satisfactory way of treating matte, where it is wished to produce a more finished product, is to crush it to a 4-mesh size and roast in a hand-roaster. It is next sent to a blast-furnace and again smelted with silicious and oxidized ores of high grade in copper. There results a matte 65% Cu and a proportion of bottoms, the result of the reduction of the copper from the matte. The bottoms are charged into a reverberatory furnace through the side- doors, and the coarse-broken matte is put on top. The doors are closed, and the charge is fired with an oxidizing flame, as in the Welsh process of 'roasting.' The charge having melted, a reaction of the cuprous oxide on the cuprous sulphide takes place, as described in the Welsh process of making blister copper; and the charge becomes reduced to an impure copper containing arsenic, bismuth, and antimony, as well as the gold and silver that were contained in the matte. The copper is then poled to reduce the cuprous oxide, and ladled into anode-molds. The anodes are sent to an electrolytic copper refinery for final treatment. 136. COST OF TREATMENT OF LEAD ORES. To illustrate the method of calculating the actual cost of treating an ore, as in the Colorado or Utah silver-lead smelting practice, we take the case of a so-called neutral ore (SiO 2 equal to Fe). The ore is assumed to be oxidized, to contain less than 5% sulphur, and at least 10% lead. It is to be treated at a works having an output of 400 tons of charge daily. The cost of treating a ton of charge, and of treating a ton of the ore including the flux, is as follows: Cost per ton of charge. of ore. Labor $1.10 $1.54 General expense, assaying, and management.... 0.20 0.27 Fuel for power 0.07 0.10 Coke (15% of the charge) at $6.50 per ton 0.97 1.36 Interest, improvement-fund, and repairs 0.26 0.36 Limestone (0.3 ton at $1.25) 0.37 Iron ore (0.1 ton at $5 ) '. ' 0.50 $2.60 $4.50 396 THE METALLURGY The figures in the second column are obtained by multiplying the total weight of the charge, 1.4 tons, by the cost of each item per ton of material, and then adding the cost of the flux. This corresponds to the figure above obtained, and may be stated again as follows : 1.4 tons of material (1 ton ore, 0.3 ton lime- stone, and 0.1 ton of iron ore) at $2.60 per ton of charge for smelting $3.63 Cost of fluxes, 0.3 ton of limestone at $1.25. . . 0.37 Cost of fluxes, 0.1 ton of iron ore at $5.00 0.50 $4.50 In case the ore contains sulphur in quantity to require roasting, $2 should be added for the treatment. For sulphur over 5% and up to 10%, add 30c. per unit to cover the expense of iron ore for disposing of the extra sulphur, and roasting the matte made by it. The application of this method of calculating costs may be illustrated with a silicious ore. We make a charge-calculation and find that for one ton of ore we need 1800 Ib. iron ore and 1450 Ib. limestone. Thus we have : 5,250 Ib. or 2.625 tons of material of the charge at $2.60 per ton $6.82 1,800 Ib. or 0.9 ton ore at $5 per ton 4.50 1,450 Ib. or 0.725 ton of limestone at $1.25 1.08 $12.40 This $12.40 represents the actual cost of smelting a ton of silicious ore when the fluxes are paid for outright, and when run in a furnace also with suitable lead-bearing ores. In general, the lacking iron for the charge is made up in part by the excess of iron over that needed for the sulphur and silica contained in the roasted iron- sulphide, and in the other irony ores. A ton of the above silicious ore makes 1.5 tons of slag which, containing 0.6% Pb and 0.6 oz. Ag per ton causes a loss in these metals of $1.26 per ton. The self-fluxing ore above cited produces 1250 Ib. slag and carries off only $0.52. If we were treating this ore alone, of 100 tons treated daily but 38 tons would be ore, and from this all the profits of smelting would have to be obtained. In the case of the self-fluxing ore, 71 tons are metal bearing. Thus the room taken by the fluxes, called 'displacement,' is taken into account in figuring the profit of operation. PART VIM. ZINC PART VIII. ZINC. 137. PROPERTIES OF ZINC. Zinc is a white, brittle metal of 7.1 to 7.2 specific gravity. Commercial zinc, called spelter, contains lead, iron, and cadmium as common impurities. Lead is found in spelter to the extent of 2 to 3%, but the amount can be reduced to 1% by refining. A small amount increases ductility, but it is injurious in the better grades of brass. Iron may be present up to 0.05%. Cadmium seldom occurs in spelter and has no deleterious effect, except when the metal is used for making zinc-white. It then discolors the product. 138. ZINC ORES. The principal ores of zinc are blende and calamine. In New Jersey occur deposits of franklinite. The minerals seldom occur pure ; besides the earthy gangue, and sulphides of iron, lead, and copper, ffi Q y are fl " Q/ in^ntly centaminntcd with impuritiryh combined. chemically, nnpnmnllj - irun> which cannot be separated by ore- dressing methods. Blende, or sphalerite, when of a yellow color as in the ore of the Joplin district in Missouri, is called rosin-blende. When dark in color, due to chemically-contained iron as in the ore of the Rocky Mountain States, it is called black-jack. It is from blende that most of the spelter of commerce is extracted. It needs roasting before it can be retorted or smelted for extracting the zinc. Calamine is a term applied commercially to the carbonate (smithsonite) and to the hydrous-silicate of zinc. It is an oxidized or sulphur-free ore that needs no preliminary roasting before smelt- ing. On being heated in the retort, the CO 2 of the carbonate is expelled, leaving zinc oxide. Willemite, the anhydrous silicate occur- ing with franklinite in New Jersey, mixed with coal is decom- posed at the high temperature of the retort, yielding zinc. Never- theless it is generally found advantageous to calcine calamine for the purpose of driving off C0 2 and water, which are undesirable in retorting because of their oxidizing action on zinc-vapor. However, 400 THE METALLURGY the preliminary calcination is often omitted, but when performed, it is done in kilns much like those in which lime is burned. 139. METALLURGY OF ZINC. In outline, the metallurgy of zinc consists in grinding the ore (generally blende) and roasting to convert it into zinc oxide, then charging the roasted ore, intimately mixed with fine coal, into horizontal, cylindrical, clay retorts, heated to a white heat, where the zinc, reduced by the coal, volatilizes, and the vapor, entering the cool, tapering, clay extension of the retort (called the condenser), collects there. As it accumulates it is tapped into a ladle from time to time, skimmed, and cast in molds. When distillation is complete the condenser is removed and the content of the retort taken out and generally thrown away. The cycle of operations takes 24 hours. 140. ROASTING BLENDE. The aim is to dead-roast the ore, generally to 1% sulphur or less. For every 1% sulphur remaining in the roasted ore, 2% zinc is held back in the retort-residue after distillation. The ore is ground to 8-mesh size then slowly and carefully roasted with frequent stirring, finishing the roast at a high tem- perature to decompose the zinc sulphate formed at the lower tem- perature. The ore is generally in the form of concentrate, still containing a little gangue, galena, and pyrite. To remove the final 1% of sulphur would require a long time and would not be commercially profitable. The ore is accordingly considered to be finished when it contains no more than that amount of 'sulphur. 141. CHEMISTRY OF ROASTING ZINC ORES. We have, in the roasting of blende, the following reactions : (1) ZnS + 40 = ZnO + S0 3 43,000 86,400 71,000 = + 114,400 cal. (2) 2ZnS + 70 = ZnO + ZnS0 4 + S0 2 2x43,000 86,400 230,000 71,000 === + 301,400 cal. Thus in an oxidizing flame, blende is roasted to oxide and sulphate, both reactions being exothermic. As indicated in the reactions given in the chapter on roasting, pyrite or chalcopyrite assists in the reactions. At a cherry-red heat the zinc sulphate is decomposed into basic sulphate (3ZnO,ZnS0 4 ) thus: (3) 4ZnS0 4 = 3ZnO,ZnS0 4 + 3SO 3 The basic sulphate, exposed to a bright-red heat for a time, reacts thus: OF THE COMMON METALS. 401 (4) 3ZnO,ZnS0 4 = 4ZnO + SO 3 Finally zinc oxide is obtained and the sulphuric anhydride is eliminated. At a high temperature (900C.) the latter in part decomposes into sulphurous anhydride and oxygen as follows : (5) S0 3 = S0 2 + O When limestone or calcite is present it is converted in part to sulphate and in part to oxide. Galena also roasts to a sulphate, and tends to envelop particles of blende, and to prevent their roasting. Much of the blende from Leadville, Colorado, contains silver, and it consequently often pays to treat the retort-residues after the zinc has been removed. There is a loss of silver in roasting that may be given at 10%, and also a loss of zinc as dust and volatilization at the final high temperature that may be reckoned at 2 per cent. 142. ROASTING FURNACES. The roasting of blende has been performed in hand-rabbled reverberatory furnaces as well as in a great variety of mechanical furnaces. These are described in the chapter on roasting. The latter are gradually supplanting the former because of the saving of labor. It should be noted, however, that the wear is great on mechanical furnaces that have iron-work exposed to the heat because of the high final heat needed in blende-roasting, and consequently the types of furnace have been preferred where the rabble is exposed but a short time to the action of the fire, and where iron parts are not exposed or can be water-cooled. Thus, the Brown horseshoe furnace, where the rabble is drawn through a circular hearth, then allowed to cool, or the Wethey furnace, where the rabble is exposed to the fire but half of the time and the moving iron parts are outside the furnace, have been successfully used in blende-roasting. Of the recent types, the Hegeler furnace as used by the Matthiessen & Hegeler Co., La Salle, Illinois, and by the U. S. Zinc Smelting Co., Pueblo, Colorado, has proved most successful for the above reasons. It is a multiple- hearth furnace closed by swinging sheet-iron doors at the ends, and stirred by rabbles drawn quickly through the furnace by means of rake rods, so that the parts are outside the furnace most of the time, and no iron parts, except the end swinging doors, are affected by the fire. The hearths being superimposed make a compact furnace, and the radiation is greatly lessened, so that there is economy of fuel. A 75-ft. hearth Hegeler furnace roasts 48 tons of blende in 402 THE METALLURGY 24 hours, yielding 40 tons of roasted ore. The raw ore contains 30% sulphur, and the final roast 1.2%. The consumption of coal is approximately 20% of raw ore. 143. THE SMELTING OR DISTILLATION OF ROASTED ZINC ORES. The recovery of zinc from the ore consists in distillation of the roasted ore in refractory clay retorts after intimately mixing it with 40 to 50% the weight of fine coal. The whole is brought to a white heat which is maintained during an entire day. Before ore and coal are charged into the hot retort the mixture is moistened for convenience in charging, the water promptly being driven off by the heat. The light hydrocarbons of the coal come away next; then the iron oxide is reduced to protoxide and part of it to a porous iron or iron sponge. The final reaction is the reduction of the zinc oxide of the ore by the carbon to metallic zinc. The reaction commences at 1060C., but practically a temperature of 1300C. is reached. In the United States the Belgian system of retorting is chiefly used. The retorts 4 ft. long by 8 1 /-? i n - inside diameter are set horizontally in the hot furnace. The retort and charge being poor conductors of heat, 1.5 to 2 hours is necessary for the heat to pass from the hot exterior to the cold core. For ore high in iron, the furnace in that time becomes heated to a uniform temperature (1020C.) and the iron oxide is reduced. If the heat rises above this point rapidly, the reduction of the oxides of iron and zinc occurs simultaneously and the CO and C0 2 gas resulting from the reduction of the iron oxide tends to sweep the zinc-vapor through the condenser, and cause it to burn at the mouth. There is a difference in the reductibility of iron-bearing blende, so that a hard ore, difficult to roast, is naturally more difficult to reduce than a porous one that has been roasted at a low temperature. If, as sometimes happens, a zinc ore is roasted at so high a temperature as to form an incipient slag, especially an iron silicate, the reduction of the ore is retarded. The iron silicate collecting at the bottom of the retort attacks other bases and quickly cuts a hole through the retort. The iron sponge, where formed, is not, however, detrimental, but where sulphur is still left in the ore, it may combine with it to form iron sulphide thus: (6) ZnS + Fe = FeS + Zn " The zinc is released, it is true, but the iron sulphide formed at 1100 to 1200C. is corrosive in action upon the retort. OF THE COMMON METALS. 403 When the content of the retort has been brought to the reduction temperature for zinc oxide we have : (7) ZnO + C Zn + CO 86,000 29,000= -57,000 cal. Thus the zinc oxide is reduced by carbon to zinc vapor with the formation of carbon monoxide. The carbon monoxide escaping- from the mouth of the condenser, burns with a characteristic light blue flame, masked, however, by a small amount of zinc-vapor that escapes and burns at the same time. Any undecomposed ZnS, left in the roasted ore, remains as such unless iron sponge is formed to reduce it as above described. Zinc sulphate is reduced to sulphide according to the following reaction : (8) ZnSO 4 -f 2C = ZnS + 2C0 2 In the presence of an excess of carbon the carbon dioxide may be changed to carbon monoxide thus : (9) CO 2 + C = 2CO If, however, any dioxide remains unreduced, it tends to oxidize the zinc vapor back to zinc oxide, which then escapes from the mouth of the condenser and is lost. Lead sulphate, or oxide in roasted ore is reduced to metallic form and remains in part in the residue of the retorts and in part is volatilized and condensed with the zinc, contaminating the spelter. Cadmium oxide is present to the extent of a few tenths of a per cent in certain zinc ores. It is more volatile and easily reduced than zinc oxide, and condenses with the first of the zinc. It has been known to be present in spelter to the extent of 0.5%, but usually there is but a trace. It has not been found injurious to spelter. Silver and gold remain in the residue after retorting. In Missouri, where there is little or no precious metal, the residue is thrown away, but Colorado ores, carrying silver and gold, yield residues suited to further treatment, in the blast-furnace, for the extraction of the metals. To give an idea of the result to be expected from roasting and retorting an iron-bearing blende we take the following: A charge consisted of roasted ore containing Zn 44.8%, Pb 4.7%, Fe 18.0%, CaO 1.1%, MgO 0.7%, SiO, 6.0%, and S 2.9%. It was retorted in a charge of 60%) ore and 40% fuel for 24 hours, and there resulted 35% of residue computed on the total charge. The residue contained 0.73% Zn. 24.2% Fe, and 4.07% S, and of the total zinc,^95% was /rotaincfl. Where the same ore was imperfectly roasted and contained 7.7% S, it was found that 6.7% Zn was lost in the residue. 404 THE METALLURGY This shows that within commercially profitable limits, too much care cannot be exercised to effect a good roast. Now according to the rule stated in the chemistry of roasting, the loss should have been 15.4% Zn. We must accordingly infer that the greater extrac- tion of zinc is due to the reducing effect of the iron present reacting on sulphates and sulphides containing zinc. 144. THE ZINC FURNACE. Fig. 165 and 166 represent, in transverse section and in front elevation, a Western zinc-smelting furnace after the Belgian system, Fig. 165. ZINC-SMELTING FURNACE (SECTION). Fig. 166. ZINC-SMELTING FURNACE (ELEVATION). for the distillation of roasted blende. It contains 265 retorts, or 128 on each side. As shown in the transverse section, there are 8 rows of retorts, the rear ends of which rest on the middle wall of the furnace. The front ends rest on the front walls. Beneath the retorts are the fire-places, each 18 in. wide by 8 ft. long, two under each bank of retorts. These are fired from the four end doors. Bituminous coal is used, the fire-bed being 4 ft. deep, so that the coal is burned with a long flame that completely fills the interior of the furnace and surrounds the retorts. The products of com- bustion pass through outlet-ports in the roof to one of the two stacks, one at each end of the block or 'massive' of the furnace. Beneath OF THE COMMON METALS. 405 the grates ample room is left for a man to work, and with a long bar, to 'grate' or clean the fires, removing the clinker and ashes as they form. Two or three fire-bars, each 2 in. square, sustain the fire. This underground passage also serves for a tram-car in which the ashes are carried away, and the residue from the retorts removed. The end fire-doors have the sills at the ground level, 2 ft. below the lowest retorts. In Fig. 141 and 142 only the three lower rows of retorts shown are provided with condensers. In full action all are so provided. The block proper is 14 ft. high, 24 ft. long, and 12 ft. from face to face. The stacks or chimneys are 45 ft. high, one for each side of the block. Fig. 107. ZINC-SMELTING RETORTS. Fig. 167 shows in detail a zinc-retort in place in the furnace. It is made of fire-clay, 4 ft. long by 8.5 in. diam. and with walls 1.25 in. thick. In the figure a is the retort, which rests on a ledge on the rear wall of the furnace and extending just through the thin (4 1 /2-in.) front wall. The wall is held by buck-staves c which carry the tiles upon which the retorts rest. The whole is firmly bound together with tie-rods. When the retort has been charged, the clay condenser & is set in place, in which the zinc vapor issuing from the retort is to condense. As seen in the front view, the space between two buck-staves is divided by shelves which form 'pigeon-holes,' each of which contains two retorts. The retorts having been set in place, the opening around them is bricked up with pieces of brick and with clay. When a retort becomes cracked or otherwise useless, it can be readily removed by breaking away the temporary wall and another retort set in the place without disturbing the adjacent retorts. 145. OPERATION OF THE FURNACE. The roasted ore is thoroughly mixed with fine coal upon the floor of the retort-house with shovels, or better in a horizontal pug-mill, using water to dampen it. The charge for a retort con- sists of 60 Ib. ore and 40 Ib. coal, and is skilfully and rapidly 406 THE METALLURGY thrown into the retort, empty from the last charge and already red hot, using the special scoop shovel. Fig. 168, and filling it clear to the back. An iron rod is now run in along the top of the charge to form a channel for the escape of the gas, and the condenser _&, Fig. 167, a conical clay tube 2 ft. long is adjusted with, a piece of brick to sustain it in the exact position. The j j'mt * a 6 Fig. 168. CHARGE SCOOP. between the condenser and the retort is luted with a loamy clay. To make the joint tight, a crescent-shaped 'stamper,' Fig. 169, is used, by which the loam is compressed and the joint made tight around the condenser. The tool, before using, is heated to dull redness in one of the lower rows of retorts, called 'cannons.' These cannons are not used in retorting, being left in place to modify the direct intense heat of the fire below. Fig. 169. STAMPER. The retorts are now charged and brought up to a white heat (1300C.), the temperature in the retort being about 100 lower than that outside. As the charge becomes heated, the moisture and volatile gas come away, reduction of the iron takes place, and finally, the zinc oxide is reduced to zinc vapor, coming away together with the CO formed at the same time, and the reactions tt/uM/v^take place as already shown. The zinc vapor, escaping from the charge by its own tension, enters the cool J?sfcsit and liquefies by being cooled below the boiling point. In practice the tem- perature in the condenser is 550C. The end, or mouth, of the condenser, which is about 2 1 /4 in. diam. is loosely plastered with a handful of the charge-mixture. From time to time (four times in OF THE COMMON METALS. 407 the 24 hours) the accumulated zinc is scraped into a ladle held below the mouth, with a button-headed scraper. The opening is at once loosely plastered up again. Toward the end of the distil- lation the firing is more vigorous to remove the last trace of metal. The charge is put into the retorts early each morning and is under the action of the fire nearly 24 hours. After this time, the final zinc is removed, and the men take down the condensers, breaking the joint formed between the retort and condenser, and remove each condenser to be carefully scraped out and again used. The front layer of residue in the retort, which retains zinc, is again retorted. The rest of the content of the retort is scraped out with a suitable scraper or rabble. In Missouri it is often done by inserting a steam pipe to the back of the retort and blow- ing out the content. The residue is rejected if it contain no precious metal, otherwise it is smelted. The operation of charging, as already described, is at once begun, except when retorts are broken that have to be replaced. Replacing retorts. The average service of a retort, in good practice, is 40 charges. In the case of the block above described it would be necessary to replace six daily, though this will vary greatly from the average given. Some retorts break after a charge or two ; others have a long life. Thus a retort may be corroded by iron that eventually perforates it. A retort may have a crack, or may become cracked, so that the zinc vapor escapes ; and it must be replaced by a fresh one. Such an accident is indicated by the disappearance of the slight flame and smoke that generally issue from the condenser, there then being an inward current, due to the chimney-draft drawing the air through a crack of the broken retort. In such a case the charge may as well at once be removed from the retort. Retorts are not set in place when cold, but are heated red-hot in a coal-fired annealing-furnace. A supply of them is ready in this furnace each morning for use. The old retort is first removed by breaking out the adjoining wall of the 'pigeon-hole.' The new one when needed is brought to the zinc-furnace by three men, and at once put in place. The work is hot and arduous. The fresh retort, having been set, is quickly bricked in, and is ready for charging. A charge of 60 Ib. of 60% zinc ore if well roasted contains 45 Ib. zinc oxide or 36 Ib. zinc, and yields but 85 to 90% of the zinc. The residue of a 100-lb. charge containing ZnO that 408 THE METALLURGY persistently remains in the residue, and ZnS that is completely reduced contains: Pounds. Zinc sulphide and zinc oxide 2.5 Gangue and ash (from ore and burned coal) . .25.0 Unburned coal (two-thirds having been used in reduction) 13.0 Total . ...40.5 The reduction in the total weight (100 Ib.) is thus 59.5 per cent. 146. MANUFACTURE OF RETORTS AND CONDENSERS. Retorts to withstand the high temperature and corrosive action of the charge are made of the most compact and durable material. The material consists of a mixture of 'chamotte' 'grog' or 'cement/ composed of burned fire-clay, or made from old broken retorts, fire- brick, or tile free from slag. It is ground to 3 or 4-mesh size and mixed with an equal amount of raw fire-clay. The mixing is done in a pug-mill using 10% water to form a stiff mud or 'adobe,' which is allowed to stand 2 to 4 weeks covered with wet sacking to season and to develop the plasticity. It is again put through the pug-mill, and finally made into retorts in a hydraulic retort-making machine under the pressure of 3000 Ib. per sq. in. A machine of this kind makes 8 retorts per hour at a calculated cost of 50c. per retort. The success of the retorting operation depends upon the durability of the retorts, and thus upon the composition and manufacture. Since the bases in the charge attack the retort, we either must exclude them by selection of ores, or use less silicious material in the retort mixture. A retort carrying an excess of silica is more corroded than one containing a large proportion of clay or aluminous material. On the other hand, if we increase the proportion of clay, the retort is liable to shrink and crack in making, and it is not so infusible nor as stiff as one made of a highly silicious mixture. Good air-dried clay for retorts consists of 30% A1 2 3 , 50% Si0 2 , 15% combined water, and 5% bases. The requirements for a good retort are : that it be made of refractory material to resist the intense heat, that it be not quickly corroded by impurities, that it be dense so as to prevent the zinc- vapor from penetrating it, that it be strong to keep the shape when loaded with the charge of ore and coal and heated to a OF THE COMMON METALS. 409 white heat (1300 to 1400C.), and that it have thin walls, not to exceed l 1 /^ in., to permit the heat to be readily transmitted to the contents. It is evident that upon increasing the diameter of the retort the heat enters the charge more slowly, and hence we limit the diameter. The distance of 4 ft. between the supports is not exceeded because the retort would sag under the load at the high temperature. Condensers. Condensers are made of less refractory clay than retorts. They are not subjected to high temperature but must withstand much handling and severe treatment. They last 8 to 12 days, and cost 3 to 4c. each. Drying the retorts. The finished retorts are removed to a drying house and dried 6 to 25 weeks. Long drying gives better quality. They are first put into a room of ordinary temperature and kept 10 to 15 days until they can be safely handled. They then are taken to a hot-room and further dried at a temperature 30 to 35 C. until all the mechanically contained water is eliminated. Annealing. Before a retort can be put into the furnace it must be carefully annealed to remove the chemically combined water. Accordingly retorts are put into the annealing furnace and gradually heated, so as to be ready and at the required temperature on the morning of the day that they are to be used. As soon as they have been taken out others are put in. 147. LOSS IN DISTILLATION. The loss varies from 10 to 25% of the contained zinc, and occurs in several ways. The residue or ashes discharged from the retorts retains 2 to 10% zinc in the form of undecomposed zinc sulphide and zinc oxide that escaped reduction. The residue is higher in zinc at the extreme front of the retort. The lower retorts in a direct-fired furnace are hotter,, and hence the zinc is better expelled from them. Thus in a Belgian furnace, the residue from the upper row contains 9.15% zinc, from the middle row 4.67%, and from the lower row but 2.28%. On the other hand the retorts of the topmost row last 90 days, while those of the lowest row last only 6 days. Since in good work the residue contains 3 to 5% zinc, it follows that the lower the grade of the ore, the greater is the percentage of loss figured on the original content. Thus, on a Joplin ore of 60% zinc, we expect a recovery of 90%, while in upper Silesia, where the ore carries 25%, but 75% is recovered. A new retort does not begin to give the maximum output of zinc until it has 410 THE METALLURGY been in use several days, due to the fact that the clay absorbs metal, forming zinc aluminate, an artificial zinc spinel. It imparts a deep purplish blue color to the retort, and an old retort may contain 6% or more of zinc. Zinc is lost by filtration through the walls of the retorts as a result of the porosity and chimney-draft in direct-fired furnaces. With gas-fired furnaces, in the combustion chamber there is a little pressure which is one of the advantages of gas-firing. Filtration loss is greatly diminished by using retorts made under high hydraulic pressure, and by the addition of coke- dust to the clay mixture, to make the walls of the retort dense and impermeable. The escape of zinc vapor through a cracked retort is an important cause of loss. In direct-fired furnaces the loss is indicated by the cessation of the small flame issuing from the condenser, and between firings by the white smoke of zinc oxide escaping at the top of the chimney. In a gas-fired furnace in such cases, the flame at the condenser is greater, and becomes brownish red instead of the usual bright bluish-green. The action of corrosive slags on retorts results in making small holes in them, and zinc may escape in this way before the defect is discovered. The action on the retorts is more destructive at nights when most of the zinc has come away, and the firing, especially on the lowest retorts, is severe. Retorts weakened by the action of the corrosive slag are apt to develop holes, and such retorts are said to have been 'butchered.' When the condensers are removed, at the end of the distillation, there is a loss of the zinc-vapor still remaining, which escapes and burns to zinc oxide. The escape of zinc-vapor during the retorting is another loss, and the temperature of the condenser must be nicely regulated. If cold, a portion of blue-powder results, due to the . condensation of the zinc-vapor in powder instead of in liquid form. If hot, then zinc escapes condensation and burns at the mouth of the condenser with the greenish-white flame that is characteristic of zinc. Thus the regulation of the temperature of the condenser is no easy matter. The eu ^^^jL^ to maintain a temperature of jflf500C. or fifctP above the a3g7point of zinc. Escaping zinc may be partly saved by the use of prolongs or sheet-iron cones that fit over the end of the condenser. There is less advantage in using them than one would expect, since the material thus removed is in the form of blue-powder that must be re-distilled. 148. COST OF SMELTING. We give herewith the cost in Kansas of smelting blende- OF THE COMMON METALS. 411 concentrate per ton of raw ore. The roasted ore is assumed to be 86% of the weight of the raw ore. COST OF SMELTING PER TON OF DRY UNROASTED CONCENTRATE. Coal. Natural gas. Hand Mech. Hand Mech. roast- roast- roast- roast- ing. ing. ing. ing. Labor (except for repairs and renewals) $6.62 $5.82 $4.70 $4.25 Fuel (3 tons at 75c ) 2.25 2.25 Reduction material ($1 per ton) 0.47 0.47 0.60 0.60 Clay for retorts, condensers, etc., at 0.1 ton, , per ton of ore 0.26 0.26 0.26 0.26 Repairs, renewals, and sundry supplies (in- cluding starting furnace in operation).... 0.75 0.75 0.75 0.75 $10.36 $9.56 $6.31 $5.86 In the above table the assumption is made that the gas (natural gas being used) costs nothing. Since we have to meet the cost of acquiring the land for natural gas and that of pipe line, etc., additional cost for gas must be allowed. 149. THE SADTLER PROCESS. To counteract the corrosive action of the bases already referred to, Prof. Benjamin Sadtler patented a process that consists in lining the retort with basic material, preferably chromite, making a layer on the interior of the retort l /$, in. thick. The chromite is crushed to pass a 20-mesh screen and the dust is screened out, leaving a coarse granular powder. The interior of the retort is painted with a solution of water-glass, and while wet, sprinkled with the powder, in the same way that a house is painted and sanded to imitate stone. When the first coat is dry another is put on, the two coats giving the required thickness. The further drying and annealing is done as in making regular retorts. The lining fuses to the interior surface of the retort, and doubles the life (making 80 days), so that the retort fails by cracking, or by gradually accumulating residue rather than by corrosion. The retorts are especially adapted to treating ore containing iron and lead (such as the Leadville ores). The residue, still retaining the precious metals as well as lead and the excess of iron, is a desirable product for silver-lead smelting. The cost for putting in the lining is estimated to be 25c. per retort. The retorts have been used at the works of the Cherokee-Lanyon Spelter Co., in eastern Kansas, where Colorado ores, containing iron and lead, are successfully treated. PART IX. REFINING PART IX. REFINING. 150. PHENOMENA UNDERLYING THE REFINING OF METALS. It is found in analysis that the separation of a metal from other metals or from contained impurities is seldom complete. It is difficult and commercially impracticable to obtain metals entirely pure, so that metals that come on the market still contain small amounts of impurity. Metals thus prepared are graded according to quality, and command prices according to the gr.ade. Thus Lake copper commands the highest price of copper because of the purity and toughness, while casting-copper, which sells at ^c. less per pound, is used for making brass. In silver-lead smelting practice the slag, no matter how thoroughly settled and separated from the matte, still contains 0.2 to 0.3 oz. silver per ton and 0.3 to 0.4% lead. In copper refining, in the reverberatory furnace arsenic, antimony, and bismuth, occur- ing in the crude or blister copper, are retained in traces after refining, and where the blister copper is impure, no high-grade product can be expected. In the separation and deposition of copper by electrolysis at low current-density, the copper is of high grade even though impurities are in solution in the electrolyte ; nevertheless traces of impurity find their way into the cathode copper, though to less extent than by any other system of refining. In the refining of pig-iron to make steel, in order to obtain satisfactory quality, impurities must be removed until not more than 0.10% phosphorus and 0.05% sulphur are present, otherwise the steel lacks toughness and tenacity. 151. REFINING LEAD BULLION. Lead containing silver, commonly called base-bullion, is refined by the Pattinson or by the Parkes process. Commonly the Parkes process is used. The object in either process is to effect the. separation of the silver and gold from the lead. To get a clear idea of the principles of refining base-bullion (or. work-lead as it "is called in Europe) we first must know the 416 THE METALLURGY composition. An especially base quality is represented by the following analysis : Per cent. Lead 96.59 Per cent. Impurities : Cu 0.82 As 0.38 Sb 0.71 Fe 0.02 S 0.14 2.07 Precious metals: Ag (322 oz. per ton) 1.07 Au (0.20 oz. per ton) 0.0007 1.07 99.73 It is seen that base-bullion is principally lead. The problem is to soften the lead by removing the impurities, and then to separate the gold and silver from the purified or softened lead. In studying the process, the student should refer to Fig. 170. 152. SOFTENING. The furnace. Softening is performed in a water-jacketed reverberatory-furnace, Fig. 171. The rectangular hearth of the furnace, 7 by 14 ft. in size, is surrounded by a sheet-steel double water-jacket, shown in section at a in the sectional elevation, c. The jacket resists the action of the molten litharge formed from the lead in the operation. The furnace is heated by the fire-box, having a grate 4 by 5 ft. in dimensions, so that a high temperature can be attained in the furnace. The letters c, c indicate the rear working doors and &, & ; & the front doors in which the base-bullion is charged. At the front of the furnace the tap-hole e is provided, through which the lead is tapped when the charge is finished. The furnace communicates to a stack 50 ft. high by a flue at the front end. Operation. The work is done in two stages. In the first stage at a low temperature, the copper is removed. In the second, at a high temperature, the arsenic and antimony are expelled, after which there is left only the softened lead containing the precious metals. The base-bullion, in charges of 30 tons, is placed in the furnace by means of a long-handled paddle or 'peel' having a blade 2 ft. long by 6 in. wide. The bars are laid one at a time upon this and placed as desired in the furnace being piled in a heap on the hearth. The doors are closed and the bars are gradually melted down, the dross contained in the bullion rising to the top. When melted the heat is maintained slightly above the melting point, OF THE COMMON METALS. 417 but not higher. In about two hours the dross that has risen to the top is carefully skimmed, by means of a long-handled perforated skimmer, and removed through the door to a wheelbarrow placed for it. The dross, residue, or skimming, called the 'copper skim', consists of a drossy lead containing the iron, sulphur, and 418 THE: METALLURGY (especially important) most of the copper of the base-bullion. The removal of these completes the first stage of the process. The liquated dross thus skimmed, which may amount to 5% of the charge or 1.5 tons, consists of Pb, 62.4% ; Cu, 11.91% ; Ag, 0.17% (49 oz. per ton) ; As, 2.32% ; Sb, 0.98% ; Fe, 0.43% ; S, 4% ; and O, 1.87%. Slag, ash, and hearth material also are contained and must be reckoned in. The heat of the molten bath is now raised to a bright red ; and Fig". 171. SOFTENING FURNACE. the flame, made oxidizing by the admission of an excess of air through the thin fire, sweeps over the surface. Litharge forms, and the antimony and arsenic oxidize and enter the litharge slag. The litharge at this temperature has a corrosive action upon the brick lining, hence the need of a water-jacketed furnace. This stage of the process lasts 12 hours, until a sample of the lead taken from the furnace and placed in a mold and skimmed, shows by the appearance that it is free from arsenic and antimony. Before OF THE COMMON METALS. 419 the antimony is removed the surface of the molten lead will 'work,' or show oily drops moving upon it. A similar phenomenon is seen in the first stage of cupelling base-bullion high in antimony and aisenic. As the softening proceeds the drops become fewer and smaller, and finally a coating is seen to dull the surface of the hot molten lead, indicating the completion of the softening. For impure base-bullion this stage is of more than 12 hours' duration, and the thick layer of litharge formed retards further oxidation. It is best then to draw the fire and to cool the charge, to allow the litharge slag on the top to solidify above the liquid lead beneath. The slag is skimmed with a long-handled perforated skimmer (compare with Fig. 175), and the charge is fired again if necessary until the impurities are removed. The 'antimony skim' consists of the antimonate and arsenate of the lead with a large proportion of litharge. It is in fact an impure litharge. The softened lead to be treated either by the Pattinson or the Parkes process for the removal of the contained gold and silver, is now tapped into the desilverizing kettle, 8 ft. diam. and capable of holding 30 tons of lead. 153. THE PATTINSON PROCESS. When a kettle containing molten lead is allowed to cool slowly, as it approaches solidification, crystals of lead low in silver separate. The metal that remains liquid contains the larger part of the precious metal. The crystals are removed with a perforated ladle, melted in another kettle, and allowed to cool. Once more, crystals separate that are low in silver, the mother liquor becoming high in silver. If the liquid portion first mentioned be transferred to a kettle and likewise heated and then allowed to cool, the same segregation of the silver into the liquid part continues. We can accordingly arrange a series of kettles containing, at one end, low-grade lead, and at the other high-grade, all from one product. A series of this kind, as illustrated by practice at Eureka, Nevada, gave the assays quoted in the table below. Another crystallization would reduce the silver of the market lead to half the value given. The rich lead could be directly cupelled in an English cupelling furnace, or better, treated by the Parkes process to get rich silver-zinc crust for retorting and cupelling. The process has been modified -recently by Tredennick who raises each kettle by hydraulic power above the adjoining one so that the mother liquor drains from one kettle to the next through 420 THE METALLURGY a strainer, the lead being cooled near the solidification temperature by introducing steam upon the surface. The cost of operating has been greatly decreased in this way. The chief advantage of the Pattinson process over the Parkes is that it gives a product free from bismuth. In the Parkes process the bismuth follows the lead. Bismuth is injurious in lead that is to be corroded to make white lead, and it may be necessary to employ the Pattinson process for making a corroding lead from bismuth-bearing ores. Market Lead Kettle oz. Ag per ton. No. 1 1.25 No. 2 2.5 No. 3 5.0 No. 4 9.0 No. 5 .* 18.0 No. 6 30.0 No. 7 50.0 No. 8 75.0 No. 9 100.0 No. 10 150.0 *No. 11 450.0 *Rich Lead. 154. THE PARKES PROCESS. Operation. The softened lead from the softening furnace is tapped into a hemispherical cast-iron kettle, shown in Fig. 172 and 173, which holds 30 tons or more of lead or the full charge from the softener. The kettle is set in brick-work, and is heated from a fire-box below. In modern practice kettles are made large and are 10 ft. diam. and 2 ft. 10 in. deep, holding 60 to 65 tons. The principal of the separation of silver from lead depends on the affinity of silver for zinc which is greater than for lead. Upon adding and thoroughly mixing in a small amount of zinc (\% & 4ho whole), the 'nna takes up most of the silver. Zinc has a greater affinity than lead not only for silver but for gold and copper. When the molten bath is allowed to stand a while, the zinc being lighter separates and rises to the surface. At a temperature below the melting point of zinc but above that of lead, a crust forms that can be skimmed off. Thus the silver is concentrated in a small bulk of metal, and is later separated from the rich metal by further treatment. OF THE COMMON METALS. 421 The lead from the softening furnace flows along a cast-iron Another crystallisation could reduce tho GJlvor of the markci trough to the kettle. In so doing a litharge dross, called 'kettle dross' forms, and collects on the surface of the metal, and is skimmed off. The molten bath is next heated to an incipient red heat, well above the melting point of zinc and cakes or ingots of spelter equal to about 1.2% of the weight, or 720 Ib. are added. The quantity required varies with the richness of the base-bullion in silver. The added zinc quickly melts. Desilverizing machinery is now much used. The most approved machine is the Howard, used both for mixing and skimming. Fig. Fig. 172. HOWARD MIXER. 172 represents, in section and elevation, the kettle and the apparatus used for intimately mixing the molten zinc with the lead. The machine is brought to the kettle by an overhead crawl h and is lowered into it by a chain-block hoist. When lowered into position (shown in section Fig. 172) the screw propeller b is set in motion by a steam-driven mechanism to produce a downward flow of molten lead in the sheet-iron cylinder a. The cylinder has neither top nor bottom, and being submerged in the lead, a circulation is started, the lead flowing in over the top of the cylinder. Thus a 422 THE METALLURGY thorough mixing of the content of the kettle is assured. In a few minutes the engine is reversed, and the flow is made upward over the edge of the cylinder, then downward to the bottom. ^The mixing Fig. 173. HOWARD PRESS. Fig. 174. HOWARD PRESS (SECTION). is continued about 11 minutes, after which time the stirring apparatus is bodily hoisted and moved to one side. Several kettles can be thus served by one mixer. The content of the kettle is now allowed to cool two hours or OF THE COMMON METALS. 423 more. The light zinc rises to the top and carries the silver, gold, and copper with it. Finally when the temperature falls below the melting point, a half-fused, mushy crust or layer forms upon the lead. The crust consists of 65% Pb, 10% Ag and Au, 3% Cu, and 22 to 24% Zn. Fig. 173 is an elevation of the Howard press by which the zinc crust is removed from the lead. Fig. 174, another elevation, shows also a section of the cast-iron pot into which the press is about to be lowered. In principle the machine is like a cheese-press. The apparatus is lowered into the lead until the top edge of the cylinder a is but slightly above the surface. The plunger or follower c is raised, and the zinc-crust, as it is skimmed from the surface, is put in it by means of the perforated skimmer, Fig. 175. Fig. 175. SKIMMER. The press thus expedites the skimming. While one man is skimming and putting the skimming into the press, another assists by pushing the crust to one place with a wooden rabble. When full, the press is raised and the surplus lead begins to run out of the half-inch holes in the hinged bottom 1). The plunger c is brought down, squeezing out more of the lead, and leaves the remaining, mushy half-fluid mass nearly free from lead, of the composition given above. The press is now run to one side over a floor paved with cast-iron plates, the hinged bottom & is dropped by releasing the catch, and the zinc-crust is pushed out by continuing the down- ward movement of the plunger. The crust falls upon the cast-iron floor-plate, and while soft is readily broken with hammers into lumps the size of the fist. Meanwhile the hinged bottom is closed, and the press returned to the kettle and is opened to receive more skimming. These operations continue until the surface of the lead is well skimmed. The crust amounts to 3000 Ib. and contains 90% of the silver originally in the softened base-bullion, resulting in a concentration of twenty into one. This first 'zincking' removes all the gold and copper for which zinc has a great affinity. It does not remove all the silver, and the operation must be repeated once or twice more before the silver 424 THE METALLURGY content is diminished to the fraction of an ounce per ton beyond which it does not pay to go. Of the 1.8% zinc needed, first is added % of the zinc or 1.2%, then ~y, or 0.45%, and finally the remaining Vi 2 > or 0.15% ; or 900, 270, and 90 Ib. respectively. The desilverized lead remaining in the kettle after the last skimming retains 0.6 to 0.7% zinc and traces of arsenic and antimony, all of which must be removed before the lead is suitable for market. This is done by siphoning or tapping the metal from the kettle into a reverberatory furnace similar in construction to the softening furnace. Here the charge is brought up to a bright- 176. MOLDING MARKET-LEAD. red heat, the zinc is volatilized and burned off, and litharge forms as a slag upon the surface of the lead. The operation takes six hours, and the completion when the zinc has been expelled, is known by taking a sample of lead in a mold and observing the appearance of the surface as the metal solidifies. The furnace is allowed to cool until the litharge-slag is solid and can be skimmed. Finally the lead is tapped into a market-kettle similar to the desilverizing kettle. This is the reservoir from which it is drawn to be cast into molds. The molds, 50 in number, standing in a semi- circle as shown in Fig. 176, hold 100 Ib. of lead each, and are con- veniently mounted on two wheels by which, when full and cool, they are transferred to the adjoining floor. There the lead is tilted OF THE COMMON METALS. 425 out and the molds at once returned to the semi-circle to be used again. The lead is withdrawn from the kettle by means of a siphon. It descends into a small cast-iron pot, into which is screwed the 2 1 /2-in. pipe that delivers it to the 50 molds, the pipe being quickly moved from mold to mold as filled, without interrupting the flow. (b) ' FABER DU FAUR RETORT. At the end of each round the flow is interrupted only to carry back the end of the pipe to the first mold of the series, which mean- while has been emptied and replaced. The 100-lb. pigs, or bars, are the desilverized lead of commerce. Dry steam may be blown into the molten lead in the kettle to refine it. It is introduced by means of a pipe inserted deep beneath the surface. The constant agitation produced by the steam brings 426 THE METALLURGY the metal in contact with the air and oxidizes it and the remaining impurity. It is softer than ordinary desilverized lead, and is easily corroded by the acetic acid used in making white lead. It is accord- ingly called 'corroding lead.' Fig. 1' ENGLISH CUPELLING FURNACE. Section C.D. V3 =t Fig. 179. ENGLISH CUPELLING FURNACE (DETAILED SECTIONS). Retorting. Referring to the flow-sheet (Fig. 170), we see what becomes of the crust or skimming that results from the first zinck- ing. The material is in lumps, containing 22 to 24% zinc. It is charged with charcoal breeze, into bottle-shaped retorts /, Fig. 177, OF THE COMMON METALS. 427 each holding 1200 Ib. zinc-crust. The figure represents at a a sectional elevation through the retort, at b a transverse section, and at c an elevation of a Faber du Faur tilting retort-furnace. The retort rests upon a narrow arch, and carries a grate upon which rests a coke fire that fills the furnace and covers the retort /. The products of combustion escape by an outlet-port at the back to a stack of good draft. The coke is fed through a hole in the roof of the furnace, and is poked down, and kept in vigorous combustion, so that a yellow heat (1000C.) is attained. A condenser, made by cutting off the end of an old retort (as shown at c Fig. 177), collects the zinc vapors distilling from the charge, the condensed zinc being drawn into molds x through a one-inch hole, bored through the bottom edge of the condenser. When distillation is complete, the condenser and the supporting truck is removed, the furnace is tilted or revolved by means of a lever on the trunnions, and the remaining 'rich-lead' is poured into molds like those used for molding market- lead. The rich lead still retains zinc, copper, and impurities, taken into the crust at the first zincking. Cupelling. To obtain the silver (and gold) from the alloy, the English cupelling-furnace, Fig. 178, 179, is used. The principle of the action is much like that of cupellation in assaying, except the litharge here not only saturates the cupel, but flows from it as fast as formed. Fig. 179 is a sectional elevation of the furnace, showing the fire-box a where a long-flaming coal is burned, the products of combustion passing to the chimney by the port ~b and an underground flue. The flame plays over the hearth called a 'test,' a large cupel, not shown in Fig. 179, but shown mounted on the carriage in Fig. 178. The test is lowered by the jack screws, removed on the carriage, and another is put in, when the first is consumed. The test is hollowed like a cupel, to hold a shallow bath of molten lead 3 in. deep. .Fig. 180 shows two views of the test and the supporting truck including a view of an inverted truck and test. In the elevation of the furnace, Fig. 178, is seen the overhead pipe, branching to the ash-pit, to supply undergrate-blast, and to an opening at the back of the furnace where a tuyere is inserted, by which a stream of air is brought to play upon the surface of the molten red-hot bath of rich lead. The air oxidizes the lead to litharge. Other impurities are oxidized and enter the litharge-slag and are carried away with it. The molten litharge, as it forms, runs along a shallow groove or channel in the top of the front edge of the cupel or test. A door at the front can be lifted to inspect the oneration, or to cut the channel as needed. At the rear are 428 THE METALLURGY provided two ports of a size to permit inserting two bars of rich lead that are pushed in as fast as the cupellation proceeds. The ends of the bars melt and supply the lead. The litharge stream is the size of a lead pencil, and falls into a small slag-pot beneath. The lead is fed at the rate of one to two tons daily until the bath has become rich in silver, when the feeding of the lead must be stopped. Oxidation then is continued, cutting the channel deep to allow the remaining litharge to flow out, and finally the mirror-like bath of silver appears. The fire must keep the temperature above the melt- ing point of the silver. At the last, a shovelful of bone-ash is thrown on the bath to absorb the remaining trace of litharge as it fo^ms on the surface. This is skimmed, and the silver is then ready to be ladled out or tapped, commonly into cast-iron molds each holding sft i Fig. 180. TEST FOR ENGLISH CUPELLING FURNACE. 1000 oz. silver. The silver is then subjected to the acid-parting operation to be described later. The copper-skimming, which is the first obtained from the softening furnace, is returned to the blast-furnace where the sulphur of the charge combines with the copper and removes it as matte. The rest of the skimming, mostly lead, containing silver and gold, is reduced to base-bullion. The third skimming of the softening furnace, if any, is returned to the blast-furnace since it contains but little antimony. The second softening skimming or antimony-skim, containing 15 to 25% antimony, goes to a small reverberatory furnace, 8 by 12 ft. hearth dimensions and 10 in. deep, built like a softening furnace and called a precipitating furnace. Here it is melted, with a reducing flame into a slag. Charcoal is added, and stirred, to reduce or precipitate part of the lead of the slag. The lead, falling OF THE COMMON METALS. 429 to the bottom of the bath, carries down the silver of the slag. When the reaction is complete the supernatant slag is tapped into slag- pots, and the lead is tapped into a kettle at a lower level and molded into bars. This precipitated lead-bullion is returned to the softening furnace to be softened and desilverized. The antimonial slag, containing about 6 oz. silver per ton, when accumulated, is smelted in a small blast-furnace to reduce it to antimonial lead of 20% Sb, which is sold to the type founders. The slag is rejected. 155. VARIATIONS IN METHODS OF REFINING BASE- BULLION. Instead of removing the gold and silver together, in the first zinc-crust, it is possible to take the gold, copper, and a part of the silver by a separate zincking, in which a moderate amount of zinc (250 to 350 Ib.) is used. The separation is due to the fact that zinc has a greater affinity for gold and copper than for silver. We thus obtain a zinc-crust containing all the copper, gold, and some of the silver. From this crust, upon retorting and cupelling, a dore bar (a silver bar containing the gold) is produced. The second zincking accordingly needs but 600 Ib. zinc, and removes most of the remaining silver. A final zincking makes the lead practically clean. The second zinc-crust yields a silver bar free from gold, on which the expense of parting is saved, and the full price of commercial bar-silver is obtained. In the practice of early days, on clean base-bullion, made from clean ores, softening was performed in the kettle. The bars were melted at a low temperature, and the dross was made as dry (free from lead) as possible before skimming. The skimmed lead was brought to a red heat by vigorous firing, and stirred to furnish contact with the air. Arsenic and antimony were present in limited quantity and could be removed in this way. For ordinary base- bullion the method is inadequate, and the softening-furnace must be used. When lead has been desilverized, the contained zinc can be removed in the desilverizing kettle without the use of a reverberatory calcining furnace. This is accomplished by heating the lead in the kettle to a high temperature and rabbling by means of dry steam or compressed air, admitted into the kettle through a pipe that leads to the bottom. As the steam rises it agitates the hot metal and brings it in contact with the air above the kettle. The zinc is thus in part volatilized, in part oxidized with a little 430 THE METALLURGY of the lead. A powdery litharge is formed that is skimmed off, leaving market lead. In the old way of stirring the zinc into the charge of the desilverizing kettle, the work was done by means of paddles (shown in Fig. 181) having a handle 7 ft. long. Two men worked with the paddles, one on each side of the kettle. The paddle blades were Fig. 181. STIRRING PADDLE. moved as in the operation of rowing, thoroughly mixing the zinc into the lead in 15 minutes constant stirring. Where the Howard press was not used, the first or silver-crust, from the desilverizing kettle, was transferred to another kettle and kept at a low-red heat. In consequence there was a separation of a liquid portion that was retained and sent to the next charge, the dry residue being the copper skimming. The bars were put into the softening furnace to complete the refining. 156. COST OF REFINING BASE-BULLION. The actuual cost of refining base-bullion is as follows : Prime or flat-cost, of softening and refining $5.00 to $6.00 General expense 3.00 to 3.00 Loss in metals and- incidental expenses 1.70 to 3.00 Total $9.70 to $12.00 157. COPPER REFINING. Blister copper, or black copper, whether produced in the blast- furnace, the converter, or the reverberatory furnace, or by melting the concentrate from the native copper ore of the Lake Superior region, still contains impurity, principally arsenic with sulphur and iron, and all impurity must be removed by refining. If the copper contains gold and silver in quantity to warrant (20 to 40 oz. silver per ton), it is melted without attempting to refine, and cast into anodes that then are subjected to electrolytic refining. If there is but little precious metal in the copper it may be directly refined in the copper refining-furnace. Fig. 182 is a sectional OF THE COMMON METALS. 431 3levation, and Fig. 183 a sectional plan of a 40,000 to 50,000 Ib. copper-refining furnace. It is 14 by 19 ft. hearth dimensions, and has a fire-box 5% by 6^2 ft., or 30 ft. area, and carries a fire-bed L 4% ft. thick. The hearth, 2V 2 ft. deep, has a brick or a sand bottom. If of sand, the bottom is carefully smelted in. Beneath, the hearth is vaulted for ventiliation. The bridge, 5 ft. wide, is strengthened by a double conker-plate, and on either side and in the roof over the fire bridge, are ports that are opened when an oxidizing flame 432 THE METALLURGY is desired. In the elevation, at the front end, is to be seen the outlet-flue that leads to the stack or chimney. The chimney is close to the furnace but is not shown in the plan. The charge of ingots of blister-copper is put in at the side door. The door is then tightly closed, and vigorous firing follows. The charge melts after several hours. The front door is then opened, and whatever slag formed during the melting is skimmed. Next follows the rabbling, the object of which is to oxidize a portion of the copper and the impurities with it. The operation years ago consisted in striking the surface of the bath with a rabble in such a way as to splash the metal and agitate it, thus exposing it to the action of the air. The present way is to insert a %-in. pipe just beneath the surface of the metal and force compressed air through it to agitate, and at the same time to oxidize it. The air-ports of the furnace also are opened and the flame is made an oxidizing one. The action proceeds to the stage of 'set copper,' Cu 2 O having been by this time formed, and in part dissolved in the copper. Iron, sulphur, and arsenic partly volatilize, and partly oxidize and enter the slag that is formed at the same time, this is skimmed off. The copper oxide must be removed by poling. This is a reducing action in which the air-ports are closed to give a reducing flame, and spruce or poplar poles are inserted at the front door into the metal. The outer end of the pole is raised to force the butt-end beneath the surface of the metal. At the same time a wheelbarrow-load of charcoal is thrown in to cover the surface, to exclude air, and to reduce cuprous oxide. As the hydrocarbon of the wood is evolved and the moisture evaporates, that is as the wood burns, reduction takes place. The operation requires an hour or two. Additional poles are inserted to replace those consumed. Samples of a few ounces of the copper are removed in a small ladle from time to time and examined to note the progress of reduction. The 'tough pitch,' (the point at which the cuprous oxide is completely reduced to metallic copper) is the end in view. The charge is now ready for dipping or ladling. Hand-ladles, holding 25 Ib. or, large 'bull ladles,' holding 200 Ib. and carried by an overhead trolley or crawl, are used. The dipping or molding consumes three hours. The copper is kept hot by occasional firing, and by keeping the surface of the metal covered with charcoal. The charcoal serves also to keep the copper in pitch, or in the condition of tough copper. The molds into which the copper is poured from the ladles are of the shape required by the trade. There are required; ingots or bars suited to re-melting for making OF THE COMMON METALS. 433 brass ; wire bars of a form convenient for roiling into wire ; and rectangular cakes, often 18 in. square and 4 in. thick, but also of dimensions giving 2000 to 4000 Ib. weight. The size of the cake is suited to the size of the sheets of copper into which they are to be rolled. 158. MELTING AND REFINING LAKE SUPERIOR COPPER. The product from which copper is made, in the Lake Superior region, is a concentrate (called locally 'mineral'), which averages 70% copper in native form, accompanied with a self-fluxing or fusible gangue. In addition there occur pieces of copper of different sizes, from that of the fist to several tons in weight, called mass- copper. The small pieces are handled easily, and are shipped to the smelting works in barrels. It is called barrel-work. The larger pieces called 'mass copper' or simply 'mass,' are 70% copper. For the large pieces that cannot be charged at the side doors, a hatch- opening with a clamped brick cover is provided. Large pieces are raised by a crane and charged through the hatch, also concentrate or mineral to make a charge of 36,000 to 40,000 Ib. The charging takes place immediately after the dipping and the repairing or fettling the furnace. The furnace is now closed, and firing proceeds for several hours. As the charge melts and slag forms, it is skimmed until the metal is completely melted and the surface is clear. The operation of refining then continues as has been described above. The slag contains 15 to 25% copper, partly as entrained prills and flakes, and partly as cuprous oxide. This slag is smelted in a blast-furnace with added limestone to make the resulting slag fusible, using anthracite coal and a portion of coke for fuel. The charge consists of slag 2000 Ib., limestone 600 Ib., and anthracite coal 400 Ib., and the slag resulting contains 41% SiO,, 25% CaO, and 22% FeO. It is the practice at one works to add much small mass-copper and barrel- work to the charge, the idea' being that as the copper melts and sinks into the crucible of the furnace in the form of drops, it carries down particles of reduced copper with it. The products of the furnace are slag, of less than 1% copper, and cupola blocks (impure blister-copper) . In recent practice in the Lake Superior country the operation of melting is performed in one furnace and refining in another. A furnace, 18 by 40-ft. hearth area melts 100 tons of mineral of 67% copper in 24 hours using 30 tons of coal. The charge is supplied in two portions, the second following as soon as the first is melted. 434 THE METALLURGY Slag is removed several times during the period, and the copper when free from slag is tapped into a refining furnace at a lower level. The refining is done in a 14 by 22-ft. furnace carrying a deep charge of copper. The charge is made from copper scrap, high- grade 'mineral' of over 80% copper, and mass-copper. It is melted, and when ready the charge from the melting furnace is added, so that the total charge consists of 275,000 to 300,000 Ib. copper. The large quantity of copper is refined by rabbling or poling. To cast it in a reasonable time, a casting machine is employed which trans- forms the copper into commercial shapes at the rate of 50,000 Ib. per hour. The cupola-blocks, referred to above, w T hich constitute the product of the blast-furnace smelting of the slag from the rever- beratory refining-furnace are melted and refined to produce a low- grade copper called 'casting-copper.' The cupola-blocks are impure and contain so much arsenic that it is practically impossible to remove it all. Other impurities (iron and sulphur) are eliminated. 159. ELECTROLYTIC COPPER REFINING. Blister-copper from reduction works in the Western States is profitably treated by electrolytic refining. Not only can a pure copper be produced from impure material, but the gold and silver in the blister-copper can be separated and recovered. The copper from the reduction works, in the form of rough ingots, is sampled to determine the content of precious metal and the copper. The electrolytic refining of copper is well illustrated in the practice of the Raritan Copper Co., Perth Amboy, N. J. Sampling is done by drilling each ingot or bar with a %-in. drill, the drillings being then mixed for the sample of the lot. The copper after being sampled is sent to the re-melting-furnace which holds 150,000 Ib. copper. It is melted, poled, and cast into smooth anodes by means of a Walker casting-machine. The anodes, weighing 300 Ib. each, coming from the machine, are trimmed to remove fins or projections of copper, and are loaded on trucks holding 22 anodes in a rack. The trucks are moved to the tank-house where the anodes are picked up, and a group of 22 forms the charge for a tank. There are 420 tanks in all, each tank being in electrical connection with the adjacent ones. The current passes through each of the 22 plates or anodes to the cathodes placed between, and anodes and cathodes are 2 in. apart. Assuming that the anodes are 2 by 3 ft. in size, we have, in each tank, a total area of 264 sq. ft. through OF THE COMMON METALS. 435 which the current is passing with a density of 15 amp. per sq. ft. We thus have a total of 3960 amperes. The plates are immersed in a solution (electrolyte) of copper sulphate (blue-stone) in dilute sulphuric acid, the CuSO 4 being 15% (4% Cu) and the H 2 S0 4 10% of the solution. The pressure in passing through this 2 in. of electrolyte to the cathode is 0.25 volt. The tanks in series would therefore give a presure of 105 volts. For the cathodes, so-called 'stripping plates' are prepared as follows: Certain tanks are reserved for this duty and are provided with insoluble anodes of lead plates. Opposite the anodes are inserted greased copper plates upon which a layer of copper precipitates. When the layer becomes 1 / 32 in. thick the plates are taken out and the sheets of copper are stripped from the surface, the grease upon which prevents adhesion. Some of the sheets are cut into strips, and made into loops, and are riveted to the plates that are suspended in the regular tanks. There are 23 plates for each tank and these form the beginning of the cathodes. One ampere deposits one ounce avoirdupois in 24 hours. Each cathode, 2 by 3 ft., increases in weight 11 Ib., and all the plates of a tank 248 Ib. in 24 hours. The anodes correspondingly decrease in weight and become thinner by the same amount. At the end of two weeks the cathode, now weighing 160 Ib., is removed, and another starting-sheet is put in the place to receive all the remaining copper that can be taken from the anode. Not all the copper of the anode is dissolved. The part of the anode above the solution, and at least the skeleton of the plate, must remain to transmit the current. Shreds and pieces of copper drop to the bottom of the tank, and these with the fragmentary anode, still retaining 15% of the original weight, are removed, re-melted, and cast into new anodes. Current density. The greater the density (in amperes) of the electric current through the anodes, the more rapid is the deposition, and hence the smaller is the stock of metal that needs to be carried for a given output. At Great Falls, Montana, the current density is 40 amp. per sq. ft., while at Perth Amboy, N. J., the density is 15 to 17 amp. At high density there is a liability of short-circuiting, and the cathode-plates take on more impurity such as arsenic and antimony. It commonly takes a month to dissolve anodes. Hence in a large refinery, a half-million to a million dollars is tied up in copper and precious metals undergoing treatment. Treatment of the slime or anode-mud. Blister-copper, made into anodes, is commonly 97 to 98% copper. The remaining 2 to 3% is largely substance insoluble in the electrolyte. The insoluble 436 THE METALLURGY substance falls to the bottom of the tank, forming a thin mud, valuable for the gold and silver contained. To remove the mud the tank is by-passed. Being cut out of the system, the clear electrolyte is decanted, and the anode mud washed into a tight tram-car placed below the outlet-plug of the tank. The mud contains fragments of copper of all sizes that have dropped from the anodes when nearly dissolved. They are removed by passing the mud through a 40- mesh screen into a settling tank. From here it is drawn to a pressure-tank, then to a filter-press, and is dried by steam to 2% moisture and broken up. The broken material is then ready for treatment in an English cupelling-furnace. It is charged into the furnace on a molten lead-bath into which it melts, the lead taking up the silver and gold. The impurity in the dried material, such as copper, arsenic, antimony, and tellurium, are partly volatized, partly absorbed by the litharge-slag that forms under the action of the air-blast used in this type of furnace. The lead having been finally cupelled, the dore silver is left and is cast into bars. The scrap copper, screened from the anode mud as above described, is re-melted and re-cast with other anode-scrap into anodes, all amounting to 15 to 20% of the copper treated. Another method of recovering the precious metals consists in digesting the anode mud in dilute sulphuric acid which dissolves and removes the copper. Purifying the electrolyte. The best proportion of acid and copper for the electrolyte is 10% H 2 SO 4 and 15% CuS0 4 (equivalent to 4% Cu). When the copper exceeds this quantity, the resistance .increases; hence copper is removed from the circulating electrolyte or solution if in excess, to bring the amount to the required pro- portion. The quantity of iron, arsenic, antimony, and tellurium gradually increases, and a time comes when the electrolyte becomes foul with them and the excess must be removed. Antimony can be kept low by the daily addition of a small amount of salt, which precipitates as an oxychloride. To purify the electrolyte the following method is used. A portion of the electrolyte is diverted in a constant flow to tanks reserved for the purpose of purification. These have insoluble sheet- lead anodes and copper cathodes. A strong current is used so that not only is copper deposited but also the impurity. The deposit collects loosely upon the copper plates and falls to the bottom of the tanks. Every two months the accumulated mud, containing 40 to 60% copper, is cleaned out and reduced in a reverberatory refining OF THE COMMON METALS. 437 furnace to form impure bars of copper. The purified electrolyte is returned to the main system. Circulation of the electrolyte. To avoid short-circuiting, and to increase the activity and regularity of deposition, the electrolyte is made to flow or circulate through the tanks, entering the top of each tank near the end, passing downward between the plates, and finally rising and flowing away through an overflow pipe at the other end. After the solution has flowed through two tanks in this way, it enters a launder that returns it to the collecting or sump-tank. Thus every pair of tanks has an independent circulation. The sump tank receives all the electrolyte. Here the electrolyte is heated by means of a steam-coil to 40C., the effect of the warming being to decrease the electrolyte resistance. It is then pumped up to a distributing or stock-tank, and thence once more enters the circulation. Testing the current. Besides the voltmeter and ammeter to be found at the switch-board in the power-house, it is customary to use a voltmeter for constantly testing the drop in potential between the anodes and cathodes. For this a forked rod is used, which touches the two plates and takes a small current through a portable voltmeter. A slight drop of pressure indicates short-circuiting. 160. COST OF ELECTROLYTIC REFINING. In the early days of refining, costs were high ($20 per ton), and refineries charged $40 per ton of blister-copper treated. At present, the cost of refining 98% copper anodes has been reduced to $4 to $5 per ton, and contracts have been closed at $7.50 per ton. To this cost is added the charge for re-melting, if the copper comes to the refinery in the form of ingots that need it. Anodes, wherever made, should be re-melted, since casting them directly from the converter leaves a rough and porous product that is undesirable for the elec- trolytic tank. The cost of re-melting is generally stated at $5 per ton. The economy now possible is the result of cheaper power and the use of casting and handling machinery, as well as improved operation in other respects. Capital in plant. Not only is capital invested in the buildings and the equipment of the plant, but it is required for : (1) The stock of anodes in process of treatment. (2) The stock of anodes awaiting treatment. (3) The copper constantly contained in electrolyte. (4) The copper needed for the heavy conductors transmitting the current. 438 THE METALLURGY The result of this large demand upon capital is to restrict the operation of plants to places near financial centers, like New York, where cheap money is available, the copper near the market, and the labor abundant. These considerations outweigh the advantages of having the plant near cheap water-power. 161. REFINING IMPURE SPELTER. In general the spelter, as made at the zinc furnace, is sufficiently pure and is used without refining. If, however, the ore contains lead, the spelter will also contain it, and this must be removed by refining. A small quantity of iron also remains in the spelter, and this must be removed. It is observed that the spelter first made and expelled by distillation is purer than the part obtained later, and often the first of the zinc is kept separate, molded into smaller plates, and is put on the market as a special brand. The principle of the method consists in re-melting the spelter in a reverberatory furnace with a reducing flame, and letting the molten bath stand until the metal separates into layers according to the specific gravity of the different metals, the lower part of the bath consisting of a leady zinc and the upper part of spelter nearly free from lead. The lower layer is then tapped, or removed otherwise. The separation or refining must be done at a temperature near the melting point of zinc since the higher the temperature the more persistently does the zinc retain lead. Under the most favorable circumstances the lead-content of the spelter is reduced to 1 to 1.25 per cent. To refine spelter, a furnace resembling that shown in Fig. 182 and 183 is used, but the fire box is in two parts and provision is made to charge the spelter close to the bridge. It holds 30 tons of spelter when full, and in it 10 tons can be refined in 24 hours. The metals separate into layers. At the bottom is the lead; the iron forms with the zinc and part of the lead a difficultly fusible alloy that floats on the lead, and uppermost is the stratum of pure zinc. By means of an iron rod inserted into the bath, layers are distinguished, the zinc being soft, the iron-lead-zinc alloy (called 'hard zinc') being mushy, and the molten lead at the bottom soft. The underlying lead is removed weekly. A cylinder or pipe closed at the lower end is sunk below the lead layer. The plug is then knocked out, and the lead, rising in the cylinder, is ladled into molds. The zinc of the top layer is ladled out daily into molds, and it retains 1 to 1.25% lead. The hard zinc layer is removed when opportunity offers. To do this the zinc is ladled out first, the lead OF THE COMMON METALS. 439 is next removed, and finally the mushy mass of ferruginous metal is removed with ladles perforated so that the lead drains off. This hard zinc is sold for the manufacture of Delta or Sterro-metal. Spelter produced in the United States is generally pure and needs no refining. American high-grade spelter contains only 0.01 to 0.02% lead, and 0.01 to 0.02% iron. American Western spelter ranges from 0.4% lead and 0.02% iron in the better brands to 1% lead and 0.05% iron in the poorer. 162. PARTING GOLD-SILVER BARS. The bars from reduction works contain gold and silver commonly alloyed with copper, but sometimes also zinc and lead. It is custom- ary to re-melt the bars and assay them, buying them on the result of the assay. The bars are parted in nitric or sulphuric acid. Sulphuric acid, being cheaper, is the acid commonly employed. Bars containing a large proportion of gold are inquartated by melting them with silver in order to decrease the proportion of gold to silver ratio of at least 2.5 to 1, otherwise the acid fails to attack the silver. In parting with sulphuric acid the copper should be less than 10%, but in nitric-acid parting more than 10% is allowed. To adjust this percentage, bars low in copper are melted with those high in that metal. Nitric-acid parting. This method, still practised at the United States Mint, Philadelphia, is an efficient way of parting, especially on a small scale. The bars having been melted and proportioned as above described, the molten metal is granulated by pouring it into a tank of water. The granulated metal is transferred to porcelain, glass, or platinum vessels, and treated with nitric acid, 1.20 sp. gr., until action ceases. The solution is allowed to settle, then is decanted, and fresh acid is added for the purpose of dissolving the remaining traces of silver. This is again decanted, and the gold residue is washed thoroughly with hot water. It is then ladled out, drained, dried, and mixed with a little flux, and melted in a graphite crucible. From the decanted silver solution the metal is precipitated by adding common salt. The silver chloride thus formed is washed, thoroughly granulated, zinc is added to reduce the silver to metal, and the zinc chloride resulting from the reaction is washed from the precipitated silver. The silver is pressed into cakes, melted, and cast into ingots for bar-silver. Sulphuric-acid parting. Since for this method of parting there 440 THE METALLURGY should be less than 10% copper present in the gold-silver alloy, and not less than 2g* parts silver to one of gold, the bars to be parted that exceed the required limit, are so selected and melted with others as to afford the required proportion. The melted metal is cast into flat ingots and parted in this form. The ingots are placed in a cast-iron kettle, covered with a sheet- iron hood that is connected with a chimney so that the acid fume from the kettle is carried away. Here they are treated with sulphuric acid of full strength (66Be). When action has ceased the solution is allowed to settle, after which the clear supernatant part is decanted, being drawn off by a lead-pipe siphon into a lead- lined precipitating tank. The residue in the kettle is treated six or seven times with fresh boiling acid. In this way the silver completely dissolves, the acid solution being removed after each treatment. The brown gold residue is finally boiled with water, being heated and agitated by live steam from a pipe inserted in, the water. In this way the gold is 'sweetened.' The residue is removed from the kettle, dried, melted in crucible with a little borax for flux, and cast into a bar of gold of 999 fine. The acid solution from the kettles, which flows to the precipi- tating tank, is diluted with water, and the silver is precipitated by hanging copper plates about one inch in thickness in the solution. The copper replaces the silver in the acid solution, which becomes blue in color. When precipitation is complete the clear solution is decanted, and the cement silver at the bottom of the tank is washed with hot water to remove the acid copper-solution. The 'cement' silver, or precipitated silver, is removed to a box, then pressed into cakes or cheeses in a hydraulic press. Thus compressed, it is ready for melting in plumbago crucibles after adding a little borax flux. In large establishments the silver is melted in a small reverberatory furnace where it can be conveniently fluxed, skimmed, and ladled into bars for the market. The bars weigh 35 Ib. or 500 oz. each. On refining, each bar is marked with a number and the exact weight, and fineness, and the name of the refinery that pro- duced it. 163. REFINING CAST-IRON TO MAKE WROUGHT-IRON AND STEEL. Because of the large amount of carbon in cast-iron, it is too weak and brittle for many engineering purposes. Three-foiirths of the pig-iron in the United States is made into steel or wrought-iron, about 3% being manufactured into wrought-iron. Wrought-iron is OF THE COMMON METALS. 441 used in preference to steel for certain purposes because of the welding quality, ductility, and toughness compared with bessemer or open- hearth steel, but for most engineering purposes steel, that now is as cheap as wrought-iron, is superior. 164. PUDDLING PIG-IRON TO MAKE WROUGHT-IRON. In this process the pig-iron is melted in a reverberatory furnace lined with iron ore, using an oxidizing flame. During the melting there is an elimination of silicon and manganese which are oxidized in part by the flame and in part by the lining that with the silicon produces a slag. After melting, the heat is reduced and reaction starts between the iron oxide of the slag and the silicon, carbon, phosphorus, and sulphur, of the bath, whereby the impurities become oxidized and absorbed in the slag. The removal of the impurities or metalloids leaves the metal in the state of wrought-iron, but it is so nearly infusible that the heat of the furnace fails to keep the charge molten, and the metal 'comes to nature' or becomes pasty. The puddler collects the iron in the furnace into several balls weighing 125 to 180 Ib. each, that are removed, dripping with slag, and carried to the jaws of a squeezer by means of which the slag is squeezed out of them. The squeezed balls are sent to the rolls and are rolled into bars. The puddled bar, called the 'muck-bar' is cut into lengths, and the pieces are made into bundles, half the bars being piled cross-wise. They are wired together and heated in a re-heating furnace to a welding heat, then rolled into bars of a smaller size than the first. The re-rolled material is known as 'merchant bar,' and the effect of the further treatment is to eject more slag and cause a cross- fibre structure in result of the position of the cross-piled bars. A sample of hand-puddled bar has been found to contain carbon 0.296%, silicon 0.12%, sulphur 0.134%, and phosphorus 0.139 per cent. 165. STEEL-MAKING BY THE ACID BESSEMER PROCESS. The pig-iron used in the bessemer process preferably contains 1% silicon and 0.5% manganese, but to make a salable steel, the phosphorus should be below 0.10% and the sulphur below 0.08%, since neither element is removed in the converter. If the silicon is above 1% the large quantity of slag produced carries away iron. If far below 1%, the charge does not blow hot. When manganese is high (1.5%), it makes the charge sloppy, the slag then being highly fluid and easily ejected during the blow. 442 THE METALLURGY The converters in a large plant are supplied from several blast- furnaces, and to insure a good average pig-metal, it is customary to collect the product of the several furnaces in a single tilting reverberatory furnace, called after the inventor the Jones mixer, capable of holding 300 tons of pig-metal. From the mixer it is drawn to the converters as needed, and a regular supply is thus insured. 166. THE ACID BESSEMER PROCESS. The conversion is done in an upright converter, lined as shown in Fig. 184, with a lining of silicious material held together with Fig. 184. SECTION OF BESSEMER CONVERTER IN UPRIGHT POSITION. fire-clay. Fig. 184 represents a converter in vertical section. It is 9 ft. diam. by 15 ft. 6 in. high, and is capable of treating a charge of 15 tons of pig-metal. It is swung on trunnions, through one of which the compressed air needed in operation enters to the tuyeres at the bottom. The tuyeres are 19 in number and each is provided with 12 holes % in. diam., or 228 holes in all. The slag made in a converter is high in silica, and has but little effect on the lining, so that the lining lasts several months. The mouth of the tuyere, however, comes in contact with the iron oxide formed during the OF THE COMMON METALS. 443 blow, and hence this part of the converter lasts only 20 to 25 hours. This bottom accordingly is made so that it can be replaced by another causing a delay of 20 minutes in changing. Operation of the converter. The hot converter, from which the metal of a blow has just been poured, is placed in a horizontal position and 15 tons pig-iron is poured into it by means of a ladle that is brought from the mixer. When the converter is in this position no metal can flow into the tuyeres and obstruct them. After the metal is poured in, the blast (or 'wind') is applied at the rate of 25,000 cu. ft. per min., the converter being at this time turned to the vertical position. The blast now blows in fine streams upward through 18 in. of molten metal. Active oxidation of the manganese and silicon results and in about four minutes they are oxidized by the oxygen of the air and have become slag. The carbon now begins to oxidize to CO, and this also streams upward through the metal and issues with the air from the mouth of the converter in a body of flame. After another six minutes the flame shortens or drops, and the operator, knowing that the carbon has been elimin- ated, turns the converter into horizontal position, the wind being at the same time shut off. In anticipation of this, a weighed quantity of spiegel iron or 'spiegel' has been tapped from the spiegel-cupola, where it is kept melted, into a ladle. The ladle is transferred by the traveling crane and poured into the converter. So great has been the heat evolved by the oxidization of the impurities of the pig during the ten minutes of the blow that the temperature is higher than at the start, and we have a white-hot liquid consisting of comparatively pure metal. Oxidation-products remain in the bath, and the carbon and manganese of the charge tend to reduce these, the unused carbon being in sufficient quantity to impart the desired strength to the steel. Silicon, which also is introduced, tends to dispose of gas contained in the metal. After the spiegel or 'recarburizer' has been added and the reactions have ended, the steel is poured from the converter into a ladle. The ladle, after a short interval, is carried to a position over the ingot molds into which the steel is to be teemed or poured. The teeming ladle is 'bottom-poured', that is, a tap-hole and plug are arranged in the bottom, so that when the ladle is brought over the ingot mold a stream of metal drops straight downward into it until it is filled; and so on, the molds are filled successively until the ladle has been emptied. The metal remains until solid, after which the molds are stripped, leaving the ingots standing. The ingot is picked up, and conveyed to a re-heating furnace, and finally sent to the rolls to 444 THE METALLURGY be formed into the shapes desired for market use. Fig. 185 illustrates graphically by curves the progress of the reactions, and the elimina- tion of impurities during the blow. From it we see the rate at which the easily oxidized manganese and silicon are burned and also the carbon, which is but little acted upon until these disappear, but which after they are gone oxidizes rapidly. The pig contains at the beginning 3.5% C, 1.0% Si, and 0.5% Mn, all being removed. The recarburizer adds to it, as Fig. 185 indicates 1% Mn, 0.7% C, and 0.15% Si. The manganese is added to take from the metal the oxygen absorbed during the blow; the carbon is to give the steel to II 12 13 Fig. 185. ACID BESSEMER BLOW, AMERICAN PRACTICE. the required strength and hardness, and the silicon to dispose of the gas contained in the bath. 167. THE BASIC OPEN-HEARTH PROCESS. During the past fifteen years the bessemer process has been gradually giving way to the basic open-hearth process, due to the OF THE COMMON METALS. 445 fact that low phosphorus ore is being exhausted. Phosphorus imparts the quality of brittleness to steel if present in excess of 0.1%, and, since no phosphorus is eliminated in the bessemer process or in smelting, the iron ore supplied must be low in phosphorus. Suitable ore is called 'bessemer ore,' and ore having phosphorus above the desired limit is called non-bessemer ore. The limit may be considered 0.085 Ib. phosphorus per 1000 Ib. of iron in the ore. It is claimed that for most purposes open-hearth steel is better than bessemer, but the latter gives the most satisfactory product for tin-plates, and is well suited to the manufacture of rails. An im- portant advantage in the basic open-hearth process is that it can be used for making steel from pig-iron and ore high in phosphorus. To make steel in this way, lime is added to produce a basic slag, Fig. 186. LONGITUDINAL SECTION AND ELEVATION OF OPEN-HEARTH FURNACE. the hearth is lined with basic material to withstand the action of the slag, and impure iron and scrap that contain phosphorus are used. To a limited extent, sulphur is removed by the operation. Fig. 186 is a sectional elevation, and Fig. 187 a plan of a basic open-hearth furnace, having a hearth of 30 ft. 6 in. by 14 ft. wide. With the furnace two pairs of regenerating chambers are connected, one pair taking the heat from the product of combustion or waste gas after it leaves the furnace, and the other pair pre-heating the producer-gas and air entering the furnace. Every twenty minutes the current is reversed, the escaping gas going to the other pair of regenerators, while the air and gas go through the regenerators that have just been heated by the escaping product of combustion. In this way, not only is heat utilized for pre-heating, but the gas and air, thus pre-heated, naturally give a high temperature in the furnace. The heat in fact is so increased by successive reversals, 446 THE METALLURGY that it could be made to melt the roof of the furnace itself, and indeed care must be taken that it does not do this. As shown in the plan, a set of valves is provided. Certain valves are shown to be open, and others are closed, to adjust the currents as desired. Air and gas enter at the chambers at the left and the waste gas escapes Main Gas Flue i 1 Casting Pit *o Fig. 187. SECTIONAL, PLAN OF OPEN-HEARTH FURNACE. through the other pair to the stack. The temperature of the checker work near the furnace becomes 1000C., and near the stack, 400C. Thus the air and gas entering the furnace become heated to 1000C. The highest temperature of the furnace is 1600 to 170()C. The hearth of the furnace is lined to the depth of 24 in., and OF THE COMMON METALS. 447 carries a bath of molten metal of that depth, its surface being even with the level of the side door. The lining is of calcined dolomite (CaOMgO) held together with 10% its weight of anhydrous tar. The tar burns to a strong coke, that firmly unites the mass into a hard substance. Pure magnesite (MgO) is more expensive than dolomite, but it lasts longer and sometimes is used. It is the basic lining that gives the basic open-hearth furnace the name. The roof and side, above the slag level, are made of silica brick on account of the infusible nature of the material, and it is customary to put in a layer of neutral material, chromite brick (FeOCr 2 O 3 ), o Q iO hour* Fig. 188. CHEMICAL CHANGES IN BASIC OPEN-HEARTH PRESSES. between the acid brick above and the basic lining below. A basic furnace lasts 350 hours (18 to 24 weeks) without radical repairs. Operation. In present practice the charge for a basic furnace consists of steel scrap (steel trimmed in the process of manufacture, old steel rails, and steel collected by junk dealers) ; of pig-iron containing less than 1% Si, more than 1% Mn, and up to 2% in P; of calcined limestone (quicklime) 8 to 30% of the charge; and of iron ore. As shown in Fig. 188, it takes 4 hours to melt a charge, and 6 additional hours to complete the manipulation, so that in 10 hours the charge is ready to draw. During the 3 to 4-hour melting period, the carbon, manganese, and silicon we can see are reduced. The 448 THE METALLURGY reactions are controlled by the melter, who sees that the carbon is eliminated last, and if it is oxidizing too fast he must 'pig up' the charge by the addition of pig-iron to increase the carbon. On the other hand, if phosphorus is oxidizing too fast, the oxidation of the carbon can be hastened by 'oreing down' (adding iron ore) to produce the following reaction : Fe 2 3 -f- 30 = 2Fe + 3CO If carbon is eliminated too soon, much iron becomes oxidized. With the oxidation of silicon and phosphorus to silica and phosphoric acid, acids form with lime and iron oxide a basic slag containing 10 to 20% -Si0 2 , 5 to 15% P, 45 to 55% CaO, and 10 to 25% Fe. The slag does not attack the basic-lined hearth, and retains the phosphorus and the sulphur, but the CaO must be as high as possible for this, and yet not so high as to render the slag infusible. After melting, active oxidation begins, and the bath boils by the escape of gas. Upon the completion of the operation the charge is ready for tapping into a 50-ton ladle, the metal filling the ladle and the light slag overflowing and being thus removed. If the slag remained, phosphorus would be reduced from it, upon addition of the recarburizer, and would again enter the steel. The recarburizer is now added to the metal in the ladle. It consists of charcoal or coke contained in a dozen paper bags, each holding half a bushel. These are tossed into the ladle, then ferro- manganese is added, and then a pound of aluminum. Half the fuel is burned, half is absorbed by the steel; and the ferro-manganese supplies the necessary manganese and silicon and a part of the carbon. The aluminum absorbs oxidation products from the metal. The ladle is now picked up by the 75-ton traveling crane, brought to the ingot molds, and turned into them. The further treatment of the steel has been described under the 'Bessemer Process.' The furnace is patched or repaired where needed with a mixture of dolomite or magnesite and tar, and is ready for a new charge. As shown in the diagram Fig. 188. the resultant steel contains 0.35% manganese, 0.12% carbon, and still retains 0.050% phosphorus and 0.020% sulphur. 168. THE BETTS PROCESS FOR THE ELECTROLYTIC RE- . FINING OF LEAD. The principle of this process depends upon the solubility of lead in an acid solution of lead fluosilicate, which is used as an elec- trolyte. The solution is formed by diluting hydrofluoric acid con- OF THE COMMON METALS. 449 taining 35% HF with an equal volume of water and saturating with powdered quartz according to the reaction : 2 + 6HF = H 2 SiF 6 + 2H 2 Cxfe, In the hydrofluosilicicTead is dissolved until the solution -is- 8% lead, after which there remains 11% H 2 SiF 6 in excess. The anodes are plates of the base-bullion to be refined, cast 2 in. thick, resembling ordinary copper anodes. The cathode-sheets that receive the deposited lead are 'stripping plates, ' obtained as in the case with copper cathodes. They are made by depositing lead upon steel cathode-plates, prepared for use by cleaning, coating with copper, lightly lead-plating them in the tanks, and greasing with paraffin. On them is deposited the lead, and when the coating is of the desired thickness the steel cathodes are removed from the bath, and the lead coating or sheets are stripped off for use as cathode. Another method consists in casting the cathodes in the form of thin sheets. The anodes and cathodes are placed l l /2 to 2 in. apart in the tank. As in copper refining, the anodes are in multiple, and the tanks in series. The fall of potential between anode and cathode is but 0.2 volt, and the current strength is 15 amp. per sq. ft. One ampere deposits 31/4 oz. lead per 24 hours following the ratio of the atomic weights of copper and lead which is 63 to 207. In the process the impurities remain as an adherent coating on the anode, and consist of the copper, bismuth, arsenic, gold, and silver. The zinc, iron, cobalt, and nickel dissolve in the electrolyte. As compared with ordinary refined lead, electrolyticalty-refined lead is pure, being practically free from bismuth, even when much is present in the base-bullion, and it is remembered that bismuth is harmful to ' corroding lead. ' The residue or anode-slime, averaging 8000 oz. or more of silver and gold per ton, is treated by boiling it with sulphuric acid, using a steam pipe inserted in the solution to boil and agitate it with free access of air. The washed residue is melted in a small basic- lined reverberatory-furnace, the copper is removed by using nitre as a flux, and the antimony by the addition of soda. The dore bars finally obtained are parted in the usual way with sulphuric acid. PART X. PLANT AND EQUIPMENT PART. X. PLANT AND EQUIPMENT. 169. PRIMARY PLANT AND EQUIPMENT. A mine has a practical value only when it has been developed sufficiently to show what the future has to offer in erecting a reduction works. Where the value of the ore justifies, the beginning of metallurgical operations is kept simple, relying upon the skill of the workman, and not using a complicated plant. After knowledge is obtained in this way, the introduction of an efficient plant may be undertaken. This can be carried out to the best advantage, at times, by the erection of a single unit, adding other units when the first has been brought to a practical success. 170. LOCATION OF WORKS. The location of a reduction works depends upon whether the ore of a single mine is to be treated or a custom-works is to be built. If a mining company builds a works to treat the ore of its own mine it is usual, in order to save the expense of trans- portation, to place the plant as near the mine as the securing of a suitable site and water-supply permits. In the case of a smelting works, where fuel and flux must be freighted, and where the product is to be shipped, then in addition to the above requirements, a site must be considered with reference to a railroad. A custom-works, which buys ores from different mines and localities wherever it can, requires a place as convenient as possible to the chief source of supply and to the coke, flux, etc., that it must use. For such a plant, a point should be chosen where several railroads give rise to competition in freight rates. Such a center supports a large population, and this affords an abundant supply of labor. The location of a works is also affected by the cost of freight on the ore and on the supplies such as fuel. It is also affected by the railroad and labor conditions, the local market, and the capital or credit for obtaining money at a low rate for operating. Thus iron and steel manufacture has centered about Pittsburg, Pennsylvania, because coke, coal, and natural gas is abundant, and because a good market is found there for the products. On the other hand the iron smelter at Pittsburg must pay for freight from mine 454 THE METALLURGY to furnace, $2.25 per ton, and must carry a large supply of ore to last through the winter months when navigation is closed. The United States Steel Corporation, the largest manufacturer of iron and steel in the world, is to erect a plant near the iron ranges at Duluth, Minnesota, for reasons shown below. Vessels carrying iron ore to Lake Erie ports can return with cargoes of coke or coal to supply the Duluth furnaces. They then have a local market, and it is not necessary to 'stock up' with a winters supply of fuel. Figuring roughly that 2Vi> tons coal, made into coke and into producer gas, is required to make a ton of steel, there is a slight advantage, as to fuel, in making coke in by-product ovens at Duluth, Minnesota, and using gas engines which utilize the blast-furnace gases to the best advantage. Nearly two tons of iron ore must be sent to Eastern furnaces to produce this one ton of steel. It is seen from the cost of producing zinc, that 3.5 tons of coal are needed per ton of ore. Thus it is cheaper to convey ore to fuel, than coal to the mine where ore is produced. Near Joplin, Missouri, there is ore and also fuel, we expect, therefore, to find the zinc smelting works working there to the best advantage. The region is made more favorable by the fact that natural gas is to be had there. With respect to silver-lead works using lead as a collector of other metals, the favored places have been found to be railroad centers, such as Denver, Pueblo, and Salt Lake. From 12 to 15% coke is used in the charge in such smelting, so that nearness to coal-fields is not the all-important condition. On the other hand, ores are available there in proportion favorable to combining profit- ably with one another. The lead of one ore and the iron of another being combined serve the requirements of smelting. In treating ore by milling and cyaniding, the amount of fuel and other supplies required is small, and hence the natural place for the work is near the mine that produces the ore, provided the extraction, or recovery of the precious metals, is high. When, however, the ore is refractory and the recovery is low, it pays to ship the ore to smelting works that guarantee a high* extraction. The site. The advantage in a side-hill (or terraced) site as related to the level or flat site has been much discussed. It has been claimed as an advantage, that the former permits the ore to advance by gravity from one operation to the next, and that, as it becomes reduced, there is no necessity to again elevate it. On the other hand the flat site has the following advantages: OF THE COMMON METALS. 455 (1) The first cost of the works is small, since grading and retaining walls, that would be needed on a side-hill site, are reduced to a minimum. (2) The arrangement can be more convenient, since there is no need, as in a side-hill plant, to place the different parts of the plant in definite order to obtain the required fall. When it is desired to expand the works, it is possible to extend in any direction. (3) Every square foot of ground can be made the equivalent to a lower or an upper terrace or can be left level. Hence the parts of the plant that must be far apart, on a terrace site, can be side by side on a level one. Ventilation is good and the plant is accessible for supervision. Of course on a level site one must use elevators or other hoisting appliances, but it is seldom that we find a terrace site where elevators are not used. Certain products often must be returned for re- treatment, and the cost of elevating is less than 0.5c. per ton. Iron and steel plants, tL^ largest in the world, are constructed on level ground. The ore is unloaded direct from vessels to the stock pile, using grab-buckets holding 5 to 10 tons each. If trans- ported in cars, the cars are loaded in a similar manner. The custom is to use hopper-bottom cars, from which the ore drops into the charg- ing bins, and thence by charge-cars is conveyed to the furnace-skip. By the skip it is hoisted 100 ft. to the furnace-top. Many recent silver-lead smelting plants occupy level sites, but the dumping- ground is at a lower level. For iron works little attention is paid to the location of the dump. There is no hesitation in sending the slag, if necessary, a mile away by locomotive to be dumped. Mill-sites. On the unclaimed mineral lands of the Western United States, title is secured from the general government for a mill-site for reduction works, five acres in extent, either in connection with a mining claim (on a theory that each mining lode is entitled to a mill-site) or as a site for an independent or custom reduction plant. A reduction company, operating a mill, must dispose of the tailing it produces, and of the water discharged, not encroaching upon the property of other people, and it is responsible for all damages. A company must not let tailing flow into a stream that, at a reasonable cost, can be impounded, nor run into waters where liable to interfere with navigation. The right or custom of dumping on the valueless land of lower mining claims is general, except that the practice must do no damage to the property of owners below. 456 THE METALLURGY A reduction company can take up lands for a ditch or flume from unappropriated public land, and the claim can not be interfered with by later locators; but the owner of such a ditch or flume is responsible for damage arising from breaks or overflows. The same rule holds with respect to roads and trails. In Colorado, mining claims are subject to the right-of-way to parties hauling ore over them, but in other States the location gives exclusive control, except that a water, electric, or railroad company can take it under the law of eminent domain by giving a fair compen- sation for it. The smoke from the smelting works, especially those treating sulphide ore in quantity, delivers into the atmosphere many tons of sulphur-fume daily, as well as fine flue-dust carried out of the stack by the draft. This diffuses through the atmosphere and is carried by the wind to trees and the crops of the land. If not diluted, it blights vegetation, and naturally the farmers organize to secure damage, or to close the works. The question of what to do to avoid the difficulties is a serious one, and today when pyrite smelting and extensive roasting of sulphide ores is in practice, the trouble can not be altogether overcome. Thus far the solution has consisted in locating the works in places where there is little vegetation to be damaged, or in discharging the fume into the atmosphere from high stacks. It may be said that the latter expedient lessens but does not altogether obviate the difficulties. The metal- lurgist must give serious consideration to the matter therefore, otherwise, after erecting and starting the operation of a plant, he may find that he is compelled to close it, to the ruin of the entire enterprise. 171. INSTALLATION OF PLANT AND EQUIPMENT. Preliminary to building a plant and operating a works, an investigation is made of the process, the requirements of the plant, and all limiting conditions. It includes, besides the general matters outlined above, the questions of supplies, markets, railroad facilities, freight rates, sufficient and suitable labor not liable to strikes, and reliable civil conditions unaffected by revolutions or oppression by the government under which the plant must operate. Next comes the organization of the operating company and financing of the enterprise, or obtaining capital to build and operate the plant until it pays the operating costs. Often the promoters, besides owning the mine for which the reduction works are built, have acquired the necessary real estate and OF THE COMMON METALS. 457 the rights that go with it. Provisions should be made for access by railroads, for the necessary trackage, and for the common roads to the plant. Not only must water and power be provided, but right- of-way for securing them. If fluxes are needed then the proper quarries or deposits must be found. Construction. Before beginning construction, plans should be fully worked out by competent engineers. Cost estimates are made in detail, good materials are accumulated, and the labor-force is properly organized. In the design of the plant provision for duplicate parts is made, so that in case of break-down no interruption of operation occurs. On beginning construction the hydraulic works, where needed, are put in under skilled supervision. This includes the building of dams, reservoirs, the water-power plant, and transmission line. For a long-distance power-transmission line there may have to be sub-stations and a distributing system. Money must be provided for the salaries of officers of the company that are to receive pay during the period of construction, and all money expended must be accounted for, and cost-records kept by a skilled accountant. The money needed for legal expenses, general expense, traveling expense, and all expenses incurred during con- struction must be included. Equipment. This includes the machinery, furnace-tools, and appliances used in operating, but excludes land, buildings, and track- age. Labor-saving machinery, when reliable, effects a saving in costs, but it is remembered that this saving must not sacrifice the efficiency of operation. The question 'how much' often arises, and we may even come to the conclusion that it is not desirable (considering the cost of installation) to put in the labor saving appliance. 172. HANDLING MATERIALS. The application of machinery to the handling of materials has of late years received much attention. It has been rapidly developed because of the resulting economy in labor. For handling on one level, 100 tons or less of material daily, especially where the ore is to be distributed to various places, one or two-wheeled buggies, or barrows, on a good floor, have been found to be economical, elastic, and low in first cost. For small quantities the metallurgist is not led into installing machinery, for he finds in practice that it effects no saving. For large quantities barrows or buggies may be used, or hand-propelled tram-cars, as in mining. For still larger quantities, power-propelled cars are 458 THE METALLURGY used, that can be handled also on up-grades and sent from level to level. The idea is well carried out at the Washoe plant of the Anaconda Copper Mining Co., where industrial locomotives move thousands of tons of material daily from level to level. Indeed the Fig. 189. VERTICAL BELT ELEVATOR. OF THE COMMON METALS. 459 plant is an admirable example of how a side-hill site becomes effec- tive, where locomotives can be used. Appliances for mechanical handling are divided into continuous machines, and appliances for intermittent handling, such as light railways and cranes. Continuous machines. These carry a distributed load, so that the sub-structure upon which they rest is light compared with one upon which the load is concentrated as in a car. They deliver material continuously, and no time is lost in loading and unloading. Intermittent conveying, on the contrary, if we increase the load of the skip or bucket, becomes slow and awkward, whereas in the continuous conveyor it is possible to increase the capacity by widen- ing the conveyor and providing the correspondingly increased feed. We divide continuous machines into elevators, conveyors, and conveyor-elevators. Elevators are used for vertical or nearly vertical lifting. The belt elevator, Fig. 189, is of this type, and consists of an endless Fig. 190. STEEL ELEVATOR BUCKET. belt having sheet-steel buckets, Fig. 190, attached by flat-headed elevator-bolts at 18-in. intervals. To allow for the stretching of the belt, the upper pulley shaft is carried in take-up boxes, by which the shaft can be raised. It is generally preferable, however, to use the take-up boxes for the lower pulley-shaft. The lower pulley is enclosed in a boot, the ore delivering into the buckets at the rising side, shown at the left. Ore, not caught by the buckets, falls into the boot and is there scooped out by buckets. The boot has a hinged drop-bottom, so that it can be cleaned out when desired. The drop-bottom is of particular advantage for elevators used in a sampling mill in cleaning up between samples. The speed of the belt is 275 to 300 ft. per min. to insure that the material delivers to the discharge spout by centrifugal action, as the buckets pass over the top pulley. At this speed a belt-elevator, having buckets 8 in. wide, delivers 7 tons per hour, and a 10-in. bucket-elevator, 18 tons per hour. The head and boot-pulleys are 30 in. diam. and, 460 THE METALLURGY as well as the belt, are made 1 to 2 in. wider than the buckets. The whole is inclosed in a wooden housing to prevent the escape of dust. Fig. 191 represents a single-strand endless-chain elevator. The chain is carried by head and foot sprocket-wheels with sprockets spaced to take links of the chain. In this case the 'take-ups' are Fig. 191. SINGLE-STRAND END- LESS-CHAIN ELEVATOR. Fig. 191. SINGLE-STRAND ENDLESS- CHAIN ELEVATOR. carried by the boot, As is the case with the belt-elevator, the velocity should be sufficient to insure an efficient discharge from the buckets ; but chain elevators, because of the numerous joints, are not well suited to run at a high velocity. Fig. 192 is a double-strand endless-chain elevator having two head-pulleys arranged so that the buckets discharge into a spout between them. In this way the elevator can be run at the low velocity suited to the type. To take up the slack, the shaft of the second head-pulley is carried by horizontally moving take-up boxes. OF THE COMMON METALS. 461 Either of the endles-chain elevators can be housed, in wood as in Fig. 189, or in a sheet-iron. Conveyors are used for the horizontal transfer of materials, and can be modified easily to carry up an incline. Of all conveyors, the belt-conveyor is most widely used. To give it capacity it is troughed by running on pulleys that raise the edges of the belt forming a shallow trough (See Fig. 193). The simplest form is an endless belt running over end-pulleys, the load being fed at one end, delivering into a chute or into a bin at the other. The conveyor carries a load not only on a level, but on as steep as 24 incline. Fig. 193. CONVEYING BELT WITH TRIPPER. The capacity is large and the conveyors are simple and durable. A 12-in. belt, traveling at the rate of 150 to 350 ft. per min., delivers 10 to 35 tons per hour. A 24-in. belt, traveling at the extreme velocity of 600 ft. per min., has a capacity of 250 tons per hour of crushed ore, and requires 6 hp. per 100-ft. length, the power needed varying with the length of the belt. When the ore ascends an incline we add the power for lifting the load. It is desired at times to deliver the ore into bins situated at different points along the belt. This is accomplished by using the movable tripper shown in Fig 193, which also shows the belt loaded with ore. To discharge the ore the belt goes around the upper pulley, as shown, then around a second one just below, and continues the course to the front end- pulley. The ore shoots from the belt into spouts, that deliver on either side of the track upon which the tripper moves. Indeed the tripper is sometimes made to travel continuously back and forth 462 THE METALLURGY from end to end of a long bin at the rate of 200 ft. per minute, evenly distributing the ore and bedding it. The bin when full supplies the furnace, and the ore is of an even constitution throughout. The worm or screw-conveyor. This is convenient for delivering crushed ore short distances, and it thoroughly mixes the ore con- veyed. Fig. 194 represents a screw-conveyor delivering ore from the trough along which it has been conveyed, into another at right angles. The ore drops from the first to the second, and is conveyed by the screw in the second, shown at the left. The bottom of the trough is lined with smooth sheet-steel bent half round to conform to the worm or screw. A screw conveyor is shown in Fig. Ill illustrating a dry-crushing silver-mill. The disadvantages of this Fig. 194. SCREW CONVEYER (QUARTER TURN). type of conveyor are, that much power is needed, and that the ore grinds on the conveyor resulting in wear. It is well adapted to carrying a moderate and continuous supply of ore short distances. Endless-chain conveyors, These are much used, since they con- vey ore not only on a level, but vertically if necessary. Being entirely of metal they successfully convey hot materials. Fig. 195 represents an endless-chain push-conveyor, consisting of a series of plates or 'flights' attached to a double endless-chain carried at each end by sprocket-wheels like the double endless-chain elevator Fig. 192. The ore, drawn from any desired storage-bin as shown in the figure, is pushed up an incline by the moving flights in a fixed steel-lined trough, and is taken by a double-strand endless- chain elevator to a floor above. If desired, slides may be provided in the bottom of the trough. When the slide is opened, the ore drops into the desired bin beneath. Sometimes in place of flights, a continuous series of buckets OF THE COMMON METALS. 463 or trays is used. These overlap so that the ore can not drop between them. They operate upon the principle of the Howden pig-casting machine, Fig. 126. Indeed the conveyors lend themselves to a great variety of applications, as the examination of a catalogue of elevating and conveying apparatus will show. The chief draw-back to them is that they have numerous joints to wear, and that the troughs, flights, or buckets are subjected to serious wear. They must run slower than the belt-conveyor. In Fig. 38 and 39, the Edwards roasting-furnace, we have an example of a swinging push-conveyor. The flights are bladed and so hinged from the vibrating carrying-beam as to swing over the ore Fig. 195. ENDLESS-CHAIN PUSH CONVEYOR. in the conveying trough on the backward motion, but to push the ore along when moving forward. As is seen, the bottom of the trough is provided with slides to deliver the ore where it is needed. Vibrating-trough conveyors. These are of a simple type, consist- ing of a sheet-steel flat-bottomed trough, 50 ft. long by 2 ft. wide, supported by spring legs, and receiving a throw of 1 in. from an eccentric making 300 rev. per min. The conveyors can be arranged to deliver into one another, in series, a distance of 500 ft. if desired. Owing to the inclination of the spring legs, the trough rises in the forward motion and drops in the backward motion, so that the material is propelled to the discharge end. The conveyors are simple, inexpensive, and need few repairs. A conveyor, having a trough 24 in. wide, has a capacity of 20 to 25 tons per hour. When provided with a double or false bottom, the upper one being a screen, the conveyor can be made an effective screening apparatus. 464 THE METALLURGY Hoists. Of these, the commonest about reduction works is the platform elevator, which takes buggies, wheelbarrows, or tram-cars from floor to floor. It may have a platform of a size (6 by 6 ft.) to receive two cars or wheelbarrows at a time, and it raises a one- ton load 60 ft. per min. They are often run in balance, but it is better to have two independent counterweighted platforms. Neces- sarily, time is lost in loading and unloading, so that the estimate of the capacity is 25 tons hourly. Fig. 122 shows an iron blast-furnace having a platform-hoist, and Fig. 121, one having a skip-hoist adapted to mechanical charging. The capacity of the skip is 2 to 5 tons of ore to half the quantity of coke, and they are run in balance. It takes 34 seconds actual time for raising, dumping, and returning the skip to pit ; but the total time including the waits is 4 minutes, this furnishing the supply to a furnace producing 350 to 500 tons of pig-iron daily from a total burden of 1150 to 1650 tons. Industrial railways and tram-tracks. We already have referred to the use of these in industrial work. Where an industrial locomotive can be used it is possible to convey on up-grades, taking advantage of side-hill grades, or of trestles. Trestles are used about level sites, and by means of them ore can be conveyed in hopper-bottom cars and dumped into storage or charge-pockets. Fig. 137 shows an electric power system with a motor for handling a 31 cu. ft. or 3!/2-ton slag-car, as used at the United Verde Smelting works, Jerome, Arizona. Grabs. In large establishments hoisting rigs are used that are provided with large clam-shell buckets or grabs. They take 5 to 10 tons of ore at a time, and are used for unloading vessels, and for transferring ore to stock-piles for storage, or to the furnace storage bins as desired. It is noticed that the movable frames or bridges are made heavy to carry the large loads safely. The traveling crane. Fig. 143 gives in elevation a traveling crane as commonly used. It is for handling ladles and converters as described under copper-converting and for bessemer and open- hearth practice. For handling materials inside a building it is coming into general use. The cranes are operated by electricity, and move in any direction, horizontally or vertically, over the floor of the building commanded by them, and they avoid obstacles on the floor. Provided with large mushroom-shaped electro-magnets, they are now used to unload pig-iron or handle steel sheets weighing a ton or more, and by using them no time is lost as in older methods, in passing chains around objects to be lifted. PART XI. COMMERCIAL PART XI. COMMERCIAL. 173. KINDS OF WORKS. There are two classes of metallurgical works; those constructed at mines, and custom-works. A mine-works gets the ore-supply from the mine of the company. In such case, it does not pay for the ore it receives, and the operation- cost, being only that of reducing the ore, is low. A custom-works, that includes the purchase of ores among the items of cost, has here a serious item of expense. This is especially applicable in the case of the treatment of precious-metal ores. The stock of ore that must be carried depends upon the distance from the mines, and upon the certainty of the supply. In the case of the Lake Superior iron-ore supply, for example, a stock must be accumulated by the beginning of the winter to supply the furnaces until the opening of navigation in the spring, a period of six months. The silver-lead and copper smelting-works of the Rocky Mountain States carry a supply to last two to six weeks. We have already alluded, under 'Electrolytic Refining,' to the capital locked up by the process. On the other hand, at a mine-works, the supply can be replenished as used, so that provision is needed only for one to two days' running, and often for but half-a-day's run. 174. ORGANIZATION OF A METALLURGICAL COMPANY. Metallurgical operations on a commercial scale require, gener- ally, the organization of a company, or if the company is already organized, the establishment of a department to provide the addi- tional function. Where a metallurgical company is to be organized, the promoters or organizers obtain a charter, or articles of incor- poration, from the State in which they desire to incorporate. They next hold a meeting at which they receive the property that is t< be taken over by the company, adopt a set of by-laws for the guidance of the company, and elect the directors that are to manage the affairs. The directors proceed to the election of the corporate officers of the company from their number. The officers, of a small 468 THE METALLURGY company, are the president, the vice-president, the secretary, and the treasurer. The directors may appoint from their number a managing director, or they appoint a manager from the outside to have charge of the affairs of the company. In outlining the organization of a company undertaking metallurgical works, the manager should be guided by the following rules : (1) He should see that a supreme authority is provided over all action to be taken, and should carefully and fully outline the authority and responsibility of each position, making the duties of each conform to the capability of the party holding it. To do this he must avoid making any person subordinate to two or more ; should place the authority and responsibility together; should distribute the work and the duties not to overburden nor to under- load; and should arrange the positions so that promotion can come from them. While the manager gives his chief attention to the commercial or business affairs of the company, he generally appoints a superintendent to attend to the technical affairs of the plant. The organization, under the charge of the manager, may include (1). the supply, (2) the operating, (3) the accounting, and (4) the selling department. (1) The supply department attends to the purchase, and delivery of ore, fuel, and flux, and to the care and issuing of chemical and general supplies. (2) The operating department has to do with all that pertains to the reduction or manufacture of the ore into metal (the winning of the metal from ore) or to refining metals to bring them into marketable form, and has control of the operating forces, consisting of the foremen (and men under them), the repair force (consisting of mechanics and their helpers who keep the plant in repair and put in the needed improvements), and the laboratory or assay- office force. (3) The accounting department attends to the accounting, pay- roll, cost-keeping, and the distribution of costs. (4) The selling department attends to the disposal and sale of the product of the works. By-products, in process of further treatment, are not here included. 175. THE PURCHASE OF ORES. Smelting works purchase ores of every kind according to the requirements, provided the ores are sufficiently valuable to pay for OF THE COMMON METALS. 469 treatment. They are purchased according to the content and by a pre-arranged schedule. 176. IRON ORES. These are purchased by guarantee on the part of the shipper that they will come up to a given standard that generally is based upon the percentage-content in natural condition, thus including the contained moisture. For Lake Superior ore the prices for 1908 at Lake Erie ports, per long ton (2240 Ib.) were: Old Range bessemer, 55% iron base $4.50 Old Range non-bessemer, 51.5% iron base. . . . 3.70 Mesabi bessemer, 55% iron base 4.25 Mesabi non-bessemer, 51.5% iron base 3.50 The Old Range ores come from the iron ranges on the south side of Lake Superior, and command a higher price because of the better mechanical condition. The Mesabi ores are soft, friable, and carry much fine, which makes flue-dust. When smelting such ore., in ratio of 85% soft to 15% hard ore, as much as 6% flue-dust or dirt is made. On a non-bessemer ore, as it varies from this, a premium of 8.349c. is paid for each per cent iron over the guarantee, and a penalty or deduction is made of 8.349c. down to 50% iron, of 12.523c. down to 49% iron, and a double penalty down to 48% iron. Below 48%, the^penalty becomes 18c. per unit. On bessemer ore, provision is made for a premium only in case the ore exceeds the guaranteed 55% iron. 177. ORES USED IN SILVER-LEAD SMELTING. A schedule for the purchase of silver-lead ores in Clear Creek and Gilpin counties, Colorado, is here given, dated Feb. 1, 1905, f. o. b. cars at Denver, based on the dry-weight of the ore. The ton used is the short ton of 2000 pounds. Dry tailing and concentrate. Gold, $19 per ounce, if 0.05 oz. or more per ton; silver, 95% of the New York quotation on the day of assay, if 1 oz. or more per ton. Copper, dry-assay (wet less 1.5 units), Per Unit. 5% or less $1.25 Over 5% and including 10% 1.50 Over 10% 1.75 10% silica basis, lOc. up; 5% zinc basis 30c. up. 470 THE METALLURGY Treatment Gross value of ore. per Ton. Not over $35 per ton ..................... $3.50 Over $35 and including $80 per ton ........ 4.00 Over $80 per ton ........................ 5.00 Upon lots containing less than 7 tons ....... 5.00 Dry-silicious and copper-bearing ore. Gold, silver, and copper are paid for as in the schedule for concentrate. Treatment charge $8 on a 40% silica basis, 5c. down and lOc. up to a maximum charge of $11 on ores not exceeding $25 gross value; 5% zinc limit 30c. up. Lead ore. Gold $19.50 per oz., silver 95% of the New York quotation on the day of assay, copper $1 per unit dry (wet less 1.5 units) when ore assays 2% or over wet; 10% zinc basis, 50c. up. NEUTRAL SCHEDULE. Lead Inclusive Per Cent. 5 to 10 Per Unit Cents. 25 Treatment Charge. $800 10 to 15 25 700 < 15 to 20 25 5.00 20 to 25 25 400 25 to 30 30 400 30 to 35 . . . 30 300 35 to 40 30 250 40 to 45 32 2.00 45 to 50 35 200 Over 50 . 40 2.00 FLAT SCHEDULE. Lead Inclusive Per Cent. 5 to 10 10 to 15 15 to 20 20 to 25 25 to 30 30 to 35 35 to 40 40 to 45 45 to 50 Over 50 . Per Unit Cents. 25 25 25 25 30 30 30 32 35 40 Treatment Charge. $12.00 10.50 8.50 6.50 6.00 4.50 3.00 2.00 2.00 2.00 OF THE COMMON METALS. 471 Neutral basis, lOc. up or down. The schedule most favorable for the shipper to be used. Oxidized irony ores. Gold and silver are valued as in the schedule for concentrate. Lead 25c. per unit for 5% or over. Treatment charge $2 on a neutral basis, lOc. per unit up. Lead concentrate. Gold $19 per oz. if 0.05 oz. or over per ton ; silver and copper as in lead ores. Silica basis 10%, lOc. up ; 5% zinc limit, 30c. up. Lead Inclusive Per Cent. 5 to 10 Per Unit Cents. 25 Treatment Charge. $4.75 10 to 15 24 4.00 15 to 20 30 3.50 20 to 25 32 3.25 25 to 30 , 35 3.25 Upon concentrate assaying over 30% lead apply either the neutral schedule or the flat schedule as under 'Lead Ores,' whichever favors the shipper. Copper ores were formerly purchased upon the result of the fire- assay, which experience shows to be about 1.5% low. It was considered that the fire-assay expressed what was to be obtained from the ore when treated on a full scale in the furnace. If analysis shows that the percentage of SiO 2 is equal to that of Fe, the iron present is considered able to flux the silica, and the ore is called self-fluxing and is said to be neutral, the base (iron) neutralizing the acid (silica). If more silica than iron is present then the difference in percentage is called the silica excess, and the reverse when iron is in excess. Zinc is detrimental to smelting, and if an ore has more than 5%, a charge is made against the ore or a penalty, as it is called, is exacted by the smelting company. Flat prices are paid for the lead, that is, they remain the same whether the price of lead rises or falls. Concentrate (the granular product made in dressing or concentrating ore) containing no lead but consisting of iron or copper sulphide, is called dry. The value is in the contained silver and gold. Lead concentrate is valued also for the contained lead. As an example of the use of the above schedule, let us take an ore containing SiO 2 , 14% ; Fe, 6% ; Zn, 11% ; Mn, 4% : S, 10% ; Pb, 21% ; with 60 oz. Ag and 0.2 oz. Au per ton. The ore is evidently a lead ore, and we will figure it on both the neutral and the flat schedule given under 'Lead Ore.' 472 THE METALLURGY We find a silica excess of 4% over the iron and manganese. The excess of zinc over 10% is one per cent. On the neutral schedule we have : Gold, 0.2 oz. at $19.50 $ 3.50 Silver, 60 oz. at 95% of 62c. (New York quotation) 35.34 Lead 21% (21 units) at 25c 5.25 Total metal value $44.49 Deducting treatment, $4 -f (4% Si0 2 excess at lOc.) + (1% Zn excess at 50c.) 4.90 Net returns to the shipper f. o. b. Denver $39.59 On the flat schedule we have : Total metal value as before $44.49 Deducting $6.50 treatment + 50c. zinc penalty 7.00 Net returns to shipper f. o. b. Denver $37.49 Therefore the neutral schedule is more favorable to the shipper and is the one used. 178. SCHEDULE FOR COPPER ORES. Low-grade copper sulphide. Ores figured f. o. b. Tacoma, Wash- ington, 1907, per short ton dry-weight. Gold $20 per ounce, silver 95% the New York quotation, day of assay, copper 3c. per pound less than New York quotation. Treatment Charge. When net value is less than $5 per ton $3.25 " " more than $5 per ton 3.45 While the above schedules give an idea of prices on which to base estimates of the value of ores, most buying is done by special arrangement, and time-contracts are made for the delivery of ore. The skilled buyer, well informed in the details of the business, is apt to have an advantage over the shipper, and can make profitable contracts for his company. Often there is but one buyer, who has much his own way, but the seller, who can show to the buyer that he is well posted, makes better terms. Other supplies. With the exception of ores, fuel, and fluxes, supplies are kept in a store-room, and issued by the supply depart- ment only on a written order from the foreman or other responsible party that needs them. In this way is known where the supplies are to be distributed on the cost sheets. An account is kept of all OF THE COMMON METALS. 473 supplies received and issued, so that from it can be learned how much and when to order to keep properly stocked. It is detrimental to the business to so run out of supplies as to cause delays. Much knowledge is required in the purchase of supplies, to buy when prices are low or on a rising market, but only in small quantity on a falling market, and to obtain the best discounts, but care is taken to avoid the purchase of inferior goods. 179. THE OPERATING DEPARTMENT. The choice of men to operate a plant is often the important condition to the success of the enterprise. The superintendent not only must be informed as to the actual technical operations, but he must know how to arrange and organize his forces. He should be able to handle men effectively, and must possess tact, discretion, and firmness. He must be strict and just, and able to encourage as well as to drive. He is often the metal- lurgist as well as the superintendent, and has direct management of the furnaces and the metallurgical machinery. When thing go wrong it is he that is called upon to correct them, at whatever hour of the day or night. If a furnace is in bad condition or a machine out of order, he is responsible. If all is going smoothly his duty may be light, but when troubles occur, or when the company is losing money, his work is hard. If he fails in adjusting difficulties, no excuse is accepted. He must succeed or must resign. Much of his success depends upon his subordinates, and first in importance among them we must place the foremen. In early times the success of the operation depended upon the skill of the workmen. Today, manual skill must be supplemented by intellectual skill to insure the certain and exact operation of the plant. At first, the puddler worked up his charge of iron by hard labor. Today he operates levers and watches the progress of the successive operations. In place of muscular effort the workman now must supply intelligent direction. The payment for the work of supervision and control, as for the office force, superintendent, assayer, chemist, and foremen, is made monthly. In certain cases an additional premium is paid. In plants where the exhaustion of a mine or of a mining district may cause the closing of a works, premiums can hardly be assured. The premium method, however, has been found to give good results, especially where rated on tonnage, or in the case of the superintendent, on the lessening of costs. There is a decided difference between the in- centive of a man who draws his salary in any case, and one who 474 THE METALLURGY knows that his compensation increases with his diligence and effort to practice economy. Engineering is the art of doing with one dollar what ordinary man may, after a fashion, do with two ; hence a competent man, who is a trained expert, is necessary to best results, and where the margin of profit is small, he is absolutely essential. When the price of the metal being produced is high, and the margin of profit is large, tonnage is more important than economy. The capacity and efficiency of a plant depends on the intelligence and reliability of the men in charge of the different departments, and they are chosen with these qualities in view. The low-grade labor is used where hard routine work comes in, the high where judgment is essential. While low-grade labor may be faithful, it is stupid and liable to blunders ; when trained, however, if faithful, it becomes reliable. The employment of cheap men for supervising costly machines is offset by the loss of time or by actual disaster. Qualities of a foreman. A foreman often is a man that has been advanced from a subordinate position in the works. He may be selected from among several applicants from the outside. In quizzing an applicant, note the names of his former employers, ask him if he 'gets on' well with his men, if he scolds them, if he has been threatened by them, or if he treats them with consideration. A foreman must not be thought to be a 'good fellow' (an 'easy fellow') by his men. He may, himself, sometimes lead off in the work and show his men how to do things, but in general he has enough to do in seeing that they are all busy. The foreman not only must watch the progress of the work, but plan ahead, to be sure that everything is provided and ready at hand as needed. In such work the foreman need not be always so strict a disciplinarian. On routine work, and in order to drive, especially when he has many men to handle, he must be just, having reserve of manner, but looking to the welfare of the men while at work. The chemist, or assayer. The chemist or the assayer is called upon for results regarding the operation of the plant, and these he must furnish with promptness and accuracy. Besides, in his routine work, his skill and value is found in investigation under the director and advice of the metallurgist or superintendent. The construction force. The various mechanics (blacksmith, carpenter, engineer, and machinist) not only have repairs, but also new construction to attend to, under the personal direction of the superintendent, who may, where the work requires, employ a drafts- OF THE COMMON METALS. 475 man or constructing engineer to attend to construction. It is a rule, in case of a break-down or other emergency, when the general foreman in charge of operating needs aid, that this work has pre- cedence, and other work must be dropped to expedite it. Skilled and unskilled labor. The ten-hour labor at the reduction works is largely unskilled or common labor. These laborers are called 'outside men' or 'roustabouts.' They do the work requiring the use of pick and shovel, such as unloading cars, handling the products of the works, and assisting in the construction or repairs. Skilled laborers working in shifts of eight to twelve hours, called also 'inside labor' receive higher pay per hour than common laborers. The pay varies according to the skill needed. These men are responsible for the successful performance of the duties given them, and they are expected to work until they are relieved by their partners on the following shift, or until the foreman provides some- one to fill their places. For keeping discipline, and to prevent slackness in the work, certain rules, the result of long experience, have been laid down for the guidance of the men. These are : The men must be promptly at work, and must work the full time, or (for inside hands) until relieved. For ten-hour men the working-day is from 7 a. m. to 12 m. and from 1 p. m. to 6 p. m. in summer, while in winter the noon-hour is shortened to 30 minutes, and the day ended at 5 :30 p. m. Inside men must be on hand the entire time of their shift, and must eat their luncheon when there is time while watching their work. Charge wheelers must keep up the supply, even when the furnace is running fast, but may rest at intervals, and are not called upon to do other work than sweeping up at their own places before going off shift. The inside man can leave when relieved by his partner, but must wait for his partner if delayed in arriving. If the latter fails to appear the foreman provides another man, who then holds the place, the absent man losing it unless he has a good excuse, or if sick, he is required to notify the foreman by message, who then provides a man in his place. When the first man desires to return to work he must notify the foreman one shift in advance, so that the substitute is not put out of a shift for which he has come prepared. When men are sick on shift, if not too seriously, they should be held if possible until the end of the shift. It is impressed upon them that it is detrimental to the work for them to leave, and that it is difficult at short notice to get someone to fill their place. 476 THE METALLURGY Men must obey orders, and flat disobedience is followed by discharge, irrespective of who the man may be. Otherwise discipline is weakened. Be strict but just. It helps in discipline to discharge the poorest man occasionally, and if much time elapses without this you may be sure that you are becoming less strict. Do not entrust men to do work without supervision and inspection. They do it wrong, or become careless when they realize that they are not watched. In certain respects metallurgical work differs from that of other industries. The work is carried on by a crew who work together, each man having a part of the operation to depend on him alone. All are directed by the foreman who sees to the regular operation of the mill or furnace. Thus we have about a silver-lead blast- furnace, the feeder and his helper, the weigher, and the wheelers or trammers who bring in the stock, all being men that work on the feed-floor. At the slag or lower-floor there is the furnace-man, the tapper, and the pot-pushers. The men on both floors are called inside men. They work in shifts of 8 to 12 hours, and one crew replaces another. The men are by no means paid an equal wage. The furnace-man or feeder, for example, is paid more than the others, and the pay varies with the skill and knowledge required. Provision is often made for the care of the men in case of sickness. A charge of one dollar per month holds against every man, and this entitles him to medical attendance and care at a hospital in case of accident. If a man has worked five days in any given month, the one dollar is deducted from his pay for hospital dues. While accidents occur, especial care is taken by the foreman, for the safety of his men. Men are apt to grow careless and must be warned, or if disobedient, discharged. Any carelessness on the part of the foreman renders the company liable for damages. The tendency at modern mills and metallurgical works is to reduce the amount of labor per ton of output, this being done by the use of mechanical appliances, so that the labor of the individual becomes less strenuous, less trying to the strength and endurance, and yet better paid. In result, the workman becomes more able and is more desirous of retaining his job, and his services are more reliable. A large mill is run with less labor, proportionally, than a small one. An easily-treated ore requires less labor, since there are few operations and the skill and care is less. Eight-hour shifts need one-half more men than twelve-hour shifts. Steam-power requires more labor than water-power, the latter OF THE COMMON METALS. 477 often requiring only an occasional adjustment of the supply-gate. In dealing with Mexican employees the American engineer must be scrupulously honest. Honesty, there as elsewhere, is expected of him, and no training nor circumstances alters the restraint he expected to put upon himself. The mining and metallurgical engineer is intrusted with the control of large enterprises and the care of precious metals, and his success depends upon his responsi- bility. In starting a new reduction-works, a list of the places and occupations of the men should be prepared, and the men that are engaged are quickly assigned. All men thus engaged should be questioned as to their qualifications, and given places that they can fill. When a new works is to be started, men skilled in the operation and duties apply at the plant and often are willing to accept common labor awaiting the starting of the works. In this way they may be retained until needed. Quality and cost of labor. American labor is characterized by intelligence, energy, and responsibility, as compared with Mexican and much foreign labor. Mexicans that have been trained to duties often make excellent inside-men. When well bossed or directed, they do well, though often, from being insufficiently fed, their physical endurance is low. They need watching to prevent them from idling their time, and they should be carefully directed, since they are apt to work at disadvantage on account of their lack of thought. Treat them like grown children, have patience with them, but be judicious and strict. An employer of Mexican labor may find a man that excels in his work. There is a danger that he may be spoiled, by raising his wages. In directing Mexican labor, the familiar style, not the formal or polite one of the books, is to be used. Mexican, and in a less degree, other labor is disposed to pilfer. To such an extent is this true that in Mexico a tool or other portable article can hardly be laid down, without danger of it being stolen. Men must be held responsible for the tools that are served to them, returning them at night to be locked up. It is also well to brand tools. In the spring, when prospectors everywhere are start- ing into the hills for their summer occupation, the winter's job is left at the reduction-works, and tools needed in prospecting may be secretly taken. In determining the rate of wages in a new mining district in Mexico, it is well to start with low pay, as the people have been accustomed to it. It is easy to raise wages later if advisable, but difficult to reduce them when a certain rate has been paid. Among 478 THE METALLURGY Mexicans, whose tastes are simple, and wants easily supplied, the result of paying high wages is to give more 'laying-off' time, and this results in inconvenience at the works. Monthly payments interfere with the steady operation of the works, because at the time of the monthly payment the men are disposed to 'lay-off' to spend their money. To overcome the difficulty two methods have been tried. One is that of daily payments, by which a man, who spends his money when he gets it, has only sufficient to supply his daily wants and those of his family, and none remaining for drunkenness or gambling. The other system is to pay a wage to which is added a premium that increases with the time worked. It is paid at the end of the month if the man works through the month, but other- wise not. This tends to keep him steadily at work. 180. DUTIES AT A LARGE GOLD STAMP-MILL. The men needed will be : Foreman, night foreman, pipe-fitter, two engineers, mill- wright, two firemen, one head amalgamator, four amalgamators, two oilers, two feeders, two laborers. The foreman has general supervision of the mill, and looks after the handling, cleaning, and retorting of all amalgam collected. The amalgamators dress the chuck-blocks and plates, and keep them in good condition. They set tappets, regulate the water supply, and make renewals. The feeders attend to the uniform feeding of the batteries, and assist the amalgamators in renewals and at the clean-up. A good feeder is a valuable man about a mill. The vanner- men attend to the vanners, or concentrating tables. They must be men with experience, and commonly should first serve at the vanner as 'sulphide-pullers.' The crusher-men feed the crushers with the mine-ore as it comes to the mill. Oilers oil the machinery. Sulphide- pullers remove the concentrate or sulphide from the vanner boxes, and store it for shipment. Engineers run the power-plant, and have charge of the firemen. Firemen fire the boilers and remove the ashes. Coal passers wheel in coal from the coal pile to the boilers. On repairs there are carpenters, with laborers to help them. In repairs on the vanners there is. a special vanner-man to assist. 181. THE ACCOUNTING DEPARTMENT. This department takes charge of the accounts and transactions, makes up the pay-roll, and attends to the payment of the men. It has also to collect and collate costs. The costs are prepared for three purposes, as follows : OF THE COMMON METALS. 479 (a) To enable the owner of the property to judge as to the value and the efficiency of management. (b) To add to the economy of management by showing where economies can be practised and leaks stopped. (c) To prevent dishonesty. Costs are divided into flat, or prime-costs, and general expense, sometimes called fixed charges. The flat-cost varies directly almost with the tonnage, while general expense remains almost constant. Cost may vary as follows : Flat-costs (per ton of ore put through) diminishes as the output increases. The percentage of labor-cost to general expense is at a minimum in years of average activity. It increases in prosperous times because tonnage is more important than close saving, per ton, and it increases in dull years because general expense must be divided by a smaller tonnage. The flat-costs include labor, fuel, operating, and supplies. General expense includes cost of management and superin- tendence, office force, insurance, indemnities and damages, hospital and medical expenses, taxes and rates, and costs of selling the product. In connection with the matter of costs there are two further considerations which have a bearing on the returns of the capital invested. These are depreciation of plant, and interest on the invested capital. A plant will depreciate at the rate of 10 to 15% annually, so that even if kept in repair, it will have become obsolete and worn out at the end of ten years. If, however, a company sets aside from the earnings an amount sufficient to put in the needed improvements, and so keeps the plant up to date, then depreciation, except for the buildings, may be disregarded. In case this is not done, then dividends must be sufficient, not only to pay interest on the capital invested, and a profit besides because of the risk involved, but also to meet depreciation. Otherwise the investor is not recovering his capital and the compensation or interest due. The following are costs often not considered: Expressage on gold or silver bars to the mint or market. Costs of selling, generally a percentage on the gross amount of the sale or the metal costs. Interest on bonds and a sinking fund to meet the payment of the principal when these mature an expense that must first come out of the profits before the stock-holders are paid a dividend. Home- office expense due to the fact that the office may be in a large city where the stock-holders live, distant from the works. The object of such an office is that the corporate officers and stock-holders 480 THE METALLURGY may be in touch with the property in which they have invested. Royalty, when a percentage must be paid for the use of a process. 182. LABOR COSTS. In a 40-stamp silver amalgamation mill having 24 pans and a capacity of 150 tons of ore per 24 hours the inside labor for that time was: Four pan-men (12 hours) at $4 $16 Two helpers (12 hours) at $3 6 Ten tank-men (12 hours) at $3 30 $52 The quoted prices prevail in California and in other high-price camps in the Rocky Mountain region. Men needed and the cost of labor for a single silver-lead blast- furnace plant for 24 hours, having two 12-hour shifts are as follows : On feed floor : Two feeders at $2.50 $ 5.00 Two feeder's helpers at $1.80 3.60 Six charge-wheelers at $2.40 10.80 Two weighers at $2.40 4.80 On slag floor : Two furnace-men at $2.50 5.00 Two tappers at $1.80 3.60 Four pot-pushers at $1.80 7.20 Two engine-runners at $3.50 7.00 Two foremen at $4.25 8.50 Ten-hour men : Two dump-men at $1.50 3.00 Four sampling mill-men at $1.50 6.00 One sampling foreman 2.50 Mechanics : One blacksmith 3.00 One carpenter 3.50 $73.50 The above is based on a wage of $1.50 per day for common labor, and for a minimum force, that must be increased when putting in improvements, or for emergencies. If they can be profitably em- ployed, it is well to have men to supply vacancies caused by sickness, accidents, or men leaving. The following is the labor cost at the Treadwell stamp-mill, (240 stamps) Douglas Island, Alaska. OP THE COMMON METALS. 481 Inside men : One foreman at $150 per month $ 5.00 Four amalgamators (12 hr.) at $90 per month. . . . 12.00 Eight feeders (12 hr.) at $70 per month 18.64 Four vanner-men (12 hr.) at $65 per month 4.34 Two sulphide-pullers (10 hr.) at $2.00 4.00 Two sulphide-shovellers (10 hr.) at $2.00 4.00 Two engineers (12 hr.) at $2.50 5.00 Two foremen (12 hr.) at $2.50 5.00 Two coal passers (10 hr.) at $2.00 4.00 Four crusher-men (10 hr.) at $2.25 9.00 $79.66 Repairs : One carpenter (10 hr.) at $4 $ 4.00 One vanner-man (12 hr.) at $1 per month 3.34 One laborer (10 hr.) at $2 2.00 $ 9.34 $90.00 To these labor costs is added, for board and lodging furnished by the company, an actual cost of at least 50c. per man daily. 183. PROFITS. Profits from the operation of a plant, whether independent or connected with a mine, may be defined as the difference between the total gross costs and the returns on the metal-product sales. The profits may be increased either by better extraction (recovery) from the ore, by economy of treatment due to methods permitting a saving in labor, supplies, or fuel, and by faster running, by which the output is increased. The profit of a custom works, which buys ores outright, is the difference between the charge made for treatment and the actual cost of treatment. In the early days of smelting at Leadville, Colorado, charges of $60 per ton were made, the actual cost being $20, and the profit $40. Today this margin is as low as $1 to $2 per ton because of competition. Another profit is made on extraction. The works makes a deduction for losses in extraction. If they extract more than this, the difference is a clear gain to them. Thus they pay for 95% of the silver in the ore, but if they recover as they generally do, 98%, then the 3% difference is a profit. We will take, as an example, the profits that may be expected from the following ore : 482 THE METALLURGY Treatment charge per ton $6.25 Actual cost " " 4.50 $1.75 Gain in extraction of silver $0.25 " " " " gold 0.10 " " " " lead 0.35 0.70 This shows a total profit per ton, from both resources. .$2.45 In milling, this calculation remains the same whether the ore is highly silicious or not, but in smelting the silicious ore would make more slag that would contain more of the metal, and this would diminish the gain in extraction and perhaps convert it into loss. The following figures represent the profits of a company owning a mine, the Robinson company, on the Rand, South Africa : Gold recovered at the stamps $20.50 Gold recovered by cyaniding 5.70 Total recovery $26.50 Cost of mining. $6.55 Cost of milling 0.98 Cost of cyaniding 0.97 $8.50 Net profits per ton . $17.70 184. THE SELLING DEPARTMENT. Selling the product of the works is not difficult. There is a ready market for the commercial metal, but advantage can be taken of the market, and there are companies, aside from the producing company, that can be employed to perform this. 185. FUEL AND METAL MARKET. In examining market quotations of metals we must understand clearly the kind of ton meant. The short ton (2000 Ib.) is used in the Western United States for ore and metal. For coal, iron ore, pig iron, and steel the long ton (2240 Ib.) is understood, particularly in the Eastern States in the wholesale trade. In England, copper, tin, and spelter are weighed and sold by this ton. On the continent of Europe, in Mexico, and, in fact, wherever the metric system is in use, the metric ton of 2204 Ib. is taken, and this approximates to the long ton. In the United States, New York is the chief market for metals, OF THE COMMON METALS. 483 and the sale of ore and of metal is based on the prices there. The quotations of other markets, as San Francisco, St. Louis, and London, are also often given. For coal, coke, iron, and steel other centers are understood. Referring to a technical publication, as for example the Mining and Scientific Press, of San Francisco, or The En- gineering & Mining Journal of New York, we find the quotations published. Taking as an example the quotations of September 1908, the necessary explanations are as follows : Anthracite coal, New York market. Schedule prices are $4.75 for broken and $5 for egg, stove, and chestnut. Steam-size prices are unchanged, pea $3.25@3.50, buckwheat No. 1, $2.35@2.50, buck- wheat No. 2 or rice, $1.60@2, barley, $1.35@1.50. All prices are per long ton f. o. b. (free on board) New York harbor points. Run of mine anthracite coal is sent to breakers, large buildings containing the necessary crushing machines and revolving screens, to sort or screen the coal to the sizes above specified. The smaller sizes, called 'steam sizes', are cheaper, and are utilized in burning under steam boilers. Bituminous coal, New York market. Good grades of steam-coal are quoted at $2.50@2.65, with the poorer qualities selling around $2.40. At the mine, bituminous coal is quoted, f. o. b. cars, for run- of-mine, 85c., %-in. gas-coal, 90c., and slack at 50e. The %-in gas- coal is obtained, by screening the slack through a bar-screen having %-in. spaces. The coal that drops through the 'spaces is called slack. The difference between the prices at the mine and at New York represents chiefly the cost of freight to New York. Coke, Pittsburg 1 market. At the coke ovens the prices f. o. b. are for furnace coke, $1.65@1.85, and for foundry coke, $2.10@2.25. To be well suited for cupola use coke should be firm and in larger pieces than otherwise is required. Iron ore, market at Lake Erie ports, as Loraine, Cleveland, Con- neaut, and Ashtabula. Old Range bessemer, $4.50, non-bessemer, $3.75, Mesabi bessemer, $4.25, non-bessemer, $3.50. The difference in price between the Old Range and the Mesabi ore is because the latter is soft and contains much fine, and thus is less acceptable for blast-furnace work. Bessemer ore is low in phosphorus, not to exceed 0.045% phosphorus to 55% iron. Pig-iron, Pittsburg market (Quotations for carload lots). Stand- ard bessemer, $15 ; malleable bessemer, $14.50 ; basic, $14.25 ; No. 2 foundry, $14.50; gray forge, $13.50. Standard bessemer is used for making steel in the bessemer converter, malleable bessemer for malle- able iron castings, basic for steel suited to the basic open-hearth furn- 484 THE ace, foundry for making foundry castings suited to machining, and gray forge for wrought-iron. In the Chicago markets, both Southern and Northern pig-iron are quoted. The first, from the great iron center at Birmingham, Alabama, though cheap, is high in phos- phorus. The Northern iron from nearby points, is made from Lake Superior ores. Pig-iron, cast in sand, is weighed to 2260 Ib. for a long ton, the 20 Ib. excess being an allowance for the sand that sticks to the pigs. Steel, Pittsburg market. Bessemer and open-hearth billets are quoted at $25. These are ingots 4 in. square by 6 ft. long that are re-heated and rolled into the required merchant-steel bars. Mer- chant-steel bars remain at 1.40 to 1.60c. per pound. This is equal to $31.36 to $35.84 per long ton. Structural steel includes angles, channels, and I-beams, and is quoted much as are merchant-steel bars. Steel sheets are quoted at 2.50c. for black and 3.35c. per pound galvanized, No. 28 gauge. The 'black' means that the sheets are rolled, but not galvanized. The gauge referred to is the No. 28 wire-gauge, and is the basis from which thickness is reckoned accord- ing to a fixed scale. Steel railroad rails are held at $27. Scrap or old steel material, held at $10 to $12 per ton, is added to the charge in basic open-hearth work. Silver. At New York the quotations are on silver bars, per troy ounce of silver, 1000 fine. It takes 14.58 troy ounces to make 1 Ib. avoirdupois. London prices are for sterling silver, 925 fine. The value of the pound sterling is also given, so that with the London quotation, we may compute the equivalent price in cents there. Let us say that sterling exchange is $4.86, and that silver is selling at 25l per sterling ounce, we have then : i^S^S" == 54 - 8c - P er ounce of fine silver Copper. The New York price is expressed in cents per pound, quotations being given for Lake copper cast in the form of cakes for rolling into sheets, or ingots for re-melting to make castings, brass, and bronze, or wire-bars for drawing into wire. A sample- quotation is as follows: "The market closes steady at 13%c. for Lake, 13 1 /4@13%c. for electrolytic." (Here is noticed a difference of % to yQ. per Ib. in favor of Lake copper). "Casting-copper has averaged 13@13%c. during the week." Casting-copper is not as pure as that which is to be rolled into sheets or drawn into wire. Electrolytic copper is made by re-melting cathodes (the product of electrolytic refining) into ingots, cakes, or wire-bars. Cathodes are held at %c. less than electrolytic copper, the difference paying for the OF THE COMMON METALS. 485 re-melting. In London copper is sold by the long ton in English money, and is of various brands. A sample set of prices is as follows : English tough copper, 63 10s. Best selected, 62 10s. @ 63 10s. Standard, 59 17s. 6d. for spot ; 60 13s. for 3 months. The last quotation has reference to whether the copper is for immediate delivery, or whether the customer will take it at the expiration of three months, in which time the reduction-works will have produced it. The making of best-selected copper is mentioned under 'the refining of copper'. Standard copper, formely called g. m. b. (good merchantable bars), is the grade upon which the others depend. Besides these brands we have : Strong sheets (rolled copper) 72 10s. India sheets (a rolled brass) 68 10s. Yellow metal (a grade of brass) 5%d. per Ib. It is the business of dealers, and others interested in copper, to keep statistics of the supply of available copper, which is called the 'visible supply'. When the visible supply is small the price naturally rises, and the reverse is true when it is large. Manufactured copper is quoted at 19c. per Ib. for cold-rolled, and 18c. for hot-rolled, while wire has a basis price of 15i4c. in carload lots at the mill. Tin. Like copper, tin is quoted for immediate or for future de- livery at a specified time. A sample quotation would be 30%c. for spot and 29% to 30c. for future (three months, for example). The price is the same whether the metal is from Burma, near the Straits of Malacca (Straits tin), from Bolivia, from Australia, or elsewhere. When sold, as in the London market, we quote 134 for spot, and 130 15s. for three months per long ton of 2240 pounds. Zinc is commercially called spelter. Quotations in the United States are given in cents per pound, thus : 4.60c., St. Louis ; 4.75 to 4.80c., New York. The London market is quoted at 19 15s. for good ordinaries (ordinary brands) and 20 for specials (the purer zinc). St. Louis is near the zinc-producing district of Kansas, Missouri, and Illinois, and hence has a lower price for spelter than New York. Prices for zinc ore are given per short ton at Joplin, Missouri, on a basis for ore assaying 60% zinc, and the price varies with the zinc content above or below this figure. It once was the custom to obtain the basis price per ton in dollars, by multiplying the price of spelter per pound in cents by 7.5, but this is only an approximation the 486 THE METALLURGY buyer often being willing to pay a high price to secure the ore when it is scarce. With zinc quoted at 5c. per pound, the price for zinc ore would be $37.50. Antimony. Sample quotations per pound are, 8 l / to 8%c. for Cookson's, 7% to 8%c. for Hallett's, and l l / 2 to 7%c. for ordinary brands. The brands named are those of well known smelters of antimony. Quicksilver. The New York price is $42.50 per flask (75 Ib. net) for large lots. San Francisco makes nominal prices per flask at $42 for domestic orders, and $40 for export. The London price is 8 5s. per flask. Formerly the weight of a flask was 86 1 /4 Ib., but this has been decreased to 75 pounds. Precious metals. Gold is sold to the mints at an unchanging price of $20.67 per troy ounce, 1000 fine. From this the mint makes a deduction of 2c. per oz. to cover the cost of melting and assaying. Platinum is a commercial metal, at present more valuable than gold. We quote $22.50 for hard platinum, $20 for ordinary, and $16 for scrap. Silver has already been discussed. INDEX Page. Accounting department ....... 478 Acid bessemer blow 444 Refractories 48 Agitation vats 211 Alkaline earths in slags 383 Amalgam safe 227 Amalgamating pan 222 Amalgamation 118 Breaking ore for 68 Of silver ores 218 American ore hearth 368 Analyses of coke 385 Of copper matte 315 Of dolomite 383 Of flue dust 391 Of fuels. .33, 34, 35, 36, 39, 43, 45 Of gold ores 137 Of iron slags 293 Of lead ores 361, 362 Of limestone 383 Of matte Ill, 315 Of pig-iron 294 Of refractories.. 50, 51, 54, 55 Of slags 382 Of silver precipitate 259 Anaconda, Montana, matte smelting 334 Annealing zinc retorts 409 Anode-mud 435 Anodes, sampling of 65 Anthracite 35 Market 483 Antimony in base bullion 384 Prices 486 Argentite 217 Arsenic in silver-lead smelting 384 Assayer, duties of 474 Augustin process 248 Automatic charging of blast- furnaces 276 Feeders 126 Sampling 59 Azurite 298 B. Bag house 392 Barrel chlorination 143 Base bullion 390 Bullion, sampling of . . . 64 Page. Metal leaching. 255 Metal ores 18 Bases in slags 382 Basic open-hearth process 444 Refractories 48 Bedding lead ores 362, 369 Beehive ovens 39 Belt elevators 458 Berthelot law 22 Bessemer converter 442 Iron ores 18 Process 441 Best-selected copper 332 Betts process 448 Black Pine, Nevada, treatment at 243 Blaisdell excavator 183 Biake crusher 69 Blast for copper 303 Roasting 112 Blast-furnace, breaking ore for 66 For silver-lead 371 Plant 275 Blende 399 Blister copper 332 Blowing engines 283, 285, 346 Blowing-in 286, 377 Bodie, California, treatment at 209 Bone-ash 48, 54 Bornite 297 Bosqui, F. L., acknowledg- ments to 13 Boss process 229. Bottoms, treatment of 333 Boulder county, Colorado, gold ores 137 Breaking ore 66 Brick kiln 49 Machine 53 Mold 52 Briquetting press 394 British thermal unit, definition 20 Bromo-cyanogen process 210 Brown agitator 265 Brown-horseshoe furnace 97 Brown-O'Harra furnace 97 Bruckner furnaces 97 Butters distributor 182 Filter J86 488 INDEX. Page. By-product charcoal 37 Ovens 39, 40 C. Calamine 399 Calciners 92 Calcining 331 Calculation for matte charge. . 324 Of charge in pyrite smelt- ing 322 Of charge in silver-lead smelting 385, 389 Of iron-furnace charge... 291 California gold ores 137 Stamp-mills 133 Callow cones. , 178 Calorie, definition 20 Calorimeters 22 Cams 124 Cananea ores 137 Canda cam 124 Canyon City, Colorado, lime- stone 383 Capacity of roasting furnaces 111 Of stamps 133 Carbonates of copper 298 Carbon-brick 48, 54 Cast-iron refining 439 Cast-steel ladle 348 Casting-lead 424 Machine 287 Pig-iron 287 Centrifugal pump 173 Cerargyrite 217 Chalcocite 297 Chalcopyrite 297 Charcoal 36 In silver-lead smelting... 385 Charge floor in lead smelting. . 378 For zinc smelting 407 Scoop for zinc 406 Sheet for iron furnace. . . . 292 Sheet for matte smelting " 315, 324 Sheet for pyrite smelting. 322 Sheet for silver-lead smelt- ing 387, 389 Charging cyanidation vats 164 Chemical reactions in blast- furnace . 288 Page. Chamist, duties of 474 Chemistry of blast-furnace... 381 Of cyanide process 154 Of roasting 79 Of roasting zinc ores 400 Chilean mill 207 Chimneys 28 Chlorination 36, 135, 139 Chloridizing ores 18 Roasting of silver ores 235 Chlorine generator 140 Chrome-iron 48, 51 Chrysocolla 298 Cinder from blast-furnace 286 Classification of cyanidation methods 156 Of gold ores 115 Classifiers 177, 189, 201 Clean-up at Maitland mill 203 In Cyanidation 170 In silver cyanidation 270 Pan 128 Coal 31, 34 Market 483 Coarse crushing 69 Coke 31, 38 In silver-lead smelting. . . . 384 Market 483 Coking coal 34 Combination M. & M. Co., treat- ment at 243 Silver mill 234 Combustion 24 Commercial considerations. . . . 467 Composition of matte 315 Comstock ores, treatment of . . . 219 Concentrate, cyanidation of... 209 Definition 18 Purchase of 469 Roasting of 244 Sampling of 63 Condensers for zinc smelting. . 408 Coning 58 Construction of reverberatory furnaces 94 Work 457 Converter, bessemer 442 Plant for copper 342 Converting copper matte 336 INDEX. 489 Page. Conveyor belts 461 Copper 295 Converter 336 In cyanidation 208 In lead smelting 384 Ingots, sampling of 65 Ores, purchase of 469, 471, 472 Quotations 484 Refining 430 Sulphides, pot- roasting of. 113 Sulphides, roasting of 84 Cost at Lexington mill 242 Of chlorination 142, 150 Of converting 349 Of cyanidation.. 191, 193, 210 Of electrolytic refining. . . . 437 Of gold ore treatment 213 Of grinding 74 Of heap roasting 90 Of labor 477, 480 Of lixiviation 260 Of mechanical roasting... Ill Of refining base bullion. . . 430 Of reverberatory furnaces. 93 Of roasting in reverbera- tories 96 Of roasting stalls 90 Of Russell process 262 Of sampling 63 Of silver cyanidation 270 Of silver milling 229 Of smelting lead ores 395 Of stamp milling 134 Of zinc smelting 410 Cowper stoves 281 Cripple Creek ores 137 Ores, cyanidation of 205 Ores, treatment of 143 Crushing 66 For cyanidation 151 For patio process 246 Lead ores 364 Cupola furnace 26 Cupelling 427 Furnace 426 Cuprite 298 Current density in refining. . . . 435 Cusihuiriachic, Mexico, treat- ment at. . 262 Page. Cyanide, Colorado, treatment at 205 Cyanidation of gold ores.. 135, 151 Of silver ores 263 D. Dale, California, -process at... 208 Dead fluxes 386 Decantation 156, 180, 185 Diehl process 210 Direct process 333 Discharging cyanidation vats. 165 Discipline 475 Distillation, breaking ore for. 68 Of zinc 402 Dolomite 48, 54, 383 Double-discharge mortars.... 123 Treatment in cyanidation 153, 181 Draft in chimneys 28 Dressing plates 127 Drop of stamps 133 Dry-crushing, flow sheet for.. 73 Silver mill 238 Dry ores 18 Ores, purchase of 469 Silver milling 238 Drying cyanide precipitate... 173 Ducktown, Tennessee, smelt- ing practice 321 E. Edwards furnace 97, 101 Electric trolley slag pots 326 Electrolyte, circulation of 437 Purifying 436 Electrolytic copper refining. . . 434 Refining of lead 448 El Oro cyanidation method... 193 Elevator buckets 459 Enargite 298 Endless-chain conveyor 462 Endothermic reactions 21 English cupelling furnace.... 426 Equipment of plant 453 Extraction 19 And mesh in cyanidation. 194 By Russell process 262 In silver milling 235 Of copper 299 490 INDEX. Page. Of copper from copper sul- phides 350 Of silver 218 Exothermic reactions 21 F. Faber du Faur retort 425 Feeders 126 Filter presses 171 Pressing 172, 185 Pressing in cyanidation 156, 188 Fine crushing 72 Fineness of gold 486 Firebrick 51 Fire-clay 48, 51 Flash roasting 137 Flow-sheet in Maitland mill. . . 200 Of Augustin process 249 Of barrell chlorination. . . . 151 Of dry-crushing 73 Of lead refining 417 Of Russell process 261 Of Ziervogel process 250 Flue dust 391 Gases, temperatures 28 Fluorspar in slags 383 Fore-hearth 309 Foreman, qualities of 474 Fractional selection of samples 59 Free-milling ores 18 Fuel 31, 43 In silver-lead smelting... 384 Market 482 Furnace, cupola 26 English cupelling 426 Reverberatory matting.... 327 Wind 26 Zinc 404 Fuse box 93 Fusion of copper ore 331 G. Galena 361 Roasting of 112 Gangue, definition 18 Ganister 48, 49 Gates tube-mill 196 Genesis of fuels 32 Globe bag house 393 Plant . 370 Page. Gold bars, sampling of 64 Metallurgy of 115 Native 115 Silver, parting 439 Solution tank 168 Tellurides 115 Golden Cycle mill 207 Grabs 464 Grab samples 57 Grading ore 19 Pig-iron 294 Graphite 35, 48, 50 Grinding 66, 74 H. Hand reverberatories 92 Sampling 57 Handling materials 457 Harrisburg, Arizona, treatment at 210 Heap-roasting, breaking ore for 67, 84 Heat evolved in reactions 21 Of formation of compounds 23 Reactions in roasting 81 Hematite 273 Henderson process 356 Hersam, E. A., acknowledg- ments to 14 Heyl and Patterson casting- machine 287 Hoffman by-product oven 41 Hoists 464 Holthoff furnace 97 Homestake cyanidation meth- ods 188 Hot-blast stoves 282 Howard mixer 421 Howard press 422 Hunt ammonium process 208 Hunt and Douglas process 350, 353 Huntington-Heberlein process. 112 Hydraulic accumulator 347 Hydro-metallurgy of copper.. 349 Of gold 134 Of silver 248 Hyposulphite treatment of sil- ver ores 253 INDEX. 491 Page. I. Impurities in lead bullion.... 416 Industrial railways 464 Ingots, sampling of 64 Installation of plants 455 Iron in slags 382 Ore, market 483 Ores 273 Ores, purchase of 460 J. Jeffrey elevators 459 K. Kiln, brick 49 For making charcoal 37 Krupp ball-mills 137 L. Labor costs 480 Lake copper, refining 433 Lathe for cutting zinc 169 teaching base metals 255 Grinding ore for 68 In cyanidation 159, 165 In Ziervogel process 252 Ores 18 Silver ores 257 Lead bullion, refining 415 Copper matte 395 Electrolytic refining 448 Ores 361 Ores, purchase of 469 Leadville ores 361 Lexington mine, treatment at. 241 Lignite 33 Limestone 383 Lining converters 341 Lixiviation of silver ore 253 Location of works 453 Loomis-Pettibone gas-producer. 45 Loss in converting copper.... 349 In distilling zinc 409 Losses in roasting Ill Lump, ore, roasting of 84, 89 M. Machine sampling 59 Magnesite 48, 54 Magnetite 273 Page. Mahler calorimeter 22 Maitland cyanidation mill..... 199 Malachite 298 Manganese in slags 382 Market lead 420 Matte 314 Analyses of Ill Composition of 315 Concentration 324 Heap-roasting of 87 Lead copper 395 Smelting 302, 304 Smelting at Anaconda. . . . 334 "Smelting, charge sheet... 316 Treatment by Augustin process 249 Treatment by Ziervogel . process 250 Matting blast-furnace 303 McArtnur-Porrest process 135, 152 McDougall furnace ... 91, 97, 108 Mechanical roasters 97 Melaconite 298 Melting furnace for silver. . . . 228 In lead 378 Lake copper 433 Lead ores 364 Silver residues 245 Mercur, Utah, cyanidation method 156 Mercury, see quicksilver Merton furnace 97 Mesh and extraction in cyan- idation 194 Metal market 482 Metallic Extraction Co., treat- ment 205 Metallics, sampling 63 Metallurgical treatment of ores 19 Metallurgy of zinc 400 Metals, sampling of 64 Mill-sites 455 Missouri dolomite 383 Moisture sample 57 Molding market lead 424 Monadnock mill 207 Mount Morgan ores, treatment of 137 Muffle roasting furnace... 358 492 INDEX. Page. N. Native copper 298 Silver 217 Natural gas 31, 36 Solid fuels 33 Neill process 355 New York metal market 483 Nitric-acid parting 439 Non-bessemer iron ores 18 O. Open-hearth process 444 Operating department 473 Operation of bessemer conver- ter 443 Of blast-furnace 283 Of converters 338 Of copper furnace 86 Of lead smelting 368 Of reverberatories 94 Of stamp-battery 125 Ore-buying schedules 469 For matte furnace 312 Hearth 368 Ores, classification of -17 Definition of 17 Mixed 17 Of copper 297 Of gold 117 Of iron 273 Of lead 361 Of silver 217 Of zinc 399 Simple 17 Organization of metallurgical company 467 Oxidation of lead ores 366 Oxides of copper 298 Oxidizing roasting 79 P. Parkes process 420 Parting gold-silver 439 Patera process 253 Patio process 246 Pattison process 419 Pearce-turret furnace 97, 104, 137 Peat 33 Petroleum as fuel 31, 35 Page. Pig-iron, market 483 Sampling of 66 Smelting 274 Pipe sampler $6 Pittsburg markets 483 Plant and equipment 453 Plattner process 135, 139 Plumbago 50 Polybasite 217 Pot-roasting 112 Pound-calorie, definition 20 Precipitation after chlorination 148 After lixiviation 258 At Maitland mill 203 In cyanidation 169,191 Preparation of ores 19 Prices of pig-iron 294 Producer gas 4& Profits 481 Properties of zinc 399 Puddling 441 Pulp, sampling of 63 Pulverized ore, roasting of... 91 Punch sampling 65 Purchase of ores 468 Push conveyor 463 Pyrite matte smelting 318 Roasting 79 Smelting in two stages. .-. 321 Q. Quality of labor 477 Quartering 58 Quicksilver fed to battery 126 In silver milling 244 Prices 486 Trap 127 Quotations of metals 482 R. Rabbles in Edwards furnace.. 107 Rand cyanide practice 153, 156, 180 Raymond furnace 97 Reactions in basic open-hearth process 447 In bessemer converter... 444 In blast-furnace .... 283,288 In bromo-cyanidation . . . 213 In chlorination 141 493 353 257 380 247 320 Page. In copper converter 339 In cyanidation 154 In Hunt and Douglas pro- cess In leaching silver ores.. In lead smelting 366, In patio process In pyrite smelting In reverberatory matte smelting 330 In Rio Tinto process... 351 In roasting 8<< In roasting concentrate... 246 In roasting copper sul- phides 84 In roasting silver ores. . . . 236 In roasting zinc qres 400 In Russell process 262 In silver milling 224 In Ziervogel process 251 Receiving lead ores 362 Ore 56 Reduction of lead ores 366 Reese river process 238 Refining 413 Refractory materials 47 Re-lining converters 341 Replacing zinc retorts 407 Retorts for refining 425 For zinc smelting 405, 408 Retorting amalgam. 129, 22S, 245 Reverberatory lead smelting.. 365 Matte smelting. 326 Roasting furnalces 92 Smelting, breaking ore for 67 Revolving screen 74 Rio Tinto process 351 Page. S. Re-pressing machine. Roasting Blende Breaking ore for. 50 .... 77 400 67 Concentrate 245 For Ziervogel process..., 251 Furnaces for zinc ores. . . . 401 Gold ores 137 Silver ores 135 Rolls 72 Russell process 260 Sadtler process 411 Sampling 55 Base bullion 390 Lead ores 362 Works 61 Sand 48 In silver cyanidation 267 Treatment at Maitland mill 201 Schedule of copper ores 472 Scoop for zinc charging 406 Screen for stamp-mills 124 Revolving 74 Tests 69 Screw-conveyors 4C2 Segregation 19 Selling department 482 Settler for silver milling 225 Settling tanks 177 ^Siderite 273 Silica brick 48, 49 Silicates of copper 298 Silver 215 Bars, sampling of 64 Lead ores, purchase of... 469 Lead smelting 369 Milling 218 Quotations 484 Single-discharge mortars 122 Site for works 454 Skilled labor 475 Skimmer 423 Slag analysis 293,315,382 Floor in lead smelting... 379 From blast-furnace 286 From copper furnace 325 In silver-lead smelting... 381 Pot 310, 326 Slime in silver cyanidation... 268 Plants 192 Treatment at Maitland mill 202 Sliming, grinding for 74 Smelter plant for copper 342 Smelting copper ores 299 For pig-iron 274 Gold ores 213 Lead ores.. 365 494 INDEX. Page. Ores 18 Silver-lead ores 369 Zinc ores 402 Softening furnace 418 Lead bullion 41b Sombrerete, Mexico, treatment at 262 Specific heat 30 Speiss 384 Spelter refining 437 Sphalerite 399 Spitzkasten 177 Split shovel 59 Stacks 28 Stall-roasting, breaking ore for 67, 84 Stamp-mills 118 Men for 478 Stamper 406 Standard plant, Bodie, Cali- fornia 209 Starting blast-furnace 312 Steel cyanidation vats 163 Making 441 Market 484 Stephanite 217 Stetefeldt furnace 97 Stirring paddle 430 Suction filtration 185 Sulphatizing roasting of matte 251 Sulphide ores 18 Roasting 79 Sulphuric-acid parting 439 Supplies, purchase of 472 T. Tailing, leaching of Purchase of Sampling of Taylor gas-prouucer Tellurides of gold Temperature of combustion. . Testing current for refining. Tetrahedrite Thermo-chemistry Time in cyanidation Tin quotations Tram tracks Traveling cranes 159 469 63 44 115 29 437 298 20 167 185 464 464 Page. Treadwell stamp-mill, costs at 480 Trench sampling 58 Tripper 461 Trommel ?4f Tube-mills 75, 196 Tuyere 311 U. Unskilled labor. 475 V. Vacuum-filtration in cyanida- tion 156, 185 Vats for cyanidation 160 For hyposulphite leaching 256 Vezin sampler 60 Vibrating trough conveyor. . . . 463 W. Washoe process 219 Water-jackets 307, 375 Weighing ore 56 Weight of stamps 133 Welsh process 330 Western Australia ore treat- ment 210 Wet pan 345 Silver mill 220 Wethey furnace 97,100 White briquetting press. 206, 394 White-Howell roaster. 97, 137, 239 White metal 332 Wind furnace 26 Wood 32 Wooden cyanidation vats 161 Wrought-iron manufacture. . . 440 Z. Zince boxes 167 Furnace 404 In slags 383 Lathe 169 Metallurgy of 400 Ores 399 Quotations 485 Zone of fusion in blast-furnace 284 Of preparation in blast- furnace 284 Of reduction in blast-fur- nace . ... 284 YC 634 ^ THE UNIVERSITY OF CALIFORNIA LIBRARY