UNIVERSITY OF CALIFORNIA 
 LOS ANGELES 
 
 GIFT OF 
 
 Ann V* Adams
 
 MANUAL OF MINING 
 
 BASED ON THE COURSE OF LECTURES ON MINING 
 
 DELIVERED AT THE SCHOOL OF MINES 
 
 OF THE STATE OF COLORADO. 
 
 BY 
 
 M. C. IHLSENG, C.E., E.M., PH.D., 
 
 Formerly Dean of the School of Mines of the Pennsylvania State College? 
 
 AND 
 
 EUGENE B. WILSON, 
 
 Mining and Metallurgical Engineer* 
 
 FOURTH REVISED AND ENLARGED EDITION. 
 SECOND THOUSAND. 
 
 NEW YORK : 
 
 JOHN WILEY & SONS. 
 
 LONDON : CHAPMAN & HALL, LIMITED. 
 
 1907
 
 Copyright, 1892, 1898, 190$, 
 
 BY 
 
 M. C. IHLSENG. 
 
 ROBERT DRUMMOND, PRINTER, NBW YORK.
 
 J .'' Try 
 
 14 
 
 ]T a 5 
 t> 
 
 PREFACE TO THE FOURTH EDITION. 
 
 BASED originally upon lectures delivered at the Colorado 
 State School of Mines, between 1880 and 1888, the various 
 editions received only such additions and changes as were neces- 
 sary in its use as a text-book in the mining schools of this 
 and other countries. The Author of the MANUAL OF MINING 
 presents in this edition a complete revision of the work, which 
 now assumes the form of a collaboration with the authors men- 
 tioned below. 
 
 Mr. E. B. Wilson, who needs no introduction to the mining 
 public, brings to the original work a wide range of experience 
 and ability, an accuracy of detail and a discrimination in de- 
 scription, contributing highly to the value of the revision. He 
 has devoted himself particularly to Chapter III of Part I and 
 Chapters V, VI, VII, and VIII, of Part II. Originally planned 
 for Metal-mining, the first issue of the book gave small place 
 to Coal and its extraction. This new edition has of necessity 
 been enlarged to include Coal-mining in all its phases, with 
 full descriptions and many illustrations of modern methods and 
 machinery. The later devices in power generation and dis- 
 tribution, as the steam-turbines, oil, and compressed-air engines, 
 with such appliances as have proven themselves meritorious 
 in economy and safety, are elaborately treated. The chap-
 
 IV PREFACE TO THE FOURTH EDITION. 
 
 ters upon Electricity in its Application to the Mining Industry, 
 prepared by Mr. Rolin W. Hutchinson, Jr., whose work on " Long- 
 distance Transmission of Electricity" has secured recognition in 
 this field, will prove invaluable to the student. 
 
 The Author desires to acknowledge the courtesy of the follow- 
 ing parties for use of illustrations or tables as below: 
 
 Crocker- Wheeler Electric Company, Fig. 66. 
 Cyclone Manufacturing Company, Figs. 340, 341. 
 De La Valle Steam Turbine Co., Figs. 48, 49. 
 "The Engineer," Chicago, Figs. 39, 40, 45, 142, 143. 
 General Electric Company, Fig. 149. 
 
 Prof. F. R. Hutton, Mechanical Engineering of Power Plants, 
 Figs. 38, 41, 47. 
 
 Mr. H. C. Reagan, Locomotive Mechanism, Figs. 42, 46. 
 Mr. Frank Richards, Air Compression, Tables pp. 308, 309, 
 
 3H- 
 
 Scientific Publishing Company, New York City, Figs. 185, 
 186, 187. 
 
 Scientific Text-Book Company, Figs. 95, 104, 178, 188, 193. 
 
 Westinghouse Electric Company, Figs. 49, in. 
 
 Worthington Pump Company, Figs. 151, 152. 
 
 MAGNUS C. IHLSENG. 
 
 NEW YORK, November, 1905.
 
 PREFACE. 
 
 THIS treatise is an abbreviation of a course of lectures 
 upon mining, delivered at the School of Mines of the State of 
 Colorado, and is issued with the advice of its Board of Trustees, 
 which recognizes the importance of having, within a moderate 
 compass, the best information obtainable upon this subject. 
 In its presentation the writer has followed what his own ex- 
 perience has taught him to be the natural sequence, and has 
 endeavored to introduce such matter as sixteen years of lectur- 
 ing and field work have suggested as requisite. Part 1 contains 
 a brief geological review and a discussion of such points as the 
 engineer must include in his report, i.e., the preparatory and 
 development work, systems of mining and the plant for power, 
 hoisting, pumping, and ventilation. Part II embraces the 
 practice of prospecting, drilling, blasting, shafting, tunnelling, 
 and timbering, in addition to some remarks upon the examina- 
 tion of mines. 
 
 The work is designed as an elementary treatise for the use of 
 those desiring a reference-book. The complexity of the subject, 
 its extent, and the variety of machines to be described and repre- 
 sented, demand an elaborate discussion that would fill several 
 quartos. Descriptions of obsolete and expensive systems or 
 machinery are relegated to the historical works on mining. 
 American and foreign practice is described, and suggestions for 
 lines of future progress are offered herein. The principles of 
 the construction and operation of machines used in mining are
 
 vi PREFACE. 
 
 explained with a perspicuity and conciseness compatible with 
 the field in which this publication is to be sown among students 
 and mining men, to whom a knowledge of the fundamenta of 
 their work is valuable, but whose acquaintance with the theory 
 is slight. 
 
 The wants of the latter class have been kept in mind, and 
 the writer hopes that the manual may prove of some benefit 
 to the intelligent reader, of whom it presupposes an elementary 
 knowledge of the sciences and of the simple machines. 
 
 The author regrets his inability to deal with the subject of 
 "electricity in mining" as it deserves. Two reasons account 
 for this: insufficient data, as yet; and the large space which a 
 satisfactory explanation of the principles would demand. 
 
 The writer would also beg leave to say that the literature 
 of mining and its cognate branches has supplied much of the 
 material contained herein. References could not be made for 
 each hint obtained, but obligations are acknowledged to the 
 authors of the publications mentioned, to which the reader is 
 referred for further details. The information has been gar- 
 nered from the best available sources and condensed. The 
 Engineering and Mining Journal, the Colliery Engineer, and the 
 transactions of the American Institute of Mining Engineers have 
 been copiously drawn from, as also the experience of the prac- 
 tical men, to a long list of whom the Author is indebted for many 
 courtesies. Finally, to the manufacturers and engineers thanks 
 are rendered for the use of the electrotypes, which have so largely 
 contributed to make the work attractive. 
 
 MAGNUS C. IHLSENG. 
 
 GOLDEN, COLORADO, Nov., 1891.
 
 PREFACE TO THE THIRD EDITION 
 
 IN the present issue the text has been increased by more 
 than fifty pages in order to introduce additional matter per- 
 taining to the design of cars, hoisting appliances, and fans, the 
 added illustrations having been specially prepared with that 
 end in view. At the end of each chapter will be found a list 
 of memoirs which have been carefully selected to comprehend 
 the latest literature on the subject. The original intent to also 
 enter at some length upon a discussion of the application of 
 electricity to mining was abandoned after a number of fruitless 
 efforts to compress the matter into the small compass which it 
 should occupy in a work of this nature. 
 
 It is gratifying to observe the progress made in the increased 
 economy in the utilization of power. The growth of the young 
 giant among the motor fluids in the anthracite regions is par- 
 ticularly noticeable; while that of electricity in its various ap- 
 plications in bituminous mines is rapid and sure. It may be 
 true that the adoption of these motor agents by mining engineers 
 has been a little slow, but this wise conservatism is dictated by 
 the knowledge that a single accident resulting from the intro- 
 duction of an innovation might precipitate a peremptory legis- 
 lative prohibition which would render the outlay entirely useless, 
 and it is with reluctance that they must frequently forego the 
 advantages of some possible economic installation. 
 
 It is gratifying to observe the extended employment of arti- 
 ficial methods of ventilation in the metalliferous mines of the
 
 viii PREFACE TO THIRD EDITION. 
 
 West, and when the American method of "square sets" shall 
 have been supplanted entirely by the method of rock-filling, 
 or of flushing with waste, the dangers to life and property will 
 have been further reduced and the economy of mining materi- 
 ally increased. 
 
 Regarding the enlargement of the chapter on explosives 
 and the stress laid upon the necessity for an absolute prohibi- 
 tion of black powder from the coal-mines, I hope they may be 
 fruitful of results. 
 
 Finally, there has been added at the end of each chapter a 
 list of the more important memoirs dealing with the subject- 
 matter of that chapter, and this, carried to the close of the year 
 1897, has brought the work up to date. It should prove of 
 good service as a syllabus. 
 
 The author hopes that the book in its revised form will con- 
 tribute to a better circulation of the knowledge of the principles 
 upon which mining engineering is founded. 
 
 MAGNUS C. IHLSENG. 
 
 \STATE COLLEGE, PA.. Dec. ao, 1897.
 
 CONTENTS. 
 
 PART I. 
 
 MINING ENGINEERING. 
 
 CHAPTER I. 
 
 PACK 
 
 PROSPECTING i 
 
 Introductory; native metals; ores; gangue; definition of a 
 vein; fissure veins; mineral beds; the theory of vein-filling; 
 surface prospecting; geological aids to prospecting; explora- 
 tions; divining-rods; the proof of the existence of a vein; the 
 extra-lateral rights; the dimensions of a mining claim; side 
 line vs. end line; maintaining possessory right; the application 
 for a mining patent; chaotic condition of the mining law. 
 References. 
 
 CHAPTER II. 
 PREPARATORY WORK 18 
 
 Mine development; the buildings at the mine mouth; the 
 selection of a site ; choice of character of entry; the preparatory 
 work ; adits ; cross-cut tunnels ; vertical shafts ; inclines ; block- 
 ing out the ore body; dead- work; lifts in metal-mines; the 
 lifts in coal-mines; the dimensions of horizontal opening; re- 
 serves; pillar supports for mines; influence of cleavage planes; 
 faults; Schmidt's rule; open cuts; the steam-shovel; hydraulic 
 mining; the duty of giants; nozzles; peat; salt-mines. Refer- 
 ences. 
 
 CHAPTER III. 
 METHODS OF MINING 34 
 
 Exploitation; open-cut mining; milling; stripping; longwall 
 advancing; longwall retreating; comparison of systems; pillar-
 
 X CONTENTS. 
 
 PAGE 
 
 and-room system ; strength of coal ; influence of cleats upon the 
 resistance of coal; size of coal-pillars; relative dimensions of 
 pillars and rooms; robbing pillars; mining steep, thick coal- 
 seams; systems for gaseous mines; rock-chu'te mining; mining 
 the mammoth coal-seam; sqware work; flushing; underhand 
 stoping; overhand stoping; square sets; top-slicing and caving; 
 subdrifting and caving; slicing and filling. References. 
 
 CHAPTER IV. 
 POWER GENERATION, 87 
 
 The power plant; the boiler rating; the standard of boiler 
 comparison; the feed- water; boiler scale; the feed- water heater; 
 types of boilers; water-tubular boilers; heat losses; flue- gases; 
 the air-supply; calorific value of fuels; the rate of combustion; 
 the methods of firing; stokers; fuel consumption; weathering 
 coals; liquid fuels; draft; chimneys; blowers; boiler installa- 
 tion ; the steam-engines ; classification of engines ; the behavior 
 of steam; the mean effective pressure; table, value for the 
 constant, with various cut-offs for determining the M.E.P.; 
 the use of indicators; indicator cards; steam condensation in 
 the cylinders; remedies; the compound engine; governors; 
 automatic cut-off; the sliding steam- valves ; the Corliss valve; 
 the reversing-link ; the condenser; influence of clearance; the 
 speed of engines; the horse- power of a cylinder; the horse- 
 power of an engine; table, the horse-power for each pound of 
 mean effective pressure per square inch of piston area; the 
 steam consumption of an engine; table, the mean effective and 
 terminal pressures and steam consumption in Corliss non-con- 
 densing- and condensing-engines; setting the engine; the rotary 
 effort; the steam-turbine; the De Lavalle wheel; the Parsons 
 steam-turbine ; the advantages of the steam-turbine ; oil-engines. 
 References. 
 
 CHAPTER V. 
 HOISTING MACHINERY 127 
 
 Hoisting machinery; windlasses; balancing the loads; whims; 
 the hoisting-engine; first- and second-motion hoisters; second- 
 motion engines; the brake; the unbalanced load on the engine; 
 internal gear connection; friction-engines; V friction- wheels; 
 friction-clutches; balancing the dead load; balancing by conical 
 drum; the conical drum; the fusee; the reel; counterbalance; 
 the Koepe system of winding; chain counterbalance; the 
 Camphausen system; designing hoisters; data necessary in 
 designing; formulae for determining the number and duration
 
 CONTENTS. xi 
 
 PAGE 
 
 of trips; determining the size of the rope; the bending stress on 
 the rope; the available strength of a rope; example; deter- 
 mining the dimensions of cylinders ; the hoisting capacity of an 
 engine; hoisting on inclines; the starting moment of an engine; 
 table of maximum and minimum rotary efforts for certain 
 ratios of connecting-rod to crank-arm; determining the size of 
 the drum; the general formulas for hoisting resistances; table, 
 the starting moments, working loads, etc., with various hoisting- 
 engines and drums; electric hoisters; calculating the size of wire 
 for a hoister; example. References. 
 
 CHAPTER VI. 
 
 ELECTRIC GENERATION AND WATER-POWER 166 
 
 The application of electricity to mining; electrical units; 
 electromotive force; current force; capacity; electric energy; 
 table, to develop one mechanical horse-power at the motor- 
 shaft; Ohm's law; circular mils; electric currents; electric 
 generator; the winding of generators; theory of the dynamo; 
 alternating-current machines; the frequency; the double- 
 current generator ; table of properties of copper wire ; examples ; 
 general wiring formulae for alternating-current distribution; 
 example; resistances; the power factor; table of values of M 
 for wires 18 inches apart; electric systems; converters; the 
 transmission difficulties; the most economical area of wire; 
 the allowable voltage in mines; alternating-current motors; 
 rheostats; impulse- wheels ; flow of water through pipes; fluid 
 friction in pipes. References. 
 
 CHAPTER VII. 
 
 HOISTING MACHINERY AND UNDERGROUND CONVEYANCES 196 
 
 Underground conveyances; hoist-ropes; steel- wire ropes; 
 standard ropes; the lay of the rope; locked-coil ropes; flat 
 ropes; taper- ropes; the strength of a wire rope; table of 
 strength of cast-steel hoisting- ropes and their working loads; 
 the minimum radius of curvature; table of the radius of curva- 
 ture R or steel- wire rope in inches having i deflection and i 
 pound tension ; bending rope around curves and influence upon 
 its durability; the working load; resistance due to bending; 
 elastic connections; rope sockets; he head-frame; the height 
 of derricks; sheave; cage indicators ; signalling; the telephone; 
 means of increasing the safety of ascent and descent ; detaching- 
 hooks; engine controllers; buckets; kibbles; shaft-guides; 
 skips; gunboats; the slope-carriage; the cage; securing the cars
 
 xii CONTENTS. 
 
 PAGB 
 
 on the cage ; multiple-deck cages ; landing-doors at shaft stations ; 
 cage-chairs; self-dumping cages; mine-cars; car details; wheels; 
 car resistance ; lubricants; testing lubricating-oils by the hydrom- 
 eter; viscosity test; the flash-point; fire-proof oils; handling 
 cars at the tipple; dumping cars; cross-overs; the Ramsay cag- 
 ing apparatus; cradle-dumps; ladders; the man-engine. Ref- 
 erences. 
 
 CHAPTER VIII. 
 
 UNDERGROUND HAULAGE SYSTEMS 257 
 
 The extent of haulage requirements; the grade and direction 
 of a haulage-road; haulage- wavs in metal-mines and in coal- 
 mines; track construction; the gauge of track; turnouts; 
 curves; the degree of curvature of a curve; elevation of the 
 outer rail; formulae; conveniences at landings; safety devices 
 along haulage-ways; catches at the top of the incline; stops 
 along the incline; catches at the bottom of the incline; the 
 choice of haulage systems; tramming; animal haulage; the loco- 
 motive; the haulage capacity of a locomotive; formulae; the trac- 
 tive force; table of hauling capacities of steam-locomotives on 
 various grades; comparison of types of locomotives; the steam- 
 locomotive; the compressed-air locomotive; table of tractive 
 efforts of compressed-air locomotive; the electric locomotive; 
 details; the rheostat; the draw-bar pull of an electric locomo- 
 tive; locomotive rating; notation; storage-battery locomo- 
 tives; examples; cost of electric haulage; self-acting planes; 
 engine-planes; rope-haulage systems; the tail-rope system; 
 the endless-rope system; increasing the tension of the drums 
 at the driving end; supporting the haulage-rope; deflecting the 
 main rope around the curves ; supporting the tail-rope ; example. 
 References. 
 
 CHAPTER IX. 
 
 WIRE-ROPE TRANSMISSION 295 
 
 Ropes for power transmission; rope-driving systems; the 
 transmission of power; sheaves and end carriages; maintaining 
 uniform tension; the tension of the rope; the sag of rope; 
 formulae; examples; aerial tramways; the single-rope tramway; 
 the ore carriers; capacity; the cable supports; the Bleichert 
 system. References. 
 
 CHAPTER X. 
 
 THE COMPRESSION OF AIR 307 
 
 Compressed air; table of the weights of air at various tempera- 
 tures: table of absolute pressures, boiling-points, etc., at differ-
 
 CONTENTS. 
 
 ent heights above the sea-level ; free air; adiabatic compression; 
 isothermal compression; formulae; table of volumes, mean 
 pressures, temperatures, etc., in air compression; examples; 
 air indicator cards; the work of adiabatic compression; the 
 work of isothermal compression; table of hyperbolic logarithms; 
 table showing the horse-power required to compress and deliver 
 one cubic foot of free air per minute to various gauge pressures ; 
 the mean resistance to compression; cooling the air during 
 compression; water cooling; two-stage compression; table of 
 horse-powers necessary to compress 100 cubic feet of free air 
 to various pressures, with two-, three-, and four- stage compres- 
 sors; diagrams; the work of compressing moist air; table of rise 
 in temperature and work of compression of moist and dry air; 
 the air-compressor; the valves of the compressor; the effect of 
 the clearance space; the horse-power of the compressor; the 
 receiver; transmission of compressed air; factional resistance 
 in pipes; table of volumes and pressure- head losses of trans- 
 mission through pipes ; the loss of energy due to friction ; values 
 for a and D* for various diameters of pipe; the most economical 
 size of pipe ; examples ; compressed-air pipes ; table of weights of 
 standard and extra-strong pipes used for compressed-air haulage- 
 plants; the power value of compressed air; the work performed 
 while expanding; reheating; the efficiency of compressed air; 
 the Cummings system of air transmission. References. 
 
 CHAPTER XL 
 
 PUMPING 335 
 
 The water seepage into mines; methods of unwatering mines; 
 the hydraulic ram; hoisting water; mine-pumps; mine rod- 
 pumps ; pump-pipes ; valves on pump-pipes ; pump- valves ; the 
 working-barrel; formulas; the suction length; the lift-pump; 
 lift-pump plunger; the capacity of lift-pumps; force-pump 
 rods; the Cornish pump; the Cornish rod; the balance-bob; 
 the Cornish engine; the speed of the Cornish pump; Cushier's 
 double-acting drive-rod system; sinking pumps; the recipro- 
 cating pumps; pump- valves; single-cylinder pumps; the plunger 
 pump ; the comparative merits of the steam-pump ; the capacity 
 of a pump; examples; the duty of the pump; pressure- regu- 
 lators for pumps; the speed of the steam-pump; the suction 
 height; motor fluids other than steam; the displacement air- 
 pump; the air-lift pump; water-pressure engines; formulas 
 for pump calculations; examples; power-driven pumps; the 
 Reidler pump; electrically driven pumps; rotary pumps; the 
 centrifugal pump; the theory of the centrifugal pump; the
 
 XIV CONTENTS. 
 
 PAG* 
 
 compound centrifugal pump; efficiency of the centrifugal 
 pump; table of horse-powers and fuel required for two- and 
 three-stage centrifugal pumps for hydraulic mining or for 
 pumping purposes; the pulsometer; siphons. References. 
 
 CHAPTER XII. 
 MINE-GASES 386 
 
 Ventilation as an economic proposition; comparison of 
 metal- and coal-mines; gases in coal-mines; composition of the 
 atmosphere; oxygen; nitrogen; carbonic acid; sulphuretted 
 hydrogen ; carbonic oxide ; light carburetted hydrogen or marsh 
 gas; the occlusion of gas in coal; outbursts of gas from the coal; 
 the effects of mine-gases upon life and flame ; explosive gaseous 
 mixtures; black damp; after-damp; treatment for asphyxia- 
 tion ; the force of the explosion ; the barometric relation of ex- 
 plosions; the diffusion of gases; testing mine air for gas; the 
 height of Che flame in gas ; the height of the cap ; testing lamps ; 
 the Shaw gas-testing apparatus; the amount of air required for 
 combustion; the water-gauge; the mine resistance; the equiva- 
 lent orifice of the mine ; table of the water-gauge in two mines 
 carrying 30,000 and 100,000 cubic feet of air; table of air per 
 minute for various equivalent mine orifices. References. 
 
 CHAPTER XIII. 
 
 METHODS OF VENTILATION 405 
 
 The ventilation system; the ventilation of single entries; 
 ventilation by double entries; planning airways; the under- 
 ground temperature; natural ventilation; the tension of the 
 atmosphere; the weight of air; formulae; the production of 
 draught; the motive column; the systems of producing a 
 ventilating current; furnaces; the limits of furnace ventila- 
 tions; examples; types of fans; the trompe; pressure-blowers; 
 the Fabry blower; the Baker rotary force-fan; Cooke's fan; the 
 Lemielle blower: the centrifugal fans; open running- fans; closed 
 running- fans; the shutter regulator; the influence of shape, 
 dimension, and speed of fan upon its capacity; automatic speed- 
 and pressure- recorders ; comparison of fan and furnace; the 
 theory of the fan ; the equivalent orifice of the fan ; the relation 
 between mine and fan; the efficiency of the fan; the design of 
 a fan; table of theoretical water-gauge depression in inches 
 for the corresponding peripheral speeds in feet per second; 
 formulae; examples. References.
 
 CONTENTS. XV 
 
 CHAPTER XIV. 
 
 PAGB 
 
 DISTRIBUTION OF AIR 44* 
 
 The effective motive column; the frictional resistance of air 
 in mines; shape and dimensions of airways; table of pressure in 
 pounds per square foot due to the flow of air through one mile of 
 airway ; table showing lengths of roads offering a resistance equal 
 to one inch of water-gauge; formulae; calculation of the mine 
 resistance; multiple air-circuits; the process of balancing the 
 resistances; the power required for ventilation; examples; 
 goaves; the velocity of the air-current; the anemometer; the 
 efficiency of a ventilating system ; the regulation of multiple cir- 
 cuits; structural appliances required for controlling the air- 
 current; the means for deflecting the current; the stoppings; 
 air-crossings, or bridges; overcast; undercast; how should air- 
 bridges be made ? mine-doors ; regulator-doors ; the effect of an 
 obstruction in the airway; safety-doors; brattices; aid to 
 ventilation; example. References. 
 
 CHAPTER XV. 
 
 THE ILLUMINATION OP MINES 468 
 
 Illumination in mines; illuminants; the Davy safety-lamp; 
 defects of the Davy lamp; Stephenson's lamp; the Marsaut 
 lamp ; the Muesler lamp ; the Hepplewite-Gray lamp ; the Dick 
 patent port-hole lamp; the Clifford lamp; the Beard-Mackie 
 lamp; the Wolf benzine safety-lamp; the requisites of a safe 
 lamp; bonneted lamps; locking lamps; lighting and relighting 
 locked lamps; use of lamps; examining and testing lamps; 
 cleaning lamps ; electric lights in mines ; examples. References. 
 
 CHAPTER XVI. 
 
 ACCIDENTS IN MINES 483 
 
 The inherent dangers; comparative hazard in nations; tables 
 of ratio of deaths to tonnage and number of employees ; the death 
 rate in mines; statutory provisions governing underground 
 operations; table of average number of accidents in mines in 
 the United Kingdom in semi-decades; accidents due to falls; 
 remedies; accidents due to cars, in shafts, etc.; accidents from 
 use of explosives; gas explosions; remedial measures; the 
 theory of coal-dust explosions; ankylostomiasis; underground 
 fires; fires from blasting-agents; spontaneous combustion; 
 extinguishing fires; protective measures ; rules for guidance after 
 explosions; aerophores. References.
 
 xvi CONTENTS. 
 
 PART II. 
 
 PRACTICAL MINING. 
 CHAPTER I. 
 
 PAOB 
 
 SHAFTS 53 
 
 Shafts; how deep can we mine? a list of some of the deep 
 shafts, and mines of the world; sinking a shaft; extending 
 shafts by pentice; the service of shaft timbers; the character 
 of the timbering of shafts; timbering a shaft; timber joints; 
 shaft linings; square-set timbering; setting the timbers ; timbei 
 ing in firm ground ; repairing timbers ; building landing-stations, 
 forepoling; wood tubbing for circular shafts; masonry walling 
 of shafts ; shaft pillars. References. 
 
 CHAPTER II. 
 
 SINKING IN RUNNING GROUND 534 
 
 Excluding watery strata; tubbing; masonry curbing; sink- 
 ing through running ground; Triger's method by pneumatic 
 tube; the Kind and Chaudron process of boring shafts; the 
 trepans; the tubbing and the false bottom; making a water- 
 tight joint at the bottom; Lippman's drill; Haase's system; 
 the freezing system; conducting the seepage through the walls. 
 References. 
 
 CHAPTER III. 
 
 TIMBERING ROOMS AND GALLERIES 545 
 
 The service of timbering; the life of timber; timber consump- 
 tion; elements of timbering; props; mill-holes; braced stulls; 
 framed sets ; joints ; the dimensions of sets ; timbering roads in 
 large deposits; lagging; the use of iron under gound ; masonry 
 lining for roads; arch centres; tubular walling of galleries; 
 dams ; timbering soft ground ; the square set ; reinforcing square 
 sets; objections to the system; cribs; protecting underground 
 chambers; timbering landings; timber-cutting tools; mine 
 timber framing machine. References. 
 
 CHAPTER IV. 
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS 577 
 
 Levels; adits and levels; the alignment of a tunnel; rock 
 characteristics; driving in creviced rock; driving by air-drills; 
 driving slopes; driving tunnels; the use of supplementary
 
 CONTENTS. xvii 
 
 PAGB 
 
 shafts; driving a tunnel; tunnelling systems; comparison of 
 the four systems; tunneling soft grounnd; iron shields at the 
 tunnel-face; the hydraulic shield; the Anderson pilot tube. 
 References. 
 
 CHAPTER V. 
 
 DRILLING- AND BORING-MACHINES FOR EXPLORATIONS 596 
 
 Classes of machines; prospecting-machines ; bore-holes; 
 Poetsch method; percussive drills; punch-drill; oil-well rig; 
 the jars; Oennhausen's chisel; invention of M. Kind and 
 Chaudron the jar universally used; the drill-tools; sinker-bar; 
 spudding-bit ; auger-stem; bull wheel; the Mather and Platt 
 system; the temper screw; the sand-pump ; recovering lost tools; 
 the progress; hand-boring of deep holes; casing a bore-hole; 
 artesian flows; moss box; the Keystone prospecting-machine; 
 shooting the well; removing the casing; carbonators; borts; 
 the diamond-drill; the core-barrel; regulating the feed of the 
 drill; accidents; rotary drills; the diamond-bit; drill-bits; 
 removing the cuttings and the core ; the core ; deflection of the 
 drill-hole; fire-setting system; miners' tools; rotary drills; 
 hand-boring machines; the Chapman hydraulic pipe rotating 
 drill; the Davis calyx drill. References. 
 
 CHAPTER VI. 
 MINERS' TOOLS 615 
 
 The texture of rocks ; metamorphic rocks ; the massive rocks ; 
 underholing bituminous coal ; shooting off the solid ; shovels ; 
 picks; j adding; grip-drill; grip-drill illustrated; the post-drill; 
 portable power borers; anchor-pick or poll-pick; breaking 
 dimension stone; wedges; the plug and feather; hydraulic 
 wedges; gads or moils; blockholing; the jumper; the drill; 
 the depth of holes ; the stent ; removing the drillings ; hammers ; 
 single-hand and double-hand work; itemizing the drilling ex- 
 penses; the blacksmith-shop ; forge fuel; welding; steel; harden- 
 ing steel; tempering; the correct temper for tools. References. 
 
 CHAPTER VII. 
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS 635 
 
 Machine rock-cutters; quarrying; the quarrying of dimension 
 stone; channelers; the broach-bit; lewising; power drills; the 
 automatic steam- valve ; the tappet- valve ; the little Rand giant 
 drill; the Sergeant drill; the fluid-driven valve; eclipse drill; 
 the Schram drill; the Darlington drill; the drill-rods; the drill-
 
 CONTENTS. 
 
 PAGB 
 
 feed; cushioning the blow; the drill supports; column support; 
 the drill tripod; progress and cost by machine-drills; systems 
 of drilling holes; the centre-cut system; American method of 
 tunneling; simultaneous firing ; drilling in benches; Brain's radial 
 system; the continuous process of diamond-drilling; boring 
 headings; electric rock-drills; the solenoid drills; the box 
 electric drill; coal-cutting machines; the requirements of the 
 machine; the percussion machines; the breast machine; long- 
 wall machines; shearing coal-cutters; the comparative advan- 
 tages of coal-cutting machines; percussive machines versus 
 chain machines; comparison of hand-mining with machine- 
 work. References. 
 
 CHAPTER VIII. 
 
 BLASTING 671 
 
 The principles of blasting; blasting-agents; blasting-agents 
 of the first class; explosives; igniting and detonating explosives; 
 black powder; methods of charging the hole with powder; 
 tamping and ignition of the powder; the fuse; the cap; pre- 
 cautions in the handling of black powder; lewising; powder 
 consumption; the manufacture of nitroglycerine; dynamite; 
 the storage of nitroglycerine; methods of charging with nitro- 
 glycerine; the secondary explosion of powder; smokeless 
 powders; the theory of blasting; the line of least resistance; 
 flameless explosives; detonators; simultaneous firing; expand- 
 ing bits; the electric fuse. References. 
 
 GLOSSARY OF MINING TERMS 696 
 
 SIGNALLING 711 
 
 USEFUL INFORMATION 712 
 
 TABLE OF WEIGHTS OF VARIOUS SUBSTANCES 714 
 
 EQUIVALENTS OF FRENCH AND ENGLISH MEASURES 714
 
 MANUAL OF MINING, 
 
 I. 
 
 MINING ENGINEERING. 
 
 CHAPTER I. 
 
 PROSPECTING. 
 
 Introductory. The search for the useful and precious min- 
 erals has been diligently prosecuted since the early days of civili- 
 zation; their discovery and application have made nations pow- 
 erful exponents in the world's history. And nowhere is this 
 fact better exemplified than in our own land, in the wonderful 
 opening and rapid settlement of the Western mining States. 
 
 No subject is more entrancing, no occupation more exhila- 
 rating, than mining, with its wonderful kaleidoscopic changes. 
 In early times excavations were made and mines worked only 
 to a small depth and in easy rock, and that, too, only for sub- 
 stances of high intrinsic value, notwithstanding the myriads 
 of slaves to furnish the labor. The attempts at systematic min- 
 ing were few and far between ; but since the advent of the steam- 
 engine, mining has been acknowledged an important profession, 
 requiring technical education. Competition with the whole 
 world, brought about by the improved means of communica- 
 tion, the paucity of bonanzas and their rapid exhaustion, compel
 
 2 MANUAL OF MINING. 
 
 a skilful utilization of all the aids to a cheap extraction of our 
 immense wealth. 
 
 The accessibility of the mine and the vendibillty of its prod- 
 uct are the ever-ameliorating features in the mining history of 
 rations, districts, camps, and individuals, gradually divesting 
 mining of its risks and rendering it more and more akin to manu- 
 facturing. Each new camp, untrammelled by tradition to keep 
 it in the rut of prejudice, displays its genius for organization 
 and absorbs the latest devices, tried and true. Nevertheless, it 
 must be admitted that in each camp an adequate solution of the 
 problem involves intricate questions of environment. The 
 economy of mining is a function of many variables, as geological 
 stratigraphy, subterraneous uncertainties, wages, water, timber, 
 transportation, and treatment. The constants are few. The 
 proper relation of these it is our province herein to discuss. 
 
 Hitherto a gambling spirit has frequently controlled invest- 
 ments in metal mines. Speculative tendencies, not technical 
 economies, have dominated some of our operators; their heavy 
 aggregate outlay may have proven unprofitable, for the present, 
 because of salted mines, attractive prospectuses, or incompetent 
 management. It must be remembered, however, that they have 
 contributed to the prosperity of the country, and at some later 
 date their abandoned exploitations will be pursued to. profit t 
 when the potential investment of to-day will have been resolved 
 into future kinetic dividends, the cost of production being con- 
 tinually on the decrease. 
 
 Native Metals. The occurrence of the useful minerals in the 
 s:ate of native purity is rare. Still less often are they found 
 superficially: they must be delved for. In the extraction of 
 this subterraneous material, and its delivery to the surface, con- 
 sists the art of mining. The legal definition of a mine includes 
 such "workings as must be artificially lighted." 
 
 Gold and platinum are found native in the placer accumu- 
 lations of ancient and modern river-beds, which furnish fully 
 75 per cent of the total output of these metals. Gold occurs 
 in segregated veins, alloyed with telluirum, and always asso-
 
 PROSPECTING. 3 
 
 dated with pyrites and titaniferous iron; also intercalated 
 between the sheets of slate or shale, or finely disseminated in 
 eruptive rocks. The only extensive native copper deposit is 
 the remarkable product of the Lake Superior region, where 
 the irregular masses are mined out of the amygdaloid trap and 
 sandstone. Singular masses of metallic iron ore are found in 
 several localities, but they are curiosities and casual, if not 
 meteoric. Native silver is rare and occurs in Peru, Mexico, 
 Norway, and in the Lake Superior copper mines. 
 
 Ores. With these few exceptions the metals are found in 
 chemical union with non-metallic substances, more or less com- 
 pletely segregated to constitute mineral. Any accumulation of 
 mineral of good quality and in sufficient concentration to war- 
 rant the expenditure of energy for its extraction is an ore. Mani- 
 festly this is a fickle term, since it depends for its stability upon 
 the casual conditions of the market as well as upon the mineral- 
 ogical features. 
 
 The most common substance is iron, entering as it does 
 into almost all rocks and veins. Its most frequent, and value- 
 less, combination is with sulphur. Magnetic and specular 
 oxide and the carbonate constitute the entire supply. These 
 occur as irregular masses in the rocks of every geological age, 
 or in veins mixed with other minerals, but are chiefly in the 
 metamorphic crystalline, Archaean rocks. Zinc is obtained from 
 calamine, franklinite, and blende, which are quite extensively 
 distributed in the Carboniferous strata. With very few excep- 
 tions, galena is exclusively the ore of lead The carbonate and 
 the sulphide, in the lower Silurian and Carboniferous strata,. 
 mostly occur in irregular shoots and pockets, and rarely argen- 
 tiferous. In the older metamorphic rocks the galena is con- 
 fined in fissure veins carrying silver and gold. The main supply 
 of silver is from its minerals, more or less intimately associated 
 with other ores. Similarly with them, it has a wide geological 
 distribution, and is also found "dry" in fissures. Copper, as 
 chalcopyrite, bornite, and cuprite, is disseminated in and along 
 slates and sandstones, rarely above the Triassic. Many galena
 
 4 MANUAL OF MINING. 
 
 veins in the metamorphic rocks change with depth to copper. 
 Mercury comes from cinnabar, which is found in true veins and 
 in contacts. It is not commonly encountered. Tin has a 
 characteristic occurrence in but one form, as an oxide, and 
 only in gash or segregated veins, or "stockwerke" of the older 
 rocks. 
 
 Tin lodes are of the segregated type, and gold or silver bear- 
 ing, pyrites and cassiterite being the common minerals. 
 
 Millerite and pyrrhotite are nickeliferous and occur in gash 
 and segregated veins, rarely deeper than 500 feet. Rich films 
 of genthite in talc veins often constitute a commercial supply. 
 
 Manganese ores (standard contains 44 per cent of the metal) 
 are generally associated with limonite and occur in pockets 
 usually embedded in clay as contacts or beds or permeating 
 slates. Films of manganese appearing in moss-like forms on 
 the face of rock give it the name of "landscape" rock. 
 
 Mica is generally in bedded veins, . instances of contacts and 
 true lodes being rare. They are simply and always dikes in 
 coarse granite. Hitherto only large slabs were sought, but now 
 the fine, clean mica has a ready sale for lubrication and other 
 purposes. 
 
 Phosphate rocks for fertilizers, the practical value of which 
 is determined by the amount of phosphoric acid contained, are 
 found as beds of irregular thickness; veins or lodes transversely 
 to the strike of the strata, or superficial deposits. Apatite 
 occurs concretionary in a clay matrix between limestone and 
 clay. These are more frequent in the Miocene. 
 
 Many of the metals are incidentally obtained from their 
 mineral compounds while smelting for other metals with which 
 they are associated. 
 
 Gangue. The metalliferous portion of a lode comprises only 
 a small portion of its contents. The argentiferous galena, bor- 
 nite, blende, or their oxidized derivatives in grains, pockets, 
 or streaks, more or less connected, are associated with a "gangue" 
 of clay, quartz, fiuor, calc, or heavy spar. These earthy mate- 
 rials sometimes are intimately mixed with the mineral, and again
 
 PROSPECTING. 5 
 
 lie in layers contiguous with it, or the different constituents may 
 even manifest a ribbon-banded structure. 
 
 Definition of a Vein. The metalliferous and earthy contents 
 of a deposit constitute a bed or a vein, and may exist under such 
 circumstances as to render it workable. The term vein is in- 
 tended to describe a regular unstratified deposit in a fissure that 
 traverses the country for a considerable distance, longitudinally 
 and vertically. The Supreme Court has defined it as "any zone 
 or belt of mineralized rock lying within boundaries clearly sepa- 
 rating it from the surrounding rock." This demands a well- 
 defined crevice of ready identification, and two solid walls to 
 give it individuality. Its lead must be metalliferous. A vein is 
 the filling of a pre-existing fissure. The term has lost its 
 original significance, for formerly, the mineral system was sup- 
 posed to bear a resemblance to the human circulatory system. 
 True, the fissures have originated during periods of great dyna- 
 mic movement, producing folds and fissures which are supposed 
 to have extended deep into the earth's crust, but the main artery 
 has yet to be located. Though argentiferous lead veins are 
 quite persistent, no evidence exists for the dogma, so tenaciously 
 held, that they increase in richness with depth. They may or 
 may not become richer, or change, in constituents. Examples 
 can be cited for either side of the argument. In folded strata 
 the deposit inclines to be thicker at the ridges, or troughs, and 
 thinner at the sides of the folds. But this is not generally the 
 case in massive rocks. 
 
 Usually the vein matter is crystalline. It is commonly sepa- 
 rated on either or both walls from the surrounding rock by a 
 sheet of clay (called "selvage" or "gouge"), or by other quite 
 distinct lines of demarcation. The surface of contact of the 
 deposit with the adjacent rock is called a wall, roof, or floor, 
 according to its relative position to the miner. Not infrequently 
 the walls are polished surfaces ("slickensides"), due to grinding 
 caused by the slips during nature's contortions. Sometimes 
 portions of the vein have slid on one another, causing "false 
 walls"; therefore the miner is advised to occasionally break
 
 6 MANUAL OF MINING. 
 
 into the walls to assure himself as to the fact. On the other 
 hand, a vein may have only one or even no wall. In the process 
 of mineralization, the original face or faces of the fissure may 
 have become disintegrated, and all evidences of the looked-for 
 wall obliterated. In such cases, economic, not geologic, or legal 
 conditions define the vein. 
 
 Fissure Veins belong to regions of metamorphic action, and 
 are the principal repositories of the precious metals. And it 
 is a striking fact that they are rarely found singly, rather in groups 
 of parallel veins, often in congeries. Stockwerke is a term used 
 to describe a condition of affairs in which the country rock is 
 creviced in all directions, so that the whole mass must be mined 
 out. Some are filled with eruptive matter, others with vein 
 matter, still others were subsequently closed without any depo- 
 sition. The mineral components are markedly dissimilar, and 
 indicate different sources. Those filled with the same variety 
 of mineral were doubtless produced by contemporaneous forces. 
 Those fissures which interrupt the continuity of the older veins 
 are called cross-courses. The manner in" which the intersec- 
 tions occur determines their relative age. Their absolute age is 
 not ascertained, unless in stratified rock. Drags are more com- 
 mon than is supposed, and should not be confused with intersec- 
 tions. The latter are usually richer, the former not necessarily 
 so, at the point of juncture. Many of the older veins are broken 
 and displaced by faults. Not only do veins "pinch and shoot," 
 but the pay streak will vary in thickness, plunge from wall to 
 wall, or split up into numerous feeders and ramifications, and 
 even disappear in a thread. 
 
 Gash veins hold a subordinate position to fissures, but 
 they are of small extent, and are usually confined to a single 
 member of the formation in which they occur. Their habitat 
 is unmetamorphosed sedimentary rock. They have no distinct 
 walls or gouge, and are unreliable. 
 
 Mineral Beds. The important sources of mineral are the 
 metalliferous deposits which occur in the sedimentary strata, 
 and are termed beds. While the geologists may classify them,
 
 PROSPECTING. 7 
 
 the group is sufficiently identified by this term for mining pur- 
 poses. It includes deposits, somewhat irregular in dimensions, 
 occurring in the transverse joints of the rocks; as cementing 
 material to the remnants of shattered or insoluble rock; as layers 
 conformable with the strata; as isolated impregnations of grains 
 or bunches in porous rock; or as a metasomatic replacement of 
 porous rock. They may be found similar to fissures in a certain 
 formation, then as a blanket contact parallel to the stratification, 
 to again plunge into a lower series of rocks like a fissure, or branch 
 out into a chamber. They are more easily mined, but are less 
 persistent in depth, than veins. Their mineral contents are very 
 compact, seldom crystalline, and the gangue hardly distinguish- 
 able from the country rock. The mineral is more or less con- 
 centrated along certain lines called "ore-shoots," which probably 
 constituted the channels of communication with the ultimate 
 source. The same is also true of veins. 
 
 The Theory of Vein-filling. One requisite condition for 
 mineral deposition is a crevice, a porous or soluble rock conduit 
 for the fluid from which local action has precipitated the mineral. 
 Open cavities were not necessarily pre-existing, for a vesicular 
 rock would allow of an easy flow to the magma, or it might be 
 equally well secured by dissolving action on the rock and a sub- 
 sequent replacement. This is independent of its geologic posi- 
 tion. In every age are rocks which will satisfy this condition. 
 Besides this, a long train of circumstances has preceded the 
 vein-formation involving dynamic agencies, heat and meta- 
 morphism, and even eruptive action, as important factors. These 
 disturbances having been often repeated through the different 
 ages, the older rocks were more frequently shaken up. Beyond 
 this no reason exists for the prejudice which favors certain geo- 
 logical formations as ore-bearing. 
 
 The geognostical relations between veins and their contents 
 are of importance to the mining engineer, but our limited space 
 will not admit of any discussion here. The various works on 
 geology will supply the information as to the vagaries mani- 
 fested by ore occurrences and the numerous theories held. Some
 
 8 MANUAL OF MINING. 
 
 isolated examples exist under such circumstances as to suggest 
 the same origin for the ore as for the adjoining rock formations. 
 Many of the beds and veins have been impregnated by perco- 
 lating waters, perhaps at high pressure and temperature, con- 
 temporaneously with the country rock. Their metallic contend 
 may have been carried in solution or they may have been in a 
 molten or a gaseous state when the way for their passage was 
 opened. This is a matter for conjecture, as is also the ultimate 
 source of the mineral. The evidences frequently point to their 
 deposition as sulphides, the oxidized forms being accounted for 
 by long-continued action of atmospheric agencies. In the 
 presence of coal and bitumen in many lead and zinc beds is 
 suggested a theory of cause. The "water-line" theory has 
 served its day and is no longer tenable.* The current theories 
 have offered more or less satisfactory explanation of the genesis 
 of some of our ore deposits. Some one theory may explain 
 some of the capricious examples of lodes or their anomalous 
 fillings. But when we find contiguous depositions contrasting 
 widely in point of density; narrower parts of fissures filled by 
 denser or richer ores; superior minerals higher up than the more 
 volatile or lighter ones, even alternating with them, it is impos- 
 sible to advance a theory that is specifically applicable to all 
 ore occurrences. Aqueous currents have apparently conveyed 
 the minerals from their source and by evaporation chemical 
 decomposition or electrolytic action deposited them simultaneously 
 with or subsequent to the formation of the cavity in which they 
 are found The veins we find, but not always the silver; and 
 this inability to formulate a general law by which to locate the 
 hidden bonanzas has led to the compounding of the numerous 
 witcheries, and divining-rods of every conceivable form, for 
 imposing upon the credulity of the prospector who seeks a quicker 
 means of acquirement than is afforded by the use of the pick, 
 shovel, and patience. 
 
 There is no particular angle of dip or bearing of trend that 
 is universally favorable to rich veins. Rules based upon such 
 observations are local only. The same may be said as to the
 
 PROSPECTING. 9 
 
 supposed "live "-ness of certain rocks to mineral. Attempts to 
 formulate indications of "quickening" mineral by associations 
 with general gangue matter or minerals have failed of general- 
 ization. The mineral is where you find it. The Cornishman's 
 adage, "riding a zinc horse to fortune," has no verity in this 
 country. Each locality has its own peculiarities of mineraliza- 
 tion, which the careful and systematic engineer will observe 
 and regard. 
 
 With the two classes of rocks, stratified and massive, are 
 coexistent the two classes of mineral deposits, beds and veins. 
 Though many occurrences are of a nature that admits of question 
 as to classification, for mining purposes a sharp line of distinc- 
 tion is not sought. Legal technicalities have so confused the 
 definitions of deposits and veins as to obliterate all semblance 
 to the original intent of geologists and mining men. Of this, 
 more later. At present we shall consider some rules to assist 
 the prospector in his search for mineral. And while it must 
 be admitted that many a find has been made through accident, 
 the existence of the ore would be found not to be at variance 
 with the cumulative rules of geologic science. 
 
 Accordingly, the prospector will depend upon geological 
 data. In regions of stratified rock the- matter is simple. Coal 
 is found in three geological horizons, and the presence or absence 
 of the rocks belonging thereto is indicative of the pros- 
 pects. 
 
 The metals and their minerals are distributed, geologically 
 and geographically, over a large extent. The zinc ores in this 
 country occur in the Carboniferous and along the Mississippi valley. 
 The Archaean and Silurian are most prolific of the other ores. 
 The precious metals are chiefly found in the mountainous dis- 
 tricts, because the phenomena attendant upon their formation 
 were conducive to the filling of veins, and the forces which gave 
 character to the mountain also impressed themselves upon the 
 vein, which is exposed to view and subject to location. With- 
 out some such providential occurrences to change the monoto- 
 nous topography of the preadamic surface, bedded veins of the
 
 10 MANUAL OF MINING. 
 
 stratified districts would have been revealed only by boring,. 
 while those in massive rocks might never have been formed. 
 
 Surface prospecting is confined, therefore, to the seeking for 
 an outcrop. In igneous rock the outcrop is easily found. For, 
 unless the hill is covered with slide rock, it is indicated by a 
 jutting ledge (if the vein matter is harder than the country rock), 
 or by a sag (if it is decomposable). In heavy timber this may 
 go unnoticed. At high altitudes snow in the sags calls atten- 
 tion to the leads. 
 
 The same is true of coal, which is located by the terraces 
 which mark the outcrop. The trend of the terrace, relative to 
 the topography of the hill, gives a good idea of the slope of the 
 coal. The bench itself may give the desired information, but 
 usually it will be found that the coal dips with the hill, when 
 the terrace or depression deflects outward toward the bottom of 
 the hill, and the reverse for a coal dipping inward, when the 
 outcrop will be concaved toward its top. 
 
 Substances foreign to the rock deserve notice. Alternations 
 in the color of the slide rock covering the hill are good indications 
 of the presence of oxidizable minerals above. So, too, vegeta- 
 tion is a guide. Iron springs often accompany the outcrop of 
 coal; the ochreous covering of the rocks and soil is noticeable 
 near some of the anthracite seams, and is common in the semi- 
 bituminous districts. Masses of highly oxidized matter, broken 
 from the veins, compose what are called "blow-outs," and are 
 common in galena regions. 
 
 If no evidences of outcrop are thus found, "booming" may 
 disclose it. During winter or a wet season, snow or water is 
 collected in a reservoir upon the hill, and, at a convenient time, 
 turned loose to plough its way over the soil in its fall. Many 
 a vein has been thus discovered without great expense. 
 
 In stratified regions the order of the geological series may 
 be observed, and certain fossils furnish the guide. Or, if the 
 prospector is examining new ground, he has but to look for min- 
 eral in the float on the surface or in creek-bed. The appearance 
 of material derived from erosion is indicative of the character of
 
 PROSPECTING. 1 1 
 
 the rock from regions higher up. Therefore the bed of the 
 stream, or the hill slope, is minutely examined for fragments of 
 ore, or blossom, and followed as long as mineral is found. If 
 the float or shode boulders are pebbly or rounded, or in vegetable 
 soil, they have come from afar and the lode is not at hand. If 
 the shode is large and angular, it has not come very far, and the 
 discovery of a point beyond which no float or blossom is detected 
 is presumptive evidence of approach to the vein. The lode will 
 be found above the point of discovery, and the prospector will go 
 in the direction of the drainage and thoroughly search the ground. 
 
 In high altitudes the oxidation of the minerals in, and the 
 electric manifestations of, the vein outcrops have assisted the 
 prospector by the light playing over them. This is of continued 
 occurrence in Colorado above timber line, and particularly in 
 regions of arsenical veins. - 
 
 When found, the vein should be examined, and its value 
 confirmed at several points; most grievous disappointments 
 have ensued from testing of the lode at one point only. If the 
 country is stratified, care is taken to ascertain all the data of 
 thickness, etc. Frequently the ore oxidizes and rots away, to be 
 crushed by the overlying strata, showing only in a small streak; 
 or the outcrop may fold back, "tail out," and give false impres- 
 sions of great thickness. 
 
 Geological Aids to Prospecting. Maps are serviceable as 
 showing the important features, and a systematic plotting of all 
 data, geological and otherwise, is necessary. Dr. H. M. Chance, 
 in the Second Geological Survey of Pennsylvania, has an admir- 
 able discussion on the construction of geological cross-sections, 
 to which the reader is referred. Prospecting for oil or gas is 
 speculative. Drill-holes must be carried down to the oil-bearing 
 sands, and there is no surface guide to determine the site for a 
 well to be bored within the known limits of a given field. 
 
 Explorations. If the surface fails to reveal the mineral body 
 sought, and there remains reasonable expectation of finding 
 it, a tunnel, a shaft, or boring may be resorted to. The two 
 former are more expensive but safer guides than that offered
 
 12 MANUAL OF MINING. 
 
 by boring. Shafting is slower and more costly than tunnelling, 
 but more quickly reaches a flat seam at a point suitable for devel- 
 opment. The steep pitching vein is perhaps best reached by a 
 tunnel, if the depth of vein so gained is great enough to com- 
 pensate for the length of tunnel. The choice between them 
 depends upon local conditions. Both are advisable for shallow 
 explorations, while drilling may be employed for deep work. 
 The latter is very commonly employed on account of its cheap- 
 ness. But even when it has determined the data previously 
 doubtful, the shaft or tunnel must subsequently be driven. So 
 drilling has its limitation of use. It is rarely employed as a 
 seeker for mineral, but merely to give confirmation to, and assist 
 in a rational estimate of, the value of the undertaking. Many 
 properties owe their rehabilitation to the results of the diamond- 
 drill exploitation, and none should be abandoned until after a 
 careful surface examination has been made and followed by 
 numerous bore-holes. 
 
 Either the punch or the diamond-drill method may be used 
 for the boring. The former is cheaper, but the pulverulent 
 material brought up by the sludger is unsatisfactory; it may 
 indicate the constituents of the rocks pierced at different depths, 
 but can give little of its physical character or dip. The diamond- 
 drill core yields a little more information, but even its indications 
 are hardly trustworthy. It affords an opportunity to identify the 
 rock, but some of the soft strata are worn away or the core may 
 be turned in its tube, so its revelations are not much better than 
 those of the sand of the punch-drill. At best, the results obtained 
 from either cutter are not conclusive, for it may have just missed 
 the mineral, or have struck a solitary, small, soft chunk of ore, 
 which would supply cuttings to discolor the sands for a long 
 distance and give amazing report. Very important deductions 
 cannot be based solely upon the indications of the borings. 
 Only after numerous holes have been bored and a thorough sur- 
 face examination has been made can a conclusion be reached. 
 
 Good, hard common-sense, observation and pluck win, and 
 they alone. There is no mystery about the finding of mineral
 
 PROSPECTING. 13, 
 
 deposits. Nature is bountifully supplied with precious metals 
 and valuable minerals, but her secrets are hid. Only the cumu- 
 lative information of geological experience gives any clue as 
 to the habitat. Neither witchery nor magic charm can disclose 
 the whereabouts of an ore body or deposit. 
 
 Divining-rods. The wizard with the hazel wand, or the 
 spirit medium who is controlled by some disembodied Comanche 
 chief, is an impostor. He affects a versatility and occult power 
 that transcends combined scientific knowledge, but to a paltry 
 amount of "filthy lucre" he is not averse, when he plays upon 
 the credulity of natures which are duped into making extensive 
 explorations upon his purported previsions. This would be 
 ludicrous, were it not also painful, to see the number of mis- 
 guided men who have squandered hopes and possessions in their 
 search for a short cut to wealth. 
 
 The Proof of a Vein. The discovery of mineral at the sur- 
 face must be followed up by the disclosure of the vein. In 
 order to secure possessory title to it, the United States laws 
 require the proof of the existence of a lode or vein. Test pits or 
 shafts are sunk to reveal the mineral in place, having a definite 
 direction of outcrop. A single shaft is not regarded as sufficient 
 evidence of a vein. The mineral must be in place. Even if 
 disintegrated near the surface, it is still a vein if a crevice pre- 
 vails carrying mineral matter between rock of a nature and 
 origin different from it. The U. S. statutes divide mineral 
 ground into veins and placers only, hence the presumption would 
 be that any well-defined metalliferous crevice capable of ready 
 identification by the miner is a vein, whether fissure or not, 
 only it cannot be a placer. 
 
 The Extra-lateral Rights. The difference in the grants 
 under the two cases, besides a difference in acreage, is that the 
 mining of ore within placer ground is confined to the vertical 
 planes through the boundaries (sec. 2329, U. S. Rev. Statutes), 
 while vein deposits may be pursued along their dip, "through- 
 out the entire depth," even if they "so far depart from the per- 
 pendicular" "as to extend outside of the vertical side lines of
 
 14 MANUAL OF MINING. 
 
 the claim"; and the extent of the miner's right is determined 
 only by the vertical planes through the end lines, which should 
 therefore be properly drawn. Extra-lateral rights are, therefore, 
 accorded to vein locations. 
 
 The Dimensions of a Claim. The prospector announces his 
 discover) 7 by recording the fact, together with the data of dis- 
 covery and the direction of the outcrop of the vein, upon a stake 
 placed at the property and upon the record books of the county. 
 This done, he may proceed to work. 
 
 Locations 1500 feet in length are permitted upon the public 
 domain to the discoverer of the lode. But for access thereto, and 
 for convenience of working, the U. S. grants, as incident to the 
 principal feature, surface ground which, measured from the middle 
 of the vein, shall not exceed 300 feet on either side. Some States 
 have reduced this to 150 feet on each side, while in some Colorado 
 counties only 25 feet was, and 75 is, the outside limit. The 
 claim must be essentially a parallelogram. It may be 1500 
 feet, or less, in length, located substantially along the middle 
 of the apex, across which are drawn two parallel end lines and 
 side boundaries, within the limit prescribed, parallel in pairs 
 following the contortions of the outcrop. However else the Act 
 may be vague, it certainly is not upon the fact of the parallelism 
 of the exterior boundaries. Excessive locations are valid as to 
 the legal limit and void as to the excess. 
 
 Maintaining Possessory Right. It is incumbent upon the 
 locator to define the boundaries of his claim, by placing stakes 
 at all corners and intersections, to notify others that the ground 
 is entered upon and being exploited. These, with the filing of 
 a location certificate in the county, maintain possessor}' right 
 from the moment of posting a location notice of discover)' upon 
 the lode. Within a reasonable time thereafter, sixty days usually, 
 the locator is required to sink a "discovery" shaft at least 10 feet 
 into the vein. This satisfies the regulations regarding discovery, 
 and maintains a mining right against all comers until the expira- 
 tion of the calendar year. 
 
 From that time on, an "assessment" of $100 must be ex-
 
 PROSPECTING. 1 5 
 
 pended annually as evidence of mining intent. A failure to 
 expend such sum constitutes a forfeiture, by which the claim 
 reverts to the public domain, and is subject to relocation. 
 
 A prospector is not confined to a single entry upon a dis- 
 covered lode. He may appropriate as many claims as he chooses, 
 contiguous or otherwise, with that of the first discovery. Upon 
 each 1500 feet, or less, of length he must show the intent to. mine, 
 by a discovery shaft and the assessment work. 
 
 For the development of the mine, the annual assessment 
 work may be done upon the surface or upon the vein, and all 
 efforts outside of the limits of the location with a bona-fide intent 
 to work the claim are justly considered as if upon the claim 
 as, for instance, development by tunnel instead of shaft. 
 
 This concession is further extended by the U. S. Supreme 
 Court; for where one person owns several contiguous claims 
 capable of being advantageously worked together, one general 
 system of development may be adopted, after the discovery 
 shafts are driven. This encourages more economic work and 
 subserves the best interests of all concerned. 
 
 The Application for Patent. When development work to 
 the value of $500, exclusive of buildings, is completed within 
 five years, by an annual assessment of $100, he may proceed to 
 obtain the "patent" or legal title from the government. For 
 this purpose an approved survey of the property claimed by 
 him is necessary. An announcement of his intention to apply 
 for patent is placed on the property, in the local land office and 
 in the newspapers, for sixty days. If no opposition is filed 
 against this request to the government, a patent will be issued 
 to him when he shall have proven the previous facts, and that 
 he is a citizen of the United States and has paid $5 per acre for 
 the land enclosing the vein. 
 
 According to the laws, the applicant may obtain as many 
 claims as he desires, provided he satisfies the requirements for the 
 annual assessment prior to making application to the government. 
 
 Chaotic Condition of the Law. The government has been 
 exceedingly liberal to its citizens in throwing open the mineral
 
 1 6 MANUAL OF MINING.' 
 
 lands for exploration and occupation. The original intent was 
 to legalize the possessory system, which had grown up in the 
 absence of Federal legislation, by which the possessor claimed 
 a vein for mining purposes and was allowed the use for right 
 of way to the surface overlying his mine. This entailed the 
 granting of the mineral vein free but the enclosing non-mineral 
 land by sale. 
 
 An exclusive right of enjoyment "of all veins" cropping 
 inside of the boundaries is given with the claim. If they are 
 discovered and entered upon in adjacent territory, the subse- 
 quent locator, according to the laws of the State of Colorado, may 
 have right of way through the cross- vein to his ground on the 
 other side of the prior claim, but none of the mineral. In every 
 case it is intended that priority shall govern. Sec. 2326 grants 
 to the senior locator the mineral at the intersection, and to the 
 junior the right of way through it. 
 
 By the interpretation of the U. S. Statutes, easement and 
 title were clearly intended to be conveyed for all forms of metal- 
 liferous deposits, in the use of the terms "veins, lodes, or rock in 
 place." The Act recognizes any mineralized rock in place, 
 enclosed in the general mass of the mountain, as a vein. 
 
 The Statutes favored the miner and assumed to cover all 
 lodes whose indications were sufficiently marked for the miner 
 to continue explorations thereon. A crevice, crevice matter, a 
 fair wall, and mineral are the essential conditions. 
 
 It has been seen that a lode claim, whether patented or not, 
 carries with it all that is beneath the surface-ground claimed, 
 with a servitude upon the adjoining territory obtaining the right 
 of following the dip of the vein, and subject to a like easement 
 granted to the locator on adjacent ground to pursue his vein 
 wherever it may go. This obtains until some one can show a 
 better right. The common law as to realty is modified when 
 applied to mining property. 
 
 It has, however, happened that rulings were so made and 
 construed that a party may locate vacant ground and maintain 
 ownership to the mineral covered by it, unless it is shown that
 
 PROSPECTING. 17 
 
 the mineral body belongs to a lode cropping elsewhere within legally 
 claimed ground. The proprietor who calmly continued work 
 upon his discovery found himself breaking into the subterraneous 
 workings of others who had stolen a march on him. To secure 
 his right he had to bring action to eject. To vindicate his title 
 he had to prove the lode to be in place and continuous from the 
 point of his discovery to, into, and through the ground of the 
 trespasser. Failing to do which, his claim was defeated and with, 
 it all incidents thereto attached. 
 
 Naturally the train of reasoning led farther and farther away 
 from the original intent of the law to reward the discoverer of an 
 apex, until the accepted idea is that, although the "defendant's 
 location may appear to you to be along the line of the top, apex, 
 or outcrop of the vein, it cannot prevail against a senior location 
 on the dip of the lode." 
 
 To what absurdities the law has led us, by reason of the 
 vagarious interpretations, the reader may learn by referring to 
 Dr. R. W. R. Raymond's articles in the Transactions of the 
 American Institute of Mining Engineers. 
 
 The remedy is to repeal the present enactment, or else to so 
 prescribe and define the subjects of the United States grant that 
 purchaser shall have a warranty title to the entry. The side- 
 line law of Leadville is far preferable to the present uncertain 
 grant of extra-lateral right. The risks of mining are sufficient 
 without adding this unnecessary one. It is singular that, what- 
 ever amendments have been proposed in recent years, they per- 
 sist in retaining the essential feature of the extra-lateral right 
 depending upon the apex. So long as this basis is retained they 
 will do no good. There is no reason for longer maintaining this 
 abnormal and indefinable privilege in the grant of mineral land 
 made by the U. S. Government. 
 
 Justice W. E. Church, in a concluding and conclusive sen- 
 tence of a decision, said: "The present laws are a hotbed of 
 litigation and a fruitful source of error." Judge Bradley de- 
 clared them "imperfect," and those who have had any experience 
 with them will agree.
 
 CHAPTER II. 
 
 PREPARATORY WORK. 
 
 Mine Development. Assuming that the question "Can it 
 pay?" has been answered affirmatively, the next feature to be 
 considered comprises the extent and character of the surface 
 buildings, the nature of the preparatory work, the exploitation 
 plan, and the machinery to be installed for hoisting, pumping, 
 ventilating, and for the treatment of mineral. 
 
 The Buildings at the Mine Mouth. These comprise the 
 structure for the power plant and such additional buildings as 
 are needed for the preparation of the mine product. If the latter 
 be pure, or rich enough to be marketable, or if the coal is to be 
 coked, a simple storage building is sufficient. If it is to be sorted, 
 crushed, or screened and shipped, a tipple or a breaker will be 
 required. If the mineral requires treatment and the local water- 
 supply be deficient, or too acid, the washery, stamp-mill or con- 
 centrator will be located elsewhere and the mineral delivered to 
 it by tram. 
 
 Collieries provide standing room for a day's run-of-mine 
 cars, a storage track for empties and a side track for handling 
 the railway cars, wagons, or coke-larries. If the coal is screened, 
 three or four tracks, side by side, are connected with the main 
 track at both ends. 
 
 All buildings should be as nearly fire-proof as the risk demands 
 and the first cost will permit. A large colliery with, say, 1000 
 acres of ground to work, or a well-developed metal-mine, will 
 thus erect very substantial buildings. These should be as near 
 to the shaft as safety to the mine will allow and yet be remote 
 
 18
 
 PREPARATORY WORK. 19 
 
 from the mouth, as far as is consistent with economy. The 
 hoisting engineer should have an unobstructed view of the dump- 
 ing platform, and his engine should be far enough away to give 
 the rope a good fleet angle. The hoist-frame is open and not 
 housed. 
 
 The Selection of a Site. The location for the offices, the 
 buildings, and the point of shipment for the mineral must satisfy 
 the surface conditions, must be convenient to the shipping- 
 tracks, and at the same time one which does not require much 
 grading for the buildings. The location of the railroad and its 
 switches fixes the location of the ore-sorting or coal-screening build- 
 ing. As the surface plant may become quite extensive, its loca- 
 tion must be considered with a view to the economy of handling 
 in such a manner that the mineral shall descend naturally in its 
 progress through the works. The location of the screens fixes 
 the site for the mouth of the mine. The latter must therefore 
 be in the best position for both the track and the mine. 
 
 The opening into the ore body must satisfy the underground 
 conditions of haulage and drainage, and it must be centrally 
 located to the lowest point of the vein. It would be an ideal 
 condition if the shaft which would satisfy the shipping require- 
 ments could also be placed in the centre of the basin with the 
 haulage and drainage toward it from all sides. As this is not 
 likely, some compromise must be made between the various 
 demands of surface construction and mine operation. 
 
 Choice of Character of Entry. The method of attacking the 
 vein, or bed, varies with the inclination of the mineral body and 
 its depth. Undoubtedly, when the mineral occurs at considerable 
 depth, the sole method of attack is by a vertical shaft. If the 
 depth be not very great, a slope from some convenient point may 
 be preferable. When the vein, or the coal-seam, crops out at 
 the surface, within the property lines, an entry to the lower depths 
 is obtained by a slope following the inclination of the vein. If 
 the outcrop so exposed is quite long and extends along the side 
 of the hill, several slopes may be driven to advantage. When 
 the outcrop occurs on the hillside, the vein occupying a position
 
 20 MANUAL OF MIX IXC. 
 
 somewhat vertical, undoubtedly the best method of entering the 
 vein consists in driving one or more horizontal openings, or 
 adits, into the vein at various heights according to the system to 
 be employed in the subsequent explorations. 
 
 Owing to the difficulties and the expense of sinking shafts 
 and of hoisting the material through them later, conditions 
 sometimes arise when it may be preferable to drive a tunnel from 
 some point as far below the outcrop as possible or desirable 
 and toward the vein with the shortest entry. Such a tunnel is 
 called a cross-cut or a cross-country tunnel, and in some regions 
 is known as a rock tunnel, the significance of the titles being 
 easily understood. 
 
 Certain conditions in coal-mines warrant the driving of rock 
 slopes as a compromise between the horizontal tunnel and the 
 vertical shaft. They are employed where only a shallow shaft 
 would be required, but where a high speed of hoist is regarded 
 as desirable. 
 
 Naturally the endeavor is to secure the most direct line to the 
 point of exploitation, within the shortest time and at a minimum 
 cost, while at the same time constructing a permanent opening 
 which would serve for all the work during the life of the mine. 
 The opening of the mine is near to the mill-building where the 
 mineral is screened, weighed, washed, or treated, preparatory 
 to being marketed. The character of the opening can be decided 
 only after knowing the exact nature of the underground condi- 
 tions of the vein. Its extent, thickness, depth and general inclina- 
 tion must be known. Some clue to this can be obtained by 
 prospecting the entire field with the use of bore-holes and open 
 cuts, and geological maps are exceedingly serviceable, but in 
 metalliferous districts it is hardly possible to foresee the character 
 and lay of the ore body. Without this full knowledge, it is likely 
 that the opening made for the mine during exploration would 
 later be abandoned as being in an undesirable position or too 
 small to meet the final demands upon it. In coal-mines the 
 eccentricities of the vein are not so pronounced as to give the 
 same difficulty to the engineer.
 
 PREPARATORY WORK. 21 
 
 Coal seams have less irregularities than do the metalliferous 
 veins. Their contents are quite uniform in character, and the 
 thickness of the seam varies to a less degree than does a vein. 
 Both are found dislocated by faults or shattered by movements 
 of the strata. The metalliferous vein, however, not only changes 
 in thickness and inclination but its mineral contents will vary in 
 percentage and even in chemical composition. Frequently the 
 mineral at a lower depth bears no resemblance to that at the 
 surface. 
 
 The character of the mine opening depends upon whether the 
 ore body or bed of coal is revealed at the surface or not, and 
 whether or not the outcrop is included within the boundaries of 
 the property. It may occur on the hillside; its strike of the 
 outcrop may follow the hill, nearly horizontal; or it may lie in 
 the direction of the slope somewhat approaching the vertical. If 
 there is no outcrop the depth of the ore body is known. 
 
 The Mine Opening The entry into the mine may begin on 
 the mineral at its outcrop and, continuing, follow it wherever it 
 may go; or the opening may begin at some point on the surface 
 away from the vein to intersect it at a selected point. 
 
 The openings are horizontal, vertical, or inclined, the choice 
 between them being determined by the conditions named above. 
 The horizontal openings comprise drifts, adits, and tunnels. 
 The vertical openings are designated as shafts. The inclined 
 openings are termed slopes, or inclines. The open cuts are 
 either quarries or strippings. 
 
 Metal-mines are operated usually from one entry or opening. 
 Only under certain conditions presenting dangers of caving and 
 other reasons requiring the hasty removal of the mineral is it 
 desirable for metalliferous mines to have many openings. Coal- 
 mines are required by law to maintain two parallel openings at a 
 certain stated minimum distance apart. These are driven simul- 
 taneously or upon the coal-seams. 
 
 The Preparatory Work. The operator of a coal-, clay-, or 
 lead-mine is confined, in his choice of site for the shaft, or 
 tunnel, within the property lines. Metal miners, in regions
 
 22 MANUAL OF MINIXG. 
 
 other than the national domain, are also confined by the vertical 
 planes through their boundary-lines. 
 
 On the national domain the miner is not restricted to his own 
 property in choosing the site or the character ofen try for his 
 mine. His property is bounded by vertical planes through the 
 end lines, but not through the sides; and he may therefore 
 operate his vein on an inclination to any depth desirable. He 
 is not confined to his surface boundary lines when selecting a 
 site for the point of attack upon the vein. He may sink a shaft 
 from outside the boundary-lines, or drive a tunnel from the foot 
 of a hill to his vein, provided only that no other existing mining 
 rights are injured thereby or are interfered with. A greater 
 latitude of choice is therefore afforded the precious metal miner 
 than the 'coal-operator. Under such conditions, the tendency is 
 naturally to drive the horizontal cross-cut tunnel to the vein 
 as being the most economical method of developing the prop- 
 erty. The great expense and the changes in the character 
 of the vein militate against using a cross-cut tunnel, notwith- 
 standing the facilities it offers for shipment. It is a question 
 whether they would outweigh the importance of " staying by the 
 mineral." 
 
 Adits. Veins cropping out on the side of a hill are often 
 exploited by driving adits, which are merely horizontal openings 
 following the vein. The cost of driving is less than that of a 
 slope or a shaft. The cost of equipping is slight, and it serves all 
 purposes of permanent hauling, drainage, and ventilation, as 
 well as during explorations. These adits are driven at distances 
 apart vertically, depending upon the thickness of the vein and 
 the facilities for shipment. They block the vein into sections, 
 each one of which above the adit can be mined at a very low 
 cost. Their dimensions, equipment, timbering, and gradient 
 are the same as those of underground galleries. 
 
 Cross-cut Tunnels. These have all the advantages of the 
 adit in furnishing a convenient haulage and drainage-way at. a 
 point far below the average of the workings. They furnish a 
 cheap, secure, and permanent entry if the mine has been developed.
 
 PREPARATORY WORK. 23 
 
 It is a question, however, if a long tunnel should be driven as 
 a means of exploration before the mine has been carried to the 
 corresponding level that the value and character of the mineral 
 may be known. 
 
 Vertical Shafts. These constitute the most secure, but the 
 most expensive, method of entry into the mine. But the shaft 
 presents many of the objections which also obtain against the 
 tunnel. The shaft is the only means of economically reaching 
 a coal-bed which is not shallow and whose outcrop does not 
 occur within the lines of the property. For reaching veins this 
 is an uncertain method of attack, particularly if the vein changes 
 materially with depth in its inclination and character of mineral. 
 The great expense of sinking a shaft requires it to be carefully 
 located in a position which is certain to be that of the permanent 
 outlet of the mine. 
 
 If a choice of position be afforded the engineer, the foot- 
 wall side of the lode is usually selected as being safer than the 
 hanging-wall side. But each successive lower cross-cut is 
 longer than the one above it, and if the country rock be hard 
 and the vein very steep, the cost of these cross-cut drifts would 
 soon become prohibitory. For this reason a majority of shafts 
 are sunk from the hanging- wall side to intersect the vein at a 
 moderate depth. 
 
 The size of the shaft is as great as the capital will allow, 
 though the tendency is to restrict its area. The dimensions of 
 the hoistway depend upon those of the mine cars, which in turn 
 depend upon the character of the underground haulage systems 
 and the dimensions given to the haulage-ways. The width of 
 the shaft compartment is equal to the length of the car plus 6 
 inches at each end. This direction is the long side of the. com- 
 partment and is laid parallel to the mine tracks and also to the 
 straight track to the tipple. The shaft may have only two com- 
 partments, both being used for hoisting; three compartments, in 
 which the third is for timbering, piping, etc.; or even more com- 
 partments, the last being for ventilation. The other dimen- 
 sion of the shaft will be determined when the decision has been
 
 24 MANUAL OF MINING. 
 
 reached as to the number of cars to be run upon each cage at a 
 time and the number of compartments required. The capacity 
 of the shaft is determined by the conditions discussed in Chapter 
 V; the dimensions of the shaft, the method of sinking and 
 timbering, in Chapter I of Part II. 
 
 Inclines. Inclines and slopes are in great favor for many 
 reasons. They usually follow the vein and explore it, thereby 
 giving a return in the mineral recovered for the cost of driving 
 The maintenance of slopes is, however, higher than the cost 
 in shafts. Slopes are provided in coal regions when the inclina- 
 tion is over 10 and the depth to be reached not over 500 feet. 
 
 When the vein is constant in pitch and has no irregulari- 
 ties, everything favors its driving. If, however, the vein has 
 occasional enlargements or a varying pitch, its pursuit becomes 
 awkward. The question then arises as to whether it is desirable 
 to follow rigidly the irregularities, to reduce them to uniform 
 pitch or to abandon the entry altogether for a shaft or tunnel. 
 Only the local conditions will settle that. 
 
 Blocking Out the Ore Body. The preparatory works are 
 far from complete when the mineral body has been reached. 
 A flat coal-bed or a slightly inclined vein is treated, for the pur- 
 pose of mining, as a tabular mass of mineral which is to be divided 
 into blocks by haulage-ways. The latter are called galleries, or 
 gangways, in coal-mines, and levels or drifts in metal- mines. 
 The product from each block is delivered down grade to its gang- 
 way, whence it is hauled to the slope-shaft or tunnel to daylight. 
 These gangways are as numerous and driven as close together 
 as the requirements for exploitation will demand or the means of 
 the operator will permit. The more numerous they are the more 
 quickly can each lift above them be exhausted. On the other 
 hand, as they cost more to excavate than the corresponding 
 volume of material in the rooms, this character of work must 
 be confined to its lowest limit. Between these two conflicting 
 requirements, each mine presents its own local conditions. 
 
 Dead-work. The gangways, or drifts, partake of the nature 
 of explorations, and, being more expensive, and in many cases
 
 PREPARATORY WORK. 25 
 
 only tentative, are termed dead- work. By dead- work is understood 
 any bore-hole into the vein, auxiliary shaft, drift, or gangway 
 and the narrow work of connections between the rooms. Though 
 primarily unproductive,' dead work bears a vital relation to the 
 economy of the mine. The location of the gangway is a matter 
 of great importance. If the mineral is of varying grade of value 
 and soft, it is laid along the foot-wall. It is in the country rock, 
 near the vein, if the former is softer than the vein matter. This 
 enables the seepage and the drainage to be taken care of and 
 reduces the risk of injury from subsidence. 
 
 Lifts in Metal-mines. Metalliferous veins are blocked by 
 driving levels, or drifts, at distances of from 60 to 100 feet apart, 
 measured along the slope. These are started on either side 
 from the slope or from each cross-cut intersection with the vein 
 and carried along the vein. The casual connections made 
 between the adjoining levels occur only when the working of 
 a lower lift has carried that opening to the level above. The 
 ventilation requirements in metalliferous mines are not suffi- 
 ciently urgent to require the continuous circuit of air through 
 the various openings in the mine. The height of the lift between 
 the levels and their ratio to the thickness of the deposit depend 
 more upon the percentage of the mineral of the vein contents 
 than upon the method of mining. In soft ore the levels are 
 close together and the lifts not greater than 60 feet. In hard 
 rock, in thin deposits and in ore of low grade, the distance between 
 ihe levels is greater. The maximum, however, is 100 feet. 
 
 The Lifts in Coal-mines. The coal-bed is divided into lifts of 
 300 feet to 600 feet along the slope, the greater length for flat 
 beds. Double entries are started upon an inclination from 
 shafts,- or slopes, with such grade as depends upon the system 
 of haulage. These gangways are driven side by side in pairs 
 with an untouched rib of coal of at least 20 feet between them 
 (Fig. 7). Their direction may also be indicated, if not deter- 
 mined, by the cleavage planes of the . coal. One entry in each 
 pair of gangways serves as the intake airway to its lift of coal, 
 and the other for the outgoing air-current.
 
 26 MANUAL OF MINING. 
 
 In thick and steep coal-seams the haulage-way is usually 
 built near the floor to facilitate the landing of the cars and the 
 parallel returning airway similar in area, above it or nearer the 
 roof. From these gangways the rooms are driven upward on 
 the slope, at distances apart of about 40 feet, to the gangway 
 above. 
 
 The Dimensions of the Horizontal Opening. The dimen- 
 sions are governed by the sen-ice which the tunnel, gangway, or 
 adit is to give. They are not unnecessarily large, for, in addition 
 to the great cost of excavating, there is added the additional cost 
 of supporting the increased opening and the expense of main- 
 taining it. If it is to serve as a ventilating way only, its area is as 
 great as the requirements of the ventilation systems demand, 
 the height and width depending on the special conditions. If it 
 is to be a haulage- way, a single track is provided, unless the num- 
 ber of cars passing through it for a given output demand a double 
 track. The latter would probably be used only for an endless- 
 rope system of haulage. The height of the gangway is made 
 just sufficient for the passage of men through it and only as much 
 more as is required to furnish the area necessary for the ventila- 
 tion current, provision having been made for the probable inter- 
 ference with the current by moving cars. The width must be 
 sufficient to provide ample clearance space beyond the cars for 
 the safety of the men walking along the way. 
 
 Reserves. The number of blocks opened at a time is only 
 enough to maintain a regular output. Whatever the method 
 may be, it aims to concentrate the men as much as possible, while 
 giving them working places of ample size to break down the 
 mineral over as large an area of free face as possible, to reduce 
 the length and cost of gangways to a minimum, leaving them 
 open only as long as needed, and robbing the supports as 
 promptly as the rooms are abandoned. 
 
 The distance to which the gangways are carried beyond the 
 shaft or slope depends upon the relation between the cost of 
 their maintenance and of the haul to that of sinking an entirely 
 new entry some distance. A mile is as far as ordinary condi-
 
 PREPARATORY WORK. 2J 
 
 tions of haulage allow; beyond that, a new opening is neces- 
 sary. But to maintain a constant output for the demand of the 
 market, the levels, or gangways, are kept in advance of the imme- 
 diate places of work. Their revelations furnish a guide for 
 future action. This is particularly desirable in precious-metal 
 mines, the value of whose mineral contents is so variable. These 
 advance blocks, together with those of exceptionally high grade 
 or of medium quality, previously developed, constitute the mine 
 reserves. When the vein rock is soft and requires considerable 
 timbering for its maintenance, this character of development 
 work is not carried far beyond the work of mining. 
 
 Pillar Supports for Mines. All permanent ways in horizon- 
 tal deposits are protected from the crush of the roof to secure a 
 safe outlet. Shafts are surrounded by pillars of untouched vein 
 matter or coal, whose dimensions depend upon the depth to 
 the bottom of the shaft; haulage- ways in coal have massive 
 pillars on each side undisturbed except for the narrow, short 
 openings connecting the rooms with the haulage- ways; and the 
 rooms are protected by pillars of undisturbed coal on either 
 side. These reserves afford abundant support and are undis- 
 turbed until the working places they guard are to be abandoned. 
 In veins or nearly vertical beds where the wall pressure on 
 either side does not tend to close the rooms or the haulage-ways, 
 sufficient security is attained by horizontal timbers between the 
 walls above the haulage- way and by untouched blocks of mineral 
 alternating with the rooms (stopes). 
 
 The Influence of Cleavage Planes in Determining the Direc- 
 tion of Roads. All strata are more or less uniformly creviced, 
 with horizontal planes of growth and vertical joints, caused by 
 shrinkage. Coal masses are similarly divided by one or more 
 sets of planes producing rhombohedral coal. These cleavage 
 planes in coal, called "cleats," facilitate the breaking of the 
 mineral. The direction of the rooms in coal-mines is fixed by 
 the direction of the cleat, the important gangways being carried 
 with the cleat. In hard-coal regions, where explosives are ex- 
 tensively used, the cleat does not materially assist the breaking
 
 28 MANUAL OF MINING. 
 
 of the mineral, and as a consequence the rooms are not dependent 
 for their direction upon the cleavage planes. In steep bitumi- 
 nous seams cleat has less importance than grade for haulage-ways. 
 Faults. A fault is a break in the continuity of a vein or 
 stratum. During some convulsion of nature the strata were 
 broken and in the readjustment of position a vertical displace- 
 ment occurred. The plane of fracture may be empty or it ha< 
 been filled with extraneous matter. When encountered during 
 the operation of mining the fault plane or cross-course is pierced 
 and its strike and pitch noted. On the distant side, the char- 
 acter of the rock is examined. In a stratified country it should 
 be easy to identify the stratum and its geological position relative 
 to that of the ore-bearing stratum. Thus the engineer may be 
 guided up or down according to the direction of the throw. But 
 operations in massive rock assume a more serious aspect. On. 
 the distant side of the dike or cross-course no geological aid is 
 available. Hence, in attempting to follow the prolongation of 
 the vein beyond the fracture, the engineer must rely upon his 
 knowledge of the local conditions or upon the average of those 
 prevailing in similar ore-bearing fields. 
 
 Schmidt's Rule. It is a matter of record that 80 per cent 
 of intersected veins were heaved, apparently, to the right or left. 
 Those to the right are twice as many as those to the left. Hen- 
 wood also discovered that the heaving to the side of the greater 
 angle is five times as common as to the smaller angle. In every 
 district may be found a rule for finding the other end of the vein, 
 but it is purely of local application and unreliable. To formu- 
 late a general rule out of these numerous and apparently eccen- 
 tric displacements would seem well-nigh impossible; but Herr 
 Schmidt, in 1810, offered a solution to the problem, which, 
 though not infallible, is the best extant and has done valuable 
 service. " When the cross-course dips away, after going through 
 it, the drift is run along its far wall in a direction opposite to 
 that in which the vein pitches. If it dips toward the mouth, 
 the drift is carried along the far wall, to the right or to the left, 
 as the veins dip to right or left." The amount of the displace-
 
 PREPARATORY WORK. 29 
 
 ment, i.e., the distance to be drifted for the continuation of the 
 vein, cannot be premised. It varies between very wide limits, 
 and is thousands of feet in many localities Kenwood averages 
 the throw of veins at 16 feet. 
 
 Open Cuts. Open cuts are used to explore the outcrops of 
 veins and shallow mineral bodies They contemplate ' limited 
 explorations and a smaller output and life as compared with 
 that of a mine. They are also employed for stripping the sur- 
 face coverings of flat veins which are close to the surface. Quar- 
 ries are extensive open cuts for the extraction of iron ore, slate, 
 peat, building-stone, and such large bodies of soft minerals as 
 would be difficult to support by timbering in the usual methods 
 of underground mining. In quarries, large masses are dis- 
 engaged at a time and the operations are conducted upon a large 
 face at a comparatively small expense. The entire deposit can 
 be recovered to a moderate depth with a high degree of profit. 
 The point of attack for a quarry is the lowest consistent with 
 the requirements for transportation. It is operated in benches 
 and cut by mechanical picks or channeling-machines (Fig. 16). 
 Hoisting from the pit is accomplished by derrick and bucket 
 with a form of suspended cable-way. The pulsometer or a 
 centrifugal pump is used for drainage. The depth to which a 
 quarry can go is limited to, perhaps, 200 feet, beyond which 
 some more rational method must be introduced. It involves 
 some form of underground work which ultimately becomes far 
 more expensive than the mining which could have been adopted 
 prior to the operations of quarrying. Open cuts are always 
 to be deprecated where a subsequent underground work is to 
 be pursued. 
 
 The objection to quarrying is the damage which is done by 
 the influx and accumulation of surface-waters. These give 
 trouble to the miners and must be removed as promptly as possi- 
 ble to admit of sinking of the quarry as readily as the output 
 demands. Again, the difficulty of propping the sides of the cave 
 is no small matter, particularly if the deposit occurs between 
 pronounced walls with an overhanging side. The Tilly Foster
 
 30 MANUAL OF MINING. 
 
 mine was deepened only by the blasting away of 200,000 tons 
 of the overhanging wall. This character of quarry work is 
 dangerous and not economical, but is universally employed 
 for iron ores. In the case of the diamond ground of the Kim- 
 berley mines, the quarry was replaced by a method very closely 
 approximating the long wall for coal and caving of iron ore. The 
 surface-waters accumulating- in the open quarry, which found 
 their way into the mine, were constantly threatening danger, and 
 the operators finally dug tunnels around the quarry to catch and 
 divert the surface-waters. 
 
 The Steam-shovel. In soft minerals the steam-shovel with 
 its dredger derrick is employed for excavation purposes. Its 
 introduction in the iron-mines of Wisconsin and Minnesota has 
 enabled the marketing of iron ore of a very low grade at low cost. 
 The cost of quarrying, crushing, and administration of the iron- 
 ore deposits in the northwestern portion of the United States 
 and in British Columbia rarely exceeds 90 per cent per ton, and in 
 many cases is very much lower. Some of the copper deposits 
 of British Columbia are also of such phenomenal size that their 
 development by the steam-shovel has brought about the intro- 
 duction of machinery for hydraulic-, electrical-, and steam-power 
 purposes in one case up to 6000 horse-power. Air-machine drills 
 of 3^-inch size are used as auxiliaries where the rock is suffi- 
 ciently hard to demand their employment upon the immense 
 boulders which are interspersed with the softer rocks. The 
 quantity of earth which can be removed by a steam-shovel varies, 
 of course, with the supply of wagons which it has and the length 
 of their haul. But a 45-ton Bucyrus shovel having a bucket 
 carrying 1.75 cu. yds., excavated in a day about 1200 bank yards 
 from a cut whose depth was 8 feet and whose average width 
 was 22 feet. In this case automatic dump- wagons were employed, 
 and with a haul averaging 1500 feet each way. Three wagons 
 were filled per minute. 
 
 Hydraulic Mining. This -is a species of open work in which 
 water is the agent for cutting and transporting the gold-bearing 
 earth. From nozzles at the lower end of pipe lines, leading from
 
 PREPARATORY WORK. 31 
 
 reservoirs located at convenient points of great elevations, the 
 water is discharged at a high velocity against the bank of a stream 
 or an old river-bed to undermine it and to wash the material into 
 sluices conveniently placed to receive the flow. The sluices are 
 built on an inclination, with cleats on the floor and of a width and 
 depth sufficient to carry the volume of water and its suspended 
 material. The later installations of hydraulic plants use centrif- 
 ugal pumps (Fig. 122). In two or three steps the water is raised 
 from the creek and delivered to the bank at high pressure and 
 velocity. 
 
 Hydraulic mining is inexpensive because of the small amount 
 of labor involved. Four men can operate the entire plant, 
 hence the process can be conducted at low cost and with a high 
 degree of efficiency. Earth carrying 6 cents in gold per cubic 
 yard is regarded as " pay-dirt." The total cost of operation is 
 $0.082 per miner's inch. About 2^ to 3 cu. yds. are moved per 
 miner's inch of water used. 
 
 A long sluice is unnecessary. All the coarse gold, and nearly 
 70 per cent of the total fine gold, recovered is caught in the first 
 section of boxes, about 12 per cent being obtained in the second 
 and an insignificant amount in the fifth section. The grade of 
 the boxes is 5 inches per foot. 
 
 The Duty of Giants and Lift-tips. The following table 
 shows the distances which the standard nozzles can throw a stream 
 of water together with the quantity of water which they can handle 
 under the stated pressures. 
 
 Salt-mines. The getting of salt is generally by a special 
 process. It is always found in old river-beds and frequently 
 quite thick. The process of mining follows one of two methods. 
 Pure thick beds are mined systematically by a method of squares 
 or like similar bodies of soft coal, due regard being had to sound- 
 ness of roof and solidity of the mineral. When the salt body is 
 thin or somewhat impure, a different method is employed. Holes 
 are drilled to bed-rock, water poured down and the rock-salt 
 bleached out. A pump-pipe line is extended to the floor level 
 and the brine raised to the surface.
 
 MANUAL OF MIXING. 
 
 DUTY OF GIANTS. 
 
 Diameter of 
 
 Nozzle. 
 Inches. 
 
 
 Pressure at Nozzle. 
 
 Pressure in pounds 
 Head in feet 
 
 3 
 69-3 
 
 40 
 92-4 
 
 5 
 "5-5 
 
 60 
 138.6 
 
 70 
 IOI-7 
 
 80 
 184.8 
 
 90 100 
 207.9! 231 
 
 < 
 
 Gallons . . 
 
 U4 
 
 9 
 62 
 
 *55 
 109 
 
 70 
 
 173 
 126 
 
 94 
 
 189 
 142 
 1 08 
 
 % 
 
 121 
 
 219 
 168 
 13 1 
 
 232 1 245 
 178 | 186 
 140 148 
 
 Horizontal distance 
 Vertical distance 
 
 I* 
 
 Gallons 
 
 170 
 93 
 63 
 
 196 
 
 "3 
 81 
 
 219 
 
 132 
 97 
 
 240 
 148 
 
 112 
 
 259 
 163 
 125 
 
 277 
 175 
 137 
 
 294 
 186 
 148 
 
 3i<> 
 193 
 157 
 
 Horizontal distance 
 Vertical distance 
 
 ii 
 
 Gallons 
 Horizontal distance 
 Vertical distance 
 
 210 
 96 
 63 
 
 242 
 118 
 82 
 
 271 
 138 
 99 
 
 297 
 156 
 
 "5 
 
 320 
 172 
 I2 9 
 
 342 
 1 86 
 142 
 
 9 
 
 154 
 
 383 
 207 
 164 
 
 i? 
 
 Gallons 
 Horizontal distance 
 Vertical distance 
 
 253 
 100 
 
 64 
 
 293 
 124 
 
 83 
 
 327 
 146 
 
 100 
 
 118 
 
 387 
 l8 4 
 133 
 
 413 
 
 200 
 
 I 4 6 
 
 439 
 213 
 158 
 
 462 
 224 
 169 
 
 ii 
 
 Gallons 
 
 313 
 104 
 
 64 
 
 361 
 130 
 83 
 
 404 
 
 X 52 
 101 
 
 442 
 176 
 119 
 
 478 
 
 *95 
 135 
 
 5" 
 
 212 
 149 
 
 54i 
 224 
 163 
 
 571 
 240 
 
 175 
 
 Horizontal distance 
 Vertical distance 
 
 
 if 
 
 Gallons 
 
 425 
 
 no 
 
 65 
 
 490 
 140 
 85 
 
 548 
 
 161 
 104 
 
 600 
 1 88 
 
 121 
 
 649 
 
 2IO 
 138 
 
 604 
 2 3 
 '54 
 
 735 
 2 43 
 169 
 
 775 
 259 
 183 
 
 Horizontal distance 
 Vertical distance 
 
 2 
 
 Gallons 
 
 556 
 
 116 
 
 65 
 
 642 
 
 147 
 86 
 
 717 
 166 
 
 105 
 
 785 
 194 
 124 
 
 849 
 
 218 
 141 
 
 907 
 240 
 158 
 
 962 
 255 
 J 74 
 
 1014 
 274 
 189 
 
 Horizontal distance 
 Vertical distance 
 
 A nozzle if inches diameter, under 207.9 ^ eet head, can throw 
 439 gallons cf water per minute 158 ft. high or 213 feet distant. 
 
 Peat is recovered by dredging channels usually cut in the 
 heavy bogs deep enough to serve for drainage and for navigation. 
 The peat is then cut in blocks. When it occurs above water the 
 fuel is quarried in steps as are building materials.
 
 PREPARATORY WORK. 33 
 
 REFERENCES. 
 
 The Genesis of Certain Auriferous Lodes, J. R. Don, A. I. M. E., Vol. 
 XXVII, 564; The Telluride Ores of Cripple Creek, Kalgoorlie, T. A. Rick- 
 ard, A. I. M. E., Vol. XXX, 708; The Greatest Gold-producing Mines, J. H. 
 Curie, Eng. & Min. Jour., Nov. 7, 1903; The Geologist in Matters of Prac- 
 tical Mining, J. E. Spurr, Eng. & Min. Jour., April n, 1903; Mining Geology, 
 G. J. Binns, N. Z. Mines Rec., Dec. 16, 1901; Our Coal Resources and Con- 
 sumption, Profs. Hall and Jevons, Jour. Gas Light., Feb. 20, 1900; The Dis- 
 covery of New Gold Districts, H. M. Chance, A. I. M. E., Vol. XXIX, 224; 
 Connellsville Coal Region, W. C. Irvin, E. & M. Jour., Mar. 24, 1900; The 
 Ducktown Copper Mining District, S. W. McCallie, Eng. & Min. Jour., 
 Oct. 4, 1902; Some Notes on the Nome Gold Region of Alaska, F. C. Schrader 
 and A. H. Brooks, A. I. M. E., Vol. XXX, 236; The Coal-mining Industry 
 of the United Kingdom, R. A. S. Redmayne, Engineering Mag., Dec. 1903. 
 
 The Prospecting Density Rule, J. H. Pollock, A. I. M. E., Vol. XXIX, 
 281; The Equipments of Camp Expeditions, C. H. Snow, A. I. M. E., Vol. 
 XXIX, 157; Bibliography of Mexican Geology and Mining, R. A. Santillan, 
 A. I. M. E., Vol. XXXII, 605; A Glossary of Spanish American Mining and 
 Metallurgical Terms, A. S. Dwight, A. I. M. E., Vol. XXXII, 571; Histori- 
 cal Sketch of Mining Legislation in Mexico, E. M. Baco, A. I. M. E., Vol. 
 XXXII, 520; A Synopsis of the Mining Laws of Mexico, R. E. Chism, 
 A. I. M. E., Vol. XXXIII, 3; Mining Rights, Foreign: Germany, Austria, Coll. 
 Guard., Dec. 4, 1896, 1063; Laws as to Corporations, etc., in U. S. Domain, 
 Coll. Guard., July 1897, 105; Mining Gold Ores in California, general article 
 State Min. Bureau, zoth Report, 1852; Tunnel Decisions in Colorado, and 
 Editorial, E. & M. Jour., May 22, 1897, 515. 
 
 Our Coal Resources and Consumption, Profs. Edward Hall and Stanley 
 Jevons, Jour. Gas Light., Feb. 20, 1900; Development of the Connellsville 
 Coal Region, W. G. Irvin, Eng. & Min. Jour., Mar. 24, 1900. 
 
 Placer Mining, Illustrated, by A. Lakes, Coll. Guard., Eng., May 1896, 219; 
 Mining Methods in Peru, S. of M. Quart., XV, 219; Hydraulic Mining, 
 Strength of Beams for Flumes, Min. & Sci. Press, May 29, 1897, 452; A 
 Modern Dredging Plant, Min. & Sci. Press, Mar. 24, 1900; Hydraulic 
 Sluicing in Collingwood District, Logan, N. Z. Mines Rec., Feb. 16, 1900; 
 and Hydraulic Mining, E. B. Wilson. 
 
 Cleavage Development, Journal Geology, IV, 4 pages, 444. 
 
 Colliery Surface Arrangements, Ernest H. Thomas, Coll. Guard., Mar. 
 30, 1900. 
 
 A Modern Dredger Mining Plant, Min. & Sci. Press, Mar. 24, 1900.
 
 CHAPTER III. 
 
 METHODS OF MINING. 
 
 Exploitation. The plan pursued for winning minerals and 
 making them available for use is termed exploitation,, while the 
 general term mining is applied to the labor of excavation. 
 
 For the purpose of exploitation mineral deposits may be 
 classified according to their shape and inclination into those 
 which are of irregular shape and those which are tabular in 
 form. The former are mined by open cut, quarry, stripping, 
 or the hydraulic process, with or without the aid of a dredger or 
 steam-shovel, as they are usually superficial in position and 
 their contents friable or soft. The latter are composed of com- 
 paratively hard material and are extended in a horizontal or 
 vertical direction with some degree of uniformity. Their con- 
 tents are recovered by some system of Underground Mining. 
 
 The character and depth of the superficial covering may 
 determine the choice of method for some deposits which occur 
 in tabular masses. The relative convenience and cost of remov- 
 ing the cover, and the increased expense of supporting it above 
 the subterranean works, constitute the decisive element. Open- 
 cut mining may be profitable when the quantity of superficial 
 material to be stripped from the surface does not exceed one 
 half of a foot depth for each foot of thickness of material in the 
 deposit, and where the quality of the cover is such as can be easily 
 removed by the steam-shovel. 
 
 Open-cut Mining. The method of extracting the iron ore 
 in the Mesaba range in Minnesota as adopted by the Iron Mining 
 
 34
 
 METHODS OF MINING. 35 
 
 Company consists in stripping the surface material by a steam- 
 shovel of the dipper type. This machine has a radius of action of 
 15 feet and excavates and delivers the dirt to cars of the stand- 
 ard railroad gauge. A cut is first made for a certain distance and 
 width in which the track is laid to admit the machine. The 
 steam-shovel excavates the dirt on one side of the cut, delivering 
 it on the other side, and stripping the surface for the depth of cut 
 to the boundaries of the property. The shovel is returned to its 
 starting-point, and from the side cut where the dirt has been 
 stripped a new track is laid and the work proceeds, while cars 
 are loaded on the first track. The stripping advances over 
 the entire area to the level of the track. Several steam-shovels 
 strip in a similar manner until the ore body has been exposed. 
 The usual depth of a cut is 20 feet; beyond that there is danger 
 of the bank caving and causing injury to men and machinery. 
 Occasionally deep holes are drilled and chambered into the bank 
 ahead of the machine, into which several kegs of powder are 
 loaded and blasted for the purpose of loosening the ore for the 
 work of the shovel. 
 
 The attack upon the iron ore proceeds in the same manner 
 as in the surface material in benches of a width of 30 feet, con- 
 taining one track for cars and a short section of track for the 
 machine. The product per shift of ten hours is 1800 tons, 
 though there are records of a daily product of 3544 tons. 
 
 The cost of stripping and wasting the mineral of a coal-bed, 
 under the most favorable conditions, is approximately 20 cents 
 per cubic yard, whether machines are employed or hand labor 
 alone is utilized for extraction. The sole advantage in the intro- 
 duction of machinery is the rapidity with which work can be 
 prosecuted and the increased area of stripping which is possible. 
 These are the determining features in situations where the cover 
 is thick and the working season short. Inasmuch as the rail- 
 road approach to the several benches on which operations are 
 conducted must also be lowered with the progress of the work, 
 it soons happen that its grade becomes excessive and the traffic 
 is interfered with unless a corresponding increase in the length
 
 36 MANUAL OF MINING. 
 
 of the approach be made. This necessitates a deep open cut 
 for extensive work, or a tunnel through which the railroad cars 
 may be drawn to the works. 
 
 The Milling System. When a limit has been reached for 
 economical depth in the quarry an underground process may be 
 introduced in combination with the open cut. This is termed 
 milling, and has been practised to advantage in the Lake Superior 
 iron region. A mill is a hole made in the rock or ore connecting 
 an upper with a lower level of the deposit, through which material 
 is sent down to the level below. In the system mentioned above 
 the tunnel is driven large enough to admit railroad cars into the 
 deposit, which, at distances of two car-lengths apart, is connected 
 with the surface by mill-holes sunk from the surface or from the 
 level above previously worked by steam-shovels. At the bottom 
 of the mills are chutes for loading into cars the ore blasted down 
 at the surface and delivered through the mills, to be thence hauled 
 out of the tunnel. 
 
 This system of mining is very successfully loading a train of 
 twenty-five 2O-ton cars per hour. 
 
 Stripping Anthracite Coal. A lenticular coal basin at Latti- 
 more, Pennsylvania, is attacked by stripping and underground 
 work. The depth at the centre of the deposit is about 100 feet. 
 After stripping the earth cover from the coal a slope is sunk 
 on the foot-wall of the basin to the centre or lowest point of the 
 deposit. Levels are driven along the strike, and rooms turned 
 right and left, which are worked upward to the surface, the coal 
 being allowed to fall to the level below. Here it is loaded, trammed 
 to the slope, and hoisted out of the mine. The product of the 
 mine is limited only by the facilities for transportation. 
 
 Underground Mining. The deposits having a tabular form 
 are of two types: beds, or seams, occurring nearly horizontal, 
 and veins whose inclinations approach the vertical. The former 
 extend laterally over a considerable territory and under conditions 
 which are quite uniform throughout, permitting a system of 
 mining to be promptly selected. Veins present so many eccen- 
 tricities which increase with the progress of the work that the
 
 METHODS OF MINING. 37 
 
 early operations are somewhat tentative until the property has 
 been thoroughly explored. 
 
 Deposits of coal, iron ore, and minerals occurring in the stratified 
 rocks are, with rare exceptions, flat. They have a greater regu- 
 larity of thickness, inclination, and character of contents than do 
 vein deposits in the metalliferous regions which are encased in 
 hard rock, occupy a vertical position, and are subject to changes 
 not only in dimensions and depth, but even in the character of 
 their mineral. The brittleness of the two minerals is markedly 
 different. Coal must be delivered in large sizes for the market, 
 while metalliferous ores are just as valuable when pulverulent as 
 in the lump. The entire contents of the coal-bed can be utilized 
 and are salable, and hence are extracted as completely as 
 possible. On the other hand, in the precious metal deposits 
 not one tenth of the contents between the walls is of sufficient 
 value to warrant hoisting to the surface; the remaining nine 
 tenths, being valueless, must be stowed away. These radical 
 differences in the character of the deposits require distinct systems 
 of extraction for coal and for ore, and hence they will be treated 
 separately. 
 
 In one respect only is the mining of coal and the mining of 
 other minerals conducted alike. The operations progress to the 
 rise or upon the slope except on exceedingly rare occasions, 
 when precious metals may be worked downward. 
 
 The- thickness and pitch of the mineral deposit, the thickness 
 and nature of its cover, the direction and its outcrop, the mechani- 
 cal and physical character of its rock or mineral having thus 
 been ascertained, the place of opening and the character of the 
 entry into the mine are selected according to the principles dis- 
 cussed in the previous chapter. The directions of the gang- 
 ways to be used for haulage and of the rooms for economical 
 work are next determined in connection with the system to be 
 adopted. After this the vein or bed is blocked out and is ready 
 for mine operations. 
 
 Systems of Mining : Beds. According as it may be desirable 
 or necessary to support the stratum above the mineral bed or
 
 38 MANUAL OF MINING. 
 
 allow it to fall, the mineral should be removed in two distinct 
 operations, or should be extracted in its entirety at once. The 
 first is the pttlar-and-room system, in which chambers or rooms 
 are cut out at various points, isolated from one another by 
 untouched portions of the bed, and the mineral from them 
 extracted on the advance from 'the shaft or entry toward the 
 boundaries. The remainder of the deposit is attacked on the 
 return. This system is applicable where the roof is sufficiently 
 strong to be left with a partial support. 
 
 The second, longwatt system, contemplates the extraction of 
 the entire contents of the bed in one operation. It is employed 
 when the roof is treacherous or weak and cannot be economically 
 secured. No effort is then made to support it. By this plan 
 the entire width of the coal property is attacked simultaneously; 
 or mining may progress circumferentially and outward to the 
 boundaries, removing the entire mineral contents; or all coal 
 may be removed while progressing in the opposite direction 
 from the boundaries to the shaft. These methods are known as 
 the longwall advancing and longwall retreating systems. 
 
 Naturally the simplest method of attacking the mineral 
 would be to complete a series of breasts as wide as the nature 
 of the roof or the walls would permit, and opening them when 
 the main headings have been carried to the extreme boundaries, 
 leaving the mineral intact except for the necessary roads. The 
 breasts are mined out as rapidly as possible on the return home- 
 ward toward the shaft. This plan of mining retreating represents 
 the maximum degree of safety for miners, who are then working 
 in solid mineral. It is economical for operators, who may recover 
 the entire deposit. The cost of mine timber is far less, and the 
 product per acre greater, than by mining outward from the shaft. 
 But the plan requires patience and a large capital, particularly 
 if the area to be mined is large or the royalty to be paid is great. 
 It cannot be unconditionally recommended. 
 
 Systems of Mining Veins. The systems employed in recovering 
 mineral from veins are essentially the same as for beds with 
 such modifications as are due to the increase in pitch. The
 
 METHODS OF MINING. 39 
 
 rock to be supported is the vein matter itself, while the rock adjoin- 
 ing the vein exerts little pressure upon the vein matter. Hence 
 the character and strength of the vein matter determines the 
 choice of the system. In strong mineral and hard coal the pillar- 
 and-room system is adopted. In metalliferous districts the 
 rooms for the first excavations are called stopes, and the system 
 is a sloping system. The longwall system as applied to veins 
 consists in the attack of their contents over a broad area and 
 allows the friable materials to cave without restraint. The broken 
 mineral contents are then removed and replaced by additional 
 material caved from above. This continues until the lift has been 
 exhausted. This caving system is much in vogue in iron mines. 
 
 Two modifications bearing resemblances to both of the typical 
 systems are also in extensive use, one being known as the square- 
 set system and the other as the filling system. In these the area 
 of attack is broad, but the roof is not allowed to cave. The exca- 
 vations are filled as rapidly as it is possible to supply the material 
 for support. In the square-set system a massive cribwork of 
 timbers is built into position. In the filling system foreign 
 materials are lowered into the mine from above, spread and 
 packed into the rooms immediately behind the miners who are 
 engaged in extracting mineral. 
 
 Systems of Mining Coal. The two typical systems employed 
 for extraction of coal are the pillar-and-room and the longwall 
 previously described. The longwall system is employed for 
 thin seams of coal^ or for thick seams divided into thin layers by 
 partings, and having regular thickness, a uniform pitch, weak 
 roof, soft floor, and the bed deep. The inclination of the beds 
 for the longwall system must not exceed 10. 
 
 For the pillar-and-room method the vein may be thick, the 
 material for filling and gob may be absent, the vein shallow, 
 the coal soft, and the seams irregular in thickness and eccentric 
 in pitch. 
 
 As not all of the conditions favorable for either system exist 
 in a given mine, the choice must be indicated by the compara- 
 tive merits of the majority of favorable conditions of greatest
 
 40 MANUAL OF MINING. 
 
 importance from an economical standpoint. No system can 
 entirely satisfy all the conditions; hence a compromise inevitably 
 results in adopting the least objectionable system, with such 
 modifications as have been evolved from local conditions or are 
 based upon the structural difficulties encountered in the given 
 mine. The widths and directions of the rooms or the stalls for 
 attack,* the mode of recovering the mineral from the pillars, and 
 the process of filling and packing the excavations vary in locali- 
 ties and mines and cause a departure from the typical method 
 in some cases to such an extent that the identity is almost lost. 
 
 The Longwall Advancing System. This system consists in 
 driving gangways from the shaft, leaving around the latter an 
 ample pillar for security (Fig. i). When its boundaries have 
 been reached the gangways are extended, the miners proceeding 
 to the right and left from them, cutting the coal while advancing 
 from the shaft and throwing behind them such waste and debris 
 as are produced during the operations. No effort is made to 
 support the roof beyond that which is necessary for the pro- 
 tectipn of the men for ten yards back of the face. The gang- 
 ways serving as roads are carried with the progress of the work 
 by building pack-walls out of the material obtained from the 
 debris. They are advanced radially. Each length of working 
 face is a stall. When two miners are assigned to a stall of 
 about 30 feet each, the gangways or haulage roads serving two 
 stalls 60 feet long are not over 60 feet apart. This neces- 
 sitates building additional roads to reduce the miners' length of 
 haul. 
 
 The numerous stalls extend in one continuous straight line, 
 as in Figs, i and 4, or a curved line, as in Fig. 2, without 
 any corners or breaks. The roof either bends, yields, and closes 
 upon the excavation, or it breaks freely and fills the space which 
 is known as the gob. The subsidence of the roof is communicated 
 to the overlying strata and will often extend hundreds of feet 
 
 * The reader will find in the appendix the definitions of the various technical 
 terms employed in this chapter and elsewhere throughout the book.
 
 METHODS OF MINING.
 
 42 MANUAL OF MINING. 
 
 toward the surface. Hence the bed must be deep or the strata 
 firm, if the surface land is not to be injured. 
 
 The Scotch System. A system of mining known as the 45 
 system, and by some as the Scotch system of longwall, is shown 
 in Fig. 2. The main entries aa are at right angles to each other, 
 being turned from the right and left from the upcast shaft c. The 
 gangways bb are turned at an angle of 45 from the main entries. 
 
 FlG. 2. The Scotch Longwall System. 
 
 The shaft pillars p are left untouched ; the dimensions of each are 
 discussed in Part II, Chapter I. The face of the main entry is 
 advanced with a width of 30 feet, allowing 12 feet between the 
 pack -walls, or 9 feet on each side. Miners turning from right or 
 left of the head of the entry break the coal over its entire face and 
 advance outward, shipping the mineral through the roads. The 
 pack-walls are extended on either side as fast as the face pro- 
 gresses. The roads are so driven as to leave from 120 to 180 
 feet of face, according to the nature of the room. 
 
 Cutting the Coal. The process of mining the coal is shown in 
 Fig. 3 at a, where a groove is cut in the floor or near the floor
 
 METHODS OF MINING. 43 
 
 by pick or by machine. Two rows of props bb are placed parallel 
 to the face and at such distances from it as to give ample support 
 without interfering with the work of the miner. The sprag c 
 supports the corner of the underhold coal while the miners are at 
 work. The sprags are removed, allowing the roof to release the 
 coal as soon as the length of the stall has been undermined. No 
 powder is used, the pressure of the roof being employed for the 
 
 FIG. 3. A Cross-section at the Longwall Face. 
 
 purpose. The slate and debris are thrown on the gob behind 
 the men; the back line of props is withdrawn and advanced in 
 front; the track is advanced in sections parallel to the face; and 
 the operation is repeated for another 3 to 5 feet according to 
 the depth of the groove. If the roof is weak, little space is left 
 for the track and the gob-roads are nearer together, the stall 
 being shorter. 
 
 When the inclination of the coal-seam exceeds 10 the com- 
 ponent of the vertical pressure of the overlying strata transmitted 
 by the roof to the coal parallel to the bed of the vein causes a 
 side pressure or "swing" which tends to produce slip in the coal. 
 This is serious, and if the curved face of the stall will not provide 
 for it the system must be abandoned. 
 
 Usually the mining is conducted by day, while the construction 
 of the pack- walls and the driving of gob-roads proceeds, at night. 
 If it is necessary to operate in two shifts, the stalls are worked 
 alternately from the main entries. 
 
 The Demerits of the Longwall System. The success of the sys- 
 tem depends upon the nature of the roof and its behavior when 
 the coal is removed. It should be brittle and weak. A tougb
 
 44 MANUAL OF MINING. 
 
 flexible roof is unsuitable and may require blasting down to 
 relieve the pressure upon the coal. The roads which are main- 
 tained through the excavations are called the gob-roads, and here- 
 in lies the serious objection to this system. The fracturing of 
 the roof which is the essential element of success tends also to 
 close the roads, increase the expense of mining, and endanger the 
 lives of the men. The roof and floor are continually being cut into 
 to maintain sufficient height for the travel in the roads. The length 
 and number of the latter increase with the progress of the work, 
 and the cost of their maintenance becomes a serious one. If 
 the boundary is a great distance away, the expense of keeping 
 them open may be so great as to necessitate a discontinuance of 
 the system. The liability of the material in the gob to spon- 
 taneous combustion and fire is an additional serious danger. A 
 fire once begun cannot be extinguished without the entire removal 
 of the waste. It is impossible to make the pack-walls air-tight. 
 Ventilation presents difficulties, for the numerous roads must 
 receive separate air-currents, while the working face may easily 
 be ventilated by a single continuous current. 
 
 Longwall Retreating. This is considered a better and safer 
 system, as the roof pressure is supported by the solid coal while 
 the work is proceeding toward the shaft. The waste produced 
 in mining is allowed to accumulate behind the men where the 
 roof closes upon the gob. There is no danger of any section of 
 the mine being closed by collapse, nor is there any fear of spon- 
 taneous combustion. 
 
 In Fig. 4 is shown a system of opening the mine by longwall 
 retreating. The four pairs of entries aa are started from the 
 downcast shaft and driven at right angles to one another to the 
 boundaries between them; entry pillars cc are left about 40 feet 
 thick with breakthroughs at intervals of 100 feet apart. When 
 the entries have reached within 300 feet of the boundary gangways 
 dd are driven, and at right angles to them the face entries ee, at 
 intervals of 200 or 300 feet. From their heads the stalls are turned 
 off to the right and left, opening a continuous face of coal. Mining 
 is then commenced. The rails are laid and the haulage-ways
 
 METHODS OF MINING. 45 
 
 mm moved from the entries ee, where they are no longer of service. 
 Work is simultaneously commenced in the haulage-ways nn to 
 prepare for the continuance of the retreat from the gangways 
 dd as soon as the operations have been completed on the face 
 entries ee. 
 
 The main haulage-ways are in the return airways to the 
 upcast and hoisting shaft u. Entries crossing the intake air- 
 ways are carried over by air-bridges, Chapter XIV. 
 
 FIG. 4. Longwall System of Mining Retreating. 
 
 Occasionally, when the roof is very weak, the gangways are 
 driven singly and connected by cross-cuts 60 feet apart instead 
 of being driven in pairs. This method, however, is productive 
 of a greater amount of narrow work and a greater expense in 
 ventilation. 
 
 The Combination Longwall System. In Fig. 5 is illustrated 
 a combination of the retreating and advancing systems. A 
 portion of the bed above the bottom of the shaft is worked by the 
 advancing system, while that below is driven by the retreating 
 system. The seam is 16 feet thick, but -does not prove very profit-
 
 46 MANUAL OF MINING. 
 
 able because of a considerable expense incurred in driving the 
 entries while planning the works. However, when properly ex- 
 ecuted it will relieve the roof, though there is always danger of 
 a suspension of its subsidence which interferes with the " swing, " 
 and also in a collapse along an unexpected line. 
 
 The mining to the rise is conducted advancing in section A, 
 while in section B the operations are in retreat. 
 
 FlG. 5. A Combination of Retreating and Advancing Longwall Systems. 
 
 Comparative Merits of Longwall Retreating and Longwall 
 Advancing. In the matter of ventilation the retreating system 
 is more perfect than the advancing system, as it can be regulated 
 to better advantage. No danger need be apprehended from the 
 evolution of gas from the waste. The cost of mining is the same 
 in both systems, but the expense of making pack-walls and of 
 opening and maintaining the roads is far greater than the in- 
 creased cost of driving the narrow gangways through the solid 
 coal. The large initial outlay in longwall retreating for the four 
 pairs of main entries out to the extreme boundary of the property, 
 and for laying tracks the entire length, and the long period of
 
 METHODS OF MINING. 47 
 
 waiting for the returns, constitute the chief obstacle to its more 
 ready acceptance. But its aggregate advantages warrant its 
 employment under all circumstances where the longwall system is 
 at all applicable. 
 
 The Nottingham or Barry Modification is employed for 
 mining splint coal of five feet in thickness with a strong roof and a 
 clay floor. In this the gangways are 6 feet wide in the solid coal, 
 with 2o-foot pillars between them and the cross-entries 200 feet 
 apart. From these are driven flats 120 feet apart and 200 feet 
 long, the coal being attacked "long horn," Fig. 9, and to the rise. 
 The cars travel along the working-face. Eight miners are em- 
 ployed on each stall of 120 feet, and produce large coal at low 
 cost, with an excellent yield per acre. 
 
 In one mine, changing from room-and-pillar to longwall, the 
 coal was opened by butt-entries on both sides of the gangways until 
 their faces met, the stalls being 48 feet wide. The coal was 
 strong and had a regular cleat overlaid by a slate top with a hard 
 floor. It was blasted, not undercut, and 37 tons per keg of 
 powder were obtained. 
 
 Pillar-and-room System. The second system of mining coal 
 consists in driving levels and opening long narrow rooms, from 
 which the coal is extracted while advancing from the shaft, and 
 in leaving pillars on either side of untouched coal which are sub- 
 sequently recovered on the return. The rooms are frequently 
 termed stalls and breasts, though the latter term is more cor- 
 rectly applied to the face of attack. When the rooms are very 
 broad compared with their length, the term chambers is some- 
 times applied to designate them. The pillars left for support 
 are also called ribs. 
 
 This system is a survival of antiquated methods and still pre- 
 vails to a greater extent in the United States than elsewhere, 
 rather because of local custom than from any special merit 
 which it possesses. It is employed where the material for filling 
 or gob is scarce. A great diversity exists in the various localities 
 in the details of the plan. The height of the lifts, the direction 
 of the rooms, their dimensions, the forming of pillars, the rob-
 
 48 MANUAL OF MINING. 
 
 bing of pillars, the taking down of roof, etc., vary widely. 
 Many of its modifications are due to physical conditions of the 
 coal and roof, though the latter elements do not always receive 
 the consideration due them. The facility with which the details 
 of working may be varied to suit the irregularities of pitch 
 and thickness, the rolls in the strata, etc., renders it especially 
 acceptable in regions of geological disturbances. 
 
 The necessity for economy in the conservation of the mineral 
 is not imposed upon the American operator. The question of 
 output per acre is not of vital importance, and mining as a con- 
 sequence is conducted here in rather a loose manner. Indeed, 
 with the exception of culm-flushing there has been no improve- 
 ment introduced in the anthracite exploitation during the past 
 sixty years, though numerous mechanical improvements have been 
 utilized to economize the handling and preparation of the coal. 
 
 The Typical Room-and-pillar System. As stated in the pre- 
 vious chapter, the height of the lifts varies with the thickness and 
 inclination of the seams. But this determined, the gangways 
 are driven, after which the rooms are turned as fast as practicable 
 to the rise or up the slope at an angle depending upon the pitch, 
 as in Figs. 6 and 7, or on the mode of conveying the coal from 
 the breasts to the gangways. The jaws or necks, aa, of the 
 rooms B are about 6 feet wide for a distance equal to that allowed 
 for the stump-pillars A, when they are suddenly enlarged to a 
 working-face of 20 or 30 feet, as time and the condition of the 
 roof will allow. The rooms are driven regular and uniform 
 to a length of from eight to ten times their width, leaving a 
 chain-pillar for support to the upper level. The rooms of each 
 lift are opened progressively to the boundary as fast as circum- 
 stances will permit. 
 
 The cost of- driving narrow entries and headings aa and cc 
 is so high that modifications are introduced whereby the jaws aa 
 are made as short as possible and widened out to the full face of 
 the room, thus reducing the narrow work without weakening the 
 stump-pillars. The pillars P left between the rooms are unbroken 
 except for three or four small breakthroughs dd, through which
 
 METHODS OF MINING. 4^ 
 
 the return ventilating air-current passes from room to room or* 
 its way to the upcast. 
 
 Rock Pressure. Assuming the specific gravity of rock to be- 
 
 2.6, then a block having a base of i square inch and a height 
 of 12 inches will weigh i pound. A column 500 feet high will 
 exert a pressure upon its base of 500 pounds per square inch. 
 Each square foot of rock at that depth below the surface will.
 
 MANUAL OF MINING 
 
 I
 
 METHODS OF MINING 51 
 
 receive a pressure of 72,000 pounds per square inch. The load 
 which a pillar of coal will sustain is therefore proportional to 
 its depth below the surface and the area of its base. A coal- 
 pillar 25 feet wide and 200 feet long at a depth of 500 feet will 
 receive a total pressure of 180,000 tons. To sustain this load 
 without suffering injury it must have a strength at least equal 
 to two or three times the load thus imposed upon it. 
 
 The Strength of Coal. Experiments recently conducted upon 
 the crushing strength of coal have obtained values ranging 
 between 1600 pounds per square inch and 2100 pounds per square 
 inch. The value for the ultimate strength of coal having been 
 determined in a specific case, it is possible to determine the 
 dimensions of a pillar required to sustain a given load trans- 
 mitted to it from the roof. 
 
 Owing to the wide range of values obtained for the crushing 
 strength of coal, it is advisable that each engineer determine the 
 specific value for the coal of his locality, determining, thereupon,, 
 the dimensions of the pillars to be used in connection with this 
 system of mining. Sample cubes should be selected and sawed 
 from various portions of the seam, and their strength normal to 
 the bedding ascertained by actual test. They should be pro- 
 cured from the breasts rather than the pillars. 
 
 Influence of Cleats upon the Resistance of Coal. It wa& 
 shown in the previous chapter that the direction of the cleavage 
 of coal may determine the direction of driving the rooms and' 
 roads. The cleavage may also determine the resisting strength 
 of the coal. In Fig. 8 is illustrated a block of coal with its hori- 
 zontal bedding cleavage aa, and, extending from roof to floor,, 
 the vertical cleavage bb, as well as the butt-cleats cc, at right 
 angles to both. In all coals this prismatic structure is pronounced 
 in one or more directions, and along these planes a slip may occur 
 before actual crushing ensues. The load which the coal can 
 sustain is therefore less than its actual strength. 
 
 In Fig. 9 are shown arrows indicating the directions of attack 
 upon the lines representing the plane of the main cleavage bb. 
 The butt-cleat, line cc, presents usually a hackly appearance
 
 MANUAL OF MINING. 
 
 when exposed, and can be readily distinguished from the main 
 cleats, which are perfectly smooth. Adhesion along the plane of 
 the former is stronger, and hence, when conditions of haulage do 
 
 FIG. 8. The Cleats in Coal. 
 
 FIG. 9. Directions of Attack upon 
 the Cleats. 
 
 not prevent, most mining is conducted in the direction represented 
 l>y the arrow c at right angles to the main cleat, known as " face- 
 on work"; an advance in the direction of the arrow e is "end-on 
 work" mining; and in the direction of arrow d, "half end-on 
 
 FIG. 12. c c> 
 
 The Directions of Cleavage. 
 
 work." The mining of coal at an angle greater than 45 and 
 less than 90 is said to be "long horn," arrow g; and at an 
 angle of less than 45, arrow /, is called "short horn." 
 
 In Fig. 10 is illustrated the columnar structure of coal. Across 
 the bedding a it is stronger than the cleat &, represented by the 
 broken lines. In the horizontal position a is weaker than b (Fig. 
 IT). In the inclined position (Fig. 12) neither cleavage plane 
 is in its original position. The cleavages receive their respective 
 components of the thrust c, the component a acting normal to the
 
 METHODS OF MINING. S3 
 
 bedding planes or across the main cleat, while the component 
 e is parallel to the main cleat. The amount of thrust falling 
 normally upon the bedding plane is less as the inclination increases. 
 
 Size of Coal-pillars. The width of the coal-pillar depends 
 upon the strength of the coal and the depth of the rock, rather 
 than on the strength of the roof. When the rooms are open each 
 pillar must support not only the rock immediately over it, but also 
 sustain the pressure transmitted to it from the roof over the room. 
 A rigid and firm roof will transmit its entire pressure equally to 
 the sides. Hence the coal, at a depth of 500 feet, in the pillars 25 
 feet wide and 200 feet long between rooms 20 feet wide, will 
 receive a pressure equal to the weight of a mass 35 feet wide, 200 
 feet long, and 500 feet high, or about 252,000 tons. This corre- 
 sponds to 50.4 tons per square foot. So long as this is less than 
 the breaking strength of the coal the pillar will remain intact. If 
 the crushing strength be 2000 pounds per square inch, the stress 
 upon the coal is less than half its ultimate strength. The margin 
 or factor of safety is 2.8. This factor will be further reduced if 
 the planes of main cleavage are pronounced and weak. 
 
 At a depth of 1400 feet there remains no margin for safety, 
 the limit of the coal resistance being reached with its relation of 
 room to pillar. Below this depth the pillars must be wider or the 
 rooms narrower if the system is to be pursued. When a roof is 
 brittle it will break without transmitting its pressure to the pillar. 
 The pillar is relieved of some pressure, in which event its width 
 is not necessarily increased. A soft floor requires a broad pillar 
 to distribute the thrust over a large area of the floor. Otherwise 
 a heaving of the floor ensues, which disturbs the tracks and 
 " creeps" over the territory, closing rooms and airways. 
 
 The Relative Dimensions of Pillars and Rooms. The width 
 of the pillars depends upon the character and length of service 
 they are to perform. Safety demands large pillars, but inasmuch 
 as only a portion of their coal contents is recoverable during the 
 second working, a very large percentage of the total in the tract is 
 lost. A rapid production demands wide rooms which can be 
 mined by machinery and exhausted quickly, but they require a^
 
 54 MANUAL OF MINING. 
 
 good roof and, in turn, broad pillars to support them. The rela- 
 tion between the two must be determined by local conditions, 
 observing the principles mentioned above. Narrower pillars with 
 wide rooms are decidedly wasteful of coal, and in a five-foot seam 
 with ribs of 12 feet and rooms 36 feet, none of the pillar coal is 
 recovered, whereas in a 6-foot seam with 1 2-foot rooms and 
 28- foot pillars nearly 90 per cent of the coal contents are sold, 
 particularly if panel pillars are also employed in sections of the 
 mine, as illustrated in Fig. 17. The usual proportion of pillar to 
 room is about 3 to 2 for the average depth of mining as pre- 
 vailing at present in America. 
 
 As an illustration, it will be found on calculation that in a 
 mine opened with rooms of 12 feet width between pillars of 18 
 feet a greater concentration of labor is obtained, and there are 
 positive sources of economy over that possible if the rooms and 
 pillars alike are 20 feet wide. Along a mile of gangway 176 
 rooms can be opened in the former and but 132 rooms in the 
 latter design. Though the aggregate breast length is less, there 
 will be more rooms for simultaneous work. Or, for an equal 
 number of working-faces, there will be a concentration of work 
 by the narrower pillars and rooms, which saves 3520 feet of track 
 for the 132 rooms; haulage will be less; and 44 cars can be ob- 
 tained in the same time from three quarters the distance of that 
 when the rooms are spaced 40 feet apart. In robbing the pillars 
 more coal will also be recovered from the relatively thicker pillars. 
 
 The width of pillars is increased and that of the rooms decreased 
 as the depth of the vein is great. The limit to an economical 
 application of this system is reached when the rooms are reduced 
 to an unprofitable portion of the deposit. At a depth of- 700 
 feet they contain less than 30 per cent of the total contents of 
 the vein. At 1600 feet depth the pillars leave only 25 per cent of 
 the coal for the rooms; while at 2000 feet. the rooms are one 
 fourth as wide as the pillars, and are capable of producing not 
 more than 15 per cent of the total coal. 
 
 The general impression that coal-pillars and beds of steep 
 pitch are less strong than those in flat beds is untrue unless a
 
 METHODS OF MINING. 55 
 
 deformation of the strata has produced a fissility in it. Their 
 resistances are equally great and their life as great, being dimin- 
 ished only by a protracted period of pressure and by alterations of 
 temperature and humidity which injure the integrity of the pillar. 
 This is independent of the inclination of the bed. 
 
 There is an economical limit to the narrowness of the rooms, 
 based not upon physical laws but upon an artificial law by 
 which a working- face of less than 15 feet is considered in many 
 coal-fields as "narrow work," making the cost of driving such 
 rooms greater per foot of length than that of wider rooms. More- 
 over, narrow rooms do not admit of an improvement of machine- 
 cutters or large gangs of loaders, and therefore increase the cost 
 of coal. 
 
 The Mode of Delivering the Product from the Face to the 
 Entry. The most important problem in mining is the method 
 of transporting the mineral from the working-face to the entry or 
 gangway. It usually determines the plan of operations. When 
 the vein is vertical or nearly so the mineral may be dropped 
 through mill-holes or chutes to the cars in the level below (Fig. 87), 
 or lowered by means of a self-acting incline (Fig. 115). In the 
 anthracite region the working-face is termed a breast. Here the 
 broken coal is allowed to accumulate in the batteries (Fig. 13), 
 which are loaded at the coal face and more or less completely 
 emptied at the foot from a platform, the flow of mineral being 
 controlled by a gate. So long as the pitch exceeds the natural 
 slope of broken rock, 40, the coal will roll or slide on the rock 
 bottom without the necessity for lining the latter. The sides 
 may be built of plank spiked to the props, reaching diagonally 
 from the floor to the roof, and the lower portion closed by a 
 gate or door with a platform. By keeping the chute or battery 
 full of coal and emptying only the excess through the gate, the 
 loss from attrition will be slight. This, of course, is not pos- 
 sible with bituminous coal. At a grade flatter than 40 the 
 chute must be floored with wood or iron, but below 18 soft 
 coal will neither slide nor roll. If the working-room is large, 
 the breasts wide, or the vein thick, the chutes are carried up with
 
 56 MANUAL OF MINING. 
 
 the work, one on each side of the waste filling. In small rooms 
 the chutes are built centrally with the manway compartment 
 alongside of them (Fig. 13). 
 
 Beds having an inclination between 18 and 6 from the hori- 
 zontal are the most difficult to provide with an economical method 
 of transportation. They are too flat for self-acting systems and 
 too steep for any simple plan of haulage. An endless- rope way 
 plan for each room is too complex, and any scheme involving 
 shovelling would be both expensive and injurious to the coal. 
 
 FlG. 13. Battery Mining of Thick Steep Seams. 
 
 In such event buggy- roads are resorted to for the rooms (Fig. 6). 
 The track is laid in sections directly upon the floor, or upon 
 the trestle raised at the lower end to moderate the grade, accord- 
 ing to the headroom that may be afforded and the pitch of the 
 seam. Another method consists in having jig-planes where the 
 loaded car in its descent pulls up a weight which, on the next 
 down trip, raises the empty car back to the face. 
 
 Breasts driven up a grade of less than 10 are said to be flat, 
 and the accessory haulage from them presents no special diffi- 
 culty. The rooms may be driven directly up the rise for the 
 small angles, or, when the pitch reaches the limit, are laid so 
 as to secure a satisfactory grade. In such cases mules may be 
 utilized for the power, or in exceptional cases man-power is 
 employed.
 
 METHODS OF MINING. 57 
 
 Rooms in beds as flat as 3 are driven directly up on the pitch, 
 employing mules or men for transportation of the cars. 
 
 The rooms are rarely worked downward in the dip, unless the 
 pitch be very slight. For a short distance an undulating seam 
 may require this modification. 
 
 Robbing the Pillars. The rooms, having been carried to their 
 length, are then turned and "butted off," the pillar at the top 
 driving along the under side of the chain-pillar. The pillars 
 are then attacked and their contents removed as soon as possible, 
 before their coal contents have suffered in quality from pressure. 
 The operation is performed as rapidly as possible for safety, and 
 the pillars are taken in successive retreating. Their sides are 
 scaled off as much as can be secured before the pillar crushes. 
 The heading may be driven off the centre of the pillar and some 
 of the middle portion of the coal obtained, leaving two thin sup- 
 ports for the roof; or the pillar may be worked over its entire 
 width from the chain-pillar toward the gangway. 
 
 The stump- and the chain-pillars are not disturbed until the 
 lift which they support is to be abandoned, when their contents 
 are promptly extracted and the roof allowed to cave. 
 
 Fig. 14 is the plan of a flat seam operated as described 
 above, the two sections of he property being separated by the 
 double broken lines. From the outcrop at a and b the main 
 and return headings are carried 60 feet apart. The cross-entries 
 are 360 feet apart and the rooms 300 feet long. The outer irreg- 
 ular line is the outcrop. The fan is located at c. The break- 
 throughs d are midway in the rooms and on one line to serve 
 as an auxiliary haulage-way during the robbing of the pillars. 
 The jaws of the rooms are 10 feet wide and 15 to 20 feet long. 
 The room is widened in a direction away from the outcrop. 
 Pillar-robbing begins at the outcrop. This process, being slow 
 and decayed till all rooms are mined, results in an excessive 
 waste of coal. The speedy returns, good ventilation, and safe 
 work commend it. 
 
 Comparison of Longwall Advancing with Pillar-and room 
 System. The relative advantages and disadvantages of these
 
 58 MANUAL OF MINING. 
 
 two systems are easily stated. In the former the expense of 
 maintaining roadways is higher, the number of accidents 
 greater, the product of round coal much larger, the product 
 
 JJJJJJJJJJJJJJJ 
 
 FlG. 14. A Double-entry System in a Flat Seam. 
 
 per acre higher, the ventilation of the face simpler, the amount 
 of narrow work less, and the consumption of powder less. On 
 the other hand, the latter system is more advantageously em- 
 ployed where faults, horsebacks, and dikes may be encountered,
 
 METHODS OF MINING. 
 
 59 
 
 if the upper strata are wet, if the boundary is at a great distance, 
 if the mines are shallow, or if the surface land is very valuable. 
 Modified Pillar-and-room Systems. In some fields the 
 cross-entries are driven 360 feet apart and the rooms are turned 
 from each entry, making them about 180 feet long; in such event 
 the cross-entries are carried up the slope independent of the 
 direction of the cleavage. The only feature commending this 
 plan is the reduction in the amount of narrow work and the 
 yardage thus saved. 
 
 FACE HEADINC 
 
 BUTT ENTRIE 
 FlG. 15. Connellsville Method. 
 
 A modified double-entry system employed in the Connellsville 
 coke district drives gangways 1000 feet apart (Fig. 15), and 
 four pairs of headings from them at distances of 400 feet. The 
 rooms are turned when the latter entries have gone 1040 feet and 
 are butted off at the chain-pillars, after which the pillars are 
 robbed. The coal being soft, the headings and the jaws of the
 
 6o 
 
 MANUAL OF MINING. 
 
 rooms are driven 8 feet wide, the pillars 30 feet, and the stump- 
 pillars between gangways and headings 52 feet. 
 
 The double-room modification giving satisfaction in the 
 Southern States for seams of 5 feet thick with weak roof is shown 
 in Fig. 16. The main and cross entries are driven in pairs, and 
 later the rooms are started right and left 42 feet wide with pillars 
 of 36 feet; their length is 180 feet, but they are not opened fully 
 until the cross-entries have reached their boundaries. Props 
 
 FIG. 16. The Double-room System. 
 
 are set along the inner side of each roadway for the gob, and 
 waste is thrown into the centre of the room. When they have 
 reached their extreme length the rooms are widened and one 
 half of the pillar on the floor withdrawn retreating. When the 
 succeeding rooms have reached their limit each is widened in 
 turn, the remaining half pillar on one side and one half the 
 adjoining pillar on the other side are cut away. Props set along 
 on either side of each roadway are removed or broken during 
 the retreat, in order to facilitate the uniform fall of the roof.
 
 METHODS OF MINING. 6 1 
 
 Mining Steep Thick Seams of Coal. The coal in a steep bed 
 may be attacked in steps proceeding upward to the top of the 
 given lift, the excavated space being filled with rock obtained 
 from elsewhere. In such an event provision must be made for 
 planked chutes to deliver the coal without loss or admixture 
 with the waste rock. The broken coal may be allowed to accu- 
 mulate in the excavated spaces as a partial support to the roof, 
 and also for the men while mining the coal overhead. If the 
 rooms are not too wide, the coal will not suffer in quality by the 
 time the room is, exhausted. The other system of mining thick 
 seams of coal contemplates its removal over its entire length, 
 proceeding downward from the top of the lift or from the super- 
 ficial covering of the seam and allowing the material above to 
 cave upon the work as it progresses downward. These methods 
 resemble somewhat the two longwall systems as applied to the 
 flat thin beds. 
 
 The famous mammoth bed of the anthracite fields of Penn- 
 sylvania presents the greatest difficulties to the operator. It 
 varies in pitch between wide limits and within a small area 
 attains to a thickness of no feet. 
 
 A Room System for Gaseous Mines. In Fig. 17 is shown 
 the Brown panel system employed in thick seams of coal where 
 there is considerable gas and the roof poor. At distances along 
 the gangway of 180 feet are turned double rooms up on the pitch. 
 These are 24 feet wide, separated by a pillar of 15 feet in thick- 
 ness, which latter is withdrawn as soon as the rooms have reached 
 the upper airway. Midway between each pair of rooms is driven 
 a central heading 12 feet wide to the full length as a travelling 
 road and airway, as well as a coal-chute. From this, to right and 
 left at intervals of 30 feet, are turned similar headings carried 
 through the coal until the double rooms are reached. The 
 coal is mined from the rooms toward the central heading as 
 indicated in the figure, and in such order as to leave flanking 
 pillars as shown. Small cars convey the coal to the chute. A 
 heavy barrier pillar separates each pair of rooms, and is untouched 
 until the boundary of the property has been reached. The sys-
 
 62 
 
 MANUAL OF MINING.
 
 METHODS OF MINING. 6j 
 
 tern affords safety to the men, but is expensive because of the 
 large amount of narrow work performed before much coal is pro- 
 duced. 
 
 The County of Durham System is a combination of the panel 
 and the pillar- and- room methods. The breasts with their pillars 
 are laid in groups of eight or ten from the gangway, each section 
 or group being mined separately and systematically toward the 
 boundary. Between the sections are left barrier-blocks of coal 
 150 feet thick, unbroken except for airways. These give com- 
 plete isolation to the section and also localize whatever move- 
 ment may occur in the falling of the roof. As the coal in these 
 barrier-pillars suffers in quality from the crush of the roof, their 
 location is selected where the coal is of poor quality. They are 
 not touched until all of the adjacent sections have been exhausted. 
 This method is safer than the pillar or room in gaseous mines, 
 and more economical than the panel method. 
 
 Anthracite Mines. The mining of anthracite presents condi- 
 tions that are not usual in other fields. The contortions and 
 folds to which the strata have been subject during the processes 
 of mountain-making have developed so many eccentricities of 
 pitch that it is difficult to plan the works a great distance in 
 advance. For example, in a shaft-mine in which the workings 
 may have continued horizontally for some time, the coal might 
 suddenly rise or dip and change the entire topography, and hence 
 necessitate an entire revision of the system. The system gener- 
 ally adopted in the anthracite region is the pillar-and-room on 
 the double-entry plan. Levels are driven each way from the 
 shafts sunk to intersect one or more beds. The coal is invari- 
 ably worked from the lower toward the higher level, and the pillars 
 are shot through, leaving chain-pillars as a solid support for the 
 level above. The width of the rooms and of the pillars is not 
 materially affected by the pitch of the beds, though, at times, 
 elaborate timbering may be necessary to safeguard the men and 
 prevent the coal from running into the gangways. 
 
 Rock-chute Mining. In the Wyoming and Lacka wanna 
 valleys, where several coal-seams are worked simultaneously,
 
 64 MANUAL OF MINING. 
 
 special care must be exercised that the rooms in the different seams 
 are laid off in such manner as to leave their pillars in the same 
 vertical line. When the beds are thick and separated by partings 
 of slate of but a few feet, or the several beds are thin and separated 
 by a few yards of rock, they are mined simultaneously, all being 
 connected by a chute built from the lower to the upper bed for 
 delivery of the product to the lower seam. 
 
 Systems of Mining the Mammoth Coal-seam of Pennsylvania. 
 The system followed at Hazelton, Pennsylvania, in mining the 
 Mammoth bed at a point where it is 40 feet thick, with an inclina- 
 tion at tunes of as much as 60 to the horizontal, is shown in 
 Fig. 13. This is known as a single-chute breast method. The plan 
 shows the gangway a heavily timbered through the coal, and a 
 chute b leading from the battery platform to the gangway. The 
 battery c is made of open timbers hitched into the floor and roof, 
 backed by horizontal timbers. In it is a door of planks capable 
 of being removed when desired for loading the cars. A chute 
 in the middle of the room receives the coal, and manways, dd, 
 on either side, made by leaning posts diagonally from the floor 
 to the rib of coal, provide air and entry for the men. These man- 
 ways are often termed " juggler- ways. " A breakthrough, e, is 
 made between the rooms for ventilation, and the main airway, /, 
 is driven over the gangway, as usual, near the roof. Every third 
 or fourth room is connected with the intake and return airway, 
 and in extreme cases every two rooms are also connected. The 
 amount of coal which is withdrawn from the battery is only suffi- 
 cient to relieve the battery of the excessive volume, sufficient 
 being allowed to remain for the miners to stand upon while engaged 
 in work. 
 
 A modification of this single- chute breast method consists 
 in providing each room with two batteries, the manways dd 
 serving as chutes, the space between them being used for refuse. 
 This is the plan employed when a large amount of bony coal and 
 slate is present in the seam. 
 
 Propping through a Weak Floor and Roof. At the Richard 
 mine, where both the floor and the roof are weak, a variation, as in
 
 METHODS OF MINING. 6c; 
 
 r-> 
 
 Fig. 1 8, has been introduced. The props are placed 6 feet 
 apart, with their length such as to drive them through the weak 
 shale to the solid rock beneath the floor and likewise above the 
 roof slate. By this plan an ample support is given while the; 
 pillars are being extracted. 
 
 FIG. 18. Variation of System for Pillar-robbing. 
 
 The Room System with Caving. In the North Staffordshire; 
 mines is a system similar to the one mentioned above, in which 
 the gob follows downward with the mining and requires no atten- 
 tion. This has also been employed in California, and resembles 
 closely the caving system as practised in the iron-ore mines of 
 Lake Superior, as described later on in this chapter. 
 
 A Room System with Filling. At the Shamrock colliery,; 
 Westphalia, is practised another system of mining coal at 45, the : 
 bed having a thickness of seven and one half feet. Here the bed isi : 
 divided into blocks 2000 feet long on the strike and 600 feet ort 
 the pitch. Each block is subdivided into three levels (aa, Fig: ig), 
 200 feet apart, through which is driven the central heading a for
 
 66 MANUAL OF MINING. 
 
 ventilation. On each side of the pillars are self-acting inclines, 
 or coal-chutes, ee. 
 
 This method contemplates the filling of the excavation made 
 by the removal of the coal by such material as can be obtained at 
 the surface. The central headings aa serve for lowering the 
 material used for storage. At the point c is a small shaft in which 
 is accumulated the filling material and such rock as is obtained 
 in taking up the floors of the haulage-ways. From this point is 
 
 FlG. 19. A Room-and-pillar System adopted in Westphalia. 
 
 obtained the material which is delivered to the rooms through the 
 central heading aa when needed. 
 
 The plan of the mine provides three working places on each 
 side of the central heading, and during operations three of them 
 are being filled, while those on the opposite side are in the process 
 of excavation. This alternation of operations maintains regular 
 pressure of the roof. 
 
 The coal is removed by a process of overhead stoping (Fig. 27), 
 delivered to the levels a, and thence by the inclines e to the main 
 levels. On either side of the central heading is left the stump- 
 pillar, which is eventually withdrawn when the level is to be
 
 METHODS OF MINING. 6/ 
 
 abandoned. Its lower end is very carefully blocked to prevent 
 the leakage of air from the ventilating current. 
 
 Mining Thick, Steep Coal-seams with Rooms. The method of 
 mining shown in Fig. 20 is described in the Coal- and Metal-Miners" 
 Pocket-Book as having originated in California. The coal-seam is 
 7 feet thick and has an inclination of 60. The gangway a is driven, 
 along the strike 40 feet from the airway 6; breakthroughs, cc, at 
 distances of 30 feet connect the two, acting first as airways and 
 
 FlG. 20. A Calif ornian Variation of the Pillar-and-room System. . . ;Q. 
 
 later as mill-holes or chutes down which to lower the coal. The 
 door or gate at the foot of each chute permits the coal to be loaded 
 directly into the cars on the gangway. Counter airway-chutes 
 dd are driven over 30 feet apart at an angle of 35 for delivering 
 the coal to the chutes, cc, by gravity. Numerous breakthroughs, 
 40 feet apart, divide the coal and the lift between each two pairs 
 of levels into pillars. When a panel of five or more chutes has 
 been driven the pillars are robbed on the retreat from one cor- 
 ner at the top. This is possible only in comparatively strong 
 coal without partings. 
 
 Square Work. Square work is a modification of the pillar- 
 and-room system, in which very thick beds, not only of coal but 
 also of salt gypsum, puzzolana, etc., may be mined. Rooms
 
 $8 MANUAL OF MINING. 
 
 are opened from the gangway 150 feet square, with pillars 30 
 .feet thick between them. 
 
 Two sets of cross-galleries 20 feet wide are driven through 
 'the rooms, leaving pillars about 25 feet square for the support 
 -of the roof. They are driven as high as the vein-matter will 
 allow. If the roof or floor is not firm, a layer of mineral may 
 ie left for greater security. The objection to the method is in 
 the difficulty of obtaining systematic ventilation. Moreover, 
 <very little of the pillars is recovered, and the system involves great 
 risk to life, since the roof is usually too far for the eye to detect 
 any threatening disaster. 
 
 Flushing. About 15 per cent of the coal obtained from the 
 anthracite mines is of such a nature that it requires special treat- 
 ment, which may not warrant the expense of attempting to 
 recover it. 
 
 Moreover, in breaking and sizing the coal for the market a 
 large quantity of fine coal is produced and slate sorted which 
 were formerly dumped and accumulated in large piles, encum- 
 bering the surface and increasing the pressure below. These 
 culm piles are now being utilized for the filling of the excava- 
 tions by a process known as flushing. In Scranton, Pa., a stream 
 of water is discharged against the culm piles, washing the mate- 
 rial into a pipe-line and down holes driven from the surface 
 to the underground workings. The mouth of the rooms being 
 boarded up at the gangway in such a manner as to drain off 
 the water without allowing any culm to escape, the room becomes 
 filled with wet culm. The culm packs quickly and solidly and 
 offers a degree of solidity almost equal to that of the untouched 
 -coal in the pillar. The pillars are then attacked on either side 
 with safety and their contents almost totally recovered. The 
 roof is then allowed to close in the space and rest on the floor and 
 culm. 
 
 Metal-mining. The location of the point of attack of a metal- 
 jnine involves considerations which were discussed in Chapter II. 
 The comparative merits of the shaft driven on the wall side 
 <of the vein, and of that intersecting the vein, as illustrated in
 
 METHODS OF MINING. 
 
 69 
 
 Figs. 22 and 23, were likewise reviewed. The determination 
 of the system of mining depends largely on the capital avail- 
 able and the value of the deposit. In precious-metal mining 
 
 FlG. 21. Following Vein by Incline. 
 
 the mineral cannot be worked from level to level as rapidly 
 as the levels are pushed forward, for it is difficult to estimate 
 in advance the quantity of "pay ore." Hence for economical 
 
 FIG. 22. 
 
 FIG. 23. 
 
 The Two Positions for a Shaft Relative to the Vein. 
 
 work only sufficient ore is extracted to pay the expense of 
 operation until the limit of the property has been reached. 
 This assumes on the part of the operators a degree of patience 
 which is not usually found in speculative districts. Under ordi-
 
 70 MANUAL OF MINING. 
 
 nary circumstances it is desirable or may be necessary to obtain 
 immediate returns. In Fig. 24 is shown the method pursued 
 in blocking out the mine. It represents a longitudinal view of 
 the vein. Through the vein-rock extend streaks of ore of 
 high grade which are neither uniform in dimensions nor in pitch. 
 These may be encountered at any place in the vein-matter. In 
 such cases the risers a are driven, or winzers b are sunk, in such 
 of them as promise profit. If the ore is uniform in grade and 
 value, the mill-holes are spaced uniformly. These serve to 
 develop the property, enabling the future to be calculated with 
 
 FIG. 24. A Vein Blocked Out, showing Ore-shoots. 
 
 some degree of certainty. The mill-holes serve for ventilation 
 as well as ore-chutes. Risers are cheaper to open than winzes, 
 and are preferred in wet mines, though they are at a disadvantage 
 in the hot mines. 
 
 Systems of Mining Metalliferous Veins. The precious metals 
 occur usually in thin veins of hard rock, and constitute a small 
 portion of the deposit, and the systems adopted for their extraction 
 are known as underhand or overhand sloping. 
 
 In overhand stoping the mineral is recovered by upward 
 workings in a series of steps, the barren rock being stowed away 
 in the lower portion of the lift. In underhand stoping the ore 
 is mined out in steps working downward, while the waste rock 
 is disposed of on platforms built to receive it.
 
 METHODS OF MINING. 71 
 
 For thick vertical deposits of iron, copper, lead, zinc, or 
 other ores, three systems are in vogue, varying with the strength 
 of the ore. The square-set and filling systems are used in 
 large deposits and veins for removing the entire contents, whether 
 they be friable or strong. The caving system is designed for ore 
 in large masses friable enough to break and fall when undermined. 
 The order of proceeding in all these systems is upward. A 
 descending system is employed when a heavy cover can be utilized 
 as a flexible roof, which is allowed to cave with the downward 
 progress of the excavation. 
 
 The Stoping Systems of Mining. After the vein has been 
 blocked out, then such blocks as are deemed workable are mined 
 by one of two methods, overhand or underhand. In the first 
 method the miner picks or shoots down the ore in front of and above 
 him, advancing the breast as high as he can reach and as wide as 
 the vein and as far as the limit of the block. In underhand work 
 the miner removes the mineral below him and advances the breasts 
 longitudinally, with the load taking them successively downward. 
 These two methods are applicable to steep veins if the overhand 
 working is quickly adjusted in the thick flat beds. In the former 
 method they are merely divided from the falling rock, as gravity 
 facilitates all operations of breaking down the vein-matter; this 
 is a much quicker, cheaper, and less arduous method than that of 
 underhand, as the miner is compelled to raise the mineral before 
 the next lower can be attacked. 
 
 The Underhand-stoping System. This is a method of limited 
 application. When the vein has been blocked out and the several 
 blocks on any given level are attacked, they are worked away in 
 horizontal slices, beginning at the winzes and extending half way 
 on either side to the adjoining winzes. The miner standing on the 
 floor of the level (Fig. 25) removes the vein-matter for a depth 
 of nearly 6 feet, sorts out the ore, and throws upon the platform 
 behind him the waste. When these slices have been carried as 
 far as designed, the next lower is attacked, beginning at the winzes, 
 its ore being raised to the proper level and its gangue thrown upon 
 the platform behind the miner. The platforms are constructed
 
 7 2 MANUAL OF MINING. 
 
 AS rapidly as the miner advances, there being one for each hori- 
 zontal slice. If the entire contents of the vein from wall to wall 
 .are composed of pay dirt, the timbering may be dispensed with. 
 It frequently happens that instead of a single gang operat- 
 ing in a given block on one slice at a time, as many gangs are dis- 
 posed in the work as there are slices, the uppermost gang preceding 
 the one below, in each case by an amount sufficient to secure for 
 him the advantage of the firm footing on the ore. Thus the 
 mine progresses, as illustrated in Fig. 25, in which the block resem- 
 .bles a set of steps downward from the level above, whence the 
 
 FIG. 25. Underhand Stopes. 
 
 :name underhand stoping. Occasionally the winze is sunk to the 
 lower level at once, and each step is advanced as rapidly as possible, 
 the mineral from all the slices or steps being delivered downward 
 to the level below, instead of requiring to be raised. 
 
 This system of underhand stoping is economical and has 
 many advantages as compared with the overhand system to be 
 described later. Timber is used only for the support of the frag- 
 ments of rock and for building platforms. The ventilation is 
 good, and better drilling can be conducted by the men in under- 
 hand than in overhand. 
 
 The iron-ore deposits of the Eastern States are worked down- 
 ward, the vein being entirely removed except for the occasional 
 pillars left to support the walls. The deposits vary in thickness, 
 the average being probably 12 feet, and in pitch between 15
 
 METHODS OF MINING. 
 
 73 
 
 and 70. In Fig. 26 the shaft pillars ar^ at aa; those at bb 
 support the roof; at cc, the surface, and at gg, the track in the 
 upper level. At dd is the untouched ore. The latter is broken 
 down from the stopes, ee t with air-drills and loaded into cars 
 below. The levels are 60 feet apart. Each stope can furnish 
 
 ^xxx^x 
 
 FIG. 26. The Underhand Sloping of Iron Ore. 
 
 about 50 tons in a single shift. The hoisting-car track follows 
 the foot-wall except where it is too irregular, when the hollows 
 are bridged over or the projections cut away, as at ft, to obtain a 
 regular slope. 
 
 The Overhand Stoping. The ore in this case is mined from 
 below upwards, as shown in Figs. 27 and 28. The attack upon 
 
 FIG. 27. A Double-wing Overhead Stope. 
 
 the blocks of mineral begins at some point on the upraise, the 
 miners standing on the caps of the timbering in the gallery, as shown 
 on the extreme right or left in Fig. 27. The miners remove a
 
 74 
 
 MANUAL OF MINING. 
 
 horizontal slice for a height of about 6 feet, as far as desirable, 
 sorting out the ore and delivering it to the mill-hole, where the 
 chute discharges it into cars below. Meanwhile the waste rock 
 is piled on the gangway timbers behind the men. The next 
 upper s.lice follows, and is broken away in the same manner from 
 the mill-hole, the miners then standing on the waste rock from the 
 previous slice. If the vein can supply no waste rock, a temporary 
 
 """ 
 
 *w $ r 
 
 LOA/G/TVD/MAL SEC T/ON Of <S TOPE 
 
 FlG. 28. Longitudinal Section of Stope. 
 
 staging is placed conveniently for mining as is shown near the 
 top of the room in Fig. 28. The room in this case presents an 
 appearance of inverted steps, whence originates the name. In 
 the double wing stopework advances simultaneously from both 
 sides of the mill-hole, as in Fig. 28. 
 
 In shooting down the ore, sheet iron or boards are laid on
 
 METHODS OF MINING. 75 
 
 the floor to receive the mineral and prevent the pulverized and 
 friable ore from being lost in among the waste. 
 
 Comparison of Underhand with Overhand Systems. The 
 comparative merits of these systems depend largely upon the 
 amount and cost of timbering. From this point of view the 
 overhand has the preference, as requiring little or no timbering 
 for the reception of the waste. The underhand requires a plat- 
 form for each step in each room. If the vein is entirely salable, 
 there is then no waste, and very little permanent timbering 
 will be needed for underhand stoping. The underhand system 
 is accounted best for valuable ores because all rock must be 
 handled, and there is no occasion for any being lost. On the 
 other hand, as the ore is being trodden upon continually, it can- 
 not be used for brittle minerals. Overhand work affords some 
 advantages in regard to breaking down mineral. In a wet mine 
 the underhand stoping presents objections unless the mill-holes 
 and winzes have been sunk to the lower level. The overhand 
 method is practicable in a wider vein than is the underhand, 
 which is limited to a width of vein by the length and size of the 
 stulls which are available. 
 
 The Square-set System. When the vein of metalliferous 
 mineral is wider than 8 feet, the underhand-stoping method used 
 to introduce the timbers to reach from one wall to the other must 
 be exceptionally long and correspondingly stout, hence is very 
 difficult to handle with the conveniences available underground. 
 The method then employed may consist. in the substitution of a set 
 of frame-timbers (Fig. 258). This method of square-set timber- 
 ing is applicable to either hard or soft ore. It is equally applicable 
 to flat as to steep deposits, and succeeds just as well with soft ore 
 as with firm ore. 
 
 The vein is blocked out in the usual manner into lifts of 60 
 feet each by gangways or levels running longitudinally in the 
 vein, if its contents are of hard rock, or in the country rock if soft. 
 
 Along the gangways are alternate rooms and pillars, each of 
 the same width, from 20 to 40 feet long, and as wide as the deposit. 
 The breast of each room is divided into working-faces of about
 
 76 MANUAL OF MINING. 
 
 7 feet wide. The erection of the square set of timbering begins as 
 soon as possible, props and wedges being placed between the 
 timbers of the set and the roof or side rt>ck to prevent fall. With 
 the advance of the room additional timbers are laid in the frame 
 on the floor, at the side, in front and above ; thus, as the room is 
 being excavated to its limit, the framing assumes a form similar to 
 that in Fig. 258. 
 
 The square setting of this type is employed in mining cham- 
 bers or deposits when the ore is of a comparatively uniform 
 nature. It is most economical, for the steps can be worked on 
 all four sides. When the first sets are fully under way for a con- 
 siderable area, the first floor is stoped out and sets placed directly 
 over the sill-sets, care being taken to be at least two sets behind 
 those on the first floor, in order to keep sufficient ore on the floor 
 to work with while the sets are being placed in position. 
 
 The operations of mining may be confined to one level only, or 
 they may be prosecuted simultaneously from several levels working 
 across the vein and upward in each lift. The mining proceeds by 
 slices, ascending to the upper level, as in the systems of sloping, the 
 timbers of the respective sets being placed in vertical tiers. The 
 mining is also conducted in an order resembling longwall, in which 
 the entire face is attacked and the timbering placed at once, the 
 progress then being in every direction except downward. This 
 modification is favored where the mineral is of low value and very 
 soft. 
 
 When the rooms on a level have been exhausted and the timbers 
 carried to the upper level, work begins upon the intermediate 
 pillars in the same manner as practised in the rooms. 
 
 In Fig. 29 is illustrated the method of sloping with square sets 
 where the work is prosecuted from the hanging-wall side of the 
 deposit and carried across it. This plan assumes the hanging-wall 
 to be heavy and liable to cave or the ore soft. The construction 
 of the square sets, b, is then conducted as rapidly as sufficient 
 space is excavated to permit of it. The set c is next provided for 
 and erected, after which the ground above them is stoped out for 
 the high set d and the flat set g. The set e is next erected, as shown
 
 METHODS OF MINING. 
 
 77 
 
 at /, and the sets built up in succession until the hanging-wall 
 is reached. 
 
 To avoid any risk of movement, care must be exercised that 
 the posts be perfectly in the line of pressures and are securely 
 
 FlG. 29. The Square-set System, working from the Hanging-wall. 
 
 blocked by wedges at the points of contact with the hanging-wall. 
 Where there is any pressure the sets are reinforced by braces, as 
 at h. Occasionally ladders ii are carried up for the men, and 
 chutes j arranged for the delivery of the ore into the bin. All ore 
 is removed from the vein. The support for the roof depends upon 
 the timbering alone and upon its security. Care must therefore 
 be taken to prevent caves. 
 
 In Fig. 30 is an illustration of square- set timbering beginning 
 at the foot-wall. The rock is first cut for the sills a, which are
 
 78 MANUAL OF MINING. 
 
 placed in position, and the sills &, of extra lengths to secure a firm 
 base. Sometimes they are of especial lengths, as at d. Hitches 
 are cut directly under the posts of the set above and the short 
 posts c, the caps then having different lengths to simplify the 
 framing. 
 
 When the deposit contains considerable barren material 
 which can be employed as waste and thrown into the spaces 
 
 FlG. 30. The Square-set System, working from the Foot-wait 
 
 between the sets to accumulate and eventually assist the timbers in 
 sustaining the pressure, the sets at the level are reinforced by 
 auxiliary timbers g and h, and also lagged to keep the level open. 
 The mill-hole m is surrounded by a gob and serves as an airway. 
 At p are props used to temporarily support the ground and to 
 guard against accident during the operations of sloping, and the 
 single braces are introduced to prevent deformation of the frames. 
 In the Lake Superior region several of the iron-mines are 
 worked in this manner, the lower sets being subsequently filled 
 with rock or sand lowered into the mine from above.
 
 METHODS OF MINING. 
 
 79 
 
 Top-slicing and Caving. There is a system of mining which 
 has been successful in wide deposits where the cover can be induced 
 to break and follow the excavation from a higher level .to a lower 
 one. The mineral is removed from the top of the deposit in horizon- 
 tal slices 8 feet wide and 6 feet high, made by a series of horizontal 
 drivages timbered with the three-stick sets (Chapter III, Part II) 
 as the work progresses. On one side is the solid ore, on the other 
 is refuse of previous slices which had caved when the timbers 
 were removed. 
 
 The shaft is sunk in the country rock, and from it, at levels of 
 60 feet each, are driven cross-cuts to the deposits. The first cross- 
 cut is extended from one wall to the other, and from the centre of 
 it are driven levels at right angles to the outer boundary lines, at 
 intervals of 100 feet; on this level, risers are carried up nearly to 
 the cover, from which are run cross-drifts until the side walls of the 
 deposits have been reached. The horizontal slice is then removed 
 at the top parallel to the wall and the space timbered. Its length 
 on either side of the cross-cut does not exceed 50 feet. 
 
 In Fig. 31 is shown this system, in which a is the main level, 
 
 FIG. 31. A System of Top-slicing the Ore. 
 
 b the riser, d the cross-cut on top of the ore, and e the cave fol- 
 lowing the slicing. A riser is made 6 or 8 feet and divided into 
 two compartments, one as a chute and the other for a ladderway. 
 No effort is made to support the roof, and the cover is there-
 
 8o MANUAL OF MINIXG. 
 
 fore allowed to cave, the caving taking place as the men retreat from 
 the end of the walls toward the risers, and from the ends toward 
 the cross-cuts, and from walls toward the risers. 
 
 When the excavation has been carried 50 feet from the cross- 
 cut, the timbers are withdrawn, or cut or blown down, as the case 
 may be, in order to permit the roof to cover as soon as the work is 
 completed. Prior to removing the timbers, however, there are 
 placed on the floor, 3 or 4 long, 8- inch poles along the passages, 
 with split lagging or sawmill slabs across them. These support 
 the roof when the ground caves, and enable the timber-sets of 
 succeeding slices below to be placed without great difficulty. 
 
 In very treacherous vein-matter or under a bad roof, each 
 cross-cut has risers at stated intervals that take the place of the 
 level for removing the ore to the main level. The remainder of 
 the work does not differ materially from that described. 
 
 Subdrifting and Caving System. This system originated at 
 the Brotherton mine in Lake Superior, and has been adopted 
 with excellent results in many wide soft-ore mines. Cross-cuts are 
 driven from the hanging- to the foot- wall, where they connect with 
 longitudinal levels. At intervals of 50 feet risers a (Fig. 32) are 
 driven 50 feet upward to the level above. The first six feet are 
 cribbed and a couple of driven sets put on over them. Subdrifts 
 c are then driven each way to meet the corresponding drifts from 
 the adjacent risers. The riser meanwhile is continued another 
 6 feet upward, when timbers for a second subdrift d and above 
 them for the third subdrift e are laid, until finally the riser reaches 
 the level above at a distance of 50 feet. This divides the space 
 between the two levels into three subdrifts, each having a top or 
 back of ore above them 6 feet thick. The ore from subdrifts is 
 delivered through a riser, and shipped to the cars on the lower level. 
 
 Following the completion of the subdrifts the capping of ore 
 above the level g is next removed by dividing the ore into blocks 
 of 10 feet square, beginning at the end of the level g. This weakens 
 the pillars, which crush and permit the removal of the ore. That 
 above the subdrift e is caved in a similar manner, followed by d 
 and c.
 
 METHODS OF MINING. 
 
 Si 
 
 Occasionally it is necessary to support temporarily the timbers 
 above the ore, as at h, by long timbers, but in all cases all the close 
 timbering is employed to control the pressure and prevent accident. 
 In this method all timber can be lowered from the levels above 
 and handled without interference with haulage. Ventilation is 
 good, and simple means of escape are provided in case of accident. 
 
 FIG. 32. Subdrifting and Caving. 
 
 Slicing and Filling System. This system is practised in wide 
 deposits, and consists in taking two slices of mineral in successive 
 squares from a lower to a higher level and filling up the excavation 
 with rock or other material quarried for the purpose. In this 
 method the lifts are 60 feet each and the slices of attack 6 to 7 
 feet high, taken horizontally across the vein for as broad a face of 
 attack as may be decided upon. Two main levels c and d (Fig. 33) 
 extend to the end of the district to be thus mined, being at intervals 
 connected "by risers e. The first slice is removed at level c, and
 
 82 MANUAL OF MINING. 
 
 waste rock from above delivered through the opening e, which is 
 then spread to fill the excavation completely. The temporary 
 level / is made on the top of the setting g. In each successive slice 
 chutes are constructed for the purpose of letting down mineral into 
 the lower level, from which it may be trammed to the shaft. This 
 process is repeated until all of the mineral has been worked out in 
 
 FIG. 33. A Slicing and Filling System. 
 
 slices represented by dotted lines and the waste rock distributed 
 to fill the spaces. 
 
 In Fig. 34 is shown another method in which, if the exploration 
 of the deposit is determined, this form of slicing-shaft, number 
 8, may be located permanently, and the rock-shaft, number 7, 
 sunk. The drift A on the hanging-wall side of the deposit is 
 connected by cross-cuts with the level C. Cross-cuts are also 
 driven through the deposits for the drift A to the foot-\vall and 
 winzes sunk between the levels. The deposit having thus been 
 blocked out, in each lift of 60 feet high there are three methods 
 of removing the ore, according to the extent of the face, which 
 may prove to be self-supporting. 
 
 First Method. From the drift A, slices DD are taken 6 feet 
 wide on either side of the timbering, cross-cut B, whose space 
 is being filled with rock dumped into A through a winze from the
 
 METHODS OF MINING. 83 
 
 level above. The next two slices, EE, are taken while the excava- 
 tions DD are being filled and the slices FF are mined, while E is 
 being filled with rock. The filling material is lowered from the 
 surface through rock-shaft number 7 and is hauled from the shaft 
 to the winzes above. 
 
 This process is continued until a layer 6 feet thick has been 
 removed over the entire deposit. The next slice above is operated 
 
 FIG 34. The Filling System, Plan of One Level. 
 
 in the same manner, with the exception of a new drift driven 
 above A , as in timbering as at B. The miners in removing the 
 second slice stand on the waste while at work, and behind them 
 is spread the material lowered for the purpose. The fillers follow 
 the miners very closely. Mill-holes are timbered upward as 
 mining progresses, the filling being packed about them. 
 
 Second Method. When the vein-matter is dry and stands 
 firmly it admits of a wider vein than 6 feet; the attack may be 
 made from the foot of the wall side for a width of half the distance 
 between the cross-cuts BB. The ore is then removed through the 
 cross-cut levels while the filling is being spread as above described. 
 The slices are driven in a direction parallel to the strike and across 
 it, as in the first method. It may be well to mention that similar
 
 84 OF MINING. 
 
 operations are being conducted not only at other points on the 
 same level, but also in a similar manner at other levels above or 
 below. 
 
 Third Method. This is designed for veins not over 20 feet 
 wide. Here the face of the attack is the two sides of the pillars 
 cut out by cross-cuts BB. The miners, starting from the cross-cuts, 
 work toward one another, parallel with the strike, until they meet, 
 their first cut being made at the walls, the filling progressing 
 behind them as before. 
 
 This system of filling is considered a cheaper and safer system 
 than either caving or square set for the extraction of mineral, 
 whether soft or hard, in large deposits, particularly if a neigh- 
 boring quany can supply the waste rock. It is of the widest range 
 of application. Very little timber is needed, and, with the excep- 
 tion of the first cross-cut B made in the solid ore, the mining is 
 cheaply done, because the width of face is large and the miners 
 have two free faces to take advantage of while blasting. Any 
 material whatsoever will serve for filling, but the best is that 
 mixed with clay, as it packs well. It should be in pieces as large 
 as may be conveniently handled, for the "smalls" settle too much. 
 In a few weeks' time the entire mass of filling compacts so that no 
 difficulty is experienced when the stowing of the uppnr level is 
 reached. 
 
 In the caving system, extreme caution is necessary to prevent 
 sudden falls of roof, and the adoption of the method is only justi- 
 fied where there is an absence of filling material and a scarcity of 
 timber. But if the vein-matter is comparatively firm, the method 
 with filling presents greater security, and may be cheaper. In 
 such event the choice of method depends upon the relative expense 
 of procuring quarried rock and framed timber.
 
 METHODS OF MINING. 8$ 
 
 REFERENCES. 
 
 Mining Thin Seams at Escatpelle, Coll. Guard., LXXII, 365; France 
 and Belgium, Coll. Guard., LXXII, Serial; North Staffordshire Fed. Inst. 
 M. E., VIII; Mining Thick Seams in Belgium, Coll. Guard., June 1896, 
 1202; Mining Methods, Pratt Coal, A. I. M. E., XIX, 299; Connellsville, 
 XIII, 330; Report A. C. Second Geol. Survey of Penn. Anthracite; Report 
 Second Geol. Survey of Penn. Bituminous, Annual Report, 1886; Mining 
 Methods in the Connellsville Coke Region, F. C. Keighley, Eng. Mag., Oct. 
 1900; Mining Methods and Safety Measures in the Wilczek Mines in Polish 
 Ostrawa, Josef Mauerhofer, Oesterr. Zeitschr. f. Berg u. Huttenwesen, May 
 16-23, 1903. 
 
 A New Method of Coal Mining, Genie Civil, Dec. 16, 1889; Coal Mining 
 in Natal, C. J. Gray, Coll. Guard., Nov. 6, 1903; Coal Mining in New South 
 Wales, Coll. Guard., Oct. 23, 1903; Coal Production and Consumption of 
 the Principal Countries of the World, Coll. Guard., Oct. 30, 1903; Mining 
 Coal at the Face, J. T. Beard, Mines & Min. Jour., April 18, 1903. 
 
 Pillar and Stall Mine Insp., 8th Kans., 21, 13; With Longwall, Coll. 
 Eng., Feb. 1897, 317; Thick Seam Anthracite in Colo., Coll. Eng., April 
 1897; Improvements Suggested by Mine Insp., Coll., May, 438; Boundary 
 Pillars Illust, Mine Insp., Pa., 1886; Pocket, Shaft Pillars, Coll. Guard., 
 June 18, 1897, 1185; Coal Pillars, M. & S. Press, 1893; Pillars and Waste in 
 Coal, Editorial and Historical, Coll. Guard., Dec. 18, 1896, 1153; Shaft Pillars, 
 Coll. Eng., July 1897; Coal Irruptions into Breasts, Coll. Manager, Dec. 18, 
 1896, 634; Coll. Guard., Dec. 1896, in; Mine Creeps, 111. Min. Inst., I, 
 239; The Working of Seams and Faults in Coal Measures, H. Fleck, Glauck- 
 auf, Jan. 3, 1903; Mine Development Methods, B. J. Forrest, Can.-Min. 
 Rev., June 30, 1903; Surface Subsidences Caused by Mine Workings, Coll. 
 Guard., Sept. 1897, 569; Suggestions for Improvements of Methods, Mine 
 Insp., Pa., in Anthra. Fields, Coll. Eng., May 1897, 438; Causes that Prevent 
 Improvements, Coll. Eng., May 1897, 462. 
 
 Mining Soft-ore Bodies, A. I.M. E., Vol. XVI, 862; L. Sup. M. Inst., I., 
 13; Mining Method on the Mesabi Range, C. E. Bailey, A. I. M. E., Vol. 
 XXVII, 529; Iron Mining, Menominee Iron Mines, A. I. M. E., XVI, 891; 
 A. I. M. E., XVII, 107; The Iron-ore Mines of Biscay, Bennett H. Brough, 
 Cassier's Mag., April 1903; New Jersey Iron Mines, S. of M. Quart., Jan. 
 1885, no. 
 
 Tamarack Mine, Cassier's Mag., Jan. 1897, 215; Osceola Mines, Coll. 
 Eng., 1895, 217; Butte, Montana, Mineral Industry, III, 175. 
 
 Mining Methods, Mineral Industry, II, 379; Thick Deposit of Ore, 
 S. of M. Quart., Jan. 1883, 100; Zinc and Lead Mines, Iowa Geol. Survey,
 
 86 MANUAL OF MINING. 
 
 VI, 1897; Zinc and Lead Mines, Missouri, Geol. Survey, VI and VII, Silver 
 Mining in South America, Cassier's Mag., V, 435; Methods of Mining and 
 Timbering in Large Ore Bodies in British Columbia and Michigan, Norman 
 W. Parlee, Can. Soc. of Civ. Engrs., Adv. Proof, Feb. 25, 1904; Mine Tim- 
 bering by the Square Set System at Rossland, B. C., Bernard McDonald, 
 Can. Min. Rev., Sept. 30, 1902. 
 
 Coal Mining in New South Wales, Harrison F. Bullman, Coll. Guard., 
 Feb. 23, 1900; Earth Pressure in Deep Mines, Coll. Guard., Vol. LXXXIV, 
 23; Flushing Goaves, ColL Guard., Vol. LXXXIV, 1274.
 
 CHAPTER IV. 
 
 POWER GENERATION. 
 
 The Power Plant. The engine plant consists of boiler and 
 engine, occasionally supplemented, in favored localities, by water- 
 power electric plants. The plant must be installed as compactly 
 as possible and of ample size. Common prudence suggests that 
 the machinery be in slight excess of the immediate wants. This 
 is more important, the lower the value of the mineral output. 
 While the latest and best machinery will in the long run prove to 
 be more economical, the question of finance and the amount of 
 capital available is not to be ignored, and the engineer may be 
 compelled to content himself with expedients until the mine has 
 been fully developed. 
 
 The mechanical considerations will call for the following 
 machinery: In a slope mine, stationary engines or locomotives for 
 haulage ; in a shaft mine, an additional hoister; cutting machinery 
 for coal, or drills for the rock or ore; fan for ventilation; pumps 
 for the accumulating waters; machinery in tipple, breaker, or 
 mill for crushing, screening, and jigging the mineral; pumps in 
 the mill; electric lights for the mill and a portion of the mine. 
 Either steam, electricity, or compressed air may be the motor 
 agent in any of these cases; their comparative advantages and 
 economy are discussed in Chapter VI. 
 
 The fan, the steam-locomotives, and such oil-engines as may be 
 employed are alone independent of the central plant. Other 
 machinery receives its power from the central source. Except 
 the self-contained oil-engine and the rare hydraulic engine, all 
 other motors depend for their primal power on the steam gen 
 erator. 
 
 87
 
 S8 MANUAL OF MINING. 
 
 The Boiler Rating. The boiler is essentially a device for 
 converting water at low temperature into steam at a high tempera- 
 ture, the measure of its capacity being the quantity of water which 
 it can evaporate in a given time. It must have sufficient surface 
 exposed to fire to accomplish this end, hence the boiler is rated 
 on its evaporative power and the area of its heating surface. 
 The commercial requirements, demanding some definition of 
 boiler unit commensurate with that of engines which it must 
 supply, have accorded an evaporation capacity of 34.5 Ibs. of 
 water per hour from and to 212 as equivalent to a horse-power. 
 The standard horse-power corresponds to 30 Ibs. (3.55 gallons) 
 of feed water evaporated from 100 F. to steam at 70 Ibs. gauge 
 pressure per square inch. This requires an absorption by the 
 water through the boiler shell of 33,305 British thermal units 
 per hour, practically equivalent to the centennial unit just men- 
 tioned. The general term of unit horse-power thus retained, 
 though a deceptive one and unscientific, is convenient. 
 
 In calculating the size of boiler required for a certain purpose 
 the amount of steam consumed per hour in the aggregate by the 
 various engines may be known from table, page 118, from the 
 given conditions of the operation. The amount of heat, measured 
 in B.T.U., necessary to develop the given quantity of steam may 
 be obtained from the steam tables. The number of boiler horse- 
 power to give the required service is found by dividing the total 
 B.T.U., necessary for evaporating the steam required per hour 
 by 33>305 B.T.U. 
 
 This power may be subdivided in any manner if one boiler is 
 added in excess of the average demand. It would be more ad- 
 vantageous to install the required heating surface for average load, 
 remembering that any unit may easily be forced 33^ per cent with- 
 out sacrificing efficiency. Then at times of emergency one unit may 
 be repaired while the remaining units are overloaded, yet within 
 their capacity. Thus, if 800 pounds of steam are required per hour 
 as a maximum and but 3000 pounds throughout the day, three 
 units may be installed. If 12,000 pounds of steam are in demand, 
 four units may be installed, giving opportunity for one boiler to be
 
 POWER GENERATION. 89 
 
 repaired while the remainder are forced. Thus the plant is not 
 too large. 
 
 Practice has shown that to enable the absorption of 33,305 
 B.T.U. to be effected through the shell there must be a heating 
 surface of 14 sq. ft. Hence 14 sq. ft. of surface exposed to the 
 flames under a fire-tubular boiler constitute a boiler horse-power 
 or 12 sq. ft. of exposed area in the water-tubular boiler. 
 
 The grate area must be sufficient to provide space for the 
 coal and be at least -gV f the total heating surface. In other 
 words, each boiler horse-power must have from 0.24 to 0.28 sq. 
 ft. of grate area. If the grate is too large, the temperature of 
 combustion is too low. High temperature in the furnace is the 
 first condition of economy, the second being that the heat pro- 
 duced shall be absorbed as completely as possible by plenty of 
 clean heating surface. A grate surface too small causes clinker- 
 ing of the coal. A low-grade fuel requires a large grate. As a 
 limited quantity can be burned on a square foot of grate, the max- 
 imum evaporative capacity of the boiler is limited by its grate 
 area. 
 
 The Standard of Boiler Comparison. The equivalent evapora- 
 tion of a boiler is that amount which would have been evaporated 
 had the steam been produced at atmospheric pressure (=212). 
 It may be ascertained by dividing the B.T.U. necessary to boil 
 the feed-water into steam by 966, the latent heat of steam at 
 14.7 Ibs. pressure. Thus, feed-water at i6oF., converted into 
 steam at 100 Ibs. gauge pressure, requires 1057 B.T.U. The 
 factor of evaporation is then 1057-^-966 = 1.094. The equivalent 
 evaporation is 1.094 times the actual evaporation. 
 
 The rate of evaporation is about 8 to 19 Ibs. of feed- water 
 per pound of coal. 
 
 The Feed-water. Care must be taken to secure ample supply 
 of pure water for the boiler. Acids in the water corrode the 
 boiler and reduce its efficiency and life. Salts in suspension 
 or in solution give trouble in depositing within the boiler and 
 forming a compact layer which is both injurious and wasteful 
 of heat. A loo-horse-power boiler evaporating 3000 pounds
 
 MANUAL OF MIXIXG. 
 
 "\
 
 POWER GENERATION. 
 
 of moderately pure water per hour would in a year's time receive 
 248 Ibs. of solid matter, which would cover the surface inch 
 thick unless removed before being consolidated. Such a crust is 
 
 inevitably formed in all boilers, and, being a poor conductor of 
 heat, prevents heat from reaching the water. A scale of T V inch 
 thick on the tubes causes a loss of 13 per cent of the fuel heat.
 
 92 MANUAL OF MINING. 
 
 Moreover, the boiler-shell becomes excessively heated for the 
 same steaming effects and tends to blister and warp. The scale 
 may contract, and cracking, will suddenly expose the heated 
 plate to the water; the generation of steam may follow and an 
 explosion ensue. 
 
 Boiler Scale. These two wasteful and dangerous sources 
 require that care be exercised in the selection of the water-supply 
 for the boiler plant. Mine waters are usually acid and should 
 be neutralized by ammonia or precipitated by chloride of barium 
 in the tank prior to being delivered to the boiler. Other sources 
 of supply, except in the mountainous regions of the Rockies and 
 Sierras, contain organic or mineral substances which should be 
 rendered innocuous in the boiler or, preferably, removed before 
 introduction by feeding the water-supply through the econo- 
 mizer or heater, Fig. 36, prior to its delivery to the boiler. A 
 considerable deposition will ensue from mere increase in the 
 temperature of the water; the sediment there formed can readily 
 be removed. Such material as still remains in suspension in 
 boiler must be rendered innocuous by an addition of kerosene, 
 which would cover the grains with a film and prevent the sedi- 
 ment compacting too densely on the tubes and shell. Washing- 
 soda will dissolve the scale and change its nature. The amount 
 of kerosene to be fed should not exceed 2 quarts per week per 100 
 horse-power, and of soda one pound per week. An excessive use 
 of these remedies would cause "foaming" and deliver wet steam 
 into the cylinder. The best remedy for the formation of scale is 
 to blow off and to clean the boiler regularly every week. 
 
 When the condenser is used in a steam plant its supply is 
 returned to the boiler, and if lubricating-oil has been used exces- 
 sively in the engine a very injurious mixture is introduced into 
 the boiler. The oil should be absorbed from the steam before 
 it reaches the condenser, or removed from the water leaving the 
 condenser before delivery to the boiler. It is caught on screens 
 or separated by a centrifugal machine. 
 
 It is a question of no small matter for the engineer whose 
 water-supply is chemically bad to determine whether to incur
 
 POWER GENERATION. 93 
 
 the expense of a special distillation plant, or to employ the con- 
 densed exhaust-steam. Some engineers have solved the problem 
 by using the very smallest amount of lubricating-oil possible 
 in the pumps and engines, thus saving the boiler at the expense 
 of wear upon the pistons. The author favors the plan of purify- 
 ing by chemical or thermal process such water as requires it in 
 a feed-water heater, or an economizer, before delivery to the 
 boiler. It may even be desirable to sink an Artesian well for 
 water-supply if not otherwise attainable, though in that event 
 it should not be drawn from a limestone formation. 
 
 The Feed-water Heater. The feed-water heater is designed 
 to utilize the heat in the exhaust- steam or the flue- gases by bring- 
 ing the feed-water supply in close contact with them. It is 
 known as an economizer, because only waste heat is employed 
 for the purpose The simplest form of heater has a tank of 
 matched dovetailed boards in which is placed the coil of pipe, 
 one end open, blowing into the atmosphere, and the other end con- 
 nected with the exhaust from the engine. The heat in the steam 
 is communicated to water and its temperature raised, while at 
 the same time the carbonic acid or carbonates in the water are 
 released and the salts deposited. The sediment can readily be 
 removed, if a hand-plate be provided near the bottom. 
 
 A duplicate tank to this may then receive the exhaust- steam 
 during the process of cleaning. Hay is often placed in the water- 
 supply tanks, filtration through which would remove some of 
 the muddiness. 
 
 Feed-water heaters can also be procured of the closed type, 
 consisting of a steel cylinder filled with tubes through which the 
 steam flows, while the water circulates around them or the reverse. 
 The benefit derived from the use of the economizers increases 
 with the wastefulness of the boiler, and the need of puri- 
 fication warrants their introduction when the water-supply is 
 chemically bad. The life of a well-built economizer is from 
 fifteen to twenty years. The average plant will save at least 
 10 per cent of the fuel value. The gross return on the invest- 
 ment of the economizer is 48 per cent annually with coal at $5
 
 94 MANUAL OF MINING. 
 
 per ton, and 19 per cent with coal at $2 per ton, assuming twenty 
 hours of service out of the twenty-four. In addition to the sav- 
 ing of heat and the perfecting of the waters, there is another 
 advantage, that the boiler receives water at a high temperature 
 and is saved from the straining effects which result from cold 
 water. A small pump, or an injector, delivers the supply to the 
 boiler. The latter is cheaper, equally reliable, and easily handled. 
 
 Types of Boilers. It is a matter of considerable difficulty to 
 select the best boiler, because of the great variety now on the 
 market. The three types, have each their advocates. The 
 Cornish type of boiler, with the fireplace in two large tubes, 
 has a high efficiency, and is much liked where its length is 
 no objection. The multitubular boiler, with the fireplace below 
 the front end and fitted with numerous tubes through which the 
 fire returns on its way to the chimney, is a most common type. 
 The latest type is the water-tubular boiler (Fig. 36), in which 
 the fire and the gases circulate around the tubes; in these the 
 water flows and is converted to steam. These three types have 
 their separate fields of usefulness and find wide acceptance in 
 engineering circles. Their records can easily be had, so they 
 must be selected according to local requirements. The relative 
 efficiency of the three types is i, 1.6, and 2; their relative initial 
 cost is very nearly in the same proportion. 
 
 The Cornish boiler utilizes 50 per cent of the fuel value in 
 the production of steam; the fire- tubular boiler absorbs 60 to 
 65 per cent of the heat generated in the fireplace; and the water- 
 tubular boiler absorbs fully 80 per cent of the fuel value. The 
 comparative durability of the three types is about the same, with 
 an advantage in favor of the more expensive water-tubular 
 boilers. They all are good steamers, though the last-named 
 can be most readily forced when the emergency arises. 
 
 Water-tubular Boilers. The steam pressure to which these 
 types may be worked is respectively 60 Ibs., 100 Ibs., and 200 
 Ibs. pressure per square inch, as a maximum for ordinary condi- 
 tions. Inasmuch as the simplest method of increasing the power 
 of a steam-engine consists in raising the pressure of the boiler
 
 POWER GENERATION. 95 
 
 steam, it follows that when such emergency arises the water- 
 tubular boiler will best fit the case. These boilers are sectional 
 and can be easily transported. They are safer than the ordinary 
 type even at higher steam pressure, because the excessive pres- 
 sure beyond the danger line will overstrain but one tube instead 
 of the entire boiler of the other types. They are quickly cleaned, 
 easily repaired, and give drier steam and with a smaller amount 
 of fuel than the other types. 
 
 Heat Losses. All the heat developed by coal must be al> 
 sorbed in steam production. There are, however, three sources 
 of loss that due to heat carried by the flue- gases escaping up 
 the chimney, the radiation from the walls and shell of the boiler, 
 and the heat remaining in the ashes. 
 
 The waste in the ashes depends on the conditions in firing, 
 etc. Their weight may be 20 per cent of the total fuel burned, 
 and in them is also some unburned combustible. The unpro- 
 tected boiler surface radiates, per square foot, 650 B.T.U. per 
 hour. Thus the loss from a 100- horse-power boiler and its wall 
 is equivalent to the heat of 100 tons of coal a year, or at least 
 15 per cent of the fuel. A covering of magnesia, hair-felt, or 
 asbestos, i inch thick, would in five years' time save much more 
 than its cost, which is 25 cents per square foot. A brick arch 
 over the boiler, furnishing a conduit for flue- gases on their way to 
 the chimney would be still better. Boilers in batteries, if placed 
 back to back instead of facing one .another, would also reduce 
 the radiation loss, though the cost of stoking would be increased. 
 
 Flue-gases. The flue-gases have a temperature between 300 
 and 600 F., being higher with the forced draft than with the chim- 
 ney draft. In order to give place for the air-supply in the com- 
 bustion of the coal, these flue- gases are removed as promptly as 
 possible. Each pound of coal produces gases that transport 
 1500 or more B.T.U. This is always more than 10 per cent of 
 the total available heat in the fuel. The only recoverable por- 
 tion of it is that which the economizer may utilize. Smoke 
 increases the heat loss by that which is latent in the unconsumed 
 combustible.
 
 p6 MANUAL OF MINING. 
 
 The Air-supply. The aggregate of the above sources of heat 
 loss is at least 15 per cent of the fuel value, and is more frequently 
 35 per cent in the ordinary boilers. 
 
 Economic combustion will reduce the great waste of heat by 
 providing the proper supply of air, the perfect burning of a well- 
 selected fuel, and a good regulation of the draft. Conditions 
 for complete combustion require that the particles be brought 
 into intimate contact with the requisite amount of oxygen. The 
 amount of air-supply being the chief consideration, each pound 
 of coal will require an amount of air at least equal to A = 
 i.52(C+3Ho.4O), in which C, H, and O are the percentages 
 of carbon, oxygen, and hydrogen. 
 
 This is about 160 cu. ft. or 12 Ibs. of air. More than this is 
 necessary to insure intimate contact of the fuel constituents and 
 the air. Hence 24 Ibs. are usually given for each pound of fuel 
 with the chimney draft and about 18 Ibs. with forced draft. Any 
 excess over this amount chills the fire and will only carry heat 
 from the furnace to the chimney. With less air the carbon will 
 develop only 4450 B.T.U. per pound, instead of 14,500 B.T.U., 
 which is more wasteful than smoke. The production of CO in- 
 stead of CO 2 results from a deficiency of air; the fire is dull and 
 sluggish. 
 
 An occasional analysis of the flue-gases, by the D'Orsat ap- 
 paratus, for free oxygen and CO, will reveal the working of 
 the furnace. The former indicates air excess, the latter air 
 deficiency. 
 
 Calorific Value of Fuels. The heating power of a fuel depends 
 upon its combustible elements and the amount of ash. The fuel 
 should be bought on its calorific value and according to analysis. 
 This latter may be the ultimate analysis, giving the percentages 
 of carbon, oxygen, and hydrogen in the fuel, whence its calorific 
 value is P = i4,55oC+62,oooH 54oo(O + N); or an approximate 
 analysis giving the amount of volatile matter and fixed carbon in 
 the fuel. In the latter case, assume M to be the percentage of 
 
 volatile matter, V, compared with the dry combustible, M = 77-7^, 
 
 ~
 
 POWER GENERATION. 97 
 
 and A a coefficient whose values are given below. Whence the 
 calorific power: P^i^.jC+AM. 
 
 A = 252 for M= 2 to 12; 4 = 169.2 for M = 30 to 35; 
 
 A = 216 for If =12 to 17; 4 = 144.0 for Af = 35 to 38; 
 
 4 = 198 for M =17 to 24; 4 = 142.2 for M = 38 to 40; 
 
 4 = 183 for M = 24 to 30; 4 = 136.8 for M = 40 to 50. 
 
 The heat which can be developed from a pound of coal varies 
 between 11,000 B.T.U. and 15,000 B.T.U. As one pound of 
 water requires for its evaporation and conversion to steam some- 
 thing over 1 100 B.T.U. it follows that from 10 to 14 Ibs. of steam 
 can be produced for each pound of coal, though some boilers 
 realize only 4 to 6 Ibs. 
 
 Undoubtedly the semi-bituminous coals are better steam- 
 producers than anthracite coals, the amount of the ash being 
 less. But it is usually a question of local price as to which is 
 preferable in a given case. A superior coal is that which has 
 about 85 per cent of fixed carbon with some volatile matter to 
 assist in the firing. A steaming coal should kindle readily and 
 burn steadily without clinkers. 
 
 Slack is nearly equal to coal in its calorific value, but has 
 usually too much refuse to be acceptable, except where the trans- 
 portation is cheaper. The size of the fuel is selected according 
 to the relative grate openings. The smaller coals and slack are 
 cheaper than the larger size of the same coals, but the amount of 
 ash in the former increases the labor of handling. 
 
 The Rate of Combustion. The amount of coal that can be 
 burned depends upon the intensity of the draft and the air- supply. 
 The thickness of the bed should be as small as possible without 
 requiring excessive watchfulness. A 5 -inch bed of anthracite 
 pea coal or an 8-inch bed of lump coal is good practice. Only 
 12 Ibs. of anthracite or 15 Ibs. of bituminous coal can be burned 
 per hour per square foot of grate area by natural draft with eco- 
 nomical results. The results from the boilers depend more upon 
 the character of firing than on any other single element. 
 
 The Methods of Hand Firing usually adopted in practice 
 with ordinary grates are the "spreading" method, in which a
 
 98 MANUAL OF MINING. 
 
 thin layer of coal is worked over the entire area from the bridge 
 towards the front; the "alternating" method, in which the fresh 
 coal is charged alternately on either side of the fire-box; and the 
 "coking'' system, for bituminous coal only, that charges it on the 
 dead- plate at the front of the fire and pushes it back, when coked, 
 into the grate to make room for a new charge. This is advan- 
 tageous with all long-flame gaseous coals. Two stokers and two 
 engineers can fire 35 tons of coal per week under the boilers during 
 24 hours. The limit for one man firing boiler and running the 
 engine is about 10 tons of coal per week for a 50-H.P. engine. 
 
 Stokers. Mechanical systems of feeding hard coal are employed 
 to advantage. The Steam Users' Association of Boston reports 
 that in all plants stokers may prevent smoke, save coal, and 
 reduce labor as much as 30 per cent in big plants employing 
 also coal-handling machinery. Mechanical stoking reduces the 
 smoke nuisance by furnishing a continuous charging. Stokers, 
 however, cut down the capacity slightly, though they respond to 
 a sudden demand for steam as well as hand firing. The amount 
 of repairs is not excessive. 
 
 The stokers are travelling grates or moving bars fed from a 
 hopper continuously at the front and carried by power toward 
 the rear at a slow rate such as to insure a complete burning of the 
 fuel before it reaches the rear of the fire-grate where it is dumped. 
 Fine sizes of coal can be utilized and little power is required. 
 Undoubtedly they afford a betterment of evaporative results, but 
 it is a question whether the same amount of improvements can- 
 not be effected in other directions. 
 
 Fuel Consumption. The consumption of fuel per hour 
 averages 3! Ibs. per horse-power with a common boiler and 
 ordinary engine. The very best results which have yet been 
 obtained in practice require 1.5 Ibs. of good coal to produce a 
 horse-power hour. 
 
 The question of fuel economy is an exceedingly important 
 one, whether one considers the production of power at the coal- 
 mine, where the fuel is cheaper, or at the metal-mine, where it is 
 very dear. Any economy effected in this direction represents a
 
 POWER GENERATION. 99 
 
 corresponding gain. The bituminous collieries of the United 
 States consume between 1.5 per cent and 2 per cent of their total 
 output in providing power for pumping and water. The anthra- 
 cite coal-mines of the United States are estimated as consuming 
 from 8 per cent to 10 per cent of their total production for the 
 same purpose. 
 
 A cord of well-dried spruce is capable of driving a common 
 slide-valve hoister about 320 hourly horse-powers; a ton of 
 lignite, 470; a barrel of petroleum, 170; a ton of anthracite, 650. 
 
 Weathering Coals. Fuels should be properly housed from 
 the weather, and not piled out of doors, to prevent absorption 
 of moisture, the evaporation of which causes deterioration. 
 Freshly mined coal absorbs three times its volume of gas or air, 
 and this absorption produces a slight decomposition that reduces 
 its heating power. Moreover, spontaneous combustion may 
 develop from a prolonged oxidation, which can be prevented 
 only by a thorough ventilation of the coal and its protection 
 from the elements. The practice of sprinkling the coal arrests 
 the action temporarily, but deterioration soon begins, which would 
 be facilitated by the presence of fine coal. It is a good practice to 
 have a storage for 500 Ibs. of coal for each indicated horse- power 
 of engine plant. 
 
 Liquid Fuels. Petroleum has an approximate analysis of 
 carbon 85 per cent, hydrogen 13 per cent, and oxygen i per cent, 
 and a calorific value of about 20,000 heat-units per pound. This 
 fuel can be injected into the fire-box from tanks by a nozzle 
 with the aid of compressed air. This blast induces a current, 
 atomizes the oil, and delivers a spray with an excess of oxygen 
 which permits the production of an intense flame. No grate is 
 required. One pound of oil equals 2.18 Ibs. of coal in calorific 
 value. Where conditions favor it, gas from coke-ovens can be 
 employed to advantage under a boiler. 
 
 Draft. A draft is necessary to carry off the products of com- 
 bustion and to admit a fresh supply of air to the fire. The force 
 of the draft is measured by a water-gauge which varies between 
 r inch and 2 inches. It may be either natural or forced.
 
 ioo MANUAL OF MINING. 
 
 Chimneys. Natural draft is procured by a chimney of brick 
 or of sheet-steel tubing. It depends upon (he relative densities 
 of the hot gases inside the chimney and the external air. The 
 difference in the temperatures and the height of the stack will 
 determine the velocity of the flow and the pressure of the draft. 
 The inside area determines the volume of air which can be car- 
 ried. This is known from the total hourly consumption of fuel, 
 F, which is based on 5 Ibs. per horse-power. The desired horse- 
 power being known, the effective area, E, may be determined 
 for a known height of chimney, H. 
 
 0.06257? 
 
 = ~~ 
 
 The effective area E js less than the actual area by the amount 
 indicated, E=A- o.6\/Z. 
 
 The commercial horse-power of chimney is ^.^EVH. 
 
 The first cost of a brick chimney is greater than that of a 
 sheet-steel stack, but its stability is greater and its repairs are less. 
 By the use of a damper regulated by hand or automatically by 
 steam, any lower pressure than the maximum draft for which it 
 is designed may be maintained in the latter arrangement. A 
 small diaphragm raises the lever and closes the damper when 
 the steam pressure in the boiler exceeds a certain desired limit. 
 
 Blowers. Forced draft may be obtained by the use of a 
 blower driven by the engine and dynamo or belt. By it air is 
 delivered through the pipes into the fireplace at a pressure of 
 perhaps 10 Ibs. above the atmosphere. The steam-jet may also 
 be employed, acting like an injector; though it is not an eco- 
 nomical device, it softens the clinkers and aids combustion. 
 A steam-blower can be built composed of a horizontal ring with 
 holes punctured in it. Steam from the boiler is discharged 
 through the openings and creates a current that will increase 
 the draft and thereby the capacity of the boiler 20 per cent to 
 30 per cent. The ring is made of i^-inch pipe 12 inches in diam- 
 eter punctured with eighteen -inch holes for a 6o-horse-power 
 boiler. Such a blower costs 90 cents an hour for operation.
 
 POWER GENERATION. IOI 
 
 Exhaust- fans may be so placed as to draw the gases out of 
 the furnace and deliver them into short stack to aid draft. 
 
 The chimney has a higher first cost than any of these sys- 
 tems which require short stack only, but is cheaper to maintain. 
 It is, however, very wasteful of heat, since the flue-gases must 
 be hot in order to produce draft. Nevertheless, the latter adapts 
 the boiler to a wide range of power, for the fire can be made more 
 intense. 
 
 Boiler Installation. The boilers should be on a level below 
 the engine-room and set as high up as is consistent with the 
 fireman's duties. The coal should be stored above them and 
 gravitation availed of for feeding the boilers and for removing 
 ashes. This would be an ideal condition as to labor. The 
 cost of power production is not always a matter only of fuel 
 selection and economy of combustion, for the cost of handling 
 is inseparably connected with it. Generally speaking, the annual 
 coal consumption equals the initial cost of the boiler plant, and 
 economy in labor may easily be effected equal to 10 per cent of 
 the initial cost. 
 
 With the pipes hung to drain towards the boiler, covered 
 with a layer of asbestos, magnesia, or felt covering, with a steam- 
 trap properly placed, a supply of dry steam will be delivered 
 to the engine. Each square foot of exposed surface of pipe 
 radiates 1.8 B.T.U. per hour for each degree Fahrenheit differ- 
 ence in temperature between the steam and the external air. 
 Thus a 4- inch pipe carrying steam of 90 Ibs. pressure (320 F.) 
 40 feet long radiates an amount of heat equivalent to 17 Ibs. 
 steam from i2oF. to 90 Ibs. absolute pressure. Hence a con- 
 siderable loss of power ensues if pipes be uncovered Every 
 284 feet of 4- inch pipe covered with i-inch layer saves a horse- 
 power. 
 
 As the allowable velocity of steam in pipes is ordinarily 50 feet 
 per second, it is necessary to have a line of pipes ample in area to 
 furnish a copious supply of steam with but small loss. A 4-inch 
 pipe will discharge 115 Ibs. of steam per minute, the initial pres- 
 sure being 100 Ibs. gauge, or it will deliver 109 Ibs. of steam at
 
 102 MANUAL OF MINING. 
 
 80 Ibs. gauge. A 2-inch pipe will deliver 30 Ibs. and 24 Ibs. 
 of steam respectively. As it requires about 25 Ibs. of steam 
 to supply one horse-power per hour, a 2-inch pipe will serve 
 for about 60 horse-powers. 
 
 The maximum absolute steam pressures for successful opera- 
 tion are ordinarily as follows: For simple non-condensing 
 engines, 60 Ibs. gauge pressure; for compound non-condensing, 
 loo Ibs. gauge; for simple condensing, 90 Ibs. gauge; and com- 
 pound condensing, 130 Ibs. gauge. The character of the boiler 
 will largely determine the maximum pressure available. A 
 multitubular boiler cannot be forced beyond 100 Ibs. pressure, 
 while the water-tubular boiler may be carried to 200 Ibs. or more. 
 
 The Steam-engines. The steam will be conducted to the 
 underground pumps, the exhaust-fan or the ventilating steam- 
 jet, and to the steam end of all compressors, electric generators, 
 and hoisters. Its behavior in the cylinders of these engines 
 is alike in all, and the following considerations will apply to 
 all. 
 
 Classification of Engines. The engines may be classified 
 according to the manner in which they utilize the elastic force 
 of the steam. Direct-acting pumps and other engines having a 
 constant resistance do not admit of a change of pressure within 
 the cylinder of the engine; but where the resistance varies, or 
 it is possible to provide for the varying resistance, the steam is 
 permitted to expand, with the corresponding reduction of pressure, 
 with marked economy. 
 
 The Behavior of Steam. Fig. 37 illustrates the behavior 
 of steam inside of a pump in which the length of the diagram 
 corresponds to the length of the cylinder, and the height of the 
 diagram ihe pressure of steam at various points. The upper 
 horizontal line represents the steam pressure admitted to the 
 cylinder, which, remaining constant, will be straight. The atmos- 
 pheric pressure being also constant, the former will be parallel 
 to it and at a distance above it according to the scale on which 
 the pressure may be drawn. The pressure of the atmosphere being 
 14.7 Ibs. per square inch, a line may be drawn below it at that
 
 POWER GENERATION. 103 
 
 scale and represents the line of absolute zero of pressures. OX 
 is that line, MN the atmosphere line, and AB the line of con- 
 stant steam pressure. When the piston reaches the end of its 
 stroke the exhaust-valve is opened for the discharge of the steam. 
 The pressure falls nearly to atmospheric pressure. On the return 
 stroke of the piston the steam is being expelled at a constant 
 pressure, represented by the line CD, until the piston is at the 
 end of the cylinder. Steam is now admitted. The pressure 
 rises and the line DA represents such increase in the pressure. 
 
 On the other side of the piston, the operations ensue in the 
 same order, the line CD, representing the back pressure, on one 
 side being simultaneous with the steam pressure, AB, on the other 
 
 FIG. 37. Indicator Card from a Non-expansion Steam-engine. 
 
 side. If the steam be exhausted into the condenser, the line CD, 
 representing the back pressure upon the pump, will be below the 
 atmospheric line MAT". Its length above the zero line, OX, is an 
 amount representing the vacuum pressure in the condenser. 
 This line would then be at about 5 Ibs. as a minimum. 
 
 If the steam after being admitted at constant pressure for a 
 portion of the stroke is cut off, the piston will continue to move in 
 its stroke, but with a continual reducing pressure to the end. Fig. 
 38 will represent in the two lines AB and BI the behavior of 
 the steam during admission and expansion. The back-pressure 
 line, CD, will remain the same as in the pump. The discharge 
 of the steam may be cut off at some point, D, on the return stroke; 
 the steam still remaining in the cylinder will be compressed and 
 the line DE will indicate the increase in pressure for the balance 
 of the stroke. When admission occurs the line EA will rise 
 more or less promptly, according to the freedom of entry of steam. 
 The back-pressure line will be below or above MN according
 
 104 MANUAL OF MINING. 
 
 to whether a condenser is attached or not. The diagrams repre- 
 sent, therefore, the behavior of the steam at all periods of the 
 piston movements in the cylinder, and the area enclosed be- 
 tween ABIC DBA measures the work done during one stroke. 
 The vertical lines represent the pressures to scale, in pounds per 
 square inch of area piston; the horizontal line represents the 
 length of stroke, s. The area divided by the length of card and 
 multiplied by the scale of vertical pressures equals the mean 
 effective pressure, M.E.P., in the piston during the entire stroke. 
 Diagrams such as these are called indicator cards and measure 
 the indicated horse-power of the engine. The device known 
 
 102.fi* 
 
 FlG. 38. Indicator Diagram from a Non-condensing Engine or the High-pressure 
 Cylinder of a Compound Engine. 
 
 as the indicator (Fig. 40) can be procured, by which such cards 
 of steam behavior can be automatically registered by the engine 
 while at work. 
 
 In order to obtain the most economical consumption of 
 steam, the lowest point, /, of the expansion line, BI, should be 
 as near to the back-pressure line as possible. The limit, however, 
 is 3 Ibs. gauge in non-condensing engines and about 5 Ibs. abso- 
 lute when condensers receive the exhaust. 
 
 The expansion line, BI, follows practically the equation of 
 the equilateral hyperbola, PV = constant; if the cylinder walls be 
 perfect non-conductors, the expansion line would be represented 
 by the adiabatic equation py 1 - 135 . 
 
 The most desirable point of cut-off then is fixed, the conditions 
 being given. Thus, if the pressure at /be 18 Ibs. absolute and 
 the admission pressure be 90 Ibs. absolute, the cut-off will be at 
 one-fifth of the stroke. The ratio of expansion is then 5; that
 
 POWER GENERATION. 
 
 is, the volume of steam delivered to the cylinder up to point of cut-off 
 is expanded to five times this volume at the end of the cylinder. 
 
 The Mean Effective Pressure. By the following equation 
 M.E.P. may be ascertained for a given admission gauge pres- 
 sure, P, and a back pressure, B, in pounds absolute per square 
 inch. By gauge pressure is understood the pressure above 
 the atmosphere as recorded on the steam-gauge. Absolute pres- 
 sure is the pressure above absolute vacuum. Atmospheric pressure 
 is 14.7 Ibs. absolute. />=C(P+ 14.7)-^. 
 
 The value for C is obtained from the following table for given 
 ratios of cut-off. 
 
 VALUE FOR THE CONSTANT, WITH VARIOUS CUT-OFFS, FOR DETERMINING 
 THE M.E.P. 
 
 
 . 
 
 
 Ratio of 
 
 
 Ratio of 
 
 
 Percentage 
 Cut-off 
 Apparent. 
 
 Expansion, 
 No Clear- 
 ance. 
 
 C. 
 
 Expansion 
 with 3 
 Per Cent 
 
 C. 
 
 Expansion 
 with 7 
 Per Cent 
 
 C. 
 
 
 
 
 Clearance. 
 
 
 Clearance. 
 
 
 O.I 
 
 10 
 
 332 
 
 7.69 
 
 4113 
 
 5.88 
 
 '47*5 
 
 f 
 
 8 
 
 385 
 
 6-45 
 
 4435 
 
 5.128 
 
 5i37 
 
 i 
 
 6 
 
 4653 
 
 5.102 
 
 .5155 
 
 4-23 
 
 5774 
 
 i 
 
 5 
 4 
 3 
 
 5219 
 .5966 
 
 6995 
 
 4.3478 
 3-57 J 4 
 2-754 
 
 .5682 
 .6368 
 .7294 
 
 3-7037 
 3-125 
 .481 
 
 .6225 
 .6861 
 .7708 
 
 0.4 
 
 2 -5 
 
 .7664 
 
 2 -3 2 55 
 
 793 1 
 
 .1276 
 
 .8242 
 
 | 
 
 2 
 
 .8465 
 
 1.8867 
 
 .8658 
 
 754 
 
 8925 
 
 | 
 
 1.6 
 
 .9180 
 
 1-527 
 
 .9310 
 
 439 
 
 .9481 
 
 I 
 
 1-333 
 1.25 
 
 .9656 
 9785 
 
 1.282 
 1.2875 
 
 9743 
 .9807 
 
 2195 
 .149 
 
 .9842 
 .9914 
 
 I 
 
 1. 143 
 
 .9918 
 
 1.104 
 
 .9941 
 
 .058 
 
 .9961 
 
 For all practical purposes the ratio of expansion and the 
 point of cut-off are reciprocals of one another. The clearance 
 space is left at the end of the cylinder, which the piston never fills 
 and in which steam is allowed to accumulate and act as a cushion 
 to prevent the piston from striking the end of the cylinder. This 
 clearance space contains steam which is ineffective and changes 
 somewhat the ratio of expansion for a given cut-off. The clear- 
 ance is from 3 per cent of the stroke in the better classes of 
 engines to 7 per cent in the simple slide-valve engines. High- 
 speed engines may have a still larger clearance. 
 
 The shorter the point of cut-off the greater the efficiency of
 
 106 MANUAL OF MINING. 
 
 the engine. The limit, however, depends upon the character of 
 the admission-valve. With a simple slide-valve engine the cut-off 
 cannot be less than 0.4 of the stroke. With the Corliss valves, 
 or some forms of rider cut-off valves, the cut-off may be very much 
 shorter, though it rarely is less than one-fifth in a single cylinder. 
 
 The M.E.P. and the area of the card in Fig. 37 may be greater 
 than in Fig. 38, but it must be remembered that the volume of 
 steam used in the former case is equal to the entire volume of 
 the cylinder, and in the latter case to the volume up to the point 
 of cut-off. Hence the work performed per pound of steam in 
 the latter case will be greater than in the former. 
 
 The Use of the Indicators. An indicator, which is essentially 
 a stethoscope to the engine, reveals the character of the steam 
 expansion compared with the theoretical; the amount of water 
 existing in the cylinder at any moment; the point of cut-off; the 
 setting of the valves; and the amount of resistance to exhaust. 
 None of these can be detected from the behavior of the engine. 
 Leaks, tight stuffing-boxes, loose piston packing, or a priming 
 of the admission steam are likewise revealed by the indicator. 
 The machine is simple, easily attached, and requires no elaborate 
 calculation, and by the aid of the planimeter (Fig. 39) it is 
 
 FIG. 39. Measuring the M.E.P. of an Indicator Card by Planimeter. 
 
 possible to obtain a card by which to determine the indicated 
 horse-power. Fig. 40 illustrates one attached to the engine. 
 
 An indicator is a device for registering the behavior of the 
 steam within the engine cylinder. It consists of a drum and a
 
 POWER GENERATION. 
 
 107 
 
 cylinder; the latter has a plunger whose rod carries a pencil above 
 it and is connected by a pipe to the engine cylinder. When 
 desired, the steam may be admitted from the engine under the 
 plunger, which is then lifted an amount depending upon the pres- 
 sure existing in the engine at that instant. With it the pencil 
 rises, and each change of steam pressure will be indicated by a 
 corresponding rise or fall of the pencil. To control the amount 
 of lift of the piston and its pencil a spring of known strength is 
 inserted above the plunger, and thus not only each change of pres- 
 sure but also the amount of such pressure may be recorded. 
 Each rise of pressure lifts the plunger a definite amount by com- 
 pressing the spring correspondingly. A drop in the pressure is 
 indicated by an extension of the spring. The pencil then rises 
 a certain amount when the steam pressure in the engine and on 
 the plunger is 100 Ibs. per square inch. It will rise half that for 
 a steam pressure of 50 Ibs., etc. 
 
 FIG. 40. Attaching an Indicator. 
 
 The pencil bears against and records upon a sheet of paper, 
 about 6 inches long, which is slipped on a drum of about 2 inches 
 in diameter. A large wheel attached to the drum, turned by 
 a string fastened to the cross-head of the engine, pulls the string, 
 turns the drum, and the paper moves under the pencil. As the 
 piston travels in its cylinder the pencil rises and falls simultane- 
 ously with it, thus furnishing a diagram called an indicator card. 
 Various devices are employed to reduce the long piston stroke 
 to a card length of about 4 inches.
 
 joS 
 
 MANUAL OF MINING. 
 
 The plunger springs are selected according to the boiler pres- 
 sure, so that the height of lift of pencil shall not exceed 2 inches^ 
 thus for 100 Ibs. admission pressure a number 50 spring is taken, 
 which means that steam of 50 Ibs. pressure will compress that 
 spring an amount that will lift the pencil i inch; each change of 
 i Ib. steam pressure will move, the pencil -^ of an inch. The 
 scale of the spring is 50. On the diagram drawn by such a 
 spring the vertical lines represent the pressure at the rate of 50 
 Ibs. per inch of height, and the horizontal line the stroke of 
 the piston on a scale equal to the ratio of the piston stroke divided 
 by the card length. 
 
 Steam Condensation in the Cylinders Remedies. The objec- 
 tion to a short cut-off and a high rate of expansion lies in the fact 
 that the walls of the cylinder are not non-conductors, which, being 
 comparatively cool at the end of the stroke, condense some of 
 the entering steam whose energy is thus lost. The greater the 
 ratio of expansion the larger is the percentage of the steam, 
 condensed. It may reach as much as 50 per cent during some 
 portion of the stroke in some engines, and is always 25 per cent 
 with unprotected cylinders. This is reduced by wooden lagging 
 around the cylinders or steam circulating in an annular space 
 around the cylinder. Another remedy consists in superheating the 
 steam after leaving the boiler by passing it through the pipes 
 heated by flue-gases before delivering to the engine. 
 
 A better remedy is to divide the expansion into several stages 
 by compounding the cylinders. 
 In this the steam, after partial 
 expansion in one cylinder, is de- 
 15 livered to a second, and there 
 j expanded to the tower limit de- 
 termined by the pressure in the 
 condenser or the atmosphere (Fig. 
 FIG. 41. Combined Indicator Cards. 4i)- The action in the combined 
 
 cylinders is the same as in a 
 single cylinder with the same ratio of expansion. 
 
 The Compound Engine. These cylinders may be connected
 
 POWER GENERATION. 
 
 109 
 
 on the same piston-rod and have a common stroke (tandem), or 
 are side by side on different rods (cross- compound) and move 
 in opposite directions. If, however, the motion of the pistons 
 is to be at some other phase, a receiver is inserted between the 
 two cylinders, the exhaust from the high- pressure accumulating in 
 it before being discharged into the low-pressure cylinder. 
 
 Steam Exhausting 
 into.L.P. Steam Chest . 
 
 {Effective Pressure) ti. 
 pace A. A. A. A. is a Balanced Pressure, 
 between the two Piston Heads, 
 tTandem CompoitntLt'ofltaa 
 
 FIG. 42. The Section of a Compound Cylinder. 
 
 The compound engine economizes steam by permitting a 
 larger ratio of expansion, having smaller clearance volumes and 
 far less condensation in the cylinders. The cost of the com- 
 pound is, however, higher than the simple, its bulk and weight 
 are greater, but its steam consumption is very much less. 
 
 Governors. To control the engine and to economize steam 
 by supplying to the cylinder only the amount that is necessary 
 to do the work, two forms of gov- 
 ernors are employed. The steam 
 pressure may be reduced before ad- 
 mission to the cylinder by closing 
 the valve in the steam-pipe; or an 
 equally automatic regulation may 
 
 be had by reducing the volume of FIG. 43. Indicator Diagrams from a 
 
 the steam delivered to the cylinder Throttling steam-engine. 
 by hastening the point of cut-off. 
 
 In Fig. 43 are the diagrams obtained successively in the 
 throttling-engine as the steam-supply is reduced. Fig. 44
 
 HO MANUAL OF MINING. 
 
 shows the diagrams from a cut-off engine. The throttling is 
 obtained by a pendulum governor connected with the driving- 
 shaft of the engine, which closes the valve when the rate of revo- 
 lution exceeds a certain limit. The cut-off is operated by some 
 form of gearing connected with the shaft, and when the rate is 
 too high reduces the travel of the valve. The throttling-engine 
 is cheaper and suffers less from cylinder condensation than a 
 cut-off engine, but does not receive so sensitive or so uniform a 
 regulation. All engines have an independent throttling control 
 for the engineer. 
 
 Automatic Cut-off. The mechanism 'for shutting off the 
 steam-supply before the piston reaches the end of its stroke may 
 be fixed in position or variable, automatic, or adjustable. The 
 fixed cut-off is intended for constant duty, the governor being 
 set to shut off the steam-supply when the desired point is reached. 
 The variable cut-off is needed for a varying load during suc- 
 cessive strokes. In the high-speed 
 engine the shaft-governor varies the 
 length of travel of the slide-valve, 
 and thereby hastens cut-off when 
 the load is light. In the Corliss 
 ^*-i- engine the governor operates di- 
 
 FIG. 44 .Indicator Diagrams from rectly upon an arm which varies 
 
 an Automatic Cutoff Engine. ^ ^ Qf ^ Qff foy chang j ng ^ 
 
 tripping position. The latter two types are automatic in their 
 action, the Meyer being adjustable. Its rider-valves (Fig. 44) 
 can be set at each end at any time to the desired cut-off without 
 affecting the movement of the lower valve which determines the 
 duration of the exhaust. The Corliss valves are also adjustable 
 by changing the length of the rods from the wrist- plate to valve- 
 spindles. 
 
 Throttling governors are usually employed on all low-speed 
 engines and shaft-governors on high-speed engines. Engines 
 arranged to run "over" are smoother than those turning "under." 
 Facing the engine broadside with cylinder on the left and crank- 
 shaft at the right, right-handed revolution is called "over."
 
 POWER GENERATION. in 
 
 The Sliding Steam-valves. The valves for regulating the 
 admission and exhaust of the steam are of different types. Sim- 
 ple engines have one slide-valve to perform all the functions for 
 both ends. Such a valve admits of a cut-off not shorter than 
 0.4, and wastes considerable power in supplying the steam. The 
 sliding piston-valve (Fig. 144) is preferable, because it is 
 balanced and requires very little power to move it, but it does 
 not give a sharp cut-off. The inability of either of these to alter 
 the cut-off (B, Fig. 38) without also affecting the degree of 
 compression, DE, makes the Meyer cut-off valve (Fig. 124) a 
 better controller. Separate eccentrics drive the main slide- 
 valve and the rider-valve, the latter to control the point of cut- 
 off only, while the main valve regulates the exhaust and admits 
 steam when the rider uncovers its port. 
 
 The Corliss Valve consists of four rotary valves, each with a 
 single function to perform. One admission rotary valve and 
 one exhaust rotary valve serve for each end of the cylinder. They 
 are all moved by rods from a wrist-plate, the admission-valve 
 being provided with a releasing trip-gear to make a sharp cut- 
 off (Fig. 45). 
 
 FlG. 45. Section of Corliss Engine, showing valves in position for admitting 
 steam at the left. 
 
 Fig. 126 illustrates the Corliss engine and valve in section 
 on the left-hand air-cylinder. 
 
 The Reversing-link. The reversal of the direction of motion 
 <jf the engine is accomplished by one of two methods: engines 
 supplied with shaft-governors may be reversed in direction by 
 changing the position of an eccentric on the shaft and thus mov- 
 ing the valve to correspond; and slow-speed engines by the
 
 *I2 MANUAL OF MINING. 
 
 Stevenson link motion with two eccentrics (Fig. 46). Motion is 
 imparted to the steam- valve by a rod connected to an eccentric, 
 on the crank-shaft. The eccentric is placed at an angle of some- 
 thing more than 90 in advance of the crank-pin position. When 
 a reversal of motion is desired the single eccentric is turned to 
 an angle of more than 90 in advance on the other side of the 
 crank-pin. The Stevenson link arc has each end connected 
 with an eccentric When the arc is raised or lowered till the 
 block to which the valve is attached is at the extreme end of 
 the arc, the valve. will have a maximum travel. In position a 
 
 FIG. 46. The Reversing Link. 
 
 (Fig. 46) the block gives a forward motion, and when raised to 
 the position b a backward motion, to the valve. In its central 
 position the block is acted upon by both eccentrics and only a 
 slight motion is imparted to the valve. 
 
 The Condenser is a receiver fitted with tubes of copper 
 through which the steam may flow from engine to a tank or a 
 pump before delivery into the boiler. Around the tubes large 
 volumes of cold water circulate, condensing the steam and 
 creating in the tubes a vacuum, which is further maintained by 
 the air-pump. The vacuum obtained in practice is about 25 
 inches of mercury column below the atmosphere, or an absolute 
 pressure of 2.5 Ibs. per square inch. This addition to a plant 
 economizes power by reducing the back pressure, increasing 
 correspondingly the net effective pressure, as may be seen by 
 comparing the indicator cards (Figs. 38 and 47), by nearly 15 Ibs. 
 per square inch. This is the quickest and simplest means of
 
 POWER GENERATION. 113 
 
 increasing the power of a given engine and boiler plant. The 
 saving in steam consumption is indicated for Corliss engines 
 in a later table. 
 
 FlG. 47. Indicator Diagram from a Condensing Engine or the Low-pressure 
 Cylinder of a Compound Engine. 
 
 For every 100 feet of piston speed the condenser will save 
 1.03 H.P. in an engine whose cylinder is 6 inches in diameter; 
 with a lo-inch cylinder, 2.86 H.P. is saved per 100 feet of veocity; 
 with a i6-inch cylinder, 7.31 H.P.; and for 20 inches diameter, 
 11.42 H.P. Thus an engine with io"Xi6" cylinder making 
 80 r.p.m. (213.33 f ee t piston speed) saves 6.1 H.P. 
 
 Influence of Clearance. The clearance space at each end of 
 a cylinder prevents the piston striking the heads and provides a 
 space in which steam accumulates to form a cushion and assist 
 in the reversal of piston movement. A large clearance is not 
 economical. Hence high-speed and slide-valve engines are less 
 economical than the Corliss. A large clearance implies a small 
 degree of compression in the cushion and less back pressure. 
 
 The most economical mean effective pressure is that corre- 
 sponding to a minimum steam consumption. When the steam 
 supply is cut off at a point of stroke that will permit it to expand 
 to the pressure of the condenser, a complete expansion is obtained. 
 The more complete the expansion the less will be the mean effec- 
 tive pressure. With a low M.E.P. a higher economy is attained 
 than from a similar engine with a high M.E.P. A low pressure 
 means large cylinders and excessive condensation. The degree 
 of expansion and the number of cylinders are fixed by the ratio 
 of the allowable boiler pressure to that of the exhaust into con- 
 denser or atmosphere. 
 
 An engine with a fixed load is operating at its maximum effi- 
 ciency when driving its calculated load. If the load varies, the
 
 JI4 MANUAL OF MINING. 
 
 -engine must have a capacity for the maximum demand. If this 
 load is casual and seldom demanded, the engine may be then 
 forced. Its size will be determined by the demand of the average 
 load, in supplying which it may attain its highest efficiency. 
 
 With a single- valve simple engine the most economical M.E.P. 
 is about one half of the boiler pressure, which would represent a 
 steam consumption of about 28 Ibs. and a M.E.P. of 50 Ibs. per 
 square inch. A simple Corliss has its maximum economy at a 
 M.E.P. of 35 Ibs. pressure and a Corliss compound at 26 Ibs. 
 The steam consumption would be 25.3 and 19 Ibs. respectively. 
 
 Selection of Engine. Generally speaking, any expenditure for 
 a plant is warranted which will show a saving more than equal to 
 the interest on the expenditure, the depreciation and repairs of 
 the plant. But this balance is not easily obtained. An economic 
 engine is one with a low M.E.P., large cylinders, and costly at- 
 tachments. But its initial cost may be prohibitive at the time of 
 Its installation. If the fuel is cheap and the amount of repairs 
 not known, a high-pressure plant may not be desirable, particu- 
 larly as its piping system usually gives trouble and the wear and 
 leakages are large. Nevertheless the saving resulting from its use 
 generally more than offsets the increased cost. The smaller 
 boiler capacity and the diminishment of its cost should be con- 
 sidered with the other items of saving. 
 
 The Speed of Engines. Engines are classified according to 
 their piston speed, as slow-speed, averaging 100 feet per minute, 
 and high-speed, making nearly 900 feet of travel per minute. 
 This value is the product of the number of revolutions and the 
 piston stroke two independent variables. 
 
 The Corliss engine is limited to about 60 r.p.m., and electric 
 engines for lighting, etc., make about 400 revolutions. For 
 mining purposes the rate rarely exceeds 100 r.p.m. The high 
 rotary speed is not economical, because of the large clearance 
 at each end of the cylinder. The excellent workmanship and 
 the great care which the high-speed engines require make their 
 first cost and subsequent maintenance large. They have less 
 steam condensation than slow engines.
 
 POWER GENERATION. 115 
 
 The work of the piston is equal to the continued product of 
 the mean "effective pressure upon It, the stroke in .feet, its area in 
 square inches, and the number of strokes If the speed is fixed 
 for a given engine, the area may be designed proportionately to 
 s, the mean effective pressure being altered at will (by changing 
 the initial boiler pressure or the cut-off). 
 
 The Horse-power of an Engine. The theoretical horse-power 
 is the product of the work of the piston and the number of strokes 
 divided by 33,000. The area of the -piston-rod on the crank side 
 is neglected, as also the losses from friction of the engine parts, 
 steam condensation in the cylinder, clearance, and compression. 
 The value for the effective pressure, p, is obtained from the 
 formula on a previous page. 
 
 The indicated horse-power (I.H.P.) is the actual work of 
 the steam as revealed by the indicator card; the mean effective 
 pressure is obtained (Fig. 39) by the card. It is less than 
 p, obtained above by a coefficient, E, whose value is less than 
 unity. 
 
 The "brake" horse-power (B.H.P.) is a certain fractional 
 part, m, of the indicated, which allows for the mechanical losses 
 and for intermittence in running. The value for the modulus 
 varies from 0.60 in a very poor engine to 0.90 in one in good 
 working order. A fair average for the mechanical efficiency, 
 m, of the engine ratio between the brake horse-power and the 
 indicator horse-power is 0.8. 
 
 Let s =the length of the piston stroke in inches; 
 &=the diameter of cylinders in inches; 
 N =the number of revolutions per minute; 
 o.i66./Vs=the piston speed, feet per minute; 
 c=the number of cylinders; 
 /=the coefficient of friction; 
 
 ^=mean effective steam pressure, pounds per square inch; 
 P= absolute steam pressure entering the cylinders; 
 r = ratio of expansion or per cent of cut-off; 
 B = absolute back pressure of the steam during exhaust; 
 I?=horse-power developed by the engine.
 
 MANUAL OF MINING. 
 
 Then 
 
 H = (o.ooooo3g66csk 2 N)pm. 
 
 In the following table are calculated, for certain sizes of 
 cylinders and given piston speeds, the values for the parenthetical 
 expression. 
 
 THE HORSE-POWER FOR EACH POUND OF MEAN EFFECTIVE PRESSURE PER 
 SQUARE INCH OF PISTON AREA. 
 The value for o.ooooo3966s 2 ./V. 
 
 Diam- 
 
 Average Speed of Piston in Feet per Minute. 
 
 eter of 
 
 
 Cylin- 
 
 
 
 
 
 
 
 
 
 
 
 der. 
 
 240 
 
 300 
 
 350 
 
 400 
 
 450 
 
 Soo 
 
 550 
 
 600 
 
 650 
 
 750 
 
 4 
 
 .091 
 
 .114 
 
 133 
 
 .162 
 
 .171 
 
 
 
 
 
 
 5 
 
 .144 
 
 .18 
 
 .21 
 
 .24 
 
 27 
 
 
 
 
 
 
 6 
 
 .205 
 
 .256 
 
 .299 
 
 342 
 
 .385 
 
 .428 
 
 .471 
 
 
 
 
 7 
 
 .279 
 
 348 
 
 .408 
 
 .466 
 
 524 
 
 .583 
 
 .641 
 
 
 
 
 8 
 
 .365 
 
 .456 
 
 532 
 
 .608 
 
 .685 
 
 .761 
 
 .837 
 
 .912 
 
 
 
 9 
 
 .462 
 
 577 
 
 .6 74 
 
 .770 
 
 .866 
 
 .963 
 
 1.059 
 
 -154 
 
 
 
 10 
 
 571 
 
 .714 
 
 .833 
 
 952 
 
 1.071 
 
 1.190 
 
 1.309 
 
 .428 
 
 i-547 
 
 I-785 
 
 it 
 
 .691 
 
 .864 
 
 I.OOS 
 
 I.I52 
 
 1.296 
 
 1.44 
 
 1.584 
 
 .728 
 
 1.872 
 
 2. 1 60 
 
 12 
 
 .820 
 
 1.025 
 
 I 195 
 
 1.366 
 
 1.540 
 
 1.708 
 
 1. 880 
 
 .050 
 
 2.222 
 
 2.564 
 
 14 
 
 1.119 
 
 1.398 
 
 1.631 
 
 1.864 
 
 2.097 
 
 2-331 
 
 2.564 
 
 797 
 
 3.029 
 
 3-495 
 
 15 
 
 1.285 
 
 i. 606 
 
 1.873 
 
 2.I3I 
 
 2.409 
 
 2.677 
 
 2-945 
 
 3.212 
 
 3-479 
 
 4.004 
 
 16 
 
 1.461 
 
 1.827 
 
 2.I3I 
 
 2.436 
 
 2.741 
 
 3-45 
 
 3-349 
 
 3-654 
 
 3.958 
 
 4-567 
 
 18 
 
 1.849 
 
 2.312 
 
 2.697 
 
 3-083 
 
 3.468 
 
 3.854 
 
 4.239 
 
 4.624 
 
 5.009 
 
 5-/8 
 
 20 
 
 2.292 
 
 2-855 
 
 3-331 
 
 3-807 
 
 4.285 
 
 4-759 
 
 5-234 
 
 5-73i 
 
 6.186 
 
 7-I38 
 
 23 
 
 3.021 
 
 3-776 
 
 4.104 
 
 5-35 
 
 5.664 
 
 6.294 
 
 6.923 
 
 7-552 
 
 8.181 
 
 9-44 
 
 EXAMPLES. i. Required the horse-power of an engine 6"X 10" with a pis- 
 ton spaed of 300 feet per minute (183 revolutions) with M.E.P. of 40 Ibs. 
 According to the table, for k = 6 inches and the piston speed of 300 feet, a, 
 constant of 0.256 is found. For 40 Ibs. pressure and w=o.8o the horse- 
 power, #"=0.256X40X0.80=8.19. 
 
 2. Required the horse-power of a duplex engine io"Xi4.4" making 100 
 r.p.m. and having a cut-off of \ with 7 per cent clearance and initial admis- 
 sion pressure of 100 Ibs. absolute. 
 
 Let m=o.8o and back pressure 18 Ibs. according to table on page 105. 
 Then />= 0.6861(100) -18=50,64. C=o.686i. 
 
 According to the above table the constant for ic-inch cylinder and 240 
 feet piston speed is 0.5 71. Whence #"=0.571X50.64X0.80=23.13 for each 
 cylinder and 46.16 for the duplex engine. 
 
 Usually there is an approximate ratio between the diameter, k, 
 and the length, s, of a cylinder. .In a long- stroke engine s = 2k to 
 2.56. A short-stroke engine has a ratio s=k to s = i.2k.
 
 POWER GENERATION. 117 
 
 The Steam Consumption of an Engine. A one-horse-power 
 engine exercised through one hour, equals 1,980,000 ft.-lbs.; one 
 B.T.U. =778 ft. Ibs.; hence one horse-power hour corresponds 
 to 2545 B.T.U. A pound of steam represents about noo B.T.U. 
 above feed-water temperature. Hence one horse-power should 
 be obtained during one hour from 2.3 Ibs. of steam. The very 
 best on record is 12.2 Ibs. per I.H.P., and for a stationary engine 
 of the Corliss type, compound, about 18 Ibs. Hence all steam con- 
 sumption above 12.2 Ibs. may be regarded as preventable waste. 
 
 The following table shows the consumption of steam per 
 hourly horse-power with non-condensing and condensing engines 
 at various rates of cut-off and different boiler pressures. An 
 allowance is here made for ordinary cylinder losses. The slide- 
 valve engine would show a steam consumption under similar 
 conditions of nearly twice those indicated in the table. 
 
 In a slide-valve engine 22"X28" at 118 r.p.m. with a boiler 
 pressure of no Ibs. absolute, clearance at 8 per cent, a mean 
 effective pressure of 43.1 Ibs. is obtained. By calculation from 
 the indicator diagram the steam consumption was 0.64703 Ibs. 
 per stroke. The same size of Corliss engine with a j cut-off and 
 3 per cent clearance used 0.32635 Ibs. of steam per stroke. Each 
 engine making 14,160 strokes per hour, the steam consumption 
 will be, in the two cases, respectively 33.5 and 17.08 Ibs. per horse- 
 power per hour. The Corliss therefore saves nearly one-half the 
 steam consumption of the slide-valve engine. A zoo-H.P. Corliss 
 will annually save 150 tons of coal over a slide-valve engine. 
 
 Setting the Engine. A foundation is prepared for the engine 
 before its arrival. Two accurately constructed templates, ordered 
 from the manufacturers, are laid in proper position with bolts 
 connecting them long enough to reach from the bottom to the top 
 of the proposed masonry foundation. Upon the lower one a 
 foundation of concrete^ or brick is built, without disturbing the 
 bolts, to the level of the bed-plate, which is finished and adjusted 
 ready for the engine. The upper template may be removed and 
 the engine at once dropped into position without delay. Tim- 
 ber foundations are used only for second-motion haulage-engines
 
 n8 
 
 MANUAL OF MINING. 
 
 THE MEAN EFFECTIVE PRESSURES AND TERMINAL PRESSURES AND STEAM CON- 
 SUMPTION IN CORLISS NON-CONDENSING AND CONDENSING ENGINES. 
 
 
 % Cut-off. 
 
 \i Cut-off. 
 
 Initial 
 
 M.E.P. 
 
 Termi- 
 
 Water Con- 
 
 M.E.P. 
 
 Termi- 
 
 Water Con- 
 
 r^res- 
 
 
 nal 
 
 sumption. 
 
 
 nal 
 
 sumption. 
 
 sure. 
 
 
 Pres- 
 
 
 
 Pres 
 
 
 
 N. C. 
 
 C. 
 
 sure. 
 
 N. C. 
 
 C. 
 
 N. C. 
 
 C. 
 
 sure. 
 
 N. C. 
 
 C. 
 
 60 
 
 24.24 
 
 36.24 
 
 16.08 
 
 2 7 .6 
 
 20.5 
 
 29.56 
 
 41.56 
 
 19.77 
 
 27-3 
 
 21.2 
 
 65 .... 
 
 26.93 
 
 38.93 
 
 17.15 
 
 26.5 
 
 20.3 
 
 32-61 
 
 44.61 
 
 21.09 
 
 26.4 
 
 21.0 
 
 70. ... 
 
 29.63 
 
 41.63 
 
 18.23 
 
 25-7 
 
 20.1 
 
 35-67 
 
 47.67 
 
 22.41 
 
 25-8 
 
 20.8 
 
 75---- 
 
 32-32 
 
 44-32 
 
 I 9-3 I 
 
 24.0 
 
 19.9 
 
 38-72 
 
 50.72 
 
 23-73 
 
 25.2 
 
 20.6 
 
 80.... 
 
 35-02 
 
 47-02 
 
 20-39 
 
 24-3 
 
 19.8 
 
 41.78 
 
 53.78 
 
 25-05 
 
 24-8 
 
 20.4 
 
 80.... 
 
 37-73 
 
 49.71 
 
 21.46 
 
 23-7 
 
 10.7 
 
 44.83 
 
 56-83 
 
 26.37 
 
 24-3 
 
 20.3 
 
 90.... 
 
 40.41 
 
 52.41 
 
 22-54 
 
 23.4 
 
 19-5 
 
 47-89 
 
 59.89 
 
 27.69 
 
 24.0 
 
 20.3 
 
 95 
 
 43.10 
 
 55-io 
 
 23.62 
 
 23.0 
 
 19.4 
 
 50.94 
 
 62.94 
 
 29.01 
 
 23.6 
 
 20.2 
 
 too 
 
 45.80 
 
 57-8o 
 
 24.70 
 
 22.6 
 
 19.4 
 
 54.00 
 
 66.00 
 
 30-33 
 
 23.1 
 
 20.1 
 
 
 %o Cut-off. 
 
 s"Hoo Cut-off. 
 
 60... 
 
 34-21 
 
 46.24 
 
 23-37 
 
 27.4 
 
 22.0 
 
 38.18 
 
 50.68 
 
 27.04 
 
 28.3 
 
 22.8 
 
 65.... 
 
 37-6: 
 
 49.61 
 
 24.94 
 
 26.7 
 
 21. 9 
 
 41.80 
 
 53-80 
 
 28.85 
 
 27-5 
 
 22.6 
 
 70 
 
 40.98 
 
 52.98 
 
 26. 51 
 
 26.1 
 
 21.7 
 
 45-42 
 
 57-42 
 
 30.66 
 
 26.8 
 
 22-5 
 
 ::;: 
 
 44-35 
 47.72 
 
 59-72 
 
 28.07 
 29.64 
 
 25.6 
 25-1 
 
 21-5 
 
 21.3 
 
 49 5 
 52.68 
 
 61 .05 
 64.68 
 
 32-47 
 34-28 
 
 26.2 
 25-7 
 
 22.4 
 22-3 
 
 85 .... 
 
 51.09 
 
 49.71 
 
 31.20 
 
 2 4 .6 
 
 21.2 
 
 56-31 
 
 68.31 
 
 36.09 
 
 25-3 
 
 22.2 
 
 90 
 
 54-46 
 
 66.46 
 
 32-77 
 
 24.2 
 
 21.0 
 
 59-94 
 
 71.94 
 
 37-90 
 
 25.0 
 
 22.2 
 
 95.... 
 
 57-83 
 
 69.83 
 
 34-33 
 
 23-9 
 
 20-9 
 
 63-57 
 
 75-57 
 
 39-71 
 
 24-7 
 
 22. 
 
 100.... 
 
 61.20 
 
 73.20 
 
 35-90 
 
 23-7 
 
 20.8 
 
 67.20 
 
 79.20 
 
 4L52 
 
 24-5 
 
 21. 9 
 
 
 tto Cut-off. 
 
 Yz Cut-off. 
 
 60.... 
 
 41.68 
 
 53-68 
 
 30.66 
 
 28.8 
 
 23-6 
 
 47-50 
 
 56.50 
 
 37.98 
 
 29.4 
 
 25.2 
 
 65.... 
 
 45-54 
 
 57-54 
 
 32-71 
 
 28.2 
 
 23.4 
 
 51-75 
 
 63-75 
 
 40-52 
 
 28.8 
 
 25.0 
 
 70. .. 
 
 49.41 
 
 61.41 
 
 34-77 
 
 2 7 .6 
 
 23-2 
 
 56.00 
 
 68.00 
 
 43-07 
 
 28.4 
 
 2 4 .8 
 
 75 
 
 53-27 
 
 65-27 
 
 36.82 
 
 27.1 
 
 23.0 
 
 60.25 
 
 72.25 
 
 45.61 
 
 27.9 
 
 24.7 
 
 80.... 
 
 57-14 
 
 69.14 
 
 38.88 
 
 26.7 
 
 22.9 
 
 64.50 
 
 76.50 
 
 48.16 
 
 2 7 .6 
 
 24-6 
 
 85.... 
 
 61.00 
 
 73-oo 
 
 40-93 
 
 26.1 
 
 22.8 
 
 68.75 
 
 80.75 
 
 50.70 
 
 27.2 
 
 24-5 
 
 90 
 
 64.87 
 
 76.87 
 
 42-99 
 
 25.8 
 
 22.6 
 
 73-oo 
 
 85.00 
 
 53-25 
 
 27.0 
 
 24.4 
 
 95 
 
 68.73 
 
 80.73 
 
 45-04 
 
 2 5-5 
 
 22.5 
 
 77-25 
 
 89.25 
 
 55-79 
 
 26.8 
 
 24.2 
 
 100 
 
 72.60 
 
 84.60 
 
 47.10 
 
 25.2 
 
 22.4 
 
 81.50 
 
 93-5 
 
 58.34 
 
 26.6 
 
 24.1 
 
 These, reduced to coal consumption at the rate of 10 Ibs. of steam per pound 
 of coal, can be at once referred to the coal-pile. 
 
 The Rotary Effort. The "horse-power" of the engine as- 
 sumes a uniform effort of the engine against the resistance, but 
 owing to varying angles occupied by the crank-arm, which receives 
 the thrust from the connecting-rod, the rotary effort upon a
 
 POWER GENERATION. IIQ 
 
 crank-pin varies from one dead-center to a maximum and thence 
 diminishes to a minimum during each half of the revolution or 
 each stroke. Owing to this variation in effort the cylinders 
 are usually duplicated side by side, their pistons operating upon 
 their respective cranks 90 apart. This enables the piston of 
 one cylinder at the middle of its stroke to assist the piston start- 
 ing its stroke. 
 
 The governors, discussed above, check variation of speed 
 from stroke to stroke but not during each single stroke. The 
 variations in the rate of rotary effort upon the crank-pin must be 
 kept within reasonable limits by use of a fly-wheel. There is 
 a varying pressure on the piston due to expansion, a varying 
 thrust upon the connecting-rod due to its angularity, and a 
 varying tangential effort upon the crank-pin due to the angular 
 position of the crank-arm. These variations cause a fluctuation 
 in the speed of the crank-pin, the crank-shaft, and of the hoister- 
 drum which is attached to it. The shorter the connecting-rod 
 the more nearly equal are the forces throughout the stroke and 
 in each half -re volution. A diagram drawn to represent the 
 tangential effort upon the crank-pin by vertical lines and the 
 rectilinear motion of the crank-pin by the horizontal distances; 
 the curve shows the varying efforts, or torque, of the engine. 
 If diagrams be prepared representing the rotary effort of each 
 piston of a duplex engine and be superposed upon one another, 
 at the same time advanced 90 to represent the difference in 
 the angle of the crank-arms, the sum of the ordinates at any 
 one point represents the aggregate rotary effort of two pistons 
 upon the crank-shaft. 
 
 An engine must be capable of starting from any position of 
 its piston, or of its corresponding crank-pin, and have at that 
 instant a rotary effort in excess of the resistance on the crank- 
 arm. 
 
 The Steam-turbine may be employed for purposes of a motor 
 engine where a high rate of revolution is desired, as, for example, 
 with a centrifugal pump or an electric generator. For these pur- 
 poses it is particularly adapted.
 
 120 MANUAL OF MINING. 
 
 The steam-turbine is a wheel fitted at its circumference with 
 numerous blades receiving the impact of a jet of steam flowing 
 against them at a high velocity. The turbines may be classified 
 according to the path of the steam, which may be radial, outward 
 or inward, parallel flow or spiral flow. The turbine utilizing 
 the steam which flows in a direction at a constant distance from 
 the axis of the wheel is the one most highly developed at the pres- 
 ent time. Of these the De Laval single wheel and the Parsons, 
 Curtis, and Rateau multiple wheels are the types. 
 
 The De Laval Wheel. In this a jet of steam is driven through 
 a nozzle, D (Fig. 48), where it expands to the atmospheric pres- 
 sure, thus creating a high velocity of flow whose energy is imparted 
 to the blades of the wheel producing a peripheral speed of 30,000 
 feet per minute. This wheel, being 8" or more in diameter, 
 makes 15,000 or more revolutions per minute. This rate being 
 too high for motors, a reduction gear is employed. Pumps and 
 electric generators are attached to the shaft, making 1000 or 
 more revolutions. In Fig. 48 are shown also details of its 
 construction. 
 
 The Parsons Steam-turbine. This is a series of rotary discs 
 fitted with blades around their circumference (Fig. 49). Alter- 
 nating with the discs are rings of fixed blades projected inward 
 from the casing. The steam is discharged from a nozzle, thence 
 to the fixed guide-blade, which deflects it to the rotary blade. 
 This, after absorbing some of the power, again discharges the fluid 
 to the next guide-blade, etc. Thirty or more such passages are 
 encountered before exhaust, and each one of the 16,000 rotary 
 blades receives a small amount of pressure, the aggregate of 
 which produces the effective rotation of the shaft. The discs 
 are fitted on the one axle and revolve together, making 400 revolu- 
 tions per minute. 
 
 The Curtis and Rateau types of wheels are very similar, but 
 occupy intermediate positions between those already mentioned, 
 as to the method of utilizing the energy of the steam. 
 
 The Advantages of the Steam-turbine. It is a rotary machine, 
 requiring very small floor space and little or no foundation. It
 
 POWER GENERATION. 
 
 121 
 
 is much lighter in weight than is a steam-engine delivering the 
 same amount of power. These steam-turbines are employed at 
 pressures not to exceed 150 Ibs. gauge. The steam should be dry 
 and to a moderate degree is superheated before being delivered 
 to the turbine. Condensers may be added to increase their 
 power. 
 
 Its steam consumption is about the same as that of a steam- 
 engine of equal power. There being no internal lubrication, 
 
 FIG. 48. De Laval Wheel Deta'ls. 
 
 the steam does not come into contact with a running part and 
 its condensed exhaust can be returned to the boiler free from oil. 
 This is a merit of great importance where the supply of feed- 
 w r ater is chemically bad. There are no complicated valves 
 to give occasion for injury, and no pistons to leak. All bear- 
 ings are easily lubricated. The only \vearing parts are the 
 nozzles and the blades. Turbines can be had in power from 
 those of 10 horse-power up to one of 7400 I.H.P. 
 
 In point of efficiency and commercial utility it has the field, 
 which will soon be undisputed. Its thermodynamic efficiency 
 is equalled only by the gas- or oil-engine.
 
 123 
 
 MANUAL OF MINING. 
 
 v
 
 POWER GENERATION. 123 
 
 Oil-engines. When the distances are too remote for econom- 
 ical admission of steam or compressed air through the pipes or 
 of electricity by wires to pumps and fans, independent oil- or gas- 
 engines are employed. These are self-contained and are placed 
 on the same foundation with the machine to be driven, and may 
 even be portable on a truck. These engines require little atten- 
 tion and have a high efficiency in addition to the advantage of 
 portability. The motor fluid is gasoline, naphtha, or alcohol, 
 which is held in a supply-tank from which, by means of a small 
 pump attached to the engine, it is conducted to the working 
 cylinder after having been atomized or vaporized. It is then 
 mixed with air in proportions regulated by a suitable device. 
 This vaporous mixture is drawn into the cylinder during the 
 first stroke of the piston, compressed into a large clearance space 
 in the second stroke, at the end of which the mixture is ignited 
 and expanded, the explosive force of which drives the piston 
 forward on its third and outward stroke. On the return or 
 fourth stroke the products of combustion are expelled into the 
 atmosphere. The operation is then repeated, there being but 
 one working stroke out of each four strokes or every two revolu- 
 tions of the fly-wheel? This constitutes what is known as the 
 four-cycle engine (Fig. 50). Two or more such cylinders may 
 be attached on the same shaft, with their cranks at various angles 
 to produce uniform rotary effort A very heavy fly-wheel is 
 essential to carry the engine over the three non-working strokes. 
 It is started by hand and thereafter requires little attention. Its 
 thermodynamic efficiency exceeds that of the steam-engine, and 
 it is useful where the power must be continuous, as for electric 
 generation or pumps. Fig. 147 illustrates one of these internal- 
 combustion engines of the two-cycle type, the compression being 
 conducted outside of the cylinder. 
 
 It is essential that the amount of air be chemically perfect for 
 complete combustion, when the initial pressure attained will 
 be 200 or 300 Ibs. per square inch. The temperature of the 
 explosion is over 1000 F. The cylinder is cooled by the water 
 circulating in a jacket, and the exhaust- gases are delivered into
 
 124 
 
 MANUAL OF MINING. 
 
 the stack. The former absorbs from 40 to 50 per cent of the 
 heat developed in the combustion, and the exhaust carries off 30 
 to 40 per cent more, leaving about 20 per cent for effective work. 
 The oil-consumption per horse-power hour is from to of a 
 gallon. 
 
 Ignition may be attained by an electric spark or a tube kept 
 red-hot by a flame during the entire running. The efficiency of 
 
 FIG. 50. The Four Cycles of a Gas- or Oil-engine. 
 
 the oil-engine increases with the increase of the degree of com- 
 pression at the end of the second stroke. It follows therefore 
 that the Diesel motor, which raises the compression to 700 Ibs. 
 per square inch nearly the point of ignition of the gases gives 
 the highest results (Fig. 147). One I.H.P. is obtained from the 
 consumption of 0.3 Ib. oil. 
 
 The regulation of these internal- combustion engines is obtained 
 by a governor, which produces a missfire when the engine is under- 
 loaded.
 
 POWER GENERATION. 125 
 
 REFERENCES, 
 
 Boiler Erection, Adjustment, Am. Mir., July 23, 1897, 115; Water Tubu- 
 lar Boilers, Superiority, Am. Mfr., Jan. 23, 1897, 113; Root Safety Boiler, with 
 Tables of Conversions, Catalogue; Boiler-room Practice, W. T. Edwards, 
 Am. Elect'n, Feb. 1904; Corrosion of Boilers, H. Ost, Engng., Oct. 10, 1902; 
 Boiler Scale and the Transmission of Heat, Locomotive, Sept. 1002; Water- 
 tube Boilers, Sci. Am. Sup., Dec. 27, 1902, Jan. 3, 1903, and William Fyvie, 
 Aust. Min. Strand., Oct. i, 1903; Revue Technique, Feb. 25, 1900. 
 
 Selecting a Boiler Feed, W. H. Wakeman, Mod. Mach., Mar. 1900; 
 Comparative Methods of Heating Feed-water, Power, Nov. 1902; Feed- 
 water Purification, Am. Elect'n, Mar. 1900. 
 
 Coal Storage, Waldron Fawcett, Sci. Am., Oct. 18, 1902; The Valuation 
 of Coals, E. L. Rhead, Ir. & Coal Trds. Rev., April 3, 1003; Coals: Their 
 Sources and Heating Values, John W. Langley, Engrs. U. S. A., April i, 
 1903; Spontaneous Ignition of Coal, Mines & Min., Dec. 1899; Coal Con- 
 sumption in Central Stations, Alfred S. Giles, Mech. Engr., Jan. 23, 1904; 
 Smokeless Firing, Colliery Manager, Jan. 17, 1896, 24. 
 
 Mechanical Stokers, Edwin Titts, Pro. Engrs. Soc. of W. Penn., April 
 1903; Comparative Economy of Stoking and Hand Firing, Julius Geldard, 
 Mech. Eng., Jan. 3, 1903; Mechanical Stokers Depending upon the Coking 
 Method of Firing, Albert A. Gary, Engr. U. S. A., April i, 1903; Surface 
 Condensers, C. G. Robbins, Marine Engng., Oct. 1902; The Cost of Steam 
 Raising, John Holliday, Elec. Engr., Lond., Nov. 24, 1899; Mechanical 
 Draft, J. I. Lyle, Engr. U. S. A., Feb. 15, 1904. 
 
 Notes on Machinery Constituting a Mining Plant, Alfred C. Cards, Can. 
 Min. Rev., Sept. 30, 1902; The Development of Power-driven Machinery 
 in the Mine, E. H. Roberton, Eng. Mag., Mar. 1904; The Designing of Steam- 
 power Plants, Walter C. Kerr, St. Ry. Jour., Oct. 4, 1902. 
 
 The Curtis Steam Turbine, Engng., Feb. 5, 1904; The Steam Turbine, 
 William Chilton, Prac. Engr., Feb. 12, 1904; The Turbine Problem, H. F. 
 Schmidt, Am. Elect'n, Feb. 1904; The Curtis Steam Turbine, W. L. R. Emmet, 
 St. Ry. Jour., April n, 1903; A New Steam Turbine, Elec. Rev. N. Y., 
 April n, 1903; The Rateau Turbine, St. Ry. Jour., April 18, 1003; The 
 Steam Turbine: Its Commercial Aspect, Edwin H. Sniffin, St. Ry. Rev., 
 Oct. ii, 1902; The Westinghouse Steam Turbine, Edwin Yawger, Am. Mfr., 
 Oct. 23, 1902; Problems in Hauling and Hoisting, A. Brown, A. I. M. E., 
 Vol. XXXI, 265; and a serial on Steam-turbine Design, The Engineer, 1904. 
 
 Slide Valves, Haswell, 731; Automatic Valves Controlling Engines, Coll. 
 Eng., April 1896, 199; Indicator Diagrams, Calculation of, Coll. Eng., Feb. 
 1897, 320; Steam-engine Breakdowns, Coll. Guard., Nov. 13, 1896, 930; 
 Shafts, Cylinders, Valves, Coll. Guard., Nov. 20, 1896, 965; Tables, etc.,
 
 126 MANUAL OF MINING. 
 
 Coll. Guard., Nov. 20, 1896, 065; Setting Corliss Valves, Engineer, Chicago, 
 June 15, 1903; Setting Corliss Valves, The Engineer, June 15, 1903; Setting 
 Corliss Valves, The Engineer, June i, 1904. 
 
 Tables of Flow of Steam and Drop in Pressure through Pipes, The Engi- 
 neer, Oct. 15, 1902, 697; Governors and Governing Mechanism, Herbert 
 Reed Hall, Prac. Eng., Oct. 6, 1902; Hints on the Construction of Shaft 
 Governors, C. B. Risley, Power, April 1903. 
 
 Petroleum Engines, Priestman, Colo. S. of M. Quart., I, 1893, 65; Cur- 
 rent Practice in Gas-engine Design, H. Lee Koenig, and G. W. Rice, Sib. 
 Jour, of Engng., June 1903; The Balancing of Locomotive Engines, Engr., 
 Lond., Feb. 28, 1902. 
 
 Mechanical Engineering of Collieries, Serial, J. Futers, Coll. Guard., VoL 
 LXXXVLLI, 114; "Power Plants" and Details, A. Meyer.
 
 CHAPTER V. 
 
 HOISTING MACHINERY. 
 
 Hoisting Machinery. The various mechanisms used for 
 hoisting minerals in the mines are the windlass, the whim, the 
 whip, and the engine. The first-named is employed for moder- 
 ate depths of mine and in auxiliary underground openings where 
 neither the weight to be hoisted nor its distance is very great. 
 The whim and whip usually represent the next stage in the de- 
 velopment of a metalliferous mine, when the limit of the former 
 hand methods has been exceeded. These are driven by a horse, 
 or team of horses, and are employed for depths not exceeding 
 200 feet. The steam-engine is the motor for all mines having 
 a large output or a great depth of shaft, and even for a depth 
 of but 100 feet is still more economical than the simpler methods 
 mentioned above. 
 
 Windlasses are employed for shafts of moderate depth and for 
 winzes underground where manual labor is depended upon for 
 hoisting purposes. The amount that can be raised by two men 
 100 feet in eight hours cannot exceed 4 tons, allowing for delays, 
 etc. The windlass is a barrel, the wooden cylinder, 6 inches 
 to 10 inches in diameter, long enough to reach across the shaft, 
 raising by its axle on uprights, and operated at each end by cranks 
 15 inches long set at right angles to each other. Winches and 
 crabs or windlasses, in iron, can also be had in every possible 
 combination, but the simpler the machine the less is the friction, 
 and the more acceptable it is. Each additional gearing in- 
 volves a large percentage of loss, and there is little to spare from 
 the average man-power of 5300 ft.-lbs. per minute. Windlasses 
 
 127
 
 128 MANUAL OF MINING. 
 
 are, however, useful for incidental purposes in handling heavy 
 pieces of timber, machinery, and pump-pipes. A given power can 
 only equal a certain product of weight and velocity; for an in- 
 crease of speed the weight must be proportionately diminished. 
 With a 15-inch crank-arm, 12 revolutions per minute, coil 25 feet 
 of rope on an 8-inch barrel, and with this speed of hoist the greatest 
 load that may be moved under the circumstances by an average 
 laborer is 214 Ibs. Friction and stiffness of the rope will reduce 
 this. A i5o-lb. load can be raised at a speed of only 35 feet per 
 minute. 
 
 Balancing the Loads. The lengthened windlass barrel re- 
 ceives several coils of rope, from the ends of which two buckets 
 are suspended, the one falling to balance that rising. The length 
 of rope is a few coils only greater than the depth of the shaft. 
 At the start the weight to be hoisted is only the contents of the 
 tub plus the rope. This weight diminishes in rising, till at the 
 top it is the contents minus the rope. This does not, how- 
 ever, obviate the great stress from inertia which arises at 
 the moment of starting. For this reason single or double conical 
 barrels are used, on which rope is coiled in such manner that 
 the empty bucket is hung from the larger diameter, while the 
 rope from the loaded tub at the bottom is wrapped around the 
 smaller end. The tubs balance each other, but the empty acts 
 with a greater leverage than the loaded tub, and thus assists 
 the power in overcoming the inertia. After the hoisting is under 
 way, the empty and its lengthening rope uncoils with diminish- 
 ing leverage, while the load with its shortening rope gradually 
 winds on a larger diameter of the cone The buckets do not 
 meet in the middle of the shaft, but in the middle of the number 
 of revolutions of the barrel. In any event the pitch of the cone 
 must be calculated for the given conditions of depth and Toad, 
 otherwise its advantage is manifest only at the start. When 
 properly constructed the conical drum is to be recommended. 
 Not many attempts are made to apply this mechanism to hand 
 hoisting, but it is rapidly coming into favor with horse- and engine- 
 power, notwithstanding that it is dearer than cylindrical drums.
 
 HOISTING MACHINERY 129 
 
 Manual labor is manifestly too expensive to be regarded as 
 any but a temporary expedient for hoisting. 
 
 Whims. When the height of he hoist exceeds 60 feet or the 
 output 5 tons per shift, horse-power is employed to advantage. 
 The average horse develops an effort of 135 Ibs. when walking at a 
 speed of 180 feet per minute. This is much below the theoretical 
 value fixed for a horse-power, yet it represents the results of many 
 tests upon the energy of the animal, which will raise 9 tons 150 
 feet per day. 
 
 For a slightly greater quantity two horses are used. But 
 when a larger quantity is to be handled, a more efficient and eco- 
 nomic power is employed. The utilization of horse-power is 
 but an intermediate step in the history of many mines. Where 
 water is scarce, fuel dear, and the transportation of machinery 
 difficult, the horse- whim serves temporary ends as a simple and 
 tolerably satisfactory hoister. 
 
 The invariab'e arrangement is a wheel-and-axle machine, 
 consisting of a drum and driving-beam to which the horse is 
 harnessed. Two sticks 6"X6" and 9 feet long are mortised to- 
 gether at right angles to each other, with four 4-inch planks 
 trimmed to the quadrant of a circle. These are held a foot or 
 two apart by studs, and to them 3-inch plank staves are spiked 
 to form the barrel; which, though upheld on the axle, turns 
 freely about it. The axle is a round 2-inch rod stepped in z. 
 stone or iron block and held at the top in an iron socket on the 
 span-beam. The latter is 10 inches square, 36 feet long, sup- 
 ported on legs mortised to it and braced A square iron axle- 
 rod fastened to the drum and turning wi:h it is often seen, but 
 is not so good as the free axle. The entire frame can be built 
 for $ico. The drum may be above or below the driving-beam 
 (Fig. SO- 
 
 A derrick frame is necessary over the shaft at a height suffi- 
 cient for convenience of handling the hoisted tub The hoist- 
 rope passes over a sheave at the top and to the drum if set up 
 high over the driving-pole, or under another pulley at its base if 
 the drum is close to the ground. The latter arrangement is
 
 130 
 
 MANUAL OF MINING, 
 
 FIG. 51. A Whim.
 
 HOISTING MACHINERY. 131 
 
 cheaper to build, but is wasteful o" power, For lowering the 
 tub (Fig. 51), a lever is in reach of the driver, by which the driv- 
 ing-beam may be disengaged from the drum and the tub lowered 
 by its own weight, uncoiling the rope from the drum. A band- 
 brake 3" X \" regulates the speed. The brake must be set so as 
 to work with the motion, not opposite to it; and the brake force 
 exerted to produce larger tension in the driving, not in the slack, 
 portion of the band. The length of the driving-beam and the 
 diameter of the drum may be altered at will but the ratio be- 
 tween them is also the ratio of the speed of the horse to that of 
 the hoist 
 
 If circumstances permit, two ropes may be operated from the 
 same drum, one ascending and the other descending Conical 
 drums may also be employed. Iron-framed whims are on sale, 
 in which a drum is horizontal and turned by a bevel-gear on 
 its axle, fitting to another at the central end of the drive-beam. 
 While convenient and easy to erect, the introduction of the bevel- 
 gear involves additional friction. 
 
 The plane of the derrick-pulley should be tangent to the 
 drum, and the latter far enough away that the rope may coil 
 and unwind freely without charing on its adjoining coil. This 
 is accomplished by arranging the point of departure of the rope 
 from the pulley at the same height as the central coil. Where 
 the full and empty tubs are simultaneously operated, this can 
 be attained only approximately. If greater nicety is desired, 
 the pulley slides on an inclined p^ne as the coiling or uncoiling 
 proceeds, or else the drum shifts its position by turning on a 
 screw-thread. No lateral motion is allowed to the sheave, since 
 the rope must occupy a central position in its own hoisting com- 
 partment. 
 
 Tte Hoisting-engine. A hoister is a simple engine attached 
 to a drum on which a rope is wound The combinations are 
 numerous, from that of the upright boiler and engine on the 
 same base, to the vertical condensing, compound engine con- 
 r.ected with drums on separate foundations. Generally the 
 cylinders are horizontal. The choice between horizontal and
 
 MANUAL OF MINING.
 
 HOISTING MACHINERY. I3J 
 
 upright engines is chiefly one of space The horizontal engine 
 is the cheaper, the s.mpler, the easier to inspect, and the easier 
 to repair. Outside of the advantage of requiring less space, 
 the upright engine has the advantage of less wear on the cylin- 
 der, and a more direct strain upon the foundations. The types 
 are few, from the slow haulage-engines for tail-rope to the 
 high-speed hoists with large drums Most hoisters are oper- 
 ated intermittently; only in a few instances are they continuous. 
 The engine must be simple, safe, under complete control, capable 
 of quickly attaining full speed and being accurately stopped. 
 
 The selection of the type will largely depend upon the speed 
 desired for hoisting and the choice of valves, condensers, ratio of 
 expansion, compound or simple engine, and the comparative cost 
 and maintenance. Frequently the question of economy receives 
 the last consideration, simplicity of construction and safety of 
 operation being of prime importance. 
 
 First- and Second-motion Hoisters. The communication of 
 motion from piston to the drum is direct or secondary and may 
 or may not be provided with a reversing motion. In the former 
 class the piston is coupled directly to the drum through the con- 
 necting-rod. The load is hoisted, held, or lowered by steam at 
 the will of the engineer. High speed is possible with direct- 
 acting engine (Figs. 52, 57 and 58.) 
 
 Second-motion Engines.- All engines transmitting the power- 
 io the drum through one or two pinion wheels or a friction-clutch 
 are second-motion engines These engines are slower but under 
 greater control than the direct- acting engines, besides offering 
 less risk to overwinding. The pinion-wheels may be toothed 
 wheels gearing with others on the drum- shaft, or friction- wheels 
 bearing against those on the drum-shaft (Fig. 53). 
 
 The ratio of this gear varies between 4 to i and 7 to i, the 
 latter ratio being for very slow speed. Some quarry engines 
 are provided with a double set of gear-wheels for two speeds 
 
 (Fig- 54). 
 
 The geared engine is intended for slow hoisting, shallow 
 depths, and small outputs. The direct-acting is employed for
 
 134 
 
 MANUAL OF MINING.
 
 . HOISTING MACHINERY. 
 
 J 35 
 
 m
 
 136 MANUAL OF MINING. 
 
 large outputs and with a hoisting speed of never less than 500 
 feet per minute. The limit to the size of the gear-wheels is 
 determined by the peripheral speed of the teeth, which should 
 not exceed 1000 feet per minute. 
 
 Second-motion engines may be operated with a short cut-off, 
 which is not always possible with the direct-acting tyye. More. 
 over, their repairs are greater and the operation easier. A properly 
 designed second-motion engine may be employed economically 
 on steep slopes, and occasionally for short shafts where the 
 output is large. A pair of 16" X3o" cylinders on a three-foot 
 drum had an output of 1413 tons from a shaft 370 feet deep 
 in six hours' time. 
 
 Internal Gear Connection. Frequently drums are fitted with 
 pinion gearing into an internal rack on the drum. This can be 
 slipped out at will, leaving the free drum to pay out the rope until 
 the proper depth is reached, when the pinion is returned to place, 
 and the drum is ready for hoisting from that depth. This arrange- 
 ment is in ended for the variable lengths of the hoist. 
 
 Friction-engines have a pinion-wheel on the driving-shaft 
 of papier roache" bearing against an iron surface on the drum- 
 shaft. The surfaces are kept dry and free from water, steam, or 
 other substances which reduce the coefficient of friction and 
 thus require an excessive pressure to move the requisite load. 
 Friction-engines deliver the power to the drum more quietly 
 than toothed-gear or direct-connected engines, and therefore pro- 
 duce less shock upon the rope. Sometimes the engine is in 
 continual rotation with its pinion-shaft, and hoisting is effected by 
 bringing the drum with its shaft and gearing surface against 
 the pinion. For lowering, the drum is released from contact 
 with its pinion and revolves freely, being controlled only by the 
 brake manipulated by the engineer. The risk of slipping and 
 the inability of the engineer to apply the requisite pressure makes 
 the friction-engine unsafe. 
 
 V Friction-wheels. The surfaces of the drum-band and the 
 driving pinion are grooved circumferentially to give a tight grip 
 between them without excessive pressure on the bearings. These
 
 HOISTING MACHINERY.
 
 1 3 8 MANUAL OF MINING. 
 
 wheels are of cast iron. The irregular wear of the wedges, how- 
 ever, makes them unreliable in time. 
 
 Friction-clutches. These engage the drum externally or 
 internally. They are quick, noiseless, and give little shock. 
 A number of blocks attached to a band furnish the frictional 
 surface, and are directly attached to the revolving shaft and in 
 continual rotation about the drum. They engage the outside 
 of the drum when forced by a two-armed driver operated by 
 bell-crank levers and revolve the drum (Fig. 56). A grooved 
 sleeve or cone moving along the shaft may also cause a shorten- 
 ing of the arms which bring the band down on the drum and 
 impart their motion to it. For lowering, the band is released and 
 the drum revolves while held in check by the brake. 
 
 One form of internal friction consists of driving-bands with four 
 arms keyed to the drum-shaft and in continuous motion. The 
 drum is free on this shaft and will hoist as soon as the band is 
 expanded to contact with its rim, by sliding the sleeve upon 
 its shaft and engaging its arms against its inside rim (Fig. 55). 
 
 The Brake. This is either an iron band with blocks of hard 
 wood on end, or a V-grooved wheel bearing on a corresponding 
 surface on the drum. The first is less troublesome and safer. It 
 is suitably applied by a simple lever, by a hand- wheel and 
 worm-screw, or an auxiliary steam-piston, according to the size 
 of the hoist. Some large engines require a small regulating- 
 engine to stop and start them. 
 
 The Unbalanced Load on the Engine. The resistance which an 
 engine must overcome will include the unbalanced weights, the 
 friction and the bending stress of the rope. These having been 
 determined, the size and the weight of the rope become known. 
 
 The maximum speed of hoist is limited by the equipment 
 of the shaft, and the minimum by the hourly -output desired. 
 Owing to the diminishing weight of the rope as the hoisting pro- 
 gresses and the unequal work thrown upon the engine, a constant 
 speed of hoisting cannot be maintained. The engine, from the 
 beginning, speeds up with the reducing load, while the accelera- 
 tion increases almost to the end of the hoist.
 
 HOISTING MACHINERY.
 
 140 
 
 MANUAL OF MINING.
 
 HOISTING MACHINERY. 
 
 141
 
 142 - MANUAL OF MINING. 
 
 Balancing the Dead Load. The work of the engine divided 
 between the live load and the dead load is unequal because 
 of the diminishing influence of the rope during the hoist. This 
 inequality can be partially reduced by balancing the weight of 
 the rope. The cages and the cars carrying the load may be 
 balanced by attaching to another rope on the same drum and 
 engine a similar cage and car. The weights of the two, going 
 in opposite directions balance one another, though the weight of 
 the rope, its friction, and its bending stress are still unbalanced; 
 so also, the weight of the mineral being hoisted. 
 
 The balancing of cage and car may be accomplished by the 
 use of two cylindrical drums (Fig. 57), a conical or a fusee drum 
 or a counterbalancing system. 
 
 Balancing by Conical Drum. The method of balancing by 
 the use of a cone or a fusee depends on the change of leverage 
 of the two loads. The load is hoisted from the small diam- 
 eter' of the drum, while the descending rope is suspended from 
 the larger diameter of the cone. As the hoisting proceeds, the 
 ascending rope winds upon an increasing diameter, while the 
 descending rope and cage arc gradually operating from a shorter 
 radius of the drum. Equalization is somewhat maintained and 
 an easy start is thus made. 
 
 A p'ain conical drum, to be effective as a counterbalance, 
 requires an angle so great as to be dangerous on account of the 
 liability of he rope to slip. Any advantage to be derived from 
 having a safe angle being trifling, it is really not worth the risk. 
 
 The cylindrical drum admits of extension of the hoisting 
 depth but the conical or fusee drum is built for a specific condi- 
 tion of load, speed and depth. 
 
 The Conical Drum. This is used when the rope is heavy 
 and the economy of accurate counterbalance is clearly indicated 
 as practicable and will warrant the increased cost. It is much 
 larger than ihe cylindrical drum, even if its mean diameter is 
 the same. The depth of shaft or the point of hoist must be 
 fixed to maintain equalization with a given cone or fusee. Hoist- 
 ing from several landings would not be economical without the
 
 HOISTING MACHINERY. 
 
 J 43 
 
 use of additional counterpoises. The conical drum is employed 
 single or double (Figs. 53, 58, and 59). 
 
 In designing the single or double conical drum the hori- 
 zontal distance, P, between the centres of the two consecutive 
 grooves or coils, and the vertical distance, p, between them must 
 be in such ratio that the inclination shall not exceed 30; p+P 
 must be less than 0.577. The minimum diameter, 2r, of the 
 smaller end is fixed by the rope, and the larger diameter, 2d, is 
 dependent upon it. 
 
 Let c = the circumference of rope circle = 3. 1416(27+ #); 
 if = 0.2830^; 
 L = length of the drum between the first and last grooves. 
 
 The lenglh can be determined when the number of grooves 
 has been calculated from the conditions given. 
 
 FIG. 59. A Double Conical Hoist. 
 
 The Fusee. This drum is designed to give an equalization 
 of the dead load, including the rope, at all points throughout the 
 journey of the cage. It may be determined by the use of the 
 following formula:
 
 I44 MANUAL OF MINING. 
 
 After one revolution (RD+A and the weight of one turn of rope 
 at the small end) p must balance (4 + the weight of one turn of 
 rope at the large end) p' and still equal Ad = M. For each 
 revolution throughout the journey this balance must be main- 
 tained. This is satisfied by the following formula, in which 
 Z is the total arc of revolution described by a point on the rope 
 at the small end from the beginning to the end of the hoist: 
 
 R is the weight of one foot of rope. 
 
 The curve of the drum is then constructed by substituting 
 different values between r and d, which are the limits of p, placing 
 the second member equal to 27tn, and solving for n, the number 
 of grooves. The various assumed va'ues for p are radii of the 
 curve at the various points along the axis which are at a distance 
 from the initial point equal to the horizontal pitch, P, multiplied 
 by n. The curve so plotted is the section of the drum which 
 will fulfil the conditions, furnishing an equalization throughout 
 the journey. 
 
 Let D = 2ooo feet, R = $ Ibs., A =4000 Ibs., r=$> inches; 
 then M = 40,000 ft.-lbs., and d becomes 10 feet. Solving the 
 equation and equaling with 2rw, ^ = 55.8 revolutions between the 
 initial and final points of winding. Where d = g. n becomes 
 53.3 revolutions from the start; where d = 8, ^ = 49.5: where J=7, 
 n=44-4; and where d=$, here have been 35.1 coils of rope. 
 
 The Reel. This is a barrel with spider arms on either side, 
 between which a flat steel rope is wound in consecutive layers. 
 It may be so devised that the moment of resistance will be nearly 
 constant The thickness of the coiling rope may be so selected 
 that it will increase the leverage of the load at the same rate as 
 the weight of the load decreases, in which event the work of 
 hoisting is uniform. If however, uniform hoist'ne; cannot be 
 obtained throughout the trip with a reasonable thickness of rope,
 
 HOISTING MACHINERY. 
 
 '45 
 
 FIG. 60. Plan of a Geared Reel Hoister, showing the Brake Connections.
 
 146 MANUAL OF MINING. 
 
 the design must be carefully considered and the engine tested 
 after being placed (Fig 60). 
 
 The smallest allowable diameter of barrel is fixed by the rope, 
 and the number of layers, n, of the rope is equal to (dr) + t. 
 The diameter of the coil, 2d, when the cage is at the top of he 
 hoist may be ascertained by the use of the formula P/ = 3.i4i6 
 (d 1 r 2 ). Given / (f" to "), the value for d may be obtained; 
 or, having decided upon d, t can be ascertained. 
 
 Counterbalance. The balancing of the rope weight and its 
 resistance is accomplished either by the varying weight of an 
 attached counterpoise to the cages or drum, or by varying the 
 leverage on the hoisting-shaft of a fixed weight. The Koepe 
 and Whiting systems and the chain balance are of the first type, 
 and that installed at the Camphausen mine is of the second type. 
 This makes the moment of the static load during the hoist con- 
 stant and equal to that of weight of mineral The best results 
 are obtained from a counterbalance when the hoist is always 
 from one and the same level 
 
 The Koepe System of Winding. This form of counterpoise 
 consists in extending a tail rope from the bottom of one cage, 
 under a sheave at the bottom of the shaft, up to the floor of the 
 other cage. The usual hoist-drum is replaced by a simple grooved 
 sheave Fig. 61. Not only is the weight of rope, cage, and car 
 in each compartment constant and the work of the engine nearly 
 uniform, but the comparative lightness of the sheaves and the 
 absence of heavy coils revolving with the drum give a correspond- 
 ingly less inertia, and hence require less engine power, than with 
 the ordinary systems. The net load which can be carried on the 
 rising cage cannot exceed the friction caused by the aggregate 
 weights suspended from the upper sheave, or else it will slip. 
 A positive grip must be taken on the rope by the driving mechan- 
 ism to prevent its slip or creep on the sheaves. Winding the 
 main rope on a pair of cylindrical drums, or wrapping it a few 
 times over a pair of multiple-grooved sheaves, as is the European 
 practice in power transmission, Chapter IX, would accomplish 
 this. Should such a slip take place, the indicators would fail 
 to give the correct location of the cage.
 
 HOISTING MACHINERY. 
 
 147 
 
 This system meets almost all the requirements of a perfect 
 equilibrium, is highly efficient, insures against overwinding, de- 
 creases wear, and disposes of an enormously heavy drum by using 
 a sheave instead. But the lower sheave blockades the bottom 
 of the shaft. The tail-rope may be a discarded hoisting-rope, 
 
 FIG. 6 1. The Koepe System of Winding. 
 
 for it has no work to do beyond supporting its own weight. The 
 method is of limited application as to depth. 
 
 In a shaft 1260 feet deep a saving of n per cent in power is 
 effected. With sheaves 8 feet in diameter and 8-inch axle, 
 main rope 175 Ibs. per foot and tail-rope 1.55 Ibs., cage and car 
 2100 Ibs., load 2750 Ibs., depth of shaft 2500 feet, and velocity
 
 1 4 8 
 
 MANUAL OF MINING 
 
 of hoist 2500 feet per minute, 250 horse-power are saved over 
 that required to operate the same loads without this balance. 
 The Whiting system (Fig. 62), a development of the Koepe 
 
 FIG. 62. The Whiting System of Balance. 
 
 system, uses two sheaves and dispenses with the heavy coils of 
 rope revolving with the drum. 
 
 Chain Counterbalance. Another form of counterpoise is a 
 heavy chain wound on a secondary drum on the main shaft by 
 a cable-rope (B, Fig. 63). The chain hangs in an auxiliary 
 shaft, or a ladderway in the hoist-shaft, and is equal in weight 
 to the full depth of hoist-rope A. As the loaded cage begins its 
 FIG. 63. FIG. 64. 
 
 Counterbalancing the Rope by Chain. 
 
 hoist the entire weight and length of chain is hanging from its 
 drum to assist the engines in balancing the rising load. With 
 the continuance of hoist the cable B pays out chain, which coils
 
 HOISTING MACHINERY. 149 
 
 on the bottom of the auxiliary shaft. As the two cages approach 
 one another the hoist-ropes nearly balance, the chain is coiling 
 at the bottom, and the cable-rope is being paid out. When the 
 cages are midway, the entire chain is coiled there and all the 
 cable is also paid out. The counterbalancing cable-rope is n^w 
 wound under the drum and lifts the chain, whose weight now 
 balances the descending rope to the end of its journey. At all 
 points in the hoist the amount of suspended chain balances the 
 differences between the weights of the two pendent hoist-ropes A. 
 
 In a given case the chain counterbalances a load of 3500 Ibs.,. 
 :age 1300 Ibs., car 1540 Ibs., in a shaft 3400 feet deep; the radius, 
 of the hoist-drum is 6 feet, and that of the chain-drum 27 inches. 
 At the beginning of the hoist 8 tons of chain are suspended from 
 the auxiliary drum. After 400 feet of hoist, 12,134 Ibs. are sus- 
 pended; at 400 feet further, 8400 Ibs.; at 2200 feet below the 
 surface, 4666 Ibs.; at 1800 feet down, 933 Ibs. are yet uncoiled; 
 and at the end of the hoist the full weight of the chain is again 
 suspended. The bore-hole down which the chain hangs is 638 
 feet deep; 51 H.P. are required to overcome the friction due to this 
 increased weight, but 345 H.P. are saved by balancing the rope. 
 In another shaft, 2200 feet deep, a counterbalance-rope 700 feet 
 long and a chain 580 feet, weighing 4320 Ibs., are employ edj 
 the size of the chain is graduated to meet the varying weights, 
 and for 104 feet is of f-inch links, for 162 feet of f -inch, and for 
 314 feet of f-inch. 
 
 In Despre's method an endless cable has the balancing chain, 
 and is given one or two turns on the drum between the full and 
 the empty rope (Fig. 64). It passes half-way down an auxiliary 
 shaft. One-quarter the way down the chain is swung on a 
 pivot, and fixed to the cable so that the other end of the chain is 
 at the top of the auxiliary shaft at the beginning and end of the 
 hoist. 
 
 The Camphausen System. A counterpoise (W, Fig. 65) is 
 suspended from a supplementary spiral drum mounted on the 
 engine-shaft. The cable-ends are attached to the small and large 
 ends of the spiral. It winds itself on the descending side simul-
 
 I 5 o MANUAL OF MINING. 
 
 taneously with the winding on the ascending side, and thus main- 
 tains a balance by the differential lever-arm of its weight. Until 
 the cages meet in the hoistway, the lower branch of the cable 
 travels faster than the upper. During the balance of the hoist the 
 upper winds faster than the lower. For a depth of 1900 feet 
 the auxiliary shaft was 250 feet. The installation is inexpen- 
 sive and economical. 
 
 FIG. 65. The Camphausen System of Counterbalancing. 
 
 A very extended discussion of the operations of hoisting is 
 to be found in a monograph of the Institute of Mining and 
 Metallurgy, XI, Part I. 
 
 Designing Roisters. This step consists in calculating the 
 dimensions of the two cylinders, which are capable (i) of starting 
 the load from the weakest position of their combined crank- 
 pins, and (2) of maintaining an average speed of hoist with it 
 commensurate with the desired output. The mine output is 
 controlled by the capacity of the shaft or the haulage-way, which 
 in turn depends on the size of the car and the number of cars 
 delivered per hour.
 
 HOISTING MACHINERY. 151 
 
 Data Necessary in Designing. From a shaft or a steep 
 slope the cars arrive at the surface singly; from a gentle slope 
 the cars come in trains. The dimensions of the carriers (car, 
 bucket, or skip) and their capacity are known or assumed. The 
 load, q, which a single car can carry is also known, hence the 
 load which falls upon the hoisting- or haulage- engine can be 
 determined. This may be divided into a dead load and a live 
 load, the former being the cage, car, skip, or bucket, while the 
 live load is the mineral and the friction resistances. 
 
 The speed depends upon the character of timbering. A 
 mine employing buckets has usually a poorly timbered shaft, in 
 which the speed cannot exceed 300 feet a minute; with skips or 
 slope carriages the timbering must be of a better character, and 
 the maximum rate of 1000 feet is allowed; if cages are em- 
 ployed, the shaft is well timbered and the speed may reach 3000 
 feet per minute. 
 
 The depth of the shaft is known; hence the time required per 
 trip is definitely fixed; and if to it be added the time lost at the 
 top and at the bottom while the buckets, cars, or carriages are 
 being exchanged and loaded, the time occupied in each round 
 trip in each hoistway will be known. The time lost in filling an 
 attached bucket at the bottom or dumping it into the top is from 
 3 to 5 minutes. If the buckets are detached and promptly replaced 
 by loaded buckets, the time is much less. Skips loaded from 
 chutes are almost as time-consuming as buckets. The skip 
 occupies 2 to 3 minutes to side-track. A car may be run on or 
 taken off the cage or a slope carriage in 25 seconds. These 
 delays seriously reduce the capacity of the shaft and the possible 
 output. Hence for large mines cages and cars are resorted to. 
 The influence of this delay on the aggregate output is diminished 
 by increasing the number of compartments in the shaft or by 
 increasing the velocity. The former is preferable. 
 
 Formulae for Determining the Number and Duration of 
 Trips. By the following equations may be determined the time 
 occupied per round trip in each hoistway and the number of 
 trips possible per hour in each hoister:
 
 I 5 2 MANUAL OF MINING. 
 
 Let 2 =the number of hoisting- compartments; 
 D =the depth of the shaft in feet; 
 v = velocity of the hoist, feet per minute; 
 i/ = velocity of lowering per minute; 
 Q =output, tons, per hour; 
 0=load, tons, per trip; 
 /=the minutes to load and unload at the top or the 
 
 bottom ; 
 
 T =4z=minutes per round trip in any hoistway; 
 A =the interval in minutes between the arrival of each 
 
 cage, skip, or bucket at the shaft top; 
 w=number of trips per hour in each hoistway; 
 
 n=Q + zq, A~ and T^~+^+zt. 
 
 Thus in a poorly timbered shaft, if only one bucket be run up and 
 down without detaching, the output from a 3oo-foot shaft is about 
 3.6 tons per hour. With three buckets in constant use, each 
 holding 600 Ibs., the hourly product cannot exceed 6 tons. With 
 excellent timbering double the speed may be permitted, in which 
 event three buckets will deliver 7.2 tons per hour at the surface. 
 So becomes evident the influence which the loss of time at land- 
 ings has on the output when even doubling the speed increases 
 the former only one-fifth. An additional hoistway were better. 
 
 EXAMPLE. For an output of 360 tons in ten hours a rate of 360 feet is 
 necessary with but one hoistway; with two hoistways and cars holding 3000 
 Ibs. each the rate need be only 144 feet per minute, because of the time saved 
 at the changes. 2=1; /=o.4i; D=3oofeet; <7=i.5; =36; ^=24; ^ = 250 
 min.; 7=2.5 m i n -'> whence ^=^, = 366 feet. With Z=2, A becomes 2.5 
 and =i2. T=5 min.; whence 7>=7/,= i44 feet. The horse-power in 
 the two cases would be 100 and 32 respectively. This, at the coal -pile, amounts 
 to 40 tons coal per month. 
 
 Determining the Size of Rope. The rope supports the total 
 load, the friction, and the starting stresses. The latter is larger 
 than is usually suspected. A gentle start gives a stress of 2 per 
 cent only, whereas a start from 12 inches of slack rope would 
 make the stress 1.4 times as great as the actual load. The stress
 
 HOISTING MACHINERY. 1 53 
 
 at starting is usually assumed at 10 per cent of the gross load, 
 including friction. The total tension produced by these resist- 
 ances must not exceed one- third of the ultimate strength of the 
 rope. 
 
 The Bending Stress on the Rope. The rope must be sufficient 
 for the maximum stress falling on the rope after hoisting begins, 
 including the stress induced by bending the rope over its sheave 
 or drum. This sum is not to be over one-third of the ultimate 
 strength of the rope. Frequently, with a small drum, the bend- 
 ing stress may far exceed that due to the load. The value for 
 the bending stress is represented by the formulae in Chapter VII. 
 
 The Available Strength of a Rope. After deducting for 
 bending stresses, this should be at least one-eighth of its ultimate 
 strength. The size required may be ascertained from manu- 
 facturers' tables. 
 
 EXAMPLE. A 6-strand, ig-wire rope of cast steel i inch diameter will be 
 required if it is to support a load of 7270 Ibs. over a sheave 44 inches diameter. 
 k 14,606, load and friction =8000, and maximum load is 22,606. Ultimate 
 strength must be 67,800 Ibs. Had the drum been 84 inches diameter, k 
 would have been 5583 Ibs. and have required an ultimate strength 40,752 for a 
 J-inch rope. 
 
 Determining the Dimensions of Cylinders. The size of the 
 rope, the dimensions of the drum, and the values for the various 
 loads and resistances to be overcome having been determined, 
 the area and stroke of the piston is made sufficient to start the 
 load with the allowable initial pressure and hoist it at the desired 
 speed. The work of the engine must also be calculated for the 
 average resistances at the mean velocity per minute. The cylinders 
 should not be too large, because the throttling that must ensue 
 for the economic work is wasteful of steam, for the condensation 
 would be excessive. Neither can they be calculated for an over- 
 load, as then they might not meet the probable emergency. The 
 cylinders should be apportioned largely according to the conser- 
 vative judgment of the engineer. Care, however, must be taken 
 to assure a sufficiently large starting moment of the engine to pick 
 up the load. The mechanical efficiency, m, may be used as 0.7
 
 154 .MANUAL OF MIXING. 
 
 for a slide-valve engine and a cut-off of one-half. The Corliss 
 engine may be calculated to have an efficiency of 0.85 with a cut- 
 off of one-third, for a normal load. 
 
 The Hoisting Capacity of an Engine. The following equa- 
 tions will then be employed for determining the depth to which 
 a given engine can work within the normal limits of boiler pres- 
 sure which is available and the average r.p.m. for that class 
 of engine. They will also serve to calculate the required dimen- 
 sions of one or two cylinders for given conditions: 
 
 Z, = the length of the drum, inches; 
 r=the smaller diameter of the drum, inches; 
 </=the larger diameter of the drum, inches; 
 w = the number of grooves in the drum; 
 X = the diameter of the rope, inches; 
 R = the weight of the rope in pounds per foot. 
 
 v=the velocity of hoist, feet per minute, ' - ; 
 
 6 
 
 A =the weight of bucket, cage, car, skip, or carriage Ibs.; 
 M = unbalanced load on the drum at the start; 
 
 K = the moment of resistance at starting; 
 TF=the average load during the hoist; 
 
 F= the friction of the gross load, including the acceleration 
 of the moving parts; 
 
 =gear ratio, second-motion engine, = -; 
 
 y= number of teeth on the gear-wheel; 
 x = number of teeth on engine-shaft pinion. 
 
 The capacity of the engine may be known from its indicator 
 cards or from a knowledge of the cylinder dimensions and the 
 proposed boiler pressure, P, and engine speed. The velocity 
 for a given output, number of trips, and depth will have been 
 determined. Then 
 
 H =o. 0000039666$ k 2 mpN = Wv.
 
 HOISTING MACHINERY. 155 
 
 Or, the value for k and s may be ascertained for a given set of 
 conditions of load and speed as well as for P and N. The ratio of 
 s to k is assumed between 1.3 and 2, according to the class of 
 engine desired. 
 
 With a direct-acting engine g=i. In all cases the mean 
 
 sVg 
 piston speed is 0.318411 . 
 
 Hoisting on Inclines. If the hoist is not in a vertical shaft, 
 the load, M , must be multiplied by the sine of the angle of the 
 slope (the percentage of inclination), and the value for F is multi- 
 plied by the cosine of the slope angle, in calculations for the 
 cylinder dimensions. 
 
 The coefficient of friction, /, is 4 per cent with a bucket in a 
 vertical shaft, 6 per cent with a cage or skip, and 8 per cent with 
 a bucket in a slope. All of these are reduced to the pull at the 
 circumference of the drum. 
 
 The Starting Moment of an Engine. In designing an engine to 
 start the unbalanced load, M, it is necessary to determine by 
 the equilibrium of moments the value for the moment of the 
 load, friction, and starting stress, and equate this with the mini- 
 mum rotary effort upon the crank-pin. The starting moment 
 of the engine is ascertained by the substitution, in the following 
 formula, of the value for c, corresponding to the given cut-off 
 and ratio of connecting-rod to crank-arm. 
 
 The minimum rotary effort, R, of a duplex single-expansion 
 engine = CsPJPmg = R, inch-pounds. The maximum rotary effort 
 =R' = C 1 'sP i k 2 mg. s and k are in inches, P x is the net initial 
 pressure, PB, and C, the coefficient representing the minimum 
 rotary effort as ascertained from the following table for the 
 particular ratio of connecting-rod, l } to crank-arm, a, and a piston 
 clearance of 7 per cent. 
 
 The values given are the rotary efforts exerted upon the 
 two crank-pins for each pound of net initial pressure for each 
 inch of one piston diameter, the pair having their cranks at 
 quarters. The clearance is assumed to be 7 per cent, and a back 
 pressure of 18 Ibs. absolute is allowed. If the engine is a con-
 
 '56 
 
 MANUAL OF MINING. 
 
 densing engine, with 3 Ibs. absolute back pressure, the values for 
 C and C are somewhat larger than those in the table. A pair 
 of cylinders of 10 inches diameter and 20 inches stroke with 
 connecting-rods 50 inches long, receiving an initial steam pres- 
 sure of 100 Ibs. absolute, will have a minimum rotary effort 
 equal to 48,480 in.-lbs., if the cut-off be assumed at 'and the 
 back pressure be 3 Ibs. 
 
 TABLE or MAXIMUM AND MINIMUM ROTARY EFFORTS FOR CERTAIN RATIOS 
 OF CONNECTING-ROD TO CRANK-ARM. DUPLEX ENGINES. 
 
 Values for C l and C. 
 
 Apparent 
 Cut-off. 
 
 /=4.5<J. 
 
 l=5-5d. ' 
 
 ;= 7 a. 
 
 o.oo 
 
 o.s;6i5 
 
 0.3074 
 
 0.5600 
 
 0-3974 
 
 0.5580 
 
 0-3974 
 
 \ 
 
 0.4664 
 0.4380 
 
 0-3633 
 0.2472 
 
 0.4606 
 0.4365 
 
 0.3688 
 0.2499 
 
 0.4586 
 0-4350 
 
 0.3696 
 
 0.2550 
 
 \ 
 
 0.4256 
 
 0.1878 
 
 0.4193 
 
 0.1923 
 
 0.4124 
 
 0.1966 
 
 
 0.3893 
 
 0.1469 
 
 0.3818 
 
 - I 473 
 
 0.3786 
 
 0.1512 
 
 The hoisting-engine must be able to start the load from any 
 point in the shaft and at any point in its stroke, and hence should 
 have in its weakest position a rotary-effort moment greater than 
 that of the maximum load which may fall upon it. Doubtless 
 all the masses involved will have attained sufficient velocity 
 in one revolution to produce an average effort on the applied 
 moments from both cylinder-cranks. The engine-friction moment 
 decreases during the hoist nearly commensurate with the load 
 moment. 
 
 Hoisters cannot be regulated by fly-wheels, and their sole 
 means of maintaining a uniform rotary effort is that which is 
 afforded by the drum or a weight of the connecting-rod. If 
 these are made heavy, the velocity may be somewhat uniform. 
 But inasmuch as the engine must furnish an excess of power 
 at the beginning which is 50 per cent more than the average, 
 and the inertia of the rope, sheaves, drum, etc., must be over- 
 come, and, later, the kinetic energy absorbed by the brake, some
 
 HOISTING MACHINERY. 157 
 
 form of energy accumulator should be devised for the purpose 
 of economizing power. 
 
 Determining the Size of the Drum. The minimum allowable 
 diameter of the cylindrical drum is fixed by the size of the rope 
 (Chapter VII), it must be more than forty times the diameter of 
 the rope one hundred times the diameter would be preferable, 
 since thereby it would reduce the bending stress of the rope and 
 increase the net working load. The smaller end of the cone 
 or fusee, and of the barrel of the reel, is also fixed by the same 
 ratio. The desired hoisting speed also influences the final de- 
 cision as to the diameter. The drum or drums are geared to the 
 same shaft and placed centrally between the cylinders. 
 
 The length of the drum is controlled by the depth of the hoist 
 and its location with reference to the shaft. On cylindrical 
 drums the rope coils are contiguous to one another. Conical 
 drums are provided with spiral grooves. Reels have the suc- 
 cessive coils of rope winding on one another. The length of 
 the drum and its distance from the shaft must be such that the 
 acute angle from the sheave to the ends of the drum must not 
 exceed 6. When this fleet angle exceeds the limit, the rope 
 mounts its adjoining coil, and provision must be made for guid- 
 ing the rope upon sheave and drum, or the drum must be set 
 back from the shaft. In the first case the engine may be set 
 on a carriage to move with a varying position of the rope on 
 the drum. 
 
 The cases are rare where but a single rope with a single cylin- 
 drical drum is sufficient. Two ropeways are the rule with two 
 independent cylindrical drums; or a double cylindrical drum, two 
 conical drums, or a double cone, the fusee, or two reels are 
 employed according to the requirements in the given case. These 
 may be operated independently of one another, though provision 
 is always made for locking the pair of drums together to obtain 
 simultaneous hoist and lowering. This is true whether the engine 
 be a direct-acting or a second-motion engine. 
 
 The General Formulae for Hoisting Calculations. The follow- 
 ing table then gives the value for the starting resistance of the
 
 158 MANUAL OF MINING. 
 
 load for any variety of drum, the moment of which is equated 
 with the starting effort of the engine, to determine the dimensions 
 of the cylinders. The value for the unbalanced load to be hoisted 
 throughout the trip at a requisite velocity is also given for the 
 selected drum. Then the resistance M+F, multiplied by the 
 velocity v, is equated with the value for engine power, to deter 
 mine the engine dimensions for the dynamic load. 
 
 THE STARTING MOMENTS, WORKING LOADS, ETC., 
 ENGINES AND DRUMS. 
 
 WITH VARIOUS HOISTING 
 
 
 Maximum 
 Unbalanced 
 Load. M. 
 
 Friction of 
 Load, F. 
 
 Minimum Admissible 
 Radius of Drum d, 
 Inches. 
 
 One-cylinder drum.. . 
 
 A +KD + ^ 
 
 if/ 
 
 
 if)ogDg 
 
 ***" N 
 
 Two- " " ... 
 
 RD + 2oooq 
 
 (M + 2A)/ 
 
 " 
 
 Double conical drum. 
 
 a 
 
 " 
 
 zA + 2Oooq 
 
 Fusee 
 
 ,000, 
 
 (M+2A + RD)/ 
 
 RD + A 
 
 
 Reel 
 
 
 
 " 
 
 r+ n t;V 3 .8 2 Dl + r* 
 
 
 
 " 
 
 (M + 2A + 2RDV 
 
 25* 
 
 
 
 Number of 
 Grooves, n. 
 
 Length of 
 Drum, L, 
 Inches. 
 
 Starting Moments, 
 K. Inch-lbs. 
 
 One<ylinder drum. . . 
 
 1. 0098!) m 
 d 
 
 nX 
 
 (M+F)* 
 
 Two- " " . . . 
 
 3.81960 m 
 
 2 nX 
 
 V+p 
 
 Double conical drum . 
 
 3.8it)6> 
 
 ~.) 
 
 -A(l-t)i 
 
 Fusee . 
 
 \\TF) ~ f4 >(jf) ~ d * 
 
 
 
 v+'+StM 
 
 
 12.566^ 12.566^ 
 
 Reel 
 
 d~r 3.82!) 
 
 
 
 - 
 
 
 
 Koepe system . . 
 
 
 
 
 IX+IV 
 
 

 
 HOISTING MACHINERY. 159 
 
 The following examples will suffice to illustrate the use of 
 the formulas and tables for calculating the work of hoisting. 
 
 i. Duplex engine, single drum. 
 
 In a shaft 400 feet deep an output of 100 tons per ten hours 
 is desired from a single compartment. The load in a car is 
 | of a ton; weight of car and cage, 1800 Ibs. Assuming a duplex 
 engine, single drum, and full steam pressure available for starting, 
 required the size of the cylinders, number of trips, etc. 
 
 A iQ-wire rope at starting (Chapter VII), on an 8-foot sheave, 
 suffers a stress K=A + 2oooq+F+RD. Assume the rope to 
 weigh i Ib. per foot. The total load (1800+1500+ 400) (i.i) = 
 4070 Ibs. This requires a f-inch cast-steel rope weighing 0.9 Ib. 
 per foot, because the total load must be less than \ the ultimate 
 strength, which must be 28,490 -Ibs. 
 
 1200. 
 
 , 
 r+o.^x 
 
 3200 Ibs. is less than one-third the ultimate strength. 
 
 Output per hour, 10 tons; <?=f; hence, with single cage, = i4 
 trips; time, 4.3 min. for round trip. Allowing for loss at top 
 and at bottom of 30 seconds each, the hoisting time is 1.65 each 
 way and ^=360 feet per minute. A drum of 4 feet diameter 
 makes 28.6 r.p.m. Let /=o.i, ^=0.75, the engine stroke =1.5^, 
 and the mean cut-off =. 
 
 (1800+400+ 1 500) (1 + 0.1)360 =44.4 H.P.; 
 0.0000039668 & 2 Nmp X2 =44.4; 
 
 whence k a p =174,012 and =13.5 inches. 
 
 With 7 per cent clearance ^=0.8925(100) 18=71.25 Ibs. per 
 square inch. Cylinder is i7"X26", nominally, with counter- 
 bore at each end of if inches. 
 
 For starting, the dimensions should be (page 10) 
 
 CsP l k 2 mg=R=g'j,'jio inch-pounds; 
 
 C =0.3688; g = i; Pi=82; whence k*s =4308 and = 14. 25 inches. 
 2. Double drum and duplex engine.
 
 160 MANUAL OF MINING. 
 
 A shaft is 2000 feet deep; car and cage, 5000 Ibs.; load, 2\ 
 tons; velocity of hoist, 2000 feet; gear ratio, i; drum, 60 inches 
 radius; engine-stroke, 48 inches; modulus, 0.8; admission 
 steam pressure, 100 Ibs. gauge; back pressure, 5 Ibs.; cut-off 
 average, ; sheave, 12 feet; and /, o.i. Required the size of 
 ip-wire rope, diameter of engine cylinder, and output of mine. 
 
 At the same velocity of lowering as of hoist, the round trip is 
 made in three minutes, including stops. In each hoistway 20 
 trips are made hourly. The output is 1000 tons daily. 
 
 The starting stress is (5000+ 5000 +.&D)(i.i) + .K'; assuming 
 a weight of 7- wire rope of 3 Ibs. per foot, and a diameter of if 
 inches, the load is 17,600 Ibs., and the bending stress 13,060 Ibs, 
 The aggregate starting stress must be less than one-third the 
 ultimate strength. That for a if -inch cast-steel rope is 62 tons. 
 
 Dimensions of cylinders for normal running: 
 
 .F = (6ooo+ 15,000)0.1 =2100 Ibs.; 
 
 (6000+5000+2100)2000 = 26,000,000 ft.-lbs. =793.6 H.P.; 
 2 Xo.ooooo3g66sk 2 Nmp = 793.6. 
 
 The depth of shaft equals in this case the velocity of hoist, 
 hence the number of coils = the r.p.m. 
 
 N =63.6 r.p.m.; piston speed =508.8; whence k 2 p= 40,966. 
 For an average cut-off \ with 3 per cent clearance, 
 
 p =0.8658(100+ 14-7)- 5 =99-3; 
 
 k 2 =397 and diameter of cylinder =23 inches. 
 
 At 63.6 r.p.m. for ^=2000 feet, the diameter of drum is 9.91 
 feet. The starting resistance is 17,600 X59-5 =0.3688 X48 Xio9-7 
 Xo.8XiX 2 ; whence* =26 inches. 
 
 With these dimensions of cylinders the M.E.P. at the average 
 rate may be 56 Ibs., corresponding to a cut-off of 0.18 for 3 per 
 cent clearance: 
 
 56=0(114.7)- 5, whence C =0.532. 
 
 Or, at a cut-off of , an initial pressure of 71 Ibs. absolute will 
 suffice. 56=0.8658(^-5; P = 7 i.
 
 HOISTING MACHINERY. 161 
 
 The normal rate of running being at \ cut-off,, the Corliss 
 engine of 26" X4&", giving 733.3 H.P., will consume 17,600 Ibs. 
 steam per hour. 
 
 3. Double conical drum and duplex engine. 
 Same specification as in 2. 
 
 The smallest diameter of the drum is assumed at 80 inches,, 
 for the rope of if inches. The angle of the cone is 14 30';. 
 />=o.47; P=i.825; c= 230.6 inches; ^'=4.178. 
 
 To maintain equal moments of resistance when the load is at 
 top and at bottom, 
 
 , 5000 + 2 X6ooo + 2 X 5000 
 
 a = 40 =72 inches. 
 
 2X5000+5000 
 
 n = 9 ^=68.2 grooves; Z= 1.825X68.2=10 ft. 4.5 in. 
 
 This is rather a long drum. To obtain a fleet angle of 6 it 
 must be 100 feet from the shaft. 
 
 The horse-power required for the normal rate of running is, 
 as before, 763.9; ^=68.2 r.p.m.; piston speed, 545.6 feet; and 
 />=99-3 Ibs. per square inch; k becomes 19.5 inches. 
 
 The starting resistance, neglecting the lengths of rope above 
 the surface landing to the head sheave, requires a diameter, k> 
 of 15.6 inches 
 
 (5000+6000+ 5000X1.1)40 5000(0.9)72 
 
 = .3688 X48 X 109.7 Xo.8 Xk 2 ~ 
 
 4. Reels with duplex engine. 
 Same specifications as in 2. 
 
 Barrel diameter is 40 inches and flat rope is jXsJ, weighing 
 4.2 Ibs. The bending stress is 
 
 89465* 
 
 R+ 2.231 
 the diameter of the outer length of rope is 
 
 d=V^S~2Dl+r 2 =6s inches; 
 and 
 
 n = 2(65 20) =90 coils.
 
 162 MANUAL OF MINING. 
 
 If the cylinders are direct-connected, the piston speed will 
 be 720 feet per minute, whence =17.08 inches. The above 
 speed is rather high. If the barrel be enlarged to 60 inches, 
 then d becomes 97.5 inches, AT =60 r.p.m., and & = 2of inches. 
 
 The starting moment of the engine at an assumed cut-off of 
 would be 
 
 R =0.3688 X48 X82 X57I.8 Xo.8 X2 =1,328,000 inch-pounds. 
 The moment of the starting resistance is 
 
 (16,000) (1.1)20- (5000X0.9)65 =94,700 inch-pounds. 
 
 5. Koepe system with duplex engine. 
 
 Same specifications as in 2. 
 
 The load upon the engine is at all times 5000+ the friction of 
 the gross weight, moving at 2000 feet per minute. Neglecting 
 the weight of the 1 2-foot sheaves, and assuming the rather large 
 frictional coefficient of o.i, 
 
 .F= (5000+ 6000+ io,ooo)(o.i) =2100 Ibs.; 
 
 (5000+2100)2000 
 
 33,000 
 
 430.3 =2 
 
 JV= 53.05 r.p.m., if g = i; and the piston speed =424.4 feet, 
 whence k becomes 17.05 inches. 
 
 If the drum is geared with a ratio of 2, k would become 14 
 inches. Then (5000+2700)72 =o.^6S8pk 2 smg, the minimum 
 rotary effort of the engine with a cut-off of \ being 0.3688. 
 
 Electric Holsters may be located in relation to incline or 
 shaft exactly where they can be operated to the greatest advan- 
 tage, considering convenience of position, without the pipes insep- 
 arable from air- or steam-hoisters. Their speed is constant, 
 the motor is economical, and the only power consumed is 
 that actually required to handle the load. One objection to 
 their use is the large gear ratio which is required to bring the
 
 HOISTING MACHINERY. 163 
 
 speed of motor down to speed of bolster (Fig. 66). But they 
 have none of the ills of the engine, with its tremendous vibration, 
 rattle, heat, and noise of exhaust, or of the compressed-air hoist, 
 which has, in addition, the extremely cold exhaust. 
 
 The motors for electric hoist are either of the continuous- 
 current, series- or compound-wound types, or of the alternating- 
 current induction type, designed for either constant or variable 
 speed, and geared to the drum. They are fitted with appropriate 
 rheostats. The direct- current series motor is particularly adapted 
 to cases which demand a very high starting torque and where 
 closeness of speed regulation is not imperative. The compound- 
 wound motor should be employed when there is liability of the 
 motor racing on throwing off the load. It has a high starting torque, 
 and its speed will not rise above a certain definite value when the 
 mechanical load is thrown off. 
 
 The variable- speed induction motor is best used for hoisting. 
 Its starting torque is quite high and its speed regulation sufficiently 
 flexible to make it very desirable.
 
 1 64 MANUAL OF MINING. 
 
 Calculating the Size of Wire for a Hoister. One of the three 
 variable quantities are to be determined: the horse-power of the 
 motor, the load, M, and the speed of hoist, v. 
 
 The mechanical efficiency, m, for the direct-current motors 
 is 0.85 and may be regarded as unity for the induction motors. 
 
 The effective horse-power of the motor must equal 
 
 0.00003032; (M + F) 
 and 
 
 M+ F = 33000 X horse-power X . 
 
 In determining the size of wire for the direct-current circuit 
 the formulae given in the next chapter will be used. 
 
 EXAMPLE. It is desired to operate a 30-horse-power moto at a distance of 
 500 feet from the generator by a direct current of 440 volts. Then D= 500; C, 
 for i -horse-power (Chapter VI), is 2 amperes for 440-volt circuit. Let V, the 
 loss in volts, be 10 per cent, or 44 volts. Then, for a 30-horse-power motor, 
 30 X 2 amperes ( = 60 amperes) will be required. Substituting in the formula, the 
 wire must have 14,300 circular mils and, according to the table, will be of 
 No. 8 size. The weight of the wire will be 69 Ibs. in the total, and its cost 
 $17.25. 
 
 If the current be a three-phase alternating, the value for K being assumed 
 as 17, a No 10 wire with 11,600 circular mils will be sufficient. Its total 
 length of 1500 feet, weighing 73 Ibs , will cost $18.25. 
 
 REFERENCES. 
 
 Foundations, Lecture, Coll. Eng., April 1896, 207; Hoisting, Trans. Am. 
 Soc. C. E., XXIX, 695; Mining Machinery, Henti Deschamps, Rev. Univ. 
 d. Mines, Jan. 1902; Some Notes on Ceryan Underground Hoisting Prob- 
 lems on the Whitwatersrand, A. W. K. Pierce, Trans. Am. Inst. Elec. Engrs., 
 Oct. 1903; The Mechanical Engineering of Modern Collieries, William 
 Bardill, Coll. Guard., Mar. 20, 1903; The Mechanical Equipment of Col- 
 lieries, George W. Winstanley, Coll. Guard., Feb. 26, 1904; Mining Machin- 
 ery, Butte, Montana, Coll. Eng., 1896, 25. 
 
 Compound-winding Engine, description of one at Cardiff, Trans. M. & 
 M. Eng., XLV, 205; Electrical-winding Engine, Dr. R. Hertzfeld, Engng., 
 Feb. 12, 1904; Hoisting Engine, Direct-acting, 111. Min. Inst., I, 145; Reel 
 and Hoist at Boston and Montana, E. &. M. Jour., Mar. 20, 1897, 285; A 
 Single-engine Plant, L. Sup. Min. Inst., IV, 1896, 81; Engine Breakdowns,
 
 HOISTING MACHINERY. 165 
 
 Michael Longridge, Mech. Engr., Sept. 26, 1903; and Trans. A. I. M. E., 
 XXXI, 348. 
 
 Equalizing Load on Hoister, A. Rogers, A. I. M. E., XVII, 305; Koepe, 
 A. I. M. E., XVII, 429; Koepe, Col. Eng., Oct. 30, 1896, 150; by Balance 
 Chain, Trans. N. of Eng. M. Inst., XLVI, Part III, 56, with calculations of 
 Methods, S. of M. Quart., 1889, 259; Counter Balancing, Tapering Rope, 
 S. of M. Quart., April 1889, 260; Air-shait ior Winding, Coll. Guard., Nov. 
 27, 1896, ion; Electric Hoist Description, Coll. Guard., Feb. 26, 1897, 397. 
 
 Loss of Load Due to Motion in Shaits, Coll. Guard., Vol. LXXX, 530; 
 The Dynamos of Winding Engines, ColL Guard., Vol. LXXXVI, 842.
 
 CHAPTER VI. 
 
 ELECTRIC GENERATION AND WATER-POWER. 
 
 The Application of Electricity to Mining. Though the 
 utility and economy of electricity for underground work have long 
 been recognized by the mining engineer, the conservatism forced 
 upon him by restrictive legislation has made its installation under- 
 ground one of slow growth. Its economy and convenience have 
 been recognized; its adaptability to all operations of mining, 
 lighting, .blasting, coal-cutting, hoisting, haulage, etc., its flexibility 
 to all distances, is conceded; it never vitiates the air, as do steam- 
 engines, nor does it fog or chill it, as does compressed air; and 
 it is'capable of automatic regulation, requiring but a small number 
 of employees. The low cost of maintenance and repairs, its high 
 economy, as compared with other motors, and the fact that there 
 is no leakage of power when the motor is not in use, gives positive 
 assurance of future benefit, especially where the power is to be 
 intermittently applied. It is capable of transmission over great 
 distances with small loss, and to localities where the haulage of 
 coal for steam generation is exceedingly difficult. In this field 
 it is unrivalled. The development of the steam-turbine and 
 impulse-wheels makes possible direct connections with the electric 
 generator, forming an ideal combination in point of economy of 
 power and space. 
 
 Two serious objections obtain against electric installations 
 for mining purposes, and they are all important. The first is the 
 risk of fire and shock of those coming in contact with the electric 
 current. They cannot be wholly eliminated without very great 
 cost of copper and its insulation. Nevertheless it is possible to 
 
 166
 
 ELECTRIC GENERATION AND WATER-POWER. 167 
 
 insulate to a high degree, to place the wires remote from acci- 
 dental contact by the men, and to reduce the fire risk by the use 
 of non- sparking devices and enclosed machinery. The partial 
 success in this direction has removed the chief objections to 
 underground installations. A less potent objection is that, elec- 
 tricity not being applicable for every operation of mining, the 
 owner is compelled to maintain two separate power systems 
 where percussion-drills are needed and large bodies of water are 
 to be pumped. 
 
 Electrical Units. The electric units are the ampere and the 
 volt, corresponding to quantity and pressure. The unit of resist- 
 ance, R, to the electric flow is the ohm, whose value corresponds to 
 a unit electromotive force and permits the flow of a unit of current. 
 The standard ohm is the resistance offered by a column of pure 
 mercury at o C., of uniform cross-section, 106.3 cms. long and 
 14.4521 grs. weight. 
 
 Electromotive Force. This is the electric pressure which forces 
 a current through a resistance, or is the difference of potential 
 between its termini. The unit of electromotive force, E.M.F., 
 is that pressure which will force a unit current through a unit 
 resistance. The unit is the volt, which is 0.6974 of the difference 
 of pressure between the poles of a Clarke cell at 15 C., whose 
 E.M.F. is 1.434 volts. 
 
 Current Capacity. An electric current, /, has unit strength 
 when a resistance of one ohm will afford an E.M.F. of one volt 
 between its ends. The unit is the ampere, that current which 
 will electrolytically deposit silver at the rate of 0.0011118 gr. 
 per second. 
 
 The quantity of electricity which passes through a given 
 cross-section of an individual circuit in / seconds when a current 
 of / amperes is flowing is equal to // units. The unit is the 
 coulomb, whose value equals the flow of one ampere for one 
 second. One ampere for one hour equals 3600 coulombs. 
 
 Capacity, C, is the property of a material for holding a charge 
 of electricity. A condenser has unit capacity when a constant 
 of electricity will charge it to a potential of one volt. The farad
 
 1 68 
 
 MANUAL OF MINING. 
 
 is the name of this unit, but since its value is much greater than 
 actual practical values, its millionth part, the micro-farad, is 
 used as the practical unit. 
 
 Electric Energy, W, represents the work done in a circuit 
 or conductor by a current flowing through it. The joule is the 
 name of this unit, its value being the work done by the passage 
 of one ampere through one ohm for one second. Electric power, 
 P, is one joule per second and represents the passage of one 
 ampere of current under a pressure of one volt. Its value is 
 measured in watts. One watt per second is a joule. 
 
 i watt =0.7373 ft.-lbs. 
 746 watts = i horse-power. 
 
 In order to avoid the use of large numbers, the term kilowatt, 
 or looo watts, is universally used in power and lighting. Allowing 
 for 15 per cent drop, the following table has been computed, show- 
 ing the quantity of current necessary to deliver one horse-power 
 at the various stated pressures: 
 
 To DEVELOP ONE MECHANICAL HORSE-POWER AT THE MOTOR SHAFT. 
 
 8 amperes 
 
 7 
 
 4 
 
 3-5 
 
 2 
 I.8S 
 
 i-75 
 1.6 
 
 no-volt circuit. 
 125 
 
 220 
 250 
 440 
 
 475 
 500 
 
 55 
 
 Ohm's Law. Resistivity, p, is the specific resistance of a 
 substance, and is the resistance, in ohms, of a cubic centimeter 
 of material to a flow of current between opposite faces. In 
 any electrical circuit, the current which flows equals the electro- 
 motive force divided by the resistance. Expressed in the form 
 of an equation, Ohm's law is IR =E, in which 7 is the number of 
 amperes flowing in an undivided circuit, E the algebraic sum 
 of all the E.M.F. in the circuit, and R the sum of all the re- 
 sistances in series in the circuit. 
 
 Since conductors offer more or less resistance to the passage
 
 ELECTRIC GENERATION AND WATER-POWER. 169 
 
 of current, there will be a drop or fall of potential, E d , along the 
 circuit, the amount of which will depend upon the resistance of 
 the conductor and the current that is flowing: 
 
 E d =IR. 
 The resistance of a conductor may be found from the equation 
 
 in which p is a constant, termed the resistivity,-whose value depends 
 upon the material and the temperature of the conductor; L is 
 the length in centimeters, and A the cross-section in square cen- 
 timeters. The resistance of pure metals increases with a rise 
 in temperature. A rise of i C. increases the resistance by 0.004 
 times that at o C. 
 
 Circular Mils. A circular unit is a circle o.ooi inch in diam- 
 eter. A conductor one foot long and one circular unit cross- 
 section is called a mil-foot. 
 
 Electric Currents. Two forms of currents are generated 
 and transmitted for power purposes the continuous current 
 and the alternating current. In the former, known also as the 
 direct current, the electric fluid maintains a flow in one direction 
 through the circuit. An alternating current is one which changes 
 the direction of its flow at regular recurring intervals. Dynamos 
 generating electrical energy are of two classes, according to the 
 currents to be produced. 
 
 Electric Generator. The continuous-current dynamo con- 
 sists of field- magnets, an armature, and a commutator. The field- 
 magnets are poles of iron arranged in pairs. The dynamo may 
 have two field-magnets, when it is known as the bipolar machine, 
 or several pairs of field-magnets, then called a multipolar ma- 
 chine. In the first case the armature revolves between the oppo- 
 site ends of two magnets, the other ends being directly connected. 
 In the multipolar machine the magnets project inward around 
 a ring, or casing, inside of which, and nearly touching, the arma- 
 ture revolves. Each pole is wound with wire, through which a
 
 MANUAL OF MINING. 
 
 current flows to maintain the magnetism in the poles. The 
 number of poles may be 2, 4, 6, 8, or more, the wire being wound 
 right-handed about the first pole and left-handed about the second 
 in order to produce with the same current the alternate polarity 
 in the respective piece . 
 
 FIG. 67. A Drum Armature. 
 
 The armature may be of the drum or the ring type, the former 
 being stronger and generally employed for power transmission. 
 A rotary shaft (Fig. 67) is built to a solid core of thin plates of 
 soft iron projecting radially and insulated from one another by 
 mica or paper. They are bolted together 
 with the surface slotted longitudinally. In 
 the slots are laid coils of wire whose project- 
 ing ends are connected with copper strips 
 composing the commutator. Fig. 68 illus- 
 trates the ring type of armature. 
 
 The commutator has an equal number 
 
 FiG.68.-Rh^Armature. of c PP er insulated strips, held firmly in 
 place by a bronze sleeve. Brushes collect 
 the current from the commutator, one set for each pair of poles. 
 They are composed of wire, or of small blocks of carbon coated 
 wi.h copper.
 
 ELECTRIC GENERATION AND WATER-POWER. 
 
 The Winding of Generators. According to the method of 
 winding the coils, there are series-wound machines,, shunt-wound 
 machines, and compound-wound machines. 
 
 In the series-winding, the arrangement of coils provides for 
 the whole current flowing continuously and directly from one 
 brush through the winding of the magnets, then through the 
 external circuit and back to the other brush. The wire around 
 the field-magnets is of large diameter and low resistance. This 
 form of winding is suitable only where a constant current ia 
 required, as for arc lighting (Fig. 69). 
 
 In the shunt-winding, two paths are open to the current. 
 Leaving one brush, one branch flows through the external circuit 
 and the other through the field-magnets. Both join at the other 
 brush before returning to the armature. This also gives constant 
 E.M.F., but with varying conditions in the external circuit. 
 The fields are wound with a large number of turns of small wire 
 (Fig. 70). 
 
 FlG. 69. Series Dynamo. FlG. 70. Shunt Dynamo. FIG. 71. Compound 
 
 Dynamo. 
 
 The compound-winding is a combination of both the shunt 
 and the series. The dynamo gives a constant E.M.F. The shunt 
 coils are of numerous turns of thin wire overlaid by a few turns 
 of series coils of thick wire. Both are coupled up (Fig. 71). 
 
 The Theory of the Dynamo. The dynamos are driven by 
 belts from the engine, or may be directly coupled to an engine, 
 water-wheel, or steam-turbine. Belt connection is cheaper than 
 the direct connection. 
 
 The rotation of the armature presents successively the plates.
 
 172 MANUAL OF MINING. 
 
 of soft wire in apposition to the poles of the field-magnets. This 
 develops in each coil about the corresponding plate a small 
 current of electricity, which is conveyed to their ends at the com- 
 mutator. The aggregation of the currents thus successively pro- 
 duced in the multitude of coils by a high rotary speed is a cur- 
 rent of high potential and large quantity, which may be drawn 
 by the brushes to the external wire circuit. The E.M.F. developed 
 is proportional to the number of turns of wire in the rotating 
 armature. 
 
 Two wires from the brushes to the electric motor or lamps 
 complete the external circuit for the current flow. 
 
 Alternating-current Machines. The wiring of the alternating 
 machine is similar to that of the direct-current, but the commutator 
 is replaced by two insulated rings on the armature shaft, one 
 end of the numerous armature coils being attached to one ring 
 and the other end to the other ring. The field-magnets on a 
 bronze base around the rotor constitute the stator. Their north 
 and south poles alternate, their windings being excited by a 
 similar continuous-current machine driven from the alternator. 
 
 The armature of the alternating machine is called the rotor. 
 It carries the secondary circuit and has as many coils systematic- 
 ally arranged around the laminated soft steel discs as there are 
 pairs of field-magnets. In its revolution an electric current is 
 generated in the armature coils, which changes its direction as 
 often, during each revolution, as there are pairs of field-magnets. 
 This gives rise to the alternating currents. 
 
 If a second set of coils, similar to the first, but independent 
 of them, be wound in the spaces between the first set on the 
 armature, then, by the revolution of the rotor, are induced two 
 distinct sets of currents, one leading the other by one-half a period. 
 Such a current is known as a two-phase current. If three sets 
 of independent coils be arranged equally around the armature, 
 then a three-phase current will be obtained, with its pulsations 
 having their maximum one- third of a period from one another. 
 The three-phase current is sometimes called a polyphase or 
 multiphase current.
 
 ELECTRIC GENERATION AND WATER-POWER. 173 
 
 The Frequency. The value of the current varies in the 
 different intervals from zero to a maximum, diminishes with a 
 like regularity to zero, and rises to the same maximum in the 
 opposite direction, returning finally to zero. One such com- 
 plete interval is a cycle, the tilde (~) being used to denote it. 
 The time in which the cycle of changes of such currents is com- 
 pleted is known as its period. The number of cycles com- 
 pleted per second is its frequency. The frequency usually 
 adopted in practice for power transmission is usually between 
 25 and 40 cycles. The number of cycles completed in a minute 
 is known as its rate of alternations. 
 
 As the points of maximum pressure are removed 120 from 
 that of the other circuits, the algebraic sum of the phase cur- 
 rents, when balanced, is at every instant equal to zero. Two 
 wires should be employed for each of the currents and three 
 return and three direct wires for the three-phase system. But, 
 for the reason stated above, the three return wires may be dis- 
 pensed with. Three wires then serve for the practical three- 
 phase circuit. 
 
 The Double-current Generator. This is wound to produce 
 either or both varieties of current, .and is used where the cost 
 of two separate dynamos has deterred many from introducing 
 electric power. There is a definite ratio between the voltages 
 on the direct and alternating sides. For the two-phase it is 
 0.76 of that of the direct current which it could produce, and 
 0.65 of the voltage if a three-phase current. Hence the double- 
 current dynamo is a low-voltage machine. 
 
 The difference of potential which can be obtained from a 
 D.C. generator cannot exceed 1000 volts, because of the ina- 
 bility to prevent short-circuiting at the commutator. The alter- 
 nating-current machine is not limited in voltage from this source 
 of leakage. 
 
 A serious defect of the D.C. machines is the sparking of 
 their brushes, due to faulty alignment, an insufficient bearing 
 or imperfect contact, and the short-circuiting in the windings. 
 These faults suggest simple remedies without further mention.
 
 MANUAL OF MINING. 
 
 The heating of the coils occurs when any current is forced through 
 the coils, but no harmful results ensue unless the temperature 
 rises above 80. When exceeded, the insulation softens and is 
 burned. This may be detected by the odor of burning shellac 
 or rubber. Dampness may produce the same results. 
 
 The copper bars of the induction machine are all short-cir- 
 cuited, rendering attention to insulation unimportant. 
 
 Properties of Copper Wire. The manufacturers' tables fur- 
 nish the data for wires of various diameters by which may be 
 known their resistances to the passage of a current. Below is a 
 brief table which will aid in the solution of problems connected 
 with the wiring of a mine for motors. 
 
 Wire No. 
 B.&S.; 
 American 
 and Bir- 
 mingham 
 Gauges. 
 
 Circular Mils. 
 
 Weight per 
 1000 Feet of 
 Common Insu- 
 lated Wire. 
 Pounds. 
 
 Safe Current- 
 carrying Capa- 
 city in Shaft or 
 Tunnel Work, 
 Amperes. 
 
 Resistances per 
 i ooo Feet in 
 Ohms, at 60 F. 
 
 B.&S. 
 
 B. 
 
 B.&S. 
 
 B. 
 
 B.&S. 
 
 B. 
 
 B.&S. 
 
 B. 
 
 000 
 OO 
 
 
 I 
 
 a 
 
 4 
 6 
 8 
 10 
 
 12 
 
 167,806 
 133.076 
 105.534 
 83.6Q2 
 66,371 
 41,741 
 26,251 
 16,509 
 10,383 
 6,529 
 
 180,625 
 144,400 
 115,600 
 90,000 
 80,686 
 56,644 
 41,209 
 27,225 
 17,956 
 11,881 
 
 603 
 483 
 385 
 308 
 246 
 152 
 
 102 
 69 
 
 49 
 35 
 
 647 
 
 517 
 415 
 
 III 
 198 
 148 
 
 101 
 
 7i 
 5i 
 
 295 
 250 
 
 210 
 
 180 
 J.S 
 no 
 80 
 60 
 40 
 30 
 
 315 
 272 
 227 
 
 193 
 161 
 118 
 85 
 64 
 43 
 
 .0606 
 .0764 
 .0964 
 .1219 
 
 1529 
 .2446 
 
 3879 
 .6214 
 
 9785 
 1-5520 
 
 .0564 
 
 755 
 .0881 
 .1131 
 .1262 
 .1797 
 .2471 
 3769 
 5670 
 8569 
 
 The Electric Symbols as applicable to motors are : 
 E = volts at the terminals; 
 V = volts loss in transmission ; 
 -|-y=E.M.F. at the generator terminals; 
 
 / = current required at the motor to deliver H, mechanical 
 
 horse-power at the shaft of the motor; 
 D = single distance between motor and generator in feet; 
 L = aggregate length of conductor in feet; 
 fl"=the number of mechanical H.P. delivered at the motor 
 
 shaft ; 
 A =the area of cross-section of conductor in circular mils;
 
 ELECTRIC GENERATION AND WATER-POWER. 175 
 
 d=the diameter of the wire in circular mils; 
 
 R = the total conductor resistance in ohms for the length, L y 
 or zD\ 
 
 W=the weight in pounds of copper conductor; 
 
 m = the commercial efficiency of motor in per cent ; 
 
 " " generator, 90 per cent; 
 
 S= " " " whole circuit, per cent; 
 
 p=per cent of energy lost in system. 
 
 Continuous-current Distribution. The following rules are 
 serviceable as guides in calculating the dimensions of D.C. con- 
 ductors. The E.M.F. varies directly with the amount of energy 
 transmitted. 
 
 Knowing the work to be done, line loss and E.M.F. at the 
 terminals and point of distribution, then the cross-section of the 
 conductor varies directly with the distance, and the weight, as 
 the square of the distance. 
 
 The conditions remaining as given, the weight of conductor 
 will vary inversely as the square of the E.M.F. at the terminals. 
 Given the amount of power which is to be transmitted by a stated 
 weight of conductor and loss in distribution, then the distance 
 over which the power can be transmitted varies directly as the 
 E.M.F. For a given amount of power and a given conductor it 
 varies as the square of the E.M.F. 
 
 The Resistance of Conductors. The resistance of a mil-foot 
 of pure copper at o C. = 9-59 ohms. 
 
 The resistance of a mil-foot of 96 per cent conducting Cu 
 (which is the commercial conductivity usually specified at 70 F.) 
 = 10.81 ohms. 
 
 TT 
 
 Then the electrical horse-power at the motor terminals = . 
 
 m 
 
 __ IE , _ 
 T 7 and / = 5; amperes. 
 740 hm 
 
 The resistance of the conductors, both ways, is 
 21.62!) lo.SoZ,
 
 176 MANUAL OF MINING. 
 
 The line-drop 
 
 v IR 2I * 2lD an d A 
 
 Emv ' 
 
 EXAMPLE. A motor of 50 horse-power of 0.93 efficiency is to be installed 
 in a mill with the terminal volts at 500. The transmission distance is 500 
 feet, and the allowable loss in that line 6 per cent. Required the size of con- 
 
 coo 
 
 ductors. The E.M.F. at dynamo=- = 531.9 volts. 
 
 0.94 
 
 The line-drop V= 531.9 500=31.9 volts, and the area of the conductor 
 16,128.5X50X500 
 
 500X0.93X31.9 
 
 27,863 circular mils. 
 
 The current /= - = 80.6 amperes. 
 500X93 
 
 The National Code allows only 97 amperes for No. 5 B. & S. gauge con- 
 ductor. To transmit 80.6 amperes safely, at least No. 4 wire would be re- 
 quired. For underground work a No. 6 wire would be permissib e. 
 
 The resistance of No. 46. & S. wire =0.2480 ohms at 20 C. 
 
 The resistance of 500 feet of No. 4 wire =0.1240 ohms at 20 C. 
 
 Then the volts-drop =7^=80.6X0.1240= 09.94 volts; hence volts at dy- 
 
 OQ.Q 
 
 namo = 500+ 99.94 = 599-94 volts; and per-cent drop = = 1.8 70 per cent 
 
 599-9 
 
 General Wiring Formula for Alternating-current Distribu- 
 tion. 
 
 Let D= distance of transmission one way in feet; 
 W = total watts delivered to consumer; 
 p = per cent of power, W, lost in line ; 
 E = voltage between main conductor at the consumer's 
 
 end of circuit; 
 K= constant, whose value depends on the kind of system 
 
 and the power factor table, page 178; 
 A factor for determining weight of conductor (table on 
 
 page 178); 
 F= variable, whose value depends on the kind of system 
 
 and the nature of the load; and 
 M = variable, which depends on the size of wire, the power 
 
 factor, and the frequency.
 
 ELECTRIC GENERATION AND WATER-POWER. 177 
 
 Then the area of conductor = pE2 in circular mils. 
 
 Volts lost in line =0.0 1 pEW. 
 
 WF 
 
 Current in main conductor s = -g-. 
 
 D*WKA 
 
 Pounds of copper = ^^. 
 i,ooo,oooPE 2 
 
 EXAMPLE. Find the size of conductor necessary to transmit 40,000 watts to 
 a distance of 20,000 feet, the line voltage being 2000 and the permissible loss 
 10 per cent. Transmission is to be by the three-phase three-wire system, 
 at 60 cycles, power factor of 85 per cent. 
 
 20,000X40,000X1500 . 
 
 Area of conductor = ' = 30,000 circular mils. 
 
 10X4,000,000 
 
 The nearest size, No. 5 B. & S. gauge. 
 
 Volts lost in line=o.oiX 10X2000X1.06= 212 volts. 
 
 . T . 40,000X0.68 
 Current in main conductors (/) = = 13.6 amperes. 
 
 Resistance. In alternating current resistance is the same 
 in kind as in continuous current, but is generally considered as 
 negligible in comparison with other alternating-current factors. 
 The effective value of an alternating current is 1.41 times its 
 apparent value. 
 
 The Power Factor of an alternating-current circuit is the 
 ratio of the kilowatts, as indicated by a watt-meter, to the ap- 
 parent watts or volt-amperes. It enables one to readily deter- 
 mine the true energy in a circuit when the apparent energy is 
 known; likewise, the resistance when the impedance is known; 
 the energy E.M.F., when the impressed E.M.F. is given, and 
 the energy current when the maximum current is known. 
 
 Electric Systems. The available electric power systems com- 
 prise the continuous-current, the polyphase alternating current, 
 and a combination of the two. When the machinery to be installed 
 is merely traction and lighting, the continuous-current system 
 will be employed. Being essentially a low-pressure current, it 
 must be used locally and cannot be transmitted over great distances. 
 
 Either a two-wire or a three-wire line may be supplied by
 
 i 7 8 
 
 MANUAL OF MINING. 
 
 
 Values of A, \ 
 
 as 1 - 
 
 Values of K, 
 Per Cent Power Factor. 
 
 Values of F, 
 Per Cent Power Factor. 
 
 IOO 
 
 95 
 
 QO 
 
 85 
 
 So 
 
 IOO 
 
 95 
 
 90 
 
 85 
 
 80 
 
 Single-phase 
 Two-phase (4-wire). . . . 
 Three-phase (3-wire).. . 
 
 6.04 2160 
 I2.o8|io8o 
 9.06 1080 
 
 2400 
 1200 
 I2OO 
 
 2660 
 133 
 1330 
 
 3OOO 
 I5OO 
 1500 
 
 3380 
 1690 
 1690 
 
 1. 00 
 
 5 
 .58 
 
 1-05 
 
 53 
 .61 
 
 i. ii 
 
 55 
 .64 
 
 1.17 
 59 
 .68 
 
 '* S 
 .62 
 
 .72 
 
 VALUES OF M FOR WIRES 18 INCHES APART. 
 
 Size of 
 Wire, 
 
 Area of 
 Wire, 
 
 Weight 
 of Wire, 
 
 Resistance 
 of Wire, 
 per 1000 
 
 25 Cycles, 
 Per Cent Power 
 Factor. 
 
 40 Cycles, 
 Per Cent Power 
 Factor. 
 
 B & S. 
 
 Circular 
 
 
 Feet at 
 
 
 
 Gauge. 
 
 Mils. 
 
 (in Lbs.). 
 
 20 C. 
 
 (in Ohms). 
 
 95 
 
 9 
 
 85 
 
 80 
 
 9S 
 
 90 
 
 8s 
 
 So 
 
 coo 
 
 211,600 
 
 640.73 
 
 .04879 
 
 2 3 
 
 I . 2C) 
 
 i-33 
 
 34 
 
 1.58 
 
 S3 
 
 1.61 
 
 1.67 
 
 00 
 
 133,079 
 
 402.97 
 
 07758 
 
 14 
 
 T.l6 
 
 1.16 
 
 .16 
 
 i. 35 
 
 
 1.35 
 
 '37 
 
 4 
 
 41,742 
 
 126.40 
 
 2473 
 
 .02 
 
 I. 00 
 
 I. 00 
 
 I. 00 
 
 1.05 
 
 .06 
 
 1.03 
 
 I. 00 
 
 5 
 
 33,102 
 
 100.23 
 
 .3120 
 
 .00 
 
 I. 00 
 
 I. 00 
 
 I. 00 
 
 1.03 
 
 I.OI 
 
 I. 00 
 
 I. 00 
 
 6 
 
 26,250 
 
 79-49 
 
 3934 
 
 .00 
 
 I. 00 
 
 I.OC 
 
 .00 
 
 I. 01 
 
 .00 
 
 I. 00 
 
 r .00 
 
 7 
 
 20,816 
 
 63-03 
 
 .4958 
 
 For sizes 7 and 8 wire the constants 
 
 8 
 
 16,509 
 
 49-99 
 
 .6250 
 
 are within .05% (average) of 5 and 6. 
 
 Size of 
 
 Area of 
 
 Weight 
 of Wire 
 
 Resistance 
 of Wire. 
 
 60 Cycles, 
 Per Cent Power 
 
 1 25 Cycles 
 Per Cent Power 
 
 Wire, 
 
 Wire, 
 
 
 per 1000 
 
 Factor. 
 
 Factor. 
 
 B. &S. 
 
 
 oare, per 
 
 Feet at 
 
 
 
 Gauge. 
 
 Mils. 
 
 (in Lbs.). 
 
 20 C. 
 
 (in Ohms). 
 
 95 
 
 90 
 
 8s 
 
 So 
 
 95 
 
 90 
 
 8S 
 
 So 
 
 OOO 
 
 211,600 
 
 640.73 
 
 .04879 
 
 1.62 
 
 .84 
 
 i. 00 
 
 2.09 
 
 2-3S 
 
 i.86 
 
 .24 
 
 3-40 
 
 00 
 
 133,079 
 
 402-97 
 
 07758 
 
 i-34 
 
 52 
 
 i. 60 
 
 1.66 
 
 1.86 
 
 .18 
 
 .40 
 
 2.57 
 
 5 
 
 33.102 
 
 100. 23 
 
 2473 
 .3120 
 
 i. 08 
 
 .08 
 
 i. 06 
 
 1.04 
 
 I. 21 
 
 35 
 .27 
 
 3 
 
 i-43 
 1.31 
 
 6 
 
 26,250 
 
 79-49 
 
 3934 
 
 i .03 
 
 .02 
 
 I. 00 
 
 I. 00 
 
 I. 12 
 
 .14 
 
 .14 1.13 
 
 I 
 
 20,816 
 16,509 
 
 63-03 
 49-99 
 
 .4958 
 .6250 
 
 For sizes 7 and 8 wire the constants 
 are within .0^% (average) of 5 and 6. 
 
 the D.C. generators. With the latter, the generator must have 
 three wires to deliver continuous current of two pressures, one 
 being double that of the other. If this type of dynamo is not 
 used, either two generators are installed in series or else some 
 complex form of balancing device is required. 
 
 When the power is required for pumping, lighting, coal-cut- 
 ting, etc., the alternating current can be more economically 
 transmitted. Its machines are easier to construct than D.C. 
 machines of similar pressures and its motors give less spark and
 
 ELECTRIC GENERATION AND WATER-POWER. 179 
 
 are of more general application. Its higher efficiency is due to 
 the high voltage to which it can be carried. This is what gives 
 the polyphase current its chief claim. Particularly is this true 
 in mountainous districts, where sources of water-power distant 
 from the mines can be developed for electric generation and 
 thence distributed to various points. 
 
 In the event that both systems must be maintained in the 
 mine, both currents may be generated and conducted, or it is 
 practicable to produce the cheaper polyphase alternating cur- 
 rent, and, by the use of converters at suitable points, lo change 
 only such portion of the energy into direct current as will be 
 required for the haulage and lighting, leaving the remaining 
 portion of the current for the other mining operations. 
 
 Converters. Changes from the alternating current to a 
 direct current, or the reverse, can be effected by the means of a 
 rotary converter. Changes from a high-pressure to a low-pres- 
 sure current, or the reverse, whether D.C. or alternating, may be 
 effected by a transformer for stepping down or stepping up in 
 pressure. As reliability is of the utmost importance, transformers 
 are always installed of twice the capacity for the expected service. 
 
 Their efficiency is about 96 per cent. 
 
 The Transmission Difficulties. The great objection to elec- 
 tricity is the high cost of wire and erection. Economy of copper 
 wire is possible only by employing high voltage in the currents, 
 as shown by the formulae already discussed. As an example, 
 195 Ibs. of copper can transmit 50 amperes at 2000 volts over a 
 5ooo-foot line with the same percentage of line loss that 795 Ibs. 
 of No. 36. & S. gauge can transmit the same quantity at 500 
 volts. But in the transmission other difficulties are introduced. 
 The insulation must be perfect and the wires separated from one 
 another sufficiently to prevent an induction of one current upon 
 the other. From both sources the line loss would be large. 
 With proper insulation, however, the limit of voltage which 
 can be carried would be fixed by the distance between the 
 wires. There must be a wide gap between them, for, with the 
 enormous voltage which is now being used, we are approaching
 
 l8o MANUAL OF MINING. 
 
 the conditions in lightning, not in a single stroke but a number 
 of strokes following each other in very rapid succession. The 
 length across the air-space increases as the voltage increases, and 
 for currents of 50,000 volts the arms carrying the wires at their 
 ends must be 12 feet long. 
 
 There is no limit to the E.M.F. which the alternating-cur- 
 rent machines can impart to the current. The majority of long- 
 distance transmission companies employ the three-phase system 
 and at as high a difference of potential as possible. A high- 
 voltage current may be obtained either directly at the dynamo 
 whence it is delivered to the distant end, or the dynamos may 
 generate a low-tension current which is transformed into the high- 
 voltage current at the dynamo station for transmission to the 
 distant point. At the mine it may be transformed wholly into 
 alternating current of tension suitable for the machines, or only 
 in part, the balance being converted to a continuous current for 
 haulage. 
 
 The Most Economical Area of Wire. The losses in trans- 
 mission are due either to resistances or to leakage. The drop 
 in pressure caused by leakage of the current from the wire through 
 the points of support is deduced by insulating the supports and 
 an enveloping tube of non-conducting material around a rubber- 
 covered wire. The drop in voltage, due to interaction of 
 neighboring parallel circuits, is averted by an ample air-gap. 
 That due to the resistance of the wire to the passage of the cur- 
 rent is controlled by the use of a larger wire. The resistance of 
 the wire varies inversely as the area of its cross-section. A small 
 wire offers a great resistance to the electric fluid. This loss 
 reduces the available pressure at the terminals, and consequently 
 the amount of power there obtained. It develops heat in the con- 
 ductor, which is not only injurious to the wire, but, without 
 secure fire protection, is also dangerous to the surrounding medium. 
 
 In calculating the dimensions of the conducting wire a balance 
 is made between the elements great first cost of a large quantity 
 of copper and the waste of power during the life of the plant by 
 use of small wire. The allowable expenditure of wire, in any
 
 ELECTRIC GENERATION AND WATER-POWER. 181 
 
 given plant, will determine the E.M.F. which can be used; the 
 efficiency required of the plant will fix the, allowable line loss. 
 Hence the most economical area of conductor is that for which 
 the annual interest on the capital outlay equals the annual cost of 
 the energy thus wasted. 
 
 The Allowable Voltage in Mines. There are many con- 
 flicting requirements for electric wiring in mines. The great 
 distances of distribution require high voltage to save copper. 
 This demands a high degree of insulation, which is difficult to 
 maintain in the moist air and the mine-gases. Moreover, the 
 conditions of work require bare wires, and their position is such 
 as make them dangerous to life. The porcelain or glass tubes 
 are convenient insulators, but they do not prevent shock. 
 
 The lowest difference of potential causing fatal accidents is 
 220 volts with a direct current and no volts with an alternating 
 current. The more the difference of potential increases beyond 
 these figures the more must all tensions be regarded as equally 
 dangerous. Though a higher voltage would be more economical 
 in the cost of copper for a given power transmission, the miner is 
 restricted to a low voltage because of the risks imposed by fire and 
 shock. The allowable maximum voltage "has been fixed at 450. 
 
 In the three-wire system one trolley wire is used on each side 
 of the traction generator or to the free sides of two generators 
 in series. In the former case the conductors from the rails should 
 connect with a small dynamo to keep the rail pressure to one-half 
 that of the main terminal. In the latter case all rail conductors 
 are taken to a common point of connection. By grounding the 
 neutral wire the shock at contact would not be so serious unless 
 both wires were simultaneously touched. The trolley circuits, 
 however, subject the maker of the contact to the full discharge 
 of pressure between the two sides of the three-wire system. So, 
 too, a ground connection of the high-pressure circuits, or a cross 
 between the high- and low-pressure circuits, inside the trans- 
 former, would subject him to fatal results. 
 
 For traction a pressure of 250 volts is qtiite sufficient, even 
 with an induction motor, for i; can be operated by transformers
 
 1 82 MANUAL OF MINING. 
 
 at every few hundred feet to keep the voltage down. Incan- 
 descent lamps on a two- wire D.C. system are operated on 125 
 or 250 volts, and at double these pressures on the three-wire 
 distribution system. On alternating-current circuits the lamps 
 are supplied by near-by transformers. Lamps and stationary 
 motors are carried under a maximum pressure of 250 or 500 
 volts, with corresponding differences of pressure between any 
 wire and the earth of one-hal? the quantity. 
 
 The Motor. A motor converts the current into mechanical 
 energy. It is directly connected, or is belted, to the mechanism 
 to be operated. The construction of the motor is practically the 
 same as the dynamo, but its function is exactly the opposite. The 
 current flows in the reverse order, passing through the coils of 
 the armature and the field-magnets. These react upon the coils, 
 carrying a current into the armature and thus produce rotation. 
 The force turning the armature is called its torque and is its 
 capacity to do work. Around the field, magnetized by the cur- 
 rent flowing around the magnets, a current, called the back E.M.F., 
 is induced in the opposite direction. This increases with the 
 current intensity and the number of coils and their speed. Taking 
 current in proportion to the work performed, the motor is self- 
 acting. 
 
 The continuous motors are wound in three ways, correspond- 
 ing to those of the dynamo, each having its own field of useful- 
 ness. In the series motor there is but one continuous circuit 
 through the stationary and rotary pieces. This exerts a greater 
 starting torque, but its speed will change with every variation 
 of the load. It is intended for constant load only. If it is over- 
 loaded, its speed is slackened. Its torque is increased by its 
 greater draft of current, and if it is overloaded, it runs slower, 
 the field-magnets heat up and burn out, or a fuse blows. Re- 
 moving the load suddenly causes it to race, because the field 
 is weakened and the back E.M.F. is diminished. The increased 
 speed which ensues may ultimately injure the motor. The shunt 
 motor will maintain a constant speed and a back E.M.F. even 
 with variations of the load, but its starting power is low. Designed
 
 ELECTRIC GENERATION AND WATER-POWER. 183 
 
 for the maximum load which it is expected to drive, it is employed 
 in mining work. It is heavier than the series motor. The 
 compound motor is used about the mines and combines the 
 shunt and the series features of motors. 
 
 Alternating-current Motors. These motors are either syn- 
 chronous or non- synchronous. The speed of the former type 
 bears a constant relationship to the frequency of the driving 
 current. Being unable to start until it synchronizes with that 
 of the generator, miners use it very little. 
 
 In the non-synchronous motor, also called an induction" motor, 
 the speed varies with the load and is independent of the frequency 
 of the current. It is self -starting. Induction motors may be 
 obtained of the two-phase four-wire type, or the polyphase type 
 with three wires. In the two-phase motors each coil of a set in the 
 rotor receives its own current. One induces a current in the 
 conductors of the rotor, causing attraction between them and 
 the stator coils. The slight rotation that follows changes the 
 current to the next set. Each set of windings produces similar 
 results and adds to the rotary effort. In the three-phase system 
 continuous rotation occurs under similar conditions, except that 
 the entering current is divided among three separate sets of coils. 
 A back E.M.F., as an opposing pressure, is induced in the con- 
 ductors of the rotor, as in the D.C. machine. When the load 
 is increased the speed falls, the opposing pressure becomes less. 
 There is an increase in the available driving pressure, or torque, 
 and an increase of current in the rotor. This alters its speed 
 relative to the rotating field. Every alteration of current in the 
 rotor conductors reacts on the stator coils and reduces the strength 
 of the magnetic field, until the energy supplied to the motor is 
 insufficient for the work and the motor stops. In this respect 
 the induction motor differs from the D.C. motor. An increase 
 of load brings the former to a standstill, but results in the burning 
 of armature coils in the latter. An induction motor can take 
 an excess of 20 per cent over its normal load without injury. 
 Having neither commutator, induction brushes, nor movable 
 contacts, it is the ideal motor for mining machinery where the
 
 1 84 MANUAL OF MINING. 
 
 air may be laden with dust, gas, or moisture. The power factor 
 is over 90, with an efficiency proportionately high. 
 
 The objections urged against the polyphase induction motoi 
 are its excessive starting current, which results in a drop of voltage 
 with consequent bad regulation to other apparatus in current, 
 and its inflexibility in the matter of speed regulation. The first 
 objection against the short-circuited secondary type is negligible 
 if the power be increased at the generating station on starting. 
 The machine is essentially a constant- speed motor. When 
 variable speed is required it would be more advantageous to 
 employ a variable resistance with the secondary type of induction 
 motor effected by a rheostat. 
 
 Rheostats. There is no back E.M.F. in the circuits at the 
 moment of starting a motor, and, to prevent the heavy entering 
 currents from doing damage to the armature, extra resistances 
 must be introduced into the current until such times as the motor 
 has got up speed. The starting resistance consists of coils of 
 German-silver or iron wire contained in fire-proof boxes. A mova- 
 ble contact, turned to one side or the other from its central position, 
 determines the direction of rotation. It is operated by an attend- 
 ant.' On closing the circuit the current flows through the entire 
 resistance of all the coils. As the motor speeds up, one after 
 the other of the coils in the circuit is gradually cut out. The 
 diverter of columns of iron and mica in five separate portions is 
 used for locomotives. 
 
 The resistance, on shunt- and compound-wound machines, is 
 placed in the shunt winding of the fields to increase or decrease 
 the E.M.F. There is usually on pump-motors an additional 
 safeguard of an automatic release for the rheostat contact lever. 
 The contact arc is worked against a spring, and is held, when 
 fully on, by means of a magnet, which is connected in series with 
 the shunt fields of the motors. 
 
 For alternating-current machines the starting resistance con- 
 sists of two rings connected to the conductors and insulated from 
 one another and from the motor shaft. Fitted to them are brushes 
 with a resistance inserted between them, which can be automatic-
 
 ELECTRIC GENERATION AND WATER-POWER. 185 
 
 ally cut out as the motor attains its speed. These slip- rings 
 are covered in mining machines to confine the spark. Another 
 class of starter is that in which the current passes through water 
 into which is lowered a lead cone from the surface to the bottom 
 where the other terminal is located. This form of switch is free 
 from spark, though it might be dropped too quickly and thus be 
 almost valueless. 
 
 The variable speed induction motor has a two- drum con- 
 troller for resistance and reversing operated by a single handle. 
 The resistance is in series with the rotor windings, and is cut out 
 in the usual way until the approximately constant speed is attained. 
 
 Reversing-motors. When the direction of the motor is to be 
 reversed, its brushes should bear vertically on the commutator 
 and a re versing- switch be used. This appliance for changing the 
 direction of the current in the field-magnets is arranged as a part 
 of the starting resistance. 
 
 Enclosed Motors. If possible, motors should be run open, 
 for if totally enclosed a much larger motor will be required for the 
 same work on account of the difficuLy of dissipating the heat 
 developed in the coils. If precautions are. observed not to over- 
 load the machine, the danger is reduced and certainly would not 
 be greater than that existing from either cause in the daily opera- 
 tions. They are rarely placed near gaseous mixtures, and the 
 danger of igniting neighboring timbers by their spark is very 
 slight. For safety in mines, however, electric motors are enclosed 
 to protect them from fall of rock, dust, and dampness. 
 
 The Efficiency of a Motor. Not all the current which the 
 motor receives is usefully applied. Part of it is spent in heating 
 the circuit of the motor and part in overcoming frictional resist- 
 ance. The remainder, which is available for mechanical work, 
 is still, however, quite large. As large motors have high efficiency, 
 it is possible to attain very high results, if their capacity be large 
 enough. This is, however, a question of weight and first cost. 
 The efficiency of a machine is not far from 90 per cent under 
 average conditions, and will reach 92 per cent when operating 
 under the load and speed for which it was designed. The
 
 i86 
 
 MANUAL OF MINING. 
 
 table below gives the number of watts necessary to produce one 
 effective horse-power at the terminals of a motor with stated 
 commercial efficiencies. Thus a motor with 10 effective horse- 
 power having an efficiency of 85 per cent will require 8780 watts 
 of current. 
 
 Efficiency 
 of a Motor. 
 
 Watts to be 
 Delivered to it 
 per Horse-power. 
 
 T.OO 
 
 0.98 
 
 74 6 
 764 
 
 o-9S 
 
 774 
 
 0.90 
 0.85 
 0.80 
 
 830 
 878 
 933 
 
 o-75 
 
 1000 
 
 EXAMPLE. The efficiency of a 50-H.P. motor is 0.90. It is to be driven a 
 half mile from a dynamo whose efficiency is 0.85. The modulus of the engine 
 driving the generator is 0.80. Required th horse-power of the engine. 
 Volt ge, 400. 
 
 The motor requires 830 watts per horse-power or 41,500 watts total: 
 
 41,500-^400=103.75 amperes. 
 
 For this the wire must be No. 4. Its resistance, per thousand feet, is 
 0.245 ohms. 
 
 The watts lost per thousand feet are 
 
 PR= 104 X 104X0.245= 2650. 
 
 For the mile of line in the circuits 13,992 watts are lost. 
 The dynamo must therefore furnish 
 
 41,500+13,992=54,492 watts. 
 
 This corresponds to a voltage at the dynamo of 533, and a <oss of 33 per cent 
 55,492 
 
 B.H.P= 
 
 = 87.6 nd 
 
 746X0.85 
 
 A No. 2 wire, B. & S. wire, will show a loss of 20 per cent, a dynamo voltage 
 of 481, and an I.H.P. of 100. 
 
 Water-power. In the immediate vicinity of the wheel water 
 the power has been employed for various operations, but not 
 until the modern successful utilization of electricity has a 
 cheap efficient means been discovered for its transmission to 
 great distances. The installation of electric plants is opening 
 the possibilities of water-power to an enormous degree.
 
 ELECTRIC GENERATION AND WATER-POWER. 187 
 
 The gross power possessed by water and the potential energy 
 which is capable of being transmitted and converted into kinetic 
 energy is measured by the product of the weight of water dis- 
 charged by the height, h, through which it has fallen. Assuming 
 the weight of a cubic foot of water at 62.5 and Q as the number 
 of cubic feet discharged per minute, then a horse-power equals 
 o.ooi6i<2&. The net power obtained from this in effective work 
 is from 40 to 90 per cent according to the character of wheel 
 used. 
 
 This energy develops power by driving one of several types 
 of wheels, of which the undershot and the overshot were the 
 earliest. They consume, however, enormous volumes of water 
 and are used only for small machinery. The turbine-wheels, 
 placed horizontally, revolving under a pressure due to a head 
 of water above, give a high efficiency with the use of a large 
 volume of water. These wheels are not very large and are en- 
 cased in a globe or cylindrical casing above which is a penstock 
 through which the water flows. The water enters centrally 
 and discharges circumferentially and produces a high rate of 
 revolution. The heads at which these turbines are used rarely 
 exceed 200 feet. 
 
 Impulse Wheels. The more recent forms of wheels are of 
 the impulse type, in which a small wheel receives at its circum- 
 ference the impact of a stream flowing with a high velocity due 
 to a head which in some cases is as great as 1900 feet. The 
 periphery of the wheel carries a number of small cup-shaped 
 vanes whose curves and position are such as to receive the full 
 impact of the stream and to discharge the water at a velocity of 
 nearly zero. The efficiency of these wheels is therefore high, 
 and their construction is very simple ; the inferior limit of head 
 with which they may operate to advantage is about 30 feet. The 
 maximum limit of head is determined only by the strength of 
 the material in the wheel. As is the case with the De Laval 
 steam-turbine (Fig. 48), multiple jets are also used on water-tur- 
 bines. 
 
 The diameter of the wheel is proportioned to the rate of
 
 1 88 MANUAL OF MINING. 
 
 revolution desired for the main shaft. Usually the latter is that 
 for which the electric generator has been wound and designed. 
 The diameters vary between 18 and 90 inches. 
 
 There are several types of impulse motors in America, of 
 which the Doble, Pelton, and Knight are excellent patterns. 
 The efficiency of these wheels when properly regulated by gov- 
 ernors is 85 per cent of the theoretical head, due to the velocity 
 of discharge from nozzles against the wheel-cups. In the moun- 
 tainous districts, where the numerous creeks do not carry much 
 water but have a large fall, these wheels may be used to advan- 
 tage by laying a pipe-line from the wheel to some elevated reser- 
 voir in the creek constructed for the purpose. Not infrequently 
 two or even three nozzles feed the wheel when the quantity is 
 large enough to permit it. 
 
 The supply of water delivered to the wheel is usually esti- 
 mated in miners' inches, representing a flow of something near 
 1.5 cubic feet per minute. The miner's inch is known as the 
 volume of water which can be discharged through each square 
 inch of an aperture 2 inches high and 4 inches long which is cut 
 through a plank 1.25 inches thick, the lower edge of the aper- 
 ture being 2 inches above the bottom of the measuring-box and 
 the upper edge 5 inches below the level of the water. 
 
 Flow of Water through Pipes. Owing to the roughness of 
 the internal surface of a pipe and the restrictions which occur at 
 elbows, joints, and valves, *the actual velocity of discharge of 
 water from a pipe is not equal to that due to the pressure at its 
 inlet. Any change of area or of direction affects the flow of the 
 water and produces frictional resistances which are directly pro- 
 portional to the length, inversely as the diameter, and increase 
 with the velocity. Though a theoretical formula may be evolved 
 which expresses the relation between the discharge of water and 
 the head producing such discharge, empirical formulae are em- 
 ployed with coefficient inserted to provide for the various elements. 
 
 The formulae below are given to assist the engineer in deter- 
 mining the dimension of a pipe requisite for the given flow of 
 water with a stated resistance.
 
 ELECTRIC GENERATION AND WATER-POWER, 
 
 189 
 
 Fluid Friction in Pipes. Assuming very long clean pipes of 
 uniform size, the resistance which is afforded by the interior 
 surface of the pipe to flow is measured in terms of the head. 
 
 Let Q =the quantity of water discharged, cubic feet per second; 
 v=the velocity in feet per second;
 
 190 MANUAL OF MINING. 
 
 d=ihe diameter of the pipe in feet; 
 /=the length of pipe in feet; 
 /=the coefficient of friction; 
 Jf=the gross head of water the difference in elevation 
 
 between the two termini of the pipe; 
 h =the frictional loss measured in terms of feet of head, and 
 h t =the effective head =H h. 
 
 In clean pipes of smooth bore, / =0.004 an d in ordinary mine 
 pipes / =0.006. 
 
 >.55S 
 | 
 
 Qi.802 
 
 h =0.000606/^857 
 
 /)i.8o2-io.aoo 
 
 d = \ o.ooo6o6/ 
 
 L h J 
 
 Q, for the maximum horse-power, 
 
 o(.ooi 7 0- 5SS 
 The maximum horse-power to be obtained is equal to 
 
 H.P.=: 
 
 In Fig. 73 are curves based upon the above formulae, giving 
 the loss of head due to the frictional flow through pipes which 
 are ordinarily used in pumping and for transmitting power to 
 impulse wheels. On the right-hand side will be found a verti- 
 cal line giving the theoretical horse-power necessary to raise a 
 given quantity of water indicated on the left vertical line through 
 a height of 100 feet. Thus, raising 7 cu. ft. per second through 
 any pipe for a height of 100 feet requires 80 horse-power theo- 
 retically. When the height is other than 100 feet the result 
 obtained from the diagram is a multiple of the loo-ft. eleva-
 
 ELECTRIC GENERATION AND WATER-POWER. 
 
 191 
 
 tion. To raise the same quantity of water in the same pipe 
 through 600 feet of pipe will therefore require (80X6 = ) 480 
 horse- power 
 
 FEET OF HYDRAULIC HEAD LOST PER 1,000 LINEAR FEET OF PIPE. 
 
 FRICTIONAL HORSE-POWER PER 1,000 LINEAR FEET OF PIPE. 
 
 FIG. 73. Loss of H\draulic Head in Pipes. 
 
 The horizontal line at the top gives the amount of head lost 
 by friction, and the horizontal line at the bottom the horse-power 
 needed to overcome the friction. Thus let it be required to find 
 the head lost and the horse-power while raising 7 cu. ft. pei
 
 IQ2 MANUAL OF MINING. 
 
 second in a 1 2-inch pipe 5000 feet long on a slope of 12 per cent. 
 The curve shows that 1000 feet of 1 2-inch pipe carrying this 
 volume of water consumes 20 feet of head. Thus 5000 feet will 
 consume 100 feet of the head. Likewise 1000 feet will consume 
 in this lost head about 28 horse-power, while 5000 feet will con- 
 sume 140 horse-power in overcoming friction. 
 
 The frictional loss in a lo-inch pipe 1000 feet long, carrying 
 2,ico,ooo gallons per day (3.245 cu. ft. per second), is 12.8 feet. 
 An 8-inch pipe would have consumed 38.5 feet of head under 
 the same conditions of length and discharge. 
 
 EXAMPLE. A pipe 500 feet long is to deliver 4 cubic feet per second. Re- 
 quired its diameter if its frictional loss of head allowed is 30 feet. For 30 feet 
 hydraulic head per thousand feet of pipe and 4 cubic feet of flow the diameter 
 should be 9 inches. 500 feet of pipe consuming 30 feet will correspond to 
 a length of 1000 feet consuming 60 feet of head. For this condition and 
 4 cubic feet of discharge the pipe may be of 8 inches diameter. 
 
 Kutter's Formula for Ditches and Flumes. Fig. 74 contains 
 curves plotted according to the Kutter formula, which is accept- 
 able for all ditches and flumes of medium dimensions. The 
 formula is based on the average condition of the frictional sur- 
 face. The upper curves are for flumes with hydraulic radii of 
 from i to 9. The coefficient of roughness, N, for flumes is taken 
 at 0.01 1, and for ditches at 0.026. On the left are laid off the 
 gradients in terms of length of line giving a fall of i foot, and the 
 computed velocity of flow is on the top line. Then the four 
 lower curves on the left represent the relation between velocity 
 and grade in ditches having hydraulic mean radii, H.M.R., of 0.5, 
 i, 2, and 3, respectively. By the hydraulic mean radius is under- 
 stood the quotient obtained by dividing the cross-sectional area 
 of the ditch by the length of the watered surface of the ditch. 
 
 EXAMPLES. i. What would be the loss of head in pumping 2000 gallons of 
 water per minute through an 8-inch pipe 600 feet high? (^ = 4.44, and 
 ^=47.6, or 38.3 feet, according to the equations employed. 
 
 2. What horse-power is consumed in overcoming friction in the previous 
 example? Assume h to be 38.3. 19.4 H.P. 
 
 The flow is 4.44 cubic feet per second, and the horse-power is 0.1134 QH.
 
 ELECTRIC GENERATION AND WATER-POWER. 
 
 193 
 
 3. What horse-power will be given out by the discharge of 400 cubic 
 feet of water per minute from a pipe of 13 inches diameter, 2600 feet long, 
 with a head of 400 feet? 
 
 The loss of head is 32.53 feet, and the horse-power available is 276. 
 
 VELOCITY, FEET PER SECOND. 
 
 \ ELOCITY, FEET PER SECOND. 
 
 FIG. 74. Flow of Water in Flumes and Ditches, according lo Kutter's Formula. 
 
 4. Let it be required to determine the flow in a trapezoidal ditch whose 
 area is 72 square feet and whose wetted perimeter is 24. The H.M.R. is there- 
 fore three. If the fall of the ditch is i foot in 1500 feet, the velocity of the
 
 194 MANUAL OF MINING. 
 
 flow is 3.5 feet and the discharge is (72X3.5=) 252 cubic feet per second. 
 What should be the grade of a flume ia'X6' which has a H.M.R. of 3, when 
 the quantity desired is to be 547 cubic feet per second? This corresponds 
 to a velocity of discharge 7.58 feet per second. 
 
 In the event that a ditch under computation has a mean hydraulic radius 
 other than that platted among the curves it may be easily interpolated. Thus 
 one of an h.m.r. of ij on a grade of i foot in 400 feet will be found to have 
 a velocity of about 3.8 feet per second. 
 
 5. A pipe is 500 feet long and 3 inches diameter. What should be the 
 head to produce a discharge of 180 feet per minute? 
 
 Here Q=3, 1 = 500, ^=0.25, and assuming / to be 0.00566, 
 
 h= 0.1007 X 0.00566 X 500 X 9 X 1024= 2624 feet. 
 
 H=h + 7^ = 2624 + =3124 feet. 
 
 6. What diameter should it have to deliver the same quantity of water 
 with a head of 82 feet? 
 
 7. Required the flow of water through a pipe 2000 feet long, 13 inches 
 in diameter, and 200 feet head. For the maximum horse-power we have 
 
 (20o)- tM (i.o8)*- tM T =0.864 cubic feet per second. 
 
 REFERENCES. 
 
 Electric Power in Mines (Economy), Carl Pfanauch, Am. Mfr., May i, 
 1896; Electric-power Appliances in the Mines of Europe, Emile Guarini, 
 Eng. Mag., Sept. 1003 and Aug. 1903. 
 
 Enumeration of Electric Plants in Rocky Mountain Region, I. Hale, Amer. 
 Inst. M. E., 1897; Electric Mining in the Rocky Mountain Region, J. Hale, 
 A. I. M. E., Vol. XXVI, 1071 and 402; Enumeration of Electric Plants in 
 Colorado, Colorado S. & M. Quart., Vol. I, 1892, 56; in Rocky Mountain 
 Region, by I. Hale, A. I. M. E., 1897; Electric Mining Machinery in the 
 British Collieries, Sidney F. Walker, Eng. Mag., Aug. 1900. 
 
 Electrical Development, Practical and Impossible, William Baxter, Eng. 
 Mag., Oct. 1806, 113. 
 
 Electric Power Transmission in Mining, A. Bloemendal, Stahl & Eisen, 
 Nov. 15, 1899; and Long-distance Transmission, R. Hutchinson. 
 
 Electric Transmission in Mining, H. C. Spaulding, Amer. Inst. M. E., 
 XIX, 258; M. B. Holt, Amer. Inst. M. E., XX, 316; F. O. Blackwell, Amer. 
 Inst. M. E., XXIII, 400; Electric Transmission Lecture, Colliery Manager, 
 1894, 35; Siemens, Trans. M. & M. Eng., XLIV, Part II, 205; Tests of 
 Condensing Engine and Generators, Trans. M. & M. Inst., XLIV, Part IL,. 
 207; Electric Problems in Mining, Coll. Eng., Dec. 1856, 217.
 
 ELECTRIC GENERATION AND WATER-POWER. 195 
 
 The Dangers of Electricity in Mining, Gluckauf, Jan. 30, 1904. 
 
 Electric Motors, Lecture, Colliery Manager, 1894, 34; Electricity VS. 
 Compressed Air in Colorado, L. Searing, Elec. Eng., Nov. 1896, 528. 
 
 The Construction of Small Water Motors, Elec. Lond., Mar. 9, 1904; 
 Water Wheels, General, Knight Catalogue, Tables of H.P.; Min. Bureau 
 of California, i3th Report: Regulators, Water Wheels. 
 
 The Testing Water Wheels, Elect. Power & Gas Jour., Vol. XII, 32; 
 The Tangential Water Wheel, W. A. Noble, A. I. M. E., Vol. XXIX, 852. 
 
 Electric Power in European Collieries, Cass. Mag., July 1904, 243; Long- 
 distance Transmission, Cass. Mag., July 1904, 112; Distribution of Elec- 
 tricity, Cass. Mag., July 1904, 180. 
 
 Flow of Water in Pipes, Proc. Am. Soc. C. E., XXIX, 821.
 
 CHAPTER VH. 
 
 HOISTING MACHINERY AND UNDERGROUND CONVEYANCES. 
 
 Underground Conveyances. Where possible, the mineral is 
 loaded at the face into cars or buckets to avoid frequent handling. 
 But in metal-mines and in steep-pitching coal- veins this is not 
 possible, and the product is delivered through shutes, or bat- 
 teries, to the cars at the lower level. 
 
 Where the output is small and the depth of shaft is slight, 
 metal-mines employ buckets which are frequently drawn through 
 the haulage-way on the trucks, and thence hoisted to the surface 
 to be dumped. All coal-mines and the larger metal-mines em- 
 ploy cars as the common carriers for the mineral. They are 
 loaded as stated and drawn through the haulage- ways to the foot 
 of the shaft, or slope, through which they are hoisted to day- 
 light. 
 
 Occasionally when the pitch of the slope exceeds 40 the cars 
 are emptied at the level landing into skips, which are hoisted to 
 the surface, while the car returns to its loading station. Slope 
 carriages are also used, upon which the car is delivered and 
 drawn to the surface to be emptied. In vertical shafts the car 
 is hoisted to the surface on a cage, and returned underground 
 when emptied. 
 
 The several conveyances named will be here discussed together 
 with the hoist-rope, signalling and safety devices, and the method 
 of dumping at the surface. 
 
 Hoist-ropes. Hemp, aloe, fibre, and iron or steel wire are 
 used for hoisting purposes and for the transmission of power in 
 
 196
 
 HOISTING MACHINERY, ETC. 197 
 
 mines. They are wound into strands and ropes to furnish a 
 strong, flexible, light means of haulage. The hempen ropes for 
 hoisting purposes have been supplanted by the flexible steel wire 
 for many reasons, prominent among them being the great weight 
 of the former compared with its strength and the unreliability of 
 its fibres. So frequently are the hempen ropes built up of strands 
 of discarded ropes that long service is never obtained from 
 them. This fact, together with their absorption of moisture 
 while in use, contributes to their rapid decay. 
 
 Steel-wire Ropes. The successful production of fine iron or 
 steel wire and its conversion into flexible ropes for hoisting has 
 led to the abandonment of various vegetable fibres. The usual 
 number of strands in a multi-wire rope is six, with nineteen 
 wires in each strand, or sometimes only seven wires are used 
 to the strand. The latter are more frequently employed for 
 guide-ropes than for motor purposes, being less pliable than 
 the former. 
 
 The increased strength imparted to metallic wires by draw- 
 ing, and the high grade of steel wire as compared with iron, at 
 once recommend it for any purpose where strength and light- 
 ness are requisites. Good crucible-steel wire has an ultimate 
 breaking strength of 75 to 100 tons per square inch. Tem- 
 pered steel, usually known by the indefinite name of plow- 
 steel, is a high-grade cast steel, used for ropes of the greatest 
 tensile strength. 
 
 Standard Ropes. Hoisting-ropes in the United States are 
 made with hemp centres, around which are twisted six strands of 
 nineteen wires each, the individual wires being about one-fifteenth 
 the diameter of the rope. The hemp centre imparts flexibility 
 and furnishes a soft yielding core upon which, with a minimum 
 of injury or friction, the strands may "creep" as the rope passes 
 around the drum and sheave. When great flexibility is re- 
 quired, each strand has also its own centre, but this is unneces- 
 sary for ordinary hoisting service. On account of the hardness 
 and diminished flexibility of plow-steel wire, hoisting-ropes of 
 this metal are often made with eight instead of six strands, the
 
 198 MANUAL OF MINING. 
 
 wires being of smaller gauge and therefore more flexible. Though 
 the enclosing figure of the hoisting- rope is a polygon, the rope is 
 usually called a round. 
 
 The Lay of the Rope. By this is meant the twist or pitch of 
 the wires in the strands, and of the strands in the rope. As 
 ordinarily constructed, the lay of the wires is opposite to the lay 
 of the strands. The former is as long as possible in order that 
 the fractional wear on each rope shall be distributed over a great 
 length rather than to have a short twisted strand on which the 
 wear is concentrated. The so-called Langlay ropes have a 
 longer twist to each wire, thus distributing the wear over a greater 
 'ength and giving a smoother surface. 
 
 Locked-coil Ropes. In order to increase the smoothness of 
 the rope and thus to reduce wear, ropes are frequently made of 
 wires which are locked together or dovetailed. The locked- 
 coiled rope is of this type, and prevents the ends of the broken 
 wires from protruding. Its wires are built of a specially shaped 
 cross-section and wound in concentric layers, the successive 
 layers being twisted in opposite directions. The exterior layer 
 presents a perfectly smooth surface. 
 
 Flat Ropes are employed with winding reels where it is 
 desirable to aid the engine without the use of a long conical drum. 
 Moreover, the tendency to uncoil which exists in the several 
 strands and wires of round rope develops a whirling tendency 
 upon the cage as it travels in the shaft. This the flat rope prevents. 
 Flat ropes are heavier than a round rope of the same strength, 
 but are of shorter life, costing more and requiring greater care. 
 A flat rope is composed of an even number of loosely twisted four- 
 stranded round ropes, without hemp centre, laid side by side and 
 sewed through the wires. The lay of any two adjacent strands is 
 in opposite directions. The sewing wires, from eight to twelve in 
 number, are passed from side to side through the centres of the 
 strands. Flat ropes range in width and thickness from 2^'Xf" 
 to 9"XJ". 
 
 Although the breaking strength of a flat rope is nearly the 
 same as that of a round rope of equal weight per foot, its safe
 
 HOISTING MACHINERY, ETC. 1 99 
 
 working load is considerably less, due to the practical impossibility 
 of so constructing the rope that all the strands will stretch uni- 
 formly and be equally loaded in service. 
 
 Taper Ropes. As the depth of shafts increases, the use of 
 round wire rope of uniform cross-section is attended with many 
 difficulties. The diameter of the rope becomes greater, the in- 
 equality of the work upon the engine is more marked, and an 
 excessively large of drum will be required. There is a limit beyond 
 which a rope of uniform section cannot safely carry its own weight. 
 At 12,000 feet of depth, exclusive of any live load, the rope has 
 reached the limit of its safe working stress. The additional weight 
 of car and load reduces this materially. With the margin provided 
 for safety, the maximum depth is 3000 feet. A given rope for 
 carrying men having a factor of safety of 15 when less than 1000 
 feet deep has a factor of but 10 at 2500 feet and only 7 at 3000 
 feet. 
 
 Meanwhile the cross-section of the rope at the bottom is exces- 
 sively heavy, having little load to support. If the cross-section 
 be such as will support the live load, cage, and car, then the cross- 
 sections of the rope to the top may increase in area by an amount 
 equal to that sufficient to support the additional weight of rope 
 below that point. Hence tapering ropes are employed in special 
 cases, being of uniform strength, with such a factor of safety as may 
 be desired. Theoretically, the use of a tapering rope increases 
 the limit of depth which can be reached within mining possibili- 
 ties, but practically it is not a popular hoister. The round tapering 
 rope is made by discontinuing at intervals, throughout the length, 
 a single wire at a time. The cross-section of the rope is uni- 
 formly decreased toward the lower end, the thickness at any 
 point being sufficient to carry safely the load at that point. 
 One round taper rope is of steel of five sections, varying in 
 diameter from i| to ^ inch. The sections are spliced to- 
 gether and the ends of the wires soft-soldered. In some Eu- 
 ropean districts flat taper ropee of Manila fibre are still employed 
 for deep shafts. 
 
 The taper rope is not well adapted to long slopes, for its lower
 
 200 
 
 MANUAL OF MINING. 
 
 portion, which is the smaller end, traverses a greater distance than 
 any other portion of the rope and is subjected to the greatest amount 
 of abrasion. This is the reverse of the situation in a shaft, where 
 the thickest part of the rope has the wear. 
 
 The Strength of a Wire Rope. This is greater than that of a 
 steel rod of equal cross-section and same material, because of the 
 increased strength imparted to wires during the process of draw- 
 ing; but it is not equal to the sum of the strengths of the indi- 
 vidual wires. 
 
 The approximate weight of a wire rope in pounds per foot is 
 ascertained by multiplying the square of its diameter in inches 
 by 1.58 
 
 CAST -STEEL HOISTING-ROPES: 
 
 THEIR ULTIMATE STRENGTH, MAXIMUM SAFE STRENGTH, AND WORKING LOAD. 
 
 
 
 
 
 6 Strands, 
 
 6 Strands, 
 
 Diameter 
 
 Approxi- 
 mate Cir- 
 
 Esti- 
 mated 
 
 Proper 
 
 79 Wires each. 
 
 1 9 Wires each. 
 
 in 
 Inches. 
 
 cumfer- 
 ence in 
 
 Weight 
 per Foot 
 
 Working 
 Load, L. 
 
 Ultimate 
 
 Maximum 
 
 Ultimate 
 
 Maximum 
 
 
 Inches. 
 
 in Lbs. 
 
 
 Strength, 
 
 Safe 
 
 Strength, 
 
 Safe 
 
 
 
 
 
 Lbs. 
 
 Stress, 5. 
 
 Lbs. 
 
 Stress, S. 
 
 if 
 
 4* 
 
 3.00 
 
 
 116,000 
 
 38,667 
 
 124,000 
 
 41,333 
 
 J-i 
 
 4 
 
 2-45 
 
 I 
 
 96,000 
 
 3,2,000 
 
 100,000 
 
 33,333 
 
 ij 
 
 
 2.00 
 
 """! co 
 
 80,000 
 
 26,666 
 
 84,000 
 
 28,000 
 
 J 
 
 3 
 
 1.58 
 
 %% 
 
 64,000 
 
 2I >333 
 
 68,000 
 
 22,000 
 
 i 
 
 $ 
 
 2 
 if 
 
 I. 2O 
 0.89 
 O.62 
 
 o.:;o 
 
 0.39 
 
 um safe stn 
 bending sti 
 
 48,000 
 37,200 
 26,400 
 
 21,200 
 16,800 
 
 16,000 
 12,400 
 8,800 
 7,067 
 5,600 
 
 5?,CXDO 
 
 38,800 
 27,200 
 
 22,000 
 I7,600 
 
 17,333; 
 12,933 
 
 7.335 
 
 5,867 
 
 $6 
 
 1^ 
 
 0.30 
 
 J 
 
 13,200 
 
 4,400 
 
 I3,600 
 
 4,533 
 
 
 
 j i 
 
 0.22 
 
 a 
 
 9,6OO 
 
 3,200 
 
 IO,OOO 
 
 3,333 
 
 % 
 
 ! 
 
 0.15 
 
 
 6,800 
 
 2,267 
 
 6,800 
 
 2,267 
 
 * 
 
 i 
 
 O. IO 
 
 
 5,600 
 
 1,867 
 
 4,800 
 
 1,600 
 
 The Minimum Radius of Curvature. The minimum drum 
 diameter admissible for ordinary service is from 50 to TOO times 
 the diameter of the rope. By using the larger ratio the average 
 life of hoisting-ropes would be materially lengthened. The 
 English make the sheave diameter a multiple of X 2 . The length 
 of the arc of contact between rope and sheave and the angle 
 between the two branches of rope are not the same as would be
 
 HOISTING MACHINERY, ETC. 
 
 the case with a belt or Manila rope. A large part of the total 
 strength of the rope is consumed in the mere strain of bending, 
 and this is wholly a question of radius of curvature depending 
 upon the diameter of the sheave. 
 
 This bending resistance increases the tension on the wires 
 of the ropes and reduces the available load which it can safely 
 carry by that amount. The continual bending naturally reduces 
 the life of the rope. 
 
 In the following table are given the values of the radii of 
 curvature, R, in inches, assumed by new steel ropes of stated 
 diameters having a tension of one pound upon them and having 
 an angle of deflection or of bending of i. 
 
 THE RADIUS OF CURVATURE R OF STEEL-WIRE ROPE, IN INCHES, HAVING 
 i DEFLECTION AND i LB. TENSION. 
 
 Diameter. 
 
 7-wire. 
 
 iQ-wire. 
 
 1 
 
 257,270 
 
 84,210 
 
 ^ 
 
 592,97 
 
 220,950 
 
 I 
 
 1,241,510 
 
 454,120 
 
 I 
 
 2,315,030 
 
 860,400 
 
 I 
 
 3,97^290 
 
 1,430,850 
 
 4 
 
 6,386,740 
 
 2,246,676 
 
 i| 
 
 10,049,650 
 
 3,369,820 
 
 To obta'n the radius of curvature under similar conditions 
 for any other angle of deflection, it is only necessary to divide the 
 corresponding quantity by the number of degrees of bend. Thus 
 a new i-inch ip-wire rope, for a tension of one pound upon 
 it, would have a curvature whose radius is 1,430,850 inches 
 when bent through an angle of i. If, however, it is bent 90 
 over a given sheave, the radius of the curve which it will assume 
 then becomes 15,898 inches. When the tension becomes 1000 
 Ibs. the curvature which it will assume will have a radius of 
 16 inches. 
 
 Bending Rope Around Curves. When a rope is carried over 
 a sheave with an angle of contact or wrap exceeding 30, the 
 diameter of the sheave must be of liberal dimensions When, 
 however, the rope is slightly deflected by a pulley or sheave,
 
 202 MANUAL OF MINING. 
 
 the stress put upon the rope in bending is small. When it is 
 bent over a succession of sheaves at close intervals, as in carry- 
 ing a rope around curves in a roadway, the aggregate bending 
 stress may become significant. The rope assumes between each 
 pair of sheaves a curvature depending upon the tension it carries 
 and the amount of each deflection. 
 
 FlG. 75. Turning the Haulage-rope at Curves. 
 
 The distance between the sheaves may be determined from 
 the formulae below when the tension is known and the radius of 
 curvature has been ascertained by the preceding table. 
 
 Let R = radius of curvature i bend with i Ib. tension for rope 
 
 of diameter X (from the preceding table); 
 r= radius of curvature assumed by rope under tension T\ 
 d= diameter of individual wires in rope, inches; 
 w= number of wires in rope; 
 T= tension on the rope, Ibs.; 
 s= distance between sheave centres, inches; 
 a = angle of deflection over a pulley, and 
 p= radius of centre line of sheaves, inches. 
 Then 
 
 The sheaves should be laid on a curve of radius r that
 
 HOISTING MACHINERY, ETC. 203 
 
 is, r=p't d=o.inX for a y-wire rope and 0.066 for a 
 rope. 
 
 R 
 
 s =2p sin %a. 
 
 The Influence of Bends on the Durability of a Rope. The 
 
 diameter of the sheave over which the rope is bent must in all 
 cases exceed the minimum diameter of curvature assumed by 
 the rope under the conditions as calculated above. Not only 
 is the life of the rope increased, but also its bending stress is con- 
 siderably diminished and the safe working load of the rope 
 correspondingly increased. Under ordinary conditions of wear, 
 the life of a rope for mining purposes averages about seventeen 
 months, after which time the abrasion of some wires and the 
 breaking of others render it unsafe. This assumes proper care 
 in lubrication, in testing, care during hoisting to avert shocks, 
 and the use of liberal-sized sheaves. Tests upon wires of various 
 sizes passing over pulleys of various dimensions have demon- 
 strated that a wire subjected to repeated bending soon breaks, 
 the time of rupture being hastened as the diameter of the sheave 
 or the angle of the bend is small. For example, a wire of No. 20 
 B.W.G., subjected to repeated bending over a 5 -inch pulley, 
 broke after making i5j2oo turns; one of the same size and pre- 
 sumable strength made 453,000 turns over a pulley of 24 inches 
 in diameter before breaking. 
 
 EXAMPLE. It is desired to carry a ii-inch ip-wire steel rope around a curve 
 (Fig. 75) whose track radius is 50 feet, by guide-sheaves 44 inches inside 
 of the track centre. What should be the distance between the sheaves? 
 
 T=66o, X=i.25, r= 1,430,900 inches, and p= 554 inches. Then 
 a =3 oo' and 5=2 feet 5 inches. 
 
 The Working Load. This is a fraction of the ultimate strength 
 of the rope, depending upon the margin of safety desired. For 
 ordinary conditions of hoisting the factor of safety is taken at 7 for 
 hoisting-ropes and 6 for those employed on slopes. The maximum 
 stress to which the rope can be subjected should be less than
 
 204 
 
 MANUAL OF MINING. 
 
 PLIABLE HOISTING WIRE ROPE 
 
 6 STRANDS- 19 WIRES EACrf-CAST STEEL 
 
 DENVER ENGINEERING WORKS 
 
 DECEMBER 1, 1903. 
 
 NT* NT i*i 100 
 
 FIG. 76. The Net Working Load of Ropes over Sheaves.
 
 HOISTING MACHINERY, ETC. 205 
 
 one-seventh or one-sixth, respectively, of the ultimate strength 
 of the rope. A larger factor of safety would unnecessarily in- 
 crease the size of the rope, which in turn increases the dead load, 
 due to its own weight and the diameter of the sheaves over which 
 it must pass. 
 
 The stresses to which a rope is subjected, in addition to the 
 live load to be carried, are those due to its own weight and friction, 
 its bending over the sheave, its rigidity, and the shock induced 
 at the time of starting. Some of these can be calculated, but 
 the tension induced by the rigidity of the rope is not so easily 
 ascertained. An initial strain is set up in the wires by the bend- 
 ing around the drum or sheave which is not equally distributed 
 throughout the rope. Its value depends upon the sharpness of 
 the bend and is ascertained by the formula given below. A 
 small sheave and a large angle of deflection produce a large 
 bending stress. 
 
 The Resistance Due to Bending. This must be determined 
 before the available load which the rope can carry can be ascer- 
 tained. Usually the live load represents the difference between the 
 safe or minimum allowable tensile stress and the stress due to 
 bending. In deep shafts regard must also be taken for the 
 weight of the rope, which is to be subtracted from the working 
 load. 
 
 The bending stresses, k, which are developed in a rope when 
 bending over a sheave or hoisting-drum through an angle ex- 
 ceeding 30 are indicated below. 
 
 Let X = the diameter of the steel rope in inches; 
 r=the radius of the sheave or drum in inches; 
 /=the thickness of the flat wire rope in inches; 
 k = the bending stress in pounds. 
 
 Then for a y-wire rope 
 
 r+o. 3 X ' 
 for a ip-wire rope 
 
 36 7 ,cooX 3 
 r+o.$X '.
 
 206 MANUAL OF MINING. 
 
 for a flat wire rope 
 
 89,465* 
 r+2.23/' 
 
 The diagram in Fig. 76 graphically shows the net working 
 load which a given rope of pliable steel can carry over a pulley of 
 a given diameter. This diagram is based on both formulEe, and 
 is taken from a treatise of the Trenton Iron Co. on Wire-rope 
 Transmission. As an illustration of its use let it be required to 
 determine the net load for a i-inch rope on a ico-inch sheave. At 
 the intersection of the vertical line from the point on the top line 
 marked 100 inches to the curve representing the i-inch rope, the 
 horizontal line carried to the right-hand edge shows 15,500 Ibs. 
 as the net load. It will be noticed that a 2-inch rope on a 76- 
 inch sheave has a net working load not greater than that of a 
 f-inch rope on the same sheave. This diagram, therefore, indi- 
 cates the reduction in the capacity of the rope due to bending 
 stress. 
 
 Elastic Connections between Rope and Cage. Usually while 
 lowering to a platform an excess of rope is paid out and tends to 
 coil upon the top of the cage. This operates disastrously upon 
 the wires of the rope, which become bent and ultimately 
 broken. A flexible connection, therefore, is provided of 
 a length sufficient to allow for the probable slack. This 
 consists of a short length of chain with a snap- hook (Fig. 
 77), which can be released when desired. Owing to the 
 shock which these links receive, they must be carefully 
 examined at short intervals. In some mines it is the 
 practice also to anneal the iron by heating to a red heat 
 and cooling slowly. This is done once a month. 
 
 These connections should be as short as practicable 
 in order to diminish the amount of shock induced at 
 starting. Carelessness in paying out excessive slack 
 ' reduces the life of the rope and increases the stress at 
 starting. 
 
 Results of observations in Prussian mining districts during
 
 HOISTING MACHINERY, ETC 207 
 
 the past twenty years showed the effect due to repeated shocks 
 at starting to be more disastrous to the flat than to the round 
 steel ropes. In certain districts 6 per cent of flat steel ropes 
 broke suddenly out of 1419 in use; 12 per cent of flat charcoal- 
 iron ropes gave way suddenly; 7 per cent out of 124 flat Manila 
 ropes broke without previous warning; 2 per cent of 5527 round 
 steel ropes gave way, and 10 per cent of 1350 round charcoal 
 ropes failed without giving previous warning. 
 
 The tension due to jerks is twice that due to the dead load 
 with 6 inches of slack, and more if the slack is greater. 
 Let T=the total tension at starting; 
 
 a= acceleration at beginning of hoist; 
 
 h= slack in inches; 
 
 b = elongation of rope due to total load, W. 
 
 W= weight of car, cage, mineral, and rope. 
 
 Rope Sockets. There are two types of sockets for round ropes 
 the conical and the double-pin, the former being the stronger. 
 The conical socket (Fig. 78) is slipped on to the rope, the wires 
 are untwisted, hemp centres cut out, the wires bent back and 
 forth into a tangled mat to fill, as nearly as possible, the conical 
 socket, which is then slipped into place. This is slightly 
 heated, and soft lead poured in to solidify the mass. 
 The socket and rope are surrounded with wet clay to 
 prevent heating of the wires beyond. The double-pin 
 is treated in the same manner, but its connection with 
 the chain is by a pair of pins through the links, instead 
 of a ring for hooking, as in the former case. A "goose- 
 neck" socket consists of a pair of trough- shaped tongs, 
 bent to a loop, and riveted to the rope by three or 
 four rivets, driven cold. Flat ropes have riveted to 
 them shackles with eyes, which receive the first link g 
 of the chain. Six inches of the end are untwisted and 
 doubled back, bound with wire, the shackles slipped on, riveted
 
 208 
 
 MANUAL OF MINING. 
 
 ?. ?. 8
 
 HOISTING MACHINERY, ETC. 209 
 
 through the rope, and the hoops finally slipped on and driven 
 tight. 
 
 In securing the rope on the drum it is only necessary to con- 
 tinue several extra coils of the rope, insert the end through the 
 wooden lagging, and fasten it on the hub or shaft; or, instead, 
 the end may be bolted to me arm of the casting. If the fastening 
 have but a lo-lb. grip on the rope, it will resist a weight of 90 
 Ibs. if there is only one coil around the drum; if there are two 
 extra coils, 800 Ibs. will not budge the zo-lb. grip; with three 
 extra coils it requires 7300 Ibs.; while with four it has a 65,000- 
 Ib. resistance and can support that amount of tension. 
 
 Wire rope is spliced in the same manner as hemp. The 
 strands are unlaid for 3 feet, and each passed over one and under 
 another of its corresponding strands on the opposite rope, for 
 a like distance; the free ends are then trimmed off close. Short- 
 twisted rope is more easily spliced than one of long twist. 
 
 The Head-frame is designed to withstand the stresses from 
 the weight of the dead and live loads on the rope and their fric- 
 tional resistance operating vertically, and from a pull of the engine 
 along the inclined rope. The latter .exceeds the former by an 
 amount equal to the frictional resistance at the sheave and its 
 bending stress. The frictional resistance may be taken as one per 
 cent of the gross load on the journals, and the bending stresses 
 of the rope may be determined from the preceding formulae. 
 The combination of these two forces produces a resultant operating 
 in a direction which nearly bisects the angle between the two 
 ropes. From these, with the load, may be obtained the direction 
 of the resultant. Its line should fall nearly central within the 
 faase of the head-frame, when the latter is of wood. It is not so 
 essential if the derrick is of steel. 
 
 The frame over the shaft is built in one of two patterns. The 
 form of the derrick is essentially two vertical right- angle triangles, 
 each upright and brace being respectively parallel to the two 
 directions of the rope. The apex of the triangle should be at the 
 centre of the axis of the sheave. The triangular frames are 
 connected in a transverse direction by braces and ties. Instead
 
 2io MANUAL OF MINING. 
 
 of single sticks for the uprights, a vertical frame is built of four 
 uprights surrounding the shaft- opening with two or three struts 
 projecting from their top toward the hoisting-engine and inclined 
 to furnish ample base for the derrick. 
 
 In the other pattern the four posts are batired outward to 
 afford an ample square base of support without the use of any 
 additional struts (Fig. 80). The frames may be built of wood or 
 steel, the latter being extensively employed in collieries where 
 the speed of hoist is great (Fig. 81). The steel frames are 
 built by construction companies to order and are composed 
 of the usual structural shapes in steel. 
 
 The foundation may be either masonry pillars or concrete, 
 extending some distance outside of the lines of the shaft. The 
 lower base consists of three longitudinal sills framed with three 
 cross-sills. Bolted to the former are cast-iron shoes to receive 
 the uprights and braces. The base-frame is bolted and anchored 
 to heavy timbers buried in the ground or in the concrete. 
 
 The Height of the Derrick. The essential requirements of the 
 head-frame are stability and safety. The former is obtained by 
 a broad base and a rigid construction of frame; the latter is 
 obtained by its construction of fire-proof metal and a height 
 sufficient to satisfy the conditions of hoisting. The risk of fire, 
 the effect of weather, the working of the joints, and the difficulty 
 of securing sound sticks of requisite length render the adoption of 
 material other than wood advisable. Height, which is an essen- 
 tial feature of the head-frame, is difficult to attain without a 
 rigid frame, and this implies steel construction. This is the 
 rule at large mines. Steel frames are more desirable and immeas- 
 urably more rigid than wooden ones. To withstand injury from 
 the continual vibration, pin connections are more suitable than 
 riveted connections. 
 
 The height of the derrick should be such as will furnish 
 security against overwinding by providing sufficient margin within 
 which the engineer may stop the hoist. The greater the speed 
 of the engine the greater is the necessity for a high derrick. The 
 minimum allowance is that which corresponds to the length of
 
 HOISTING MACHINERY, ETC. 211 
 
 two coils of rope on the drum plus the height of the cage. The 
 distance from the landing station for cars to the sheave would 
 give the engineer four or five seconds in which to secure control 
 of his engine. Usually the derrick is of a height twice the amount 
 stated. In anthracite regions it is often more than this. 
 
 Steel frames in some instances are 65 feet in height, the base 
 of the first-mentioned type being about 16 feet square with the 
 sides extended for the braces. Of the second type there are 
 numerous steel frames with base of 25 to 30 feet square, with a 
 height of 50 feet. Derricks are frequently extended to' a height 
 above that of the tipple or breaker platform to permit of a con- 
 tinual descent of the mineral from the time of unloading to the time 
 of shipment. In such cases as these the head-frame is over 100 
 feet in height. Those in the zinc-mining districts of Missouri 
 .are of the four-post type, 30 feet high, with suitable braces. Some 
 in the anthracite district are as high as 130 feet. 
 
 The hoist-frame is rarely enclosed, and preferably is not 
 housed because of the danger from fire. A fire breaking out in 
 the building over a shaft would be uncontrollable with a draft 
 from below, besides risking the communication of the fire to 
 timbers of the shaft. A fire communicated to the hoist-frame 
 from below would endanger the only means of exit which the 
 men have from underground. 
 
 The tipple type of head-frame is built of bents, suitable for the 
 screens and pockets. It connects with and extends from the 
 head-frame in a direction opposite to that of the back-stays. It is 
 of wood or steel with all members carefully calculated. The 
 rhythmic motion set up by the vibrating screens must be pro- 
 vided for, and the design varies radically between one having a 
 dump in the plane of the hoist-rope and one in which the shocks 
 of dumping are at right angles to that plane. 
 
 Slopes require no high head-frame, for the line of track extends 
 directly from the shaft mouth to the elevated structure, and in 
 reality forms the back- stay for resisting the overturning tendency 
 of the structure. This is the strongest possible type of frame, 
 for the resultant falls inside of the base of the tipple.
 
 212 
 
 MANUAL OF MINING.
 
 HOISTING MACHINERY, ETC. 
 
 213 
 
 The Sheave. Upon the top of the uprights is a frame which 
 supports the sheave on a short horizontal axle. It may be built 
 directly upon the braces and uprights of the head-frame or upon 
 the platform above it. The dimensions of the sheave are deter- 
 mined by the size of the rope. It consists essentially of a spider 
 frame carrying a grooved circumference filled with some material 
 
 FIG. 81. A Steel Head-frame at a Bituminous Colliery. 
 
 which has a large coefficient of friction, as, for instance, wooden 
 blocks on the end, or rubber. These latter prevent slip and 
 render the hoisting more secure, as well as protect the rope from 
 excessive wear. Its hubs are double, connected to the rim by 
 rods let into sockets. The entire construction is made as light as 
 possible consistent with strength, to reduce the inertia and avoid 
 the abrasion that occurs with a heavy wheel continuing to revolve 
 after hoisting has ceased.
 
 a 14 MANUAL OF MINING. 
 
 The cost of the high steel head- frames of a mine averages 
 $32.20 per ton of material, including the design and details. The 
 erection averages $13.00 per ton when the connections are riveted. 
 
 As regards the cost of wooden head-frames, the following may 
 afford a suggestion for average conditions. The four- post frame 
 for small mines of 31 feet requires 1800 feet of board measure 
 and costs from $100 to $150 for labor and supplies. Larger 
 frames without back- stays for the height of 50 feet consume 
 14,000 feet of lumber and cost $550. Tipple head-frames for 
 vertical shafts, with 60 feet of height, require 15,000 feet of board 
 measure and cost about $800. Slope tipples for inclined shafts, 
 built to a height of 31 feet, are completed for $3300, using 40,000 
 feet of board measure. 
 
 Cage Indicators. Several independent devices are used to 
 indicate the position of the cage in the shaft at all times during 
 its hoist. Automatic indicators are installed in view of the 
 hoisting engineer. An indicator has a pointer which moves 
 around a dial or vertically along the side of a column, the dial 
 or the column being graduated to the depths of the shaft- or 
 slope-landings. A string winding on a miniature counterpart 
 of the hoisting-drum moves the index a distance commensurate 
 with that of the cage and shows the latter's position after starting 
 and its proximity to the landing station during the hoist. The 
 approach of the cage is also indicated by an electric communi- 
 cation automatically established by the cage in its ascent. In 
 this arrangement the cage makes an electric connection as it 
 passes a given point in the shaft below the point of landing, 
 which is signalled to the engineer. A second notice is given in 
 a similar manner when the cage is just about to reach the land- 
 ing. This notification is independent of the automatic devices 
 for the guidance of the engineer in braking the engine. In addi- 
 tion to these means, the brakeman or landing-man at the head 
 of the shaft having charge of the cars may. signal the engineer 
 as the cage approaches the top and furnish another control, 
 which, being sentient, is always to be preferred to any mechanical 
 device, no matter how perfect.
 
 HOISTING MACHINERY, ETC. 215 
 
 
 
 Signalling. The communication between the employees below 
 and the engineer above may be had by the use of a wire with a 
 bell or some electric annunciator, or by a speaking-tube or tele- 
 phone. Whatever may be the means employed, it is essential 
 that it should be quick and accurate, presenting no opportunity 
 for mistakes. The early clumsy arrangements depending upon 
 the signal of the gong, bell, or triangle which is struck by a lever, 
 operated from below by a rope or wire, are untrustworthy. The 
 pull on the rope may be too light to strike the blow, or more strokes 
 may be rung up than are intended, and there is no means of making 
 a correction before the engineer will have proceeded to answer 
 the call. Accident may then ensue. The annunciator is ano'her 
 method which is quick and convenient for signalling to the engineer 
 or between stations, but its scope is limited to a very small set 
 of signals. Unquestionably electricity offers the best means 
 of signalling between stations in dry shafts. One wire or a couple 
 of wires may be placed within easy reach of those in the cage, 
 furnishing a means for its occupants to signal to the surface 
 or elsewhere while in motion. Such wires would also be 
 convenient for those engaged in examining or repairing the 
 shaft, being connected to electric gongs located at each landing- 
 station. 
 
 A uniform code of signals among mining men is an eminently 
 desirable feature. A hoisting engineer at a new mine may err 
 with serious results by an interpretation of a signal which else- 
 where represents some other operation. 
 
 The Telephone. There need be no argument to convince 
 the reader of the advantages of the speaking-tube or telephone. 
 This safe natural means is largely employed. It is rapidly sup- 
 planting the electro-mechanical gongs. The expense of laying 
 out a mine telephone system is generally much less than that 
 of a surface system, since in most cases the wires can be sup- 
 ported by means of porcelain or glass insulators, either from 
 the roof or from any timber which is used on the sides of the 
 road. Their greater expense is overbalanced by the great 
 saving in time, and their efficiency in case of accident. For
 
 216 MANUAL OF MINING. 
 
 deep shafts, telephones are almost imperative; in fact, mining 
 legislation in some States compels their use. 
 
 Bridging telephones, Fig. 82, connected to the metallic sig- 
 nal-line, used in connection with rope haulage, reduce the cost 
 of installation to that of first cost of the telephones and the nomi- 
 nal cost of connecting up. For insulation, rubber-covered wire 
 
 FIG. 82. A Mine Telephone. 
 
 will serve generally in localities where moisture or contact with 
 timber might cause an interruption of service. When an in- 
 stallation of mine telephones will warrant the extra cost, lead- 
 covered cables should be employed. 
 
 Means of Increasing the Safety of Ascent and Descent into 
 the Mines. The necessity for the introduction of safety appliances 
 of this kind needs no argument when one considers that the engi- 
 neer, from a point somewhat remote from the elevator, is raising 
 the latter, at an hourly rate of speed reaching 30 miles, and the 
 load, whose weight may be nearly one fourth of the strength of 
 the rope. This is to be landed promptly, accurately, and without 
 shock at the precise point. In this work, which is repeated scores 
 of times each hour and with an engine of several hundred horse- 
 power, with no guide except the index moving about a small circle, 
 any defect in the operation of the indicator, or any momentary 
 distraction of the engineer's attention, may result disastrously 
 and the engine raise the cage a few feet beyond the point in- 
 tended, and indeed may raise it over the sheave at the top of 
 the frame. Again, some defect in the operation of the throttle 
 or of the brake may prevent the engineer from bringing the engine
 
 HOISTING MACHINERY, ETC. 217 
 
 to rest as promptly as desired, and one revolution or two, which 
 may wind 30 to 50 feet of rope, may bring the cage to the danger 
 limit. Recalling that these operations are performed 50,000 
 times per year, the mental strain and the anxiety of the engineer 
 where no overwinding device is provided is unquestionably 
 serious. Again, coal and men are hauled on the same vehicle; 
 and if no proper method of signalling exists, the engineer may 
 hoist the latter at the speed and to the place of dumping the 
 former. 
 
 Though the number of accidents arising from overwinding 
 is small, nevertheless it is sufficient in amount to warrant the 
 introduction of some device. There are two general classes of 
 these. The first type is an ample detaching-hook, and the second 
 an automatic control for the engine. 
 
 Detaching-hooks. Safety contrivances are installed at the 
 head-frame for preventing a cage from falling when severed 
 from the rope through overwinding. Their design contem- 
 plates the release of a link between the rope and its cage, and 
 the simultaneous action of some support which take? hold of a 
 portion of the framework of the headgear. These give effective 
 service and are employed very extensively in Europe, but to a 
 lesser extent in America. 
 
 Ormerod's hook is one of the best of these safety links. The 
 apparatus when in ordinary use, as in 2, Fig. 84, is wider at the 
 bottom than at the top. When overwinding occurs the link is 
 drawn into the bell-mouthed cylinder just below the sheave. 
 This lower part comes into contact with the cylinder, thereby 
 closing the bottom part of he link and expanding the top part. 
 Its projections catch over the top of the cylinder, while at the 
 same time the rope shackle A is forced out of its seat, thus being 
 allowed to go free; the bottom shackle, B, drops into a slot and 
 locks the link firmly in its position. The cage, being suspended 
 from the chain, cannot fall back. To prevent the possibility of 
 the link becoming disarranged in ordinary work, a small pin is 
 inserted centrally through the plates, which pin is sheared off as 
 the apparatus passes into the cylinder.
 
 218 
 
 MANUAL OF MINING. 
 
 Middleton's hook, 3, Fig. 84, operates on the same principle. 
 
 King's hook, one form of detaching-hook, is illustrated in 
 Fig. 83. It consists of two outside plates, enclosing two inner 
 ones, capable of oscillating about a strong pin which passes 
 through the series. When hoisting has reached the danger- 
 
 FiG. 83. King's Detaching-hook. 
 
 point, the wings on the sides of the latter pair strike the plates, 
 the jaws permitting the rope to be hoisted over the sheave and 
 the cage to fall. At the same time, however, the projecting 
 wings on the hook catch on the protruding edges of the platform 
 and hold the cage. 
 
 Walker's safety attaching-hook, 4, Fig. 84, has a loop which 
 encircles the hook and is bound by it to two copper rivets. These 
 are sheaved when the hook is down. The jaws then open and 
 release the rope, locking the suspension jaws on the disengaging- 
 plate and holding the cage. 
 
 Engine Controllers. This class is a more or less flexible con- 
 nection between the shaft and the engine by which the rage
 
 HOISTING MACHINERY, ETC
 
 ^220 MANUAL OF MINING. 
 
 -automatically shuts off the supply of steam or applies the brake 
 when it reaches a certain point. 
 
 The Visor, 5, Fig. 84, consists of a governor suitably driven 
 by a crank-shaft, which, if the speed of the engine exceeds a cer- 
 tain point at the end of the hoist, applies the foot-brake through 
 a combination of levers, and closes the steam- valve. 
 
 Buckets Kibbles. The simplest conveniences for mineral 
 Jioist are employed during sinking and during the mining opera- 
 tions of small mines. These are sheet-steel tubs, or buckets, 
 fitted at the top with a bale of round iron and at the bottom with 
 a small ring. The bale is hooked into eyes opening into a strap 
 which is bolted to the sides and under the bottom of the tubs. 
 The snap-hook (Fig. 85) on the hoisting-rope catches into the 
 ring of the bale. When, however, buckets are to be frequently 
 
 FlG. 85. Snap-hook. FIG. 86. Serpentine Hook. 
 
 cletached from the rope, serpentine hooks (Fig. 86) are used 
 instead. They are plain cylinders of sheet steel or boiler-iron, 
 18 to 33 inches in diameter, 30 to 54 inches deep, having 
 a capacity of 600 to 3000 Ibs., and a weight of about one 
 quarter of their contents. Frequently they are bellied and used 
 In slopes on skids. The buckets are loaded from chutes at the 
 mill-holes in the mine (Fig. 87), and transported on the trucks 
 to the shaft and raised to the surface. 
 
 The dumping of the bucket at the surface may be accom- 
 plished in one of two ways. It may be lowered upon a truck 
 and removed from the hoist-rope and taken to the point of empty-
 
 HOISTING MACHINERY, ETC. 221 
 
 ing, meanwhile being replaced by an empty bucket which is 
 lowered below. On the other hand, the bucket may be emptied 
 at the surface without removal from the rope. In the becket 
 and high under the bucket the shaftman inserts the hook of a 
 short length of stationary chain The tub with its hoist-rope is 
 
 FIG. 87. Loading-gate for Buckets and Cars. 
 
 then lowered and inverted, its contents being emptied upon a 
 grating at the side of the shaft or into a car which is run over 
 the landing- doors which have been lowered across the shaft. In 
 Fig. 89 is another device, not so safe, in which the bale is pivoted 
 a little below the centre of gravity of the bucket and is held in 
 position by a loose ring on the bale, slipping over the pin at the 
 upper rim of the bucket. To dump, the brakeman merely raises 
 the ring and allows the bucket to reverse. It is easily righted 
 and again fastened. 
 
 Aside from objection to hoisting of buckets on account of its 
 low speed and insecurity, there is a more important one in metal- 
 mines, namely, the production of considerable powdered mate- 
 rial. This soft product, though it may reach the surface, prob- 
 ably never reaches the smelter and, whatever its value, is lost.
 
 222 
 
 MANUAL OF MINING. 
 
 As the softer minerals are also the richer, one can readily see 
 the necessity for avoiding too much handling of the mineral. 
 In addition to this is the risk of collisions, unless each tub has 
 
 FIGS. 88 and 89. Mine Buckets. 
 
 its own hoist compartment with smooth-lined guides. These 
 slides permit of more rapid hoisting, but the damage and 
 accidents arising from contact with buckets when the end 
 of a plank protrudes make the risk and cost quite consider- 
 able. 
 
 Shaft-guides. To facilitate rapid hoisting, the shafts are 
 fitted with conductors or guides, along which the conveyances 
 move. With a bucket-hoist on the slopes the floor is planked 
 with 4-inch wooden guides spiked on either side of a runway, 
 between which the bucket may slide in its passage. The ends 
 are finished so that they will not easily loosen and cause 
 accident. . Experiments with steel rails for guides have proved 
 them far too noisy for comfort.
 
 HOISTING MACHINERY, ETC. 
 
 223 
 
 When a bucket is hoisted in a shaft it swings freely, but is 
 guided by two or four ropes stretched from top to bottom; these 
 pass through holes at the ends of two or four arms which make a 
 horizontal frame. This frame rests freely above the shackles of 
 
 FIG. 90. 
 
 the hoist-rope and keeps the bucket in a central position during 
 all times of the hoist. Cages are guided by two 4-inch timber 
 strips which are bolted to the buntons, or cross-pieces, fixed 
 across the pit. These are securely wedged against the rock, or
 
 24 
 
 MANUAL OF MINING. 
 
 are framed into the timbering of the shaft. One guide on each 
 of the two sides is sufficient for a cage compartment. 
 
 Skips, or Gunboats, are commonly seen in the inclined shafts 
 
 ''--.-. I 
 
 FIG. 91. 
 
 FIG. 92. 
 
 for hoisting purposes. They consist of an iron box weighing 
 from 900 to 1500 Ibs. set on four wheels held by bolts riveted
 
 HOISTING MACHINERY, ETC. 22$ 
 
 at the sides; the front view is rectangular, but the side view is 
 trapezoidal (Fig. 90). The inclined end is uppermost with auto- 
 matic dumpers, while in those discharging by door it is below. 
 The hoist-rope is attached to the bale, which rotates on a pin 
 passing through the side of the skip, often back of the centre 
 of gravity, so as to dump automatically. The charge of one 
 or two car-loads is shovelled or chuted into the skip, which emp. 
 ties at the surface into a bin or on grizzly. In one variety the 
 contents are discharged by the mouth at which it is loaded, while 
 the other form has a swinging door opening at the upper side. 
 A vertical safety skip is shown in Figs. 91 and 92. 
 
 The automatic dump is simple; the rear or lower wheels are 
 of wider face than the fore wheels. As the dumping-plat is 
 reached, the guide-rails on which the wheels have been travel- 
 ling gradually bend to horizontality, and these the front wheels 
 follow. As the hoisting continues, the wider rear wheels catch 
 and roll on a pair of outer guides, and continue up the slope. 
 By this means the lower end is elevated and the skip emptied. 
 
 The brake is generally a drag, consisting of a bar about 4 feet 
 long, trailing on the floor, and only catching if the skip breaks 
 away on its up-trip. Often the wheels are confined between two 
 guides on each side, as, for instance, where the direction of the 
 slope changes. The sole objection to these skips is the double 
 and treble handling involved. The car from the mill-hole empties 
 into the chute, whence the skip is loaded, and at the surface 
 the reverse operation takes place. 
 
 The Slope-carriage dispenses with two handlings, taking the 
 car at once to the surface. This is simply a double triangular 
 frame, large enough to accommodate a car with two rails on 
 its horizontal top, and two wheels on each hypothenuse. A 
 hook or lock holds the car while riding. For convenience the 
 loading and unloading gangways are not on the same level, 
 the track for "empties" being 6 feet higher than the "full" 
 track. If the seam is too thin to allow of this, each track has 
 a curved roadway connection with the gangway. 
 
 The head-room required for the slope having this means of
 
 226 
 
 MANUAL OF MINING. 
 
 haulage implies large area and extensive timbering, particularly 
 as a double trackway is necessary if the output is considerable. 
 There is not much necessity for a carriage, except in slopes over 
 60. 
 
 The Cage. The cage is essentially a wrought-iron or steel 
 platform for the cars traversing the pit. It is also the usual 
 
 FIG. 93. A Safety Cage with Keeps. 
 
 means of transport for the workmen and all other materials 
 ^between the surface and the different landing-stages in the shaft. 
 The pit timber, workmen, tools, fodder and water for the horses,
 
 HOISTING MACHINERY ETC. 227 
 
 and frequently the horses themselves, are lowered by means of the 
 cage. Its form is governed by the shape of the compartment of the 
 shaft in which it runs, and its size by the number of cars. Some 
 cages are fitted with two platforms or decks. On each deck are 
 rails for one or two cars, and so placed as to allow of those on the 
 cage being displaced by others to be hoisted or lowered. The 
 platform is held on the sides next to the guides by two stout iron 
 frames united at the top by a cross-bar to which the hoist-rope 
 is attached (Fig. 93) . Two iron ears at the top and bottom on 
 each side confine the cage to the guides. An iron roof or bonnet 
 over the cross-bar shields the passengers from falling rocks. 
 
 Some cages are also provided with extra covers, either fixed 
 or adjustable, above the tops of the cages, to protect the men while 
 examining or repairing the shaft. 
 
 An automatic device controlling the delivery of the cars in 
 and out of the cage is illustrated in n, Fig. 84. It serves for 
 the intermittent delivery of cars without an attendant. 
 
 Securing the Cars on the Cage. There are various modes of 
 keeping the car secure in the cage during its ascent or descent. 
 One of these is a "fabe bottom" in the cage, which sinks an 
 inch or two below the other and outer portions of the deck, but 
 is lifted when the cage is at a landing-station. The deck floor 
 is then on one level, allowing the tub to be changed. 
 
 Another device is a bar running across the cage, and at either 
 end is placed a short lever, which turns down or up on being 
 pushed; when down it covers the ends of the tubs and prevents 
 their moving; when up it allows them to be changed. Some- 
 times a latch-lock on the floor secures the cars in place. 
 
 Multiple-deck Cages. Single-deck cages are almost exclu- 
 sively used in America, and are sufficient except for narrow, deep 
 shafts, when heavier loads are necessary for large output. Then 
 cages with two or more platforms are used. The landing-stages 
 are arranged in the same number of tiers, from and upon which 
 cars are simultaneously run, without moving the cage. With 
 ample facilities for "decking" the cars, the saving in the trip- 
 time, per car, increases the capacity of the plant. The necessarily
 
 228 
 
 MANUAL OF MINING. 
 
 complicated underground stations are an objection to multiple- 
 deck cages. The double-platform 
 carcage is relatively lighter than 
 the single- deck two-car cage. 
 
 For inclines of uniform slope, 
 as well as for vertical shafts, cages 
 may be had, the platform being 
 hung on an adjustable lever; but 
 the slope carriage would be better. 
 Safety Clutches. Various con- 
 trivances are proposed for guarding 
 against accidents resulting from the 
 breaking of the rope. The sudden 
 starting and stopping of the hoist 
 produces a shock which strains the 
 cable more severely than does the 
 mere weight. Without care in hoist- 
 ing and rigid timbering the rope 
 cannot be insured against the jerks. 
 To avoid the accident sure to follow, 
 some form of safety appliance must 
 be thrown into action, to arrest the 
 fall of the cage or skip. The safety 
 catches differ in design and effi- 
 ciency, but depend generally upon 
 a spring, so held between the rope 
 and the cage as to be compressed, 
 while the weight of the cage strains 
 the rope, but acts on a clutch that 
 grasps the guides and stops the 
 fall, if relaxed by rupture of the 
 rope or otherwise. The clutch is 
 either a pair of sharp-pointed steel 
 levers, which are thrust outward 
 into the timbers, or a serrated 
 cam, the wider part of which will 
 
 FIG. 94. A Double-deck Cage.
 
 HOISTING MACHINERY, ETC. 229 
 
 be turned against the guides and clutch them on either side 
 (Fig. 93). The heavier the weight the stronger will be the 
 grip after they have once taken hold. 
 
 In Callow's attachment, 13, Fig. 84, a heavy weighted lever 
 on top of the cage, held up by a spring only so long as the rope is 
 intact, is set into action and engages the clutches upon the guides 
 when the rope breaks. Many a life has been saved by them, 
 but in many instances also they have failed to operate. A momen- 
 tary check, or any sudden change in the speed, often unnecessarily 
 throws them into action. On the other hand, they are rarely, 
 if ever, in order when the emergency arises, or the guides are so 
 wet and dirty that the clutches fail to catch, if the momentum of 
 the falling cage is great; besides, they are costly and troublesome. 
 Though useful adjuncts, which the law requires, yet it is not sur- 
 prising that the distrust of them is strong; they give rise to other 
 and more serious causes of alarm. The tendency following their use 
 is to rely upon them solely and to neglect the rope. Others deem 
 them of no value and treat them accordingly. Fortunately, the 
 rope more often breaks at the moment of starting at the bottom 
 than at any other time, and the point of rupture is where the 
 rope enters the socket. If, however, the rope snaps as it turns the 
 sheave and this is of common occurrence there is nothing to 
 prevent the inevitable and frightful calamity that . follows the 
 entire length of the rope falls and crushes cage and cbntents. A 
 simple appliance might be introduced at the top to grasp the rope. 
 
 Landing-doors at Shaft Stations. The landing of the buckets 
 is effected on a hinged door, of double 2-inch planks, lined on top 
 with iron. This is swung against the far side of the compartment, 
 closes it completely, and standing as it does at 45, the bucket 
 slides into the drift. When not in use it is hooked upright, closing 
 the mouth of the drift and leaving the hoistway clear. That 
 miners may go from one drift to the opposite, an escapement 
 roadway is provided around the shaft in the rock. At the mouth 
 of the shaft are two heavy doors, opened for or by the rising 
 bucket and closed at all other times. These are convenient, 
 and prevent accidents arising from stones dropping down the
 
 230 MANUAL OF MINING. 
 
 shaft, especially while dumping the bucket. A similar arrangement 
 in slopes a movable drawbridge, lowered at pleasure receives 
 the car for attachment to, or after detachment from, the rope, and 
 closes the gangway when the car is above the plat. 
 
 Cage-chairs. At the surface and at every landing- station the 
 cage is supported upon chairs, known also as "keeps," "fans," or 
 "shuts." They are counterbalanced levers (Fig. 93), which pro- 
 ject into the shaft, giving way for the ascent of the cage, and 
 immediately resume their position after the cage passes them. 
 The latter rests upon them during loading. To lower the cage 
 an attendant raises them out of the way while the cage is being 
 lifted off. 
 
 With double- decked cages the top deck of the cage at the 
 shaft bottom rests on keeps, while the bottom of the other cage 
 stands on similar supports. When the cars have been changed 
 a signal is given, the cages are lifted, while the attendant at the 
 surface pulls them clear of the shaft until the bottom deck has 
 been lowered below them. Releasing the handle thrusts them 
 into the shaft to hold the cage at its top deck. 
 
 Those at the shaft bottom are necessarily handled differently. 
 As the loaded cage leaves the shaft bottom the attendant there 
 pulls the handle of the keeps back and secures it, thus preventing 
 them from protruding into the pit. Holding the lever working 
 the keeps, when the lower deck of the cage has passed below 
 their level, he allows them to spring out to the support of the top 
 deck. The car is then changed, while the lower deck of the 
 other cage is being changed at the surface. The keep-handle 
 will not require further attention from the attendant below until 
 the cage has left for the surface, when he secures the handle back 
 in its place. 
 
 Stauss's patent keeps, shown in 15, Fig. 84, have been de- 
 signed to lower the cage into the pit without a previous lift. They 
 are fixed on wooden spring cantilevers, to take the shock of the 
 cage when it is lowered on them. At 16, Fig. 84, is the end view 
 of another style of lowering-keep. These reduce the wear and 
 tear of the winding ropes and engines. The cage can be low-
 
 HOISTING MACHINERY, ETC. 
 
 2 3 I 
 
 ered immediately without the lifting jerks, which are so grea at 
 cause of deterioration of ropes and engines. When these keeps 
 are used the winding rope must of course be carefully adjusted 
 in length so that the cage does not fall after drawing back the 
 keeps. The sinking of the cage should not amount to more 
 than the slack of the rope when unloaded. The length of the 
 rope can easily be adjusted by means of cage-adjusting hangers, 
 which are essentially of the turnbuckle type. 
 
 Self-dumping Cages. The self-dumping cage (Fig. 95) is a 
 bucket supported below the cage, whose floor is divided and 
 hung at the sides like two doors. Below it is loaded from cars, 
 and at the surface it is dumped very easily. Unless the coal is 
 
 FIG. 95. A Self-dumping Cage. 
 
 to be coked, tender coal cannot be treated in this way. Where 
 all the coal is to be coked, or at a small mine marketing its coal, 
 self-dumping cages are used, only one man being required for 
 the work. The capacity of a pair of them is limited to 1000 tons 
 of run-of-mine coal if screened over one set of bars. 
 
 Mine-cars. The dimensions of cars are reduced to a mini- 
 mum commensurate with the size of the haulage-way, and they
 
 *3 2 MANUAL OF MINING. 
 
 must be large for purposes of economy. To obtain the highest 
 efficiency they are made as large as the natural conditions of 
 grade in the haulage-ways will allow. Their weight is reduced 
 to a minimum in order to economize on haulage-power. Indeed, 
 the amount of power available determines their maximum allow- 
 able gross weight and grades suitable for the haulage-ways. 
 
 In small mines where man-power is employed for tramming, 
 the limit of weight and size is fixed by the inclination of the return 
 track; so, too, in buggy-rooms the weight and capacity of the 
 car depends upon the power there employed. These cars are 
 
 FIG. 96. A Metal Mine-car. 
 
 therefore, limited to small dimensions. Coal-cars are much 
 larger and heavier when mechanical systems of haulage can be 
 utilized. If the conditions permit automatic delivery of the cars 
 from the face to the locomotive or the haulage-way, their weight 
 is that which the mule or horse can conveniently return empty 
 to the face. Evidently the car should be as light as is com- 
 patible with its strength. Its weight is about one half that of its 
 contents. In metal-mines the proportion is somewhat greater. 
 The car weighs from 600 Ibs. to 1500 Ibs. Some coal-cars carry 
 as much as 120 cu. ft. of mineral, but the capacity of those in 
 metal-mines is less than 30 cu. ft. For strength, compactness,
 
 HOISTING MACHINERY, ETC. 
 T" 
 
 ~T 
 
 233
 
 234 MANUAL OF MINING. 
 
 and tightness sheet-steel car-bodies are used where the mineral is 
 hard, as in metalliferous mines. Otherwise a wooden box is em- 
 ployed. If the mineral is heavy but soft, it is necessary to employ 
 a sheet-steel boxing in preference to wood 
 
 The width of the cars depends upon the gauge and its set. 
 The bodies resting on the axles are preferred, but may not be 
 advantageous because of the wide gauge of track. On the other 
 hand, with a narrow-gauge track and given capacity the car- 
 body is elevated and its centre of gravity too high. A compro- 
 mise is frequently taken by which a low wide car of great capac- 
 ity may be raised on a narrow track. In this case the axles are 
 elbowed for large wheels and the car-body is set down on them, 
 but is widened above the wheels to full width of the track. Their 
 width must provide ample clearance for passing men in the haul- 
 age, and their length allow for space in the cage compartment. 
 The length is limited by sharpness of the curves. In the collieries 
 the maximum is 9 feet, and in metal-mines it is not over 6 feet. 
 
 Their height depends upon the conditions of loading. In 
 metalliferous mines the cars run only in main haulways, and are 
 filled from chutes, provided with a spout and gate, easily manip- 
 ulated at the bottom; if also hoisted on a cage, their height is a 
 matter of indifference. When the seam is thick and flat and the 
 roof good, they are carried to the face of the work, in which case 
 they are filled by hand. If so, or if raised on carriage, the diffi- 
 culties and expense are greater with a high car To shovel one 
 ton into a 3-foot car requires over 7300 ft.-lbs. ; into a car 4 feet 
 high, 9500. The average man can exert a continuous shovelling 
 effect of 28,100 ft.-lbs. per hour. Allowing for the weight of the 
 shovel, delays, throwing the mineral forward, a shoveller may 
 load about 20 and 14 tons, respectively, in the cars per shift. 
 Even for a medium output the economy is manifest. In metal- 
 liferous mines a low car is used, but in collieries heights of 4 feet 
 9 inches and over are common. For stability, too, a low car is 
 desirable. 
 
 Car Details. The steel cars are purchased of the manu- 
 facturers and may be had to order of any size and design for the
 
 HOISTING MACHINERY, ETC. 
 
 235 
 
 
 
 
 
 
 
 
 
 
 II 
 
 
 1 l! ;] -- 
 
 "Ljj-jjJ < \ 1 
 
 I 
 
 r
 
 MANUAL OF MINING. 
 
 111
 
 HOISTING MACHINERY, ETC. 237 
 
 given requirements. They are of steel plate \ inch thick, with 
 their sides and back end in one piece. The bottom plate is 
 fastened to the sides by iron angle-arms. The standard coal- 
 cars are of the shape, size, and details as shown in Figs. 96 to 100. 
 
 The metal cars are supported on one four-wheeled truck. 
 The wooden cars, on two trucks, are almost all of them arranged 
 with a swing-door at the end and provided with a trip-catch for 
 a quick release. It is rare for companies to make their own 
 cars, though all make their own repairs. 
 
 The Wheels are as large as circumstances will permit (the 
 larger the wheels and the smaller the axles, the less is the fric- 
 tion). The wheels may revolve loosely on the round or the 
 square axle, or they may be fixed to the axle and revolve with 
 it. Some are capped with a recess in the hub, to receive the 
 collar on the axle, and thus prevent admission of grit (Fig. 100). 
 They may be "inside" (below) or "outside" (beyond) the 
 body of the car. As to the relative merits of the inside and out- 
 side, or loose, wheels, it must be admitted that engineers are 
 not unanimous, though the former have the larger number of ad- 
 herents. Outside wheels are more easily oiled, are cheaper, and 
 admit of the body of the car being set lower down; they do not 
 run as smooth or last as long as those fixed under the body of 
 the car. Loose-wheel cars may be better for short roads with 
 sharp curves, but they are harder to pull. With fixed wheels, one- 
 of the mutually dependent wheels, in travelling about curves, must 
 slide. For this reason, and because they are easily lubricated, loose- 
 wheels, or cone-fixed wheels, are preferred by many. A great 
 many mines have abandoned loose wheels after careful trial. 
 At the Drifton anthracite mine a compromise is effected by- 
 using a pair of fixed and a pair of loose wheels. 
 
 The coal-car wheels are of cast iron, between 16 and 18". 
 inches diameter, and those of ore-cars about one half that and' 
 solid. The former have hub and arms to allow of "spragging." 
 
 The wheel-base of the car-trucks is a trifle greater in lengths 
 than the width of the track. Its width is f inch less than, the- 
 track gauge.
 
 338 MANUAL OF MINING 
 
 The life of a car depends upon conditions too numerous to 
 detail here, but it may be given as averaging three years in coal- 
 mines and a trifle longer in metal-mines. They all grow larger 
 'with age. The wooden cars become loose and swell; the iron 
 ones are quickly battered out of shape. 
 
 Brakes are provided on all coal -cars which go to the face of 
 the workings. Otherwise they are not required. 
 
 Car Resistance. The frictional resistances offered by the 
 wheels at their circumference on the track and at their journals 
 amount to about 2 per cent of their total weight. This quantity 
 may be smaller where care has been given to their construction, 
 as, for example, those which are employed in connection with the 
 electric mine-haulage. This friction can be reduced by the 
 liberal use of lubricants and care in the finishing of the journals. 
 Jf the latter are boxed and kept free from dust, the frictional 
 resistance will be materially reduced. 
 
 The lubrication of the bearings may be obtained in the ordinary 
 -way by the use of oil -soaked waste in the boxes forced against 
 the journals. Self-oiling journals would further reduce the fric- 
 tion. They are inexpensive, and are desirable if saving in power 
 is to be considered. By their use one third of a pint of oil will 
 last from two to four months, according to its quality and the dis- 
 lance traversed. The saving in power is $10 per year per car, 
 compared with the inefficient methods. This includes also the cost 
 of axle replacements. Self-oilers are also economical of lubri- 
 cant, as compared with the crude method ordinarily employed 
 of pouring oil into a box and bearing. The oil used is usually 
 one of the heavy varieties of lubricant with the addition of some 
 axle-grease. 
 
 Lubricants. An efficient lubricator must have sufficient 
 body to keep the surfaces between which it is interposed from 
 coming into contact. It must have the greatest fluidity con- 
 sistent with its needed body. A small coefficient of friction is 
 necessary, as well as a maximum capacity for conducting the 
 heat. It should be free from any tendency to oxidize or gum, 
 and free from acid-
 
 HOISTING MACHINERY, ETC. 
 
 239
 
 240 
 
 MANUAL OF MINING. 
 
 Testing Lubricating-oils. The engineer should have means 
 for testing, as oils intended for lubricating purposes are often 
 adulterated with foreign and even injurious matter. A hydrom- 
 eter and flash-test cup are sufficient for this purpose. 
 
 THE FLASH-POINT AND SPECIFIC GRAVITY, BEAUME, OF LUBRICATING OILS. 
 
 
 >, 
 
 1 
 
 A 
 
 
 
 
 .,- 
 
 . 
 
 
 Is 
 
 P 
 
 I 1 
 
 P 
 
 
 II 
 
 a 
 
 11 
 
 
 
 Ac* 
 
 ill 
 
 Castor-oil (pure). 
 Linseed-oil 
 
 0.9639 
 .9299 
 
 15 
 19 
 
 55 
 
 Tallow-oil 
 Heavy engine-oil . . 
 
 0.9080 
 .9032 
 
 24-5 
 25-5 
 
 540 
 
 414 
 
 Cottonseed-oil. . . 
 
 .9210 
 
 jg 
 
 518 
 
 Castor machine-oil 
 
 .8919 
 
 27 
 
 324 
 
 Lard-oil 
 
 9*75 
 
 2 3 
 
 505 
 
 Light engine-oil. . . 
 
 .8917 
 
 27 
 
 342 
 
 Rapeseed-oil 
 
 9155 
 
 23 
 
 
 Sperni-oil 
 
 .8815 
 
 29 
 
 
 Cvlinder-oil 
 
 
 
 
 Spindle-oil 
 
 .8^88 
 
 3? 
 
 ? j 2 
 
 
 
 
 
 
 
 
 
 
 The Hydrometer. All oils are lighter than water and their 
 purity may be ascertained by the use of a Beaume hydrometer. 
 If a contract calls for a heavy engine-oil of 25 Beaume, the hy- 
 drometer at once reveals whether the same is received. If adul- 
 terated, a lower Beaume reading reveals the blending. The 
 adulterants used are, of course, cheaper than the pure oil. A 
 "lard-oil," for example, of 23 Beaumf, sp. gr. 27, must have 
 been adulterated with a lighter oil, and the only cheaper ones 
 possible are the light petroleum or the neutral oils. 
 
 Viscosity Test. The viscosity and gumming tendency of oils 
 may be detected by noting the time required to traverse a given 
 distance down an inclined plane, and the time when this ceases. 
 The oil remaining limpid the longest has the lowest gumming 
 tendency, and that which flows the farthest is the most viscous. 
 A nine-days' test of oils gave the following results: A heavy 
 mineral-oil, for engine use, traversed 87 inches on the ninth 
 day; common sperm-oil, 63 inches on the ninth day; olive-oil, 
 22 inches on the ninth day; linseed-oil, 18 inches on the seventh 
 day; and lard-oil, 12 inches on the fifth day. In each case the 
 day given is that on which the oil ceased to travel. 
 
 The simplest method of testing the fluidity is to dip blotting- 
 paper into the oil and hold it up to drain. An oil of good fluidity
 
 HOISTING MACHINERY, ETC. 241 
 
 presents symmetrical drops. One of high viscosity spreads 
 readily over the paper. The rate of gumming may also be 
 ascertained by its retention on the paper, while at high tem- 
 perature, for some hours. 
 
 A Ring Test. An oil which is composite in nature will show 
 on a clean sheet of blotting-paper two or more rings, the outer 
 one indicating the poorest oil. The central denned ring, dark 
 in color, is due to the heavier oil. If the rings disappear, it 
 is mineral-oil. Paraffine stock in the oil will be revealed by a 
 well-discerned ridge near the centre. A light hydrocarbon oil 
 does not give any permanent translucency to the paper. 
 
 The Flash-point Test. This consists in finding the tempera- 
 ture which an oil will endure without being vaporized and ignited. 
 An oil for cylinders should have a high flash-point; those not 
 intended for lubricating steam surfaces may have a much lower 
 flash-point. The flash-point can be determined by heating 
 a sample of the oil in a cup over an alcohol-lamp, on a tripod. 
 A thermometer inserted in the oil will note the degree at which 
 ignition takes place. Care should be taken that the thermometer 
 touch neither the bottom nor the sides of the cup. As the oil 
 reaches 300 F. a match or lighted twine should be passed across 
 the surface of the oil until a blue flame shows the time of ignition. 
 Lard-oil has a high flash-test. If the sample submitted burns 
 below 500, it is adulterated with cottonseed or petroleum; lin- 
 seed-oil will have its flash-test reduced by the same adulterants. 
 
 Fire-proof Oils. These are valuable on bearings inclined to 
 heat. They derive their virtue from the fact that they contain 
 considerable water, held in suspension in the oil by means of a 
 mineral soap. When applied to heated bearings the water 
 serves to absorb the frictional heat, leaving the oil for lubrication. 
 
 Light-running machinery is lubricated with oil of 300 flash- 
 test, heavy machinery with oil of 400, and engines with oil of 
 
 550. 
 
 Oil intended for the bearings of motors and generators ex- 
 posed to low temperature should be limpid and free-flowing. 
 
 Cylinder lubrication is best obtained from a petroleum-oil of
 
 242 MANUAL OF MINING. 
 
 high flash-test (530) and viscosity. It is compounded with 
 5 per cent of tallow-oil if the steam is wet. 
 
 The air-cylinders of air-compressors and of gas-engines should 
 be lubricated only by a pure petroleum-oil. 
 
 Geared wheels are oiled by some preparation composed of tar, 
 wax, and similar substances combined with rosin, which, when 
 applied to the teeth in a molten state, will cool to a tough, elastic 
 coating, reducing friction and noise. 
 
 Handling Cars at the Tipple. To facilitate the delivery 
 of the mineral at a minimum expense, and with as little hand- 
 ling as possible, the landing-station or the tipple is fitted with 
 quick-dumping appliances for the cars and suitable grades for 
 their prompt return. The level of its floor is below that of the 
 mine mouth or the landing-station in the derrick, and far 
 enough above the railroad track or the point of shipment to per- 
 mit the mineral to have a continual descent between the time of 
 dumping and the time of shipment. The tracks to the tipple are 
 laid favorably for the load, and the character of the conducting 
 line between them will depend upon the distance and the slope 
 of the ground. In the case of a slope mine the loaded cars or 
 the train can be hauled directly to the tipple and returned thence 
 by gravity. Any grade exceeding one per cent will allow the 
 empty car to return. 
 
 Dumping Cars. The dumping of cars may be accomplished 
 by opening one side or the end of the car by inclining the car 
 on its axle or by tipping it upon some form of cradle. Mineral 
 which is to be subsequently broken or sorted is dumped upon 
 a screen or the grating to the sorting-floor. Some tipples are 
 provided also with self-indicating weighing-machines, by which, 
 as the car is emptied, the machine registers its weight, turns 
 over, shoots out the coal, and returns to its normal position for 
 another load. 
 
 The iron-body car of metal-mines is set on a swivel near the 
 centre and held in position by a lever. When it is to be emptied 
 the lever is unhooked, the car turned to position, the door raised, 
 and the car tilts because of the slight overhang at the discharging
 
 HOISTING MACHINERY, ETC. 
 
 243 
 
 end. It empties without further effort from the trammer (Fig. 
 101). It is fastened by the latch as soon as the car body has been 
 
 FIG. loi. A Steel Ore-car. 
 
 swung into position on its truck. Cars intended for tunnel 
 use empty at the end, the door being hung by strap hinges from 
 
 FIG. 102. A Side Dump-car 
 
 an iron rod at the top. It latches in the same way as mentioned. 
 Another variety of tunnel car consists of double iron frame,
 
 .?44 
 
 MANUAL OF MINING. 
 
 pivoted at the middle of the top line on the side. It is on two 
 trucks for dumping. The latch, located at each side, is raised, 
 the car opens at the middle and empties between the trucks 
 (Fig. 103). The doors on the ends of coal-mine cars are locked 
 by latches similar to those of the iron-frame cars. In Figs. 
 
 Cost $100 + 
 
 Wt. 900-1000 IDS. 
 Capacity % cord - 2 Tons 
 
 FIG. 103. Tunnel Dump-car. 
 
 105 and 1 06 are illustrated some typical latches used in the coal 
 regions. 
 
 The empty cars can be returned to the mine by being raised 
 by a short chain-haulage to an elevation sufficient for them ta 
 return by gravity to the mouth of the mine. A chain-haul can 
 be used to return the empty cars to a "kick-back" beyond the 
 shaft if the distance from the dump will permit. This is then 
 worked like the Ramsay caging apparatus, but is much cheaper 
 than the latter. 
 
 Cross-overs. Various combinations of tracks and machinery 
 are employed to expedite the loading and unloading of cars on 
 the cage and economize time at landing-stations. The plan is to 
 displace the empty car upon the cage by one to be placed thereon. 
 For example, the loaded car pushed by hand upon a cage bumps
 
 HOISTING MACHINERY, ETC. 
 
 245 
 
 off the empty car at the bottom of the shaft, or the empcy car to 
 be lowered into the mine is made to replace the loaded car, which 
 is simultaneously run off. The automatic cross-overs are of sev- 
 eral types and very effective. The usual plan consists in having 
 the car approaching the dump release dogs holding the empty 
 car, and allow it to pass over the dump to the "kick-back" (Fig. 
 104), whence it may be returned by the track on the left. 
 
 The Ramsay Caging Apparatus. This consists essentially of 
 two steam-rams placed back of the shaft and two transfer-tracks, 
 running on a track across the tipple in the rear of the head-frame, 
 which are operated by a steam-cylinder. Its operation is as 
 follows : 
 
 FIG. 104. The Ramsay Caging Apparatus. 
 
 After a car has been dumped it runs back by gravity past one 
 side of the head-frame into one of the transfer-tracks, which is 
 then moved by means of the steam-cylinder to the rear of the com- 
 partment where the next loaded car is coming up. When the 
 cage with the loaded car is at the landing, the empty car is pushed 
 by means of one of the steam-rams against the loaded car, which 
 is then taken off the cage and the empty car left on the cage 
 ready to descend, so that the loaded car is shoved from the cage 
 and the empty one pushed on at one operation. When one 
 transfer-track is back of the frame with its empty car, the other 
 track is in position to receive the next empty car as it comes 
 from the dump.
 
 2 4 6 
 
 MANUAL OF MINING.
 
 HOISTING MACHINERY, ETC. 
 
 o 
 ft 
 
 3 
 
 lo, a 
 
 CO 
 
 u
 
 2 4 8 
 
 MANUAL OF MINING. 
 
 Cradle-dumps. Various forms of cradles are provided at 
 dumping-stations to facilitate the emptying of cars with or with- 
 out hinged doors. These consist of some form of balance-frame 
 upon which the loaded car is run and held. When the catch is 
 released the frame, with its car, tilts and empties the contents as 
 soon as the centre of gravity has passed slightly beyond the 
 pivot, or axle, of support. Another form (Fig. 107), which is the 
 
 FIG. 107. 
 
 Behr device, is much used. The car is held on a platform and 
 revolves with it around the point of support, as indicated by the 
 dotted lines. The car in this device is almost completely in- 
 verted and thus emptied. Another form of cradle, used with
 
 HOISTING MACHINERY, ETC. 249 
 
 larger cars having no swinging doors, consists of a skeleton barrel 
 having tracks in such a position as to receive the car, which is 
 then completely enclosed. When the trap is released, the barrel is 
 revolved by gear-wheels, and the car is inverted. In the anthra- 
 cite mines are employed different methods of dump, as illustrated 
 in Figs. 1 08 and 109. The shock of the loaded car running 
 upon the truck carriage supported on the gravity tracks causes 
 the front wheels of the latter to roll down the curve, while the 
 end wheels are elevated and the car is tipped sufficiently to empty 
 its contents, the door having been previously released. 
 
 Means of Access for the Men. For purposes of ingress and 
 egress mines are provided with ladders or man-engines, where 
 the cage or bucket is not used. The statutes of the several States 
 contain various requirements for the accommodation of the men. 
 Some require the maintenance of substantial ladders in a separate 
 compartment as the sole means to be used by the men for entry 
 and exit. In other States operators are relieved of the necessity 
 of keeping up a ladder-way, if safety- carriages are employed. 
 The laws of many States forbid the use of buckets by the miners, 
 while the general tendency in all regions is to insist upon two 
 well-equipped escapement-ways. 
 
 If the angle of entry is below 30, no special provision is 
 necessary. The mudsills of the timbering break the descent 
 into sufficiently convenient steps. For steeper angles, up to 
 about 60, some variety of tread is necessary. When the pitch 
 exceeds this the compartment must be provided with ladders 
 isolated from the hoistway. They should be inclined, uni- 
 'form in direction, at an angle of not less than 10 from the 
 vertical, to diminish the fatigue of climbing and enable the 
 men to carry tools with them. At equal distances down the 
 ladder-way (20 to 40 feet down a vertical shaft, and at greater 
 distances on an incline) platforms are built of 2 X6 beams and 
 2-inch planks, closing it, except for a manhole, at the foot-wall 
 end. The ladders extend up through the manhole, and are 
 fastened by staples or toe-nailed to the shaft-timbers, and rest 
 on the far side of the plats. They are made of 2 X6 standards,
 
 250 
 
 MANUAL OF MINING. 
 
 ** 
 
 7\ 
 
 
 
 
 
 FIG. 108. Anthracite Car-dump.
 
 HOISTING MACHINERY, ETC. 
 
 1 8 inches apart, with iron or wooden rounds or rectangular slats 
 12 inches apart. The last-named are cheaper, last longer, and 
 
 give better toe-hold than wooden rounds, which, in turn, are 
 easier to use than the more durable iron. 
 
 In slopes having an inclination of less than 60 steps and 
 rails are provided, the tread being 6 inches and the slope dis-
 
 252 MANUAL OF MINING. 
 
 tance 12 inches. The headroom provided is at least 6 feet verti- 
 cally; when the pitch is less than 45 the rule obtains making 
 the rise plus the tread in inches equal 24; below 20 pitch noth- 
 ing is required except occasionally a plank. 
 
 Though used in Europe for 1200 to 1500 feet depth, and in 
 this country in deep mines, ladders are certainly not ad- 
 visable. According to the Cornwall Society, their use deranges 
 the respiration and shortens life by ten years. The miners 
 reach the workings more or less exhausted, and the operators 
 have lost the benefit of a proportionate amount of energy. Unques- 
 tionably, an element of success worthy of attention by mine mana- 
 gers a pecuniary as much as a humanitarian question is the 
 proper treatment of and the conveniences for the men, who 
 unconsciously reciprocate in an equivalent of work. Ladders 
 waste time. It takes 15 minutes to go down 300 feet, and the 
 ascent is twice as slow. A shift of forty men, following one 
 another at intervals of 8 feet, entails a loss to the company of 31 
 minutes each shift. With buckets and cages the loss is not so 
 great; eight men at a time, lowered 1200 feet, consume 40 
 minutes for every shift of 100 men. An additional loss occurs 
 at tally-time from the reduction of the hoisting capacity, which, 
 with the impatience of the men, leads to the crowding of the 
 cage; but in most States the limiting number of men permitted 
 on the cage is named. 
 
 The Man-engine. Movable ladders or man-engines, invented 
 by D'Orrell, of Clausthal, were adopted as acceptable substitutes 
 for ladders and were once popular in deep mines. When intro- 
 duced into Cornwall by Mr. Lorn, the engine was declared by 
 the Royal Polytechnic Society a very tl great boon to miners." 
 Its introduction involved the addition of some machinery, and, 
 though easy to operate, it is now obsolete as an economic 
 arrangement. 
 
 Two rods, of decreasing cross-section from the top down, receive 
 at the surface an oscillatory motion from balanced bobs, operated 
 by an engine having a fly-wheel and other regulators. The 
 dimensions of each rod at any point must be such that it will have
 
 HOISTING MACHINERY, ETC. 
 
 2 53 
 
 the requisite tensile strength to support the weight of the part 
 below it, loaded with men. The rods play between roller-guides 
 50 feet apart, and are provided with wings and catches, after the 
 manner of the Cornish pump-rods, which may, in fact, be utilized 
 as"Fahrkunst" rods without much extra power. 
 
 Each rod has a small platform (Fig. no) about i2"Xi2" or 
 1 8" at every 12 feet double the length of the stroke. A 
 handle 4 feet above the platform gives support to the miner, who 
 is carried up 6 feet on one rod, which brings 
 him opposite a platform on the companion 
 . rod; upon this he steps, to be lifted 6 feet 
 more, to meet a plat on the first rod, which 
 has been coming down to receive him. A 
 miner stepping from one to the other is car- 
 ried up or down at a rate of from 48 to 96 feet 
 per minute (each rod makes 4 to 8 double 
 strokes, delivering one man each time; those 
 at the Calumet and Hecla make five strokes). 
 As there is no limit to the depth to which these 
 may be carried, and as they are capable of 
 working alike in slopes as in shafts, it is not 
 surprising that they "take" so well. They 
 replace hoisters, and require little additional 
 power or space. Tools and supplies cannot 
 be carried by the miner, but may be delivered 
 by the cage or bucket. 
 
 A single rod is also used, its companion 
 being replaced by stationary platforms at- 
 tached, 6 feet apart, to the shaft-timbers. 
 Upon these the ascending men wait during the 
 down-stroke of the rod. The single-acting man-engine requires 
 chains and counterpoises at intervals, to balance it and to prevent 
 the shock incurred at the end of the stroke. 
 
 From the fact that a misstep would be fatal, it would seem 
 as though man-engines were extra-hazardous, yet the accident 
 record does not confirm this belief. Some confusion is caused 
 
 FIG. no.
 
 254 MANUAL OF MINING. 
 
 by a man missing his plat and riding on, to the annoyance of 
 those following him; but this is of rare occurrence unless his 
 light goes out, for there is a halt of several seconds at each change 
 of motion. In Prussia, out of an average of 100,000 men employed 
 for ten years, only 57 were injured on the man-engines; in 
 Cornwall, 17. This is more than compensated for by the increased 
 length of life of the miners using them. 
 
 The cost of machinery, etc., for a 1200- foot man-engine is 
 $18,000, upon which interest and depreciation may be figured at 
 $2500, amounting to 10 cents per man daily on a gang of 100 
 men. The running expenses at the Dolcoath mine are 4 cents 
 per man for 2400 feet. 
 
 REFERENCES. 
 
 Winding Ropes in Belgium, Coll. Guard., Jan. 16, 1903; Glossary of 
 Rope Terms, Mines & Min., April 1904; Notes on Steel-wire Cables, Prof. 
 F. Soule, A. I. M. E., Vol. XXIX, 550; The Testing of Winding Ropes in 
 the Province of Anhalt, Germany, F. H. Probert, A. I. M. E., Vol. XXX, 
 1020; The Hand Auger and Hand Drill in Prospecting Work, C. Catlett, 
 A. I. M. E., Vol. XXVII, 123; Mine Ropes: Strength, Care, etc., C. Aquil- 
 lon, Colo. S. of M. Quart., I, 1892, 20; Vegetable-fiber Ropes, M. & M., 
 Vol. XXIV, 530; Testing Wire Ropes, M. & M., Vol. XXIV, 405; Cable 
 Wires, M. & M., Vol. XXIV, 411; Stresses in Wire Ropes, M. & M., Vol. 
 XXII, 316; Due to Shock, M. & M., Vol. XX; Mine Ropes: Strength, Care, 
 etc., Colo. S. of M. Quart., I, 1892, 20; Hoisting Ropes, George S. Whyte, 
 M. & M., April 1904; Splicing Wire Ropes, M. & M., April 1904. 
 
 Head Frame Illustrated, Penna. Mine Insp. Packet, 1886, 1889, 104; 
 1891, 395; 1889, 240; 1890, 389; Head Works Framing, Charles H. Fitch ) 
 Min. & Sci. Press, June 13, 1903; Steel Head Frames, M. & M., Vol. XX,' 
 292; and Steel for Coal-mine Tipples, Jour. W. Soc. of Eng., IX, 343, and 
 X, 383- 
 
 Combined Safety Clutch and Stopping Device or Winding Cages, O. 
 Schenck, Coll. Guard., April 12, 1900; Safety Clutch, 111. Min. Inst, II, 109, 
 159, 180, 212; Safety Catches, Automatic Appliances in Winding, C. Roemer, 
 Coll. Guard., Nov. 1896, 916; Safety Clutch, Mine Insp., Pa., 1886, T\b; 
 Attachments, 111. Min. Inst., II, 109, 159, 180, 212. 
 
 Tipping Cars, Illustrations, Coll. Guard., April 30, 1897; Spring Coup- 
 lings (Cage to Rope), Trans. M. & M. Eng., XLVI, 54; The Tipping and 
 Screening of Coal, J. Rigg, N. Staff., IV, 103; Dumping Cradles for Mine
 
 HOISTING MACHINERY, ETC. 255 
 
 Cars, A. I. M. E., XVII, 564; Coll. Guard., April 30, 1897, 824; Colliery 
 Surface Works, E. B. Wain, Coll. Guard., Dec. 1904, 1076; A Steel Coal 
 Tipple, Eng. News, June 1904, 577. 
 
 Slope Carriage, Pa. Mine Insp., 1882, 39, 18. 
 
 Steel Mine Cars, M. & M., Vol. XXII, i 4 , : Steel Wheels, M. & M., Vol. 
 XXII, 364; Cars, A. I. M. E., Vol. V, 502; Latches, Dogs, etc., Hildbrandt, 
 essay, 23; The Comparative Merits of Large and Small Trams for Colliery 
 Use, James Brogden, So. Wales Inst., VI, 173; Car Clutches, Hildbrandt, 
 essay, 42; Wheels, Self-oiling, Mine Insp., Pa., 1889, no; Note on the 
 Friction of Car Wheels, R. Van A. Norris, Amer. Inst. M. E., XVIII, 508; 
 Packing and Lubrication, R.R. Gaz., Jan. 26, 1900; Car for Handling Rock 
 at Mines, L. L. Logan, M. & M., Oct. 1903. 
 
 Ladders, Strength of: Tables, etc., R. G. Browning, Eng. & M. Jour., 
 June 1897, 602. 
 
 Mine-car Wheels, Coll. Guard., LXXX, 392; Self -lubricating Wheels, 
 Coll. Guard., Vol. LXXX, 392; Automatic Stop for Cars, Coll. Guard., Vol. 
 LXXIX, 66; Life of Cables, Coll. Guard., Vol. LXXIX, 224; Pit Ponies, Coll. 
 Guard., Vol. LXXIX, 1092; Detaching Hooks, Coll. Guard., Vol. LXXXI, 
 369; Testing Hoist Ropes in Germany, Coll. Guard., Vol. LXXXI, 456; 
 Telescope Door for Cages, Coll. Guard., Vol. LXXXII, 415; Working Loads 
 for Manilla Rope, Coll. Guard., Vol. LXXXIII, 143; Belgian Hoist Ropes > 
 Coll. Guard., Vol. LXXXI V, 123; Changing Winding Ropes on Koepe Pul- 
 leys, Coll. Guard., Vol. LXXXVI, 1130, and LXXXV, 425; Safety Measure* 
 against Overwinding, Coll. Guard., Vol. LXXXVI, 1197. 
 
 Detaching Hooks for Cages and Ropes, 111. Min. Inst., II, 109. 159, 180, 
 212; Shaft Accidents and Preventions at Level Landings, N. Staff. Inst., 
 VIII, 204; Fencing Gates for Winding Shafts, W. Hay, Inst. M. E., X. 
 
 Overwinding Precautions, Coll. Guard., Jan. 22, 1897, 174; Its Preven- 
 tion at Collieries and Furnaces, W. Grimmitt, Fed. Inst. M. E., II; Its Pre- 
 vention by a Patent Apparatus, Fed. Inst. M. E., I; Overwinding Precautions, 
 Joel Settle, Coll. Guard., Jan. 1897, 174; Accidents from Overwinding and 
 other Cage Accidents in Mines, Coll. Mang., May 1893, 88; Overwinding 
 Safety Devices, Mining Bulletin, P. S. C., 1896; M. & M., Vol. XXII, 295. 
 
 Colliery Shaft Signalling, G. W. Smith, Min. Inst., Scot; Closing Top of 
 Upcasts, A. Reid, Fed. Inst. M. E., X; Electric Signals for Collieries, C. 
 McLaren Irvine, Min. Inst., Scot., IV, 47; Cages, Illustrated, Pa. Mine 
 Insp., 1884, 563; Mallisard-Taza Tipping Cage used in French Collieries, 
 Eng. & Min. Jour., I, 129; Stops for Mine Cages, Coll. Guard., 1896, 1173, 
 916; Signalling between Levels, F. B. Allen, N. Z. Mines Rec., Aug. 17, 
 1903; Hoisting Ropes, Robert Peele, Mines & Min., Mar. 1900; Mine Sig- 
 naling by Compressed Air, Bernard McDonald and William Thompson, 
 Con. Min. Rev., Sept. 30, 1002; Signal Device for Mines, C. S. Hertzig,
 
 256 MANUAL OF MINING. 
 
 A. I. M. E., Vol. XXX, 314; Electric Mining Signals, M. & M., Vol. XXII, 
 167. 
 
 Kips or Landing at Shafts, M. H. Douglass, Brit. Soc. Min. Stud., I, 433; 
 Shaft Timbering, J. C. Jefferson, Chest. Inst., VIII, 209; Some Arrange- 
 ments for Preventing Accidents at Level Landings in Cage Dips and Shafts, 
 A. R. Sawyer, N. Staff. Inst., VIII, 204.
 
 CHAPTER VIII. 
 
 UNDERGROUND HAULAGE SYSTEMS. 
 
 The Extent of Haulage Requirements. In any coal-mine of 
 moderate size there are fully 500 miles of track and headings 
 into the rooms, representing perhaps 2000 miles of road, which 
 at some time during its history had been laid down and given 
 service. It has been estimated that in the production of 200,- 
 000,000 tons of coal at least 5000 miles of track extension are 
 made each year. The great amount of it indicates the necessity 
 for giving special attention to its construction. The cost of 
 delivering coal from the working face to the tipple when the 
 tracks are level and the distance not over a mile is between 6 
 and 10 cents per ton. During a single month an average of 
 thirty-one mines using rope haulage showed a cost of 7.5 cents 
 per ton. In twenty mines employing electric haulage the run- 
 of-mine coal cost 6 cents per ton, and 8 cents in an average 
 of twenty-one other mines having mule haulage. Probably, 
 over the United States, the average cost of animal haul is nearly 
 15 cents. The cost of haulage increases materially as the distance 
 of travel exceeds 3000 feet. Beyond this length of haul it may 
 be advisable to sink another shaft if a convenient point of ship- 
 ment can be obtained at the surface and the depth of shaft be 
 not too great. 
 
 The Grade and Direction of a Haulage Road. A haulage 
 road is as straight and direct as circumstances will allow, so 
 designed as to utilize the natural forces to the closest degree. 
 The shaft and slope are carried down to the lowest point, that 
 to it may be delivered the mineral from the working face on a 
 
 257
 
 258 MANUAL OF MINING. 
 
 continual descent. Dropped through the mill-holes from stope 
 to cars, the ore is moved by hand- or horse-power to the foot 
 of the shaft; and coal, leaving the face in cars, should have 
 the advantage of the down-grades in its delivery to the shaft 
 bottom. 
 
 From the workings to the haulage-way the inclination of the 
 track will depend solely upon the mean inclination of the floor, 
 but the cars should be capable of moving by gravity without 
 the use of blocks or sprags. A sprag is a billet of wood 
 inserted between the spokes to prevent the wheels from revolv- 
 ing, thus converting rolling friction into sliding friction and 
 checking the speed of the car. On a 3 per cent grade only one 
 sprag is used, but on a 10 per cent grade four sprags will be 
 necessary to keep the speed of the car within moderate control. 
 A grade exceeding one in five is positively dangerous. Roads 
 are often designated as one-, two-, three-, or four-sprag roads, 
 according to the number of wheels which must be blocked. 
 
 The best grade for any haulage road is the gradient of equal 
 resistance that on which the work expended in hauling a loaded 
 car down equals that of the returning car up. A steeper grade 
 than this is wasteful of power both ways. As evidencing the in- 
 fluence of grade in the reduction of useful power, it may be cited 
 that a locomotive can pull up an incline of 2 feet in a hundred 
 only 80 per cent of what it can haul on a level, and on a 4 per cent 
 grade but 30 per cent. This is correspondingly true of animal 
 haulage. While a mule can haul three loaded cars on a reason- 
 ably level track, it can haul only one on a 6 per cent grade. 
 
 Haulage-ways in Metal-mines. Metalliferous mines rarely 
 present any difficulty for the road engineer. Very few roadways 
 feed to any given landing. In a vein there is but one line on 
 either side of the shaft for which provision is to be made. In large 
 ore bodies only a few tracks will radiate from the point of outlet, 
 and but few cars will travel over each. 
 
 Any desired grade may be laid in vein mines. It is as much 
 greater than that required for drainage as the available motor 
 power will allow for returning with the empty cars, or controlling
 
 UNDERGROUND HAULAGE SYSTEMS. 259 
 
 the speed of the loaded cars out. The direction of the roadway 
 is fixed by the sinuosities of the vein wall. For animals due 
 regard must be had to economy of size and weight of car and 
 its contents; hence the grade will probably not exceed 1.5 in 
 100 feet. For locomotive haulage the grade may be carried to 
 4 per cent as a maximum. 
 
 Haulage-ways in Coal-mines. Here the engineer must accom- 
 modate himself to the stratigraphical conditions. The thickness 
 of the coal-bed and the frequency and amount of the existing 
 faults, folds, and inclinations must be reckoned with, and a uni- 
 form grading would demand their elimination, according to the 
 character of the enclosing rock. 
 
 In the main headings the hillocks may be removed and the 
 sumps in the floor filled, or the roof material over the depressions 
 may be ripped down and the sumps filled with the waste. The 
 latter is more economical. It is far more expensive to take up 
 a foot of bottom than to blast down 3 feet of thickness in the roof. 
 Again, the cost of maintenance is not so great where the road-bed 
 filling is of loose slate from the roof, as in the customary floor of 
 fire-clay or limestone. Often the soft floor of the coal-seams 
 cannot be disturbed, and it may not be desirable either to fill 
 up the depressions or to remove the hills, because of inconvenience 
 in making connections to rooms. It is probable, too, that the 
 cleavage planes in the coal will determine the directions of the 
 haulage-ways, and the average dip of the vein will determine 
 their grade. 
 
 In coal-mines there are always two parallel headings for 
 ventilation purposes, the engineer having his choice as to 
 which shall drain the water and which be used for haulage. 
 It is rare that the same heading would be required to give 
 both services. Usually the intake is the one employed for 
 haulage, because of the better quality of air and, in consequence, 
 a less cost of maintaining the track. This roadway is never 
 obstructed by the air-crossings, which usually are placed on the 
 return-airway. When no other condition determines the choice, 
 the roadway higher in elevation is employed for haulage and
 
 260 MANUAL OF MINING. 
 
 the lower one for drainage, with frequent communications be- 
 tween them to allow of the accumulated waters of the former to 
 be promptly carried to the latter. 
 
 The haulage gangway is of single-track width only when the 
 system employed is the tail-rope or the locomotive. The double- 
 track gangway is employed where the traffic is heavy, or for an 
 endless-rope haulage-way. The advantage in the latter is the 
 great area given to the air-current, and also the opportunity 
 afforded for utilizing the excess of the power in the down-going 
 car or train to assist the engine in raising the load on the other 
 track. 
 
 The requirements for ventilation demand a roadway of liberal 
 area which will permit a large quantity of air to pass through with 
 the minimum of friction, and also reduce the effect of the dis- 
 turbance to the current by the passage of the train through it. 
 
 The enlargement of the area necessary to provide for the cars 
 may be obtained by increasing the width or the height of the 
 roadway. The former is preferable unless the roof is very firm 
 and the bed quite thick. The objection to it, however, is the 
 increased length of sills and the timbers necessary to maintain a 
 good level on the ordinary floor of the coal-mine. It is prefer- 
 able to increase the height unless the roof will not bear cutting 
 down. It is a cheaper method, for the dimensions of the sills 
 would be only of moderate size, and those of the posts would not 
 be any greater than for a wide roadway. 
 
 Track Construction. The tracks on the haulage-ways are 
 constructed with a view not only to diminish the working ex- 
 penses, but also to maintain a large capacity with a given amount 
 of power. The difference in cost and the time of laying between 
 a heavy rail on a substantial track and one which is poorly 
 laid is not large, nor is it commensurate with the difference in 
 iheir efficiencies. A reduction in the grade of \ per cent saves 
 annually in track maintenance an equivalent of 30 tons of fuel 
 for motor power for each car running through the mine. The cost 
 of such construction might, however, prove prohibitive of any 
 operations. The most that can be done is to improve the roads
 
 UNDERGROUND HAULAGE SYSTEMS. 261 
 
 within reasonable limits of cost after careful surveys and level- 
 lings of all roads intended for main arteries. 
 
 The road-bed is made as elastic as possible by the use of sand, 
 dump-coal, or ashes, without being too soft for support. The 
 choice of ballast depends upon the character of the mine floor. 
 If the latter is inclined to be wet or rough, then rocky material 
 would be preferable to sand or ashes. When the floor is very 
 strong, clay can be used to advantage with fragments of rock 
 grounded in. The dry road-bed can be secured by rounding 
 off the surface between the tracks to a slight arch. This will 
 assist the removal of water from the track to the drains. 
 
 The ties are laid 2 feet apart between centres for rails of 15 
 to 30 feet in length. To these the rails are spiked, four spikes 
 being used on each tie. Care is taken that they are not all driven 
 near the centre of the tie, causing it to split. For all rails of 
 16 to 20 Ibs. per yard the ties are 5 inches on the face, 4 inches 
 deep, and 5^ feet long. For the heavier rails in the mine they 
 are 6"X$" deep. They are laid normally to track, and on 
 curves are laid in the direction of the radii of the curve. In 
 some mines the ties are creosoted to increase their durability. 
 
 The rails are the common steel T rails, having the same speci- 
 fications as are adopted for railroad service. Their weight 
 varies from 12 Ibs. per yard in the rooms to 40 Ibs. per yard in 
 the slopes. They cannot be too heavy within reasonable limits. 
 The greater the pitch of the slope and the greater the speed of 
 haulage, the larger and heavier should be the construction of the 
 track. The rail for locomotive haulage-ways requires a weight 
 of 10 Ibs. per yard for each ton that falls upon the drivers. The 
 cost of one mile of i6-lb. rails is about $1600. 
 
 The track in the room and at the long wall-face is laid as 
 rapidly as the mining operations progress, and with as great care 
 to secure a firm run as in the main ways. It is often laid in sec- 
 tions composed of two rails of 15 feet each. These sections are 
 portable and carried up to the face of the workings with the 
 progress of the room, the expense of such track-laying usually 
 being included in the price paid to the miner per ton of coal.
 
 262 MANUAL OF MINING. 
 
 Their advantage lies in their simplicity, strength, and durability. 
 The gauge does not depend on the skill of the workman, and 
 will be preserved as long as the line lasts. The lightest rail 
 advisable for such purposes is that which weighs 16 Ibs. to the 
 yard. Wooden rails are no longer employed even for tempo- 
 rary service in the rooms. 
 
 The Gauge of Track is rarely less than 30 inches or more 
 than 44 inches. Local conditions determine the choice. A 
 broad gauge gives a greater stability to the cars and a reduction 
 of haulage expenses, but is more costly to build. The minimum 
 gauge of 30 inches affords easy haulage, opportunity for sharp 
 curves, and a cheaper track. 
 
 Turnouts. The sidings are formed by a set of points and 
 crossings, either right- or left-hand, a section of curved line, and 
 a number of sections of straight line. 
 
 Switches from one track to another are obtained by the use 
 of movable points similar to those employed on the surface 
 roads, though shorter in length. The point is simply a spike 
 driven into one of the ties supporting the rails. It is held in posi- 
 tion by and turned upon an iron pin passing through the plate 
 welded to the base of this point and thrown out of line to the 
 branch side when the car is to be switched. When two of these 
 points are connected by a flat piece of iron passing under the 
 rails of the track, the pair can readily be thrown right or left by 
 a bell-crank lever, and both rails of the track are opened for the 
 switch. Wherever possible single sets of points are arranged 
 for, the wear being less on two pairs of single points than on one 
 set of double points, besides which single sets are simpler to lay 
 down and keep in order The use of the three-way points and 
 crossings for two opposite side branches are avoided underground 
 by turning the right- and left-hand lines from the straight track 
 in positions which are not quite opposite to each other. The 
 points and crossings are made of standard sizes for curves of any 
 special radius required to suit the gauge of the rails and the wheel- 
 base of the engines. All frogs and switches at the shaft-landings 
 or slope bottom are of pointed sections of standard rail sizes,
 
 UNDERGROUND HAULAGE SYSTEMS. 263 
 
 manipulated by a spring or a hand-lever. For locomotives on a 
 main tramway the curves should not, as a rule, have a radius 
 less than from 100 to 125 feet. 
 
 When one rail intersects another, frog crossings are used. 
 They are either plain rail intersections, broken at the points of 
 crossing, or they are the standard frogs furnished by manufac- 
 turers. The points, rods, and levers used underground are usu- 
 ally made and fitted by the colliery blacksmith, the plate-layer 
 assisting in the fitting. 
 
 In passing from a heading or entry to the room a "parting " 
 is used. This is similar to a switch, the wings running back- 
 ward from the frog end near the rails, but are cut in a slanting 
 direction in order to provide a sufficient clearance between the 
 point and the rail to allow the wheels of the car to pass along 
 the main track when it is not to be deflected into the room. There 
 are two forms of partings, one of rails and another a combination 
 of rails and cast-iron pieces, the latter being much inferior to 
 the former. The former has less pieces, gives more solid track, 
 and is less liable to derail the cars. It is cheaper and can be made 
 in the blacksmith-shop. 
 
 Turnouts are obtained by the use of plates or turntables. 
 Of the former the simple iron plates laid on stout planks 
 revolving around a vertical pivot are employed where animal 
 power is used. These are common in metalliferous mines, 
 being preferred to self-acting switches. Not infrequently is 
 seen at intersections of roads a large fixed iron plate, raised upon 
 suitable supports, level with the road-bed, and having curved 
 guide-rails at each of the four corners. These may be em- 
 ployed as turntables by dragging the car around on the plate 
 until it faces the desired track, when it is run off. This crude 
 device serves well, but has nothing to commend it. 
 
 Curves. Curves in the track are built in the same way as 
 on the surface, the radius of curvature, however, being very 
 much shorter. Connecting two entries on the main haulage- 
 road, the minimum radius for narrow-gauge tracks is 60 feet, and 
 for a gauge of 45 inches 100 feet. At the shaft-landings the
 
 264 MANUAL OF MINING. 
 
 radius is as short as 30 feet for 36-inch gauge, and 50 feet for 
 44-inch gauge. 
 
 The Degree of Curvature of a Curve. By this term is under- 
 stood the number of degrees of circular arc measured at the 
 centre which will be subtended by a chord of 100 feet length. 
 Thus a 10 curve is one in which a loo-foot chord at the circum- 
 ference is comprised between two radii making 25 at the center. 
 The radius of a one-degree curve is practically 5730 feet. Hence 
 the 10 curve will have a radius of 573.0 feet. This is only 
 approximately correct, but is sufficiently accurate as a standard of 
 measure for engineers. 
 
 All curves on underground roads are laid by transit with 
 care equal to that employed on the surface for larger work. The 
 central line between the rails of each track is staked out at inter- 
 vals of 20 feet, from which on the other side are laid down the 
 two rails. The latter are now bent to conform to the curvature 
 of the circle on which the rails are laid. 
 
 Elevation of the Outer Rail. Provision is made for neutral- 
 izing the effect of centrifugal force, or the pull of the rope, on the 
 curve by elevating one rail an amount depending upon its sharp- 
 ness. On curves used by locomotives the outer rail is the one 
 which is elevated, as also on the slope haulage-ways when the 
 weight of the cars draws the main rope off its drum. In a slope 
 haulage the outer rail is raised i inch on a 9 curve, 2 inches on 
 an 1 8 curve, and 4 inches on a 36 curve, the latter being the 
 maximum. It never exceeds 4^ inches in height on a curve of 
 50 feet radius, or of 114 curvature, when the gauge'is 42 inches. 
 This would be suited to the ordinary speed of locomotive haulage, 
 or that of cars running down the slope by gravity. 
 
 The inner rail is elevated above the outer one only on such 
 curves and under such conditions as those in which the action 
 of the centrifugal force is less than that of the pull of the deflected 
 rope. In this case the rope passes over rollers near the inner 
 rail and tends to pull the car inward. It is possible, by comparing 
 the value for centrifugal force under given conditions, as cal- 
 culated below, with that of the pull of the rope upon the car, to
 
 UNDERGROUND HAULAGE SYSTEMS. 265 
 
 determine the amount of the inclination which must be given 
 to either the inner or the outer rail. 
 Let g= gauge of the track in inches; 
 
 e = elevation of the outer rail in inches; 
 
 R= radius of track in feet; 
 
 v= velocity in feet per second; 
 
 D = degree of curvature of centre of track; 
 
 c= centrifugal force due to the speed around the curve. 
 
 Then e=' 3 l gV =o.ooooo$4V 2 Dg. 
 
 K 
 
 Conveniences at Landings. Whatever may be the system of 
 haulage, arrangements are made for the prompt handling of 
 cars at the shaft bottom. When the coal is hauled from the dip 
 by rope or locomotive, the slope is continued, if possible, on the 
 same side of the hoisting-shaft at a convenient distance from it, 
 where the requisite number of loaded cars can be accumulated 
 and allowed to run by gravity on a one per cent grade to the 
 shaft. 
 
 The empty car can be run off the cage on to a short down- 
 grade of about 4 per cent, and, with the momentum given it by 
 the loaded car being run on the cage, would descend far enough 
 to rise up on an opposing grade of about 2 per cent to an auto- 
 matic switch, which delivers it to the empty-car siding. Here 
 it may be taken up by the locomotive, or a rope passing around 
 the shaft through an entry designed for this purpose. Some- 
 times a short chain haulage operated by an electric motor is 
 employed to draw the empty cars to the main haulage-way when 
 the gravity system cannot be employed. 
 
 The landings for the cars in the mine at the end of the end- 
 less-rope haulage-way are reached by switches from the main 
 track by slide-knuckles, a place being allowed for the rope to 
 pass through the switch without injury from the flanges of the 
 wheels passing over it. 
 
 Safety Devices Along Haulage-ways. To prevent accident 
 from runaway cars on a grade, whether underground or at the 
 surface, automatic devices are employed either to check the speed
 
 266 MANUAL OF MINING. 
 
 of the car or to automatically shunt it off the track. Cars are 
 frequently fitted with a drag projecting below the car, which, when 
 the latter breaks away from the hoist-rope, catches in the floor of 
 the track and prevents the runaway. A similar device is employed 
 whereby a car trips a block which is thrown across the track 
 before the car reaches it and stops the car, preventing further 
 damage being done by its descent to the lower end of the plane. 
 
 Catches at the Top of the Incline. These are to prevent cars 
 from breaking away at the top of the plane. Reaumaux's catch 
 consists of a bent lever with two unequal arms turning around a 
 vertical axis. It is placed flat on the ground and is parallel to 
 the tramway and at such a distance that at least one of the ends 
 always bars it, with the short arm next the side of the inclined 
 plane. As the empty car reaches the top of the incline and passes 
 forward its front wheel touches the long arm after the hind wheel 
 has passed the short arm, and the way is blocked. Another run 
 on the same track cannot be made until the protruding arm is 
 turned back by the planeman. 
 
 Blocks or stops are simpler appliances. There are two pieces 
 of wood at right angles to one another moving on an upright pin. 
 One arm is thrown across the track to hold the car in place until 
 the latter is to be released and lowered (6, Fig. 84). Then it is 
 knocked out. The same plan is adopted in 10, Fig. 84, by the 
 weighted lever. 
 
 Stops Along the Incline. To guard against accidents from 
 runaways along the inclines a heavy iron bar is swung as shown 
 in 9, Fig. 84. It interposes no obstruction on the track until 
 some one pulls the wire and releases it, as shown by the dotted 
 lines. A similar balanced block, Fig. 84, is used, the car being 
 caught by the end, b. 
 
 The Mortier safety- catch consists of a movable axle with 
 levers placed in the axis of the roadway and supported upon 
 sleepers, 8, Fig. 84. In one extreme position it is opened by 
 the axle of the rising car; in the other, it closes after the train 
 passes. In the middle position both levers bar the passage 
 The planeman can also move the catch by a pedal.
 
 UNDERGROUND HAULAGE SYSTEMS. 267 
 
 Catches at the Bottom of the Incline. At the bottom of the 
 incline are additional safeguards. A "seizer," shown in 12, 
 Fig. 60, is placed at the foot of the incline. An iron bent lever- 
 arm supported on a horizontal axis at the apex projects vertically 
 above the rails between which the frame stands. A strip of iron 
 swinging horizontally keeps up one arm, which is inserted into 
 the link of the chain. Even if the top man should push his train 
 over the plane before the bottom train is ready, the chain would 
 not move until the upheld arm is released and allowed to fall 
 into a recess, and thus free the cars. Various other provisions 
 of a similar nature guard against the ill effects of a broken end- 
 less haulage-chain. 
 
 The Choice of Haulage Systems. The choice of a system 
 depends upon the grades of the haulage-ways. When the latter 
 are horizontal, power will be required in both directions, and 
 obtained from the horse, locomotive, stationary engine, or rope 
 haul. When the grade is toward the shaft, the loaded train may 
 be pulled with any of these powers The horse or mule is, 
 however, used to advantage only on a favorable track of less than 
 2.5 per cent. The locomotive may be employed for grades not 
 exceeding 4 per cent. Above that it is impossible for it either to 
 pull a long train up the inclination or to check its descent. The 
 endless-rope system or the tail-rope method is used when the 
 grade does not exceed 20 per cent. Beyond this inclination the 
 plant will be self-acting, for the loaded train going toward the 
 shaft pulls the empty cars back into the mine. When the grade 
 of the haulage-way is slightly against the loaded car, the horse 
 may do for a short haul. If it exceeds 2.5 per cent, the locomotive 
 can no longer be used advantageously. On a grade exceeding 
 5 per cent a traction locomotive is totally incapable of pulling any 
 load. Up to a limit of 10 per cent, a track locomotive may be 
 employed. For all steep inclines some rope system must be 
 employed. 
 
 Tramming. Man-power is employed for incidental work on 
 the short hauls on level ground, or in rooms. The weight of car 
 and degree of inclination for such power are necessarily very
 
 268 MANUAL OF MINING. 
 
 small. The average man is capable of exerting a push of 27 Ibs. 
 at the rate of 2 feet per second. This would move a 2-ton car on 
 a level at a speed capable of making from three to six round 
 trips of a ton per mile. The condition of the roads and running 
 gear will definitely fix this. 
 
 Animal Haulage. Horses and mules are employed under- 
 ground for limited purposes of transportation singly or tandem, 
 according to the length of the trip and the weight of the train. 
 It is customary to work them in traces; though in some mines 
 shafts are used which give considerable holding-back power not 
 to be had from traces. There can be no objection to traces 
 on level or nearly level roadways. But with loaded cars down- 
 hill sprags must be placed in the wheels so as to act as a drag 
 for trace- mules. The inclination of the roadway for animal 
 haulage power is not over 1.5 per cent, as the mules rarely 
 take the cars to the surface unless the inclination of the entries 
 especially permits it. 
 
 The utility of the animal is confined to haulage in the secondary 
 ways, to the rooms, and to switching where economy in height must 
 be practised. Here the mule has the superiority over the horse 
 because of its shorter stature with equal strength. A thickness 
 of coal-seam of 3.5 feet will accommodate the mule, while one 
 of 4.8 feet is essential for the horse. The kind and size of horse 
 purchased for colliery working will of course depend upon the 
 specific work for which it is required and the nature of the road- 
 ways, size of cars, etc. Before being introduced into the pit they 
 are worked on the surface for three or four weeks to reduce the 
 risk of importing infectious diseases into the underground stud. 
 The possibility of new horses being up to the work expected of 
 them may thus be ascertained. 
 
 The ordinary speed of the animal is two miles per hour. It 
 has a tractive force of about 150 Ibs., at which a single horse 
 or mule may be estimated as an equivalent to from 8 to 9 gross 
 tons of load per hour, averaging four round trips of a mile each 
 per day. The condition of the track, the ventilation, and the 
 amount of delay at the termini of the trip would reduce this
 
 UNDERGROUND HAULAGE SYSTEMS. 269 
 
 average and make the daily capacity of the horse or mule about 
 4.0 tons. This represents a gross yearly capacity of 9000 ton- 
 miles. Under ordinary conditions one mule will serve for the 
 haulage of the output of ten miners. A mine of 1000 tons daily 
 output will require 90 mules, which with equipment will cost 
 $8000. 
 
 The average cost of mules, as well as their keep, is nearly 
 the same as for horses. Their care averages 60 cents per day, 
 inclusive of feed. The daily allowance of food varies with the 
 work done. It is not far from 18 Ibs. of grain and 12 Ibs. of 
 hay per animal. A less economical fodder now used is " ensilage." 
 This is fodder cut in the field in the usual way, but, instead of 
 being dried, is chopped up small and stored green in brick or 
 cement-lined pits, or silos. Dry food is a more convenient feed, 
 but dearer than ensilage. 
 
 The number of horses under one ostler or keeper should 
 not exceed ten. They are generally stabled underground, and 
 many, once below, never see daylight again. Where the work- 
 ings are at a considerable distance from the shaft, the stables 
 are built well back in the interior of the mine, to avoid unneces- 
 sary travelling. The cost of a comfortable stable is about 
 Sio per stall. Special attention must be given to proper venti- 
 lation, drainage, and cleanliness. The opening must be not 
 less than 18 feet wide. 
 
 Animal haulage is expensive and is fast being supplanted 
 by mechanical systems. The number of working days at a 
 colliery seldom exceeds 200 in a year. The cost of maintaining 
 the animal during the remaining 165 days, besides those result- 
 ing from depression in trade, strikes, accidents, and diseases, 
 adds materially to the cost of production. Moreover, horses are 
 subject to epidemics, and it may happen that in times of greatest 
 prosperity some or all of the horses may be unable to leave the 
 stables. Delays arising from the dropping of shoes while at 
 work are costly, for, while the farrier is at work, both horse and 
 driver- are standing idle. Out of every thirteen animals in the 
 mine an average of ten may be regarded as a full working
 
 270 MANUAL OF MINING. 
 
 complement, the remainder being on the retired list for various 
 causes. 
 
 Compared with the mechanical methods, animal haulage 
 is more expensive. It is slower and therefore used on long hauls 
 only when numerous ventilating-doors intervene along the line. 
 
 The Locomotive. For great distances and a large output 
 the locomotive is much more efficient, more rapid, and cheaper 
 than animal power, being equally flexible and having, moreover, 
 the advantage in that during strikes and lockouts there is no 
 expense in maintaining it. It can be accommodated to varying 
 demands of the different sections of the mine. The motor is in 
 direct charge of the engineer, who is always on hand, so that 
 it can be stopped promptly. This would avert such accidents 
 as are common in the rope-haulage systems. Here the con- 
 ductor of the train must first signal to the surface before the 
 train can be stopped, and any incumbrances on the track, or 
 the jumping of a car from the track, would result in demolition 
 of the entire train before the rope could be brought to rest. 
 
 The locomotive is operated by air, steam, or electricity with 
 nearly equal efficiency and in any desired size for traction pur- 
 poses. It is difficult to estimate the number in use in the United 
 States, but in Pennsylvania alone are 950 locomotives, 503 being 
 steam-locomotives, 71 operated by compressed air, and 376 elec- 
 tric. The anthracite region uses mostly the steam-locomotive, 
 while the electric motor is more common in the bituminous 
 district. 
 
 All traction-locomotives, depending solely upon the adhesion 
 of their drivers to the rails, are limited to grades of 4 per cent. 
 They are not economical on steeper grades, except for short dis- 
 tances and where the up-grade is approached by a down-grade, 
 enabling the train to acquire momentum enough to help it up the 
 steeper grade. - 
 
 The Haulage Capacity of a Locomotive. This is measured 
 by the weight of the train which it can haul upon a level straight 
 track. Its tractive force, or its draw-bar pull, is the tension exer- 
 cised by the locomotive upon its first connection with the train.
 
 UNDERGROUND HAULAGE SYSTEMS. 271 
 
 The draw-bar pull must exceed the total frictional resistance of the 
 train upon the track, including that between the wheels and the 
 rail, the resistance of the curves and of the grade. The draw- 
 bar pull is about one eighth the weight of the locomotive. Hence 
 the weight of a locomotive should be the maximum for given 
 minimum dimensions. 
 
 The Tractive Force. The tractive force, T, of a steam- or 
 air-locomotive, expressed in pounds, is measured by the formula 
 dT = i2k 2 sp, wherein d is the diameter of the driver in feet, and 
 k, p, and s have values as in Chapter V for steam-engines. The 
 traction of a locomotive, expressed in pounds, must not be con- 
 fused with its horse-power, which is a unit of dynamic force, 
 embracing the elements of weight, distance, and time. 
 
 The train resistances R are : (i) the frictional, which in mines is 
 not less than 30 Ibs. per ton of train, and equals 30 Y, Y being the 
 number of tons weight of train and load ; (2) that due to grade, 
 which is 2ogY, g being in feet per 100; (3) that due to the curve, 
 which is ^ Ib. per ton per foot width of gauge, z, per i curvature; 
 and 4, that due to speed, which is 0.257+3, V being the speed in 
 miles per hour. For any degree of curvature, D, and weight, w, 
 of cars which are at one time on the curve, the curve resistance is 
 \wzD. As the radius of a i curve is 5730 feet, then, when the 
 
 radius of the curve, p, is known, the curve resistance =7 16.2 
 
 Often the equivalent grade resistance is calculated; that is, the 
 percentage of grade which offers the same resistance as the 
 curve. This = 1.32!}. 
 
 EXAMPLE. A locomotive has a draw bar pull of 2400 Ibs. How many 
 tons will it pull up a grade of i in 50 if the friction is 32 Ibs. per ton? 
 The work against gravity is 1 750, or 45 Ibs. per ton, making a total resistance 
 of 77 Ibs. The available pull being 2400, the gross load is 31 tons, less the 
 weight of the locomotive. 
 
 The influence of grade in reducing the haulage capacity of 
 the locomotive is shown in the following table. For any other 
 weight of locomotive the capacity is directly proportional.
 
 272 MANUAL OF MlNIXG. 
 
 HAULING CAPACITIES OF IX>COMOTIVF.S ON VARIOUS GRADES, IN TONS OF 2000 LBS. 
 
 
 
 Frictional 
 
 Grades. 
 
 
 Draw-bar 
 
 Car Resist- 
 
 
 Weight. 
 
 Pull on 
 Level. 
 
 ance per 
 Ton on 
 LeveL 
 
 Level. 
 
 HPer 
 Cent. 
 
 i Per 
 Cent. 
 
 iHPer 
 Cent. 
 
 2 Per 
 Cent. 
 
 2^Per 
 Cent. 
 
 
 ( 
 
 20 
 
 23 
 
 15 
 
 1C 
 
 8 
 
 6-3 
 
 5-2 
 
 4,000 
 
 500 ] 
 
 3 
 
 IS 
 
 II 
 
 8.4 
 
 6. 7 
 
 5-4 
 
 4-5 
 
 
 ( 
 
 40 
 
 12 
 
 9 
 
 7.0 
 
 5-7 
 
 4-7 
 
 4.0 
 
 
 
 Frictional 
 
 Grades. 
 
 
 Draw -bar 
 
 Car Resist- 
 
 
 Weight. 
 
 Pull on 
 Level. 
 
 ance per 
 Ton on 
 Level. 
 
 Level. 
 
 Cent. 
 
 3^ Per 
 Cent. 
 
 Cent. 
 
 5 Per 
 Cent. 
 
 6 Per 
 Cent. 
 
 
 ( 
 
 20 
 
 23 
 
 4-2 
 
 3-5 
 
 3-0 
 
 2.2 
 
 1.6 
 
 4,000 
 
 5M 
 
 3 
 
 15 
 
 3-8 
 
 3-2 
 
 2-7 
 
 2.0 
 
 1-5 
 
 
 ( 
 
 40 
 
 12 
 
 3-4 
 
 3-o 
 
 2-5 
 
 1.8 
 
 1-4 
 
 Comparison of Types of Locomotives. At first the steam- 
 locomotive was used underground, but the objections to it, which 
 were soon recognized, led to its removal from gaseous mines or 
 from those whose ventilating system may become disturbed by the 
 passage of the locomotive through the airway, and it has been 
 supplanted by the later forms of motor-power. 
 
 The compressed-air locomotive is the only cheap form of 
 haulage for gaseous mines. It introduces no element of dan- 
 ger other than that common to all large haulage motors the 
 interference with intake air-current during the outward passage 
 of the locomotive. 
 
 The electric locomotive has now established itself in coal- 
 mines, and may be regarded as a safe form of haulage except 
 in special cases. It may be had of three different types the 
 overhead-wire trolley, the third-rail, or the rack-and-pinion when 
 the grade is excessive for the simple traction-engine. 
 
 On the other hand, a very decided objection obtains against 
 the air- and the steam-locomotive, and to a less degree against the 
 electric locomotive, in the area of the haulage-way which they 
 occupy. Unless the latter is excavated to a large area the passage 
 of the locomotive interferes markedly with the air-current. When
 
 UNDERGROUND HAULAGE SYSTEMS. 
 
 273 
 
 -
 
 274 MANUAL OF MINING. 
 
 travelling with the current, at least 20 per cent more air enters 
 the mine than when the train is moving out against it. This 
 may be averted by disconnecting the locomotive haulage-way 
 from the main ventilation system of the mine and introducing 
 the inlet current at a point beyond the inside terminus of the 
 locomotive gangway. The cost of these improvements, how- 
 ever, is very large. 
 
 The Steam-locomotive has not as ready acceptance in under- 
 ground work now that it had in former days, because of the 
 objection raised to the presence of the gaseous products which 
 it discharges into the airway. Whether coal-burning or oil- 
 burning, there is given off a considerable percentage of car- 
 bonic oxide gas as well as smoke. The former is a menace, and 
 the latter befouls the ventilation besides introducing the risk of 
 fire. Moreover, the great heat and moisture in the exhaust 
 tend to rot the timbers and to soften the roof. These are about 
 the same objections that obtain against any underground steam- 
 engine. In long haulage-ways steam-locomotives can be used 
 only when a special means of ventilation is provided. There are 
 many instances where lives have been lost by the inhalation of 
 the gases from the locomotive and underground engines. The 
 steam-locomotive, however, is less expensive than the air-engine 
 and very much lighter in weight. 
 
 The capacity depends on the driver adhesion, and a steam- 
 locomotive of 6"Xio" cylinders will haul 28 tons of train at a 
 rate of twenty miles per day up a grade of 105 feet per mile with 
 600 Ibs. of coal. One of 10" Xi4" cylinders has a duty of 46 
 tons per hour at 28 miles speed on grades of 52 feet per mile. 
 The coal consumption is 1000 Ibs. per day. 
 
 The locomotives are made of a shape to suit the mine open- 
 ing, for narrow gauge (36 to 40 inches) rarely over- 78 inches 
 high, have four to six wheels (for curves of 50 to 75 feet radius), 
 weigh 4 to 13 tons, and carry 125 to 350 gallons of water. Their 
 cylinders are from 5 Xio to 10X14, on 22-inch to 28-inch drivers 
 running over 16- to 28- Ib. rails. These engines cost from $2600 to 
 $4000. They have a traction of from 150 to 600 tons on a level.
 
 UNDERGROUND HAULAGE SYSTEMS. 275 
 
 There is no difference in the price between the wide- and the 
 narrow-gauge locomotive of the same design and size of cylinders. 
 For narrow tunnels locomotives with inside cylinders may be had. 
 The Compressed-air Locomotive. This consists of one or 
 two storage tanks, in which is retained compressed air sufficient 
 in amount for the entire run of the trip, and two cylinders with 
 the usual driver connections. When the air-pressure is as high 
 
 FIG. 112. A Mine Locomotive. 
 
 as 1000 Ibs. per square inch, the high-pressure tank, placed 
 along the side of a larger tank, is quite small, with a reducing- 
 valve between them. The cylinder receives air from the low- 
 pressure tank. As the radius of action depends on the capacity 
 of the tank, the latter is double that calculated to be required for 
 the trip. The cars are hauled to the foot of the slope, or, if the 
 grade is a natural one, all the way to the tipple. The locomotive 
 goes as far into the mine as the height of the seam and the local 
 conditions will admit. 
 
 These locomotives are operated precisely like the steam- 
 locomotive, but are of greater weight, with larger bearings, frames, 
 cylinders, wheels, etc., though the general pattern is much the 
 same. The early objection to the air-engines the fear of ex- 
 plosion is no longer warranted. Their dimensions are about 
 as follows: 16 to 20 feet in length, 6 to 7 feet in width, and 
 5 feet in height. The cylinders are not lagged as are steam- 
 cylinders; indeed their exterior surface is corrugated to increase 
 the exposure.
 
 MANUAL OF MINING. 
 
 The haulage capacity is calculated in the same manner as for 
 steam-locomotives, the percentage sometimes allowed for adhesion 
 of drivers to the rails being one fourth. Their tractive efforts 
 are calculated in the same way; after ascertaining the mean 
 effective pressure from a given ratio of expansion and from the 
 size of the cylinder, the volume of air required to deliver the given 
 power at a known rate of speed is found. In determining the 
 mean effective pressure, the second table in Chapter X will be 
 found convenient for conditions in which the air is at a pressure of 
 not to exceed 200 Ibs. per square inch. Above that the formulae 
 may be employed for the purpose. 
 
 The following table shows the tractive effort of certain air- 
 locomotives. 
 
 TRACTIVE EFFORTS OF COMPRF>SSED-AIR LOCOMOTIVE. 
 
 Cylinder. 
 
 Diameter 
 
 Weight 
 
 Tractive Effort for each loo-lb. Gauge Cylin- 
 der Pressure at Various Cut-offs, for 
 
 
 
 Driver, 
 
 Driver, 
 
 
 
 Stroke, 
 
 
 
 
 
 
 
 
 
 
 Inches. 
 
 Inches. 
 
 
 
 H in., 
 
 ^in., 
 
 Kin., 
 
 Hin., 
 
 ?*in., 
 
 y 4 in.. 
 
 Hin., 
 
 
 
 
 
 0.98 
 
 0.95 
 
 0.88 
 
 0.80 
 
 0.68 
 
 0.51 
 
 0.31 
 
 5 
 
 10 
 
 24 
 
 6,000 
 
 1020 
 
 990 
 
 920 
 
 835 
 
 710 
 
 530 
 
 325 
 
 6 
 
 IO 
 
 24 
 
 8,500 
 
 1470 
 
 1425 
 
 1320 
 
 I2OO 
 
 IC2O 
 
 760 
 
 445 
 
 7 
 
 12 
 
 26 
 
 13,000 
 
 2200 
 
 2150 
 
 1990 
 
 1810 
 
 1540 
 
 1140 
 
 700 
 
 8 
 
 12 
 
 26 
 
 18,000 
 
 2880 
 
 2750 
 
 2600 
 
 2360 
 
 2000 
 
 1510 
 
 900 
 
 9 
 
 14 
 
 26 
 
 25,000 
 
 4340 
 
 4140 
 
 3840 
 
 349 
 
 2960 
 
 2220 
 
 135 
 
 10 
 
 14 
 
 26 
 
 32,000 
 
 S28o 
 
 5150 
 
 4740 
 
 431 
 
 3660 
 
 2630 
 
 1670 
 
 ii 
 
 16 
 
 28 
 
 42,000 
 
 6770 
 
 6450 
 
 SQo 
 
 5440 
 
 4620 
 
 347 
 
 2440 
 
 12 
 
 16 
 
 28 
 
 52,000 
 
 8050 
 
 7800 
 
 7200 
 
 6550 
 
 5580 
 
 4150 
 
 2850 
 
 Aside from its low efficiency, the main objection to the com- 
 pressed-air locomotive is in the great dimension of its tanks. 
 The exhaust is laden with snow if the ratio of expansion has been 
 high, for the air is usually filtered through water previous to com- 
 pression, to remove the dust. This charges it thoroughly with 
 moisture, which causes the trouble. 
 
 Compound pneumatic locomotives are being introduced in 
 order to utilize the expansive effort of the air. They have an 
 advantage over the simple locomotive of lower piston speed for
 
 UNDERGROUND HAULAGE SYSTEMS. 277 
 
 the same consumption of air. A higher initial pressure is then 
 used. 
 
 The -cost of a compressed-air plant, with motors, boilers, etc., 
 is $33,000 for a mine having an output of 900 tons per day with 
 a mean haul of 2000 feet. The cost of haulage per ton mined is 
 12.6 cents. 
 
 The storage tanks on the locomotive are charged at either end 
 of the line from stations where large storage tanks are located, 
 which latter are in pipe communication with the air- compressor. 
 The capacity of these stationary tanks exceeds twice the capacity 
 of the locomotives comprising the system. 
 
 FIG. 113. A Pneumatic Locomotive. 
 
 The Electric Locomotive. This is the most advantageous 
 form of electric application for underground work both in point 
 of economy and convenience. Its introduction has been of slow 
 growth in coal-mines because of the conservatism enforced upon 
 engineers by adverse legislation, which has always compelled 
 them to go slow in the introduction of any new form of machinery. 
 Since its first installation it has established itself and is more 
 popular than the compressed-air motor, because it is more com- 
 pact and more economical. Its compactness enables it to serve in 
 entries too small for compressed air by engines and too sinuous 
 for cable-ways. The results from its introduction have not proved 
 any more destructive or injurious to life than from other forms of 
 haulage power, and it is used in any gangway where the naked 
 lamp can be carried. It possesses the same advantage as other 
 locomotives in having a practically unlimited radius of action, and 
 has an additional advantage in being cheaper to operate than its
 
 278 MAXUAL OF MINING. 
 
 competitive traction machinery, having nothing to get out of 
 order. Because of this little trouble ensues. 
 
 Two systems of electric locomotives are used; one with over- 
 head-trofley wire and the other with the third rail. 
 
 The mechanism of the underground locomotive is similar in 
 kind to that of the familiar street-car trolley in which hard-drawn 
 copper wire is laid along the roof of the road. This form of 
 the overhead wire on which the trolley of the locomotive arm 
 rests conducts the current to the motor, whence it may return by 
 the rail to the generator. 
 
 The third- rail locomotive has a rack between the two main 
 rails in which engage two sprocket-wheels driven by the elec- 
 tric motor by means of suitable gearing. These sprocket- 
 wheels serve the double purpose of driving the locomotive 
 along the track and taking the electric current from the third 
 rail to feed the electric motor. These machines are of recent 
 introduction. 
 
 Locomotive Details. It is impossible to indicate the average 
 dimensions of the locomotive, because each one represents a 
 special type designed for local conditions only. Any desired 
 height, power, or gauge may be 'procured. Its height is rarely 
 more than 4 inches above the tops of the wheels; its length is n 
 feet, and width 57 inches, for a gauge of 36 inches. Its average 
 weight is 10 tons. This would have 50 horse-power capacity 
 and would be capable of taking 25 cars of 4.5 tons gross weight 
 each on the level. 
 
 The driving motors are made as small as is consistent with 
 good design, and are suspended as low as safety warrants in the 
 space between the wheels. They combine both high power and 
 good speed; their mechanism is of simple character and easily 
 accessible. A housing affords protection from dust or injury. 
 They are enclosed in a heavy cast-iron frame firmly bolted 
 and carrying axle-boxes. To each axle a steel-clad motor is 
 geared. The gears are of the best cast steel and designed for 
 maximum strains which are sure to occur when, in the darkness 
 of the mines, it may become necessary to reverse from full speed.
 
 UNDERGROUND HAULAGE SYSTEMS. 279 
 
 Sand is carried in boxes, fore and aft, fitted with spouts in front 
 of each wheel; the brake is designed to be applied by one hand. 
 The screw-brake is preferred to gear or ratchet. 
 
 The wheels are generally of chilled iron, 30 inches in diame- 
 ter, and the trucks of the prevailing gauge. Steel tires are more 
 expensive, but furnish a greater tractive effort than chilled 
 wheels. The length of wheel-base is governed by the existing 
 curves. This is 5 feet for a gauge of 36 inches and a radius of 
 25 feet. 
 
 The trolley-pole is reversible and adjustable to maximum as 
 well as minimum heights. It may be placed on either side or 
 end of the car as conditions require, the centre of the socket 
 being over the gauge line and on the ditch side, to reduce the 
 element of danger. The motor is of the slow-speed four-pole 
 type with windings suited to the given line pressure and a speed 
 of eight miles per hour, taking into consideration the line loss. 
 If it happens that the line loss is less than was assumed, the loco- 
 motive speed will exceed this. 
 
 The Rheostat. The controlling device is of the five-point 
 rheostatic type of electric railway motor, so designed as not to 
 heat to the danger-point when using the maximum current. This 
 form of diverter, or rheostat, is necessary because of the great 
 variation in load and grade which the locomotive encounters in 
 a run. 
 
 The resistance-box is placed in front, where it will receive a 
 strong circulating current of air. Powerful magnetic blowouts 
 are always provided to insure against sparking. 
 
 The Draw-bar Pull of an Electric Locomotive. This is 
 about 50 Ibs. per horse-power of the motor, and the weight of the 
 locomotive at the ordinary speed of running is about 400 Ibs. for 
 each horse-power of motor, making the tractive effort of an elec- 
 tric locomotive about one eighth of its weight. For continuous 
 work the effective tractive effort should be assumed as 20 per 
 cent less than this rated capacity. This will prevent overheating 
 of the wires. 
 
 The average bar pull, P, of different weights, W, of loco<
 
 280 
 
 MANUAL OF MINING. 
 
 motives expressed in pounds on varying grades, g, is expressed 
 as follows: 
 
 P = W (0.125 g) = horse- power of locomotive X (50 4^). A 
 ten- ton locomotive can pull 1900 Ibs. on a 3 per cent grade or 
 1300 Ibs. on a 6 per cent grade, as against 2500 Ibs. on a dead 
 level. The starting draw-bar pull is nearly 60 per cent greater 
 than the ordinary running rate when sand is used. 
 
 Locomotive Rating. The length of time for which a motor 
 can give a stated ampere rating, 7, is based on the expression tP > 
 
 II- 
 
 ELECTRIC MINING LOCOMOTIVE CURVE 
 BOOVOITS 
 
 \ 
 
 
 FlG. 114. The Power Curves of a Locomotive. 
 
 in which t is the number of minutes of running, and 7 the number 
 of amperes for regular work. Every overload causes heating 
 beyond the allowed limit. This varies with the square of the 
 current. The rating of a given locomotive being known, ft is 
 possible thus to ascertain its hauling capacity.
 
 UNDERGROUND HAULAGE SYSTEMS. 281 
 
 Thus a motor rated at 50 amperes for 18 minutes can work 
 only 4| minutes at 100 amperes. The length of haul also has 
 an influence on the capacity of the motor. One intended to pull 
 a maximum load for a long period must have a greater capacity 
 than one which is operating for short hauls, allowing time to 
 cool between stops. 
 
 Notation. A system has been adopted for notation of elec- 
 tric locomotives similar to the Whyte system for steam-locomo- 
 tives. The number of wheels under the locomotive, the number 
 of motors and horse-power of each, and the number of pairs of 
 drivers geared to the motor are indicated in the order named. 
 The letter indicates the number of drivers: B, for one-geared 
 axle; C, two pairs of drivers; D, three axles geared to motors. 
 4/o-C means 4 wheels, 2 motors of 50 H.P. each, and 2 geared 
 axles. 
 
 Storage-battery Locomotives. Owing to the numerous acci- 
 dents from contact with the wires, whereby the mule becomes an 
 unnecessary evil in the operation of the mine, the animal may be 
 eliminated from the haulage systems by employing manual 
 labor for pushing the car from the parting into the entry. A com- 
 bined storage and trolley locomotive has also been devised for 
 the purpose and employed with success. Storage batteries obvi- 
 ate the necessity for wires, but as yet are too expensive and also 
 too heavy for efficient use. A 25-lb. battery will furnish one 
 horse-power hour; and a space of 40 sq. ft. will accommodate 250 
 elements, which would be able to give n horse-power during a 
 full ten-hour shift. This weight and the adhesion is more than 
 twice that of the steam-locomotive of equal horse-power, but 
 requires a heavier rail. 
 
 EXAMPLE. It is desired to move sixty cars of a capacity of 3 tons each and 
 weighing i ton over a track having the following profile, favoring the loaded 
 cars: 1920 feet of 1.3 per cent grade, 1000 feet of 1.8 per cent grade, 600 feet 
 of 2 per cent grade, 400 feet of 2.5 per cent grade, 425 feet of 3.5 per cent 
 grade, and 1000 feet of 0.9 per cent grade. The locomotive rating is 64, 
 voltage 500, amperes 32 for each motor, and speed 6 miles an hour. 
 
 The maximum resistance is on the steepest grade, or 4^ per cent. 
 
 If the car friction be assumed at i per cent, the cars will require 5400 Ibs
 
 282 MANUAL OF MINING. 
 
 draw-bar pull to do this work. To provide for the friction of the locomotive 
 itself, which is 1280 Ibs., the total tractive effort must be 6680 Ibs. With 
 cast-iron chilled wheels the weight of the locomotive must be 53,280 Ibs. 
 Dividing the draw-bar pull by 50 shows that the locomotive must be equipped 
 with io8-horse-power motors to do the work without overload. For con- 
 tinuous work we have, for the two motors, 
 
 P 
 
 - = 2X32. 
 
 T=the total time that should be taken for each trip. Calculating for the 
 several sections the time of haul in minutes and the required amperes, 
 respectively, as below, with a voltage of 500: 
 
 Time = 3.8 2 i.i 0.8 0.8 2 minutes 
 
 Amperes = 114 128 134 148 182 100 
 
 and tP = 41,400 32,768 19,751 17.523 26,500 20,000 
 we have // 2 = 165,942; (64)^=165,942; and ^=40. 
 
 Cost of Electric Haulage. The life of the average electric- 
 haulage plant is estimated as twenty years, and hence in dis- 
 tributing the cost of haulage an allowance of 5 per cent depreci- 
 ation per year divided by the number of working days per year 
 will give the allowance to be made for this item per day. 
 
 Comparing some instances of the cost of electric haulage with 
 that of mules, the evidence preponderates in favor of the former. 
 On the item of track cost alone it shows a saving of 25 per cent 
 on the investment per year. 
 
 Self-acting Planes. Conditions occasionally occur when 
 gravitation can be utilized to deliver the loaded car to the bottom 
 of the plane. A double line of rails is laid the entire length of 
 the track for two trains of cars. A rope connects them and 
 passes around the sheave at the top of the incline fitted in a recess 
 above the landing where the cars are started. In some cases a 
 single rope is used, the sheave being horizontal and its axis ver- 
 tical; in other cases a drum, whose axis is horizontal, carries two 
 ropes wound in opposite directions, each one having one or more 
 cars attached to it when desired. A single rope may be used ori 
 the drum with five or six extra coils upon the drum between 
 the two branches of rope to give friction enough to prevent the 
 rope from slipping. This form of plane is used in haulage- ways,
 
 UNDERGROUND HAULAGE SYSTEMS. 283 
 
 and occasionally in individual rooms for the delivery of one or 
 more cars at a time. 
 
 Sometimes three rails are laid with a suitable turnout mid- 
 way along the incline. This, however, is false economy, as com- 
 pared with the full four- rail line. 
 
 The lowest gradient at which the plane is self-acting depends on 
 the state of the road and the load, but under the most favorable 
 conditions the gradient requires to be about 1.25 inches per yard, 
 or i in 29. At about 6 per cent pitch it works freely. The most 
 satisfactory grade is i in 6. When the grade is as large as i in 3 
 there is an excess of motor force above the resistance, and a suitable 
 brake must be supplied to control the speed. Some inclines have 
 a counterbalance-car. The average safe velocity on such a plane 
 is 400 feet per minute. A two-car plane on 10 incline requires a 
 f -inch rope. On a 45 slope the rope is f inch ihxdiameter for 
 three cars. 
 
 The lower portion of the inclination is flattened somewhat to 
 diminish the momentum of the descending cars, and in some 
 cases a reverse grade is built in order to check the train. The 
 speed of the car or train is regulated by a brake applied on the 
 band bolted over the sheave- wheel. It. is of ihc usual type, fitted 
 with good durable surface having a high coefficient of friction. 
 The safety attachments provided on cars are essentially con- 
 fined to some form of trailing-block. 
 
 Single-track self-acting inclines for single cars have a trailing 
 counterbalance-car to regulate the speed, with a turnout at the 
 middle of the line. 
 
 Engine Planes. When the slope is against the load and a 
 locomotive cannot be employed to advantage, engine planes are 
 used. A stationary engine with drum is located at the head of 
 the plane to raise the cars in the train at any speed which the 
 timbering of the roadway will permit. If the traffic is large 
 enough to warrant it, the engine is in continual motion, the drum 
 being thrown in or out of gear when desired. If the drum may 
 turn freely, it will pay out the rope for the descending cars, and is 
 geared to pull them up. On a grade of 1.7 per 100, gravity will
 
 28 4 
 
 MANUAL OF MINING.
 
 UNDERGROUND HAULAGE SYSTEMS. 285- 
 
 take the loaded cars down with a reasonable velocity (empties on 
 a 2.25 grade), pulling the rope behind them. On a 10 per cent 
 grade a brake will be necessary for the empty cars. 
 
 The track may be single or double, according to the output 
 desired. On single-track planes the engine is non-reversing.. 
 The cars usually travel in trains, ten to thirty cars to a train, each 
 in charge of a conductor, who operates a dead-fall timber block to 
 hold the train while the cars are being shunted. The size of 
 the engine can be calculated as for hoisting, by determining the 
 amount of power requisite to overcome the vertical component 
 of the weight of the rope and train and the horizontal resistance 
 of the friction. The engine plane is eminently suited to an oil- 
 engine or an electric installation. It is well adapted for the 
 delivery from side entries at different levels, and may be used on 
 slight curves by curving the rope on iron guide-wheels. Ordi-. 
 narily the rope will last four years. A 14" X3o" engine with a 
 3-ton fly-wheel, T \-inch steel rope on a plane 4600 feet long and 
 80 feet rise, has a daily output of 950 tons of mineral in trains of 
 25 to 30 cars. 
 
 Rope-haulage Systems. The two classes of rope-haulage 
 systems are known as the tail-rope and the endless rope. Both 
 are extensively used in horizontal seams, and while each has its 
 special adaptation to a given grade, local conditions may make 
 it possible of application under steeper grades. The tail-rope 
 can be employed for grades not exceeding 3 in 100, either with 
 or against the loaded cars. 
 
 The Tail-rope System. This consists of a stationary engine 
 with two drums which are thrown alternately in and out of gear. 
 From each drum is carried a rope. The main-rope has a length 
 equal to that of the road, and is fastened to the front end of the 
 train. The tail-rope, of double this length, passes from the drum 
 around a sheave at the end of the road and up to the rear end of 
 the train. 
 
 When hauling the loaded train out, the main- rope drum is 
 thrown into gear and that carrying the tail-rope is free. The 
 loaded cars are then driven to the outlet or the point of distribu-
 
 286 
 
 MAX UAL OF MIXIXG.
 
 UNDERGROUND HAULAGE SYSTEMS. 287 
 
 tion, dragging behind them the tail-rope. When the train has 
 been replaced by empty cars, the main- drum is released and the 
 tail-drum engaged. The engine is started and the cars are de- 
 livered into the mine. The lower sheave is a grip-pulley similar 
 to that used in the power transmission, Fig. 118. The main- 
 rope is rarely over f inch in diameter. The tail-rope is usually 
 an old discarded main-rope. Swivel connections are used from 
 the rope to the train. 
 
 The roadway is usually a single track, and the sheave is ver- 
 tical with the tail-rope carried along the roof, while the main- 
 rope is supported along the floor. The length of haul is limited 
 only by the engine power and the track resistances. The most 
 advantageous pitch is about 3 per cent outward from the workings. 
 The velocity of haul may reach as high as 10 miles an hour. 
 Each trip may take a train of from 10 to 50 cars, usually ac- 
 companied by one trailer in charge of a conductor. 
 
 This system is very much used in American collieries. It 
 is the best plan for operating branch ways. Each branch has 
 its own rope passing over sheaves at the ends. When a train is 
 to be hauled out of the room to the point of discharge, the principal 
 ropes are opened at the proper points and connections with the 
 branch rcpes are made; meanwhile the other branches are idle. 
 This is an inexpensive plant to build and to maintain. It is 
 advantageous for undulating roadways and has very largely 
 replaced animal power. 
 
 An engine of i8"X3o" cylinder with 75 Ibs. steam-pressure 
 draws thirty trains of 17 cars each in a day over a slope of 2800 
 feet long with a grade of i in 200. The drums are 4.5 feet in 
 diameter, the main-rope is f inch and ihe tail rope \ inch diame- 
 ter. The velocity of travel varies between 8 and n feet per 
 second. 
 
 A tail-rope p'ant for 1000 tons daily for a mile of haulage 
 will require about 130-!!.?. boiler and engines; 6000 feet of 
 main-rope i inch diameter; 16,000 feet of f -inch tail-rope; 600 
 rollers; 500 cast-iron sheaves; 10 return sheaves and drums. 
 The first cost would be nearly $7500, and the maintenance 
 about $20 per day.
 
 288 MANUAL OF MINIXG. 
 
 The Endless-rope System. This system is much in vogue. 
 A rope or a chain receives continuous motion in one direction 
 from a wheel or drum, and draws the cars out on one line and in 
 on the other. The tension requisite to carry the load is arti- 
 ficially produced by a carriage provided with some form of coun- 
 terpoise. It requires a double-track line for communication 
 and considerably less rope than in the tail-rope system. Cars 
 can be attached at any point on the line, but branch connections 
 are not possible with this system. 
 
 The wire rope or chain cable may rest on the top of the car 
 or be supported below it. In the cheapest and most universal 
 method the chain rests in forks riveted on the top of each car. 
 For uniform grades and sharp curves the endless rope may be 
 suspended above the cars, to which it is attached by short lengths 
 of chain. If the chain cable is underneath the cars, it runs on 
 rollers, connections being made by hooking a short length of 
 chain from each car. With a rope running over rollers or sheaves 
 on the floor, like the surface cable roads, connection is made by 
 Si hand-grip attached to the car. 
 
 The engine of the double-track system is in continual rotation, 
 ^driving a shaft fitted with a heavy fly-wheel, and with a friction- 
 clutch for emergencies. From the thaft a belt or rope drives a 
 .parallel shaft with the sheave about which the endless rope 
 travels. A similar sheave at the far end of the line, with a 
 balance-car or tightener, constitutes the entire mechanism. The 
 speed of the rope rarely exceeds four miles per hour. 
 
 The cars are attached singly at the landings as fast as they 
 arrive. The connection is made by wrapping a few turns of 
 chain about the rope and hooking the ends on by the hand-grip. 
 The latter is more convenient and secure than the former, par- 
 ticularly if the cars are in irain. The grip is a pair of clamps 
 attached to one end of the car, the end of the movable-jaw lever 
 (being closed on the rope by a ring slipping on to its mate. Rais- 
 ing the ring opens the clamp and drops the rope. The car is 
 released and switched to its proper track. In the same way it 
 can be as easily clamped on the rope to pull the car. For a 
 train of cars a leading grip-car is attached by a gripman, who
 
 UNDERGROUND HAULAGE SYSTEMS. 289 
 
 takes the train out or in. Clevis couplings are also used for 
 single cars and at the front and rear ends of trains. Sometimes 
 a "knock-off" hook is used at the rear to aid in uncoupling while 
 strain is on the rope. 
 
 Increasing the Tension of the Drums at the Driving End. 
 The difference in tensions due to the two loads pulling upon the 
 two branches of rope is slight. The amount of friction at the 
 sheave must at least equal the difference of loads. Should any 
 excessive loading or resistance occur, the sheave will slip and fail 
 to drive. A tighter grip must be taken on the rope by the sheave 
 to communicate power. It would also prevent excessive wear 
 of the rope. This is possible where the sheave has several grooves. 
 The number of coils will be such as will provide the necessary 
 friction to prevent the rope from slipping off. This may be 
 determined by the formulae for power transmission. Two drums 
 in tandem Serve much better, each one having four or five 
 grooves, the rope making a half-coil alternately upon each drum 
 in each groove before returning along the other trackway. 
 They are both geared to the engine for equal speed. 
 
 The Balance-car. The variations of load and changes of 
 temperature produce varying tensions and elongations of the 
 rope. To insure a uniform tension the sheave at the other end 
 of the line is mounted on a carriage and capable of sliding freely. 
 Beyond it is a sunken pulley over which is the rope from the car- 
 riage to a suspended counterpoise. When the loads on the rope 
 become heavy, the carriage is pulled by the pendent weight to 
 increase the rope tension. With a light load the weight of 
 the rope draws the carriage. If the grades favor the load, the 
 carriage is at the end of the run of the loaded cars, the power 
 being inside of the mine at the top of the empty track. On 
 aerial tramways the same regulating device is used, and in 
 Fig. 118 is an illustration of one. 
 
 Instead of the plain grooved sheave there may be a grip- 
 pulley, several feet in diameter, carrying radial forked arms to 
 seize the rope. The arms are capable of being lengthened or 
 shortened as the rope stretches and drags. When the arms have
 
 290 MANUAL OF MINING. 
 
 been extended to their limit, a short length of rope or a few links 
 of cable are removed and the forks readjusted. 
 
 An i8"X42 <; Corliss engine, with two 36-inch driving-pulleys,. 
 1 8-inch belts to a pair of 84- inch pulleys, and two pinions of 
 18 inches diameter with 26 teeth, gearing into 72-inch spur-wheels 
 with 91 teeth, drives two 72-inch rope-drums. The drums with 
 four grooves each operate an endless-rope system one and a 
 half miles long, delivering 1200 tons daily over an average road. 
 The fuel, labor, etc., of a 45<>ton endless ropeway is about $7.50 
 per day. The initial cost is about $6700 for a plant of 4200 feet 
 double track (1500 feet having a rise of 40 feet, the rest being 
 level). The rope is | inch diameter. 
 
 If only a single track is possible in the haulage- way, the oper- 
 ation is intermittent. The engine alternately pulls the rope 
 with its train inward and outward, being reversed for each trip. 
 This resembles the tail-rope system with the two ropes linked 
 together. One line of rope is supported on overhead sheaves, 
 and the other, to which the cars are attached, rests on the floor 
 rollers. The cars are in train. 
 
 The road may be more readily extended with this system 
 than with the tail-rope. It costs less for power and can be carried 
 around sharper curves and over steeper grades. It can be used 
 with practically every condition of roadway with greater econ- 
 omy than any of the other systems. The delivery of cars is 
 uniform and continuous, regardless of the length of haulage. 
 The amount of car and rope repairs is less than in the tail-rope 
 system. The risk of accident to men along the roadways or 
 damage to property is also less, because of the slow speed. 
 Though applicable to sinuous undulating roads, it is best adapted 
 to uniform grades throughout the entire length of the roadway, 
 because of the varying load which would otherwise be thrust 
 upon the engine where the grades are steep and variable. This 
 would be particularly true when the intermittent attaching of 
 cars is very irregular. 
 
 Supporting the Haulage-rope. Rollers and sheaves along 
 the haulage- way support the rope to reduce its wear and to dimin-
 
 UNDERGROUND HAULAGE SYSTEMS. 291 
 
 ish the frictional resistance. They are placed midway between 
 the rails on suitable blocks spiked to the ties, and 20 to 30 feet 
 apart. They are also employed for supporting the tail-rope 
 near the roof by being suspended from the roof-timbers. The 
 rollers are usually of wood, 12 to 18 inches long, 5 inches in 
 diameter, and carefully turned with an iron journal for support. 
 They are spaced on a level track from 15 to 20 feet apart (closer 
 together if the rope is large or the grade steep). The bearings 
 also are of oak, well oiled. Hollow rollers may be obtained of 
 cast iron with curved faces, to keep the rope as central as possible. 
 They are shorter than the wooden rollers and offer much less 
 resistance. 
 
 Deflecting the Main-rope Around the Curves. This is accom- 
 plished by a number of small sheaves laid on the centre line of 
 the track, concentric with the rails. They are of a diameter of 10 
 to 15 inches and are spaced 3 feet or so apart to give a slight 
 deflection of each one for the entire length of the curve. On 
 reverse curves they are laid nearer the inside rail than the outer 
 rail, with their axis somewhat inclined and leading outward. 
 The formulae for determining their spacing are to be found on 
 page 177. 
 
 Short vertical rollers are also placed on the ties when the rope 
 is to be deflected around a gentle curve. For a sharp curve the 
 rollers are frequently placed at the side of the track, with their axes 
 slightly inclined from the vertical to carry the rope. These are 
 spaced 3 feet apart outside of the inner rail of the curve. They 
 are usually of cast iron, chilled, with an upper flange. Guide- 
 blocks are laid at intervals from the inner rail toward the sheave 
 on all simple curves in order to guide the rope into its position in 
 the groove, and also to prevent its catching in the rail as the 
 train moves around the curve. These blocks are of oak, spiked 
 to the ties. 
 
 Supporting the Tail-rope. Tail-rope sheaves are hung on 
 the posts at the sides of the haulage- way, 25 to 30 feet apart. 
 For these a perfect alignment is necessary, in order to reduce 
 friction and increase the security to the rope in the groove. They
 
 2Q 2 MANUAL OF MINING. 
 
 are of chilled cast iron or case-hardened, having wooden filling 
 in the grooves, and are from 10 to 16 inches in diameter. 
 
 The tail-rope is turned out of a side heading by a horizontal 
 guide-sheave of a diameter as large as convenient. Its pattern 
 is like that of the clip-pulley described under Power Transmission, 
 Chapter IX. A large diameter is necessary because the angle of 
 contact exceeds 90. 
 
 An electric system along the tramways is requisite for safety 
 as well as for signalling, though the malicious destruction of 
 insulation, etc., has caused its abandonment by many operators. 
 
 EXAMPLES. i. A one-mile endless rope, travelling at two miles an hour 
 over an average grade of 3 in 100, delivers 50 tons per hour. Required the 
 total resistance on the line and the size of the cylinders under 50 Ibs. boiler 
 pressure and 160 feet piston speed. Each car weighs 800 Ibs. and carries 2000 
 Ibs. Along the line are distributed 25 loaded and 25 empty cars, about 
 100 feet apart, and a ton is delivered every 36 seconds; with a coefficient of 
 friction of 0.02, the frictional resistances on the halves of the line are 1400 
 and 400 Ibs. respectively. To raise 25 tons up the plane requires a force 
 of 3 per cent of 50,000, or 1500 Ibs.; the gravity component of the cars is 
 600 Ibs. The gravity components of the cars and the rope, up and down, 
 balance each other, leaving the work of the engine to be that of overcom- 
 ing 1500+ 1800 (the drag of the rope is assumed at 3000), which, carried at a 
 rate of 176 feet per minute, requires 1,109,800 ft. -Ibs., or 33.6 I.H.P. The 
 piston diameter, k, is 7$ inches, and for 130 strokes, s, nearly 15 inches. 
 If the driving-sheave is 6 feet diameter, it makes 9.3 revolutions per minute, 
 and is therefore geared i to 7. 
 
 2. A haulage-engine having cylinders 24"X48", at the head of a plane 
 of 10 per cent grade, is directly connected to its drum. With an effective 
 steam pressure of 45 Ibs. per square inch, a piston speed of 350 feet per 
 minute, and a train speed of 8 miles an hour, what is the size of the drum 
 and the capacity of the plant? 
 
 The total piston pressure is 27,150 Ibs., assuming a modulus of , which, 
 at 350 feet per minute and a train speed of 704 feet per minute, represents 
 a continuous load of 13,497 Ibs. This tension upon the hauling-rope, due 
 to the component, parallel to the plane, of the weight of the rope, cars, and 
 load, plus their friction, is therefore limited to 13,497 Ibs. For this working 
 load the rope may be ij inches diameter, weighing 2 Ibs. per running foot; 
 the cars may be assumed to weigh one half that of their contents, W, and 
 the frictional resistances may be allowed for at the rate of 50 Ibs. per ton 
 of normal pressure. Then, by formulae in Chapter V, 
 13.497= TVVi2,ooo+fHO(i + 0.025).
 
 UNDERGROUND HAULAGE SYSTEMS. 293 
 
 W becomes 79,840 Ibs., and the hourly capacity less than 160 tons, without 
 allowing for delays. 
 
 3. To ascertain if the engine is properly proportioned for starting the 
 given load, we examine the table on page 131. Note that the coefficient 
 is 0.3974, which multiplied by 45X(24) 2 X4 gives a minimum moment of 
 41,202 ft.-lbs. With a direct-acting engine the drum must have a diameter of 
 5 feet 6 inches. The engine can therefore start from the bottom a load, W, 
 greater than 99,300 Ibs. 
 
 4. If the plane were double-tracked and the engine supplied with two 
 drums fast on the same shaft, the only work falling upon the engine would 
 be that due to the contents of the cars plus the total friction on the double line. 
 In this case seven full-length trips might be made per hour, thus increasing 
 the hourly capacity to 377 tons. 
 
 5. Required the length of a haulage plant for a tail-rope system to 
 deliver icco tons in ten hours over a road the first 600 feet of which has 
 a dip of 4 feet per 100 with the load, the next 2100 feet a grade of 3 feet per 
 100 against the load, and the lower 600 feet 2 feet per 100 against the load. 
 
 Assume that each car weighs 1000 Ibs., and that its capacity is 2000 Ibs.; 
 assume also a piston speed of 200 feet per minute and an average train-speed 
 of 6 miles an hour. 
 
 On the lowest section gravity produces a tension on the main-rope of 
 5665 Ibs., and the engine has to perform 2,991,120 ft.-lbs.; on the next 
 upper section for 4 minutes 122.6 horse-power is necessary to pull the train 
 up-grade. At the head of the gangway the grade favors the loaded cars, 
 which produce a tension on the tail rope of over 6000 Ibs. With a drum 
 of 132 inches diameter, making 15.3 revo utions per minute, gearing to a 
 pinion of 0.40 its diameter, a modulus of |, and an effective steam pressure 
 of 40 Ibs., the diameter of each cylinder would be 21 inches and the stroke 
 30 inches. 
 
 REFERENCES. 
 
 Electric Haulage in Metal Mines, W. N. Clark, Eng. & Min., Feb. 25, 
 1904; Benzine Locomotives for Gaseous Mines, L. Volf, Oesterr. Zeitschr. f. . 
 Berg. u. Huttenwesen, Oct. 4, 1902; Electric Haulage Plant, Berwind-White 
 Collieries, T. W. Sprague, E. & M. Jour., LIX, 508; Electric Mine Loco- 
 motives, William L. Affelder, Third-rail Friction Locomotive, M. & M., 
 Vol. XXIV, 383. 
 
 Notes on the Compressed-air Haulage Plate at No. 6 Colliery of the Sus- 
 quehanna Coal Co., Glen Lyon, Pa., J. H. Bowden, A. I. M. E., Vol. XXX, 
 566. 
 
 Rope Hauling, Tail-rope, Double Tail-rope, Pa. Mine Insp., 1889, 340; 
 Underground Wire-rope Haulage, Sci. Am. Sup., Dec. 27, 1902; Wire-rope
 
 2Q4 MANUAL OF MINIXG. 
 
 Slope Haulage, M. & M., April 1904; Slope Haulage, M. & M., Vol. XXIV, 
 413; Rope Transmission, M. & M., Sept. 1900, 89; Aerial Ropeways, 
 A. I. M. E., Vol. XIX, 760; Hildbrandt Essay, 53. 
 
 Engine Planes at Wearmouth Colliery, W. R. Bell, Trans. M. & M., Eng., 
 XLV, 219. 
 
 Report of Tail-rope Committee, N. E. I., XVII; Tail-rope System, De- 
 scription of Plant, R. S. Williamson, Coll. Guard., Nov. 1896, 1096; Third- 
 rail Tail-rope Haulage, M. & M., XXIII, 465. 
 
 Endless Rope, Hildbrandt Essay, 37; The Endless-rope Haulage System, 
 M. & M., Vol. XX, 251, 520, and 566. 
 
 Brake for Inclines, Coll. Guard., Vol. LXXXII, 846; Steam Locomotives 
 for Mining, Cassier's Mag., July 1904, 291; Air and Internal Combustion, 
 Aug.; and Electric, Oct. 1904. 
 
 Mules, Care of: Protection in the Mines, M. & M., Vol. XXIII, 272, 568; 
 Shoeing of Pit Horses, J. A. Longden, Brit. Soc. Min. Stud., IV, 127; De- 
 scription of Improved Colliery Stahles, J. P. Kirkup, Brit. Soc. Min. Stud., I, 
 328; Selection and Care cf Horses, M. E. Boissier, Coll. Guard., June 1897, 
 
 "73- 
 
 Tracks, Latches, etc., Hildbrandt Essay, 18, 42; Self-acting Switch, X, 
 Coll. Eng., Nov. 1895, 78, 80; Roads for Mines, James W. Abbott, Eng. & 
 Min. Jour., May 26, 1903; Grades in Rooms, M. & M., Vol. XXII, 227; 
 in Roads, M. & M., Vol. XXII, 180; Maintenance of Track, M. & M., 
 Mar., 199, 338. 
 
 Different Methods of Mine Haulage Compared, B. F. Jones, M. & M., 
 Aug. 1902; Comparison of Haulage Methods, M. & M., Vol. XXIII, 8. 
 
 Timbering in Mines, Mineral Industry, 740. 
 
 The Underground Ry. of the Stassfurt Salt Mines, J. Westphal, Glauckauf, 
 Oct. 4, 1902; The Longest Mine Haulage, F. Z. Shellenberg, A. I. M. E., 
 Vol. XXIX, 101. 
 
 Haulage Tests, Dynamometer Truck Measuring Resistances, Coll. Guard., 
 Nov. 20, 1896, 981; Sheaves, Pulleys, Tracks, etc., Hildbrandt Essay, 38; 
 Mechanical Devices at Coal Mines, L. L. Logan, M. & M., Feb. 1904; 
 Sheaves, Pulleys, Tracks, etc., at the Leith Coal Mine, Pa., H. L. Auchmuty, 
 Coll. Guard., Eng., Aug. 1896, 6; Checks for Sharp Curves on Engine Planes, 
 Idris Thomas, Coll. Guard., April 6, 1900. 
 
 Curves in Roads, M. & M., Vol. XXIII, 280, 456; Track Switches, 
 M. & M., Vol. XXIV, 331.
 
 CHAPTER IX. 
 
 WIRE-ROPE TRANSMISSION. 
 
 Ropes for Power Transmission are of different materials, 
 cotton, hemp, manilla, and steel wire being the most common. 
 For mining purposes the cotton ropes are relatively weak com- 
 pared with manilla, and have given place to the latter, which are 
 still largely employed for some purposes of driving. If care is 
 taken in the proper selection of a vegetable-fibre rope and in pro- 
 viding it with large pulleys, its durability is increased to an extent 
 that will make it a good competitor of wire rope, which is the 
 most extensively used for power transmission. The steel-wire 
 rope is of the y-wire, 6-strand type of a diameter of f to f inch, 
 according to the amount of power desired. 
 
 Rope-driving Systems. Motion may be imparted to an 
 endless rope by its adhesion to a sheave around which it passes 
 and thence to a distant sheave. If the distance between the 
 sheaves is great, the rope is supported at intermediate points. 
 The amount of power which can be delivered to the far sheave 
 depends solely upon the amount of adhesion between the rope 
 and the surface of the sheave. In the American system one 
 endless rope is wound several times around the driving and the 
 driven pulleys. In the English system the sheaves have mul- 
 tiple grooves. Several independent ropes run side by side, each 
 driving in its own groove. The latter, or multiple, system more 
 easily transfers power between the generator and motor, and is 
 more secure against the possibility of breakdowns. On the 
 other hand, the ropes have an unequal stretch and are liable to 
 drive at different velocities and tensions, thus introducing loss 
 
 295
 
 296 MANUAL OF MINING. 
 
 of power, greater wear, etc. In the American, or continuous, 
 system there is only one splice in the rope and no differential 
 driving is likely to exist. 
 
 The Transmission of Power. The amount of adhesion deter- 
 mining the power transmission depends on the friction between 
 the surfaces in contact, the angle of wrap, and tensions on the 
 rope. Being of slow speed, the tensions in the two branches 
 of the rope are due solely to, and depend directly upon, the stretch 
 produced by the weight of the sections between the supports and 
 the amount of sag. 
 
 Sheaves and End Carriages. The sheaves around which 
 the power-rope travels are horizontal, the upper (Fig. 94) being 
 on a fixed frame, and the lower one at the end, remote from the 
 power, on a carriage-tower frame. These sheaves are of large 
 diameter keyed to a heavy shaft, the latter running in a step at 
 its foot and held in a box at the upper end. The rate of revolu- 
 tion of the sheaves is 125, or less, per minute, the speed of the 
 rope being maintained below 4200 feet per minute. The rim 
 of the pulley holds in its groove a number of wooden blocks, 
 or a strip of rubber, to furnish as large a friction as possible. 
 In some cases the rim is composed of a number of gripping- 
 jaws, which are capable of locking on the rope. The brake- 
 wheel is bolted above the arms of the grip-sheave, and is controlled 
 by an adjusting-screw and hand-wheel. In power transmission 
 both frames carrying the end sheaves are fixed, and provision is 
 made for the stretch of the rope by a sheave with arms which 
 are capable of extension, thus enlarging the diameter of the 
 sheave. The rope may also be cut out and respliced. 
 
 At intermediate points on the line sheaves are supported 
 on the pillars or towers to relieve the rope of the excessive stress 
 of a long rope. These are of cast iron, 1 8 or 20 inches in diame- 
 ter, and fixed on the ends of cross-arms. They run loose on 
 their axles. The grooves are deep and wide to admit the rope 
 nearly to the bottom, the sides being curved to let the rope slip 
 easily. Occasionally the grooves are lined. The distance be- 
 tween the sheaves is governed by the allowable tension and sag.
 
 WIRE-ROPE TRANSMISSION. 297 
 
 For power purposes the maximum is 450 feet; for aerial tram- 
 ways it may be only 200 feet. If they merely support the ropes, 
 they are small in diameter, but if intended to deflect the rope, 
 their diameter is increased. Care must be taken in their align- 
 ment in a vertical plane and that the axles each have as little 
 friction as possible. On steep inclinations a second sheave 
 vertically above the lower supporting sheave is placed to prevent 
 the rope from flying from the wheel. Where it is possible for 
 power transmission, the upper branch of the rope is made the 
 driving portion, and the lower branch its slack part. 
 
 Maintaining Uniform Tension. All single-rope power-trans- 
 mission lines require some form of tightener to maintain a con- 
 stant tension upon the rope and a uniform driving power. Wire- 
 rope lines for endless-rope haulage and aerial tramways are 
 illustrations of this type. In the English multiple-rope system 
 tighteners are not used. 
 
 The tightener may be a sheave or a pair of drums on a mov- 
 able carriage. When the drums are at the power end of the 
 line they are placed on the same base with their axes parallel. 
 The rope is wound consecutively from one to the other of the 
 drums and around all of the grooves. Either or both may then 
 act as the driver if they are geared together. The two may also 
 be mounted on a carriage, in which case they are at the driven 
 end of the line. The carriage is free to travel upon a track and 
 is fitted with a heavy pendent weight, such that when the rope 
 stretches the counterpoise will pull the carriage to balance the 
 tension, as in Fig. 93. The length of the travel of the carriage 
 must be sufficient to allow of a freedom of movement within 
 limits of variation in rope length caused by the changes of load. 
 
 The Tension of the Rope. On the tight side the tension is 
 that which is due to the centrifugal action of the pulley-sheave, 
 the bending stress of the rope, and the power which the latter 
 receives from, or remits to, the respective sheave. This must 
 never exceed the maximum safe stress of the rope and should 
 be less than one third of the ultimate strength. The tension on 
 the slack side is equal to that on the driving side less the power
 
 -298 MANUAL OF MINING. 
 
 taken off by the sheave. It is from a half to one quarter of the 
 tension on the driving side, and the relation between these three 
 forces is indicated by the following formulae: 
 
 Let v= velocity in feet per second; 
 d= diameter of the rope, inches; 
 R = weight of the rope per foot; 
 Z=distance between stations, feet; 
 TT=weight on the journals; 
 ^^revolutions per minute; 
 S= maximum safe stress of the rope; 
 k= bending stress of the rope; 
 
 7\= available working tension of the rope, maximum; 
 C = stress due to centrifugal force; 
 T= tension on the slack side of the rope; 
 P= power transmitted to the pulley, Ibs.; and 
 H= horse-power transmitted. 
 
 Then 
 C=o.i3^ 2 ; Tt=S-C-k; P = T 1 -T; and H =0.00182 Pv. 
 
 The influence of centrifugal force is negligible in power trans- 
 missions for underground haulage or aerial tramways. In these, 
 too, the difference in tensions is not great, for the dead load of 
 cars, buckets, or conveyors is exerted nearly equally upon both 
 branches of the line. 
 
 The Sag of a Rope. The tension upon the rope may be 
 regulated by the amount of sag, or deflection, permitted between 
 the two points of support. The relation between them is indi- 
 cated by the formula below. Having calculated the sag for a 
 given power condition, a mean between h and A t is taken. The 
 end sheaves are separated sufficiently to reduce the sag at rest 
 to this amount. When in action the desired tensions will be 
 obtained. If more power is desired, the sheaves are separated 
 somewhat. If the desired amount cannot be obtained, the rope 
 is cut and spliced. This must be done every three months, as 
 wear and use weaken the rope.
 
 WIRE-ROPE TRANSMISSION. 299 
 
 The deflection of a power- rope at the centre of its span in 
 feet being h for the slack and h^ for the tight portion, then 
 
 RV RL 2 
 
 h and h= 
 
 Where possible, the upper branch of the rope should be the 
 driver and the lower branch the slack portion. An average 
 sag on spans of 150 to 250 feet is about 3 feet when the rope is 
 at rest, this amount being about one half of the sag permitted 
 for the slack portion when in action. 
 
 EXAMPLE. Power is to be transmitted by a steel 7-wire rope, i inch in 
 diameter, at a velocity of 4200 feet per minute over i2-ft. wooden grooved 
 pulleys 400 feet apart. Required the horse-power which can be trans- 
 mitted and the deflections to be given the rope. 
 
 Let 5=2i,ooo Ibs. for the rope, and assume the tension on the slack 
 side to be one half that on the tight side. 
 
 Here ^=70; ^=0.5^; and =1.58. 
 
 Then C=o.i3^ 2 =637; =10,073 Ibs.; 7^=10,290; ^=0.5^ = 5145; 
 P=5i45 Ibs.; #=646.47 horse-power as a maximum; ^ = 8.2 feet; and 
 =4.1 feet. 
 
 Aerial Tramways. Elevated endless ropes are a special form 
 of power transmission. They have long been used as a cheap 
 mode of transportation, serving as feeders to establish systems 
 of communication. They consist of a ropeway elevated above 
 the surface, with carriers for mineral swinging free above the 
 ground. The elevated structure is essentially a series of pillars, 
 or towers, of a strength depending upon the load and its speed 
 and the inclination. It is employed where the contour is too 
 rugged for economical road-building, and is therefore attractive 
 to the miner and the quarryman. The mineral is charged into 
 the carriers at the mine and these are allowed to descend, tak- 
 ing with them the rope to which they are attached if the grade 
 is sufficient, or are pulled downward if power must be employed. 
 A succession of carriers is loaded at the top and delivers the 
 contents at the bottom, where they are automatically emptied 
 without stop, and thence returned on the empty line. The rope
 
 300 MANUAL OF MINING. 
 
 may be straight and of uniform grade, or follow the surface con- 
 figuration over undulating profile and around curves for dis- 
 tances of a mile or more, provided the power be sufficient to 
 overcome the resistance. 
 
 The power required to operate aerial roadways depends upon 
 the grade and average inclination of the line. In the simplest 
 case the grade may be uniform to the point of shipment. Then 
 the inclination alone will determine if auxiliary power is necessary 
 either to assist the load or to check its speed. If the inclination 
 exceeds 20 per cent, the system is self-acting and only a slight 
 braking power is requisite at the top of the line. If it exceeds 
 this amount, power is required to check the speed. The brake 
 must overcome the excess of the downward effort of the load 
 over the resistance of the rising empty bucket and the total 
 friction. If the grade is below 20 per cent, power may be re- 
 quired to supplement the force of gravity. It is not necessary 
 that the grade be uniform, provided there is ample fall from the 
 mine to the roadway. In this event great care must be exer- 
 cised in determining the value of the motor force as well as all 
 the elements which contribute to the resistance. If power is 
 required in such a case, it may be applied at either end of the 
 line upon a grip-sheave. The motor attached or belted to 
 the driving- sheave must be of the slow- speed type. The super- 
 structure is light and inexpensive. There is no interference with 
 or from surface travel. Climatic conditions do not interrupt 
 operations. The time for their construction is comparatively 
 short, and they have a capacity equal to that of the average mine 
 usually located in such districts. There is no machinery along 
 the line except the towers with their sheaves. In elevated rope- 
 ways only such driving and controlling apparatus is built as 
 may be required for maintaining a constant speed and uniform 
 tension. The expense of operation is comparatively low. 
 
 The Single-rope Tramway. The earliest of these tramways 
 was a single endless moving rope, of which the Hodgson and 
 Hallidie are the representatives. The buckets or the carriages 
 are attached to saddles which ride on the rope. They are either
 
 WIRE-ROPE TRANSMISSION. 
 
 301- 
 
 attached permanently to the rope or may be adjustable. In both 
 these patterns one and the same rope moves the load (Fig. 117),- 
 
 The Ore-carriers. The mineral is transported in buckets? 
 of various designs, according to the character of the material to 
 be handled. They are suspended by hangers or clips, which are 
 either inserted into the rope or clinched around the outside of it,, 
 and attached at intervals determined by the amount of material
 
 3 C2 
 
 MANUAL OF MINING 
 
 FIG. 118. The Power End of the Halliclie Svstem. 
 
 FIG. 119. The Ore-buckets of a Single-rope Tramway.
 
 WIRE-ROPE TRANSMISSION. 303? 
 
 to be delivered. Usually they are wrought-iron rectangular 
 buckets holding about TOO Ibs. each (Fig. 119). For the trans- 
 port of very large outputs the buckets may be nearer together 
 than the average 200 feet, or larger, and the rope may be heavier 
 than the ordinary size of f inch. The buckets may be loaded at 
 any point along the line, automatically or by hand. They are un- 
 loaded at the lower end by an automatic device in which the 
 carrier strikes a lever, which opens a catch, releasing the swinging; 
 bottom and discharging the ore. The bottom is returned to 
 place by a heavy counterpoise weight on the end of a projecting; 
 arm and is automatically locked. The hangers are so made 
 that they may pass uninterruptedly over the rims of the support- 
 ing sheaves and around the terminal pulleys. 
 
 Capacity. A ropeway running 200 feet per minute, cany- 
 ing 100 Ibs. per bucket every 100 feet, will deliver 60 tons per 
 shift. With a descent sufficient for gravity to supply the power,, 
 three men can manage all of its operations. It requires some' 
 supervision, and delivers ore at 20 to 35 cents per ton-mile (in- 
 clusive of all allowances), and about 60 cents per cord-mile for 
 wood. The line can be completed for $1.30 per foot, and $2000 
 for the machinery at the terminals. Curves and long stretches 
 increase the first cost; grade does not. The greatest item in 
 the maintenance cost is that for renewals of the clip-hangers. 
 It amounts to $100 a year on a line of half a mile in length. 
 
 The Cable Supports. These are simple frames or towers- 
 built up from a rectangular base. The former construction is 
 rigid and not liable to get out of line, nor can its cross-arm be 
 pulled down on the loaded side of the rope. In Figs. 117 and 
 1 20 are two typical constructions. The height of the stations 
 depends upon the distance between them. It is about 20 feet 
 when the stations are 200 feet apart. The frames are locatei 
 at the higher points always when there are any undulations irr 
 the grade of the hill, as shorter towers can be used with equal 
 result and the inclination of the rope is more nearly natural. 
 On top of each is bolted a cross-arm, at the ends of which are 
 boxed the carry ing- sheaves to support the rope and allow free-
 
 304 MANUAL OF MINING. 
 
 dom of movement. In some instances two sheaves are carried 
 an upper and a lower one. The object of the upper one is tc 
 ;prevent the rope jumping out of place from its groove in the 
 jlower. At a curve the standards are near together and the rope 
 jis slightly deflected at each standard to conform to the curve 
 -desired. The sharp turns or horizontal angles, as, for example, 
 .around bluffs, are also made by using large sheaves. One only 
 jis necessary for the turn when the clips holding the buckets pro- 
 ject outward from the rope, but two will be required when the 
 clips project inward. In the latter case the ropes cross in pass- 
 ing to the first pulley and also to and from the second pulley. 
 
 The end sheaves and carriages are similar to those described 
 for power transmission, with modifications dependent upon local 
 conditions. In Fig. 118 is illustrated the upper frame of the line, 
 while the lower end is illustrated in Fig. 117, on the left. 
 
 The Bleichert System. In the two-rope system there is a 
 pair of separate stationary cables over which numerous trolley 
 carriages, supporting buckets, are drawn by a light endless mov- 
 ing rope, called the traction-rope, to which the buckets are 
 Attached by patent grip (Fig. 120). The cables constitute the 
 roadway for the trolleys. The tubs are dumped automatically 
 into a bin or wagon at the lower end of the line. 
 
 The several varieties of two-rope systems differ mainly in 
 the mode of suspending the rope. In one system the bucket is 
 .attached to a clutch, which seizes the running rope and rails and 
 attaches by friction alone. It locks and releases automatically, 
 .and exerts a uniform friction whatever may be the slope of the 
 ropes. The rope permits them to be disconnected and run off 
 on suspended rails when they reach their terminal, or any turn- 
 out, on the line. This enables them to be loaded and discharged 
 without interference with the travel. 
 
 The double- rope system has a greater capacity than the 
 ; single rope. It is capable of much heavier traffic, for the indi- 
 vidual loads may be as great as 1000 Ibs. each, and the distance 
 between the towers 1000 feet, though the average span is 300 
 .feet. The individual loads carried by the Hallidie lines cannot
 
 WIRE-ROPE TRANSMISSION. 
 
 3S 
 
 exceed 150 Ibs., nor the distance between the towers 300 feet. 
 The buckets cannot be closer than the distance between the towers. 
 The speed of the rope is limited to 300 feet per minute, because 
 of the danger of the rope jumping off the carriage- sheaves. 
 
 TRAMWAY CAR FOR THE TRANSPORTAffON -OF GOAL. ORES. SANDS, fcc. 
 SHOWING LUG COUPLING 
 
 FIG. 120. A Double-rope Aerial Roadway. 
 
 The operating cost of the double- rope system is higher than 
 that of the Hallidie single ropeway, but its capacity is also larger. 
 The materials, and supplies, in one instance, for a two-rope sys- 
 tem was $1.90 per foot, exclusive of the machinery. The net 
 cost of a single ropeway in the immediate vicinity of the above 
 is $1.21 per foot of line, plus the cost of the two end structures. 
 The cost of repair of either system is not heavy, but the double- 
 rope system is less than that of the single- rope. On one of the 
 former, two miles in length, the cost of maintenance is 1.5 cents 
 per ton put of the total operating cost of 17.5 cents per ton. 
 
 A self-acting aerial tramway, like the self-acting underground 
 plane, may, by stretching one or two ropes the full length of 
 the line, serve as a guide for large skips holding a ton or so 
 which are attached to an endless rope. The latter rope is of a 
 length equal to that of the roadway, and is operated from a clip- 
 pulley at the top by gravity with or without power, according as 
 the grade is small or large.
 
 306 MANUAL OF MINING. 
 
 REFERENCES. 
 
 Aerial Tramways, at Sandor, C. T. Finlayson, B. C. Min. Sci. Press, June 
 1897, 544; Various Types of Aerial Ropeways, W. Carrington, Coll. Guard., 
 Mar. 1897, 556; Wire-rope Tramways, E. & M. Jour., LXI, 208; The Trans- 
 mission of Power by Wire Rope, W. H. Graves, M. & M., April 1904; Wire- 
 rope Tramways, J. H- Janeway, Jr., M. & M., April 1904; Aerial Ropeways, 
 A. I. M. E., Vol. XIX, 760; Hildbrandt Essay, 53. 
 
 New Rope Clip, Coll. Guard., Vol. LXXX, 823. 
 
 Compression of Air, Peabody, Thermodynamics, Coll. Eng., XVII, 173; 
 Air Compressors, Valve Movements, Coll. Eng., Feb. 1897, 319; A Study in 
 the Economical Arrangement of Compressed-air Haulage, E. & M. J., Feb. 
 28, 1903. i
 
 CHAPTER X. 
 
 THE COMPRESSION OF AIR. 
 
 Compressed Air is employed in the same manner as steam for 
 motor purposes in driving rock-drills, pumps, locomotives, and 
 coal-cutters. The advantage over steam lies in its transmission 
 without condensation and giving cool, dry, ventilated rooms 
 instead of hot rooms from exhaust where steam is used. It is, 
 however, dearer than the other motor agency. It can be trans- 
 mitted over great distances where electric transmission is not 
 desirable or economical. 
 
 When air is subjected to pressure its volume is proportion- 
 ately diminished, and the energy thus expended in the compression 
 is retained by the air and capable of being applied as is the ten- 
 
 FIG. 121. A Water-jacketed Air-compressor. 
 
 sion of steam. To obtain air at an absolute pressure of 100 Ibs. 
 per square inch a unit volume must be reduced to 0.147; a ^ a 
 pressure of 200 Ibs. per square inch, to 0.074. In order to obtain 
 the same pressure from steam it must be heated to 338 F. and 
 388 F. respectively. 
 
 The air-compressor is an ordinary cylinder provided with a 
 piston and suitable valves for admitting and delivering air. During 
 
 307
 
 308 
 
 MANUAL OF MINING. 
 
 the out-stroke air flows into the cylinder through the inlet- valves 
 AA, Fig. 121, which close when the return-stroke commences. 
 The air in the cylinder is then compressed until it opens the 
 delivery-valves, BB, through which it is forced into a receiver. 
 
 The Weight of Air. The following table of weights of a 
 cubic foot of air at different temperatures has been calculated 
 on the assumption of a constant pressure of 30 inches of mercury, 
 approximately sea-level pressure: 
 
 Tempera- 
 ture /. 
 
 Weight in 
 Decimals of 
 a Pound. 
 
 Tempera- 
 ture *. 
 
 Weight in 
 Decimals of 
 a Pound. 
 
 3 2 
 
 .0809749 
 
 120 
 
 .0686678 
 
 35 
 
 .0804831 
 
 125 
 
 .0680799 
 
 40 
 
 .0796767 
 
 130 
 
 .0675020 
 
 45 
 
 .0788863 
 
 135 
 
 .0669338 
 
 5 
 
 .0781113 
 
 140 
 
 .0663751 
 
 55 
 
 0773515 
 
 145 
 
 .0658256 
 
 60 
 
 .0766063 
 
 IS 
 
 .0652852 
 
 62 
 
 .0763122 
 
 155 
 
 0647535 
 
 65 
 
 075 8 753 
 
 160 
 
 .064^305 
 
 70 
 
 .0751582 
 
 165 
 
 .0637158 
 
 75 
 
 .0744544 
 
 170 
 
 .0632093 
 
 80 
 
 0737 6 38 
 
 175 
 
 .0627108 
 
 85 
 
 .0730858 
 
 180 
 
 .0622201 
 
 QO 
 
 .0724202 
 
 185 
 
 .0617371 
 
 95 
 
 .0717666 
 
 190 
 
 .0612614 
 
 100 
 
 .0711246 
 
 195 
 
 .0607931 
 
 105 
 
 .0704941 
 
 200 
 
 .0603318 
 
 no 
 
 .0698746 
 
 205 
 
 0598775 
 
 "5 
 
 . 0692660 
 
 212 
 
 .0592529 
 
 The table on page 309 shows the weight of air at various 
 altitudes above the sea. 
 
 Free Air. Free air is understood as air at the ordinary tem- 
 perature and an atmospheric pressure of 14.7 Ibs. per square 
 inch or 2116.8 Ibs. absolute per square foot. By "absolute pres- 
 sure" is meant the pressure, above the vacuum as distinguished 
 from gauge pressure, which is measured above the atmosphere 
 by the gauge. Absolute temperature, at Fahrenheit scale, is 
 the reading of the thermometer, in Fahrenheit, plus 461, 
 and is represented by T. At 461 F. there is no pressure. 
 The absolute temperature is o.
 
 THE COMPRESSION OF AIR. 
 
 39 
 
 TABLE OF ABSOLUTE PRESSURES, BOILING-POINTS, ETC., AT DIFFERENT 
 HEIGHTS ABOVE SEA-LEVEL. 
 
 I 
 
 
 
 3 
 
 4 
 
 s 
 
 6 
 
 7 
 
 Height 
 above 
 Sea-level, 
 Feet. 
 
 Barometer, 
 Inches of 
 Mercury. 
 
 Boiling- 
 point, 
 Degrees 
 Fahr. 
 
 Absolute 
 Pressure, 
 Pounds. 
 
 Weight of 
 i Cubic 
 Foot of 
 Air at 60, 
 Pounds. 
 
 Volume of 
 Air Equal 
 to i Cubic 
 Foot of 
 Free Air at 
 
 Volume of 
 Free Air at 
 Sea-level 
 Equal to i 
 Cubic Foot 
 at Given 
 
 
 
 
 
 
 Sea-level. 
 
 Altitude. 
 
 
 
 30 
 
 212 
 
 14-7 
 
 .0765 
 
 I 
 
 I 
 
 5" 
 
 29.42 
 
 211 
 
 14.41 
 
 07499 
 
 .02 
 
 .98039 
 
 1,025 
 
 28.85 
 
 210 
 
 14.136 
 
 07356 
 
 .04 
 
 .96154 
 
 ^539 
 
 28.29 
 
 209 
 
 13.86 
 
 .07213 
 
 .06 
 
 9434 
 
 2,063 
 
 27-73 
 
 208 
 
 I3-587 
 
 .07071 
 
 .08 
 
 9259 
 
 2,589 
 
 27.18 
 
 207 
 
 I3-3I8 
 
 .0693 
 
 .10 
 
 .90909 
 
 3> IT 5 
 
 26.64 
 
 206 
 
 I3-054 
 
 .06793 
 
 .12 
 
 .89285 
 
 3, 6 42 
 
 26. II 
 
 20 5 
 
 12.794 
 
 .06658 
 
 .14 
 
 .87719 
 
 4,169 
 
 25-59 
 
 204 
 
 12-539 
 
 -06525 
 
 17 
 
 8547 
 
 4,697 
 
 25.08 
 
 203 
 
 12.289 
 
 06395 
 
 .I 9 
 
 .8403 
 
 5, 22 5 
 
 24-58 
 
 2O2 
 
 12.044 
 
 .06267 
 
 . 22 
 
 .8197 
 
 5,764 
 
 24.08 
 
 2OI 
 
 i i . 799 
 
 .0614 
 
 .24 
 
 .8064 
 
 6,304 
 
 23-59 
 
 2OO 
 
 11 -559 
 
 .06015 
 
 27 
 
 7874 
 
 6,843 
 
 23.11 
 
 199 
 
 11.324 
 
 05893 
 
 2 9 
 
 7752 
 
 7,38i 
 
 22.64 
 
 I 9 8 
 
 11.094 
 
 05773 
 
 32 
 
 75757 
 
 7,932 
 
 22.17 
 
 197 
 
 10.863 
 
 0565 
 
 35 
 
 74074 
 
 8,481 
 
 21.71 
 
 196 
 
 10.638 
 
 05536 
 
 38 
 
 .7246 
 
 9,031 
 
 21 . 26 
 
 195 
 
 10.417 
 
 .05421 
 
 4i 
 
 .7092 
 
 9,579 
 
 20.82 
 
 194 
 
 10.202 
 
 05309 
 
 44 
 
 6944 
 
 0,127 
 
 20-39 
 
 193 
 
 9-99 
 
 .05199 
 
 47 
 
 .6802 
 
 0,685 
 
 19.96 
 
 192 
 
 9.78 
 
 .0509 
 
 50 
 
 .6666 
 
 1,243 
 
 J 9-54 
 
 I 9 I 
 
 9-57 
 
 .0498 
 
 53 
 
 6536 
 
 T ,799 
 
 !9-!3 
 
 I 9 
 
 9-37 
 
 .0488 
 
 -56 
 
 .64102 
 
 2,367 
 
 18.72 
 
 ISO 
 
 9.17 
 
 .0477 
 
 .60 
 
 .625 
 
 2,934 
 
 18.32 
 
 188 
 
 8.98 
 
 .0467 
 
 63 
 
 6i35 
 
 13,498 
 
 17-93 
 
 187 
 
 8.78 
 
 0457 
 
 67 
 
 .6 
 
 14,075 
 
 17-54 
 
 186 
 
 8-59 
 
 .0447 
 
 7 1 
 
 .5848 
 
 14,649 
 
 17.16 
 
 185 
 
 8.41 
 
 437 
 
 74 
 
 5747 
 
 The volume of air may be altered by a change of temperature 
 or of pressure. It increases with a rise in temperature or a 
 reduction of pressure. Correspondingly it decreases for a di- 
 minished temperature or an increased pressure. 
 
 Adiabatic Compression. When compressed or expanded by 
 the application of force, air suffers a change of temperature. If 
 the walls confining the air be absolute non-conductors of heat, 
 the increase in temperature and of pressure may be definitely 
 known.
 
 310 MANUAL OF MINING. 
 
 The relation between the pressure, volume, and temperature 
 is expressed by the following equations, in which P = the initial 
 pressure of the air in pounds per square foot; F = the volume at 
 that pressure in cubic feet; and r = its absolute temperature P, V, 
 and T correspondingly represent the pressure, volume, and tem- 
 perature at the end of the operation. 
 
 and 
 
 log T t = log 7+0.29 log Pj-o.29 log P; 
 log T, =log T+ 0.408 log V 0.408 log Fjj 
 log P + 1.408 log 7 = log P t + 1.408 log F x . 
 
 Thus, if a pound of free air at 60 be compressed to 90 Ibs. abso- 
 lute, it will attain a final temperature of 881 absolute or 420 F., 
 the final volume becoming 0.276 of the original. This is known 
 as adiabatic compression. 
 
 log ^ = 2.716838+0.29X1.954243-0-29X1.176091, 
 
 whence T t = 88i. 
 
 Isothermal Compression. If the air be compressed within 
 walls which are perfect conductors and the compression be a 
 slow one, the heat generated will be absorbed or radiated as 
 fast as developed and the air, originally at 60, will remain at 60, 
 the process being isothermal. At the final pressure of 90 Ibs. 
 absolute the volume would be only 0.1640 of the original. Thus 
 isothermal compression furnishes a smaller volume at a lower 
 temperature than is attained in adiabatic compression. 
 
 The relation of volume, temperature, and pressure is repre- 
 sented by the equation PF = 53.i8r. Thus the volume of air 
 varies inversely as the pressure when the temperature is con- 
 stant; the absolute pressure varies directly with an absolute 
 temperature if the volume remains constant; and the volume 
 varies as the absolute temperature so long as the pressure re- 
 mains constant.
 
 THE COMPRESSION OF AIR. 
 
 TABLE OF VOLUMES, MEAN PRESSURES, TEMPERATURES, ETC., IN THE OPERA- 
 TION OF AIR-COMPRESSION FROM i ATMOSPHERE AND 60 FAHR.* 
 
 I 
 
 2 
 
 3 
 
 4 
 
 5 
 
 6 
 
 7 
 
 8 
 
 9 
 
 10 
 
 ii 
 
 
 
 , 
 
 *, ^ 
 
 a 
 
 tc-M 
 
 +> 
 
 M'~g 
 
 _ 
 
 
 
 
 
 
 o 
 
 I 
 
 ^ 
 
 0.5 
 
 fc 
 
 
 
 
 
 
 
 i 
 
 g 
 
 <3 s 
 
 j 
 
 o a 
 
 
 QO 
 
 Q T3 
 
 rt 
 
 
 i 
 
 I 
 
 4 
 
 H 
 
 "< 
 
 l*i 
 
 ^ 
 
 v C 
 
 
 &1 
 
 i 
 
 I 
 
 I 
 
 c 
 
 f 4 
 
 +3 
 
 
 
 
 l.ll 
 
 1JJ 
 
 IcS 
 
 
 1 
 
 
 ii 
 
 .1 
 
 gj 
 
 5jfl 
 
 H-d 
 
 
 
 !!<| 
 
 III 
 
 -1 
 
 
 1 
 
 1 
 
 W < 
 
 E-S 
 
 |!l 
 
 3 "8 
 
 ocj 
 
 MH 
 
 -8 
 
 a cflO 
 
 ij-jjjl 
 
 ^ u 
 
 ~'3 
 
 1 
 
 o 
 
 ^ 
 
 
 
 
 ^ 
 
 
 
 
 ^ 
 
 s 
 
 fe 
 
 o 
 
 
 
 14.7 
 
 i 
 
 ! 
 
 i 
 
 
 
 
 
 
 
 
 
 60 
 
 
 i 
 
 15-7 
 
 i. 068 
 
 9363 
 
 95 
 
 .96 
 
 975 
 
 43 
 
 44 
 
 71 
 
 I 
 
 2 
 
 16.7 
 
 1.136 
 
 .8803 
 
 .91 
 
 1.8 7 
 
 1 .01 
 
 .96 
 
 .96 
 
 
 2 
 
 3 
 
 17.7 
 
 1.204 
 
 8305 
 
 .876 
 
 2.72 
 
 2.8 
 
 1.4 
 
 1.41 
 
 88^9 
 
 3 
 
 4 
 
 i8. 7 
 
 1.272 
 
 .7861 
 
 .84 
 
 3-53 
 
 3.67 
 
 1.84 
 
 1.86 
 
 98 
 
 4 
 
 5 
 
 19.7 
 
 i-34 
 
 .7462 
 
 .81 
 
 4-3 
 
 4-5 
 
 2.22 
 
 2.26 
 
 1 06 
 
 5 
 
 10 
 
 24.7 
 
 1.68 
 
 5952 
 
 .69 
 
 7.62 
 
 8.27 
 
 4.14 
 
 4.26 
 
 145 
 
 10 
 
 15 
 
 29.7 
 
 2.02 
 
 495 
 
 .606 
 
 10.33 
 
 11.51 
 
 5-77 
 
 5-99 
 
 
 J 5 
 
 20 
 
 20 
 
 34.7 
 
 2. 3 6 
 
 4237 
 
 543 
 
 12 . 62 
 
 14.4 
 
 7.2 
 
 7.58 
 
 207 
 
 
 25 
 
 39-7 
 
 2-7 
 
 3703 
 
 494 
 
 J 4-59 
 
 17.01 
 
 8.49 
 
 9-5 
 
 234 
 
 2 5 
 
 30 
 
 44.7 
 
 3-04 
 
 .3289 
 
 4638 
 
 16.34 
 
 19.4 
 
 9.66 
 
 10.39 
 
 255 
 
 3 
 
 35 
 
 49-7 
 
 3-381 
 
 2957 
 
 .42 
 
 17.92 
 
 21.6 
 
 10.72 
 
 n-59 
 
 281 
 
 35 
 
 40 
 
 54-7 
 
 3.721 
 
 .2687 
 
 393 
 
 19.32 
 
 23.66 
 
 11.7 
 
 12.8 
 
 302 
 
 40 
 
 45 
 
 59-7 
 
 4.061 
 
 . 2462 
 
 37 
 
 20.52 
 
 25-59 
 
 12.62 
 
 13-95 
 
 321 
 
 45 
 
 50 
 
 64.7 
 
 4.401 
 
 . 2272 
 
 35 
 
 21.79 
 
 27-39 
 
 13.48 
 
 15-05 
 
 339 
 
 5 
 
 55 
 
 69.7 
 
 4-749 
 
 . 2109 
 
 331 
 
 22.77 
 
 29. II 
 
 14-3 
 
 15.98 
 
 357 
 
 55 
 
 60 
 
 74-7 
 
 5.081 
 
 .1968 
 
 
 23.84 
 
 30-75 
 
 
 16.89 
 
 375 
 
 60 
 
 65 
 
 79-7 
 
 5.423 
 
 .1844 
 
 .301 
 
 24.77 
 
 31.69 
 
 J 5-7 6 
 
 17.88 
 
 389 
 
 65 
 
 70 
 
 84-7 
 
 5.762 
 
 I 735 
 
 .288 
 
 26 
 
 33-73 
 
 16.43 
 
 18.74 
 
 405 
 
 70 
 
 75 
 
 89.7 
 
 6.102 
 
 .1639 
 
 .276 
 
 26.65 
 
 35-23 
 
 17.09 
 
 19-54 
 
 420 
 
 75 
 
 80 
 
 85 
 
 94-7 
 99-7 
 
 6.442 
 6.782 
 
 1552 
 .1474 
 
 .267 
 .2566 
 
 27-33 
 28.05 
 
 36.6 
 
 37-94 
 
 17.7 
 18.3 
 
 20.5 
 
 21 . 22 
 
 432 
 447 
 
 80 
 
 85 
 
 90 
 
 104.7 
 
 7-122 
 
 .1404 
 
 .248 
 
 28.78 
 
 39.18 
 
 18.87 
 
 22 
 
 459 
 
 90 
 
 95 
 
 109.7 
 
 7.462 
 
 134 
 
 .24 
 
 2 9-53 
 
 40.4 
 
 19.4 
 
 22.77 
 
 472 
 
 95 
 
 100 
 
 114.7 
 
 7.802 
 
 .1281 
 
 232 
 
 30.07 
 
 41.6 
 
 19.92 
 
 23-43 
 
 485 
 
 t oo 
 
 105 
 
 119.7 
 
 8.142 
 
 .1228 
 
 .2254 
 
 30.81 
 
 42.78 
 
 20.43 
 
 24.17 
 
 49 6 
 
 105 
 
 no 
 
 124.7 
 
 8.483 
 
 .1178 
 
 .2189 
 
 3T.39 
 
 43 -9 1 
 
 20.9 
 
 24.85 
 
 57 
 
 i 10 
 
 "S 
 
 129.7 
 
 8.823 
 
 "33 
 
 .2129 
 
 31.98 
 
 44.98 
 
 21.39 
 
 25-54 
 
 518 
 
 ii5 
 
 120 
 
 134-7 
 
 9.163 
 
 . 1091 
 
 2073 
 
 32 - 54 
 
 46.04 
 
 21.84 
 
 26.2 
 
 529 
 
 r ao 
 
 125 
 
 139.7 
 
 9-503 
 
 .1052 
 
 .202 
 
 33-07 
 
 47-06 
 
 22.26 
 
 26.8l 
 
 540 
 
 125 
 
 130 
 
 144.7 
 
 9-843 
 
 .1015 
 
 .1969 
 
 33-57 
 
 48 ..i 
 
 22.69 
 
 27.42 
 
 550 
 
 i3 
 
 135 
 
 149.7 
 
 10.183 
 
 .0981 
 
 .1922 
 
 34 5 
 
 49-i 
 
 23.08 
 
 28. oe 
 
 
 1 35 
 
 140 
 
 154-7 
 
 IO -5 2 3 
 
 095 
 
 .1878 
 
 34-57 
 
 50.02 
 
 23-41 
 
 28.66 
 
 57o 
 
 i4 
 
 145 
 
 159-7 
 
 10.864 
 
 .0921 
 
 1837 
 
 35 9 
 
 51 
 
 2 3-97 
 
 29.26 
 
 580 
 
 i45 
 
 
 164.7 
 
 i i . 204 
 
 .0892 
 
 .1796 
 
 35-48 
 
 51.89 
 
 24.28 
 
 29.82 
 
 589 
 
 1 5 
 
 1 00 
 
 J74-7 
 
 11.88 
 
 .0841 
 
 .1722 
 
 36.29 
 
 
 24-97 
 
 30.91 
 
 607 
 
 Z 6o 
 
 170 
 
 184.7 
 
 12.56 
 
 .0796 
 
 l6 57 
 
 37-2 
 
 55-39 
 
 25-71 
 
 32.07 
 
 624 
 
 1 7 
 
 1 80 
 
 194.7 
 
 13-24 
 
 755 
 
 Z 595 
 
 37-96 
 
 57-01 
 
 26.36 
 
 33-04 
 
 640 
 
 j8o 
 
 I 9 
 
 204.7 
 
 13.92 
 
 .0718 
 
 154 
 
 38.68 
 
 58.57 
 
 27.02 
 
 34.06 
 
 6^7 
 
 1 90 
 
 2OO 
 
 214.7 
 
 14-6 
 
 .0685 
 
 .149 
 
 39-42 
 
 60. 14 
 
 27.71 
 
 35-02 
 
 672 
 
 2 oo 
 
 *From Richards' "Compressed Air," page 21.
 
 312 MANUAL OF MINING. 
 
 The table on page 311 shows the relation between volume, 
 pressure, and temperature in adiabatic and isothermal com- 
 pression. 
 
 The columns 4, 8, and 6 respectively express the relations 
 assumed by one unit of air when compressed without change 
 of temperature to the stated pressures. Column 8 shows the 
 mean effective resistance during the period of compression which 
 is offered by the air to the point of delivery only. Column 5 
 contains the ratio of volumes of air compressed adiabatically to 
 given pressures; column 9, the mean effective resistance during 
 the compression period; and column 6, the mean effective 
 resistance during the entire stroke, including the period of 
 delivery. 
 
 EXAMPLES. i. Let it be desired to determine the final volume of i Ib. of 
 free air at a temperature of 62 F. compressed to 80 Ibs. gauge. 
 Then 
 
 V= 13.08; P is 14.7X144=2116.8; P,=94-7; and 1=62+461 = 523. 
 
 Tsothermally, 
 
 2ii6.8X 13.08= 27,729=P 1 F 1 . ^,=94.7X144X7,. 
 Whence 
 
 V = 2.027 cu - ft- 
 Adiabatically, 
 
 2ii6.8x(i3.oS) 1 - 408 =63,iio=P,(F l 1 - 408 ; 94.7Xi44X(F l ) I ' 408 =63,no. 
 Whence 
 
 F,- 3.492. 
 
 According to the preceding table the volumes are respectively (columns 
 4 and 5) 0.1552X13.08=2.027 cu. ft. and 0.267X13.08=3.49 cu. ft. 
 What is the final temperature in the latter case? 
 
 log r=log 523+0.29 log 94.7-0.29 log 14.7=2.953151. 
 r=897abs. =436F. 
 
 According to the table, column 10, r would have been 432 F. from 60 F. 
 2. Required the temperature of a pound of air which at 50 Ibs. abs. 
 pressure occupies a volume of 10 cu. ft. 
 
 PF=5oXi44Xio=53.i8r; 
 whence 
 
 T= 1353 absolute= 892 F.
 
 THE COMPRESSION OF AIR. 313 
 
 Air Indicator Cards. In Fig. 122 are indicator cards 
 steam- and air-cylinders; they show also the adiabatic and iso- 
 thermal compression. On the horizontal the volumes are meas- 
 ured from the right, and on the vertical lines the corresponding 
 pressures. It is noticeable that the adiabatic curve rises more 
 rapidly than the isothermal; in other words, its pressure in- 
 creases more rapidly. Again, the volumes in the former case are 
 correspondingly larger for a given pressure, as may be seen by 
 comparing volumes obtained by cutting the two curves by some 
 horizontal line. This may also be noted by comparing the vol- 
 umes in columns 5 and 4 of table page 311. 
 
 FlG. 122. Indicator Cards from Steam- and Air-cylinders. 
 
 The initial pressure of the steam in Fig. 122 is 58 Ibs. gauge, 
 and the cut-off 0.3. The final pressure of the delivery air is 80 
 Ibs. gauge. The mean effective pressure for the entire card, 
 including the period of d scha ge represented by the top horizon- 
 tal line, may be obtained in columns 7 and 6 respectively. These 
 mean effective pressures represent the resistance to compression 
 offered by the air under the conditions assumed. Thus, for 
 example, according to the table the mean effective pressure of the 
 entire air card, assuming isothermal compression to be 94.7 Ibs. 
 pressure absolute, 80 Ibs. gauge, would be 27.33 Ibs. per square 
 inch; if carried to the same degree adiabatically, 36.6 Ibs. per 
 square inch. During the compression of the strokes only, the 
 mean effective pressures are respectively 20.5 and 17.7 Ibs. per 
 square inch.
 
 3M MANUAL OF MINING. 
 
 The Work of Adiabatic Compression of w pounds of air is 
 
 F a =i8 3 . 4 5 (TI- rJw-wwp^-PV). 
 The Work of Isothermal Compression of w pounds of air is 
 
 V+v 
 W t =P(V+v) hyp log y~~^- (P|-P)*w, 
 
 in which v is the volume of clearance in the cylinder. 
 
 TABLE OF HYPERBOLIC LOGARITHMS. 
 
 (Base 2.72.) 
 Hyp log=2.3O26X common logarithm. 
 
 Number. 
 
 Logarithm. 
 
 Number. 
 
 Logarithm. 
 
 Number. 
 
 Logarithm 
 
 .01 
 
 .009 
 
 .10 
 
 74i 
 
 4-60 
 
 1.526 
 
 .02 
 
 .019 
 
 .20 
 
 .788 
 
 4.70 
 
 547 
 
 03 
 
 .029 
 
 3 
 
 832 
 
 4.80 
 
 568 
 
 .04 
 05 
 
 39 
 .048 
 
 .40 
 5 
 
 .875 
 .916 
 
 4.90 
 5.00 
 
 589 
 .609 
 
 .06 
 
 058 
 
 .60 
 
 955 
 
 5.10 
 
 .629 
 
 .07 
 
 .067 
 
 .70 
 
 993 
 
 5.20 
 
 .648 
 
 .08 
 
 .076 
 
 .80 
 
 .029 
 
 5-3 
 
 .667 
 
 .09 
 
 .086 
 
 .90 
 
 .064 
 
 5-40 
 
 .686 
 
 .10 
 
 95 
 
 3-oo 
 
 .098 
 
 5-50 
 
 .704 
 
 .11 
 
 .104 
 
 3.10 
 
 131 
 
 5-6o 
 
 .722 
 
 .12 
 
 "3 
 
 3-20 
 
 .163 
 
 5-7 
 
 .740 
 
 13 
 
 .112 
 
 3-3 
 
 193 
 
 5.80 
 
 757 
 
 .14 
 
 131 
 
 3-40 
 
 223 
 
 5-90 
 
 774 
 
 15 
 
 139 
 
 3-5 
 
 252 
 
 o.oo 
 
 .791 
 
 .20 
 
 .182 
 
 3-6o 
 
 .280 
 
 6.20 
 
 .824 
 
 2 5 
 
 .223 
 
 3-7 
 
 -308 
 
 6.40 
 
 .856 
 
 3 
 
 .262 
 
 3-8o 
 
 335 
 
 6.60 
 
 .887 
 
 .40 
 
 .336 
 
 3-9 
 
 .360 
 
 6.80 
 
 .916 
 
 
 
 .405 
 .470 
 
 4.00 
 4.10 
 
 386 
 .410 
 
 7.00 
 7.20 
 
 945 
 974 
 
 
 
 SB? 
 
 4.20 
 4-3 
 
 435 
 .458 
 
 7.40 
 7.60 
 
 .001 
 
 .028 
 
 .90 
 
 .641 
 
 4.40 
 
 .481 
 
 7.80 
 
 54 
 
 .00 
 
 693 
 
 4-50 
 
 504 
 
 8.00 
 
 .079 
 
 The following table will facilitate calculation of engine-power 
 required for compression. These values are to be divided by 
 m, the modulus of the compressor, which is for 0.50 to 0.70, to
 
 THE COMPRESSION OF AIR. 
 
 315 
 
 determine the steam-power required to drive the piston against 
 friction. 
 
 TABLE SHOWING THE HORSE-POWER REQUIRED TO COMPRESS AND DELIVER 
 ONE CUBIC FOOT or FREE AIR PER MINUTE TO VARIOUS GAUGE PRESSURES; 
 ALSO THE POWER REQUIRED TO COMPRESS AND DELIVER ONE CUBIC FOOT OF 
 AIR AT THE GIVEN PRESSURE. 
 
 
 Compressing One Cubic Foot of 
 Free Air per Minute to 
 Given Pressure. 
 
 Delivering One Cubic Foot per Minute 
 of Air Compressed to the 
 Pressure Given. 
 
 I 
 
 
 
 Gauge 
 
 
 
 
 
 Pressure. 
 
 2 
 
 3 
 
 4 
 
 S 
 
 
 Compression at 
 Constant 
 
 Compression 
 without 
 
 Compression at 
 Constant 
 
 Compression 
 without 
 
 
 Temperature 
 
 Cooling. 
 
 Temperature. 
 
 Cooling. 
 
 5 
 
 .01876 
 
 .01963 
 
 .02514 
 
 .0263 
 
 10 
 
 033 2 5 
 
 .03609 
 
 .05586 
 
 .06399 
 
 IS 
 
 .04507 
 
 .05022 
 
 .09105 
 
 .10145 
 
 20 
 
 .05506 
 
 .06283 
 
 .12994 
 
 .14829 
 
 25 
 
 .06366 
 
 .07422 
 
 .17191 
 
 .20043 
 
 30 
 
 .0713 
 
 .08464 
 
 .21678 
 
 25734 
 
 35 
 
 .0782 
 
 .09425 
 
 .26445 
 
 .31872 
 
 40 
 
 .084305 
 
 .10324 
 
 31375 
 
 .38422 
 
 45 
 
 .08954 
 
 ,IIl66 
 
 .36368 
 
 45353 
 
 50 
 
 .09508 
 
 .11952 
 
 .41848 
 
 52605 
 
 55 
 
 .09936 
 
 .12702 
 
 .47H2 
 
 .60227 
 
 60 
 
 . 10402 
 
 .13418 
 
 5 2 855 
 
 .68181 
 
 65 
 
 .10808 
 
 .14028 
 
 .58612 
 
 .76079 
 
 70 
 
 .11245 
 
 .14718 
 
 .64812 
 
 .8483 
 
 75 
 
 .11629 
 
 15373 
 
 .70952 
 
 93795 
 
 80 
 
 .11926 
 
 I597I 
 
 .76843 
 
 1.02906 
 
 85 
 
 .1224 
 
 l6 555 
 
 83039 
 
 1.1231 
 
 90 
 
 . 12558 
 
 .17096 
 
 .89444 
 
 1.2176 
 
 95 
 
 .12886 
 
 .17629 
 
 .96164 
 
 1.3148 
 
 100 
 
 .13121 
 
 18153 
 
 1.0243. 
 
 1.4171 
 
 EXAMPLES. i. Required the number of units of work necessary to compress 
 80 Ibs. of free air from 14.7 Ibs. absolute pressure and 60 F. to a pressure 
 of 88.2 Ibs. per square inch. Here P= 2116.8; 7=80X1308=1046; 
 Fj=o.i66X 1046= 174.3. 
 
 Neglecting friction and the value for the clearance in the cylinder, 
 JF,= 21 16.8X1046X1. 79= 3,963,320 ft.-lbs. Compressed adiabatically, T= 
 60 + 461 = 521, and T I} according to the previous table, is nearly 881, or by 
 computation log t^log 521+0.29 log (88.2 X 144) 0.29 log (2116.8) = 
 294254. Whence ^=876.1, T7 a =i83.5(T 1 -T)x8o=5,2ii,5oo ft.-lbs. 
 By the preceding table the work is respectively 121.65 an d 160.77 h-P- f r * 
 final compression to 89.7 Ibs. absolute. 
 
 2. 100 cubic feet of air per minute are to be compressed to 50 Ibs. gauge
 
 316 MANUAL OF MINING 
 
 without cooling. Required the work expended in the operation. The initial 
 temperature of the free air is 60; then ^=14.7; P 1 =64-7; F=ioo-, 
 7i=35-; T= 5 2i; ^-339 F. 
 
 Neglecting clearance, the value for the work per cubic foot, according to 
 the table, is 0.11952 h.p., and for 100 cubic feet 11.952 h.p. 
 
 100 
 Wa= 183.45(800-521)^-^= 11.9 X33,ooo ft.-lbs. 
 
 If the compression were conducted with perfect cooling, then, according 
 to the table, there would be required 9.5 horse-power and the final volume 
 V, would be 22.7. 
 
 Wi= 2116.8X100 hyp log 4.401 = 9.504X33,000 ft.-lbs. 
 
 3. What should be the size of the cylinders to compress 100 Ibs. free 
 air per minute from 60 F. to 100 Ibs. gauge pressure? r=52i abs.; 
 PI= 114.7X144; 7=1308 cubic feet; v=^ per cent V; w=ioo; ^=946 F. 
 
 PPa= 183.45(946-521)100= 7,797,000 ft.-lbs.; 
 Wi= 2116.8(1.07X1308) hyp log ^^-(14,400)91.5=4,843,300 ft.-lbs. 
 
 If the modulus, m, be taken at 0.70, the work to be supplied the steam- 
 piston is 11,138,571 ft.-lbs. and 6,919,000 ft.-lbs. respectively. With an 
 actual capacity of cy Under 0.90 that of the apparent capacity, there being 
 80 strokes per minute, the cylinder capacity is, for isothermal compression, 
 16.35 cubic feet. With 7 per cent clearance the cushioned air will correspond 
 to 0.3027 of free air. Each piston displacement is then (1.3020.07)8.175 = 
 10.07 cubic feet. The cylinders are 23f"X36" stroke each. 
 
 The mean effective resistance is 30.07 Ibs. per square inch. The steam- 
 cylinder must be i2$"X36" if the average cut-off is , initial pressure 120 Ibs. 
 gauge, and the back pressure 5 Ibs. Then, Chapter V, 3 per cent clearance, 
 
 m.e.p.=o.8658(i2o+ 14. 7) 5=116.6; 
 
 g= 6,919,000 
 
 33,000 
 whence 
 
 jfe= 12.54 inches and 5= 3 feet. 
 
 The Mean Resistance to Compression of dry air during com- 
 pression can be ascertained. Let P l be the final pressure and 
 P the initial pressure; then
 
 THE COMPRESSION OF AIR. 
 
 3*7 
 
 Fig. 123 is a diagram showing on the right the temperatures 
 attained by air when compressed adiabatically to the degree 
 indicated at the bottom of the figure, and on the left the work 
 in foot-pounds expended during compression. The difference 
 in power consumption by adiabatic compression and isothermal 
 compression is pictorially revealed. Thus a point on the iso- 
 thermal compression line corresponding to 6 atmospheres indi- 
 cates the work expended on one pound as 50,000 ft.-lbs.; upon 
 
 7 8 
 
 PRESSURES IN ATMOSPHERES 
 
 nder 
 
 FIG. 123. Diagram Showing the Work Done During Air-compression u 
 Various Conditions. 
 
 one pound at the same pressure adiabatic there has been ex 
 pended 73,300 ft.-lbs. and the temperature attained is 460 F. 
 
 Cooling the Air During Compression. Isothermal compression 
 is more desirable to attain than adiabatic compression, not only 
 because the power required is less, but also because the high 
 temperature of the latter is particularly injurious to the machinery. 
 It is an obstacle to rapid running, as proper lubrication cannot 
 be maintained. 
 
 Isothermal compression is too slow. Moreover, the cylinder 
 walls are never perfect conductors. So the air is taken as cool
 
 MANUAL OF MINING. 
 
 CO 
 
 w o
 
 THE COMPRESSION OF AIR. 319 
 
 as possible chilled if may be and kept cool during all stages 
 of the compression by artificial means. As the air will cool 
 rapidly after compression, producing a great loss by radiation 
 from cylinders, receivers, conducting pipes, and all intermediate 
 appliances before reaching the motor, and as it is impossible to 
 retain this heat in spite of every precaution, it is desirable to 
 extract the excessive heat as promptly as possible. So the supply 
 of air is taken from outdoors at the coolest side of the building, 
 effecting a saving of at least 2 per cent in power by this choice 
 of site, which may be 10 cooler than on the other side. It may 
 be noted, too, in the table and the diagram, Fig. 123, that the 
 increment of heat is greater in the earlier portion of the stroke. 
 During the first 15 Ibs. increase the temperature rises 118 F., but 
 during the increment from 90 to 105 Ibs. the rise is only 37 F. 
 
 Water-cooling. Cooling is effected by a circulating current 
 of water in order to maintain an isothermal compression, in 
 which case there will be extracted from the air under com- 
 pression an amount of heat corresponding to the area of the 
 space between the two curves in Fig. 123, or an amount of work 
 represented by the difference in the vertical lines between the 
 isothermal and adiabatic curves in diagram Fig. 122. The 
 cooling is a direct waste, and heat once removed from the air is 
 never returned to it. The absorption of heat is accomplished 
 by cold air circulating through a jacket, ;', Fig. 128, surrounding 
 the cylinder. In Fig. 124, the course of the water through it is 
 represented by the arrows and passages numbered i, 2, 3, and 4. 
 
 A fine spray of water injected into the cylinder would be a 
 more effective method of cooling, were it not so objectionable 
 because of its corrosive action on the cylinder walls and its freez- 
 ing when, at some later stage, it is used expansively in the motor. 
 The spray-injectors are more efficient coolers than water-jackets, 
 as shown in the diagram Fig. 125, the former curves being lower 
 than the latter. 
 
 The circulating water should be the coolest attainable, and 
 the cylinder lining as perfect a conductor as possible, to extract 
 the appreciable amount of heat.
 
 320 MANUAL OF MINING. 
 
 Two-stage Compression. The very short time of contact 
 between the water and the hot air may be increased by dividing 
 the air- compression between two cylinders (Fig. 126), and in 
 addition passing it through the intercooler. This latter is very 
 efficient, saving as much as 10 per cent of the power, which almost 
 counterbalances the friction of the machine. This compound, 
 or two-stage, compression gives nearly treble the time of con- 
 tact for the circulating waters. In the larger cylinder the air 
 is compressed to one third or one fourth its volume, and in the 
 
 FIG. 125. Combined Indicator Diagrams from a Two-stage Air-compressor. 
 
 small cylinder to that required. The intercooler is essentially 
 a cylinder with small brass pipes through which water circulates 
 and around which air flows from the low- to the high-pressure 
 cylinder. In this way a stepped curve, diagram Fig. 125, is 
 obtained indicating economical results approaching an isothermal 
 compression. From i to 3 horse-power are saved over the 
 single-stage compression for each 100 cubic feet of free air at 
 the ordinary degrees of compression used in mining. Two- 
 stage compression is not advisable below 4 atmospheres final 
 pressure, but is imperative for high degrees of compression. 
 But the greater the degree of compression, the greater is the 
 amount of abstracted heat and the more serious becomes the 
 loss of power. For i, 2, 3, 4, 5, and 6 atmospheres, gauge reading, 
 the losses are 28%, 37%, 46%, 50%, 53%, and 56%, respectively,
 
 THE COMPRESSION OF AIR. 
 
 3 2I 
 
 of the original power. Economic work is best obtained by operat- 
 ing at as low pressure as is consistent with the work. 
 
 HORSE-POWER NECESSARY TO COMPRESS 100 CUBIC FEET OF FREE AIR TO 
 
 VARIOUS PRESSURES AND WITH Two-, THREE-, AND FOUR-STAGE 
 
 COMPRESSORS. 
 
 
 Horse-power Necessary. 
 
 
 Horse-power Necessary. 
 
 
 
 Gauge 
 
 
 Pressure. 
 
 
 
 
 Pressure. 
 
 
 
 
 
 Two- 
 
 Three- 
 
 Four- 
 
 
 Two- 
 
 Three- 
 
 Four- 
 
 
 stage. 
 
 stage. 
 
 stage. 
 
 
 stage. 
 
 stage. 
 
 stage. 
 
 100 
 
 15-7 
 
 15-2 
 
 14.2 
 
 900 
 
 36.3 
 
 33-7 
 
 31.0 
 
 200 
 
 21.2 
 
 20.3 
 
 18.8 
 
 1000 
 
 37-8 
 
 34-9 
 
 31.8 
 
 300 
 
 24-5 
 
 23.1 
 
 21.8 
 
 I2OO 
 
 39-7 
 
 36.5 
 
 33-4 
 
 400 
 
 27.7 
 
 25-9 
 
 24.0 
 
 I4OO 
 
 41-3 
 
 37-9 
 
 34-5 
 
 500 
 
 29.4 
 
 27.7 
 
 25-9 
 
 1600 
 
 43-o 
 
 39-4 
 
 35-6 
 
 600 
 
 3 1.6 
 
 29-5 
 
 27.4 
 
 1800 
 
 44-3 
 
 40-5 
 
 36.7 
 
 700 
 
 33-4 
 
 31.2 
 
 28.9 
 
 2 COO 
 
 45-4 
 
 41.6 
 
 37-8 
 
 800 
 
 34-9 
 
 32.5 
 
 30.1 
 
 2500 
 
 
 43-o 
 
 39-o 
 
 The Work of Compressing Moist Air is less than that re- 
 quired during the comp.ression of dry air, nor do the mean net 
 resistance and the temperature increase as rapidly. 
 RISE IN TEMPERATURE AND WORK OF COMPRESSION OF DRY AND MOIST AIR. 
 
 
 Temperature. 
 
 Work on One Pound. 
 
 Absolute 
 
 
 
 Pressure. 
 
 
 
 
 
 
 Dry. 
 
 Moist. 
 
 Dry. 
 
 Moist. 
 
 14.7 ' 
 
 68 F. 
 
 68 F. 
 
 
 
 
 
 22 
 
 133-8 
 
 94 
 
 13.300 
 
 13,200 
 
 29.4 
 
 185.9 
 
 in 
 
 23>5 
 
 22,500 
 
 36-7 
 
 229.5 
 
 124 
 
 30,500 
 
 29,000 
 
 44.1 
 
 266.7 
 
 135 
 
 37,000 
 
 35,ooo 
 
 51.4 
 
 3P.2 
 
 145 
 
 43,200 
 
 40,600 
 
 <S.8 
 
 330-1 
 
 153 
 
 48,500 
 
 45,ooo 
 
 73-5 
 88.2 
 
 383-5 
 428.9 
 
 167 
 179 
 
 58,500 
 67,160 
 
 52,500 
 60,000 
 
 The mean effective resistance or pressure during compres- 
 sion or expansion of moist air is 
 
 and 
 
 Kp \ 0.166 ) 
 
 P7 ~ x ?
 
 MANUAL OF MINING.
 
 THE COMPRESSION OF AIR. 
 
 323 
 
 The Air-compressor. The simple principle of air-compres- 
 sion is difficult of execution. To obtain a compact high-speed, 
 uniform, rapid- cooling, efficient compressing engine is not easy. 
 
 The essentials are partly secured in various ways by the several 
 successful patterns now on the market. 
 
 The piston of ihe air- compressor may be driven directly by 
 a steam- cylinder at the other end of the same rod (in tandem),
 
 324 
 
 MANUAL OF MINING. 
 
 or " straight line," or it may be alongside of and joined by a 
 common cross-head with the steam-piston. Such an engine is 
 called a " cross" air-compressor. It may be driven by a belt or 
 by a reducing- gear from a shaft actuated by electricity or steam- 
 power, or an impulse-wheel mounted directly upon the main 
 shaft of the compressor. The cylinders are usually horizontal and 
 
 Flo. 128. A Water-jacketed Air-compressor, IngersoU Pattern. 
 
 may be single or duplex, and the air may be compressed in one 
 stage with one cylinder, or in two stages, corresponding to simple 
 and compound steam-engines. The duplex engines are simple, 
 inhaling and compressing separate volumes of air. They have 
 none of the merits of the compound or two-stage engines, in which 
 the large low-pressure cylinder expels its air into the small 
 high-pressure cylinder. The cranks of a duplex compressor are 
 at quarters to equalize the rotary effort. 
 
 The frame, compared with that of the steam-engine, is very 
 solid and on heavy base-plate Compressors may be had sectional 
 for convenience in transportation. They are fitted with auto- 
 matic throttling governors to alter the steam pressure according 
 to the demand for power, and are so connected that when the 
 delivery air attains an excessive pressure it will operate on the 
 piston, which raises ihe regulator and closes the governor- valve, 
 thus reducing compressor speed. Their action is very much like 
 that of the automatic damper regulator mentioned in Chapter IV.
 
 THE COMPRESSION OF AIR. 
 
 325
 
 326 MANUAL OF MINING. 
 
 High speed to the piston is advantageous for the economy of 
 steam and for capacity, but because of the rapid wear and the 
 difficulties with large valve areas, as well as the inordinate resist- 
 ance developed thereby, the velocity is less than 300 feet per 
 minute, except in the larger sizes. 
 
 The Burleigh is upright, its air-cylinder is single-acting, and 
 its peculiarity lies in the admission of steam one eighth of a stroke 
 before the air. The Waring has a bonnet, or conical valve, like 
 that in Fig. 127, whose pistons are moved by a rocker on the fly- 
 wheel shaft, the steam- cylinder being set at an angle to the hori- 
 zontal air-cylinder. The Clayton has the usual poppet-valve, 
 and is a compact machine, with its fly-wheel centrally located. 
 The Delamater has an important contrivance for dropping the 
 discharge-valve from its seat. This form is very heavy. The 
 Sullivan, Ingersoll, Norwalk, and Rand are the popular pneu- 
 matic machines. 
 
 The Valves of the Compressor are of the poppet, spindle, or 
 ring pattern. Whatever their form, they should open quickly, 
 have a full lift, and be ample in size. Large inlet-valves offer 
 little difficulty, though for a short time they are subject to full 
 reservoir pressure. An unrestricted entry for the air is obtained 
 easily by the use of poppets held by springs. The Ingersoll- 
 Sergeant compressor admits air through a hollow piston and 
 rod (Fig. 128). This provides a very liberal inlet area, G, en- 
 ables the cylinder-covers, J, to be completely water- jacketed, and 
 leaves more room for the discharge-valves, H. The Norwalk 
 employs a rotary valve. The inlet-valves of the Rand are shown 
 at , g' , Fig. 127. They are provided with guards that prevent 
 their falling into the cylinder. 
 
 The valves should be positive, and this the poppet attains, 
 though the tendency to "chattering" is the serious objection to 
 it, particularly for discharge-valves. This arises from the two 
 opposing effort one, of the air, to open, and the other, of the 
 spring, to close the valve. The valve-gear shown in Fig. 124 
 does away with this trouble in the high-pressure engines; the 
 arms, a, b, relax the spring pressure and allow of the valve rising
 
 THE COMPRESSION OF AIR. 327 
 
 full-lift without dancing. Poppet-valves can hardly be im- 
 proved upon for low pressures, though their springs in time lose 
 elasticity and open too soon. This reduces their efficiency, as 
 also does any slip of the valves. In the Norwalk pattern (Fig. 
 129) a positive discharge is obtained by moving the valve by 
 cams, such that it remains at rest till the pressure is sufficient to 
 open it quickly. A difficulty about this, it would seem, is that, 
 as the reservoir pressure constantly varies (unless perfectly regu- 
 lated), the valves must receive constant attention. This may 
 be corrected by an automatic governor like the Corliss release, 
 which will open the valves at different points in the stroke as 
 desired. 
 
 The discharge-valves require careful construction, for their 
 leakage is equal to a large clearance space. They are made 
 large, and, to prevent inordinate frictional loss and wear, are as 
 numerous as possible. An excess of air pressure over the re- 
 ceiver pressure is necessary to open the valves and expel the air. 
 This unavoidable loss has an important bearing upon the uni- 
 formity of speed. 
 
 The Effect of the Clearance Space. The large clearance 
 space between the piston and the cylinder-head at the end of 
 the stroke constitutes one inevitable source of reduction of cylin- 
 der capacity. Not only is the compressed air filling that space 
 never discharged, but on the forward-stroke it will expand and 
 fill a volume that should have been occupied by fresh atmos- 
 pheric air which is being inhaled. 
 
 If the air in the clearance is compressed on the return-stroke 
 to 60 Ibs. and occupies 7 per cent of the cylinder volume, it is 
 evident that in the forward-stroke it expands to five times the 
 original volume, or nearly one-third that of the cylinder, before any 
 fresh cool air whatever can be admitted. The effective volume 
 of the cylinder is therefore materially reduced. Moreover, this 
 air is hot, and, the cylinder also being warm, such inlet air as does 
 enter is expanded to correspondingly reduce the capacity of the 
 cylinder. The clearance space cannot be made smaller than T V 
 of an inch, and is usually much more. The only remedy for this
 
 328 MANUAL OF MINING. 
 
 loss is an increased length of stroke or compounding the cylin- 
 ders. At 75 Ibs. absolute pressure a single cylinder must be 
 three times as long as a compound having the same clearance 
 loss. 
 
 The friction of the air-piston taken with that of the valves 
 is about equal to 10 per cenc of the work of the engine, and 
 may icach 25 per cent when including the losses in the steam- 
 engine. 
 
 The Horse-power of the Compressor. This is determined, 
 in the same general manner as for the steam-engine, by the indi- 
 cator cards or by assuming a mean effective pressure, obtained 
 from column 6 of the table of volumes, and substituting the 
 same in the general expression for horse-power. The mechanical 
 efficiency of the air-compressor m must be taken at not more 
 than 60 per cent for the one- stage and 70 per cent for the two- 
 stage cylinder, measured in the terms of the energy possessed by 
 the compressed air delivered from the machine, compared with 
 the steam-motor power. 
 
 By inspection of the diagram Fig. 122 it is seen that the 
 air resistance is atmospheric at the beginning of the stroke when 
 the steam-power is at the maximum. Toward the end of the 
 stroke it is at a maximum, while the expanded steam is at its mini- 
 mum pressure. This large engine excess at the beginning, and 
 resistance at the end, necessitate the use of heavy fly-wheels. 
 Indeed, the fluctuating stresses to which the entire compressor 
 is subject requires it to be built excessively heavy. In the two- 
 stage compressor this inequality of engine excess is not so marked. 
 
 The Receiver. A receiver is a necessity in all compressed- 
 air systems. It is a huge tank with regulating devices to main- 
 tain uniform pressure. In a measure it is a power accumulator, 
 if of suitable size ; it compensates for the pulsating effect of each 
 stroke of the compressor, and for this purpose should be within 
 50 feet of the compressor. Another ought to be placed near to 
 the machine drills to reduce friction losses. This will serve also 
 as a drain, if water is present, to catch the water condensing in 
 the pipes.
 
 THE COMPRESSION OF AIR. 329 
 
 Further cooling ensues as the compressed air is stored in 
 receiver or pipes, but represents a total loss of power for which 
 there is no compensation. A pound of . air not cooled in the 
 compressor, when at 80 Ibs. absolute, would radiate 81 B.T.U. 
 in cooling to its initial atmospheric temperature of 60 F., and 
 thus dissipate 63,000 ft.-lbs. of work. Perfect cooling in the 
 cylinder would have saved this amount of steam-power to the 
 motor. 100 cubic feet per minute would thus have saved 16 
 horse-pcwer, 350 Ibs. of steam, and 45 Ibs. of coal per hour. 
 
 Transmission of Compressed Air. The air is conveyed to 
 the drill, coal-cutters, hoist, etc., by pipes. With the exception 
 of electricity, no other means of power transmission can compare 
 in efficiency with compressed air. The diameter of the pipe 
 is a matter of the first importance. 
 
 The transmission losses appear in two ways : as loss of power 
 and as loss of pressure, or head, indicated by the difference in 
 gauge reading at the ends of the line. There is a distinction 
 between these two losses. The first is the larger, due to cooling, 
 of the air during compression, and is unavoidable and not charge- 
 able to transmission. Of the power remaining, some of it is 
 lost by subsequent cooling and some in overcoming the frictional 
 resistances in the pipes. The power depends upon its pressure 
 and its volume. In the process of transmission the pressure is 
 reduced by the frictional losses, but there is a certain compensa- 
 tion from a corresponding increase in the volume. The actual 
 loss of power from this cause is, therefore, slight in ordinary 
 mining conditions. 
 
 Frictional Resistance in Pipes. The loss of pressure, or of 
 head, due to frictional resistances takes place according to laws- 
 governing the flow of fluids. If the pipes be short, the velocity 
 will vary inversely as the area, and the frictional loss will be 
 directly proportiona" to the square of the velocity of the flow. 
 It will be also proportional to the periphery and the length of 
 the conduit. But in long pipes the expression becomes com- 
 plex. Tables of the loss of pressure by flow in pipes are given by 
 manufacturers, and it will be found therein that air at 32.8 feet
 
 33 
 
 MANUAL OF MINING. 
 
 per second loses 8.26 Ibs. pressure in a mile of ic-inch pipe, 
 10.04 Ibs. in an 8-inch pipe, and 20.08 Ibs. in a 4-inch pipe. 
 
 The table below, from the handbook of the Norwalk Iron 
 Company, shows the losses of pressure for given volumes and 
 velocities in pipes 1000 feet long. 
 q = volume of free air passing per minute at 60 Ibs. gauge pressure ; 
 
 ^ II II (( II II It Q_ It II 
 
 p = the loss of pressure-head in pounds per square inch; 
 v = the velocity in feet per second. 
 
 TABLE OF VOLUMES AND PRESSURE-HEAD Loss, TRANSMITTING FREE AIR 
 THROUGH PIPES 1000 FEET LONG. 
 
 Vel. 
 
 3 Inches. 
 
 4 Inches. 
 
 6 Inches. 
 
 10 Inches. 
 
 V 
 
 P 
 
 9 
 
 <f 
 
 P 
 
 9 
 
 of 
 
 P 
 
 9 
 
 ^ 
 
 P 
 
 a 
 
 <t 
 
 J.a8 
 
 .046 
 
 48 
 
 60 
 
 134 
 
 86 
 
 109 
 
 .023 
 
 193 
 
 244 
 
 .014 
 
 537 
 
 680 
 
 6.56 
 
 .209 
 
 Q6 
 
 121 
 
 152 
 
 172 
 
 217 
 
 .104 
 
 386 
 
 488 
 
 .064 
 
 i73 
 
 1359 
 
 0.84 
 
 .488 
 
 144 
 
 182 
 
 .360 
 
 358 
 
 326 
 
 244 
 
 579 
 
 633 
 
 145 
 
 1610 
 
 2039 
 
 13.12 
 
 .838 
 
 193 
 
 243 
 
 .628 
 
 343 
 
 436 
 
 .419 
 
 772 
 
 977 
 
 .20 
 
 2146 
 
 2719 
 
 16.40 
 
 I-3I7 
 
 241 
 
 34 
 
 .982 
 
 429 
 
 544 
 
 .6 S 8 
 
 96s 
 
 1221 
 
 393 
 
 2683 
 
 3399 
 
 19.68 
 
 1. 808 
 
 289 
 
 ^64 
 
 i-35 6 
 
 515 
 
 6S3 
 
 .904 
 
 1158 
 
 1466 
 
 542 
 
 3220 
 
 4079 
 
 26.24 
 32.80 
 
 3-352 
 5.270 
 
 386 
 480 
 
 4 86 
 60 7 
 
 2-513 
 3.928 
 
 687 
 859 
 
 871 
 1088 
 
 1.670 
 2-635 
 
 1544 
 i93i 
 
 1954 
 2443 
 
 1.024 
 i-573 
 
 4293 
 5367 
 
 5438 
 6798 
 
 For any degree of c.ompression p", other than 60 Ibs. gauge, 
 the quantity of free air passing per minute would be obtained by 
 the ratio 74.7 : ^'+14.^. This is approximately correct. 
 
 EXAMPLE. Required the volume of free air which at 70 Ibs. gauge pres 
 sure can be carried by a 6-inch pipe with a friction loss of only 0.658 Ibs. per 
 square inch. From the table, 965 cubic feet of free air compressed to 60 Ibs. 
 can be carried; hence, with air at 70 Ibs. pressure, the volume is 965 X 84. 7 -s- 
 74.7=1094 cubic feet free air. 
 
 The Loss of Energy Due to Friction. The measure of the 
 work possessed by air is represented by that expended in produc- 
 ing the given conditions of temperature, pressure, and volume. 
 Any loss in either of these elements will necessarily diminish its 
 capacity. During compression heat has been abstracted and 
 the capacity for work is measured by the amount of intrinsic
 
 THE COMPRESSION OF AIR. 33 1 
 
 energy remaining. Its value may be obtained from the usual 
 formulae; the pressure, P, and temperature, T, at the entrance 
 to the pipe being known. The volume occupied by one pound of 
 air is then ascertained: 
 
 P, ^ = 53.181. 
 
 Then the relation between the condition of the air at the 
 entrance to the pipe to that at the point of delivery to the motor 
 is ascertained from the formula 
 
 PI/ 1.408 _ p y 1.408 
 1^1 -^2 l/ 2 
 
 If the temperature has not changed during transmission, the 
 volume and pressure will bear the same relation at the pipe exit 
 as at the entrance, and P 2 F 2 = 53.i8r 1 . A drop in the pressure 
 ensues, however, due to friction of flow, and this produces a 
 slight reduction in the available energy; that is, P 2 is less by 
 the frictional loss in the preceding table and the volume F 2 has not 
 proportionately increased, making PyVj 1 ' 408 less in value than 
 that of PjTY' 408 by the amount of energy lost. 
 
 The friction in the pipe varies as the velocity and the volume 
 of the air. For a given power the degree of compression may 
 be increased in order to reduce the volume; or the size of the 
 pipe may be increased to reduce the velocity. The efficiency 
 of the system will be increased by either plan. A 6-inch pipe, 
 for example, will carry 800 cu. ft. of free air at 80 Ibs. pressure 
 absolute a distance of 5000 feet with a loss of i Ib. per square 
 inch. For the same volume in a 4^- inch pipe the loss of pressure 
 is 5.362 Ibs. per square inch. These pipes would therefore 
 require receiver pressures at the pipe entrance of 81 and 85.3 
 Ibs., respectively, per square inch. The saving in power with the 
 6-inch pipe over the 4|-inch pipe will therefore be 3 horse-power. 
 
 In determining the loss of head due to given conditions, the 
 following formula is available: 
 
 n _ V*L 
 io,oooD 3 a'
 
 MANUAL OF MIXIXG. 
 
 in which H =head or difference of pressure required to overcome 
 
 friction and maintain the flow of the air; 
 V= volume of compressed air delivered in cubic feet 
 
 per minute; 
 
 L = length of pipe in feet; 
 D =nominal diameter of pipe in inches; 
 a = coefficient depending upon the size of the pipe. 
 VALUES FOR a AND Z> 5 a FOR VARIOUS DIAMETERS OF PIPE. 
 
 Nominal 
 
 a 
 
 a 
 
 Nominal 
 
 a 
 
 
 D 
 
 
 
 D 
 
 
 
 l" 
 
 o-35 
 
 0-35 
 
 3" 
 
 0-73 
 
 177-4 
 
 Ij 
 
 
 I-525 
 
 
 0.787 
 
 413-2 
 
 
 0.662 
 
 5-03 
 
 4 
 
 0.84 
 
 860.2 
 
 2 
 
 4 
 
 0-565 
 0.65 
 
 18.08 
 
 1 
 
 0-934 
 
 I. 00 
 
 2918.8 
 7776.0 
 
 The Most Economical Size of Pipe. The friction in properly 
 designed pipe systems is not a serious matter and can be made as 
 small as the most exacting requirements demand, by enlarging 
 the pipe or securing a smooth interior. The capacity of a pipe 
 is somewhat proportional to its cross-sectional area, but is affected 
 by the character of its interior surface and the various couplings 
 used. In calculating the size of pipe required, due regard must 
 be paid to the commercial sizes, for their actual diameter is very 
 different from their nominal diameter. The velocity of the air in 
 the main pipes should not exceed 25 feet per second, and the 
 dimensions should be determined accordingly. At a higher rate 
 the friction becomes excessive and the power lost in overcoming 
 it too large. Though the rate of flow through the pipe is continually 
 increasing from beginning to end, all calculations for frictional 
 loss should bear this in mind, though it may be neglected if the 
 drop in pressure is small. 
 
 With due regard to economy in installation, the size of the 
 pipe should therefore be made as large as advisable to reduce 
 friction, but need not be increased beyond that requisite to sup- 
 ply a flow at 25 feet per second. This gives, for ordinary mining 
 practice, a diameter of 4 inches for the mains and not less than
 
 THE COMPRESSION OF AIR. 
 
 333 
 
 2 inches for the stopes and rooms. A 4-inch pipe with air at 
 82-lbs. gauge will supply five 3-inch drills 3000 feet away. 100 
 feet of i^-inch pipe will serve for only one drill. 
 
 Compressed-air Pipes. The pipes used are steel-riveted or 
 lap- weld, as illustrated in Figs. 131 and 132. The joints should 
 be carefully secured. Means must not be neglected for. pro- 
 viding for changes in length due to alternations of temperature. 
 Iron expands 0.000007 its length per i F. This allowance is 
 more essential above than below ground; and in shafts where 
 the temperature is inconstant compensation-joints are neces- 
 sary. At every 300 or 400 feet a copper U tube is attached; ;is 
 flexibility will allow for contraction or expansion of the pipes. 
 At the Republic mine the brass- lined expansion- joints every 
 500 feet allow for movements of 12 inches. They rest on gas-pipe 
 rollers. The Chapin iron-mine has expansion-joints at every 
 680 feet of the 24-inch pipe. 
 
 To reduce the frictional losses in the transmission large 
 elbows and bends of long radii should be used. The joints must 
 be made very carefully, to reduce leakage, unless the air-pipes are 
 laid in ventilation-passages of the mine. Leaks when discovered 
 must be plugged. The velocity of escape of compressed air 
 being 20,000 feet or over per minute, a very large loss of fluid 
 will ensue from a neglected leak. 
 
 STANDARD STEAM AND EXTRA-STRONG PIPE USED FOR COMPRESSED-AIR 
 HAULAGE PLANTS. 
 
 Trade 
 Diameter, 
 Inches. 
 
 Cubic Feet 
 in 
 i Lineal 
 Foot. 
 
 Lineal 
 Feet 
 Necessary 
 to Make 
 i Cubic 
 
 Steam. 
 
 Extra Strong. 
 
 Actual 
 Diameter, 
 Inches. 
 
 Thick- 
 
 Weight 
 
 Thick- 
 
 Weight 
 
 
 
 Foot. 
 
 ness. 
 
 per Foot. 
 
 ness. 
 
 per Foot. 
 
 
 2 
 
 .0218 
 
 45-41 
 
 15 
 
 3.61 
 
 .22 
 
 5.02 
 
 2.067 
 
 4 
 
 .0341 
 
 29.32 
 
 .20 
 
 5-74 
 
 .28 
 
 7.67 
 
 2.468 
 
 3 
 
 .0491 
 
 20.36 
 
 .21 
 
 7-54 
 
 3 
 
 10.20 
 
 3.067 
 
 3i 
 
 .0668 
 
 15 .00 
 
 . 22 
 
 9.00 
 
 32 
 
 12.50 
 
 3-548 
 
 4 
 
 .0873 
 
 11.52 
 
 23 
 
 10.70 
 
 34 
 
 15.00 
 
 4.026 
 
 4} 
 
 .1105 
 
 9-05 
 
 .24 
 
 12.30 
 
 35 
 
 I7.6o 
 
 4.508 
 
 5 
 
 .1364 
 
 7-33 
 
 25 
 
 14.50 
 
 37 
 
 2O. 50 
 
 5-45 
 
 Si 
 
 165 
 
 6.06 
 
 .26 
 
 16.40 
 
 .40 
 
 24.50 
 
 5.28 
 
 6 
 
 .1963 
 
 5.10 
 
 .28 
 
 18.80 
 
 43 
 
 28.60 
 
 6.065 
 
 
 
 ! 
 
 
 

 
 334 MANUAL OF MINING. 
 
 Sleeve couplings are used in all pipe lines except the wrought 
 iron, which are spiral-riveted or welded tubes. The joints are 
 carefully made, and leaks avoided with the greatest of care. 
 Elbows of as liberal a radius as possible must be provided to 
 have the frictional resistance small. 
 
 The Power Value of Compressed Air. In determining the 
 power value of compressed air it must be remembered that the 
 mean effective pressure of the air is lower than that of steam 
 for a given cut-off and initial pressure. 
 
 The mean effective pressure during expansion can be ascer- 
 tained by substitution in the formula page 316, P being somewhat 
 near the atmospheric pressure. It is less in amount than is 
 the m.e.p. of steam during a similar expansion, as also is its 
 terminal pressure. The work it is capable of equals the work 
 expended upon it if the friction losses be neglected and there 
 were no loss in cooling, for the cycles of changes which the air 
 would experience in expanding are duplicates of those during 
 compression. Much power has, however, been extracted from 
 the air, whose high temperature had been reduced nearly to that 
 of the surrounding atmosphere. 
 
 The Work Performed while Expanding from a pressure P^ 
 and a temperature T X , which latter is near that of the atmosphere, 
 to a temperature r and a pressure P, which is nearly 15 Ibs., is 
 
 JF= 183.45(^1- *)w. 
 
 The final temperature, T, which is far below that of the atmos- 
 phere, produces an intense refrigeration, which is objectionable 
 if moisture is present. The moisture will be frozen in the process 
 and may clog the exhaust-passages. As this is usually the case, 
 the air is reheated so high that after expansion it will still be 
 warmer than the atmosphere. The heat thus expended has the 
 effect of increasing, in proportion to the heat added, the volume, 
 and thereby the amount of work obtainable from the air. 
 
 Reheating. Reheating may be accomplished by direct fire, 
 or by steam, passing through a pipe inside of the air-pipe. Steam 
 thus used gives up all of its latent heat, which can be converted
 
 THE COMPRESSION OF AIR. 335 
 
 into an amount of work in the air-engine far exceeding what 
 could be derived from it if used directly in the steam- pump. 
 
 A reheater-jacket is sometimes used where other means 
 are not available or desirable. This method would be preferable 
 for compound air-engines which might otherwise require re- 
 heating in two stages. 
 
 The Efficiency of Compressed Air. As the efficiency of the 
 driven machine is not over 60 per cent and the efficiency of the 
 compressor is not over 75 per cent, it is evident that, with the 
 losses ensuing in pipes, the aggregate efficiency of the combina- 
 tion does not exceed 25 or 30 per cent of the original motor- 
 power. This, compared with the fuel burning at the boiler, 
 represents an efficiency of about 2 J to 3! per cent-; in other words, 
 the work performed at the drill is equivalent to from 275 
 to 385 B.T.U. per pound of coal. Notwithstanding this waste- 
 fulness, compressed air serves well for many purposes in mining, 
 and will retain its place even against electricity. 
 
 The Cummings System of Air Transmission. This is a closed 
 line of pipes between the compressor and the air-engine. The 
 initial absolute pressure of the air entering the compressor and 
 the final pressure of exhaust from the engine are about 80 Ibs., 
 and the terminal pressure in the compressor and the admission 
 pressure into the engine are 150 Ibs. This system is used in 
 some pumping plants. 
 
 REFERENCES. 
 
 Efficiency Tests of Compressed-air Machinery, W. C. Popplewell, Mech. 
 Engr., Jan. 10, 1903; Transmission of Power by Compressed Air, W. C. 
 Popplewell, Mech. Engr., March 5, 1903, April 18, 1903; Compressed Air, 
 S. of Quart., April 1897, 196; Transmission of Compressed Air for Power, 
 L. C. Bayles, Compressed Air, Nov. 1903; Hints Concerning Air Compressors, 
 H. H. Kelley, Engr. U. S.A ., Mar. 1900; and Compressed Air, J. H. Richards. 
 
 Coal-cutting Machinery, E. W. Parker, A. I. M. E., Vol. XXIX, 405; 
 The Lee Long-wall Mining Machine, H. P. Bain, A. I. M. E., Vol. XXIX, 
 474; Stoping with Machine Drills, B. L. Thane, A. I. M. E., Vol. XXIX, 
 770; A New Type of Air Compressor, W. H. Booth, M. & M., Nov. 1903.
 
 CHAPTER XI. 
 
 PUMPING. 
 
 The Water Seepage into Mines. Water gains entrance into 
 mines by many and devious ways. Into some workings it 
 flows incessantly from some watery stratum, in others the seepage 
 is intermittent. . The subterranean current is easily excluded 
 from the mine by the use of a cement lining, or an iron or steel 
 tubing to the shaft, but the seepage accumulates and must be 
 pumped off, unless the workings possess a natural drainage or 
 an easy effluence by adit or tunnel for the upper ground. A 
 gutter at the side of the track, or under the tramway path, with 
 a slope of i in 500, readily carries off the water, and not uncom- 
 monly delivers it to a small wheel to drive a ventilating-fan. 
 Generally the seepage, following the hydrodynamic law, in- 
 creases with the depth of the opening, and a very liberal sump 
 is provided for its accumulation. Often one shaft and its 
 workings become, naturally, a sump for the entire district, and 
 drain all the neighboring properties above its level, and this 
 suggests a simple means of keeping one's mine dry. Otherwise, 
 as the amount of water to be encountered is uncertain, provision 
 must be made for the handling of a large volume, according to 
 the history of similar properties. In some coal-mines as much as 
 4000 gallons of water are raised per ton of coal, and in Colorado 
 40 tons of water per ton of ore. The Chief of the Bureau of 
 Mines reports that during 1902 1721 pumps delivered to the 
 surface water at the rate of 615,013 gallons per minute from the 
 collieries. The magnitude of such work demands the employ- 
 ment of powerful machinery, and often on a plan too elaborate 
 for the means of the average operator. 
 
 336
 
 PUMPING. 337 
 
 Methods of Unwatering Mines. Some districts were drained 
 by a cooperative scheme with extremely beneficial results. A 
 long tunnel penetrating the country at a level much below the 
 lowest point of exploration drains considerable territory, dis- 
 pensing with the heavy individual plants and extending the ex- 
 ploration and the productiveness of the mines. Numerous ex- 
 amples of tunnels miles in length may be quoted, some even 
 carrying so much water as to become canals for transportation. 
 
 Upon cutting a wet cross-course to the vein, it is a common 
 practice to plaster it up; or, in encountering old workings, to 
 build a brick or stone bulkhead, arched convex towards the 
 water (Fig. 257). To provide means for the escape of the accu- 
 mulated water, which might otherwise do injury, a cast-iron pipe 
 is built into the dam near its top, and another near the bottom. 
 Either or both may be plugged as required. Similarly, in 
 approaching abandoned works, it is required by law in some 
 States that a bore-hole be kept 30 to 50 feet in advance of the 
 drift, with flank-holes on each side, to guard against dangers 
 from the sudden breaking into the reservoir. 
 
 Under certain conditions, in stratified regions, a hole is drilled 
 from the sump down to some permeable stratum, into which 
 the water is discharged. 
 
 The water entering a mine at various levels, where economy 
 rather than simplicity is the object, should be led to pumps at 
 the levels where it issues, and not be permitted to find its way 
 to the bottom, to be raised the entire height to the surface. 
 
 When the surroundings are such that a tunnel may not be 
 used for the unwatering of the mine, pumping arrangements 
 are indispensable. The earlier forms were crude, the engine 
 being of recent date. Surface waterfalls were employed to 
 operate wheels, which raised bucketfuls from below; or the 
 surface water was arranged to compress air in a reservoir at the 
 surface, from which pipes to the sump conveyed the compressed 
 air, the elastic force of which, in turn, forced the water up to 
 the surface through another pipe. This is a wasteful system 
 and intermittent, but doubtless was cheaper than any other means 
 then available.
 
 338 
 
 MANUAL OF MINING. 
 
 The Hydraulic Ram. At the Comstock mines a special 
 hydraulic ram is used, by which 1800 gallons are pumped from 
 the 26oo-foot level to the Sutro tunnel at 1600 feet. The air- 
 pressure in the accumulator is 960 Ibs. 
 per square inch, and the pipes at the 
 bottom sustain a pressure of 2000 Ibs. 
 
 The efficiency of the ram diminishes 
 with the ratio between the quantity of 
 water raised and that used. With a fall 
 of i and a lift of 4, the efficiency is 86 
 per cent; if the lift is ten times the fall, 
 it is 53 per cent; with i to 20, it is 17 
 per cent; and with i to 26, it is o. 
 
 Hoisting Water. Small volumes of 
 water are handled by buckets, obtainable 
 of any size, and with a capacity up to 
 200 gallons (Fig. 130). At the bottom is 
 an inlet-valve by which the tub is quickly 
 filled as it sinks into the sump; it is then 
 hauled up, its valve closes, and at the 
 surface it is discharged by being brought 
 down on a pin which again opens the 
 valve. In some mines the water-bucket 
 is attached underneath the cage, and travels continually with 
 it. Bailing-fanks holding 500 to 1000 gallons, balanced in pairs 
 in one compartment, are hoisted by a special drum. Slopes are 
 equipped with similarly valved skips, the emptying being done 
 from the mouth, as with ore. 
 
 Bailing-tanks should always be in readiness for immediate 
 service at every mine operated through shafts or inclines, to re- 
 lieve the pump on emergency. Bailing is not economical, be- 
 cause of the great dead weight of tank and rope compared with 
 the weight of water hoisted. They are adapted to a variable 
 inflow and give an equal, though low, efficiency with all volumes 
 of water. 
 
 Mine-pumps. The design of mine-pump plants requires the 
 greatest possible reliability against breakdowns, the greatest 
 
 FIG. 130. A Water-bucket.
 
 PUMPING. 339 
 
 possible facility for making repairs, and, if practicable, the high- 
 est mechanical efficiency at the normal rate of flow. 
 
 The pumps used in mining work include reciprocating pumps 
 driven by rods ; steam-, air-, or water-engines ; rotary pumps oper- 
 ated by electricity, water- or steam-turbines, oil-engines, or belts; 
 pulsometers, injectors, etc., employing steam; and siphons. 
 
 The reciprocating pumps include the plunger, piston, and 
 bucket. Of these the oldest is the lift-bucket, which, however, 
 affords no means of repairing the barrels and packings, which 
 wear rapidly with sandy waters. The piston-pumps are suit- 
 able for low pressures, but cannot be repaired as readily as the 
 plunger-pumps, whose packing and stuffing-boxes are outside 
 and can be replaced without stopping the pump. They may be 
 driven by rods from a motor at the surface, or by a motor directly 
 connected, receiving power through wires or pipes. The rod- 
 driven pumps have many disadvantages; nevertheless, even 
 with the competition from steam and air-driven motors, they are 
 much used. 
 
 Belts or hemp rope may also be used to drive pumps on a 
 separate foundation. They are called power-driven pumps, 
 having the additional advantage that they may be located remote 
 from the power source. Of these the rotary pumps and centrif- 
 ugals have a capacity which can be economically varied within 
 very narrow limits. 
 
 Injectors, pulsometers, ejectors, etc., are to be regarded as 
 temporary substitutes only during the repairs to the main plant. 
 
 Mine Rod-pumps. The rod-driven pumps are of two types. 
 One, the lift-pump, is a single line of rods, which reciprocates in 
 a straight line of pipe. The lift-pump is a direct-driven pump, 
 with vertical single-acting engine by which the water and the rod 
 are lifted on the up- stroke, while the weight of the latter carries 
 it down. The other, the Cornish system, is an extended rod 
 operating outside of the column-pipe upon a number of plunger- 
 pumps along the line, with a lift section at the bottom. In the 
 Cornish pump the water is forced up on the down-stroke by the 
 weight of the rods. The engine at the surface raises the rods
 
 34 MANUAL OF MINING. 
 
 and plungers besides lifting the water in the suction-lift. The 
 Cornish system employs a rotary engine, making it possible to 
 utilize the expansive force of the steam. 
 
 Pump-pipes. The pipe through which the water flows is 
 variously termed as a stand-pipe, column-pipe, and lift. It is 
 of a diameter commonly 10 inches, often as much as 20, and ex- 
 tends from bottom to top. It is of cast iron, 
 wrought-iron lap-weld steel or spiral riveted, 
 or welded, in standard lengths of 5 to 20 
 feet. The cast-iron pipe, having a smooth 
 interior and uniform diameter throughout, is 
 preferable and more convenient than tha 
 riveted pipe (Fig. 131) or the lap-weld iron 
 (Fig. 132); but as it represents too much 
 dead-weight for the strength, its days of 
 utility are nearing an end. The ideal pipe 
 is of steel, which gives the lightest, strongest, 
 and most durable tubing; this may be had 
 in four grades, light to extra heavy. It is 
 made of sheet steel spirally laid, riveted 
 at the overlapping joints or cold hammer- 
 welded. The pipes are united by bolting 
 together at the flanges, which are riveted, 
 screwed, or locked on the pipe (Fig. 132); or, 
 preferably, they are coupled on the hub-and- 
 spigot plan of sleeve (Fig. 133). This is a 
 double socket, into which the pipe is slipped, 
 "oakumed," and leaded from each side, as 
 shown. For this joint the pipes have expanded 
 ends. 
 
 A water-tight joint is secured by placing 
 rubber, leather, lead, or, best of all, corrugated 
 copper gaskets between the flanges, which are 
 
 FIG. 131. Spiral Pipe. . rr , , ,-i i > 
 
 then bolted together while lowenng. Spence s 
 metal, used as a calker, offers an excellent joint, is cheaper than 
 lead, and ought to be better known.
 
 PUMPING. 341 
 
 The pipes last from fifteen to twenty years unless the water is 
 corrosive, in which case gun-metal is used. If the water is very 
 
 FlG. 132. Pipe-joint. 
 
 bad, wooden pipes are made by hollowing the trees, fitting the 
 joints, tarring them, and strengthening by wrought-iron bands 
 at every 3 to 6 feet. In many mines recourse has been had to 
 
 FIG. 133. Pipe-join.-,. 
 
 these as the only stand-pipe that will last over six weeks. 
 
 In calculating ^ the flow of water through pipes, the effect of 
 entrance head, the friction factor, the effect of bends, elbows, 
 tees, valves, gates, etc., as well as the condition of pipe-joints 
 and the internal surface of the pipe, should enter into the calcu- 
 lations; otherwise the results obtained are mere approxima- 
 tions. 
 
 Care should be taken to give independent support to the 
 water-pipe lines to prevent motion, to relieve the lower sections 
 of the pressure from above, and to avoid the evil consequences 
 of vibration. They are always stayed laterally to keep them in 
 line. In no case should the supports be rigidly connected, pre- 
 venting expansion or contraction. All bends and elbows should 
 be supported, as the curve of the pipe receives excessive pressure 
 and is also of weak resistance. All the bands are made of cast-
 
 342 MANUAL OF MINING. 
 
 ings, and the long bands may be of riveted pipe. The acu'e- 
 angular turns are made by inserting thick wedge-gaskets. 
 
 The size of the pipe should be as large as possible, that the 
 velocity of flow may be maintained with little resistance. The 
 suction-pipe should be short and more liberal in diameter than 
 the delivery. The velocity of water in the pipe differs little from 
 the velocity of the plunger or piston. Water-hammer and seri- 
 ous consequences would be the result, if water were allowed to flow 
 intermittently through pipes at more than 300 feet per minute. 
 
 Owing to the corrosive action of most waters and their solid 
 contents, the pipes are made as thick as consistent with other 
 conditions in the problem. The varying pressures, the acidity 
 of water, the water-hammer, and the weight of the column which 
 is supported, are elements which increase the thickness of the 
 pipe. 
 
 Valves on Pump-pipes. Air-chambers are supplied at con- 
 venient points of long hydraulic pipe, if no provision is made for 
 the insertion of relief-valves or pressure-regulators. The air- 
 chambers should be made air-tight and coated with heavy asphalt 
 for heavy pressures. Relief- valves are spring-loaded or weighted 
 pistons, or valves set to open at a given pressure. The former 
 are more sensitive than the latter. Ample lift should be given 
 to allow of a prompt discharge of air or of water to relieve the 
 excess pressure promptly. 
 
 Automatic air-valves on hydraulic pipe lines are sometimes 
 used, constructed so that they will close by the combined effect 
 of buoyancy and the velocity head of the water. Some air- 
 valves are provided to admit air and prevent a collapse of the 
 pipe from atmospheric pressure when the pipe is emptied of 
 water. They do not operate, however, to let air into the pipe 
 until the pressure falls very low. The pipes should be located 
 in the up-cast shaft compartments, the steam and reheated air- 
 pipes being protected also against radiation. Steam-pipes and 
 air-pipes are provided with traps at low points for drains. 
 
 Check-valves must not be neglected where the conditions are 
 such that a stoppage in the pump may cause a reverse flow of
 
 PUMPING. 343 
 
 the water with consequent danger to machinery. All valves and 
 gates and water-pipes should close tightly and without shock. 
 
 Pump-valves. The valves for pumps are hinged, called 
 clack-valves, straight lift-valves, rising vertically from their seats, 
 and flexible valves, which alter their form on opening. Direct- 
 driven pumps employ straight lift-valves of rubber or vulcanite 
 resting on a grating. The requirements of a good valve are that 
 it should close promptly and perfectly on its seat, open easily, 
 -and remain with a minimum of pressure. It should not present 
 much resistance to the flow or divert the current from a straight 
 line. It should be simple and accessible. 
 
 If the waters carry material in suspension, the whole valve 
 should be made of some elastic material in order that the par- 
 ticles lodging on the seat shall not spoil the band. Leather can- 
 not be used if the water is acid. Rubber compositions are em- 
 ployed for hot water. Hinged valves are more liable to leak 
 than straight lift-valves. The use of wooden seats for metal- 
 faced valves is objectionable because they leak, although they 
 are more desirable than rubber or leather. For acid waters the 
 seats are usually of brass screwed into position, and thus easily 
 removed when too far gone from corrosion. The straight- line 
 lift-valve (Fig 141) of rubber composition rising from a grating 
 and held in brass cages is used for pressures up to 500 Ibs. to the 
 square inch, but above that the discs are of metal with leather 
 facing. Flexible valves are generally of rubber, suitable only 
 for moderate lifts, and round in form, or rectangular with all of 
 the corners trimmed. 
 
 The area of the valves is large to allow of a free flow of the 
 "water, but increasing the area of the valve increases the pressure 
 upon its surface and requires a heavy, thick valve. Hence it is 
 more desirable to use a large number of small valves than to 
 employ few of large area. On the other hand, it is desirable to 
 restrict the lift of the valves as much as possible and to reduce 
 the velocity and consequent resistance of the flow past the valve 
 to a minimum. Enlargement of the area increases the risk of 
 leakage and also the percentage of slip. Again, increasing the
 
 344 MANUAL OF MINING. 
 
 area of the valve increases the width of the bearing and its seat. 
 This makes the difference between the lower area of the valve 
 subject to pump pressure and the upper area of the valve sub- 
 ject to the resistance in the pipe so great that an excessive over- 
 pressure will be required to raise it. 
 
 The Working-barrel. At the lower end of the stand-pipe a 
 12-foot length of cast iron constitutes the working-barrel, in 
 which oscillates a piston carrying an upward-opening valve, 
 similar to that at the lower end of the barrel (Fig. 136). For 
 acidulous waters the barrel is bushed with gun-metal. It 
 should be thick, to admit of being bored out several times, as it 
 is rapidly cut away by the gritty waters during sinking. 
 
 The working-barrel can never be more than 28 feet from the 
 sump-level; in mountainous districts still less; at 5600 feet alti- 
 tude, 23 feet; and at 10,000 feet, 18 feet. Usually the working- 
 barrel and suction-pipe are suspended by chains from two stulls 
 resting in the cribbing, and the stand-pipe supported at intervals 
 by stout reachers. 
 
 The Suction Length. Below the barrel is a length of pipe 
 or flexible hose dipping into the sump and receiving the water 
 through a perforated strainer. During sinking this suction-pipe 
 must follow the lowering of the sump. During the blast it is 
 raised for each shot or boarded over. 
 
 The strainer prevents sand and gravel entering the cylinder. 
 The suction-lift must be as short as possible and as large as ad- 
 missible. As the water-level lowers during the sinking opera- 
 tions in a mine, a stationary pump should be provided with a 
 wire-wound flexible hose below its suction-lift, dipping into the 
 sump, or a telescopic joint allowing for a supply of 10 feet. The 
 hose is preferable because it can be adjusted to the bottom of 
 the shaft. Owing to injury and hard usage it is usually covered 
 with canvas and wire- wound. 
 
 When the water-level has receded below the allowable suc- 
 tion height of the pump a length of pipe must be added above 
 the pump which is then lowered to a degree determined by 
 fixed conditions. Under such circumstances the working-'oump
 
 . PUMPING. 345 
 
 is suspended by chains from two stulls resting in the cribbing, 
 the stand-pipe above the pump being independently supported 
 at intervals. 
 
 The Lift-pump. The piston, or "bucket," is attached by 
 an iron fork (Fig. 134) to a wooden rod, 4 or 5 inches square, ex- 
 
 FIG. 134. Single-acting Steam Lift-pump. 
 
 tending up through the pipe to the surface, where it is connected 
 either with one end of a walking-beam or to the piston of a single- 
 acting engine. As it receives a tensile strain, the joints are scarfed 
 and strapped, or, if the ends are flushed, two continuous lines of 
 strap-iron, breaking joints, are bolted together through the rod. 
 The latter plan reduces the breakage and the number of stop- 
 pages for repairs. A 4- inch rod is large enough for a 1 2-inch
 
 346 MANUAL OF MINING. 
 
 pipe, and a 5-inchrod, properly spliced and strapped, for a 1 3-inch 
 to i6-inch deliver}'. The size of the straps is easily calculated. 
 The area of each one, a, is o.ooooj$d 2 D. A 2oo-foot pump-rod 
 requires two straps 4X or 3X{ for a lo-inch pipe. 
 
 On the down- stroke the rod falls through the column of 
 water, while the valve in its piston opens and the clack of the 
 wcrhing-barrel closes. Returning, the valve's action is reversed, 
 water rises from the sump into the working-barrel, and all that 
 above the piston is lifted a distance equal to the stroke, and a 
 column of water simultaneously discharged at the surface. 
 
 At the surface the column-pipe terminates in an elbow dis- 
 charge or in a laundry box and trough, the pump-rod continuing 
 up to the framing. The mechanism by which the motion is 
 communicated to it is simple. A stout frame, with two samson 
 posts, supports a walking-beam receiving its oscillatory motion 
 
 FIG. 135. Connection for Lift-pump Rod. 
 
 from a pitman actuated by a crank-arm adjustable to a i-, 2-, or 
 3-foot radius, giving strokes of double this length, at the opposite 
 end, to the pump-rod, which requires little force besides its own 
 weight. The arm is on a shaft turned from the engine by cog, 
 geared i to 6 or 7, giving 12 to 20 strokes per minute to the rod. 
 The ironwork of this frame, inclusive of cogs, pulley, and cast- 
 ings, will cost about $250; the woodwork, including a 24-ft.XT5 
 inches square walking-beam, about $125. 
 
 Lift-pump Plunger. The valves are made of several thick- 
 nesses of oak-tanned leather cut into discs, tacked together, and 
 slipping easily on a grid at the top of a cast-iron cellular ring- 
 bucket. A perforated cast-iron guard on the grid limits the 
 rising of the valve as the water passes through the bucket. The 
 cellular-ring bucket casting is all there is of the piston, which 
 Slides freely in the barrel, and has no other packing than 'that 
 offered by the leather discs forming the valve and which are cut 
 larger than the cylinder. The rapid movement, the wear, par-
 
 PUMPING. 
 
 347 
 
 ticularly during sinking, and the heavy 
 pressure upon these valves consume 
 a set every two weeks, and, though 
 substitutes have been suggested, such 
 as flexible brass or gutta-percha plates, 
 they are not more durable, nor have 
 the brass balls or conical poppets had 
 any marked success. The valves are 
 repaired or replaced by removing a 
 bolted door-plate in the barrel opposite 
 them (Fig. 136). 
 
 The Direct Lift-pump. The frame- 
 work just described is crude. A simpler 
 connection is a vertical steam- cylinder 
 placed over the shaft wi h its piston bolt- 
 ed to a fork on the rod This dispenses 
 with framing (Fig. 134) beyond that nec- 
 essary for foundation. The piston may 
 be single-acting, though frequently the 
 steam is admitted on both sides These 
 cylinders are easily set, their cost is low, 
 and their maintenance is small. Cylin- 
 ders of i2"X3o" are operated at a rate 
 of 24 double strokes per minute. The 
 Bull pump is an example of this type, 
 well known in collieries. It is single- 
 acting, and is as large as 60 inches in 
 diameter, making 6 to 8 strokes of 10 
 feet each. 
 
 It is not certain that this form of 
 pump gives a higher duty than the drive- 
 rod pump. Except for increasing the 
 speed there is no occasion for using steam 
 on the down- stroke. An economic degree 
 of expansion on the up-stroke is not pos- 
 sible. The main objection to its more 
 general adoption is the large area of 
 
 36. Drive-rod Plunger.
 
 348 MANUAL OF MINING. 
 
 shaft-mouth it covers. Besides, to lengthen or repair the rod or 
 column-pipe the cylinder must be displaced, or the additions 
 are made below; either is slow. This pump cannot be used in 
 slopes; the irregular wear of the cylinder on one side cannot be 
 compensated for, nor the friction of the rod in the pipe coun- 
 teracted. 
 
 The Capacity of Lift-pumps. If d be the internal diameter 
 of the pipe in inches, and L the stroke in feet, the discharge 
 in gallons with each upstroke is 0.0408 d 2 L; and the work 
 done per minute, in foot-pounds, exclus've of resistance in 
 the cog-gear and the mechanism for transmitting the power, 
 is 0.3427 2d 2 LND, where N is the number of double strokes per 
 minute, and D the height of the water-column in feet. The 
 direct-connection lift-pump wastes less power in friction and 
 has an efficiency of about 85 per cent. Its least working steam- 
 pressure is represented by this equation : 
 
 k 2 p=o.34$d 2 D and ^-0.551(0.7854^ a)Z>. 
 
 With a moderate steam pressure the ratio between k and d 
 becomes large if the shaft is deep, and the area of the wooden 
 rod, a, must be large enough to support the load. When the 
 depth of the shaft has reached 250 feet the lift-pump is no longer 
 practicable, and must be altered to a single discharge-force, or 
 replaced by the more economic continuous-flow steam-pump. 
 
 Force-pumps Rods. The pump-rod of the force-pump, unlike 
 that of the lift -pump, works outside, not inside, of the stand-pipe. 
 Its lower end is bolted to one end of a cast-iron H chamber, the 
 other stem of which carries a long working-barrel, into 'which 
 plays a solid plunger-piston (Fig. 136), instead of the bucket- 
 valve of the lift-pump. Below the stem carrying the stand-pipe 
 is the suction-pipe, and in it is an upward-lift clack-valve. Then 
 during the up-stroke the lower clack opens, while water rises 
 through it into the working-barrel; as the plunger falls the water 
 is driven through the upper valve against the column of water 
 in the stand-pipe. As sinking progresses, suction lengths are
 
 PUMPING. 349> 
 
 added at the bottom until the sump is lowered beyond the suc- 
 tion distance, then the pump is lowered, while additional lengths 
 of pipe are attached between it and the H piece at the discharge- 
 station above, until a lift of 300 feet has been reached; then the 1 
 suction-barrel is removed, to do service similarly for the lower 
 lifts, and is replaced by another H piece. 
 
 The Cornish Pump. The Cornish system is a develop- 
 ment of the rod-driven pump by which the lifting section is added 
 to a series of force-pumps in a continuous line, the water being 
 driven by stages to the surface (Fig. 138). The engine raises 
 the rod, and water is sucked up into the working-barrel at the- 
 bottom, while that above the bucket is lifted to the first station) 
 above. Here the plunger is drawing water up through the lower 
 clack into its barrel, all the other plungers doing the same. At 
 the end of the stroke (6 to 10 feet) occurs a slight halt, incidental 
 to the change of direction. The rod falls by reason of its own 
 weight, and each plunger closes the lower clack in its H piece, 
 opens the upper one, and forces the water out of its working- 
 barrel, driving at the same time the entire lift-column an amount 
 equal to that of the stroke. At the surface a volume of water is 
 discharged on the down-stroke. 
 
 The Cornish Rod. The pump-rod extends down the shaft, 
 and terminates at the bottom in the piston of the bucket-lift. 
 At intervals are offsets, to which are bolted the rods carrying 
 external plungers, not pistons, to move the water, and are there- 
 fore more easily packed and admit of pumping against higher 
 heads, and remaining tight much longer, than piston-pumps, 
 being much less subject to wear. They are made of cast iron, 
 though brass is a better material, having also less friction. The 
 stuffing-boxes for packing the plunger are separate from and 
 built to top of the pump-level. The usual packing is of square 
 braids of hemp, flax, or cotton soaked in tallow or a mixture of 
 tallow with beeswax. Flax is the cheapest, but hemp is more 
 durable. Both of them offer greater friction than leather as a 
 packing. The latter is used in all pistons and bucket-pumps. 
 
 Each plunger reciprocates within its working-barrel, whiclr
 
 -.35 MANUAL OF MINING. 
 
 is one leg of an H piece (Fig. 137). At every station is located 
 ;an H piece having at the top and the bottom a hinged valve 
 opening upward. The working-barrel is about 15 feet long. 
 
 Excepting the short suction-lift length at the bottom, the 
 column-pipe is in a continuous line, broLen only for the intro- 
 duction of H pieces, or tanks, at ;he stations. 
 
 The lifts are rarely over 300 feet or less than 150 feet in 
 '.height. As the stations require heavy timbering, especially if 
 .tanks are used, their dimensions must be large and their sup- 
 ports solid and independent of the shaft- timbers. They are not 
 any more numerous than the circumstances demand. The 
 greater the number the greater the speed of pump and the smaller 
 the diameter of the pipe will be for a given first cost, exclusive of 
 that of the tank-chambers. 
 
 The pump-iod is composed of sections joined at their ends 
 by iron straps to bear the continual reversal of direction of stress. 
 Their length is as great as convenient to handle, and their sec- 
 tion as large as necessary to resist the tensile and compressive 
 forces to which they are subject. On the up stroke all sections 
 are in tension. On the down-stroke the net stress is the resultant 
 of the compression from forcing the water and the tension of 
 the pendent weight below the section, modified by the inertia of 
 the attachments and counterweights The counterbalances 
 reduce the tension and increase the compression by an equal 
 amount. 
 
 The aggregate weight of the rod is greater than that of the 
 water to be moved. Its cross-section is smallest at the bottom, 
 increasing to the top with the increase of tension upon a section 
 from the pendent weight below. 
 
 To prevent accident from breakage or buckling of the rod, 
 stout stulls are laid across the shaft at intervals close to the rod 
 to catch corresponding "wings" of heavy timber clamped by iron 
 collars on the rod. The stulls are called ' guides," or " stays." 
 
 In a deep mine requiring many superposed sets cf pumps on 
 the same rod the stroke of the lowest length is shorter than that 
 of the topmost plunger. This may require an increasing area
 
 PUMPING 35 r 
 
 of the plungers from bottom to top. Usually, however, they 
 are the same. It not infrequently happens that there is a differ- 
 
 FlG. 137. An H-piece Cylinder of a Cornish Pump. 
 
 ent displacement of the pumps, which may cause pounding or 
 other troubles. A float provided at each station to tap the pipe
 
 35 2 MANUAL OF MIXING. 
 
 into the station tank when it happens to be drained too low. 
 would maintain a supply of water and prevent overflow of one 
 tank and the draining of another. 
 
 Elastic bumpers are also needed to break the force of the blow 
 which might occur with a stroke greater than the intended limit 
 through a variation of the steam pressure or neglect in regulation 
 of the water flow. As an instance of the size of a rod that of 
 4he Maira, 2300 feet deepwe find the first 780 feet down was 
 36"X32", tapering to 12" X 24"; at 864 feet it was 16 feet square; 
 it tapered to 14 inches square at 964 feet; thence it was 13 inches 
 and at the bottom 12 inches. In well- ventilated shafts wood is 
 the preferable material for rods, neither wrought-iron rods nor 
 wire rope having the requisite resilience. 
 
 In vertical shafts the rods fall freely by their own weight; 
 in slopes they rest on friction-rollers, placed about 30 feet apart. 
 When iron ropes are used instead of wooden rods, sheaves sup- 
 port them. Changes in the slope may be provided for by the 
 use of the rocking-arm. A chamber is cut in the shaft at the 
 angle, in which is firmly set a frame on which swings a V bob 
 "by a hinge-pin at the apex. To the two arms of the angle the 
 inclined rods are attached. While this arrangement is not de- 
 sirable, because of the expense and the loss of power, still it is 
 the best to be had when slopes are sunk on contorted veins. The 
 application of the Cornish system to inclines is attended with 
 many drawbacks. 
 
 In mines utilizing the pump-rod for a man-engine additional 
 counterbalance weights are connected at intervals down the shaft 
 {Fig. 138). Sometimes two lines of rods are used in a shaft, 
 working two pumps from the same bob, in which case no counter- 
 balance is needed. 
 
 The ironwork of the rod should be protected against rust, 
 particularly the joints and the inside of the hollow iron pump-rods. 
 Pickling in acid will remove the rust, after which a coating with 
 warm oil and red lead is recommended. 
 
 The Balance-bob. At the surface the rod is directly con- 
 nected to a pin at one apex of a king-post balance-bob. If it be 
 of iron instead of wood, a link is used for a flexible connection.
 
 PUMPING. 353 
 
 The horizontal beam of the bob is about 25 feet long, with a sad- 
 dle and axle underneath near the centre. From the upper end of 
 the king-post, which is 8 feet high, a connecting-rod leads to the 
 engine. Besides the braces on each side down to the beam there 
 are two tie-rods, taking with the braces the tensile and compressive 
 stresses. All the members of the frame are of wood or iron, in 
 iron shoe-castings at the ends. The frame stands vertically in a 
 pit dug alongside of the shaft, 8 to 10 feet deep (Figs. 35 and 
 138). The rocking motion is communicated to the bob by a 
 connecting-rod, or " pitman," operated from a wheel geared to the 
 fly-wheel shaft of the engine, the work of which, during the up- 
 and down-strokes, is somewhat equalized by the bob and its 
 counterpoise. The crank-pin can be set to different radii to 
 alter the pump-stroke, increasing the leverage of the engine 
 for the greater depth of pumping. The third apex is occupied 
 by a box full of iron and boulders, to counterbalance the ex- 
 cessive weight of the rod; for it will be found that the weight 
 of the long column-rod of a strength requisite to force the 
 water up is much greater than that of the water pumped. A 
 certain pump, raising 440 gallons 1690 feet by six lifts in a 
 22-inch pipe, has a balancing weight of 33 tons on the bob. 
 
 The Cornish Engine. When once placed and its speed regu- 
 lated, the Cornish pump gives little trouble. It is the most 
 reliable and also the most expensive pump in use. It has numer- 
 ous advocates as against the steam-pumps; but in transplanting 
 the system to America its main redeeming feature the cataract- 
 engine was discarded, while persistently clinging to the worst 
 the cumbrous bob and rod. When the vertical direct-acting 
 engine was introduced it was thought to be a great improvement, 
 because of the suppression of the heavy bob; but it was soon 
 discovered to be a mistake, and the beam was quickly reestab- 
 lished, with a Corliss horizontal engine as the motor. An engine, 
 boiler, and fittings complete, with three 15 -inch plungers and 
 one 1 6-inch lift-pipe, etc., etc., for a 600- foot shaft, weighed 
 650,000 Ibs., had a capacity of 800 gallons, and cost in place 
 $54,000.
 
 MANUAL OF MINING. 
 
 
 FlG. 138. The Cornish Pump. 
 
 All the foundations about the 
 shaft should be carefully laid, for 
 the condensed steam and pump- 
 water soon makes the ground 
 yielding. In stable ground 
 heavy beams buried in the stone 
 will suffice ; in ground at all soft 
 secure foundation can only be ob- 
 tained by concreting a consider- 
 able area for 6 feet deep, and 
 erecting a rigid timber or brick- 
 work base several feet high (Fig. 
 
 138)- 
 
 The Speed of the Cornish 
 Pump. A variable regulator 
 enables the engine somewhat to 
 change its speed commensurate 
 with the water to be discharged. 
 The rod speed is about 60 feet 
 per minute, giving six to ten 
 strokes of 10 or 6 feet each. To 
 increase the speed it is preferable 
 to increase the length of stroke 
 rather than the number. The 
 maximum speed cannot exceed 
 that of the inflow, or the plunger 
 would acquire on its down- 
 stroke an acceleration before 
 the water had fully entered the
 
 PUMPING. 355: 
 
 cylinder. A shock ensues which is detrimental to the rod and 
 valves. If continued, a vibration or churning occurs with the 
 continual pound of descending plunger upon the rising water 
 column. The speed of the pump should be reduced or the engine 
 stopped entirely for a time. If after starting it is again set up 
 the fault may be with the valves giving an obstrucled flow. 
 
 The valves should afford unobstructed passage to the water 
 in one direction, and close perfectly in the other Iwo antagonis- 
 tic conditions which can be attained only partially. The strain 
 on the valves is enormous, and if tried too hard they become 
 weak, do not work properly at either stroke, and lose water by 
 their "slip." If the valves are found to be free and the vibra- 
 tion still continues, the joints of the rod are loose or the bob 
 requires resetting. Increasing the mass of the rod is sometimes, 
 though not always, a remedy. This vibration must be par- 
 ticularly guarded against if the rod is also to be utilized for a 
 man-engine. For convenience in repairing, the pumping com- 
 partment should be provided with a ladder-way, with plats and 
 chambers. 
 
 Cushier's Double-acting Drive-rod System. The use of a 
 double-acting pump, retaining therewith the advantages of the 
 Cornish, would save space in the shaft, the pipes for a continuous 
 discharge occupying less than one fourth the area of a single 
 discharge-pipe and its rod. 
 
 Cushier's system of pumps for deep mines consists in having 
 sets of two pumps, each working in concert, one above the other, 
 the suction- and discharge-pipes being common to both pumps. 
 
 The pumps are placed at intervals of about 200 feet in the 
 shaft, the power being transmitted directly through the centre 
 of the plungers. The connection with each other and to the 
 motive power is effected by means of a steel-wire cable encased 
 in wood, preserving it from external wear as well as from rust. 
 This cable is fastened to, and its length regulated by, shackle- 
 bolts. 
 
 The plunger of the lower pump in a set is double in area 
 that of the upper one, so that in working on the upper stroke
 
 356 MANUAL OF MINING. 
 
 one half the water raised fills the chamber of the upper pump, 
 the other half being forced out through the discharge-pipe on 
 the down- stroke ; the upper pump-plunger forces out, in its 
 turn, the water in the chamber, thereby causing a continuous 
 delivery. 
 
 This form of pump can be worked at any angle, to any depth, 
 and is almost perfectly balanced. The last-named advantage 
 enables it to be connected with, and worked by, a direct-acting 
 steam-cylinder, and thus does away with the complicated gear 
 and bob of the Cornish. 
 
 Sinking-pumps. During operations of shaft-sinking a double- 
 acting steam-pump may be suspended vertically from some support 
 by a chain to a bale, the suction-inlet being at the bottom pf 
 the pump. By this suspension it can be accommodated to a 
 varying water-level. A centre-packed plunger is directly con- 
 nected with the steam-piston, from which it receives its power, 
 as in the horizontal pump (see Figs. 140 and 144). The sinking- 
 pumps of the Cameron, Knowles, and Worthington companies are 
 of similar pattern. The valves are absolutely positive, and are 
 protected by a cast-iron shield serving as a yoke between the 
 steam and water ends, while those in the steam end are cush- 
 ioned to regulate the strokes. Hand-hole plates, with hinged 
 bolts, allow of easy repair of the valves and the shoes and dogs 
 (at the left of the pump, Fig. 139) of easy and simple means 
 of support. 
 
 Sometimes the sinking-pumps are attached to a sinking- 
 frame guided in the shaft. The frame and pump are raised as 
 occasion requires by chain-blocks, winches, or special hoist at 
 the surface. 
 
 The Reciprocating Pumps. A plunger driven by steam, air, 
 or water pressure in a cylinder directly connected with the water- 
 piston or plunger, the two being rigidly coupled and having a 
 common stroke (Fig. 141), constitutes a reciprocating pump. 
 It is always double-acting and may be single or duplex, and is 
 often compounded at the motor end, with, in some cases, a con- 
 densing connection.
 
 PUMPING. 
 
 i 3Q _A Sinking pump. FlG. 140. Section of a Sinking pump.
 
 358 
 
 MANUAL OF MINIX
 
 PUMPING. 
 
 359 
 
 The column-pipes are smaller in diameter than for the Cor- 
 nish system and the piston speed is greater, but, being of the non- 
 rotary type, no advantage can be taken of the expansion to any 
 great degree. The water end may have a piston (Fig. 142) or a 
 plunger (Fig. 141). The latter is used for the higher heads of 
 lift. 
 
 FlG. 142. Packing a Follower Piston. 
 
 The steam-pump dispenses with the cumbrous gear, bob, and 
 rod, having instead a small, well-lagged steam-pipe, conveying 
 the power down from a surface boiler. Its construction is sim- 
 ilar to that of the air-compressor (Fig. 126), consisting of a steam- 
 cylinder in which a piston oscillates, moving its own steam- 
 valves by rockers without the aid of any rotary appliances; at 
 .the same time it reciprocates a solid plunger centrally in a water-
 
 MANUAL OF MINING. 
 
 FlG. 144. A Steam Sinking-pump and One of the Relay 
 Pumps of a Forcing System. 
 
 FIG. 143. A Suction-pipe Fitted for a Condenser,
 
 PUMPING. 361 
 
 cylinder, at each end of which is a set of double-beat valves of 
 appropriate construction, open only so long as the water is being 
 forced, and closed with the aid of springs. One valve is removed 
 in the figure. 
 
 Pump-valves. The inlet-valves are opened back of the 
 water-piston by the pressure of the atmosphere, which forces 
 water into the space thus provided. The water is expelled 
 through valves into the dischajge-pipe on the return-stroke. 
 At the other end of the water-cylinder similar conditions prevail, 
 the delivery- valves being raised against the action of the springs 
 which hold them to their seat (Fig. 141), and the exhaust-valves 
 being raised by atmospheric action. 
 
 The steam-valves in duplex pumps are like those in the steam 
 end of engines, but have no lap. They are provided, however, 
 with considerable allowance for lost motion, which prevents one 
 port from opening as soon as the other port has been closed, 
 at the end of the stroke of the piston. The valve-stem does not 
 engage the valves rigidly, and hence imparts no motion to the 
 valve until the piston operating it has nearly completed its 
 stroke. The amount of lost motion is about one third the width 
 of the steam-port. If the piston does not make the full stroke, 
 the valve-chest cover must be removed and the lost motion 
 increased. 
 
 Single-cylinder Pumps are often preferred to duplex pumps, 
 because they deliver more water in proportion to their size and 
 the amount of steam used. There is no interference in the 
 action of the piston by the tightness of the stuffing-box on the 
 other side, as in the duplex; hence there is no tendency to reduce 
 its piston speed. In the single-cylinder pump the piston never 
 starts on its return until it has travelled its full length of stroke. 
 The steam economy is due to a smaller radiating surface for a 
 given capacity and the diminished clearance. The valve-gears 
 of the various types are automatic and give little trouble, and 
 are not complex in construction. They usually consist of some 
 auxiliary valve, which opens and closes the exhaust and steam 
 admission for the main valve. The former is moved by the
 
 362 MANUAL OF MINING. 
 
 piston-rod, and the latter controls the movement of the piston. 
 The piston of the single-cylinder pump always makes a full stroke. 
 
 The Plunger-pump. Where great pressure or the gritty 
 nature of the water renders the use of the single piston unde- 
 sirable, the water- cylinder is divided in the centre, and a pair of 
 plungers, discharging alternately, work in the opposite ends, 
 and are connected with yokes and heavy outside rods to the steam 
 piston-rod (Fig. 145). This arrangement of external stuffing- 
 boxes permits of instant detection of leaks. Strictly speaking, 
 the combination is a pumping-engine ; but this term is custom- 
 arily applied only to the double- and triple-expansion engines 
 used for city supply. 
 
 The pump-cylinders are best made with brass linings. The 
 piston is packed with hydraulic packing (Fig. 142), hemp soaked 
 in Albany compound for cold water, or square-braided cotton 
 mixed with plumbage for hot water. Such packing is held in 
 place by the follower. Rings of square rubber or double cup- 
 leathers are also sometimes used. 
 
 The plunger is usually hollow and of such thickness that it 
 will be of the same weight as an equal volume of water. 
 
 An air-chamber is used on pumps to maintain an equal dis- 
 charge. It is placed at the highest point on the discharge and 
 permits the air to rise to form a cushion which maintains uni- 
 form- pressure upon the column of rising water. Its volume is 
 about twice the volume of the water- cylinder of duplex pumps. 
 
 A vacuum-chamber on the pump will keep the water in full 
 motion, and stop it gradually and easily. It may be placed at 
 the side or the end of the pump. Its action is practically the 
 reverse of that of the air-chamber and facilitates the changing of 
 continuous into intermittent motion. 
 
 The suction-hose is connected under the inlet-chamber, and 
 the discharge-pipe to the surface on top or at one side of the 
 outlet-chamber. The water-passages are short and very direct; 
 the valves should be large, move quickly, and close tightly, that 
 little loss be experienced; otherwise the effective and suction 
 powers are both reduced.
 
 PUMPING. 363 
 
 The principal difficulty is with the disposal of the exhaust. 
 If turned off into the sump, as in Fig. 143, the temperature of 
 the mine is raised, ventilation is injured, and the timbers ruined; 
 if carried to the surface, the condensation in the pipe gives trouble. 
 The best remedy is to use a condenser, which reduces the back 
 pressure and increases the efficiency. Jacketing the cylinders 
 materially contributes to economy of steam consumption. 
 
 The condensation of the steam can be carried into the suc- 
 tion-pipe, as is the universal practice, being arranged to enter 
 the pipe in a direction parallel to the flow of the water. This 
 enables it to act like an injector, and aid in accelerating the lift 
 of the water (Fig. 143). 
 
 The Comparative Merits of the Steam-pump. The direct- 
 acting steam-pump does not equal the Cornish cataract pumping- 
 engine in fuel economy. A fly-wheel is necessary in order to 
 secure the full benefits of a high degree of expansion, which, 
 as has been seen in Chapter V, is not feasible in one cylinder, be- 
 cause the resistance (the weight of water forced up) is constant 
 throughout every stroke. A fly-wheel would distribute the 
 stearn-power excess of the first part of the stroke to the latter. 
 By the use of the compound cylinders the saving of steam-power 
 may be fully 10 per cent. The ratio of expansion is rarely over 
 3. The Deane compound pump with externally packed pistons 
 is illustrated in Fig. 145. 
 
 Direct-acting pumps are generally made as light as possible, 
 but it is not practicable to employ the steam expansively. The 
 steam is admitted during the full stroke of the plunger, as 
 is revealed in the indicator card, Chapter V, showing the con- 
 stant pressure required for the uniform water resistance. The 
 pressure may be changed within limits of the boiler supply. 
 
 For general purposes these direct-acting force-pumps are in 
 universal use. Their chief feature is their comparatively high 
 efficiency at any speed, slow or fast; they are capable of quick 
 adjustment in speed and discharge, as emergency demands; 
 but they require close watching, especially where the water 
 is "quick," or they may be drowned. The Cornish pump does
 
 364 MANUAL OF MINING. 
 
 not admit of variations in its rate: during summer and winter 
 it is run only a few hours through the shift to empty the sump 
 which has been filling overnight. The small cost, great sim- 
 plicity, and ease of repairs give the steam-pump an important 
 advantage. A plant with boiler, pipe, and fittings, complete, 
 can be installed for less than one fifth that of a Cornish outfit. 
 One for 850 gallons per minute, 400 feet, cost $15,000 in 
 place. These pumps have timber foundations, in a large, well- 
 
 FIG. 145. Deane Compound Pump with Externally Packed Pistons.. 
 
 timbered excavation alongside of the shaft and near the sump 
 level, which must practically be invariable. They are useless 
 during sinking without a sinking-pump to deliver to their tanks. 
 In coal-mines, and where the machinery can be established for 
 a permanent bed, these pumps have no rival (especially if com- 
 pounded); whereas in vein-mining the pumping apparatus, and 
 indeed all of the machinery, is under process of continual exten- 
 sion. For this reason in metal-mines the choice must be between 
 a set of relays of direct-acting pumps at each 200 or 300 feet 
 with a sinking-pump at the bottom (Fig. 144), and the Cornish 
 pump with its several force-stations and its bottom-lift (Fig. 138). 
 The Capacity of a Pump. In determining the horse-power 
 and the amount of steam necessary to operate the pump running 
 at full capacity, the calculation will be made in the same 
 manner as for steam-engines, bearing in mind that the steam
 
 PUMPING. 365: 
 
 cannot be used expansively. The volume of steam consumed.' 
 per minute can readily be determined by the continued product 
 of the area of the steam-piston, the number of strokes made- 
 by all of the plungers in a minute, and the length of each stroke. 
 This product, reduced to pounds by reference to the steam-tables- 
 and divided by 34.5, will determine the boiler horse-power of 
 its boiler. Pumps usually have boilers independent of the other 
 steam-consumers. 
 
 Steam-pumps are very wasteful of steam, the simple direct- 
 acting type using at least 100 Ibs. per horse-power per hour; the- 
 duplex may consume as much as 300 Ibs. of steam; the com- 
 pound non-condensing may not use more than 70 Ibs., and the- 
 compound condensing but 40 Ibs. per horse-power hour. 
 
 Let W = weight of steam required by the pump per hour, arid' 
 w = " ' ' i cu. ft. steam at the pressure in the cylin- 
 der. 
 
 Then k, s, and AT" being the diameter, stroke, and number of 
 strokes per minute, W =o.o^4ik 2 sNw. 
 
 When it is found that a boiler pressure exceeding 100 Ibs^ 
 per square inch is necessary for the given lift of water, either 
 the pump or the boiler pressure must be increased, or the height 
 of the lift reduced. With a given steam-pump and boiler pres- 
 sure it would be necessary to employ several pumps in relays, 
 each pump delivering to the tank above and thence to the sur- 
 face. In Fig. 144 is shown the disposition and arrangement 
 of such a set. 
 
 The Duty of the Pump. The calculation of the steam and 
 fuel economy is easily made; the necessary elements are few 
 in number. The standard of comparison of the work of a pump 
 is its duty the number of foot-pounds of work actually per- 
 formed per bushel (80 Ibs.) or per 100 Ibs. of coal. The com- 
 bustion of one pound of anthracite gives sufficient heat, theoretic- 
 ally, to do 12 million foot-pounds of work. The ratio between 
 the work actually done and the power at the steam end measures 
 its efficiency, to the consideration of which, in and about mines, 
 insufficient attention has been given notwithstanding its pecuni-
 
 3^6 MANUAL OF MINING. 
 
 any importance. The duty of a small pump is from 7 to 15 
 jnillion foot-pounds per 100 Ibs. coal; a compound pump gives 
 from 15 to 30 millions; while the higher types of pumping-engines 
 furnish from 30 to 100 million dynamic units, corresponding to 
 Ihe consumption per hourly horse-power of 28 to 13, 13 to 6.6, 
 And 6.6 to 2 Ibs. coal respectively. The consumption of coal per 
 liourly horse-power equals 198 divided by the duty (in millions). 
 A recent report of a Wo:thington engine having a capacity of over 
 -looo gallons per minute, against an equivalent of 2000 feet head 
 of water, showed a duty of 184 ft. -Ibs. per thermal unit, or 158,- 
 DOO,OOO dynamic units per 100 Ibs. of coal. 
 
 To illustrate the influence of compounding and jacketing the 
 steam-cylinder, and of condensing the exhaust, upon the coal 
 bills, two examples will suffice. As has been stated, many of 
 the collieries pump 4000 gallons of water per ton of coal hoisted. 
 To raise this only 300 feet requires the consumption, theoreti- 
 cally, of 336 Ibs. anthracite for a daily output of 400 tons. If 
 the duty be 90 million foot-pounds, as in Cornwall, or 20 million, 
 as with our average duplex compound pumps, the aggregate 
 yearly consumption is 675 and 3005 tons respectively of anthra- 
 cite, or of 900 and 4000 tons of lignite. 
 
 But duty is not the sole feature of a piece of machinery: 
 the repairs and lubricant accounts and the durability of the 
 plant are not to be overlooked; for the indicator card is a less 
 valuable guide than are the coal, oil, and packing bills. More- 
 over, the inconveniences arising from the use of steam under- 
 ground and those of the occupation of a shaft compartment 
 fcy rods, catches, etc., must receive attention. The cost of pump- 
 ing per million foot-pounds is not far from 1.6 cents with the 
 direct-acting pumps, and 2.5 to 2.9 with high-pressure rotative 
 engines. 
 
 Pressure-regulators for Pumps. A steam-pump should be 
 provided with a speed-governor and a pressure-regulator. The 
 design is to place on the pump a diaphragm or a separate piston 
 connected by a spindle with the valve supplying steam to the 
 pump. When the water pressure becomes excessive it acts upon
 
 PUMPING. 
 
 a balanced diaphragm which opens the steam-pipe, and admits 
 a greater supply to the pump. 
 
 A check-valve should be placed in the pipe between the pump 
 and the delivery stand-pipe, to prevent the return of the water 
 in the event of the pump stopping operations. A relief-valve or* 
 the water-pipes will also prevent any injury from an excessive- 
 pressure or sudden shock. 
 
 The Speed of the Steam-pump. Pumps are usually started 
 by a process called priming, which consists in removing the air 
 from the barrel, filling the pump with water from a discharge- 
 pipe until all the suction space is charged. 
 
 The speed may be varied, but is not advisable beyond 100 
 feet per minute as the standard rate. The water-pipe is calcu- 
 lated for a flow of 200 feet per minute. The piston speed is 
 limited by that of the possible velocity of entry of the suction* 
 A speed in excess of this not only results in water-ram, but reduces 
 the capacity of the pump and increases the difficulty of a prompt 
 seating of the valves. Any delay in the closing of the valves 
 results in a slip of the water into the cylinder. The slip should 
 not exceed 3 per cent of the piston displacement. A large num- 
 ber of small valves is preferable to a few large ones, so far as 
 the amount of slip is concerned. The area of the suction-valves 
 should be 40 per cent of the water-piston area. Their diameter 
 should not be over 4 inches. 
 
 The Suction Height. The size of the suction-pipe should 
 be as large and as short as possible, that the pump may be able 
 at high speeds to obtain as much water as it can deliver with a 
 given speed of plunger. Its size will depend upon the length of 
 the pipe between the pump and the water level, as well as the 
 difference in elevation between the two. The former determines 
 the frictional resistance, and the latter the maximum velocity of 
 inlet entering the pump, due to the atmospheric pressure. 
 
 As the reciprocating piston cannot produce a vacuum lower 
 than 3 or 4 Ibs. absolute pressure, the atmosphere cannot lift 
 the water higher than 28 feet in the suction length. To obtain 
 a velocity of eatry the pump must be placed nearer to the water
 
 368 
 
 MANUAL OF MINING,
 
 PUMPING. 369 
 
 level than this suction height. As it is desirable to have a high 
 speed of transmission, the pump should be lowered to such a 
 position as to give a velocity of at least 25 feet per second. Hence 
 the maximum suction-lift of 14 feet should not be exceeded if 
 high speed is desired. 
 
 The height to which the water is raised is equal to the level 
 of the suction- lift below the pump plus the height to which it is 
 forced above the pump. This height multiplied by 0.433 * s 
 equal to the head acting on each square inch of the water- 
 piston. The area of the water-piston multiplied by this weight 
 determines the total resistance. The steam pressure required 
 in the pump per square inch of steam-piston is equal to the 
 resistance just determined divided by the area of the steam- 
 piston. This quantity multiplied by m, the efficiency of the 
 pump, determines the actual steam pressure necessary to over- 
 come the resistance of the column of water. Thus a 6-inch steam- 
 piston and a 4-inch water-piston having areas of 28.27 an d I2 -57 
 sq. in. respectively, with the water pressure due to 100 feet of 
 suction and force head, or 43.3 Ibs. per square inch, the net 
 steam pressure theoretically required will be 19.4 Ibs. per square 
 inch; allowing for a mechanical efficiency of .70, the steam 
 pressure then must be 28 Ibs. above that of exhaust. 
 
 Motor Fluids other than Steam. Compressed air, oil, an4 
 water are used also as water agents in the driving of the direct-i 
 connected pumps. , The oil-engine is a simple, self-contained, 
 and portable motor of great value. An oil-engine of high economy 
 is illustrated in Fig. 147. 
 
 Compressed air is used in the same manner as steam in the. 
 power cylinders, by utilizing its elastic property. A pump driven 
 by air cannot be drowned out as is a steam-pump, and does 
 not require as much attention as the latter. A high-class mod- 
 ern compressed-air plant will compare very favorably in com- 
 mercial efficiency with underground steam-pumps. In one 
 case 280 horse-power was consumed in raising 400 gallons per 
 minute against a head of 120 Ibs. per square inch, showing an 
 efficiency of 9 per cent.
 
 37 
 
 MANUAL OF MINING. 
 
 There are two other classes .of air-pumps: the displacement 
 pump, in which the water is displaced or expelled by entering 
 
 FIG. 147. The Diesel Motor. 
 
 compressed air; and the air-lift pump, in which the water is 
 raised by pressure and the expansive force of the air.
 
 PUMPING. 371 
 
 The Displacement Air-pump. Two cylinders are provided 
 side by side with an air-valve immediately above them. Com- 
 pressed air is automatically admitted, first to one cylinder, then 
 to the other, and the air correspondingly released. The pressure 
 of the air on the surface of the water in the cylinder forces the 
 latter through its discharge-valve to a height depending on the 
 air pressure. 
 
 These displacement pumps, or tanks, may also be arranged 
 in multiple series placed at various elevations between the bot- 
 tom and the top of the shaft. At each station the displacement 
 is effected in the manner indicated by a pressure sufficient to 
 raise the water to the nearest tank above. 
 
 The Air-lift Pump. In this extension of the displacement 
 system air is delivered from a necessarily small pipe into the 
 stand-pipe, where the water column is broken into short lengths 
 by small volumes of entering air, whose expansive power over- 
 comes gravitation. The efficiency of the system is regarded as 
 low. 
 
 Water-pressure Engines are located at the mine below some 
 source of supply. The water flows under pressure against the 
 piston of the pump, which communicates the power to the rods 
 of the Cornish system, or to a direct-connected plunger which 
 lifts and forces the drainage from the mine (Fig. 148). If the 
 engine be placed underground, a still greater head is obtained 
 for power, the discharge being effected at a tunnel level inter- 
 mediate between the shaft mouth and the sump level. This is 
 the method at the Comstock mines. The great trouble with these 
 engines is in the valve-gear, which is more complex and must 
 be more nicely balanced for an incompressible fluid like water 
 than is necessary for an elastic fluid like steam or air. The sudden 
 shutting of the valves produces a not inconsiderable concussion 
 in the inelastic fluid, which is entering the cylinder at a high 
 velocity. Unless delicately manipulated the valves fail to operate. 
 
 They will operate at any reasonable rate within the limits of 
 the valve movement. The fluid, being inelastic, requires a very 
 heavy construction to resist the static pressure and the water^
 
 372 
 
 MANUAL OF MINING.
 
 PUMPING. 373 
 
 ram. Ample time must be allowed for pause at the end of each 
 stroke. The hydraulic engines are expensive plants. They can- 
 not be used for shaft-sinking, as the working pressure cannot be 
 altered to meet the increasing resistance. Their field as prime 
 motors at the surface has been reduced by the much cheaper and 
 .more durable high-pressure impulse-wheels, driving centrifugals. 
 Formulae for Pump Calculations. 
 Let c= diameter of the water- cylinder, inches; 
 N = number of strokes per minute; 
 k= diameter of the steam-cylinder, inches; 
 s= stroke, inches; 
 d= diameter of the pipe, inches; 
 
 h= height of head lost due to friction in the pipes, feet; 
 L= height of the lift including the suction, ?, feet; 
 G= discharge in gallons per minute; 
 Q = cubic feet per minute; 
 /=pump efficiency plus the slip, in per cent. 
 Then 
 
 G=o.oo34c 2 sNf; Q =0.0x2045 4C 2 sNf. 
 
 The pressure per square inch of water-piston = 0.434!,. 
 
 The pressure on the bottom of the pipe =o.^iLd 2 . 
 
 The suction vacuum in inches =30(33.8 /). 
 
 Total pressure on the water-piston =o.34ic 2 . 
 
 k*p =0.341 (L+h}c\ 
 
 The work of the pump =0.0002 $$GL + /. 
 
 The I.H.P=o.ooo253G(L+/0-w/- 
 
 EXAMPLES. i. Let it be required that 100 gallons of water weighing 
 833 Ibs. be raised 200 feet in one minute. The number of ft.-lbs. of work 
 necessary will be 166,600 ft.-lbs. When the mechanical efficiency is 0.6 
 the I.H.P. = 8.4 3 . 
 
 2. Required the horse-power necessary to furnish 60 cubic feet of water 
 through 400 feet of 6-inch pipe and a suction-lift of 4 feet. 
 
 According to the formula, Chapter VI, the head of friction is equal to 
 3 Ibs. per square inch; the total pressure due to the column of water is 
 176.6 Ibs. per square inch; the velocity through a 6-inch pipe is equal to 
 300 feet per minute; whence the work done is 179.6x28.3 X3oo = 46.1 H. P. 
 The steam end must be capable of furnishing more power than this, and if
 
 374 MANUAL OF MINING 
 
 the initial steam pressure be ico Ibs. absolute, with a back pressure of 
 18 Ibs., the diameters of the steam and water cylinders with ^=0.50 and 
 a piston speed of 200 feet per minute, will be 16 and 8 inches respectively. 
 The pump receives its supply from a boiler which is independent of those 
 feeding the hoisting-engines, because the intermittent work of the latter 
 causes such changes in steam pressure as to seriously affect the speed of 
 the pump. 
 
 3. A mine delivers ifoo gallons per minute. The depth of the shaft is 468 
 feet. Required the size of the pump-cylinders under a boiler pressure of 
 100 Ibs. (gauge) and a $ cut-off, back pressure being 16 inches of mercury. 
 Efficiency, 60 per cent. 
 
 As the ordinary piston speed is 200 feet per minute, the discharge of 4.227 
 cubic feet per second may be delivered at the same speed in the pipe, which is 
 then of a diameter of 15 inches; or if a jo-inch discharge-pipe be employed, 
 the velocity therein would be 464 feet. The loss of head would be respectively 
 1.56 and 11.86 feet. Total head being then 470 or 480 feet, the work of rais- 
 ing the 1900 gallons would be 7,441,500 and 7,599,840 ft.-lbs. respectively. 
 The two steam ends must be capable of 12,402,516 and 12,666,400 ft.-lbs. 
 From the table in Chapter IV, the mean pressure corresponding to a cut-off 
 of J is 0.726 for i lb., and 83.27 Ibs for 114.7 absolute. 
 
 Let the average effective pressure be 64 Ibs., then the diameter of each 
 steam end should be \2\ inches. Assuming a stroke of 24 inches, each stroke 
 represents 2.53 cubic feet, and the diameter of each water-cylinder would 
 be lof inches. If the minimum effective pressures upon the two steam- 
 pistons be taken (25 Ibs. per square inch on one and 64 Ibs. on the other), 
 the diameter of the water-cylinders should not exceed i\ inches. This 
 discussion neglects the inertia of reciprocating parts. 
 
 If the discharge-pipe be assumed at 6 inches in diameter, it would entail a 
 loss of head of 151.47 feet, requiring 475 horse-power. 
 
 The ratio between the diameters of the two ends of a steam-pump is about 
 i : 2 for the smaller sizes, and the steam end is three times that of the water 
 end in the large sizes. 
 
 4. What volume could be raised by a double acting steam-pump having 
 water-cylinders 8 inches in diameter, and the steam-cylinders 18 inches, with a 
 2-foot stroke ? Piston speed 200 feet per minute. 139.4 cubic feet. 
 
 5. What should be the effective steam pressure to discharge the water, 
 assuming an efficiency of 50 per cent and a shaft 400 feet deep? Assuming 
 a discharge-pipe of 5 inches diameter, p is 83.7 Ibs. per square inch, the loss 
 of head being 88.22 feet. 
 
 Power-driven Pumps. Where it is not possible or desirable 
 to connect the water-plungers directly with an engine, power 
 may be communicated to a pulley on a rotary shaft by a belt
 
 PUMPING. 375 
 
 or wire rope, or to an electric motor. If the distance between 
 the engine and the pump is very great, electricity is the agent 
 for the motor. If the distance is slight, a belt is employed, with 
 or without gearing, to reduce the speed of the motor to that 
 of the pump, which is always low. Power- pumps are built 
 with more than one trunk-cylinder whose pistons are single- 
 acting, the shaft being a multiple crank. With three plungers 
 the cranks are 60 apart, with four they are 90, etc. Trip- 
 levers can be attached to relief-valves, so that, when the pres- 
 sure becomes excessive, they will permit the water to escape and 
 simultaneously close the suction-valve. 
 
 The cylinders are generally vertical, though in the Riedler 
 pump they are horizontal. Double-acting duplex or triplex 
 pumps can also be had, of the power-driven type, which are used 
 for pressures up to 700 Ibs. in the anthracite region. 
 
 The power-pump is of practical utility and simple, occupying 
 less floor space for a given capacity than the direct-acting. It 
 is of low speed and is therefore geared down from the driving 
 source. Its loss by friction is a little greater than in the direct- 
 connected pump. Where fuel is expensive it has a distinct 
 advantage in the point of steam economy, for the direct-acting 
 type cannot use steam expansively. The power-driven pump 
 may be operated by a modern automatic cut-off engine and 
 develop power with less steam than the direct-acting. The latter 
 cannot use the steam expansively, as is possible with the engine 
 and a fly-wheel-crank-driven pump. 
 
 The Riedler Pump has mechanically operated valves which 
 will open and close at high rotative speed. Its capacity is large 
 for the space it occupies and it requires little foundation. At 
 its maximum speed of 300 r.p.m. it can raise water 1000 feet. 
 The single-acting plungers are installed two or three on a line, to 
 lighten the work and economize power. 
 
 Electrically Driven Pumps are necessarily of the geared-crank 
 type, for this power is not adapted to reciprocation. Either 
 continuous- or alternating-current motors are suited for them. 
 The main requirements are a practically constant moment or
 
 376 
 
 MANUAL OF MINING. 
 
 torque, and a nearly constant speed. With a continuous current 
 the compound- wound motor is employed; if a variable speed is 
 demanded, a rheostat in the shunt-field will suffice. Its advan- 
 tage over an ordinary shunt-motor is that its series-coils obviate 
 the wide variation of current which would occur with the latter 
 when passing through the different points of the pump cycle. 
 It will not race if the pump happens to lose its water. Prefer- 
 ably, it should not be enclosed. In the shunt-wound motor 
 the field resistance will rise as the enclosed machine gets hot, 
 thus causing higher speed and armature current. The diffi- 
 culty can be avoided by a few additional turns of series winding. 
 
 FIG. 149. Electric Power-pump. 
 
 With the alternating current the short-circuited squirrel-cage 
 type of armature induction-motor, designed for any desired turn- 
 ing moment and operated at a constant speed, is well adapted 
 to drive the pumps. 
 
 The leakage loss of electric pumps is small and their friction 
 loss low, while the expense of maintenance and repairs is a
 
 PUMPING. 377 
 
 minimum. One compartment may be saved in the shaft, since 
 the wires and the discharge-pipes can be placed in either of the 
 hoistways without interference. In common with the other 
 motor-driven pumps, it is portable. It is started and stopped 
 easily by the ordinary switch with an automatic release-switch. 
 
 Like the power end of the air- and water-driven pumps, the 
 motor may be operated even under water, for it can be com- 
 pletely housed with its sinking-pump. 
 
 Electrically driven pumps require automatic switches which, 
 when the water level reaches a given point, engage a lift that 
 moves the switch and opens or closes the circuit, as may be 
 desired. 
 
 Rotary Pumps are employed for lifting and forcing water 
 against low head. This type of pump is compact and self-con- 
 tained and will deliver more water for a given weight and space 
 occupied than will reciprocating pumps. It is usually driven 
 by belting or wheel gearing. Rotary pumps may be divided 
 into several classes, according to the forms of the pistons or im- 
 pellers, or according to the arrangement of the butments. Two 
 rotary impellers receive, in the space between them from the 
 suction below, a volume of water which is carried with the evolu- 
 tion of the impellers around to the upper side of the pump. 
 The butments receive the force of the water, and prevent the 
 latter from being carried around the cylinder, thus compelling 
 it to enter the delivery, where it is discharged. The Sturtevant 
 blower has two rotary pieces, and the Root three lobes. 
 
 In some pumps the butments are movable and are arranged 
 to be pushed back, as they revolve, to allow the piston to pass a 
 given point; in others the pistons give way when passing fixed 
 butments; and in others the pistons are fitted with a movable 
 wing, as in the fan, which slides radially in and out when pass- 
 ing the butments. These pumps have no packing or springs, 
 are quite durable, but have a tendency to become noisy as the 
 gearing wears. 
 
 The Centrifugal Pump. This is the cheapest and, under 
 certain conditions, the most efficient of the rotary variety of
 
 378 MANUAL OF MINING. 
 
 pumps where the height of lift is moderate and large volumes 
 are to be moved. Like all of the rotary type, it handles sandy 
 or muddy waters with facility, and has a very short suction and 
 only a moderate height to which it can raise water. It is a fan 
 with a number of blades attached to a shaft, and is turned by an 
 electric motor or steam-turbine at 500 to 800 r.p.m. The inlet 
 
 FIG. 150. Section of a Centrifugal Pump. 
 
 may be on one side only, or on both (Fig. 151). With the latter 
 the inlets and blades may be kept down to smaller sizes for an 
 equal capacity. The water enters at the centre and is delivered 
 at the circumference by the high speed of the blades. The 
 blades are convex, radial, or concave. The former require a 
 higher rate of velocity for a given lift than do radial blades. 
 Their inner ends are curved forward slightly to scoop up the 
 water with a minimum of shock. The inlet velocity should not 
 exceed 3 feet per second. The pump is primed and started
 
 PUMPING. 
 
 379 
 
 by some form of steam- injector which expels the air and fills 
 it with water. 
 
 The Theory of the Centrifugal Pump. The water is driven 
 through the fan partly by the pressure of its blades and partly 
 by centrifugal force, and the water will escape with that force 
 at c (Fig. 150), but with little rotary motion, and will rise to a 
 height, h, corresponding to the centrifugal force. If there be 
 no unavoidable efficiency losses, the relation between the total 
 lift, h, and the peripheral velocity, V, is expressed by 
 
 If the outlet for the water be of proper size and flared to 
 reduce the velocity of discharge gradually, the energy imparted 
 to it by the blades is converted into pressure head. With that 
 due to centrifugal force, the head ideally obtainable from a periph- 
 
 FlG. 151. Double-inlet Centrifugal Pump. 
 
 eral velocity, F, is 2h (Fig. 150). The usually imperfect pump, 
 however, shows a relation between velocity and head of 
 
 in which u is the velocity in the discharge-pipe. The head, h, is 
 only about 60 per cent of that theoretically due.
 
 380 MANUAL OF MINING. 
 
 To increase the discharge for a given head means loss of 
 efficiency, because the speed of the pump and that of the water 
 issuing from the pipe are increased unnecessarily; a reduction 
 of velocity is also attended with loss by fluid friction. Where 
 efficiency is desired the design must be just suited to the precise 
 conditions under which it is to operate. The first cost is low 
 and its efficiency increases to 70 per cent at about 40 feet of 
 lift, beyond which height the efficiency again diminishes, until 
 at a height of 100 feet the efficiency is but 40 per cent. 
 Let A/"=r.p.m. of the wheel; 
 
 H = height of delivery in feet; 
 D = diameter of wheel in feet; 
 
 K= constant; 153 for small pumps and 187 for large 
 pumps; 
 
 then N 
 
 EXAMPLE. A centrifugal pump discharges 180 cubic feet per minute at a 
 peripheral velocity of 30 feet per second. Required the horse-power to 
 drive the pump. The peripheral velocity of the water is 25 feet per second. 
 
 Then momentum lost by the wheel per second=- X 25 = 145. 5 Ibs. 
 
 The work done= 145.5X30=4365 ft.-lbs. per second=8 horse-power. 
 The efficiency being assumed at 70 per cent, the horse-power applied to 
 the pump is 11.5. 
 
 4365 + 3X62.5 = 23.3 ft.-lbs. energy. 
 
 This would raise a pound of water 23.3 feet. 
 
 The Compound Centrifugal Pump. These pumps are essen- 
 tially of a high-speed type with low head of lift. A high lift is 
 attained only by increasing the rotary speed. If, however, two 
 or three wheels, each in its own chamber, are mounted in series 
 on the same shaft, the lift of each single one is multiplied, while 
 yet keeping the speed within moderate bounds (Fig. 152). 
 
 In such a multi-stage pump the water from the discharged 
 chamber of the first impeller is led back to the suction-point of 
 the next impeller through channels in the pump-casing. This is
 
 PUMPING. 
 
 381 
 
 repeated as often as there are impellers. Each pump of the 
 series raises the pressure head from that of its suction. A three- 
 stage pump, with each member capable of 100 feet of head, will 
 then deliver water to a height of 300 feet with, of course, a corre- 
 sponding increase of motor-power. These are much used in 
 placer-mining service instead of a long pipe-line from a distant 
 lofty reservoir. 
 
 The Efficiency of the Centrifugal Pump. The efficiency of 
 these pumps drops rapidly when the difference between the out- 
 let and the inlet pressure heads exceeds 60 feet, and the efficiency 
 of a multi-stage pump is about the same (0.85) as that of a single 
 
 FlG. 152. A Three-stage Pump. 
 
 pump of the same construction, but the advantage of the former 
 lies in the reduced velocity for the given height of lift. On the 
 same work and within reasonable limits, the multi-stage cen- 
 trifugal is slightly more efficient than the single pump, due to 
 the decrease in frictional loss attaining the reduced rotary speed. 
 The steam-turbine, or the electric motor, is capable of direct 
 connection with the centrifugal pump, and this combination
 
 MANUAL OF MINING. 
 
 compares very favorably in steam consumption with the direct- 
 acting type. 
 
 HORSE-POWERS AND FUEL REQUIRED FOR Two- AND THREE-STAGE CENTRIFU- 
 GAL PUMPS FOR HYDRAULIC MINING OR FOR PUMPING PURPOSES. 
 
 Capacity. 
 
 Horse- 
 
 
 Pounds 
 
 
 
 
 Miner's 
 Inches. 
 
 Gallons 
 per 
 Minute. 
 
 Cubic 
 Feet per 
 Second. 
 
 at 70 
 Per 
 Cent. 
 
 Diam- 
 eter. 
 
 Coal, 
 
 I 2 
 
 Hours. 
 
 Short 
 Tons. 
 
 Barrels 
 Oil. 
 
 sa 
 
 5 
 
 562 
 
 1.2 
 
 41 
 
 5" 
 
 2,460 
 
 1.2 
 
 4-4 
 
 2 
 
 75 
 
 843 
 
 1.8 
 
 62 
 
 6" 
 
 3,690 
 
 1.8 
 
 6.7 
 
 3 
 
 100 
 
 1125 
 
 2-5 
 
 82 
 
 6" 
 
 4,920 
 
 2.4 
 
 8.9 
 
 4 
 
 125 
 
 1406 
 
 3-i 
 
 i3 
 
 7" 
 
 6,150 
 
 3-i 
 
 ii. I 
 
 5 
 
 150 
 
 1687 
 
 3-7 
 
 124 
 
 8" 
 
 7,380 
 
 3-7 
 
 13-3 
 
 6 
 
 175 
 
 1968 
 
 4-4 
 
 J 45 
 
 9" 
 
 8,610 
 
 4-3 
 
 15-5 
 
 7 
 
 200 
 
 2250 
 
 5-o 
 
 164 
 
 10" 
 
 9,840 
 
 4.9 
 
 17.8 
 
 8 
 
 225 
 
 2531 
 
 5-6 
 
 185 
 
 10" 
 
 11,070 
 
 5-5 
 
 20. o 
 
 9 
 
 250 
 
 2812 
 
 6.2 
 
 206 
 
 10" 
 
 12,300 
 
 6.2 
 
 22.2 
 
 10 
 
 275 
 
 393 
 
 6.8 
 
 227 
 
 12" 
 
 13,500 
 
 6.8 
 
 24.4 
 
 ii 
 
 300 
 
 3375 
 
 7-4 
 
 246 
 
 12" 
 
 14,760 
 
 7-4 
 
 26.6 
 
 12 
 
 35 
 
 3937 
 
 8.6 
 
 267 
 
 12" 
 
 17,220 
 
 8.6 
 
 28.9 
 
 14 
 
 400 
 
 4500 
 
 9.9 
 
 328 
 
 14" 
 
 19,680 
 
 9.8 
 
 35-5 
 
 16 
 
 500 
 
 5625 
 
 12.4 
 
 410 
 
 14" 
 
 24,600 
 
 12.3 
 
 44-4 
 
 18 
 
 A single-stage pump making 1500 revolutions per minute, 
 with 55 horse-power steam-turbine, delivered 1700 gallons per 
 minute with a lift of 100 feet. The diameter of the plunger is 
 13! inches. A two-stage centrifugal pump, 9 inches in diameter, 
 was making 2000 revolutions per minute with a Dela valve steam- 
 turbine running at 20,000 r.p.m., and raised 250 gallons per 
 minute 700 feet. The tests of these showed the efficiency of 
 the wheel and turbine to be 75 per cent in the first case be- 
 tween the limits of 1400 and 1800 gallons per minute. The 
 duty of the pump per thousand pounds of steam operated 
 with a condenser was nearly 62,000,000 ft.-lbs. The two-stage 
 pump showed a duty at 250 gallons per minute of 48,800,000 
 ft.-lbs. per thousand pounds of steam. Forty-two pounds of 
 steam were used per water horse-power. 
 
 A test of a single-stage pump direct-connected electric motor 
 of 20 horse-power delivering 1200 gallons per minute at 2000
 
 PUMPING. 383 
 
 revolutions for a lift of 45 feet showed an efficiency of 76 per cent. 
 The diameter of its wheel was 8.3 inches. 
 
 EXAMPLES. i. What will be the size of the discharge-pipe, d, the wheel 
 and its rate of revolution, to deliver 1000 gallons per minute for a height of 
 50 feet? 
 
 d= 0.25 v^ =0.25X31.6= 7.8 inches pipe; 
 
 wheel diameter= 2 X 7.8= 15.6 inches; 
 and 
 
 JV= 187X^-^=880.67 r.p.m. 
 
 2. Required the size of a three-throw electric pump, the horse-power 
 and current at 550 volts for a motor of 90 per cent efficiency, to deliver 240 
 gallons 300 feet high. Speed of piston 60 feet per minute. Let }=o.gj. 
 Then =5.86 inches and the stroke i|c=8 inches. 
 
 The work of pumping is 240X8^X300=590,000 ft.-lbs. Assuming the 
 inertia, friction, etc., to be 50 per cent, i requires 36 H.P. at the motor-shaft. 
 With 90 per cent efficiency, this =29,840 watts, which at 550 volts=54-3 
 amperes. 
 
 3. It is desired to raise 90 gallons of water per minute up an incline 
 of i in 5 which is 1200 feet long. Assuming a discharge-velocity of 3 feet 
 per second, what will be the size of the pipes and of the pumps and motor? 
 If the three-throw pump is employed, each pump raises 30 gallons; with 20 
 per cent of slip, 36 gallons must be raised about 60 feet per minute, with a 
 wheel-diameter of 4.25". At a stroke of 8 inches and 90 revolutions per 
 minute, the diameter o the discharge-pipe will be 3.875 inches to deliver 
 90X0.16 cubic feet per minute. A pipe of 4 inches diameter will lose, accord- 
 ing to the formula in Chapter VI, 1.22 feet of head for each 100 feet of length 
 at the assumed velocity. The loss of head due to friction is then 14.6 feet. 
 The work of lifting the water, including the above loss, is 190,980 ft.-lbs. per 
 minute and the horse-power is nearly 6. Allowing a friction of gear and 
 starting of 50 per cent, 12 B.H.P are required. The motor will run at 630 rev- 
 olutions per minute with a gear of 7 to i . If the latter has an efficiency of 80 
 per cent and receives a current at 380 volts, 24.5 amperes will be required 
 to deliver the 10 B.H.P. at 380 volts. The wire for this current must be 
 19/19 and the drop in voltage for the total length of wire down the slope 
 will be 24.5X1.786, which equals 44 volts per mile and 20 volts for 2400 feet. 
 The voltage at the bottom is therefore 400. 
 
 The Pulsometer. This useful apparatus raises sandy or 
 acid waters to limited heights where economy of fuel is less in?
 
 384 MANUAL OF MINING. 
 
 portant than quickness of installation. It is perfectly free from 
 risks of breakdowns, and is employed 
 in open works. A A (Fig. 153) are two 
 chambers into which steam enters accord- 
 ing to the position of the ball, C, which 
 oscillates from one side to another of 
 their necks. Through the inlet passage, 
 Z>, the water enters into A by opening 
 the valves, E. H is a delivery passage 
 communicating with each chamber 
 through openings and valves, G. 
 
 Steam enters at the top, passes into a 
 chamber uncovered by the ball-valve, 
 
 FIG. 153. The Pulsometer. and presses upon the sur f ace o f the water, 
 
 forcing it down and out through the discharge-valves into H. 
 When the water-line falls below the discharge outlet, the steam 
 above condenses, a partial vacuum is formed, and its pressure 
 suddenly falls again. Meanwhile, with the collapse of the steam, 
 the ball- valve is thrown to close that chamber, and admits steam 
 to the other side. Here the water is expelled, in a similar manner, 
 out through H , while at the same time water has entered through 
 D to be expelled from the other side. 
 
 The limit of lift is about 30 feet, and the capacity of some 
 pulsometers is often 1000 gallons per minute. 
 
 Siphons. Though not a water- raising appliance in the proper 
 sense of the term, since water by this apparatus can be lifted 
 or transported over an eminence not exceeding 28 feet in height 
 and discharged on the other side of it at a level lower than that 
 of the supply, a siphon can find application for forming a water 
 communication over a slight elevation between two distant points. 
 This height will be further reduced by an amount necessary to 
 cause the required velocity of flow and to overcome the frictional 
 resistance. Formulae in Chapter VI will determine the losses due 
 to friction in the given case, L and d being given. 
 
 When the siphon is not too long, and when the acceleration 
 head is sufficient to give the water a considerable velocity, the
 
 PUMPING. 385 
 
 air, entrained by the rapid current, may be carried out at the 
 end of the discharge branch, if the latter is not too steep. In 
 most cases, however, it is necessary to provide artificial means to 
 remove the accumulated gases, either periodically or continu- 
 ously. A hand-pump is usually employed for this purpose, its 
 suction being connected to the highest point of the siphon, and 
 operated as occasion requires. 
 
 REFERENCES. 
 
 Compressed-air Pumps, 111. Min. Inst., I, 100; Pumps, Breakdowns, Coll. 
 Guard., Nov. 13, 1896, 567; Mine Pumps, Breakdowns, Coll. Guard., Nov. 
 13, 1896, 930; Pump Valves and Their Care, Dugald Baird, Coll. Eng., 1893, 
 221; Mine Drainage and Pumps, Hans B. Behr, California State Mining Bureau 
 Bull. 9, 210; Pumps, Description, Pumps in England, Gears, Strokes, etc., 
 Coll. Guard., May 7, 1897, 855; Pumps, Calculations of, Bjorling, Coll. Man- 
 ager, Nov. 20, 1896, 567; Pump Cylinders, Determining Strength of, Kettell, 
 Am. Soc. Mech. Eng., XVII; Careof Pumps, M. & M., Vol. XXIV, 323; 
 The Engineer, July 1902; Sizes and Weights of Tubing, O. J. Edwards, 
 E. & M. Jour., LVIII, 387; Pumping by Compressed Air, Frank Richards, 
 E. & M. Jour., LIX, 3 4; Pneumatic Pumping Plant, Proc. Am. ?oc. C. E., 
 XXX, 445; Hoisting Water Tandem Tanks, J. W. Bowden, Amer. Inst. M. 
 E., XX, 343; Mine Pump, Breakdowns, Michael Longridge, Coll. Guard., 
 Nov. 1896, 930; The Contest with Water at the Comstock Mines, D. L. Lord, 
 U. S. G. S., IV, 230; Tapping Drowned Workings, W. B. Wilson, Jr., Coll. 
 Guard., Feb. 14, 1902; Water Hoisting in Anthracite Mines, M. & M., Vol. 
 XXIII, 390. 
 
 Centrifugal Pumps for Mine Work, W. R. Crane, M. & M., June 1902; 
 Tests of Centrifugal Pumps, Engineer, Chicago, June 15, 1904; Theory of 
 Centrifugal Pump, Am. Soc. C. E., Jan. 1904; and Proc. Am. Soc. C. E., 
 May 1903, and Vol. XXIX, 839. 
 
 Determining C O in Mine Air, Coll. Guard., Vol. LXXXIV, 1162; Grav- 
 itation of Fire Damp, Coll. Guard., LXXXI, 124.
 
 CHAPTER XH. 
 
 MINE-GASES. 
 
 Ventilation as an Economic Proposition. The ventilation of 
 a mine is a matter of very great concern, not only from a humani- 
 tarian standpoint, but from an economic point of view. No 
 shaft or tunnel can be carried more than 200 feet beyond an 
 opening without some special means of stirring and freshening 
 the stagnant air. The men should not be compelled to work 
 in the hot atmosphere of a stove or room vitiated by the variety 
 of gases given off from and by coal, powder, lamps, respiration, 
 rotting timbers, and decomposing ore. These gases cannot 
 support combustion, nor can they be inhaled with impunity; 
 and such an atmosphere is unfit for respiration, being deficient 
 in oxygen, as well as by reason of the presence of these gases. 
 In coal-mines additional peril accompanies some of these gases, 
 which with the air form explosive mixtures, which, bursting 
 into flame, destroy everything in their path and emit dense vol- 
 umes of poisonous fumes that are fatal to all who have escaped 
 the shock. 
 
 Recognizing the pecuniary value of the life and energy of a 
 miner, the statutes are becoming more and more rigorous in 
 the insistence of safety and hygienic measures. Methods are 
 specified for rendering the noxious gases harmless, and officials 
 are given sufficient authority to suggest needed improvements 
 and to punish non-compliance. Not only are the miners benefited 
 by the diminishment of risk, but also the operators, who profit 
 in the increased energy of men working under favorable circum- 
 stances. The illumination is better, smoke clears away quicker, 
 and the men are capable of a full day's work. 
 
 386
 
 MINE-GASES. 387 
 
 Comparison of Metal- and Coal-mines. There is a marked 
 contrast between the requirements of vein-mines and those of 
 coal or other mines in flat beds, which latter usually have two 
 shafts connected by a labyrinth of workings on nearly one level. 
 The ventilation of metal-mines presents by no means the same 
 difficulty as that of gaseous collieries, since as a rule the former 
 class of mines requires but a small supply of fresh air, abundant 
 enough for the health, comfort, and effective work of the men 
 and for the removal of the dead air vitiated by various causes. 
 Trust is placed in natural means of circulating the air by the 
 winze communication of the different levels. This is certainly 
 inadequate, and the lower mortality in well-ventilated coal- 
 mines is doubtless due to their better ventilation. In coal-mines 
 large volumes of fresh air, additional to that required for the 
 men, are necessary to carry off the fire-damp, choke-damp, and 
 other gases, both noxious and inflammable, though no serious 
 results may occur from their presence in large quantities. These 
 gases are continually being evolved from the coal and constitute 
 an ever-present danger to coal operations, but do not threaten 
 metal-miners. Nevertheless, the latter should remember that 
 the inflammable gases of coal-mines, from which they are com- 
 paratively free, are not the only ones to be guarded against. The 
 cloud of dust formed by blasting the mineral, and the carbonic 
 acid gas, are dangers equally insidious to the health of those 
 inhaling it, which cause as many deaths as do the small explo- 
 sions which occur in coal-mines. 
 
 Gases in Coal-mines. In addition to the gases found in a 
 normal atmosphere, there exist in mines several gases which 
 are always the result of decomposition and combustion. Pur- 
 suing the nomenclature of early mining, they are designated 
 as " damps." For example, carbonic acid, CO 2 , is known as 
 choke-damp; carbonic oxide, CO, as white damp; sulphuretted 
 hydrogen, H 2 S, as stink-damp; carburetted hydrogen or marsh- 
 gas, CH 4 , as fire-damp or fulminating damp; air vitiated by 
 breathing, having therefore a deficiency of oxygen, as black 
 damp; and the residual gases of an explosion, after-damp. In
 
 388 MANUAL OF MINIXG. 
 
 metal-mines carbonic acid and sulphuretted hydrogen are the 
 only gases met with. Both are heavier than air and naturally 
 will be found near the floor of the workings. Coal-mines, how- 
 ever, are troubled with additions and emanations, sudden or con- 
 tinual, of carbonic oxide and marsh-gas, which two, being lighter 
 than air, are to be looked for in the upper portions of the work- 
 ings, near the roof. 
 
 Composition of the Atmosphere. Every hundred grains of 
 the atmosphere contain 76.84 grains of nitrogen, 23.10 grains of 
 oxygen, and 0.06 grain of carbonic acid, occupying respectively 
 79.02 per cent, 20.94 per cent, and 0.04 per cent of the original 
 volume. These gases are not in chemical combination, but as a 
 mechanical mixture from which the oxygen may be extracted 
 under given conditions. 
 
 Oxygen, which is colorless, odorless, and transparent, unites 
 with all other chemical elements, forming various chemical com- 
 pounds. All the ordinary phenomena of light, heat, and fire 
 are the result of the union of other elements with oxygen. This 
 active principle is essential to life and to all other processes of 
 combustion, and when combined with carbon forms choke-damp 
 or white damp; with pyrites, a common mineral in coal, sul 
 phuretted hydrogen; and with marsh-gas, water or steam and 
 carbonic acid. An atmosphere deficient in oxygen will no longer 
 support combustion, and a lamp or flame immersed in it is extin- 
 guished; so, too, air-breathing animals are suffocated when the 
 percentage of oxygen in the air is below 15. 
 
 Nitrogen, the predominant constituent of unpolluted air, 
 is colorless, odorless, and incapable of supporting combustion 
 or animal life. It is inert in its effect upon the system, but extin- 
 guishes flame immersed in it. 
 
 Carbonic Acid, which is not an important constituent of 
 the atmosphere, is, however, to be expected in all abandoned, 
 and unventilated places. It is given off in quantity by the respira- 
 tion of men and animals, by the combustion of lamps, fuel, 
 explosives, timber, and organic matter, and by all substances 
 in a state of decay or fermentation. It, also, is colorless and odor-
 
 MINE-GASES. 389 
 
 less, and at a normal pressure of 30 inches of mercury and a 
 temperature of 32 F. weighs 128.45 Iks. P er thousand cubic 
 feet, as against 81 Ibs. for an equal volume of air under the same 
 conditions. It is easily detected, for, being incapable of sus- 
 taining combustion, light burns dimly when immersed in it. 
 When, therefore, the gas is revealed by this action on a flame, 
 the atmosphere is at least unfit for respiration. The gas has an 
 injurious effect upon the human system Air containing 2 per 
 cent of it produces an overwhelming depression upon those 
 breathing it; with 6 per cent, lights are extinguished; and with 
 10 per cent, it is positively fatal by producing suffocation. As it 
 accumulates in the lower part of idle workings and where the 
 air is stagnant, no one should venture into abandoned works 
 without having previously tested by a flame the condition of 
 their atmosphere, and if the light is extinguished, the poison 
 must be swept out by a strong current of air, which, in shafts, 
 may be incited by rapidly raising and lowering a bundle of hay, 
 which is then ignited. Small volumes of carbonic acid may 
 be removed from the atmosphere by the use of absorbents like 
 lime or ammonia. Its presence may be chemically detected by 
 the milkiness produced in a test-tube containing baryta hydrate. 
 
 Sulphuretted Hydrogen, which at 30 inches barometric pres- 
 sure and a temperature of 32 F. weighs 94.62 Ibs. per thousand 
 cubic feet, is an extremely poisonous and common gas occurring 
 in mines. As it is the result of the decomposition of pyrites, 
 always accompanied by heat, this gas is a warning of incipient 
 fires. In the gob and abandoned portions of the mine the pyrites 
 and coal-waste subjected to pressure from the roof, and in the 
 presence of moisture, decompose with the development of heat 
 and possibly flame, of which the presence of sulphuretted hydro- 
 gen is an important indication. Being colorless but strongly 
 odorous, its presence is readily detected. It does not support 
 combustion, but is itself inflammable. A flame will burn in a 
 mixture of it with air, but is extinguished in an undiluted atmos- 
 phere of this noxious gas, which in this state is fatal to life. 
 
 Carbonic Oxide, which is claimed to be a normal gas exist-
 
 39 MANUAL OF MINING. 
 
 ent in coal-mines, weighs 78.305 Ibs, per one thousand cubic 
 feet at a temperature of 32 F. and a barometric pressure of 
 30 inches. This gas has neither color, taste, nor smell, and is 
 exceedingly poisonous. One half per cent of this gas in air ren- 
 ders the atmosphere fatal to life if breathed for ten minutes. It 
 acts upon the system by combining with the oxygen absorbed in 
 the blood, forming a stable compound reducing the haemoglobin, 
 and insidiously and surely destroying the blood and tissues. As 
 carbonic oxide is an unstable compound usually resulting from 
 the imperfect or secondary combustion of a gas or of carbon in 
 the deficiency of oxgyen, the existence of this gas in mines is 
 doubted. It is only barely possible that it might be found in 
 the goaf, where the oxidation of pyrites and the absorption of 
 oxygen by the fine coal-dust may have depleted the air of its 
 vital element, thereby giving rise to a sufficiently high temperature 
 to incite combustion with the development of carbonic oxide. 
 
 Light Carburetted Hydrogen or Marsh-gas is a stable, never- 
 failing constituent among the products of dry distillation of 
 organic matter, and exists as the predominant constituent of 
 the compound gas known under the general term of fire-damp 
 because of its ready ignition by flame with a mixture of air. Its 
 weight at the normal barometric pressure and a temperature of 
 32 F. is about 45.22 Ibs. per thousand cubic feet. It is absorbed 
 in the coal, diffused through its pores, collected in crevices or 
 cavities, and even stored up in reservoirs, having been exuded 
 from the coal in the early stages of decomposition of the organic 
 source of the coal, or expressed during the geological movements 
 of the earth's crust. It is not a constituent of the coal, but is 
 entirely free from chemical combination with it, and continually 
 exudes with greater or less violence; it is liberated in volumes by 
 falls in the mine- roof, by squeezes or creeps, and by any sudden 
 fall in the barometric pressure. The deep portions of coal- 
 seams appear to be more heavily charged with this compound 
 gas than are workings in coal-beds so close to the surface as to 
 allow of its escape. No coal-seam should be regarded as free 
 from liability to irruptions of this gas.
 
 MINE-GASES. 39 1 
 
 The Occlusion of Gas in Coal. Marsh-gas exists free in 
 the pores of all coals, but must not be confounded with the gases 
 chemically compounded with the coal. It is under high pressure,, 
 and is given off into the workings of the mine with greater or less 
 violence from the fissures or crevices, sometimes without warn- 
 ing, sometimes accompanied by the heaving of the floor and the 
 trembling of the roof, but always more or less distinctly audible. 
 The volume so emitted from the coal varies in amount from 9 to 
 100 cu. ft. per ton. From anthracite the discharge is most copi- 
 ous, and from bituminous coal -the least. 
 
 Frequently it escapes into the mine without warning; its 
 presence is not always detected or manifested, as it accumulates 
 in a nook or under a platform until some untoward circumstance 
 brings it into contact with a naked light. The slight hiss accom- 
 panying its exudation is hardly enough to be distinguishable. 
 These "blowers," of all sizes, up to the outbursts that for a time 
 overpower the ordinary ventilating current, contain 90 per cent 
 of marsh-gas, and may be liberated anywhere and at any time. 
 For a long time it has been recognized as a constituent of the 
 gases entering mines. It escapes at various springs and salt- 
 mines; it has fed the sacred fires of Baku and the mud- volcanoes 
 of Bulganak; it has been found in the Silver Islet mine, in the 
 iron-mines of Alsace, and in the lead-mines of Tuscany; and 
 some years ago an explosive gas was met with in driving the lake 
 tunnel at Chicago. It constitutes from 40 to 90 per cent of the 
 natural gas, and is obtained among the volatile and combustible 
 constituents in the ultimate analysis of coal. While sinking 
 shafts through porous strata fire-damp has been encountered, 
 and precautions are therefore necessary in regions of natural 
 gas, or in formations immediately above the coal horizons. 
 
 Outbursts of Gas from the Coal. Measurements made by 
 the Royal Commission on Accidents in Mines have revealed 
 the pressure of gas to be frequently as great as 450 Ibs. per 
 square inch. Sudden outbursts may therefore be expected when 
 escape is afforded, and these, according to the treatise of M. 
 Roberti Linterman, preponderate in coal-seams disturbed by
 
 392 MANUAL OF MINING. 
 
 faults, foldings, or thinnings-out, and are influenced by the dip 
 of the seam. They occur without any premonitory symptoms 
 and even in districts heretofore free from gas. They are most 
 voluminous in the periods of pillar-robbing or during the ex- 
 ploration of virgin ground. The disengagement of gases in- 
 creases both in intensity and frequency with the depth of the 
 workings and with the presence of permeable masses surrounded 
 by rock-masses so hard and compact as to constitute effective 
 barriers against fire-damp. In every district are found one or 
 more infested zones of gas. 
 
 Owing, therefore, to the inevitable occurrence of this in- 
 flammable gas in all coal-seams, and the uncertain quantity 
 which may be thrust into contact with the flame of the illumi- 
 nating-lamp or of the explosive employed in the mine, fire-damp 
 is the dread enemy of coal-miners. The amount of gas which 
 renders the mine unsafe cannot be stated, for, while an atmos- 
 phere containing 2 per cent or more is neither injurious to life 
 nor dangerous in mines, nevertheless its presence to that extent 
 in the air discharged from the mine into our atmosphere indicates 
 a probably excessive accumulation in some of the workings. 
 The Coal Commission of Austria, in its classification of mines 
 according to the composition of the air at the outlet of their venti- 
 lating-shaft, regards those having more than 2 per cent of fire- 
 damp and carbonic acid in their return air as fiery, and those 
 having less than i per cent of gas admixture as safe. 
 
 The Effects of Mine-gases upon Life or Flame. Mine-gases 
 mixed with air do not have the same effect upon life or an illu- 
 minating-flame. In an undiluted state each and all extinguish 
 flame and do not sustain life. When mixed with air the carbonic 
 oxide and sulphuretted hydrogen are poisonous, while at the same 
 time they support combustion. The marsh-gas is inert in its 
 influence upon life, but is capable of ignition; while the carbonic 
 acid is depressive in its influence both upon flame and upon 
 life. There are hence two methods of discovering the probable 
 unsuitability of air for respiration the extinction of flame by 
 an excess of carbonic acid, and the flame aureole from the com-
 
 MINE-GASES. 393 
 
 bustion of carbonic oxide, the odor of sulphuretted hydrogen 
 being a sufficiently strong warning without other index of its 
 noxious presence. Fortunately, however, the presence cf car- 
 bonic oxide need only be feared after a fire-damp explosion or 
 after a blast of one of the lower grades of black powder. 
 
 Explosive Gaseous Mixtures. Marsh-gas with air will burn 
 freely, but when the proportion of gas reaches a certain specific 
 amount the ignition may take place rapidly, and if the products 
 of combustion cannot escape equally rapidly, explosion ensues, 
 the force of the explosion and the dangerous degree of dilution 
 of gas varying with the different gases. 
 
 The range of proportion of gas dilution between the lower 
 and the upper explosive mixture is least in the case of fire-damp, 
 which may vary in amount between 5 and 13 per cent of the 
 total volume of air; is greatest in the case of carbonic oxide , 
 which will explode with any mixture containing between 13 
 and 75 per cent; while with sulphuretted hydrogen the variation 
 lies between 9 and 28 per cent of gas in the mixture. The gases 
 are mentioned also in the order of the decreasing danger of 
 explosibility, the first offering the greatest risk. When the 
 atmosphere in which a flame is immersed contains a percentage 
 of gas approaching the explosive limit, the cap and nimbus become 
 large, and the flame almost invisible. The rapidity with which 
 the ignition is propagated depends upon the nearness of the 
 proportions of the admixture to the figures given, and when 
 the maximum explosive ratio as indicated above is reached, the 
 propagation is instantaneous, and the concussiv force of the 
 explosion is also a maximum; as the percentage of gas recedes 
 from these ratios or increases beyond the limiting explosive 
 proportions, the violence of the explosion decreases. When 
 either of the gases is undiluted with air, the light placed in con- 
 tact with it is extinguished. In other words, a flame may be 
 immersed in the workings filled with fire-damp; and if the line 
 of demarcation between the air and the gases is "sharp," no 
 explosion will ensue, but the flame will enlarge and, after a little 
 fluttering, become extinguished. This is equally true of the
 
 394 MANUAL OF MINING. 
 
 other gases', though the percentage of their accumulation in 
 mine-workings is so low and their diffusion so perfect that no 
 undiluted accumulation of them is likely to ensue. When igni- 
 tion or explosion takes place, the products of combustion are 
 termed the "after-damp." 
 
 But one means is available for the prevention of excessive 
 accumulations of explosive gases, and this consists in supplying 
 a copious volume of air and a thorough system of distribution 
 which will dilute the noxious emanations below the danger-line. 
 This will require careful examination daily of all suspected 
 places and the enforcement of rigid discipline. The airways 
 must be ample in area to allow the requisite volume of air 
 to pass without producing a current of high velocity. In mines 
 using the common Davy or the Clanny lamps, the maximum 
 velocity admissible is 300 feet per minute. Where bonneted 
 lamps are in use the velocity may be 800 feet without fear from 
 explosion. Immunity from explosions is possible only by ad- 
 herence to these requirements and care in the use of blasting 
 agents. 
 
 Black Damp; After-damp. This residual gas, after an 
 explosion or ignition of gas, will extinguish a flame because of 
 the deficiency of oxygen, and is theoretically composed of 52 
 per cent of nitrogen and 48 per cent of carbonic acid. The per- 
 centage of mixtures of oxygen, nitrogen, and carbonic acid in 
 an atmosphere extinctive of flame is almost identical with that 
 of the air expired from the lungs of men. Dr. F. Clowes, as the 
 result of a series of experiments in relation to the lighting of 
 mines and the behavior of lamps, has ascertained that the per- 
 centage composition of the residual atmosphere in which flame 
 was extinguished, is, oxygen 15.7, nitrogen 81.1, carbonic acid 
 3.2, while that of the average exhaled air is oxygen 16.15, nitro- 
 gen 79.9, and carbonic acid 3.95. 
 
 Respiration in such an atmosphere is difficult, and produces 
 unconsciousness, followed by marked panting in the effort to 
 supply oxygen to the lungs. The countenance becomes swollen 
 and livid, the features distorted, the eyes protrude, and as
 
 MINE-GASES. 395 
 
 asphyxia is pronounced, there is a sudden cessation of the pul- 
 sations of the heart and of respiration. 
 
 Usually some steam and carbonic oxide exist in the after- 
 damp in proportions varying with the temperature of explosion 
 and the initial proportions of air and explosive gas. In the 
 presence of the latter, even in minute quantity, is the gravest 
 danger. A reduction of oxygen to 14 per cent or below also 
 causes disastrous results. 
 
 To the presence of this carbonic acid is attributed the loss of 
 many -lives in a mine explosion more, in fact, than are the 
 result of its concussion or of contact with its flame. The per- 
 fectly natural appearance of the body, lying often by a lamp 
 still burning, proves the cause of death to be some insidious 
 poison which is combustible, not asphyxiation or concussion. 
 All members of rescuing parties entering the workings thereafter 
 should take due precautions against the inhalation of this carbonic 
 oxide gas. 
 
 Treatment for Asphyxiation. Persons overcome by any 
 gas may be revived by blowing oxygen into one of the nostrils, 
 the other being closed, and by inducing artificial breathing. 
 Epsom salts, and water acidulated with vinegar, are better than 
 alcoholic stimulants. The warmth of the body should be kept up, 
 and mustard plasters applied over the heart and around the 
 ankles. If these produce no effect, recourse must be had to 
 blood-letting from the foot or jugular vein, and, as a last resort, 
 an opening into the trachea, through which pure air is forced. 
 
 Those overcome by the inhalation of carbonic oxide can 
 be resuscitated only by prompt action and a copious supply of 
 pure oxygen to the lungs. 
 
 The Force of the Explosion. Two volumes of marsh-gas 
 (CH 4 ) combine with 19 vols. of air and develop 23,550 heat- 
 units, giving a temperature of 6064 F. (6525 absolute) and 
 a pressure of 185.3 I DS - per square inch. If m be the weight and 
 c the specific heat of a gas, the heat required to raise it / is ex- 
 pressed by mtc. To raise the 2.75 Ibs. of CO 2 , 14 Ibs. of N and 
 2.25 Ibs. of water from 52 F. to / F. require
 
 39 MANUAL OF MINING. 
 
 2.75 Xo.iyn/ =0.470/5 
 14 Xo.i74o/=2.4i8/; 
 and 
 
 2.25(160+990) + 2.25 Xo.2675/ =0.602*+ 2587; 
 whence 
 
 o.47o/+2.4i8/+o.6o2/+ 2587 =23,550. 
 
 The force of the explosion of two volumes of marsh- gas 
 developing 23,550 heat-units may be ascertained to be nearly 
 30,000 Ibs. per square foot by the following analysis: 
 
 Assuming the initial temperature of the marsh-gas and the 
 mine air to be 62 F. or 523 absolute, because it requires 3.886 
 heat-units to raise the aggregate products of combustion one 
 degree, the degrees to which the final gaseous products of com- 
 bustion will be raised are 
 
 23>55o- 3-886 =6064 F. 
 
 The volume which these products seek to occupy is 
 (523+ 6064) -5- 523 =12.6 atmospheres. 
 
 12.6X14.7 per square inch equals 26,671 Ibs. per* square foot. 
 
 The Barometric Relation of Explosions. An attempt has 
 been made to hypothecate a relation between the periods of gas 
 outbursts and the movements or seasons of low barometer, has 
 failed to show any connection. A falling barometer has not 
 invariably been followed by a heavy discharge of gas, nor does 
 a study of the tables show its unfailing precedence to the evolu- 
 tion. While laying stress on the acknowledged fact that 
 December is the worst month, there appears to be no "off day" 
 for explosions, which are equally abundant on any day of the 
 week. An excessively low barometer at the sea-level is 28.3 
 inches a fall of only 6 per cent of the total pressure, and of 
 but i per cent, or less, of the pressure of the magazine gas. Upon 
 the emissions from the pores of the coal and from goaves the 
 -effect of a barometric depression is noticeable. But even here
 
 MINE-GASES. 397 
 
 an acre of ground of standard thickness will, with a barometric 
 fall of o.i inch, exude only 18 cubic feet of mixture for every 25 
 yards of length of face exposed. 
 
 A tabulation of the barometric variations with reference 
 to mine explosions was made in Westphalia, during 1896, with 
 the result that of the 42 explosions recorded 41 are attributed 
 to fire-damp alone, while in one of them coal-dust participated; 
 out of the total number, 27 happened when the air pressure 
 showed a tendency to fall suddenly or was at its minimum, 
 while in the 15 others the air pressure was at maximum, or showed 
 a tendency to rise. As regards the places where they occurred, 
 the explosions are divided into eight in exploring or preparatory 
 workings in rock, 26 in preparatory workings in coal, and eight 
 in the actual getting of coal, while the gas issued slowly in 28 
 and suddenly in 14 cases. 
 
 The Diffusion of Gases. It is fortunate that the gases evolved 
 from coal or produced by the various processes of decomposition, 
 combustion, and exhalation do not accumulate in separate layers 
 in the workings, with the heavier gas at the bottom and the lighter 
 one near the top, except in abandoned places where the air is 
 allowed to stagnate; but instead of this even a little circulation 
 will set up an individual motion of the separate particles of the 
 gases, by which they become gradually diffused throughout the 
 mass until, after sufficient time has elapsed for the purpose, 
 they are found intimately blended, whatever may be their relative 
 densities. This is not a chemical mixture, but a purely mechanical 
 blending, depending upon the relative tensions of the gases. 
 The rapidity of this diffusion into atmospheric air is inversely 
 proportional to the square root of the density of the penetrating 
 gas. Marsh-gas, therefore, mixes most readily with the air, 
 carbonic oxide not quite so readily; sulphuretted hydrogen less 
 so, and carbonic acid making with difficulty an intimate mixture 
 with air. This principle of diffusion is an exceedingly valuable 
 one to the safety of the mine and the purity of its air, since the 
 more ready the diffusion of the gas the more easily will the 
 gas be cleared away. Thus, by the creation of an air-current
 
 3^8 MANUAL OF MINING. 
 
 throughout the workings, the gases are mixed with the circu- 
 lating pure air, are diluted, and swept out of the mine. 
 
 Testing Mine Air for Gas. The condition of the coal work- 
 ings is usually examined and tested daily by the fire-boss, one 
 of whose duties consists in ascertaining the degree of saturation 
 of the mine air at every place of work before the men are per- 
 mitted to enter. The test is made by a candle or safety-lamp, 
 the flame of which gives evidence of the presence of an accumula- 
 tion of combustible gas. This method requires a skilful, steady 
 hand and considerable nerve. Shading the flame of a candle or 
 lamp with one hand, and raising it upward, the fire-boss watches 
 the behavior of the light. If any inflammable gas is present, 
 the flame elongates and becomes smoky. In this event the 
 test ceases, the flame is lowered, and the fire-boss withdraws. 
 The face of the coal or the room showing these symptoms of 
 danger is then supplied with more air, the employees being mean- 
 while barred from entry to the place. 
 
 The Height of the Flame in Gas. A gas gives evidence of its 
 presence upon a flame immersed in it by the elongation of the 
 flame, surrounded by a blue nimbus or aureole. The more vola- 
 tile the illuminating-oil used in the lamp, the more sensitive is the 
 flame to the presence of these gases. Thus naphtha, benzine, 
 and kerosene, in the order named, are far more sensitive indi- 
 cators of gas than 13 the heavy lard-oil. 
 
 The Height of the Cap on the light of an ordinary safety-lamp 
 depends upon the percentage of gas to that of air in the mix- 
 ture. By observing the height one may determine the ap- 
 proximate percentage as follows: Divide the constant 94,000 
 by the height of the cap in eighths of an inch and take the cube 
 root of the quotient. Thus, when a blue cap is found in a mix- 
 ture under test to be 2 inches in height, then the cube root of 
 94,000 divided by 16 equals 18. In other words, there are 18 
 parts of air to i of gas. If, in another test, the length of the 
 blue cap be only J of an inch, then there will be 46 parts of air 
 to i of gas. 
 
 Testing Lamps. Ordinary safety-lamps do not reveal the
 
 MIN&GASES. 399 
 
 presence of a quantity of gas less than 2 per cent. The Hepple- 
 wite-Gray is more sensitive than the unbonneted Davy or Clanny. 
 It burns benzoline, and shows a cap \ inch high in the presence 
 of i per cent of CH 4 . The Pieler spirit-lamp is always a good 
 gas-tester, which in air containing \ per cent of combustible gas 
 will give evidence of it in a cap i inch high. The Wolf safety- 
 lamp, burning naphtha, shows a very conspicuous halo when 
 placed in a mixture of fire-damp and air; in a mixture of \ per 
 cent of gas the flame is H inches in height, and in 2| per cent 
 of gas the flame is broader, and may even be extinguished. The 
 Beard- Mackie lamp carries a graduated scale making the height 
 of the cap visible. A bent inverted U rod has platinum wires 
 stretched across the arms, at suitable intervals, which become 
 luminous by contact with the flame. The highest wire furnishes 
 the guide to the test. 
 
 The Shaw Gas-testing Apparatus recognizes the presence of 
 explosive gas and is used for an approximate test of mine air. 
 It meets with favor where the exact analysis of the mine atmos- 
 phere is not required. While the apparatus is capable of demon- 
 strating the quantity of explosive gas in the atmosphere, it is inca- 
 pable of distinguishing between them, and thus fails to furnish 
 any clew as to the variety of gas therein contained. Of other 
 detectors, those depending upon the difference in density of the 
 gases are unreliable, because changes of temperature will pro- 
 duce similar results. Aitken's indicator is ingenious, but not 
 much more reliable. Its thermometer is coated with platinum- 
 black and plaster of Paris, and when exposed to fire-damp 
 becomes heated. If the difference of temperature between it 
 and the normal air always bore a comparable ratio to the per- 
 centage of fire-damp contained, it would work well. The special 
 forms of gas-detectors do not serve for illumination. 
 
 The Amount of Air Required for Combustion. In attempt- 
 ing to specify the amount of air required for proper ventilation 
 of a mine, we are treading upon uncertain ground. Within 
 close limits we may ascertain the amount required for the vital 
 chemical purposes of horse, light, and man. A pound of carbon
 
 400 MANUAL OF MINIXG. 
 
 requires for complete combustion 25 Ibs. of oxygen, and pro- 
 duces 3 Ibs. of CO 2 . Hence the ordinary-sized mining-candle 
 burns up n.8 cu. ft. of" air, and discharges 3 cu. ft. of CO 2 
 Eminent medical authorities state that a man consumes about 
 i cu. ft. of air per minute, converting the life-giving principle 
 into 2.1 cu. ft. of CO 2 per hour. The respiration of a horse 
 is about 13 cu. ft. CO 2 per hour. The deflagration of a pound 
 of explosive produces about 2.6 cu. ft. According to Angus 
 Smith, two hewers using a J-lb. candle and 12 ounces of 
 powder produce 25^ cu. ft. CO 2 in a shift. 
 
 The amounts of air sufficient to satisfy the conditions of com- 
 bustion during the generation of the respective amounts of COj, 
 are small, and if the exhalations were instantly removed, the 
 theoretical chemical supply would suffice. But the air in the 
 confined spaces of mine-workings is somewhat stagnant, and 
 the atmosphere is further deteriorated by the exhalations from 
 man and beast. Some of these are not easily detected chemically,, 
 but are more deleterious than CO 2 , which is not the sole test 
 of vitiation. 
 
 The Amount of Air Required for Ventilation. The hot, 
 noisome emanations, the poisonous exhalations, the unconsumed 
 azotic gases, and the exuding pent-up gases from the coal must 
 be rendered comparatively harmless. This requires a large 
 volume of air for their dilution and renewal. Pure dry air con- 
 tains, by volume, 21 per cent of oxygen, O, and 79 per cent of 
 nitrogen, N; and every 1000 cu. ft of it, weighing nearly 81 
 Ibs., contains only about 18.7 Ibs. of the life-supporting con- 
 stituent, the remainder being matter inert in its physiological 
 effects. 
 
 To furnish 2 Ibs. of oxygen for a pound of carbon would 
 require 142.2 cu. ft. of air for combustion alone. To dilute 
 the carbonic acid produced to a wholesome degree, it will require 
 2260 cu. ft. for each pound of carbon. 
 
 Judging by the rough test afforded by the sense of smell, the 
 air of a room ceases to be wholesome when it contains more than 
 6 parts of CO 2 in 10,000. And to preserve the lowest standard
 
 MINE-GASES, 401 
 
 tolerated by sanitarians, i in 10,000, the supply will be propor- 
 tioned as follows: 59 cu. ft. per hour per light; 4585 per horse; 
 9192 per pound of powder; and 1500 per man employed. Com- 
 petent writers vary in this matter, and the statutes of the various 
 States differ in their requirements (55 to 300 cu. ft. per man per 
 minute). But the allowance for a mine cannot be based on the 
 single per-capita element, for it will be seen later that the friction 
 or "drag" of air, in moving through headings and along faces 
 which increase with the developments, diminishes the volume of 
 air actually allowed to move. Moreover, the emission of gas 
 from the strata, proportional to the area exposed and the char- 
 acter of the coal, constitutes another and constant source of pollu- 
 tion. In all preparatory and prospecting work in virgin ground 
 an extra allowance of fresh air is necessary. Cognizance must 
 be taken of this unfailing source to the extent of an hourly allow- 
 ance of 0.3 cu. ft. of air per square foot of working face, in a dry,, 
 dusty, fiery mine. For a non-gaseous seam o.i cu. ft. will suffice. 
 Some property-owners allow also 200 cu. ft. of air per hour for 
 every acre of goaf. For the eruptions from the magazines no 
 provision can be made except vigilance and discipline. 
 
 The Water-gauge (Fig. 154) consists of a U tube whose arms 
 contain water and are provided with measured scales graduated 
 to inches above and below the normal of the water in the 
 columns. If a gauge be inserted through the stopping (Fig. 
 155) separating the bottoms of two ventilating-shafts of the 
 mine, the water will remain at a normal level, if the tempera- 
 tures and tensions of the gas and air in the two shafts be equal;, 
 but if by any cause the tension or temperature be changed in 
 either shaft, the ensuing difference in pressures will be com- 
 municated to the connecting-arms of the water-gauge in such 
 manner that the cool or denser air will force down the water column 
 in the arm on its side of the stopping and elevate the water column 
 on the opposite side. This difference in level, m, is read on the at- 
 tached scale and represents the motive column, M, which is capable 
 of producing motion. If, now, this excessive pressure may be 
 allowed to expend itself in producing motion through the
 
 402 
 
 MANUAL OF MINING. 
 
 workings of the mine, in circulating the air which ultimately 
 is discharged through the lighter column of upcast, the 
 level of the water in the gauge will fall slightly until equilib- 
 rium will be established, when its difference in level will rep- 
 resent the difference in pressure, /, at the bottom of the two 
 shafts on either side of the stopping. The excessive pressure, 
 P, is expended in doing work of two kinds: (i) in overcoming 
 
 FIG. 154. Water-gauge. FlO. 155. The Position of the Water-gauge 
 
 on the Door. 
 
 the friction to the passage of the current of air through the work- 
 ings, from one shaft to the other, and (2) in creating motion. 
 The latter work is measured by the velocity of the outgoing cur- 
 rent, the former is measured by the height of the water in the 
 gauge, Fig. 154, or the manometric depression, m. Its dif- 
 ference in level constituting the water-gauge reading, measures 
 the force which is required to drive the air through the mine. 
 It measures the loss due to friction or the "drag" of the mine. 
 Be the quantity of air large or small, it gives no measure of that 
 volume, but, paradoxical though it may seem, only of the power 
 of the ventilator. 
 
 The Mine Resistance The resistance of the mine is a definite 
 quantity, and bears no relation to the capacity or qualities of the 
 ventilating appliances or methods. The water-gauge, there- 
 fore, which measures this resistance is a "function of the mine," 
 and by it may be determined the relative efficiency of the mine to 
 pass air through its ways. The water-gauge reading in the
 
 MINE-GASES. 403 
 
 majority of mines varies between i inch and 3 inches. Few 
 have a larger resistance. The mine with airways of large cross- 
 sectional area and with a well-distributed current should have a 
 water-gauge reading, or resistance, not exceeding i inch for each 
 hundred thousand cubic feet circulating through it. The mine 
 having a larger ratio of water-gauge reading than this either has 
 airways of insufficient size or is not receiving the current prop- 
 erly regulated. 
 
 The Equivalent Orifice of the Mine. The resistance of the 
 mine to the passage of an air-current is often expressed in the 
 term, the equivalent orifice of the mine. By this is understood 
 the area of a thin orifice which offers a resistance to the passage 
 of a current of the same volume, Q, equal to that which is cir- 
 culating through the mine. The equivalent orifice of the mine, 
 A, bears a certain ratio to the quantity, Q, and to the water- 
 gauge reading, m, which is variously expressed by different 
 authors, the general formula being 
 
 the value for C varying between 0.00037 an d 0.00066. That 
 usually taken in calculations is 0.0004. 
 
 The equivalent orifice of most mines varies between 10 and 
 100 sq. ft. Inasmuch as water is about 833 times as dense as an 
 equal volume of air, the column depressed in the water-gauge 
 corresponds to a height of 833^ inches of air at 62 F. and 30 
 inches barometer; the head, M, of air measured in feet, to which 
 the manometric depression is due, is 
 
 M = 6g.4m. 
 
 The value for the corresponding differential pressure, P t 
 in pounds per square foot, is 
 
 EXAMPLE. How does the efficiency of a mine carrying 178,000 cubic feet 
 of air compare with another having 284,800 cubic feet per minute? The 
 water-gauge readings are 2.7 inches and 3.4 inches, respectively. 
 
 0.0004X17,8000 
 
 a= - j= - =43.36 square feet; 
 V 2 . 7
 
 404 
 
 MANUAL OF MINING. 
 
 0.0004X284,800 _, 
 a= -7= - =61.78 square feet. 
 
 Their relative efficiencies are 0.70184:1, respectively. 
 
 The following table illustrates the rate of decrease of water- 
 gauge in two mines, respectively, carrying 30,000 and 100,000 
 cu. ft. of air per minute for various equivalent mine orifices. 
 
 Equivalent 
 
 Water Gauge, Inches. 
 
 Equivalent 
 
 Water Gauge, Inches. 
 
 Orifice. 
 
 
 
 Orifice, 
 
 
 
 Square 
 Feet. 
 
 30,000 
 Cubic Feet 
 
 100,000 
 
 Cubic Feet 
 
 Square 
 Feet. 
 
 30,000 
 Cubic Feet 
 
 100,000 
 
 Cubic Feet 
 
 
 Air. 
 
 Air. 
 
 
 Air. 
 
 Air. 
 
 5 
 
 56 
 
 
 25 
 
 0.23 
 
 2-56 
 
 10 
 
 44 
 
 16.00 
 
 3 
 
 0.16 
 
 I. 7 8 
 
 15 
 
 .64 
 
 7.11 
 
 40 
 
 
 I. 00 
 
 18 
 
 45 
 
 5.00 
 
 5 
 
 
 
 0.64 
 
 20 
 
 36 
 
 4.00 
 
 100 
 
 
 
 o. 16 
 
 REFERENCES. 
 
 Behavior of Mine Gases, M. & M., Vol. XXII, 229; Air in Coal Mines, 
 Prof. Clowes, Coll. Guard., Jan. 1896, 222; Respirability of Air in which a 
 Candle Flame is Extinguished, Frank Clowes, E. & M. Jour., LXI, 515; Mine 
 Ventilation, W. J. Mollison, M. & M., Feb. 1904; Air: Composition of the 
 Air in Mines, Coll. Guard., Nov. 8, 1895, 890. 
 
 Gases: Some Dangerous Gases, Coll. Guard., Jan. i, 1896, 16; Compo- 
 sition of a Mine Atmosphere in Russia, Trans. M. & M. E., XLV, Part V, 
 no; Temperature and Humidity Collieries, Coll Guard., Oct. 25, 1875, 779 J 
 Continuous Sampling of Mine Air, Coll. Guard., 591; Mar. 27, 1896; Some 
 Mine Gases, A. Heymann, Jour Chem. & Met. Soc. of S. Africa, Aug. 1903 
 and Sept. 1903; When is Fire Damp Dangerous in a Coal Mine? Coll. Guard., 
 
 Oct. 10, 1002. 
 
 Fire-damp Accumulations, Coll. Mgr., Mar. 1893, 58; Fire-damp, Com- 
 position of, Coll. Guard., Dec. 1896, 1170; On Experiments showing the 
 Pressure of Gas in the Solid Coal, Lindsay Wood, N. E. I., XXX, 163; Fire- 
 damp Apparatus for Experimenting with, Coll. Guard., Feb. 14, 1896, 317; 
 Detection of Small Quantities of Gas in Air Dr. Clowes, Collieiy Mgr., Jan. 
 18, 1895, 17. 
 
 Indicators Fire Damp, Trans. M. & M. Eng., Vol. XLV, Part V, 106; 
 Shaw Gas Tester, Mine Insp., Pa., 1889, 372; The Shaw Gas Tester, M. & 
 M., Vol. XX, 85; Gas Detectors, M. & M., Vol. XX, 01. 
 
 How to Read a Water Gauge, M. & M., Vol. XXII, 85; Setting Pipe- 
 valves, Engineer, June i, 1904, 33.
 
 CHAPTER XIII. 
 
 METHODS OF VENTILATION. 
 
 The Ventilation System. To obtain circulation through the 
 mine, a conduit must be furnished by which the warmer and 
 lighter air may ascend to be supplanted by cold or compressed 
 air entering by a different compartment; and to maintain a 
 constant air-current throughout the workings, both inlet and 
 outlet must be afforded for the air by means of two separate 
 entries or by partitions in the one shaft. 
 
 The Ventilation of Single Entries. Shafts, in process of 
 sinking, or a mine having but a single entry, may discharge 
 their vitiated air through the wooden air-tight box-pipe pro- 
 vided for the purpose, or, if there is small liability of corrosion, 
 through a galvanized- iron pipe, the remainder of the entry furnish- 
 ing the inlet. Because of the wide difference in the areas of 
 the two airways so provided, the ventilation is not likely to be 
 good, and it is far better to divide the main tunnel or shaft or 
 mine-working into two compartments of nearly equal area, one 
 of which will serve as an outgoing conduit. 
 
 From the fact that the current in a single-entry mine is con- 
 tinually interrupted by the other uses to which the compartment 
 is put, and that there is a liability to injury of the partition, box, 
 or pipe, this plan is objectionable when a large volume of air 
 is required, because the safety of a great number of men is depend- 
 ent upon this airway for escape. The wind, moreover, disturbs 
 the ventilating current; the movement of cars, cages, and rock 
 or coal in chutes is also irregular in its influence upon it; and 
 the unusual heat from underground steam-pipes, engines, etc., 
 
 405
 
 406 MANUAL OF MINING. 
 
 sets up counter-currents, though any of the causes mentioned 
 may occasionally have a beneficial effect. Thus a double entry 
 to the mine becomes not only precautionary, but also impera- 
 tive; and as the depth and extent of workings increase, the 
 insufficiency of a single entry becomes more and more manifest. 
 Even metalliferous mines should be provided with a double entry, 
 for the numerous caves that have occurred, penning in dozens 
 of men without chance of escape unless the rescuers can reach 
 them before suffocation ensues, and the fires that frequently 
 cut off the employees from the outlet and suffocate them before 
 extinguishment is effected, are sufficient arguments in favor of 
 double entry, even if the necessities for better air do not appeal 
 to the operators. 
 
 Ventilation by Double Entries. All collieries have two out- 
 lets, separated by a safe distance of unbroken rock. The upcast, 
 advisably, should terminate in a large chimney, high enough 
 that its draught be not influenced by changes of wind or the 
 surrounding buildings. The location of the two entries, in 
 reference to each other, varies within wide limits. One plan 
 consists in having them near together, thus concentrating the 
 plant. Both airways being carried with the development, the 
 current passes through to the extreme end of one and returns by 
 the other. Then as the work deepens, each lower lift is connected 
 with the airways of the upper lift, and receives ventilation with 
 its advance. The other plan is the "diagonal system," the 
 shafts being at the extremities of the workings. While this is 
 well enough for the long-wall method, the ventilation must 
 meanwhile suffer until the connection has been made. 
 
 Two compartments in a single entry may be easily obtained 
 in coal furnishing sufficient rock from the roof or from partings 
 by driving a wide gallery and walling it up centrally with the 
 waste; but if there is not rock enough for this, two entries are 
 carried, with the usual pillar between them, having connecting 
 "throughs" at intervals of less than 100 feet, each being closed 
 as fast as the next one is completed. To ventilate that part 
 of each entry between the last connection of the entries and its
 
 METHODS OF VENTILATION. 407 
 
 face, it is subdivided by a canvas brattice fastened at the near 
 side of the "through" and leading up to the work. On either 
 side of this the current flows. The faces may be connected by 
 pipes through the door closing the intake entry without inter- 
 fering with the haulage. .The practice of relying upon diffusion 
 to do the work of ventilation is pernicious. These remarks also 
 hold true regarding the "throughs" connecting the rooms in 
 pillar and stall workings, where diffusion is usually relied upon 
 for the needful amount of oxygen. 
 
 Planning Airways. A large number of the coal-mines 
 depend for their ventilation solely upon natural means, and this 
 may suffice in small mines. But as the workings are extended, 
 the numerous connections which are necessary for development 
 or convenience of handling the materials may be planned to 
 serve also for ventilating ways without additional cost. 
 
 In planning the direction of gangways and of rooms in coal- 
 mines, usually the question of haulage is of the first considera- 
 tion, unless it be that the "cleats" are so pronounced as to de- 
 termine the direction of work. At the same time due attention 
 must be given to the matter of ventilation, that the requisite 
 amount of air i>e given each working-room, and that too many 
 men be not dependent upon the same air-current circulating 
 through the mine; whenever the mining conditions require a 
 subdivision of the incoming air-current into small currents, each 
 being distributed to its own district and group of men and each 
 separately discharged, it becomes evident that the ventilation 
 of such gaseous mines must receive special attention, not only 
 as to the direction in which the airways are driven and their 
 cross-sectional dimensions, but also as to the means of produc- 
 ing the supply of air. In such cases the fresh air should be 
 carried, if possible, to the deepest point in the mine, whence an 
 ascending current may be conveyed through the workings until 
 it is returned to the surface. Especially is this advisable in 
 steep coal-seams carrying fire-damp. 
 
 The ventilation must be so arranged that as many inde- 
 pendent ventilation districts as possible be provided with sepa-
 
 4 o8 MANUAL OF MINING. 
 
 rate air-currents; and especially must each lift of workings be 
 supplied by the shortest way with the necessary quantity of 
 fresh air, while within the separate lifts of workings the air- 
 current must always be ascending except in cases in which 
 the descending air-currents are not used for any further ventila- 
 tion purpose, or when, in certain well-ventilated working places, 
 excessive thrust of the measures renders very difficult the keeping 
 up of special return airways. 
 
 In metal-mines, where the development is of slower growth, 
 the rock hard, and a comparatively few men are at work, the 
 amount of air required is small, either for inhalation or for the 
 dilution of the gases developed therein; hence a single shaft 
 with two compartments may suffice, the circulation being left 
 to natural sources. This, however, will be inadequate when 
 the shafts and workings reach a depth of several hundred feet, 
 in which case other means must be employed. The use of com- 
 pressed air for drills, pumps, etc., may supply the deficiency of 
 pure air which natural ventilation may fail to furnish, yet a fan, 
 exhausting the air from one outlet or forcing the air into the 
 other, seems imperative with extensive workings. 
 
 The Underground Temperature. Below the depth where 
 atmospheric changes have no influence, the temperature of the 
 undisturbed rock increases with every increase in depth. The 
 depth at which the temperature of the ground will be found to be 
 invariable and equal to the natural temperature of the locality 
 is about 50 feet below the surface. Beyond this it is an observed 
 fact that in all artificial openings the temperature of the rocks 
 increases for at least a moderate depth, within which the mine 
 operator is concerned, at the rate of about i F. for every 68 feet 
 of depth. This increment is not constant for all localities, nor 
 indeed for the same mine, but generally it may be said that as 
 we go down the temperature of the mine increases more or less 
 uniformly. This increased heat is often a great drawback to 
 mining, and will ultimately limit it apart from the lesser mechan- 
 ical difficulties. As to what would constitute the limiting depth 
 to which mining may be prosecuted, it can but be said that at
 
 METHODS OF VENTILATION. 49 
 
 present several mines, with the exception of the Comstock and 
 those which are in ore-bearing districts feeling the effects of 
 solfataric action, are working at over 4000 feet. Regarding the 
 exceptions stated, it is certain that unless some means be discov- 
 ered for rendering their lower levels habitable, the limit of mining 
 depth is soon reached. It is stated that a 28oo-foot level of the 
 Yellow Jacket Mines has been abandoned because of the exces- 
 sive temperature, in many rooms of which the miner is com- 
 pelled to return to a cooling station after laboring only twenty 
 minutes. 
 
 An interesting report bearing upon this question of the rate of 
 increase of temperature with the depth of subterranean explora- 
 tions, made by a sub-committee of the Royal Commission on 
 Coal, reaches the following conclusions: That the limit of depth 
 to which mining is possible depends upon human endurance of 
 high temperature, and upon the extent to which it would be possi- 
 ble to reduce the temperature of the air which comes in contact 
 with the heated rocks ; that there is no limit caused by considera- 
 tions of a mechanical nature as to the size of rope for hoisting- 
 engines, nor by any consideration of the enhanced expenditure for 
 shaft sinking, for haulage, or for pumping. Regarding the latter, 
 the experts testifying before them demonstrated that water is 
 seldom, if ever, met with in large quantities at great depths in 
 mines. It therefore appeared that this increase in temperature 
 is the only element needing consideration regarding the limits of 
 prospective sinkings or workings. 
 
 A summary of the results of temperature observations made 
 under the direction of the British Commission Committee shows 
 the mean increase of temperature per foot to be 0.01563, or i F. 
 in 64 feet, the extremes being 0.0077 ' m tne Bootle water- works 
 "bore-holes, and 0.025 in the Carrickfergus shaft. At the Adel- 
 bert shaft, Prussia, observations five times a month, in different 
 levels, for a year, could deduce no regular law of increase; at the 
 3oth level, 3200 feet, the temperature was 98 F. 
 
 Natural Ventilation. The temperature of the air inside the 
 mine differs from that outside. The mass which is the warmer
 
 410 MANUAL OF MINING. 
 
 will rise, enabling the colder mass to fall. Circulation is estab- 
 lished so long as this difference exists. 
 So that, if two openings be made and 
 connected below, a current will be estab- 
 lished down the lower and shorter open- 
 ing in winter, and up the same during 
 the summer, as the arrows (Fig. 156) 
 marked S indicate. In winter the 
 
 FIG. i 5 6.-The Circulation of direction of the current follows that of 
 the Currents in Summer and the arrows, W. The amount of air thus 
 Winter. 
 
 set into circulation by the changes of the 
 
 exterior temperature will depend upon the relative difference of 
 temperature between the mine and surface, and also upon the 
 depths of the shafts. When these differences are slight it is not 
 easy to predict the direction which the current will take. As, for 
 example, in the fall and spring it will fluctuate from one to the 
 other. When, however, these differences are great, a current 
 will be set up which tends to continue in the same direction so 
 long as these differences remain. Thus in summer the current 
 will follow (Fig. 156) the arrows, 5; in the fall little or no current 
 will be set up; in the winter the current will reverse and follow 
 the arrows, W] in the spring the conditions are again nearly 
 balanced, and little current will flow. When the shaft attains a 
 depth of 800 feet, the subterranean air is always hotter and 
 lighter than the surface air at any season; and unless the two 
 outlets have a great difference in elevation, an uninterrupted 
 current will continue, without fear of reversal, down the lower 
 and shorter opening. 
 
 While this method may be universally practised under 
 favorable conditions in metal-mines, it is evident that in col- 
 lieries one danger arises from the reversal of current, for at one 
 time the current, following the arrows marked S, carries the air 
 through the gangways, whence it is distributed among the work- 
 rooms, to be returned to the surface by way of the longer and 
 deeper shaft; but during the other season the air may follow 
 the arrows marked W, thus entering the working places first, and
 
 METHODS OF VENTILATION. 411 
 
 departing thence through the gangways, making its exit by the 
 lower or shorter shaft. If, now, there be a number of abandoned 
 rooms or goaves connected with the working-rooms, it is mani- 
 fest that in the latter season the air must pass through them first 
 before reaching the men at work, and thus carry noxious gases 
 with the current to spread calamity by explosion or fire. Again, 
 the fact that no air will circulate during the vernal seasons would 
 render the provision for supplementary means of ventilation im- 
 perative. Air-currents which have served for ventilating pre- 
 paratory workings or prospecting drifts in the virgin seam never 
 should pass over stalls or working places where men are engaged, 
 on its way to the air-level. 
 
 The Tension of the Atmosphere. The air possesses, in 
 common with all other gases, in consequence of the repulsion 
 between its molecules, a tendency to expand into a greater space. 
 This indefinite expansion, by reason of which every gaseous 
 fluid, not restricted by an extraneous force, continues to expand 
 to the tenuity of interstellar space, results in the creation of an 
 air-current whenever by an increase of temperature or a diminu- 
 tion of pressure the given mass of air expands in opposition to the 
 attraction of the earth and rises into the upper strata. This 
 upward flow will continue so long as the gas expands until the 
 resistance encountered by it is equal to, or greater than, the 
 repulsion among its molecules. It is this readiness with which 
 gases tend to adjust themselves to the varying conditions of 
 temperature and pressure that plays so important a part in mine 
 ventilation. The tension of a gas increases with its compression, 
 and the density of a given mass of air is proportional to its tension. 
 
 The Weight of Air. This may be calculated for different 
 temperatures and barometric pressures by the following formulae: 
 
 The volumes, M, assumed by a given weight of a gas are 
 inversely as the corresponding pressures per unit of surface, 
 provided the temperature remains constant : 
 
 u:u'::p':p. 
 If the temperatures change while the pressures are constant,
 
 412 MANUAL OF MINING. 
 
 the volumes, reduced to absolute zero ( 461 F.), will be found 
 to vary proportionally. 
 
 / and T being, respectively, for up and u' /, Fahrenheit readings. 
 The weight of a cubic foot of air at a temperature / and a 
 barometric pressure B, in inches of mercury, is obtained by the 
 following formula, and at a temperature T, is W, expressed as 
 follows: 
 
 1.32535 1*253 
 
 A table of weights of air for various temperatures and a con- 
 stant pressure, B, will be found in Chapter IX. 
 
 The Production of Draught. The energy in a mass of air 
 compressed to a certain degree may be measured by the work 
 restored by it in expanding, and this energy may be converted 
 into motion producing a current, or it may result in a pressure 
 when that tendency to motion is resisted, or when the motion 
 is suddenly arrested. 
 
 That portion of the energy stored up in the air which is 
 expended during its expansion in dynamic effect causes a 
 "wind" or "draught," the velocity of which depends upon 
 the difference in tension. 
 
 Whether the difference in tensions is produced by change of 
 temperature, or of pressure, the velocity acquired by any gas is 
 
 in which H is the head due to the difference between the ten- 
 sions, or densities, of the initial state of the cool or compressed 
 gas and the final state of the hot or expanded gas. Atmos- 
 pheric air, at a barometric pressure of 29.92 inches, at a tem- 
 perature of 32 F., when flowing into a vacuum, attains a velocity, 
 in feet per second, which is equal to
 
 METHODS OF VENTILATION. 413 
 
 V =\ / 2# =8.02^ 
 
 =8.02 J I4 ' 7Xl f = 10758.9 feet. 
 
 ^ O.OOIIO 
 
 The total difference of pressures per square foot is represented 
 by P in pounds, and the weight of a cubic foot of the warmer 
 or attenuated gas by W. 
 
 So, too, the velocity with which compressed air or steam 
 escaping freely from a pipe or other reservoir of the same may 
 be ascertained, the value to be supplied for H, the head to which 
 the velocity will be due being equal to the pressure in pounds 
 per square inch under which the gas exists, multiplied by 144 
 and divided by the weight in pounds per cubic foot of the ex- 
 hausted fluid. 
 
 The Motive Column. This is the head producing motion. 
 When, however, two masses of air of equal height but of 
 different tensions, p and />', are exerting a pressure upon one 
 another through a connecting conduit, the resulting difference 
 in pressure per unit of area of base measures the motive force, 
 in which case P is the total difference of aerostatic pressure in 
 pounds per horizontal square foot of sectional area of base, and 
 W the weight per cubic foot of the rising column of air. 
 
 If, then, a column of air at /F., D feet high, with a base 
 of one square foot, be heated to T F., its new height would be 
 greater by some quantity, which we may call M . If two such 
 columns be connected, being of the same depth but of different 
 temperatures, / and T respectively, the latter column would be 
 lighter than that at / by a quantity D(w W}\ and so long as 
 this difference in temperature is maintained, this difference of 
 pressure, which we may represent by P, ensues, by reason of 
 which the hot column of air would be driven upward, producing 
 a draught with a velocity, F, due to the aerostatic head, M. To 
 hold this force, P, in equilibrium would require a resistance 
 D(w W) pounds per square foot; or the pressure of an addi-
 
 414 MANUAL OF MINING. 
 
 tional column of warm air weighing W pounds per cubic foot, 
 of a height of M feet, 
 
 D(w-W) P T-t 
 
 M- TT7 TIT *^ 
 
 w w 4 6i -H' 
 
 This quantity M is known as the motive column to which 
 is due the velocity of the flow of air, and if no resistance is offered 
 to it, motion will take place. It may be represented by OT (Fig. 
 156), which equalizes the pressure of the unequally heated columns 
 of air below the level of the line GO. 
 
 When two such shafts are of unequal depth, as at O and 
 M, Fig. 156, and have equal exterior and interior tempera- 
 tures, a rarefaction of the air in either one of them not affect- 
 ing the other would result in a diminished pressure upon the 
 bottom, just as is obtained by a difference in temperatures; 
 a rising current is established therein, with a velocity depend- 
 
 p 
 ent upon the ratio ^, in which P is the difference in the weights 
 
 of the two shaft columns of a base one horizontal square foot 
 in area and a height D; and W is the weight of a cubic foot of 
 the rarefied air. 
 
 For the purpose of mine ventilation there will be required 
 a motive column much larger than that here obtained, because 
 of the enormous friction of the air in rubbing along the rough 
 surface of the workings, turning sharp corners, and squeezing 
 through small openings. The resistance due to this cause amounts 
 often to as much as 90 per cent of the power. In other words 
 only one tenth of the theoretical motive column becomes effec- 
 tive in producing a current, and the actual velocity of the air- 
 current, v, does not exceed one third of the theoretical velocity, 
 V , due to the head, If. 
 
 The principle upon which chimney draughts for boiler or 
 other heating apparatus depends is also similar to that here 
 described, excluding, of course, frictional allowance. In chim- 
 neys for boiler-furnace draughts the fire burns best when W is
 
 METHODS OF VENTILATION. 415 
 
 0.530, and the height of the manometric column in the chim- 
 ney is about one half an inch of water. 
 
 It is evident, therefore, that the height of a motive column 
 depends upon the difference in temperatures or a difference in 
 tensions, or both, of the gaseous mixture contained in the two 
 shafts or entries to the mine. A measure for this motive column 
 may be had in feet of head of pressure per square foot of area of 
 the base, or in the number of inches of a water column in the 
 manometer which corresponds to this weight. Insomuch as a 
 column of water i inch high and a square foot in area of base 
 weighs 5.184 Ibs., the height, m, of a water-gauge column which 
 will balance the pressure P is equal to P-v-5.i84. 
 
 Let M be the head corresponding to the motive column, 
 
 v* 
 v the velocity of flow of the upcast air per second; then is 
 
 the effective velocity head of the issuing air; and if W is the 
 weight of a cubic foot of the warm or attenuated rising air, the 
 theoretical energy of the moving air per second is WM , and the 
 
 v 2 
 effective or actual energy is W , or 0.015 53 WV 8 . That por- 
 
 tion of the energy which is consumed in overcoming the friction 
 of the mine is therefore W(M 0.015 53^ 2 ). It is this lost energy 
 which is measured by the water-gauge. As the mine resistances 
 are reduced, so the water-gauge reading is reduced, and the 
 efficiency of the mine increases, permitting a greater actual 
 return from the expenditure of the same potential force. 
 
 The height, w, of the water column being measured in inches, 
 the number of horse-powers, H, necessary to produce a ven- 
 tilating current of Q cubic feet per minute is known by the fol- 
 lowin formula : 
 
 The indication, therefore, which the water-gauge reading 
 gives of the ventilating force is evident in the above formula 
 that for a given quantity of air, Q, in circulation, the horse-power 
 necessary to produce ventilation increases with the resistance of 
 the mine.
 
 41 6 MANUAL OF MINING. 
 
 The Systems of Producing a Ventilating Current. The sev- 
 eral methods of accelerating the natural ventilation and dis- 
 tributing air properly throughout the mine contemplate some 
 system either of decreasing the tension of the mine air to enable 
 the return current to ascend to the surface, or of increasing its 
 tension by the use of a compressor to force atmospheric air into 
 the mine. The several means by which these results are attained 
 may be, first, a furnace built at the bottom of the outlet shaft, 
 or a fan; or, second, a blowing, propelling, or air-compressing 
 fan at the mouth of the inlet shaft. By either of these methods 
 a different state of tension is produced in the two shafts con- 
 nected below, and, in the effort to establish equilibrium, the air is 
 set in motion, a draught is created, and a current is established 
 that flows through the air-courses at a velocity dependent upon 
 the head due to the difference in pressures, as has been seen on 
 the preceding page. 
 
 Furnaces are employed for increasing the temperature and 
 are constructed in such manner as to be remote from direct 
 contact with the coal, yet in close proximity to the shaft which 
 constitutes the outlet for the mine air, and in a gallery through 
 which circulates air from the workings. The pit selected for 
 the outlet should be that one which would naturally carry the 
 flow in winter. The furnace is simply a fireplace, walled and 
 roofed by a fire-brick or common-brick arch (Figs. 157 and 158). 
 When special care is taken, a second wall is built outside and 
 over, with an air-space between, to isolate it from the coal and 
 prevent fire. If the roof is wet, a double arch must surmount 
 the furnace, as otherwise the steam generated will burst the arch. 
 If the mine is fiery, or considerable dust is floating, care must 
 be taken that the gases are well diffused, or else the current must 
 not be brought into close proximity with the fire. In such cases 
 the current is split, a small portion being heated over the .fire, 
 the remainder passing through a "dumb-channel," entering the 
 upcast 50 feet or so above. A still safer plan passes all of the 
 fiery current through the channel, and feeds the furnace by a 
 split current of fresh air direct from the intake.
 
 METHODS OF VENTILATION. 
 
 417 
 
 The size of the grate depends upon the work to be done. 
 Its bars are 3 feet from the floor, slanting upward toward the 
 
 FIG. 157. Longitudinal Section of a Furnace. 
 
 shaft i to 6, distance to the roof 4 or 5 feet. The width, wall to 
 wall, is 6 feet and its length from 4 to 12 feet, according to the 
 volume of air to be moved, which is about 1500 cu. ft. per square 
 foot of fire surface on a properly constructed furnace. 
 
 FIG. 158. Cross-section of a Furnace. 
 
 An ordinary furnace of 34 sq. ft. heating-surface, costing 
 $130, will heat a column of air such as will furnish 29,000 cu. ft. 
 per minute. A large number of furnaces 10X12 furnish 200,000 
 cu. ft. The cross-sectional area must be 50 per cent greater 
 than the upcast airway, and the shape capable of regulation by
 
 41 8 MANUAL OF' MINING. 
 
 double-sliding iron doors, to produce varying degrees of con- 
 traction and of combustion. The fire is spread over its entire 
 width, and over only as much of its length as is necessary to fur- 
 nish an adequate motive column, at a temperature of 140 to 
 160 F. Emergencies, as low barometer and high thermometer, 
 and the cleaning of the grates, require other and more heating- 
 surface. The coal consumed is 2 to 5 tons per day, spread thin 
 and evenly over the bars, and fed from both ends, on a long fur- 
 nace. This rate is 40 to 70 Ibs. per hourly H.P. of work done 
 on the air. Attendance, etc., is $5 per day. 
 
 Q being the quantity of air in cubic feet per minute, W the 
 weight of a cubic foot of return air, T being the temperature, 
 F., of the up-cast air, and t that of the air in the return airway, 
 the number of pounds of coal consumed by the furnace per hour 
 is 
 
 x= 
 
 The area, F, of the grate-surface in square feet is about one 
 tenth of the hourly coal consumption, in pounds, and its rela- 
 tion to the depth, D, of the furnace below the surface is known 
 by the expression 
 
 FQP= i,7i6,ooo\/Z>, 
 
 P being the manometric depression in pounds per square foot, 
 and Q the volume of air per minute. 
 
 The Limits of Furnace Ventilation. A method so simple 
 and cheap in construction and easy of management presents 
 advantages which have long commended it to mine operators; 
 nevertheless the difficulties with its use, the dangers which attend 
 the exposing of an open fire in gaseous districts without the 
 possibility of introducing a safeguard, the numerous calamities 
 traced to the furnace which has fired either the solid coal sur- 
 rounding it, the gases in the return air, the timbers of the shaft, 
 or even the surface plant, and its lack of economy in shallow 
 pits, were soon made manifest. The atmospheric changes of 
 the seasons reduce its efficiency, a decrease in the barometric
 
 METHODS OF VENTILATION. 419 
 
 pressure and an increase in the surface atmosphere reduced 
 the action of the furnace, and notwithstanding its great superi- 
 ority over many other mechanical appliances, it has gradually 
 been supplanted by fans. The power of the furnace increases 
 arithmetically with the temperature, and that with the amount 
 of fuel burned. The quantity of coal that can be consumed 
 upon a given area is limited. The resistance of the mine increases 
 on the other hand geometrically with the square of the velocity 
 of the current, and it is therefore manifest that between the 
 several conditions the furnace limit is soon reached. Many 
 furnaces may be cited supplying to the mine over 20x3,000 cu. ft. 
 of air per minute; and enormous as they are, their cost is very 
 little less than that of a modern fan of large size; but when we 
 contemplate the huge pile of coal thus consumed for the produc- 
 tion of the current, we are forced to the conclusion that efficient 
 furnace ventilation is a luxury which the coal trade cannot long 
 endure. Perhaps as the depth of the collieries increases to 
 about 2000 feet the furnace may be reinstated. 
 
 With the atmospheric air at 62 F., and the furnace-heated 
 air at 132 F., the water-gauge depressions, m, produced at various 
 depths of furnace are as follows: 
 
 D. 
 
 m. 
 
 D. 
 
 m. 
 
 50 feet, 
 
 0.086 inch. 
 
 1000 feet, 
 
 1-735 
 
 400 " 
 
 0.694 " 
 
 2OOO " 
 
 3-471 
 
 700 " 
 
 1.215 " 
 
 4000 " 
 
 6-943 
 
 These are in accordance with the formula for estimating the 
 manometric depression: 
 
 W I T-t \ 
 
 - 
 
 EXAMPLE. A colliery has two shafts 1000 feet deep, 12 feet in diameter; 
 temperature in the downcast is 60 F.; barometric pressure is 30 inches. 
 150,000 cubic feet of air are supplied per minute by a furnace. Required 
 the temperature of the upcast and the horse-power necessary to produce the 
 ventilation, the mine being supposed to show a water-gauge resistance of 
 2 inches of water. 520 F. and 163.5 H.P.
 
 420 MANUAL OF MINING. 
 
 Assume the coefficient of friction for the smooth shafts to be as great as 
 that of the rough mine galleries; then 
 
 O.OOOOOOO2I7X 1 000X37. 7X(i5o,ooo) 2 
 />=- (113)' =I2 ' 8 - 
 
 Each shaft therefore offers a resistance equal to 12.8 Ibs. per square foot. 
 The total resistance then is 25.6+10.368=35.968. The work done in ven- 
 tilating is 150,000X35,968=5,395,500 ft.-lbs., or 163.5 horse-power. 
 The temperature of the upcast shaft is 
 
 T-6o 
 
 T . --- - . 
 39-759 521 
 
 Or, by another method: A cubic foot of air at 60 F. and 30 barometer weighs 
 0.0766 Ib. 150,000 cubic feet of the circulating air weigh 11,490 Ibs. Since 
 the furnace is performing 5,395,500 ft.-lbs. of work upon 11,490 Ibs. of air, 
 the height through which it is moved is 470 feet in one minute. Then M is 
 470 ft., /=6oF., and D=iooo ft. 
 
 From this it is seen that the temperature of the upcast air necessary to force 
 150,000 cubic feet of air through the mine is dangerously high. The furnace 
 must be replaced by the exhaust-fan, or the frictional resistances must be 
 reduced by enlarging the entry-ways. 
 
 What should be the size of the airway shafts in the above case, that the 
 upcast air be not hotter than 190 F. ? By substitution above, M is found to 
 be 200 ft.; the work is then 11,490X200=2,300,000 ft.-lbs. (70 H.P.). This 
 value requires that p should not exceed 15.3 Ibs., which limits the shaft's 
 resistances to 4.932 Ibs., or 2.466 Ibs. each. In order to obtain so low a 
 friction the areas are enlarged to a radius of 16.66 feet. 
 
 pa*=jlmq\ or p(xr>y=jl( 2 xr}q\ 
 
 Types of Fans. Of the various mechanical ventilators tried 
 in mines, fans remain our main reliance at the present time. 
 As the furnace has in the past supplanted various mechanical 
 devices in the form of pumps and trompes, so fans built on various 
 principles have succeeded the furnace and the steam-jet. There 
 are two classes or types of fans: (i) blowers, either rotary or 
 reciprocating, and (2) fans, propeller or centrifugal. Those 
 of one type sweep out a fixed volume of air at each revolution
 
 METHODS OF VENTILATION. 42* 
 
 and are known as the definite-volume exhausters, under which, 
 head come the Root, Baker, Lemielle, Cooke, and Fabry. In. 
 the other class, acting centrifugally upon the air, we have a 
 simple revolving wheel always working in one direction, pro- 
 ducing by its rotation a pressure or a rarefaction the degree of 
 which depends upon its speed. Of these we have the Guibal, 
 Waddle, Walker, and Schiele ventilators. 
 
 The Trompe is a simple application of the injector princi- 
 ple, water falling in the cylinder and carrying with it air, creates 
 a small intake draught. The volume of air, compared with the 
 quantity of water used, is so insignificant that, unless an espe- 
 cially favorable means be provided for carrying off the water, the 
 ventilation is too expensive to be continued except as a temporary 
 expedient. 
 
 Pressure Blowers, either rotary or reciprocating in their 
 action, are of general use in America, being represented by the 
 Root, Baker, and Champion on the one hand, and air-compressor 
 and other reciprocators on the other. The blower forces the 
 air through the intake compartment of the mine, which dis- 
 charges it at the upcast. These blowers or force-fans are much 
 in vogue for small workings and as expedients in furnishing 
 a separate ventilation for stopes and drifts; but few are employed 
 in coal-mines to produce the total ventilation there required. 
 In metal-mines, however, they are largely depended upon, though 
 they supply a pressure higher than that ordinarily required to 
 overcome the resistance of the mine. They produce air by 
 reason of their high speed at a pressure often attaining 10 Ibs. 
 per square inch, whereas a mine requires an initial pressure only 
 sufficient to overcome its resistance, which is rarely greater than 
 10 Ibs. per square foot of base. The blower is a small radial 
 wheel revolving freely in a casing and nearly touching its sides. 
 By a central opening on either side the air is admitted to te 
 acted on and set into rotary motion. These blowers may be 
 had in sizes capable of furnishing as much as 16,000 cu. ft. per 
 minute, requiring from i to 15 horse-power for their operation. 
 Some of the blowers are capable of a ready alteration from a
 
 422 MANUAL OF MINING. 
 
 t>lower to an exhaust, or the reverse, which fact recommends 
 them particularly for wide shafts which are liable to freeze during 
 winter. This is particularly advantageous in metal-mines, 
 ^where it makes very little difference which way the current moves. 
 En collieries, however, as has been seen, this is not feasible. 
 
 The Root Blower or force-fan consists of two interlocking 
 timpellers revolving side by side in very close connection, without 
 actually touching one another or the enclosing case. They are 
 made of cast iron accurately bored and dressed to a true surface, 
 so that, while practically no air escapes, there is also no internal 
 wear. At each revolution a definite volume of air enters, is 
 enclosed, and discharged either at the top, the bottom, or the 
 side. They are driven by a pair of external gears at a speed 
 ordinarily of from 250 to 500 revolutions per minute. The 
 extremities of the revolving arms of the impeller section are of 
 an acorn shape, or their surfaces are arcs either of true circles 
 . or of cycloids. 
 
 The Fabry Blower, which resembles the Root, is much used 
 in the north of France and Belgium. Two fans, each having 
 three broad blades arranged radially, are hung in a chamber. 
 They revolve with equal velocities in opposite directions, the 
 blades coming in contact, isolating a quantity of air, and expel- 
 ling it into the atmosphere. The success of this blower is attrib- 
 uted to the fact that there are no joints in it. 
 
 The Baker Rotary Force-fan has inside of its casing three 
 -drums, each being an independent casting turning truly and 
 balanced perfectly to insure a steady motion. The upper drum, 
 which receives the power from the engine, does all the work 
 of blowing, while the two lower drums serve as valves to prevent 
 the air from escaping. 
 
 Cooke's Fan is a positive machine. An eccentric drum 
 revolves inside of a 1 2-foot circular case, very close to which is 
 held a swinging shutter that cuts off the entering current from 
 the discharge. The inlet and outlet portion occupies 235 of a 
 -revolution. At the Lofthouse iron-mines are seen two of these 
 side by side, the drums being placed opposite each other on the
 
 METHODS OF VENTILATION. 42$ 
 
 shaft, so that the revolving mass is balanced, the discharge 
 equalized, and the efficiency raised. 
 
 The Lemielle Blower is a species of air-pump, complicated! 
 and leaky, producing large volumes under great rarefaction. 
 It consists of a vertical cylinder, within which a second revolves, 
 eccentrically; on this latter are two or more vanes, which in* 
 one part of the revolution lie close to the shutter, and in another 
 open and expel the air. 
 
 The reciprocating blowers have been displaced almost, 
 entirely by the rotary blowers, either class being capable of a 
 reversal of rotation to force air into or exhaust air from the mine, 
 as desired. The power required to drive the force-fan depends- 
 upon the volume and pressure of air exhausted or discharged;, 
 but the rule usually followed for computing the net power in a 
 given volume at different pressures is 'to multiply the number of 
 cubic feet delivered per minute by the pressure in pounds per 
 square foot at the blower, and the product by 0.00003; the quo- 
 tient will give the net horse-power required to drive the fan. 
 
 The Centrifugal Fans, which are used almost exclusively 
 in America, may be divided into two great classes: (i) Those 
 which are called open-running, by which we mean that they 
 are free and discharge their air all around the circumference; 
 and (2) those called closed- running fans, which have but a re- 
 stricted opening for the discharge of the air. Those of either 
 class are made large in diameter and are driven at a relatively 
 small angular velocity, though a few, such as the Schiele, are of 
 small diameter, running at a high angular velocity. They 
 produce large volumes of air at a low pressure, and may be- 
 reversed in motion to exhaust or to force air. The diameter 
 of the fans of this class may be and is occasionally as high. 
 as 50 feet, those of small diameter being regarded as unnec- 
 essarily cumbrous. The action of all fans is based upon the 
 general law that bodies in motion tend to travel in straight lines,, 
 resisting any attempt at diversion from this path, in consequence 
 of which, when the fan is set in motion, its blades come in con- 
 tact with its interior air, the particles of which are at rest and
 
 424 MANUAL OF MINING. 
 
 resist rotation. When, however, the particles do move, their 
 endeavor to travel in straight lines results in their making for 
 the circumference, producing thereby in the central portion of 
 the fan a partial vacuum, which is replaced by the air external to 
 the fan. So long as the rotation of the blades continues, so long 
 will this current be produced and maintained, the pressure of 
 which will increase as the peripheral speed increases. 
 
 Open-running Fans. These fans include the Waddle, Biram, 
 Naysmith, and Hopton patterns, all of which are essentially 
 similar to the first named. The Waddle is a self-contained fan, 
 in that there is no fixed casing, and the whole machine re- 
 volves. Its form is practically that of a light hollow disc of 
 wrought iron, the blades and casing being wholly riveted together. 
 The air enters by a straight lead at one side only, and passes 
 through curved and gradually narrowing channels to the cir- 
 cumference, the blades being bent at first to incline slightly back- 
 wards, the alternate blades extending not more than one half the 
 distance between the circumference and the inlet. The passages, 
 by their contraction, are so made that the circumference at any 
 point multiplied by the cross-sectional area at that point is a 
 constant quantity. The outer circumference of the fan is bell- 
 mouthed. 
 
 A fan of 9 feet diameter circulates 80,000 cu. ft. with a water- 
 gauge of 2 inches. One of 45 feet, driven by an engine with 
 40" X42" cylinder, at a boiler pressure of 80 Ibs. per square inch, 
 has given a volume of over 550,000 cu. ft. at 42 revolutions. 
 
 The Hopton has an inlet on each side of the central diaphragm 
 with backward-curving blades, and a construction very simple. 
 The revolving portion consists of the arms and blades working 
 between two brick walls. 
 
 The open-running fans must, in order to be efficient, dis- 
 charge their air at a' very low velocity, because the energy of 
 bodies in motion increases as the square of the velocity, and 
 that passed by the discharged air is, therefore, so much use- 
 less work. It is for this reason that the passages in Jthe more 
 correct open-running fans, like that of the Waddle, are curved
 
 METHODS OF VENTILATION. 425 
 
 backward. The theoretical depression which can be produced 
 in fans of this type is equal to the height due to its peripheral 
 speed, T t in feet per second. 
 
 The Closed-running Fans are essentially of a more massive 
 structure than those of the open-running type, being of con- 
 siderable width as well as of diameter. Of this class of fans 
 the Guibal is a type, the Schiele and the Walker Indestructible 
 being similar in construction. Inside of a fire-proof housing 
 a horizontal shaft is revolved by an engine or dynamo, carrying 
 with it an hexagonal or square frame, on which are built six or 
 eight blades. The blades are flat and slightly curved at their 
 tips, sometimes radially, and often inclined backwards. The 
 clearance between the tips of the blades and the casing is made 
 as little as possible, except for a certain distance at the bottom, 
 through which the air is discharged, the amount of that opening 
 being regulated by an adjustable shutter in a gradually enlarging 
 chimney. The air enters at the centre, whence it passes into one 
 of the intervals between the consecutive blades, which form an 
 evasee canal, the speed of exit being less than the speed of entry 
 
 (Fig. 159)- 
 
 In the Schiele fan the blades are contracted in width from 
 inlet to outlet, the fan being surrounded by the usual spiral cas- 
 ing, into which the air discharges all around the circumference, 
 the space continually increasing until it reaches the chimney. 
 The blades of the Rateau fan extend to the centre of the fan, and 
 have a peculiar curvature slightly forward, and also a curvature 
 in the line of the fan-shaft. Immediately in front of them is a 
 cone terminating in a point. The Capell fan, of equal power 
 with the Guibal, is smaller, and runs at a higher speed. It has 
 two concentric shells besides its outer casing, in each of which 
 are curved blades with the convex side forward. The air enters 
 the inner shell, is forced out through ports into the second outer 
 shell, where it strikes the concave face of the outer blade, and
 
 426 
 
 MANUAL OF MINING. 
 
 thence is discharged at a low velocity through the usual expand- 
 ing exhaust-flue. 
 
 The Champion fan, which is really two fans joined together 
 by a common centre ring, is designed to propel the air with a 
 minimum resistance, the blades having a backward curvature. 
 The us-? of the inner casing or hood and attendant diaphragm, 
 
 FIG. 150. Section of a Fan. 
 
 which are hung on frames, renders it possible to change the cur- 
 rent at will, blowing to exhaust, by revolving the hood around 
 the fan without stopping the latter. 
 
 The Evase"e Chimney. The depression produced by a cov- 
 ered ventilator with an expanding chimney is twice that of the 
 uncovered or open-running type, and is double the height, due 
 to the tangential speed of its blade-tips. The use of the chimney 
 gives to this type of fan the enormous advantage over the other 
 that the air may be discharged from the fan at a higher velocity 
 without any material loss of energy. The gradually increasing 
 space into which they discharge reduces the velocity and utilizes 
 all the energy in giving motion to the air, while the air is ulti-
 
 METHODS OF VENTILATION. 427 
 
 mately sent out into the open at a speed so low that practically no 
 resistance is experienced. 
 
 The Shutter Regulator. The volumes which these fans will 
 produce vary directly as the speed of their rotation, and their 
 manometric depression varies as the square of the speed of rota- 
 tion. Though they may be run at any speed at will, the efficiency 
 of the fan materially decreases when the speed of the tips of the 
 blades exceeds, to a great degree, 5000 feet per minute, or is less 
 than this quantity. The rate, however, which is regarded as 
 normal is 4000 feet of peripheral velocity per minute. Below or 
 above the normal speed a loss of velocity ensues in the discharg- 
 ing air, which alternately is expelled into the chimney or carried 
 with the blades into the fan, there to repeat its circuit. The 
 discharge is frequently followed by a vibration in the fan, to 
 remedy which the sliding shutter (ab, Fig. 159) is introduced. 
 Its correct position is only known by experiment in each indi- 
 vidual case, determined by the point at which the throbbing 
 ceases with the given speed. Though now universally placed 
 in all fans the " anti- vibration shutter" originated with the 
 Walker fan. 
 
 Influence of Shape, Dimension, and Speed of Fan upon its 
 Capacity. The ratio between the speed and the yield seems to be 
 as the 4th power of the speed to the 5th power of the yield. The 
 relation between the volume resistance of the fan and the power 
 required to drive it depends upon the resistance of the mine 
 which it supplies. These ratios have not been accurately for- 
 mulated. 
 
 Numerous experiments have been conducted upon centrifugal 
 ventilators with the view to determining the influence which the 
 various dimensions of the fan and shapes of its parts will have 
 upon its performance; and the following conclusions are cited 
 from the results of the tests made by R. Van A. Norris, Wilkes- 
 barre, Pa., upon 25 fans, as the influence of: "ist. The diameter 
 on their performance seems nil; the only advantage of large 
 fans being in greater width and a lower speed required of the 
 engines. 2d. Width upon efficiency is, as a rule, small. 3d. Shape
 
 428 MANUAL OF MINING. 
 
 of blades shows that the back curvature is better, and dimin- 
 ishes the vibration. 4th. Shape of casing is considerable. 
 The proper shape would be one of such form that the air be- 
 tween each pair of blades would constantly and freely discharge 
 into the space between the fan and casing, the whole being 
 swept to the &vase chimney. A large spiral, beginning at or near 
 the point of cut-off, gives in every case a large efficiency. 5th. 
 The shutter on the fan is beneficial. The exit area can be reg- 
 ulated to suit the varying quantity of air, and prevent re-entries. 
 6th. Speed at which the fan is run. The efficiency is high if the 
 peripheral velocity is large." 
 
 Automatic Speed and Pressure Recorders. Some States 
 require self -recorders on all fans, by which the number of revolu- 
 tions of the fan shall be registered every hour and such data to 
 be taken and reported. In other States also is required an auto- 
 matic regulator for the water-gauge. The speed- registers are 
 generally constructed of a metal pedestal erected on blocks at 
 the side of the fan or engine-shaft, a small vertical shaft to which 
 a governor is attached. A small cog-wheel on the lower end 
 geared to a large driver on the fan or engine-shaft communicates 
 the speed to the governor, which, by a system of leverage, raises 
 or lowers the arm to which is attached a pen that presses against a 
 paper dial held in position by a light case of sheet brass. The 
 higher the speed of the fan, the more will the governor raise the 
 lever, and consequently the pen register. The time is recorded 
 by a clock to whose shaft the dial case is attached. In other 
 devices the dial case is a cylinder in which is rolled a sheet of 
 paper turning on a horizontal axis, which is also the continuation 
 of the shaft of the clock. These instruments perform the work 
 expected of them with great satisfaction. 
 
 Comparison of Fan and Furnace. The ventilation current 
 from a fan is affected by a low barometer or a high temperature. 
 Either one requires an increased degree of rarefaction just as 
 from a furnace. Moreover, as the depth of the mine increases, 
 the work devolving upon the fan proportionally increases, be- 
 cause normally the air becomes denser; with every additional
 
 METHODS OF VENTILATION. 429 
 
 thousand feet of depth, an increased rarefaction or depression of 
 0.4 inch of water-gauge is necessary. Compared with the fur- 
 nace its efficiency decreases with the depth of the upcast until, 
 at a certain depth, it becomes an open question 'between the 
 relative merits and demerits of fan and furnace, as to which will 
 be the more economical. For shallow works the exhaust-fan 
 undoubtedly takes precedence. At the depth of a thousand feet 
 a large furnace will equal a very imperfect fan, consuming 20 
 Ibs. of fuel per hourly horse-power; a good fan and condensing 
 engine will be cheaper than a furnace down to the depth of about 
 4000 feet. Taking cognizance of the objections to the furnace, 
 it must also be borne in mind that machine ventilators are sub- 
 ject to serious objections, since during the time of their repair 
 the mine must remain un ventilated, whereas with a furnace, after 
 its fire has been extinguished, a considerable circulation will still 
 continue in the upcast for some time. Auxiliary ventilating 
 appliances should be supplied against any emergency which 
 arises during the repair of the fan. 
 
 The Theory of the Fan. The theoretical depression of a fan, 
 the difference between the pressure of the entering air and the 
 discharged volume, is represented in terms of a motive column 
 of the density of the inflowing air. An ideal ventilator will 
 produce a depression which is twice the height created by the 
 tangential speed of the tips of the blades. If, then, H be the 
 height of the motive column due to the velocity, T, of the fan- 
 
 J-2 
 
 tips in feet per second, H will equal ; but imperfections of 
 
 detail prevent such an initial depression being attained, and 
 representing them by a coefficient K, which is always less than 
 unity, reaching 0.85 in Guibal fans, but more often being 
 below 0.6 in the average construction of fans, the fan approaches 
 an ideally perfect one when K approximates to unity. The 
 yield of the fan then in barometric depression, or its useful effect, 
 
 KT 2 
 is H = . Various enfeebling causes reduce the capacity 
 
 o 
 
 of the fan to 'determine the value for K. The quantity of air
 
 4,)0 MANUAL OF MINING. 
 
 which passes through an orifice is equal to the product of the 
 area and the velocity when no friction is encountered; but when 
 any fluid flows through an orifice in a plane- surface a considerable 
 diminution of the discharge takes place, because of the contraction 
 of the steam resulting from the convergent flow. The coefficient 
 corresponding to this contraction, known as the vena contracta, 
 is 0.65; hence with a given velocity, T, and a head, H, under 
 the conditions modified by the coefficient K as above, the dis- 
 charge of air per second will become 
 
 Hence it is evident that if the capacity of the mine is such 
 that it is incapable of delivering to the fan the volume of air 
 equal to the body capacity of the latter at a given speed, the 
 frictional resistances encountered in the mine will reduce the 
 efficiency of the fan by some quantity which is usually com- 
 prehended in the symbol a, representing the area of the mine's 
 "equivalent orifice" in square feet Experiments have demon- 
 strated that when a is 20 square feet, only 65,000 cubic feet of 
 air are obtained per minute for the fan peripheral speed of 5000 
 feet per minute ; but when the mine resistances have been reduced 
 until its "equivalent orifice" is as large as 100 square feet, 280,000 
 cubic feet of air are obtained from the same speed of fan. 
 Having the value for this friction, which in earlier days was 
 known as the temperament of the mine, we arc enabled to 
 understand the conditions under which the ventilator is working 
 and to provide for them. 
 
 The Equivalent Orifice of the Fan, which is designated by 
 the symbol o, may be determined in a like manner. It measures 
 or represents the orifice in a thin plate which offers such a resist- 
 ance to the flow of the current, Q, as is equal in effect to the 
 aggregate resistances encountered within the fan from its imper- 
 fections. If H is the theoretical depression which the fan should 
 produce when moving at a tangential speed, T, per second, and 
 h represents the actual or the effective depression which is pro-
 
 ELEVATION 
 FIG. 1 60. Details of a Fan House. 
 
 43*
 
 43 2 MANUAL OF MINING. 
 
 duced upon the air as measured by the water-gauge, then H h 
 is the head wasted by the fan in its construction and may be 
 represented by & , which measures the head corresponding to 
 the equivalent orifice. In large fans its value varies from 16 
 to 80 square feet. 
 
 The head lost in the fan, represented by h , is equal to H h t 
 the velocity due to which may be determined by the expression 
 
 Aslhe value of h approaches zero and that of h approaches 
 H, the fan approaches an ideally perfect ventilator. The actual 
 velocity through the orifice of entry is 0.657;,,, whence the area 
 of the orifice o, which equals the quantity flowing per second, 
 divided by the velocity of the flow, has the following value: 
 
 The density of water being 833 times that of air, the ratio 
 between the water-gauge reading and the height of the motive 
 column, H, is i : 833. To convert the water-gauge reading 
 to a height H of air-column of equal weight in feet, the height 
 of the water-gauge, m, in inches is multiplied by 69.4. 
 
 The ratio between the lost head in the ventilator and the 
 effective head represented by the water-gauge is expressed in 
 the equation 
 
 whence 
 
 The Relation between Mine and Fan. The quantity of 
 flow through a fan, depending on the relations which the area 
 of the mine airways and the condition of their rubbing surface 
 bear to the mechanical condition of the fan, is manifestly depend-
 
 METHODS OF VENTILATION. 
 
 433 
 
 SECTION ON LINE A-B 
 
 PLAN 
 FlG. 161. Plan of the Fan House and the Curve of the Fan Housing.
 
 434 MANUAL OF MINING. 
 
 ent upon a proper ratio of a to o, which ratio may be expressed 
 as the "appropriateness of the fan to the mine." When this 
 value is equal to or greater than unity, the fan would be too 
 small for the mine, and it is questionable whether any air would 
 flow under those conditions. As the ratio becomes smaller, the 
 conditions become more favorable for the fan. When approxi- 
 mating a ratio of 0.3 the orifice of discharge of the fan is to be 
 considered as having a fair working ratio. More air is obtained 
 by a given fan and at a given velocity when a is large than when 
 a is small, for, no matter how well constructed the fan may be, 
 it cannot provide a quantity equal to its body capacity unless 
 the mine can pass this quantity. The effective work done upon 
 the air is less in the latter case than in the former for a given 
 volume of air. The mechanical work of centrifugal force is 
 o.oooo34o(r 2 F 2 2 ). In this T is the circumferential velocity 
 and V 2 is the absolute velocity at expulsion, due to compression 
 from centrifugal force. As F 2 increases, so the work on the 
 departing air, and proportionately the effective work, decreases. 
 The use of the funnel-chamber reduces this quantity to or -5, and 
 the work lost to 4 or 5 per cent. 
 
 The Efficiency of the Fan. The manometric efficiency is 
 the ratio between the effective pressure, P, and that due to the 
 centrifugal velocity, while the mechanical efficiency is that of the 
 Horse-power in the air to the engine-duty. With fan properly 
 constructed, the efficiency approximates about 6 per cent. In 
 experimentally measuring the efficiency of a fan, it is customary 
 to determine the dynamometric resistance and internal friction 
 when its orifices have been cut off from any communication 
 with the mine, the air being then drawn from the atmosphere and, 
 after passing through a fan, expelled at its throat. Counting 
 the rate of revolution and estimating the volume of air which 
 is moved, the power necessary to overcome this friction is deter- 
 mined and expressed in feet of the air-column whose weight 
 equals the aggregate friction. This quantity, divided by the 
 theoretical head corresponding to the velocity of the fan, deter- . 
 mines its efficiency under the conditions named. The fan is
 
 METHODS OF VENTILATION. 435 
 
 giving its maximum efficiency when "its body capacity just 
 exceeds the quantity the mine will pass at a gauge pressure, F, 
 due to the speed of rotation of the fan." 
 
 The Design of a Fan. In designing a fan for a given mine, 
 the essential elements are Q and m. These given, the diameter, 
 the peripheral speed, and the length and width of blades, as 
 well as the direction of their inclination, must be determined 
 by the engineer. As to diameter, it may be said that the slow- 
 running fans are regarded as cumbersome and costly, requiring 
 expensive foundations. Large fans may be run at a lower rate 
 of revolution and produce the same tangential speed as would a 
 fan of small diameter. Insomuch as speed is the important 
 factor in the construction of ventilators, due consideration must 
 be given to this question, which is determined by local condi- 
 tions of place, economy, and mechanical simplicity. A con- 
 venient rate of revolution for a fan directly connected with the 
 engine is about 60 per minute. The body capacity of the fan 
 should be large enough to maintain the required pressure, P, 
 without great variations in speed. Though the practice of 
 European engineers tends towards the rate of tangential speed 
 which represents 5000 feet per minute and over, in this country 
 4000 may be considered as normal. In any event, if the calcu- 
 lation and design be made on the assumption of either normal 
 speed, it will be possible, when an emergency arises, to increase 
 the speed sufficiently to give a volume nearly one tenth greater 
 than the normal quantity. Moreover, when the mine is dry 
 and dusty, it will be possible to turn the whole volume of the 
 excess into any or each single split, through ' which it may be 
 drawn, clearing away fine dust and moisture. 
 
 The entry for the air should be made easy and large, prefer- 
 ably divided into two inlets, one on each side, with a diaphragm 
 to prevent the currents from conflicting. This necessitates a 
 wide fan, which, however, gives a volume proportionately greater 
 than what is to be had from a single fan with a single large inlet. 
 
 The length of the blades of the fan should be only a little 
 greater radially than the difference between the radii of the
 
 43<5 MANUAL OF MINING. 
 
 fan and its Inlet. With a large inlet the blade necessarily is 
 shortened, and when pressure is desired the blade length should 
 be increased to as large a quantity as possible by providing two 
 inlets. Notwithstanding that the width of the fans is much greater 
 than would be obtained by substitution in the formulae follow- 
 ing, it is certain that the latter dimensions correspond to a greater 
 efficiency. M. G. Hanarte concludes that "the Guibal fan has 
 always been eight or nine times too wide, and the Capell is nearly 
 as bad." 
 
 The shape of the blade should be such as would present to- 
 the circumference of the outlet an inclination following the 
 resultant of the movement of rotation and of the movement of 
 the air penetrating the spaces between them. The blades of 
 open-running fans curve backward. The backward curvature 
 is conceded to give a freer delivery, and the forward curvature 
 at the tips a higher water-gauge pressure. The number of blades 
 is seemingly a matter of indifference, though the limit may doubt- 
 less be determined by the inevitable friction produced by the 
 excessive surface of contact when too numerous. The friction 
 varies as the cube of the section of space between two vanes. 
 As to the shape to be given to the casing, it will be noticed that 
 the original Guibal fan had no spiral, the tips of the blades re- 
 volving but two inches clear of the casing, and the spiral enlarge- 
 ment beginning at the angle of about 67 30' from the lower 
 vertical radius. Those fans presenting a large spiral beginning 
 at or near the cut-off and increasing about six inches for each, 
 45 up to 275, and thence widening by an increasing increment 
 to the evasee chimney, appear to give larger efficiency by allowing 
 for the slackening of the speed of the air, and discharge the air 
 with less energy at the exit. M. G. Hanarte concludes that the 
 spiral envelope is not necessary. 
 
 Below are given formulae for the computation of dimensions 
 of a Guibal fan in accordance with the data indicated above. 
 All dimensions are in feet. 
 
 .D=diameter of the fan between the blade-tips: Q^-2oo=Z> 2 ; 
 /== length of the blades in feet;
 
 METHODS OF VENTILATION. 
 
 r=their radial length =2. 6im; 
 
 x = their width =A .-*- 27:5; 
 
 A = aggregate area, of the one or two inlet-ports in square feet 
 (radius of each central inlet, s) =Q-r- 1300; 
 
 JV=number of revolutions per minute; 
 
 r=peripheral speed of fan per second =1^-^19.0985; 
 
 F= theoretical velocity per second due to head H; 
 
 v l = velocity of the centre of gyration of air-column between* 
 the blades =7t(D-r}N+ 60; 
 
 p= radius of gyration of the mass of air=^(Z) r). 
 
 W= weight of the unit of revolving air-column per foot of fan- 
 width =o.o766r; \ 
 
 F= centrifugal force of the air-mass in Ibs. per square foot of dis- 
 charge area or of the housing =TF'i>, 2 -^7rg=o.ooo7578rz; 1 2 ;. 
 
 F 2 =velocity of air discharge per second =>J- ' r 
 
 Z= minimum area of discharge-port =A -r- 2 ; 
 
 h =fa.n resistance, measured in feet of head, =Hh\ 
 
 O=area of orifice offering a resistance to the flow of Q cubic 
 
 feet of air per minute, equal to that of the fan; 
 q= quantity of air discharged by the fan per second in cubic 
 
 feet=o.6sF 2 Z; 
 Q quantity of air discharged by the fan per minute in cubjc 
 
 feet =60?; 
 
 w=mine resistance in inches of water-gauge; 
 P=mine resistance in pounds per square foot =5. 184 m. 
 
 In Figs. 160 to 162 are illustrated the details of the ordinary- 
 pattern of fan which is designed in accordance with the con- 
 ditions indicated above. As, fortunately, neither the Guibal 
 fan nor the shutter is subject to patent, the working drawings 
 here given may aid the construction engineer. 
 
 When the conditions are satisfied by the revolution of the 
 fan of proper proportions, the centrifugal pressure of the fan- 
 should produce a depression, F, equal to or exceeding P, the 
 mine resistance, in order that the requisite discharge through
 
 438 
 
 MANUAL OF MINING. 
 
 Flo. 162. Details of the Fan.
 
 METHODS OF VENTILATION, 
 
 43? 
 
 the outlet should equal the desired quantity Q. When it is 
 discovered that the volume of discharge is deficient, the fan 
 dimensions D or r should be enlarged or the rate of revolution 
 increased. Below is a brief table indicating the theoretical 
 water-gauge depression in inches for the corresponding periph- 
 eral speeds in feet per second. 
 
 Speed in 
 Feet 
 of the 
 Periphery 
 per Second. 
 
 Depression 
 in Feet 
 of Air- 
 column. 
 
 Depression 
 in Inches 
 of Water- 
 gauge. 
 
 Speed in 
 Feet 
 of the 
 Periphery 
 per Second. 
 
 Depression 
 in Feet 
 of Air- 
 column. 
 
 Depression 
 in Inches 
 of Water- 
 gauge. 
 
 T 
 
 U 
 
 m 
 
 T ' r> 
 
 M ' 
 
 m 
 
 30 
 
 2 7 -,95 
 
 0.411 
 
 75 
 
 I74-.69 
 
 2.569 
 
 35 
 
 38.04 
 
 Q-559 
 
 80 
 
 I98-75 
 
 2.922 
 
 40 
 
 49.69 
 
 0.731 
 
 85 
 
 224.38 
 
 3-299 
 
 45 
 
 62.82 
 
 0.924 
 
 9P i 
 
 25I-55 
 
 3-699 
 
 5 
 
 77-63 
 
 i,. 141 
 
 95 , 
 
 280.28 
 
 4.122 
 
 55 
 
 93-94 
 
 1.381 
 
 100 
 
 310-55 
 
 4-567 
 
 60 
 
 in. 80 
 
 i . 644 
 
 i5 
 
 342.39 
 
 5-035 
 
 65 
 
 131.18 
 
 i .929 
 
 IIO 
 
 375-77 , 
 
 5-526 
 
 70 
 
 152.17 
 
 2.238 
 
 115 
 
 410.71 
 
 6.039 
 
 EXAMPLE. Required the dimensions of a fan to provide 125,000 cubic 
 feet of air against a mine resistance of -2.5 inches. 
 
 At a normal rate of 65 feet per second, the diameter becomes 25 feet; the 
 area of the inlets is 96 square feet, the diameter being n feet; the radial 
 length of the blade is 6.5 feet; the minimum \width of the blade is to be 2.8 
 feet. 
 
 As =9.25 feet and ^ = 51 feet, the centrifugal pressure, F, becomes 12.7 
 pounds per square foot of radial column; and the velocity of discharge 31.6 
 feet per second, which with a minimum area of discharge-port, Z, of 48 square 
 feet, would furnish less than 60,000 cubic feet. The' mine resistance exceeds 
 the standard allowance of i inch of water gauge for each 100,000 cubic feet 
 of air. The mine airways should be enlarged or the fan. operated at a higher 
 speed. An increased rate of 70 revolutions per minute will produce a ventilat- 
 ing pressure of 22 pounds "per 'square foot. The 'blades -may be lengthened 
 and two inlet orifices be provided, each of 48 square feet in area. 
 
 At the peripheral velocity, T, of 91.66 feet per second the theoretical head 
 of discharge is 261 feet. But the effective head, h, against which the fan is 
 operating, measured by the water-gauge, is 166.66 feet. Under the condi- 
 tions of operation, therefore, the loss of head, h , in the fan is 94.34 feet; since 
 the equivalent orifice of the mine is 21.2 square feet, the equivalent orifice of 
 the fan, O, is 41.2 feet. The ratio of a to O, nrly one half, represents a fair 
 working ratio of appropriateness of fan to mine.
 
 44 MAXUAL OF MINING. 
 
 Electrically driven blowers and ventilating-fans may be con- 
 nected either by belt, gearing, or mechanical coupling. The 
 compound-wound D. C. machine is preferred for fans on account 
 of its good speed regulation and freedom from racing. 
 
 Alternating-current induction-motors of the squirrel-cage type 
 would give equal satisfaction. 
 
 REFERENCES. 
 
 Curvature of Fan Blade and Design, M. & M., Vol. XX, 165; Fan Effi- 
 ciency, M. & M., Vol. XXIII, 406; Location of Fan, M. & M., Vol. XXIII, 
 22; Fan Ventilation, M. & M., Vol. XXIII, 260-71; Laws Governing Useful 
 Work of a Fan, H. Heenan and William Gilbert, Coll. Guard., April 1897, 
 720; and M. G. Hanarte, Coll. Guard., Mar. 1898, 552; Descriptive Lecture 
 on Fans, C. M. Percy, Coll. Mgr., 1894, 56; Fan Volume Recorder, M. 
 Murge, Coll. Guard., April 12, 1900; Modern Fans, Charles H. Innes, Eng., 
 Nov. 14, 1902; and Theory and Practice of Fan Design, Heating and Ven- 
 tilation, Serial, 1903 and 1904. 
 
 Underground Temperatures in Relation to Deep Mining, F. C. Meachem, 
 Coll. Guard., April 7, 1903; The Practical Scope of the Equivalent Orifice 
 Theory, H. W. Halbaum, Coll. Guard., Aug. 17, 1900. 
 
 Improvised Ventilating Machinery, C. M. Neyrick, Min. Sci. Press, April 
 ' n, 1903; Ventilation Furnace, Illustrated, Pa. Mine Insp., 1891, 404, 190. 
 
 Root Blower, Haswell, 449; To Compute Elements of Fan, Haswell, 448; 
 Fan Temperament, Clifford, Am. Mfr., Feb. 19, 1897, 264; Formulae, Coll. 
 Guard., Mar. 12, 1897, 489; Fan Work, Coll. Guard., Mar. 19,552; Mano- 
 metric Efficiency, Fed. Inst. M. E., 111.; Fan Tests by Van A. Norris, A. I. 
 M. E., XX, 637; The Comparative Consumption of Fuel by Ventilating Fur- 
 naces and by Ventilating Machines, J. J. Atkinson, N. E. I., VI, 135. 
 
 Guibal Fan Compared with Dynamo, Coll. Guard., Vol. LXXXIV, 847; 
 Comparative Tests of Fans, Coll. Guard., LXXIX, 218; Double Ventilation 
 Fans, Coll. Guard., LXXXI, 127. 
 
 Equivalent Orifice of Mine, Table for Finding, Coll. Guard., Jan. 1897, 
 .168; Speed Indicators, Pa. Mine Insp., 1886, 22, 24; 1891, 89. 
 ' '' JA 
 
 ih to 
 ' 
 
 :>ifr . 
 

 
 : 
 
 
 
 
 ' 
 
 CHAPTER XIV. 
 
 DISTRIBUTION OF THE AIR. 
 
 The Effective Motive Column. It has been assumed that 
 the work done upon the air is totally effective in the mine; that, 
 with a given M and P, the calculated quantity of air is obtained 
 without loss; that the momentum, once imparted to the air, 
 would be available to carry it through the mine and out. This 
 is not so. Friction consumes some of the kinetic energy; only 
 a fractional part of the theoretical velocity is realized in practice. 
 The rough sides of the galleries and rooms, their sharp corners, 
 and the diminished areas offer resistances to the passage of the 
 current that consume often 90 per cent of the power. More- 
 over, the subtle air under pressure seeks to escape at every oppor- 
 tunity, and some portion of the precious fluid is lost into the goaf, 
 through doors and at crossings. A certain mine theoretically 
 required a pressure of but 1.2 Ibs. per foot to give rise to its cur- 
 rent, yet the friction was such that u.8 Ibs. were actually neces- 
 sary to create the velocity. Not infrequently the ratio between 
 M (to which the generation of the final velocity at the top is 
 due) and M', the head actually necessary to overcome resist- 
 ances, is as low as i : 18. In other words, only 5.5 per cent 
 of the work done upon the air is usefully expended. Any means 
 of reducing this loss is to be welcomed. 
 
 The Frictional Resistance of Air in Mines. Let us examine 
 into the laws governing the movement of fluids, that we may 
 reduce this friction to a minimum, and obtain salubrity, safety, 
 and economy with the least outlay. The air which enters the 
 mine from the downcast is distributed to the rooms and chambers 
 hi proportions varying with their several needs; or the current 
 
 441
 
 442 MANUAL OF MINING. 
 
 as one mass sweeps through the main way, along working faces, 
 thence by return airway over the furnace or to the fan. The 
 resistances encountered depend upon the ratio of the area of 
 the surface rubbed to the area of the conduit, and upon the coef- 
 ficient of air friction against rock. A satisfactory value for the 
 coefficient has not been obtained: the records of experiments 
 show it to vary, as in water, according to the nature of the conduit 
 and the velocity of the flow. The coefficient varies with the 
 nature of the rubbing-surface, and consequently differs in various 
 air- passages of the same mine; nevertheless, the numerous ex- 
 perimenters have announced values for the coefficient of friction 
 of air in mines for each foot of rubbing-surface and for a velocity 
 of one foot per minute as varying between 0.000000008585 
 and 0.0000000219, with an accepted average of o.oooooooi Ib. 
 This value for /, the coefficient, is measured in the pounds pres- 
 sure per square foot. Measuring the height of an air-column 
 in decimals of a foot, the value for the coefficient of friction 
 /' has been found to be 0.00000010635 to 0.00000026881, the 
 two extremes of values for /' being heads of air-column corre- 
 sponding to the values given for dynamic pressures, /. In other 
 words, an area of i square foot of rubbing- surf ace presents to 
 the passage of an air-current moving at i foot per minute a 
 frictional resistance of o.oooooooi Ib., or an equivalent resistance 
 of 0.000000125 feet of head of air-column. 
 
 Let / be the length, m the perimeter, and a the area of the 
 gangway, through which the air is coursing at v feet per minute, 
 and the rubbing friction is found experimentally to be jlmv 2 . 
 Imagine a piston, fitting air-tight in the passage; to just move 
 it against the resistance requires the expenditure of a force pa, 
 in units of pounds and square feet. Therefore the loss of power 
 
 due to friction is pa = jlmv*, and the loss of head in feet, p= , 
 
 or in pounds, 
 
 j'lmv* 
 
 Shape and Dimensions of Airways. Cognizance must be 
 taken of these frictional resistances by ascertaining, for various
 
 DISTRIBUTION OF THE AIR. 443 
 
 portions of the mine, the elements involved and comparing the 
 value for p, thus ascertained for .a given mine, with that usually 
 considered as the standard of comparisons. The normal mine 
 should be capable of delivering 100,000 cu. ft. of air to the men 
 throughout its working without causing a frictional resistance 
 greater than 5.2 Ibs. per square foot in the aggregate. Should it 
 be ascertained that the mine shows a larger water-gauge, it is 
 evident that there is room for improvement. 
 
 The lines along which improvement may be had involve an 
 increase in the area of the .cross-section, or a change of shape, 
 such as to give a greater flow with less frictional rubbing- surf ace. 
 The conditions under which the airways are driven prevent any 
 reduction in the coefficient of friction. The friction increases 
 with the square of the velocity, and in order to increase the air- 
 current the velocity may be increased, but this increases the 
 frictional resistances rapidly. Assuming the value for / of 
 o.ooooooi, then the quantity that will flow through i sq. ft. 
 of area of passage at different velocities, as stated below, in a 
 length of one mile, will be as indicated, and with corresponding 
 resistance, p, as follows: 
 
 Feet. Cubic Feet. Lbs. per Square Foot. 
 
 Velocity i, quantity Q= 60; then^ = 0.76 
 
 " ' 2, " <2=i2o; " p= 3.06 
 
 5, <2= 3 oo; " f =19.00 
 
 " 10, " Q =600; " . = 76.03 
 
 The frictional loss is directly proportional to the perimeter, 
 and inversely as the area. Thus, a local contraction of the 
 air-passage by the use of the partly open door, or the accidental 
 accumulation of a pile of waste, or gangue, will materially dimin- 
 ish the volume of air passing through it and increase in a large 
 measure the resistance to such flow as does occur. It is there- 
 fore desirable to select such a shape for the airway as will make 
 it as spacious as circumstances will permit consistent with the 
 dimension of the rubbing- surf ace. The circular form most nearly 
 meets these requirements, but we are restricted, as a rule, to the
 
 444 
 
 MANUAL OF MINING. 
 
 rectangular or the trapezoidal cross-section. Aside from the 
 theoretical consideration, the form of airway is modified by prac- 
 tical conditions. In American mines the circular airway is 
 rarely adopted, and only in very rare instances is the Stanley 
 header used as in England. There a circular lining follows 
 the machine very closely, and the proper form can be retained in 
 the finished airway. The expense of making this form pre- 
 Tents its common use in America. The more usual form is 
 one having a semicircular arch roof and smooth walls as being 
 the nearest approach to the ideal cross-section. 
 
 As showing the comparisons between the resistance afforded 
 by a square, trapezoidal road and the round roadway of various 
 dimensions and velocities, as stated below, the following table 
 will be of service. It shows the pressure in pounds per square 
 foot due to the flow of air through one mile of airway with the 
 coefficient of o.ooooooi Ib. per square foot. 
 
 
 
 At Velocity of 
 
 At Velocity of 
 
 At Velocity of 
 
 Dimensions 
 of 
 
 Diameter 
 of 
 
 i Foot per 
 Second. 
 
 S Feet per 
 Second. 
 
 i o Feet per 
 Second. 
 
 
 
 t 
 
 Q 
 
 P 
 
 Q 
 
 P 
 
 Q 
 
 Feet. 
 
 Feet. 
 
 Pounds. 
 
 
 Pounds. 
 
 
 Pounds, 
 
 
 IX I 
 
 i 
 
 .7603 
 
 60 
 
 19.00 
 
 300 
 
 76.03 
 
 600 
 
 2X 2 
 
 2 
 
 .,3801 
 
 240 
 
 9-50 
 
 1,200 
 
 38.01 
 
 2,400 
 
 3X 3 
 
 3 
 
 2534 ' 
 
 540 
 
 6-33 
 
 2,700 
 
 25-34 
 
 5,400 
 
 4X 4 
 
 4 
 
 . 1900 
 
 960 
 
 4-75 
 
 4,800 
 
 19.00 
 
 9,600 
 
 5X 5 
 
 5 
 
 .1520 
 
 1500 
 
 3.80 
 
 7,500 
 
 15.20 
 
 15,000 
 
 6X 6 
 
 6 
 
 .1267 
 
 2160 
 
 3.16 
 
 10,850 
 
 12.67 
 
 21,600 
 
 7X 7 
 
 7 
 
 . 1086 ' 
 
 2940 
 
 2.71 
 
 14,700 
 
 10.86 
 
 29,400 
 
 8X 8 
 
 8 
 
 0954 
 
 3840 
 
 2.38 
 
 19,200 
 
 5-94 
 
 38,400 
 
 9X 9 
 
 9 
 
 .0844 
 
 4860 
 
 2. II 
 
 24,300 
 
 8-44 
 
 48,600 
 
 10X10 
 
 10 
 
 .0760 
 
 6000 
 
 1.90 
 
 30,000 
 
 7.60 
 
 60,000 
 
 It will be noted then that two galleries having a cross-sec- 
 tional area of 5'Xs' would require twice as much power to carry 
 the same amount of air as would a single gallery of lo'Xio'. 
 The pressure necessary to drive 25 cu. ft. of air per second through 
 one of the small galleries is 3.8 Ibs. over the total area of the 
 road. The pressure necessary to drive 100 cu. ft. per second 
 through a large gallery would require the pressure of 7.6 Ibs. over
 
 DISTRIBUTION OF THE AIR. 
 
 445 
 
 the total cross-sectional area. This is equal in amount to resist- 
 ances of both small galleries, which, however, are transmitting 
 but 50 cu. ft. of air per second. 
 
 The frictional resistance varies with the length of the gallery. 
 The volumes of air passing through two having the same resist- 
 ance will be inversely as the square root of their lengths. For 
 example, a gallery 1600 feet long carrying 6000 cu. ft. of air offers 
 the same resistance and consumes the same amount of power as 
 one of equal area which is 711 feet long delivering 9000 cu. ft. 
 per minute. 
 
 The following table shows the length of roads offering a 
 resistance equal to i inch of water-gauge under the conditions 
 given : 
 
 Square Airway. 
 
 Round Airway. 
 
 
 Dimensions 
 of 
 
 Area. 
 
 Periphery, 
 
 Diameter 
 of 
 
 Area. 
 
 Circumfer- 
 ence, 
 
 Length 
 of Road 
 in Yards. 
 
 Airway. 
 
 
 
 Airway. 
 
 
 tn. 
 
 
 Feet. 
 
 Sq. Feet 
 
 
 Feet. 
 
 
 
 ; ' ' 
 
 iX i . 
 
 I 
 
 4 
 
 I , 
 
 -7854 
 
 3.1416 
 
 481 
 
 2X 2 
 
 4 
 
 8 
 
 2 
 
 3-I4I 
 
 6.2832 
 
 962 
 
 3X 3 
 
 9 
 
 12 
 
 3 
 
 7.068 
 
 9.4248 
 
 1444 
 
 4X 4 
 
 16 
 
 16 
 
 4 
 
 12.566 
 
 12.5664. 
 
 1926 
 
 5X 5 
 
 25 
 
 20 
 
 5 
 
 I9-635 
 
 15.7080 
 
 2407 
 
 6X 6 
 
 36 
 
 24 
 
 6 
 
 28.274 
 
 18.8496 
 
 2888 
 
 7X 7 
 
 49 
 
 28 
 
 7 
 
 38.484 
 
 21.9912 
 
 337 
 
 8X 8 
 
 64 
 
 32 
 
 8 
 
 50-265 
 
 25.1328 
 
 3850 
 
 9X 9 
 
 81 
 
 36 
 
 9 
 
 63.617 
 
 28.2744 
 
 4333 
 
 10X10 
 
 100 
 
 40 
 
 10 
 
 78.540 
 
 31.4160 
 
 4814 
 
 In using this table it will be remembered that the volume of 
 air and the lengths of the road are directly proportional to the 
 frictional resistance, p, the other conditions of area and periph- 
 ery being the same. For example, a gallery which is 1200 yards 
 long, carrying 6000 cu. ft., would offer one half the resistance 
 that a road 2400 yards long would, or, in other words, if the cross- 
 sectional area be s'Xs'i the water-gauge would read J inch in 
 the short gallery and i inch in the long one. Again, the road- 
 way of f Xf, which is 6740 feet long, would offer a resistance 
 measured by a water-gauge of two thirds of an inch.
 
 446 MANUAL OF MINING. 
 
 Calculation of Mine Resistance. It has already been 
 remarked that the water-gauge measures the drag of the air in 
 a mine, and thus serves to indicate the pressure and head cor- 
 responding to the motive column, M. The pressure varies from 
 | inch for easy to 4 inches for difficult ventilation (from 3.9 to 
 20.7 Ibs. per square foot). In anthracite mines it is about 2 
 inches. The motive column, which is to just maintain this 
 pressure against resistances, should also be sufficient to create 
 a final or exit velocity in the shaft. If the entire current trav- 
 erses the mine unbroken, the resistance in the shaft or entry is 
 only a fractional part of the mine friction indicated by the water- 
 gauge, and the following formulae apply with sufficient accuracy: 
 
 V^P^L Q ,J*P 
 
 jlm' V jlni 
 
 The value for / is to be taken always in the same unit as that 
 for p. In other words, if the mine resistance, p, be given in 
 pounds per square foot, the corresponding value for / is taken 
 as equal to 0.0000000219; or if the value for p' is given in feet of 
 head of motive column, M, the value for /' is then 0.000000269. 
 
 If the airways in the mine, the resistances of which are to be 
 calculated with a view to determining the necessary ventilator 
 pressure to produce circulation, are all of the same dimensions, 
 the calculation of the lost pressure may be made in one operation 
 by proper substitution for the length, periphery, and area of 
 airway and the velocity or quantity concerned. The value of 
 the frictional resistance, />, thus engendered in the mine cor- 
 responding to the water-gauge height, m, and of the velocity of 
 the air-cuifent, added to that of the pressure, p", requisite for 
 the generation of the velocity, determines the motor pressure 
 required. Often p" is very small compared to p, and may be 
 even neglected without sensible error; but when it is large, the 
 actual ventilating pressure, which must be supplied by the force- 
 fan, or the manometric depression to be produced by a furnace, 
 or exhaust-fan, must be such as exceeds the sum of />+ p". 
 
 When- the airways of the mine are of various cross-sections, 
 the resistance offered by them in the aggregate must be deter-
 
 DISTRIBUTION OF THE AIR. 447 
 
 mined by adding together the separate values for p, calculated 
 for each differing cross-sectional area and length. When the 
 air-current is "split" into several smaller branches, and circu- 
 lated through an equal number of divisions of the mine, more 
 or less equal in length, with volumes of greater or less velocity, 
 the value for p must be calculated in each division or district 
 separately; and for each differing airway the aggregate resist- 
 ance in each division is the sum of the resistances encountered 
 in each of its various galleries. The sum of the several fric- 
 tional losses of head or of pressure, and that pressure or head 
 which produces the final velocity at the mouth of the mine, is 
 again equal to the ventilating pressure demanded of the motor. 
 
 Formerly it was the practice to meander the air through 
 all the galleries of each lift before expelling it (Fig. 163). This 
 involved heavy pressures, enormous airways, or a velocity dan- 
 gerously fast, and the last gang, fed by the departing current, 
 would receive an irrespirable atmosphere, vitiated by the emana- 
 tions from all previous sources. There was nothing to commend 
 this pernicious system, and it is certainly a matter of congratula- 
 tion that it has become obsolete. 
 
 Multiple Air-circuits. Mr. J. Buddie introduced a system of 
 ventilation for fiery mines that has everything in its favor. This 
 system was known at first as "coursing the air," and now is 
 termed "splitting the air," the inception of which is due to Carlisle 
 Spedding or his son, of Whitehaven, who introduced it in 1 763. By 
 it the aggregate quantity of air is increased, the dangers of explo- 
 sion are lessened by confining its train of evils to one portion of 
 the mine, and power for ventilation and haulage is saved, since 
 it goes hand in hand with the method of panel- working (Fig. 14). 
 Each panel of the mine is completely isolated from the" contiguous 
 districts by barrier pillars, and is ventilated separately by delivery 
 to it of a portion of the volume of the intake which does service 
 in that panel, to be afterwards discharged into the return air- 
 way, where it rejoins the exhaust from the other districts. The 
 electric distribution for purposes of illumination and the water- 
 supply of a town are conducted on identically the same prin-
 
 448 MANUAL OF MINIXG. 
 
 ciple, i.e., that which recognizes the tendency of a fluid to seek 
 a shorter and easier escape from confinement. With a number 
 of conduits receiving at a common point a volume of fluid from 
 a larger conductor, each will convey a fractional amount of that 
 original bulk which is inversely proportional to the resistance 
 offered by its entire rubbing-surface. If the several conduits 
 again meet to discharge their individual volumes of fluid at a 
 common point into a common reservoir, the pressure at the point 
 of discharge is the same at the mouth of each and every pipe. 
 Likewise the pressure at the point of union is the same in the ends 
 of each and every pipe. The loss of head or of pressure due to 
 the flow of the given quantity of fluid through each conduit is 
 then the same. If the original bulk is allowed naturally to sub- 
 divide, the amount of fluid in the several branches will vary in an 
 inverse ratio with the cross-sectional area of their conduits. This 
 is equally true of the circulation of air through mine-galleries or 
 districts, of water through branching pipes, or of electricity 
 through connecting wires in the circuit. 
 
 The Process of Balancing the Resistances. In planning the 
 ventilating system for a mine which is to be divided into a 
 number of districts for ventilation purposes, the practice is to 
 calculate for each separate district its aggregate resistance to 
 the flow of the volume required for a known number of men 
 employed there, and for a dilution of the gases evolved in that 
 district. Several separate values for p are thus obtained. But 
 these district resistances must be equalized or else the inlet- 
 current will be so subdivided at the point of distribution that the 
 large bulk of the air will pass through that district which offers 
 the least resistance ; while to that district offering the greater resist- 
 ance the volume there circulating will be small. This is usually 
 the reverse of the requirements; for, generally, that district offer- 
 ing the smaller resistance is the shorter one, having less men in its 
 circuit, and therefore requiring a smaller volume of air; while 
 that district presenting the greater resistance to the flow of the 
 current is either more extensive, has a greater volume of goaves, 
 or contains more working places, and hence demands a very
 
 DISTRIBUTION OF THE AIR. 449 
 
 large fractional part of the main current. In order, then, to 
 automatically deliver to that district requiring more air, which 
 at the same time offers a greater resistance, the area of the con- 
 duit throughout its course or the area of the orifice at the central 
 point of distribution must be made sufficiently large to tempt 
 through it, or into it, the requisite amount of air, leaving to the 
 smaller district, which requires less air, an orifice of entry which 
 is comparatively small. By so doing the differences in pressure 
 for each and every district, between the point at which the splits of 
 the fresh-air current are made and the point at which the return 
 currents from the same splits reunite, can be equalized to that 
 of the one offering the greatest resistance. Then the current 
 will naturally divide according to the areas of the inlets or of 
 the passages, and each district will receive its apportioned frac- 
 tion of the incoming air. Hence, whenever the ventilation of 
 the mine is to be split into several currents and the air is to be 
 apportioned in accordance with the demand, the mine foreman, 
 having calculated the relative values for the head lost in each, 
 determines by proportion, as will be seen, the comparative area 
 of inlet to be provided the several districts at the point of dis- 
 tribution, and there, by means of doors and other regulators, 
 furnish to the given district the area desired. 
 
 The measurements of the water-gauge pressure, or loss in head 
 between the beginning and the end of the split and the velocity 
 of the flow of air, are made in the intake; and while it is not 
 always possible to subdivide the current at a common point 
 of distribution, this should be done as near to the downcast as 
 circumstances will admit. The same may be said of the point 
 of reunion. Otherwise the resistance of the intermediate ways 
 and of the entries must be determined and provided for, as may 
 be seen in the example given below. The aggregate resistance 
 of the intermediate ways of the several districts through which 
 the air is circulated determines the maximum number of splits 
 which are possible. 
 
 Simple as is the theory, and satisfactory and economical 
 as is the plan when well developed, it is not easy of execution-
 
 45 MANUAL OF MINING. 
 
 The success of the plan involves an exact manipulation and 
 great skill in taking due precautions to balance the various factors, 
 to determine the equilibrium designed, and to prevent one panel 
 or district from receiving too brisk a current at the expense of 
 others. Hence, while it is eminently desirable to apply this 
 theoretical distribution, its difficulty is recognized, and it has 
 become the practice of the foreman to approximate the desired 
 conditions by making repeated tests upon the quantity of air 
 flowing in a given circuit, which, if insufficient, is provided for 
 by enlarging the inlet area for the given district and watching its 
 reaction upon the volumes in the other dependent splits. In 
 shallow workings, though the mine may be extensive, the practice 
 is an inexact one in many cases. It may often be cheaper to 
 sink a new shaft to furnish separate ventilation to a district than 
 it would be to undertake to furnish an elaborate system of splits. 
 The Power Required for Ventilation. Though it may not 
 require a demonstration to show that the subdivided splits of 
 the current are productive of greater economy in ventilating 
 power, attention will be called to the fact that the ventilating 
 force in horse-power, necessary to deliver a volume, Q, against a 
 mine resistance, p, is measured by the expression 
 
 H.P. = Q/-J- 33,000 = Qm-j- 6365.7. 
 
 The resistance which would be offered by the aggregate 
 of all the districts to the flow of the entire volume, Q, through 
 the whole length of the circuit is measured by p. Each frac- 
 tional volume, q, q f , q', etc., passing through only one branch 
 of the circuit would offer a resistance r, r f , r" , etc., which is very 
 small compared with p. When the mine boss has adjusted the 
 regulator doors at the point of distribution by altering the respec- 
 tive areas of inlets, the resistances in all of the several circuits 
 are equalized, the work performed in each split is qr, q f r, q"r y etc., 
 and the aggregate ventilating power is their sum. As 
 
 q+q f +q"+. . . =Q and r<p,
 
 DISTRIBUTION OF THE AIR. 451 
 
 the power required for the ventilation in branches will be less 
 than that for a single current, (), through the same passages. 
 
 The power required for 16,200 cubic feet of air flowing in one 
 column would be capable of producing 70,884 cubic feet of air in 
 five splits, 94,850 cubic feet in ten splits, and nearly 100,000 cubic 
 feet in fifteen splits. 
 
 EXAMPLES. i. A colliery is ventilated by a Guibal fan of 21 feet 3 inches 
 diameter, making 40 rev&lutions per minute. How many cubic feet will it 
 produce? The air must pass through a main airway 300 feet long, 6'Xi2', 
 before being split up into three separate airways, one being 12,000 feet long, 
 6'Xs'; another, n,:oo feet, of area 6X7'; while the third is 10,000 feet 
 long and 5'X8' in section. Required also the water-gauge pressure, assum- 
 ing the two shafts together to consume 0.226 Ib. per square foot in .riction. 
 
 16,380 cubic feet and 0.9 inch. 
 The splits are all drawn from a common point of junction. 
 
 Theoretically, the fan produces a water-gauge pressure of 0.902 inch 
 (4.677 Ibs.). Then the entire mine offers a resistance of 3.504 Ibs. (0.676 
 inch). The resistance of the main airway is p 0.00000000062 79(2 2 . Q, 
 the quantity of air delivered, is divided up into three separate sections, accord- 
 ing to their resistances. As p is the same for each, the quantities q may be 
 known in terms of p to be 
 
 2i7iV7, 3455^, and 3368\ / ^ 
 
 Now the resistance of the entire mine is equal to 3.504 Ibs., plus that of 
 the main airway, plus that of the split , p. From this we know ^=3.504 
 0.00000000062 79(8994^) 2 , whence p becomes 3.333 and the r sistance of 
 the main airway 0.170 Ib. Q then becomes 16,380 cubic feet, and the quanti- 
 ties received by the three splits are 3940, 6270, and 6110 cubic feet. (The 
 difference in results arises from failing to carry out the decimals beyond two 
 places.) 
 
 2. An airway 3000 feet long, 8'X4' area, is carrying 20,000 cubic feet. 
 How many cubic feet would be delivered if the air was split into three currents, 
 the power remaining the same? The sections are 3000 feet long and 8'X 4' 
 area; 3600 feet and s'Xg'; and 4800 feet 6'Xio'. 
 
 51,736 cubic feet and 11.56 H.P. 
 
 The calculated power necessary to drive the quantity of air, Q, through the 
 three sections is equal to the sum of the three powers, pav, of each section. 
 
 The benefits that may be derived from splitting the air- 
 current are manifest by inspection of the following case:
 
 45 2 MANUAL OF MINING. 
 
 EXAMPLES. i. A mine has two slope entries, 9'Xi4' in cross-section and 
 100 feet long, and such internal resistances as would be equivalent to 8000 feet 
 of a typical airway 5'Xio' in cross-section. What pressure and power will 
 be requisite to propel 16,200 cubic feet? 
 
 ^=0.2619 and u=4243 ft.-lbs., for the two entries, and 
 ^=11.19 and = 181,383 for the total. 
 
 2. Required the quantities of air that will circulate where there are 
 2, 3, 5, 10, and 15 equal splits, the pressure remaining the same as above. 
 
 After calculating the pressure p for the one current as above, then proceed 
 to ascertain the pressure p' necessary to circulate 16,200 cubic feet in the 
 several cases. These will be found 1.369, 0.405, 0.0874, 0.0109, an d 0.0032 
 Ib. per square foot, respectively; as the area of each airway is 50 sq. ft., the 
 aggregate for the equal splits are 100, 150, 250, 500, and 750 square feet in 
 the several cases, while .he rubbing-surfaces, Im, are the same (240,000 
 quare feet). The pressures are then apportioned directly to 10,935 Ibs., 
 the mine friction of one current as above. Whence, the pressures being as 
 the squares of the volumes circulating, we obtain 33,409 cubic feet, 66,354 
 cubic feet, 91,692 cubic feet, 103,755 cubic feet, and 105,255 cubic feet, 
 and 373,887 ft.-lbs., 742,504 ft.-lbs., 1,026,030 ft.-lbs., i, 161,010 ft.-lbs., 
 and 1,177,760 ft.-lbs. as the respective powers . 
 
 3. If it be desired to know what quantities will circulate with the same 
 power M, as in Ex. 2, then we have but to apportion the volumes to the 
 cube roots of the powers, . 
 
 Thus ^373,887 : 33,409: :i8i,383 126,252, the volume with 2 splits. 
 
 ^742,504 : 66,754: :3/iSi, 383 : 41, 479. " " " 3 " 
 
 "ii/ 1, 026,030: 91,692: :^ // i8i,383:5i,46i, " " " 5 " 
 
 3/i, 161,010: 103, 55: :3/ 181,383: 55,881, " " " 10 " 
 
 ^1,177, 760: 105,255 ::^/i8i,38 3 : 56,419, " " " 15 " 
 
 4. When, however, he splits are not taken from a common point of 
 functure, the procedure for ascertaining the mine resistances, and, subse- 
 quently, for balancing the delivery of air to the several sections, is not so 
 simple. The plan consists in determining the several resistances and the 
 powe- necessary to overcome them. These are then added as fo'lows: In 
 Fig. 163 D is the downcast shaft, 84^ feet deep, 8'Xio' in cross-section, 
 delivering 56,000 cubic feet per minute; by force-fan, 27,750 cubic feet go to 
 the left gangway, while the right gangwav passes 28.250 cubic feet of a?r. 
 The water-gauge stands at f inch (4.525 Ibs.). Required the volumes of air 
 received by the splits A, B, C, and D.
 
 DISTRIBUTION OF THE AIR. 453 
 
 The distances along the gangway at which the splits are taken are, a, 460 
 feet from d; b, 960 feet; and c, 1360 feet from d. Dimensions of gangway 
 6'Xi2'. 
 
 The splits are E, receiving 5000 cubic feet through 100 feet of 6'Xia' 
 gangway; 60 feet of 4'X2 f break-through; and 60 feet of return airway 
 
 c b a 
 
 FIG. 163. Diagram of a Split Air-current. 
 
 5^X14'; C, having a resistance equivalent to 2401 feet of typical roadway 
 15' X 6.87', connecting with the return-way at c v 941 feet from the upcast U; 
 B, which has 220 feet of gateway 4'X;', 1400 feet of entry 4'X8', and 1000 
 feet of room 4 / Xi2 / , delivering at b v a point 581 feet from U; and A, which 
 has a resistance equivalent to 1810 feet of typical roadway 5'Xio', connect- 
 ing with the return airway 313 feet from U. 
 
 The upcast shaft is 870 feet deep and 12 feet in diameter. It was found 
 that the velocity of the outgoing air was nearly 708 feet. This would cor- 
 respond to about 80,000 cubic feet of air. 
 
 In the return airway, along ua l , the volume was found to be 41,610 feet, 
 This shows an increase in volume of 13.870 cubic feet, which may be accounted 
 for by the higher temperature of the outtake air or increments to the original 
 volume from gas. 
 
 Beginning at E, where 5000 cubic fee. are circulating, we find the total 
 resistance to b such as requires a pressure of 0.096 Ib. per square foot at c, 
 to overcome it. Hence the same pressure must exist at the mouth of C, 
 wher? pa= 0.000000219677" Equating the two values for p (q being unknown 
 for C, and 5000 cubic feet for E) we find that the volume of air of 6410 cubic 
 feet offers a proportionate resistance to that in E. 
 
 At the point b the total resistances of E, C, and their gangways must be 
 proportional to those o f B, a,-*, the relative volumes to be circulated. At b the 
 11,410 cubic feet passing out produces a pressure of 0.57 Ib. per square.foot, at 
 which rate, by equating this with the sum of the resistances of the split B, we 
 find that q becomes 3720 cubic feet. 
 
 Tn like manner 12.580 cubic feet will be found Capable of passing through 
 A, the frictional resistance at a being 1.19 Ibs. per square foot. 
 
 The total resistances pa of the various sections are: da, 53.33 Ibs. A, 74.59 
 Ibs.; ab, 17.29 Ibs.; B, 16.04 Ibs.; be, 7.83 Ibs.; C, 9 Ibs.; E, 6.02 Ibs.; c,6,, 
 17.67 Ibs.; fejflj, 23.29 Ibs.; a,, 91.2 Ibs.; D, 323.85 Ibs., and U, 356.2 Ibs.
 
 454 MANUAL OF MINING. 
 
 The volumes of air are, respectively, 27,740 cubic feet; 12,580 cubic feet; 
 15,150 cubic feet; 3720 cubic feet; 11,410 cubic feet; 5910 cubic feet; 5000 
 cubic feet; 17,115 cubic feet; 22,725 cubic feet; 41,610 cubic feet; 80,000 
 cubic feet; and 56,000 cubic feet. 
 
 The pressure mentioned is only that necessary to overcome the rubbing 
 friction of the moving air against the sides. No cognizance has been taken of 
 the pressure requisite to air the mine. That power will depend upon the 
 velocity, and is equal to about 0.17 Ib. pressure per foot. 
 
 The power requisite to force the air into and out of the mine (11.9 Ibs. per 
 square foot) is 666,400 ft. -Ibs. 
 
 5. An airway 5'X8' across, 6000 feet long, is followed by a through 
 2'X5' in cross-section 300 feet long. To drive 10,000 cubic feet of air through 
 them requires 5.5 horse-power. Their frictional resistances are equivalent 
 to 5.2893 Ibs. and 9.114 Ibs. per square foot, respectively; the total resistances, 
 pa, being 211.57 and 91.14 Ibs. A pressure of 302.71 Ibs. is therefore requisite 
 at the inlet to overcome frictions of the passage. 302.71 -^40 X 10,000= 75,667 
 ft. -Ibs. to propel the air through the first section, in addition to which it requires 
 9.114X10,000 Ibs. to overcome the resistance of the second. 
 
 When the "through" precedes the airway, we have no differ nee in fric- 
 tional resistances, but the total pressure in the first section must be such as to 
 overcome the total resistance of the two, hence it is 302.71 Ibs. (30.27 Ibs. per 
 square foot). Then 30.27X10,000 + 5.289X10,000=355.601 ft.-lbs.= 10.7 
 horse-power. 
 
 If the "through" is midway along the 6ooo-foot airway and has 300 feet 
 of the 5' X 8' roadway on each side, we still have the same resistances, yet the 
 power requisite for propulsion is 299,047 ft.-lbs. For each end the friction is 
 2.645 Ibs. per square foot, and for the middle it is still 9.114 Ibs. per square 
 foot. The total pressure of th? first is 105.784 Ibs.; of the second, 91.14 Ibs.; 
 and of the last, 105.784 Ibs. The pressure at the inlet is 302.71 Ibs. (7.5677 
 Ibs. per square foot); at the entrance to the second, 196.924 Ibs. (19.6924 Ibs. 
 per square foot); and at the entrance to the last, 105.784 Ibs. (2.645 Ibs. 
 per square foot). 
 
 Goaves. These are the most dangerous places. Here the 
 waste "rock, broken coal, slate, dust, pyrites, etc., accumulate in 
 the presence of water which, with the stagnant air, induce a com- 
 bustion from which sulphuretted hydrogen and other gaseous 
 products exude. It is estimated that the air-space in a goaf is one 
 sixth of the volume of coal extracted, and in it most likely will 
 breed a great deal of gas, of which the sweating of the roof is 
 an infallible sign. Often a water-gauge placed in a goaf-stopping
 
 DISTRIBUTION OF THE AIR. 455 
 
 will indicate by the difference of level in its arms whether or not 
 any accumulation of gas exists behind the stopping. Spontaneous 
 combustion once begun therein, nothing will stop it. For this 
 reason, though aeration is possible, fiery coals should only be 
 worked by a method involving complete removal of the coal, or 
 its replacement by clean waste. 
 
 The Velocity of the Air-current. Spacious air-drifts will 
 then reduce the resistances and also the velocity of flow. It is 
 not desirable that the velocity exceed 500 feet per minute, nor is 
 it comfortable or safe. A speed exceeding 500 feet is equally 
 injurious with stagnation. Several mining commissions of 
 Europe have experimentally determined that, aside from the 
 chilling effect of walking against so rapid and cool a breeze, 
 lamps are not safe: a rapid current incites explosion by driving 
 the gases through, or the flame against, the screens. Many 
 lamps can resist higher-speed currents, but none are safe in 
 over 900 feet per minute. In the old country the air-currents 
 at different parts of the mines vary in velocity at the coal face 
 often as fast as 900 feet; but here our Davy, Stephenson, or 
 Clanny lamps require protection in such a velocity, which exceeds 
 that of American practice. 
 
 The Anemometer. A miner can approximately estimate the 
 speed of the current by knowing the rate at which he must walk 
 to keep the flame erect, or by noting the time elapsing between 
 the discharge of a volatile fluid or smoke and the time of its 
 arrival at a point a known distance beyond. The anemometer 
 is, however, much the simpler instrument for measuring the 
 velocity of the flowing air. In a case is a series of vanes which 
 are moved by the current, and these by proper gearing turn 
 indices over their respective dials at such a rate that the velocity 
 may be at once read. It does not give accurate results on ac- 
 count of the friction of its mechanism. Each instrument has 
 its own factor, which is not even constant. Biram's and Cos- 
 tello's patterns (Figs. 164 and 165) are most used in America. 
 Their factors are ascertained by occasional test, the anemom-
 
 456 
 
 MANUAL OF MINING. 
 
 eter being revolved on a whirling table, and its reading com- 
 pared with the actual velocity of revolution. 
 
 The point selected for observing the velocity should be in a 
 straight gallery, whose sides and roof are a fair average in rough- 
 FIG. 164. FIG. 165. 
 
 Anemometers. 
 
 ness, and where there is neither a sudden bulge nor a contraction. 
 The average of several one-minute readings is taken at the place 
 of measurement, near the roof, sides, and centrally. Then the 
 cross-section of the conduit at the observing station is taken. 
 
 The Efficiency of a Ventilating System. One object in deter- 
 mining the velocity of the air circulating in the passage is to 
 ascertain the efficiency of the motors supplying the current. 
 For this purpose the cross- sectional area of the airway at the 
 observing station is measured, and the product of this with the 
 velocity and the ventilating pressure at the given point deter- 
 mines the actual work of the effective power in the air at that 
 point. The power U to move the air is o.oooo303-yaP. The 
 ratio between the result of these observations and the simultane- 
 ous indicator diagrams of the engine gives the mechanical effi- 
 ciency of the ventilating system. The ultimate comparisons of 
 this work in terms of consumption of coal should show for a 
 good system not more than n Ibs. of fuel per useful horse-power
 
 DISTRIBUTION OF THE AIR. 457 
 
 produced by the fan and 40 Ibs. to 70 Ibs. per effective horse- 
 power of the furnace. 
 
 The Regulation of the Multiple Circuits. The air entering 
 the mine is conducted to the men by a route as direct as the 
 plan of the workings admits. To those who are at work in 
 the shaft and at the breast of the exploration entries, the current 
 is led by a partition in the shaft or drift-forming conduits for 
 an intake and return current; to the men engaged in extract- 
 ing coal, the current passes from the main entry to the head- 
 ing, thence through cross-headings, or branch headings, to the 
 rooms. The air-current passing down the entries by either 
 shaft or slope is split at each level or lift to a right and a left 
 branch. In each lift the current flows to the furthermost point, 
 along the main heading, thence rising into the most remote 
 room, along its face, thence through ''dog-holes," progressing 
 toward the hoistway till each room has been traversed, when the 
 current is led to the return or back heading which communi- 
 cates to the upcast way, there to be joined by the return current 
 from the opposite side of the lift, and finally to unite with the 
 return current rising from the lower lifts which have been fed in 
 like manner. 
 
 The main heading is usually the haulage-way, the back head- 
 ing serving only for the vitiated air. It may be used for a travel- 
 ling way. The jaw of each room, except that one which is most 
 advanced, is stopped up, which latter, in time, is closed as an- 
 other room is turned from the heading. As each room must be 
 in free communication with the haulage-way for the delivery of 
 its coal, the stopping there provided must be capable of opening 
 for the passage of the coal. The swinging doors, therefore, 
 which are placed in the entry are always closed to prevent the 
 air from returning at once down the back heading. Here doors 
 are built to deflect it into the rooms, with also a brattice in each 
 room, to direct the current it receives from its neighbor. From 
 the last room, whether working or abandoned, the air passes to 
 the return current. In the lower portions of the room not in 
 the direct sweep of the air-current the air is prevented from
 
 4S 8 MANUAL OF MINING. 
 
 becoming stagnant by the leakage from the main current and 
 the addition of such volume as is swept in or out by the travel of 
 the cars from and to the haulage-way. 
 
 The distribution in the long wall is perfectly similar, in which 
 case the gob roads are stopped up by doors to convey the current 
 to the extremities of the workings, whence the air flows along 
 the long- wall face. 
 
 In flat seams, the double entries of which have a low grade 
 for haulage from the surface, the similar split system is em- 
 ployed, the splits being taken at the point of union with the main 
 headings. When, by reason of the low pitch, three entries are 
 driven from daylight, the common practice consists in making the 
 central entry the return airway, the air being led into the mine 
 in two currents on either side. This simplifies the subdivision 
 of the currents, which may be effected with few doors. Each 
 district on either side will have but one crossing of its return 
 current with the intake current. 
 
 Structural Appliances Required for Controlling the Air- 
 current. Their object is to confine and direct the air-currents 
 in a given path. They may be permanent or temporary in char- 
 acter; they may also serve as stoppings to prevent any circula- 
 tion through a given way, or as a means of deflecting the current 
 into some other channel. 
 
 In the split system of ventilation several small currents radi- 
 ating from one given centre of distribution, or branching from 
 a main airway, are returned, after having done service, to an- 
 other airway, whence they are carried to the surface. Provision 
 must be made for carrying the return current over or under the 
 intake, where they intersect for regulating the amount of air 
 delivered to each branch of the mine; and for flexible stop- 
 pings along the gangway, admitting of the passage of trains of 
 cars. 
 
 The Means for Deflecting the Current. These are of seven 
 varieties: 
 
 Air-bridges, constructed to carry the return current over 01 
 under the airway.
 
 DISTRIBUTION OF THiL A.ii. 459 
 
 Stoppings, or walls, closing the "throughs" when their ser- 
 vice as connectors has been performed. 
 
 Partitions of wood or canvas to temporarily divide each air- 
 way into two compartments for an inlet to the face and an out- 
 let to the cross-heading nearest to the face, which in turn con- 
 nects with the return air, back heading, or adjoining room. 
 
 Brattices of canvas hung from the roof in the common 
 travelling-way to serve as a light temporary resistance to the 
 flow of air, without impeding the access of the men into the 
 rooms. 
 
 Swinging doors in the gob roads and at the jaws of the rooms 
 to prevent the entry of air from the main heading, but to allow 
 admission to the men. 
 
 Swinging doors at the return heading at appropriate places 
 to check the immediate return of the current and to deflect it 
 to a room. 
 
 Sliding doors placed in the branch airways and fixed so as to 
 offer an area for the inlets to several branches, including the con- 
 tinuation of the mainway. 
 
 The Stoppings. These consist of walls strongly built to 
 resist the pressure of the roof and floor and also to prevent the 
 leakage of air. They are of brick, with their faces set a little 
 back from the wall of the gallery in order to provide for any 
 peeling of the coal from the side of the airways. These walls are 
 built on either side of the airway two or three bricks thick, and 
 often filled with rubble to make a firm pillar and support. Oc- 
 casionally their exposed faces are plastered with coal-tar or 
 cement to make them impervious They are built about the 
 goaves and other gaseous districts. 
 
 The temporary stoppings are of wood, or canvas sheets, more 
 or less firmly framed, and may be movable, as are doors, or fixed, 
 as in the case with curtains and panels. 
 
 Air-crossings or Bridges. These are placed at the intersec- 
 tion of each pair of cross-entries to allow the pure air from the 
 intake to enter the mine without interference from the vitiated 
 air in its exit. These bridges are carried over the roadways and
 
 460 MANUAL OF MINING. 
 
 dispense with the use of doors. They are permanent structures 
 whose position is determined by the plan of the mine. 
 
 Overcast. There are two classes of air-bridges, one in which 
 the return airway is carried over the main haulage- way and intake, 
 and the other in which it may pass below the intake. Though 
 there are many conditions favoring the undercast crossing the air- 
 bridge prevails (Fig. 166). In this case, at some distance before 
 the point of intersection of the two airways is reached, an excava- 
 tion is made in the roof sufficiently high and wide to afford an 
 airway of ample dimensions, which excavation is carried above 
 the intake airway the desired distance and down on the other 
 side to the main level of the coal-seam at some distance beyond. 
 The coal close to the sides of the intake airway is left undis- 
 turbed. A bridge is then built consisting of a heavy flooring of 
 timber laid above the opening to make a roof for the intake and 
 a floor for the intake airway. It would be cheaper to remove 
 all the coal on the sides and erect stopping-walls to furnish sup- 
 port for the bridge overhead, but these are not always air-tight. 
 It is absolutely essential that the partitions between the return 
 and the intake airway should be air-tight. The cost of an over- 
 
 FIG. 1 66. A Rock-tunnel Air -crossing. 
 
 cast is about $125.00, with a month's labor of common help 
 and ten days of two masons. There would be required perhaps 
 100 bushels of mortar, 5000 bricks, and 6 cu. yds. of sand. 
 
 Many engineers prefer to drive the overcast through the 
 solid rock to insure a tight airway and one which is not liable
 
 DISTRIBUTION OF THE AIR. 
 
 461 
 
 to be influenced or affected by the accidents in the intake. These 
 are more expensive to build, but have many advantages. 
 
 The Undercast. This air- crossing is built below the floor (Fig. 
 167) of requisite dimensions and far enough beyond the intake on 
 
 
 FIG. 167. An Undercast Air-crossing. 
 
 either side to leave a secure and undisturbed pillar for a stopping 
 in the latter. It is framed and boxed as is the overcast. In a 
 few cases this construction may be cheaper than the overcast, 
 and is preferred when the roof cannot well be disturbed. But 
 the liability to undermine the coal-pillars which rest upon the 
 floor of the seam and the risk of creep constitute a serious objec- 
 tion to it. The depression which is made in the floor furnishes 
 opportunity for lodgement of water and mine-gas and may be- 
 come in time absolutely impassable. It is always subject to 
 constant flooding. 
 
 How Should Air-bridges be Made? It is a disputed question 
 whether air-bridges should be carried through the solid rock 
 or be built in the roof. Both serve equally well their pur- 
 poses under general conditions. But the effect of an explosion 
 upon the two is different. The rock terminal airway remains 
 intact and the ventilation of the mine is unimpaired. The 
 boxed bridge- crossing is usually destroyed by the force of the 
 blast, and with it the air-current. Of course the force of the 
 blast is confined to the airways, and its violence may produce 
 heavy falls which will afterwards present obstacles to the rescuing 
 party, when the work of rehabilitation begins, though the number
 
 462 MANUAL OF MINING. 
 
 of such violent explosions that will shatter the roof will be 
 small. 
 
 The timber overcast bridge furnishes a safety-valve for the 
 force of the explosion and prevents much damage. In such 
 event the reestablishment of the air-current after an explosion 
 becomes a difficult matter. The wooden partitions may be 
 destroyed by a moderate explosion, and the number of occasions 
 when the rescuing party must build bridges and stoppings is 
 larger than those in which it must reopen a blocked airway 
 through a solid overcast. 
 
 A trap-door in the floor of an overcast which would raise with 
 the explosion and fall again with the heavy reaction that follows 
 would enable the air-current to be reestablished and meet the 
 objection to the air-bridge. Such a door could be built cheaply 
 and can be made tight for ordinary emergencies. 
 
 The greater expense of the rock tunnel would result in a 
 smaller number of them being driven, and hence less split air 
 currents. As the economic distribution of air depends upon 
 the use of numerous air-currents, there would be a greater econ- 
 omy secured with the same outlay by an abundant use of the 
 cheaper air-bridges than would be obtained for a similar number 
 of solid overcasts. 
 
 The mining laws of Pennsylvania require overcasts and air- 
 bridges of masonry or in solid rock, in all mines generating fire- 
 damp in sufficient quantities to be detectible by the ordinary 
 safety-lamps. 
 
 An air-bridge passing over a slope is always built in the solid. 
 
 Mine-doors. The doors are of two classes those provided 
 with an additional valve or gate, which may be opened or closed 
 within a limited range; and those doors swinging on hinges, 
 which may be opened for purposes of ingress or egress; they 
 should open against the direction of a possible inlet current, 
 in order to completely direct the current along the main way. 
 Doors are the main dependence for the ventilation of mines,, 
 particularly those which are subdivided into districts for pur- 
 poses of ventilation. They are placed in such position as will
 
 DISTRIBUTION OF THE AIR. 463 
 
 temporarily and effectually check and deflect the current. 
 Wherever their location, they must be built with great care, of 
 matched timber, closely fitted in their frame, and are maintained 
 only so long as the draught is to be kept up. Some are even 
 provided with weather-strips on their edges. They are placed 
 in all the side entries to the rooms, swinging outward towards the 
 main passageway, and are generally in pairs when located in 
 the haulage-way. In the latter case they are located far 
 enough apart that the two will not be open at the same time, 
 and thus interrupt the principal circulation. 
 
 Regulator-doors are employed (Fig. 168) at the mouth of entry 
 to a ventilating district in the mine, the gate of which constitutes 
 an adjustable sliding door capable of being secured against 
 disturbances, and are employed to regulate the supply of air 
 to be delivered to that district in accordance with the system 
 described above. The placing of the slides is left to the mine- 
 boss or the fire-boss. Those doors which are placed in haulage- 
 ways either require an attendant or have their frames so inclined 
 that the door swings to of itself. Automatic mechanical appli- 
 ances for opening and closing the doors from a distance without 
 stopping the mules or other haulage-motors are employed in 
 many collieries, but do not stand in great favor. While there 
 seems to be no means yet offered for replacing doors or dis- 
 pensing with them entirely, there is no question as to their 
 objectionable nature. They are leaky, and offer opportunity 
 for negligent drivers, who by leaving them open divert the current 
 from its proper course or stop it entirely. The most objection- 
 able feature is their liability to destruction by explosion, and the 
 consequent annihilation of the current at the most critical 
 time. 
 
 The Effect of an Obstruction in the Airway. Any irregularity 
 in the sectional area of the airways or any fall of rock restricts 
 the passage of the air. So the movement of the cages in the 
 shaft will disturb the current, and the same occurs with- the pas- 
 sage of trains. Their effect is to increase the mine resistance, 
 which will be apparent at the fan in an increased speed, even
 
 464 MANUAL OF MINING. 
 
 though the quantity of air being delivered may be less. If doors 
 are left open in the mine, the reverse takes place. 
 
 When the number of ventilating circuits is large, the obstruc- 
 tion in one of the splits will have little effect on the fan or fur- 
 nace as revealed by the water-gauge. The greater the number 
 of splits, the less will be its influence. But a redistribution takes 
 place throughout the mine, and all the other splits will receive 
 an undue proportion of air. 
 
 The regulator-doors are then reset until the obstruction has 
 been removed. The effect of placing a regulator in any district 
 cannot be calculated without taking into consideration the whole 
 circumstances of the mine. Any change in the position of a 
 shutter affects the other airways. 
 
 Precisely the same formulae will be employed for the equiva- 
 lent orifice of the mine as will apply to mine regulators, the 
 coefficient for the vena contracta being taken at 0.65. 
 
 FIG. 1 68. A Regulator Slide-door. 
 
 Safety-doors. In some collieries an excellent device has been 
 introduced. This consists of a sheet-iron door hinged and 
 suspended from the roof at appropriate places, to be released 
 when the explosion occurs, drop, and close the opening, thus 
 replacing those disturbed by the accident and maintaining the 
 direction of the current at this critical time.
 
 DISTRIBUTION OF THE AIR. 465 
 
 Brattices are used as temporary expedients for subdividing 
 any room or airway in such a manner as to carry an inlet current 
 through the wider compartment to the breast or face of the work, 
 and return it in the smaller section to the return airway. Brat- 
 tices may be made of planks nailed on props at suitable distances 
 apart, with the interstices between the plank lathed, and the 
 whole tarred or calked with oakum to constitute an air-tight 
 partition, or the brattices may be made of canvas unrolled hori- 
 zontally and suspended from the roof and frequently adjusted 
 upon posts, according to the ventilating pressure in the airway 
 or room in which they are placed. To make canvas imperme- 
 able to air, it is usually soaked with tar, though its stench has 
 resulted in its substitution by an incombustible material, such as 
 asbestos or a soluble silicate. 
 
 Aid to Ventilation. Other methods of providing the return- 
 or inlet-way consist in cutting a ditch in the floor of a gallery and 
 boarding it over; in providing a "top sollar," the timbered roof 
 of the gangway having a little extra space for the passage of the 
 air-current above it instead of below it; in laying wooden air- 
 boxes, metal-pipes, or large canvas hose at one side for tempo- 
 rary expedients. 
 
 In the rooms, entries, and shafts the distance between the 
 working place and the last connection with the airway must not 
 exceed 22 yards, and that from the end of the brattice must not 
 exceed one fifth of that distance. 
 
 "Throughs" or "dog-holes" are driven at every 30 yards 
 through the pillar coal to connect rooms, and to allow of air 
 circulation through them in turn as the room advances. 
 
 For intensifying the permanent air-current led to a working 
 place, or for the separate ventilation of workings in seams with 
 slight disengagement of gas, compressed air may be employed 
 or hand-worked fans may be used for separate ventilation, which 
 is always brought up so near to the working place that the air 
 be not too much diffused. 
 
 EXAMPLE. A downcast 6' X 1 1', an upcast 6' X 12', 300 feet deep, supply air 
 to a mine having a gangway 3000 feet long, 50 feet sectional area
 
 4^6 MANUAL OF MINING. 
 
 five splits receive each a portion of the 40,000 cubic feet moving. Required 
 the several amounts delivered to the panels through the resistance of 6'X8', 
 700 feet long; 5'Xf, looofeet; s'X8', 1200 feet; 6'X6', 750 feet; and 4' X f, 
 800 feet. 
 
 By substitution it will be found that the pressure required to overcome the 
 downcast resistance is 1.23 Ibs. per square foot. In the mine the pressure 
 will be the same throughout; hence for each split 
 
 which, solved for each division, gives volumes q= i6,o59V^; go^o'^/J; 9679 
 V/>; io,9O9V^T; and 7543 Vp t respectively. The sum of these equals the 
 total quantity, Q, from which we get p, equal to 0.87 Ib. Assuming that the 
 splits are made as near as possible to the downcast entry and reunion to 
 upcast, which by the way is advised, there need be no further allowance for 
 resistances other than those of sudden turns or contractions. The distribu- 
 tion of the air. q, for each district is 12,070, 6785, 7275, 8200, and 5670 cubic 
 feet. 
 
 The upcast offers a resistance to the exhaust-air, the volume of which is 
 greater than the 40,000 cubic feet, because of accretions from "blowers," 
 moisture, etc. Disregarding these increments, the volume to be exhausted is 
 carried up at a velocity of 555 feet per minute, whence p is i.oo Ib. 
 
 - The total pressure to be imparted by increasing the downcast barometric or 
 rarefying the upcast is therefore 3.01 Ibs., requiring M of 57 feet, or a dif- 
 ference in temperature of about 100 at B= 25 inches. If the splits are made at 
 stated distances along the main gangway, an allowance must be made for each 
 of the several losses of friction in the various lengths thereof, remembering 
 that each branch split reduces the volume passing through the remaining por- 
 tion of the gallery, and correspondingly the friction therein. It would require 
 3.7 H.P. to do the work upon the air; a fan of 43 per cent efficiency would 
 necessitate a 9 H.P. engine. 
 
 The calculation for the ventilation of a railroad tunnel is similar. Assume 
 a tunnel 4961 feet long; sectional area, 336 square feet. Then 4961X336= 
 1,666,900 cubic feet of air to be changed every ten minutes. Velocity of cur- 
 rent, 496 feet. 
 
 p=- - = 4.8 Ibs. per square foot. 
 
 A fan 20' X 6' at forty revolutions easily meets this demand. The fan may be 
 applied, with or without brattice, at either end of the tunnel, but this is a 
 delicate matter. It would be better to place the fan at the mouth of, or a
 
 DISTRIBUTION OF THE AIR. 467 
 
 furnace at the bottom of, one of the connecting shafts used,during construc- 
 tion, and block the others off. 
 
 Required the pressure and power to get 10,000 cubic feet of air through air- 
 ways aggregating 6300 feet in length, of which 5000 feet are along a heading 
 5'X8' in area, with 1000 feet of airway 6'Xg' at the intake end and 300 feet 
 of a 2'X5' air-course at the upcast end. 
 
 REFERENCES. 
 
 Mine Doors, Coll. Eng., July 1886, 280; Balance Doors and Air Hinges, 
 Coll. Eng., April 1896, 198; Stoppings on Main Roads, E. Wain, Fed. Inst. 
 M. E., VI; Doors, M. & M. J., Vol. XX, 272. 
 
 Overcasts, Illustrated, Mine Insp., Pa., 1881, 74; Breakthroughs, Neces- 
 sity for, Ohio Mine Inst., 1892, 18; Check Curtain, Mine Insp., Ky., 1891, 61. 
 
 Improvements in Ventilation, M. & M., Vol. XXIV, 312; Air Column 
 in Ventilation, M. & M., Vol. XX, 165. 
 
 Resistances of Air-currents in Mines, Trans. M. & M. Eng., XLV, 62; 
 Resistance of Air-currents, T. L. Ewen, Inst. M. E., X; The Effect of an 
 Obstruction in the Airway, T. L. Ewen, Inst. M. E., IX; Ventilating, Splitting 
 Air-currents, Door, and Valve, Pa. Mine Insp., 889, 318. 
 
 Saftey-lamp Glasses, Coll. Guard., LXXIX, 369; Experiments in Safety 
 Lamps, Coll. Guard., Vol. LXXIX, 218; Acetylene Mining Lamp, ColL 
 Guard., Vol. LXXXII, 1127; Detecting Lamp Leakage, Coll. Guard., Vol. 
 LXXXIII, 126; Apparatus for Testing Lamps, Coll. Guard., Vol. LXXXVI, 
 842; Tests of Safety Lamps, Coll. Guard., Vol. LXXXVI, 261.
 
 CHAPTER XV. 
 
 THE ILLUMINATION OF MINES. 
 
 Illumination in Mines. In an atmosphere free from explo- 
 sive gas, the illumination may be had by any form of naked 
 light. In all metal-mines candles are used, and, occasionally, 
 the torch and kerosene-lamp. In bituminous mines known to 
 be non-gaseous the latter is employed; but in all mines which 
 are at all likely to develop gases the lamp-flame must be pro- 
 tected from direct contact with the white or fire-damp. The 
 simplicity of the means of illuminating metal-mines tends to 
 promote the comfort and safety of those engaged in them as the 
 rays of light from the open candle or lamp are shed in all direc- 
 tions, thus revealing dangerous roofs, sides, or timbering much 
 better than any form of safety-lamp can. But only in very shal- 
 low collieries should this be permitted, and then only after rigid 
 inspection by the fire-boss. 
 
 Candles and Oil-lamps. The candles which are used in 
 metal-mines are usually of stearic acid, of which Proctor and 
 Gamble's are the most uniform and will best withstand the tem- 
 perature of the heated atmosphere. They are cheaper illu- 
 minators than lamps in rooms and stopes, but not in haulage- 
 ways. The consumption averages three candles per man per 
 shift. 
 
 The candle may be placed in a loop sewn in the cap or the 
 pointed shank may be stuck into the timber, roadside, or floor, 
 where it is shielded from the current of air. With a proper sup- 
 ply of pure air it burns without unpleasant smoke emission, and 
 in places where there is a certain proportion of carbonic acid gas 
 in the air, which would surely cause its extinction if left undis- 
 
 468
 
 THE ILLUMINATION OF MINES. 469 
 
 turbed, it may be kept alight by spreading out the wick so as to 
 provide spaces between its cotton threads, and inclining the 
 candle for the tallow to feed the flame faster. 
 
 The common tin lamp with the hinged lid on top and a hook 
 and spout on either side from the spout the wicking projects and 
 is warmed is a more brilliant illuminator, and is also used 
 in coal-mines, giving a moderate light of about four candle- 
 power, with, however, the objection that it smokes. 
 
 At many collieries the pit-bank and the main roads from the 
 pit-bottom are lighted by electricity, or by gas, manufactured in 
 the colliery yard, but where this is not so, larger and more pow- 
 erful lights than those used at the face are necessary. The 
 common arrangement is a can holding about 3 to 6 pints of oil, 
 with one, and often two large wicks, having a tightly fitting lid 
 and a chain carrying a pricker, for use on the burning wick 
 when required. 
 
 Illuminants. Vegetable or animal lamp-oils were at first 
 used in safety-lamps, but now mineral oil is commonly used. 
 But mineral oils are volatile per se without decomposition. 
 Hence, even with heavy mineral oils having a flashing-point 
 high enough to warrant their use in safety- lamps, great care is 
 necessary in their storage and in the filling of lamps. Mineral 
 oils are more perfectly fluid, more combustible, and give a better 
 light than vegetable or animal oils, but unless there is a good 
 supply of air to feed the flame they are apt, owing to the large 
 amount of carbon they contain, to deposit much soot. Kerosene 
 requires an admixture of a less volatile oil. 
 
 The two staple illuminants for safety-lamps are best refined 
 rape- (colza-oil) and seal-oil. These give off no vapor under a 
 comparatively high femperature. The highest value of these 
 oils is only obtained from careful preparation, and cheap produc- 
 tions are to be studiously avoided. White lard, winter-strained 
 oil, is also much used, the consumption being one half gallon per 
 month for each lamp. In some mines a mixture of equal parts 
 of seal-oil and petroleum seems best to meet the requirements 
 of a good illumination with a minimum of smoke. In a mine
 
 47 MANUAL OF MINING. 
 
 using 260 duplex-wick lamps the annual expense for oil, repairs, 
 interest, etc., is $504. In selecting oil for illuminating pur- 
 poses, its behavior is tested not only from the standpoint of its 
 usefulness as an illuminant, but also that of its ability to burn 
 without smoke. When the oil bums and the combustion is 
 perfect, a blue non-luminous heating- flame is produced; but 
 when the conditions are such that the flame is cooled during 
 combustion or receives a deficiency of oxygen, the combustion 
 is imperfect, and the portion of the carbon in the oil is rendered 
 incandescent, thus emitting light. When the oil is very dense, 
 the amount of incandescent carbon released becomes excessive, 
 particularly in the presence of a small amount of oxygen, and 
 soot is the result. The ideal oil, therefore, should furnish a 
 maximum of light and a minimum of soot, with sufficient com- 
 bustion to produce draught. A simple test and a decisive one 
 may easily be made for the fitness of oil for use in the miner's 
 lamp by burning it, under the ordinary conditions, in a common 
 house lamp with a short chimney. The mixtures, which are 
 often used, of mineral oil with animal and vegetable oil are always 
 objectionable because of the almost unendurable odor, which 
 itself is detrimental to good air. There is little saving in their 
 employment, and they are worse than the unadulterated oil. 
 The very volatile oils and spirits, like benzine, burn with a clear, 
 uniform flame, show an easily perceptible cap in the presence 
 of gas, and are usually very sensitive, being also free from dan- 
 ger in a well-constructed lamp, even in the hands of an unskilled 
 miner.' 
 
 The Davy Safety-lamp. Much ingenuity has been ex- 
 pended in the endeavor to invent a safer means of illuminating 
 workings than that offered by the naked flame. In 1815 Davy 
 discovered that a sheet of iron-wire gauze was so good an ab- 
 sorbent of heat that the flame in contact with it could not readily 
 pass through. Further experiments indicated that for mining 
 purposes a mesh of 784 holes to the square inch was the safest, 
 and was therefore adopted as the standard. A cylinder of this 
 mesh, surrounding the light, surmounting an oil-lamp and
 
 THE ILLUMINATION OF MINES. 471 
 
 capped by a perforated top, is the form, which has been little 
 changed since Davy's time (Fig. 169). After the lamp is filled 
 
 FIG. 169. Davy Lamp FIG. 170. Clanny Lamp. 
 
 with oil and lighted, it is locked, to bar the miner against access 
 to the flame, the wick of which is trimmed by a wire passing 
 up through a close-fitting tube from the bottom. The combus- 
 tion is supported by air penetrating the gauze at all sides. 
 
 Sir Humphry Davy thus describes his invention: "The prin- 
 ciple of my lamp is that the flame, by being supplied with only 
 a limited quantity of air, should produce such a quantity of 
 azotic or carbonic acid gas as to prevent the explosion of fire- 
 damp, and which, from the nature of its operations, should be 
 rendered unable to communicate any explosion to the outer air." 
 . Defects of the Davy Lamp. This lamp has done and con- 
 tinues to do great service; but it has two defects. The first is 
 the liability of the gauze to become red-hot, and allow the flame 
 to pass through to the inflammable mixture outside. The sec- 
 ond objection is its low illuminating power. The open spaces 
 occupy only one fourth of the area of the gauze, through which 
 the light escapes horizontally; still less light gets out at the top, 
 to illumine the roof. Miners require light thrown in every dircc-
 
 47 2 MANUAL OF 111 'N INC. 
 
 tion, especially upward. The illumination of the roof is a more 
 important matter than the lighting of the working face. It is 
 also more difficult. The Davy lamp is inadequate for roof 
 inspection. These defects have been partially remedied in the 
 subsequent patterns by the use of glass, the only impermeable, 
 strong, though brittle, transparent substance. 
 
 The Clanny Lamp (Fig. 170) is a modification of the Davy, in 
 which a portion of the wire cloth is replaced by a short cylinder 
 of glass. This improves the illuminating power, and, if enclosed 
 in a tin can or shield, becomes quite safe in a gaseous mixture. 
 The majority of the later forms of lamps use a glass cylinder, 
 but an internal explosion, if it occurs, is rendered more violent than 
 in the Davy, which offers less obstruction to the escaping gases. 
 
 Stephenson's Lamp is almost as popular in America as are 
 the earlier forms, having a long cylinder of glass surrounded by 
 a wire gauze, bonneted above by perforated copper. The air- 
 feed is also through the gauze, passing underneath and into the 
 chamber to the flame, thence out at the top, as usual. This plan 
 keeps both cylinder and gauze cool, and its relative security 
 rests essentially on the regularity of the draught, for if the inside 
 air becomes overheated the light goes out; so it must be sus- 
 pended properly. This is an English favorite. 
 
 The Marsaut Lamp is an improvement upon this form, and 
 stands a fair amount of tilting safely. With care, its glass cylin- 
 der will last three years before breaking. The Marsaut lamp in 
 many mines abroad is regarded as 'the most suitable one for the 
 working miner, its construction being simple and strong. As 
 an indicator of gas it is reliable, furnishing also a good light. 
 In 370 in use, the average consumption of rape-seed oil was 2 
 gallons per year. This lamp was brought very prominently 
 before the public by the Accidents in Mines Commission. A 
 great difficulty is experienced in relighting it, and, from the wind- 
 ing path pursued by the feed-air, proper circulation does not 
 take place until the lamp gets hot. 
 
 The Mueseler, a Belgian lamp, is like Dr. Clanny's, having 
 in addition a conical chimney centrally above the flame. It is
 
 THE ILLUMINATION OF MINES. 473 
 
 highly recommended in Europe, but must be carefully handled. 
 It does not burn well in "dampy" or slow currents. The bon- 
 neted Mueseler, an English improvement, is receiving the high- 
 est encomium for use in fiery mines and high velocity. 
 
 The Hepplewite-Gray Lamp admits air at the top, down 
 four tubes, and through an annular chamber above the oil vessel. 
 The only gauze employed is that covering the outlet and annular 
 inner chamber. A serious difficulty with it is its liability to be 
 extinguished when suddenly lowered. It undoubtedly gives 
 more useful illumination than any other lamp, and as an indica- 
 tor of gas undoubtedly ranks superior to all others except, pos- 
 sibly, the Pieler or Wolf varieties. All other forms with the 
 inlet above the glass will miss, say, 4 inches of gas lying imme- 
 diately against the roof, except when they are tilted very much, 
 and then there is great danger of their going out. Many lamps 
 are now constructed to take air, if desirable, from the top, like 
 the Gray, and thus also to detect thin layers of gas; but even 
 then they will not do it so rapidly. It is possible to put some 
 modern lamps into gas and take them out again without any 
 indication being given if the test is conducted hurriedly. This 
 is quite impossible with the Gray, as the flame immediately 
 '"spires" up. Owing also to the large amount of useful light 
 given by it and the way this is directed on the roof, in addition to 
 its delicate indications of gas, this lamp is preferred to all others 
 for use by deputies, firemen, timberers, and fire-bosses. 
 
 The Dick Patent Port-hole Lamp compels all the air enter- 
 ing the lamp to go immediately to the flame, thus losing no air, 
 and is capable -of burning in a stagnant atmosphere. The air 
 entering the lamp above the case passes through the gauze, 
 thence descends to the flame, while the products of combustion 
 arise inside the lamp, to be emitted through circular holes at the 
 top of the bonnet. The bonnet is made of a seamless steel tube, 
 and is light and strong. 
 
 The Clifford Lamp is a new one and has many excellent 
 points. The light given is good. A plentiful supply of air 
 enters the lamp. It illuminates the roof as well as the sides of
 
 474 MANUAL OF MINING. 
 
 the workings. It is not sensitive to tilting or violent movement. 
 The bonnet arrangements prevent the possibility of a current of 
 any velocity from impinging directly upon the gauze, and thereby 
 render the lamp safe in explosive currents. 
 
 The Beard-Mackie Lamp has a small brass disc supporting a 
 H-bent rod, across whose standards are strung platinum wire. 
 The height of the flame is made visible by rendering the wire 
 incandescent. The percentage of gas present is known from 
 the height of wire rendered luminous. As the wires can be 
 altered in position by the mine-boss, the warning line can be 
 regulated according to the risk to be run. The lamp is enclosed 
 in a wire gauze or in a bonneted glass cylinder. 
 
 The Woolf Benzine Safety-lamp is an emphatic departure 
 from the varieties above described, in that, first, it burns benzine 
 or naphtha; second, it contains a patent self- igniter capable of 
 relighting the lamp fifty times without opening; and, third, it 
 contains a locking device which it is impossible to open except 
 by the use of an exceedingly powerful magnet. This lamp, 
 because of the sensitiveness of its illumination, is a delicate de- 
 tector of gas, and has met with very ready acceptance throughout 
 coal-mining districts, there being possibly 80,000 in use in Ger- 
 many. It furnishes a good indication of the presence of gas 
 by the height of its flame. A percentage of light carburetted 
 hydrogen, varying from i to 5 per cent in the total mixture, will 
 increase the height of the flame from 2.6 to 5.8 cm. If the 
 flame be partially turned down, a still greater degree of sensi- 
 tiveness will be manifested. 
 
 The Requisites of a Safe Lamp. The lamp must be self- 
 contained, strong, portable, and not heavy, require little atten- 
 tion from the miner during twelve hours of sustained light, and 
 capable of being placed in any position, besides giving perfect insu- 
 lation from the fiery gas. The use of the safety-lamp is to secure 
 protection to the miners from explosive mixtures, whether by 
 allowing of safe examinations by the fireman before the shift 
 begins, or by preventing explosion from sudden eruptions or 
 ordinary discharges of gas during the shift. It should afford
 
 THE ILLUMINATION OF MINES. 475 
 
 good illumination of the roof during repairs to the latter, or dur- 
 ing ordinary operations of mining, without tempting the miner 
 to uncover the flame. The wire-mesh casing reduces the pho- 
 tometric value of the light to a low degree. 
 
 The illumination from any of these lamps is very feeble. It 
 is less in any direction than the horizontal. Of all the lamps 
 the Gray sends the best light upward. The candle-power, hori- 
 zontally, of the Roberts is highest about 18, and of the Clanny 
 the lowest nearly 6. On this account a lamp must be able to 
 be held tilted without extinguishment, and be unaffected by 
 violent oscillations. The conditions dictated by safety circum- 
 scribe the lines of attempted improvement in the degree of illu- 
 mination. For, with a large mesh, the lamp is incapable of pre- 
 venting high-speed gaseous currents from entering the con- 
 struction chamber and blowing the flame against the opposite 
 gauze, or forcing gas into contact with the flame. The safe 
 lamp must have a mesh impermeable to the high-speed currents. 
 Glass improves the illumination but introduces other dangers 
 risk of breaking, if unannealed, by cold-air current or dripping 
 water, and the difficulty of securing a good joint to the gauze 
 cylinder. 
 
 Of the forty-one explosions which occurred in a certain dis- 
 trict during 1896, in four cases the immediate cause of ignition 
 was referred to a naked light or to a deterioration of the safety- 
 lamp; in twenty-five, to the passing of the safety-lamp flame 
 against the gauze through some careless movement, too high a 
 speed, or "falls in." The remaining twelve were from shot; 
 fire, or other undetermined causes. 
 
 Designers of lamps have difficulties to overcome other than 
 these. They include the rough usage lamps receive in the mine, 
 the violent shaking and tilting they are subjected to, frequently 
 causing discharges from the oil-chambers; the dust finely sus- 
 pended in the air, or settling on them at the faces, making them 
 more explosive; oblique explosive currents at the face caused by 
 falling stone or coal, close to the lamps; and the grimy hands of
 
 47 6 MANUAL OF MINING. 
 
 the collier, which clog the gauzes and inlet feed-holes with dirt 
 and grease. 
 
 An absolutely safe lamp is therefore difficult to obtain, and 
 notwithstanding the various modifications, there is as yet none. 
 The lamp which cannot ignite in an explosive mixture outside 
 of it is yet to be invented. 
 
 The defects of the Marsaut lamp are at the metallic connec- 
 tion of the bonnet with the other parts. They heat up from the 
 impingement of the flame and gases. Gas moving at 50 feet per 
 second can penetrate the two-gauze Marsaut, but a three-gauze 
 pattern will resist the current. The only source of danger in 
 the Mueseler lamp, not common to the other types, is its strong 
 tendency to smoke; otherwise it is a most efficient lamp, show- 
 ing a bright and steady flame in the strongest current. The 
 Gray lamp presents the risk of the gas burning at the cylindrical 
 strip of gauze under the glass, which in low-speed currents heats 
 the lower edge of the glass strongly, and in high-speed currents 
 allows the heated gases to pass across against the glass on the 
 other side. The top of the lamp can be easily tampered with, 
 and the gauze at the outlet is liable to be obstructed by soot if 
 the flame should smoke. 
 
 It is worthy of note that the brass lamps are less bright than 
 the iron lamps of the same pattern. A round wick is not so 
 luminous as a broad, flat wick for a given oil consumption. Seal- 
 oil is photometrically better than rape-seed-oil illuminant. The 
 insufficient light of a safety-lamp, combined with the difficult 
 and trying conditions of the bonneted forms, is proving injuri- 
 ous to the eyesight of miners, which serious evil is growing. 
 Photophobia is rare where candles are used, or where the lamp 
 is hung behind the miner. 
 
 Bonneted Lamps. In order to be safe in the highest velocity 
 of air-currents within a given mine, the flame must be enclosed 
 not only in a wire gauze, but also in a more or less impermeable 
 hood or bonnet, while the inlet area for the feed-air must be 
 reduced to the smallest allowable dimensions. Many lamps 
 now exist which appear to resist, in a highly explosive atmos-
 
 THE ILLUMINATION OF MINES. 477 
 
 phere, current velocities up to 3000 feet per minute for a period 
 of several minutes; and the four lamps which were brought to 
 the attention of the Mines Accident Commission, which re- 
 ceived special attention for their security, illuminating power, and 
 simplicity of construction, were the H.-Gray, Marsaut, bonneted 
 Mueseler, and Thomas's modification of the bonneted Clanny. 
 
 The bonnet screens the gauze cylinder from the effects of 
 draughts that blow the flame through the meshes and set up 
 a fiery heat by the excessive air and gas that enter above the 
 flame of the wick. It limits the supply of air to that required 
 for the oil-flame only. Such bonneted lamps, whose flames are 
 protected from the direct effects of the strong ventilating current, 
 may be used with safety for illumination in mines producing 
 fire-damp. Even in dry, dusty mines also developing fire-damp 
 some of these lamps are safe, though not all, for many well- 
 authenticated cases of failure are recorded where the dust has 
 proven fine enough to pass through the gauze meshes, to be 
 reduced to the state of incandescence in the inner chamber. 
 Locked safety-lamps arc insisted upon in mines when cutting 
 through clay- veins in solid workings. The Hepplewite-Gray 
 and the bonneted Mueseler have the best resistance to explosive 
 currents of high velocity, and the South Side Committee report 
 the following relative speeds at which the respective lamps and 
 the air-current can safely pass: Davy, 360 feet per minute; 
 Clanny, 600 feet; Stephenson, 780; Mueseler, naked, 1200; 
 Mueseler, bonneted, 2400; Marsaut, in a can, 2440; and the 
 Davy, in a shield, 2400. The North of England Institute of 
 Mining Engineers gives the safe velocities at 720,540, and the 
 others higher. The British Royal Commissioners of Accidents 
 approved the Gray, Marsaut, and the bonneted varieties as safe at 
 high speeds. The 'common Davy or Geordie lamps are unreliable. 
 
 Locking Lamps. All safety-lamps have locks to them. The 
 oldest form of lock is still used on most lamps, notwithstanding 
 its acknowledged insecurity. It consists of a screw-bolt with 
 a square head which is turned by a key until it has entered a 
 hole bored in the oil-chamber, which is supposed to prevent
 
 478 MANUAL OF MINING. 
 
 it from being unscrewed until the screw-bolt is withdrawn. But 
 this may be done by a couple of nails filed to fit the bolt, by 
 pieces of wire bent for the purpose, or even by the point of a 
 pick. It affords a little more security than the original device of 
 the screw-lock. 
 
 The importance of locking the lamp so that its flame can- 
 not be exposed to the gas is manifest, as there are many temp- 
 tations to the miner to open it. Most of them can be, and 
 are, opened by easily extemporized mechanical means, while 
 others are rather more difficult to open. 
 
 All manner of permutation-locks and magnetized plates are 
 offered on the market, besides the lead-plug seal with which 
 the lamp is riveted after each filling. On account of its sim- 
 plicity and ease of treatment, as well as from the measure of 
 security it affords, the lead-pin is coming more and more into use 
 at collieries. The pin is moulded with a head at one end, and 
 fits openings in the two parts of the lamp to be held together. 
 When in place it is firmly riveted and punched with some device 
 at both ends. This forms the means of detection, if the lamp 
 should be wrongfully opened. 
 
 Magnetic locks are employed to advantage. Lamps so 
 fitted cannot be opened except with the aid of a powerful mag- 
 net. They resist all the efforts of the miner, short of wreckage 
 of the lamps. The Woolf lamp and the Craig & Bidder's 
 lamps are so fastened. 
 
 Extinguishing-locks. These are devised to remove all temp- 
 tation from the collier to unlock his lamp by so contriving it 
 that the process of unlocking also extinguishes the flame. 
 
 Lighting and Relighting Locked Lamps. A great many 
 lamps become extinguished at the working faces from a variety 
 of causes, and this has led to designs of lamps which may be 
 quickly relighted without their being unlocked. The designs 
 may be divided into two kinds, in one of which lamps may be 
 relighted anywhere by the user; the other, only by an authorized 
 person at properly appointed stations by the application of 
 electricity.
 
 THE ILLUMINATION OF MINES. 479 
 
 Woolf s lamp belongs to the first class, the relighting arrange- 
 ment being adapted for the volatile illuminant, benzoline, used 
 in it. The lamp contains a reel of paper on which fulminating 
 spots are placed at intervals of about a quarter of an inch, er.ch of 
 which can, in turn, be brought opposite the wick, and at the 
 same time be struck by a small spring-hammer, operated by a 
 button at the bottom of the lamp; the composition explodes 
 and ignites the benzoline vapor surrounding the wick. 
 
 Use of Lamps. Safety-lamps provided with the best form 
 of lock to prevent their being tampered with by ignorant or 
 reckless workmen, and thoroughly tested before given out, will, 
 if carefully used, afford considerable protection in the mine. No 
 lamp is safe unless kept in thorough repair, and any infraction 
 of rules regarding careful use should be severely punished. 
 
 Examining and Testing Lamps. The lamp-room is on the 
 surface, where it is safer than underground. The safety-lamps 
 are complicated in their construction, and the work of examin- 
 ing and repairing them in large collieries is considerable. As 
 it is difficult for the eye to detect defects in adjustment, there 
 should be provided some simple means for reliably testing the 
 joints between the glass cylinder and the gauze. It is well to 
 test each lamp, before it is taken into the pit, by placing it, lighted, 
 in an explosive mixture. 
 
 Cleaning Lamps. At small collieries having a few safety- 
 lamps in use the cleaning is usually done by hand, without 
 the aid of mechanical contrivances to unscrew or remove the 
 internal fittings; an ordinary hand-lamp brush is used to rub 
 the dirt from the gauzes. Even when the workmen clean their 
 own lamps there is at least one lamp-keeper at each colliery 
 to replenish the oil vessels, renew the wicks, and replace worn 
 washers or broken glasses, except in the very rare instances in 
 which the men themselves supply all safety-lamps and after- 
 wards maintain them in a perfect state of repair. Where oil 
 is burned the gauze should be steeped in a hot alkaline solution, 
 to free it of soot, etc. Lamps burning benzine are not clogged 
 with carbonaceous deposit.
 
 480 MANUAL OF MINING. 
 
 To avoid waste, manufacturers furnish automatic fillers 
 holding the exact quantity sufficient for one lamp. 
 
 Electric Lights in Mines. The advantages of electricity 
 for lighting about a mine are its decreased cost, better illumina- 
 tion, absolute reliability, and greater freedom from accident. 
 The lights outside are arc lamps; those inside are incandescent. 
 Arc lamps are either of the open type, or enclosed. The latter 
 give less light under similar conditions, but their distribution is 
 more even. The former require from 8 to 10 amperes, and the 
 latter 5 amperes of current. Though the efficiency of the latter 
 is less, their consumption of carbon and the expense of their 
 renewal are also less. 
 
 The lamps may be connected continuously in one series, in 
 which case the dynamo is series-wound. If the lamps are in 
 series, each one is provided with an automatic cut-out allowing 
 the others to receive the current in case of its failure. When 
 connected in several short circuits from the one main, "in paral- 
 lel," they will require a compound- wound machine. A resist- 
 ance is then put into series with each circuit in order to keep its 
 current at the proper pressure. The latter connection is more 
 common for mining work. 
 
 Each circuit is provided with a fuse which bears the same 
 relation to the installation that the safety-valve does to the boiler* 
 Fuses are made of alloys of tin or lead for the 5o-ampere circuits, 
 and of copper for heavier lines. 
 
 The light wires are protected from damp by being inserted 
 in cast-iron piping or lead tubing. A double-pole switch mounted 
 conveniently on a metal base, easily reached, opens and closes 
 the circuit for the lamps. Having two contacts, it divides the 
 current and the spark, and thus reduces the risk. 
 
 Lamps of 16 candle-power on a circuit of 100 volts require 
 0.6 ampere. On the 2 20- volt circuit; 0.27 ampere. These 
 correspond to 3.75 watts per candle-power. 
 
 12 lamps of 16 C.P. require i H.P. 
 
 6 " " 32 C.P. " i H.P. 
 
 2 arc lamps of 600 C.P. " i H.P. 
 2 " " " icoo C.P. " 1.7 H.P.
 
 THE ILLUMINATION OF MINES. 481 
 
 It is usual to allow 30 candle-power for every 100 sq. ft. of 
 floor area, when they are placed to advantage. A radius of 70 
 feet is allowed for each 5-ampere enclosed lamp or a lo-ampere 
 open lamp. 
 
 For parallel distribution of current to lamps, let w= watts 
 per candle-power; W= watts per lamp; = voltage at the 
 
 W 
 lamp terminals. Then / = p-=the current in amperes per 
 
 lamp. All wiring is based on an allowable drop, of which 5 per 
 cent may be considered a good average for the voltage loss in 
 this character of wiring. The area of a conductor to furnish 
 the current for a given number of lamps, N, is 
 
 10.81 X2DXI 10.81 X2DXN 
 A.= 
 
 drop in volts R Xper cent loss* 
 
 R is the heat resistance of one lamp. 
 
 EXAMPLE. It is desired to illuminate the grounds and buildings about 
 the tipple by 100 i6-candle-power lamps on a line aggregating 600 feet 
 from the generator, 30 lamps of 32 candle-power, each at a distance of 750 
 feet, and 6 arc lamps in pairs at a distance of 600 feet from the power. 
 Required the electric horse-power and the size of conductor, the voltage of 
 the incandescent lamps to be 200 and of the arc lamps, ico. 
 
 As each lamp requires 0.3 ampere for 16 candle-power, 0.6 ampere 
 for 32 candle-power, and 5 amperes for the arc lamp, the currents will be 30, 
 18, and 30 amperes, respectively. For the small lamps 600 feet of line must 
 carry 30 amperes. This will require the wire of No. 4 B. & S. gauge having 
 a resistance of 1.5 ohms per mile. The loss on this line will therefore be 
 45 volts per mile and 5 volts for the line. 
 
 The larger incandescent lamps will be on a cabb of No. 6 B. & S. gauge 
 having a resistance of 2.5 ohms per mile or loss of 43.5 volts per mile; and 
 6.2 volts for 750 feet. 
 
 The dynamo should deliver a current at its terminals of 2.06 volts pressure 
 and 78 amperes. This equals 16,068 watts, which at 85 per cent efficiency 
 correspond to 26-brake horse-power.
 
 4^2 MANUAL OF MINING. 
 
 REFERENCES. 
 
 Location of Regulators, M. & M., Vol. XXII, 501; Mine Doors, Geo. 
 Davidson, 111. Mine Insp., II, 177; Balanced Doors and Iron Air Bridges as 
 Preventatives of Mine Explosions, W. M. Morris, Coll. Eng., XVI, 198; Stop- 
 pings on Underground Roads, E. B. Wain, Fed. Inst. M. E., VI and VII; 
 Fire Doors for Mine Shafts, R. G. Brown, E. & M. Jour., Vol. LVII, 321; 
 Illuminating Oils in Mines, Ohio Mine Insp. Report, R. Haseltine, 1895, 46; 
 Lamp Oils, Am. Mfr., LVI, 918. 
 
 Safety Lamps, Comments by Mine Insp., Coll. Mgr., 1894, 24; Lamps: 
 Discussions, Remarks on Use of, Coll. Guard., Dec. n, 1896, 1117; Safety 
 Lamp, Development of, M. & M., Vol. XX, 39; Safety-lamp Flames, M. & 
 M., Vol. XX, 417; Wolff Lamp, M. & M., Vol. XXIII, 226; Eugene B. 
 Wilson, Amer. Inst. M. E., XIII, 129; E. J. Schmitz, Amer. Inst. M. E., XVI, 
 410; Overcasts, Robert Mauchline, Pa. Rep. of Min. Insp., 1881, 74; 1885, 
 Packet. 
 
 Electric Lighting for Mines, a Paper, Coll. Mgr., 1894, 85; How to Light 
 a Colliery with Electricity, S. F. Walker, Brit. Soc. Min. Stud., XIII, 147; 
 Electric Lamps in Coal Mines, E. & M. Jour., LIX, 316; Electric Lamps, 
 M. & M., Vol. XXII, 195. 
 
 Ventilation of Thin Workings, M. & M., Vol. XXII, 141; Ventilation of 
 Deep Workings, M. & M., Vol. XXII, 273.
 
 CHAPTER XVI. 
 
 ACCIDENTS IN MINES. 
 
 The Inherent Dangers in the occupation of mining are great. 
 The uncertainties in the overlying ground, the ever-lurking foe 
 in the form of gas, the as yet imperfect methods of illumination, 
 demand the greatest vigilance as the price of comparative safety. 
 Add to these the ever-present factor of human carelessness, and 
 the equation becomes one, containing two independent variables, 
 the solution of which is hard to find. To what extent the inher- 
 ent dangers could be overcome, were the factor of carelessness 
 eliminated, it is difficult to say. The German Government, out 
 of a total of 7933 lives lost in coal-mining, attributes 5179, or 
 65 per cent, to the purely inherent dangers of the occupation; 
 2397, or 30 per cent, to the carelessness of the victims; 248 to 
 the fault of fellow workmen, and 69, or 0.9 per cent, to defects 
 in working. Reports of mine inspectors attribute 60 per cent of 
 the deaths in mines of the United States to carelessness; the 
 40 per cent to unavoidable or unforeseen causes. 
 
 Comparative Hazard in Nations. Under this head it is 
 impossible to obtain absolutely accurate conclusions, owing to 
 the great difference in methods in vogue in different countries. 
 In Europe the record of accidents, injuries, and fatalities has 
 been much more elaborate and precise than in America, while 
 the classification of accidents varies in the different countries. 
 In France an accident is not entered on the records unless the 
 person is prevented from working for a period of twenty days; 
 in Austria it is the same; in Belgium eight days is the limit; in 
 Germany the law is still more strict, as it is necessary to give 
 notice of inability to work for any period exceeding three days. 
 
 483
 
 484 
 
 MANUAL OF MINING. 
 
 Every detail accompanies the inspectors' reports in Europe, 
 such as the age of victim, nationality, character of work, length 
 of time in service, exact location of accident in mine, nature of 
 injury, character and extent of medical attendance, length of 
 time incapacitated, number of days absent from work due to 
 injury. Even family relations are reported, as number depend- 
 ent upon the man, etc. In America inspectors' reports are as 
 a rule more general or indefinite, due to the different conditions 
 and consequent differing laws existing in the several States. As 
 yet we have no Federal requirements save those locally applied 
 in territories and reservations. 
 
 There are two methods of comparison of hazards: ist, classi- 
 fication by rate of accidents per number of tons of coal produced 
 annually; 2d, by rate per accident per number of employees 
 working underground, or by total of surface and underground 
 laborers. If the estimate be made on the first basis, allowance 
 must be made for the greatest efficiency of the American laborer, 
 due largely to the shallowness of our coal-seams and the exten- 
 sive use of machine cutters. In the table below can be seen the 
 productiveness of the American compared with the English and 
 the Belgian miner. 
 
 DEATH-RATES AND TONNAGE PER EMPLOYEE. 
 
 
 Tonnage 
 per Em- 
 ployee. 
 
 Tons of 
 Coal per 
 Death. 
 
 Rate 
 
 per 1000 
 Em- 
 ployees. 
 
 United States, 1893 t< 
 ft t 
 
 Ohio, " " 
 
 Austria, 1894 ' 
 Belgium, 1896 ' 
 France, 
 Great Britain, " ' 
 Prussia, ' ' ' 
 
 > 1902, b 
 
 1903 
 1902 
 
 Dth incl 
 < < 
 
 < 
 
 
 
 usive, bituminous 
 anthracite. 
 
 660 
 381 
 
 5I 1 
 
 326 
 
 178 
 207 
 314 
 338 
 
 221,091 
 126,795 
 292,596 
 199,526 
 "3,358 
 182,525 
 238,04? 
 150,000 
 
 
 ..89 
 5.81 
 .76 
 .718 
 
 .01 
 
 23 
 31 
 .16 
 
 
 
 
 
 The second method of classification is generally accepted as 
 the more reliable and is more readily understood. It also 
 affords a comparable standard of comparison with other 
 trades. The following table requires little explanation. It 
 represents conditions in the three prominent coal-producing
 
 ACCIDENTS IN MINES. 
 
 485 
 
 States of America. The utmost care has been taken to insure 
 the accuracy of all data used, and all statistics have been offi- 
 cially verified, as a thorough and comprehensive knowledge of 
 the dangers and fatal consequences in this occupation is a matter 
 of absolute necessity to those engaged in it. 
 
 
 Penns> 
 
 Ivania. 
 
 
 
 
 Anthracite. 
 
 Bituminous. 
 
 
 
 Tons of coal 
 Number of employees inside. . 
 Number of employees outside. 
 Fatalities, total 
 Serious injuries 
 
 36,9 1 1,^4 
 
 93,377 
 49,762 
 300 
 641 
 
 98,946,204 
 110,015 
 25.371 
 *45 
 861 
 
 30,021,300 
 
 4i,5i8 
 
 44,487 
 
 99 
 406 
 
 23,929,267 
 
 27,613 
 9808 
 81 
 298 
 
 Deaths per 1000 employees. . 
 Injuries per 1000 employees. 
 Number of days worked .... 
 Tons of coal per life lost 
 Annual output per employee. . 
 
 2.25 
 4-32 
 
 220 
 123,039 
 376 
 
 3-36 
 6.36 
 116 
 217,000 
 898. 
 
 2. I 
 
 8.8 
 210.4 
 
 303,245 
 724 
 
 2.1 
 
 7-9 
 176.4 
 293,818 
 867 
 
 * Includes an explosion involving 112 lives. 
 
 The Death-rate in Mines. During the earlier period of 
 mining the loss of life annually was at a rate of i death for every 
 200 persons employed. With time this rate is appreciably 
 reduced. In Great Britain, during 1903, it was i for every 688 
 employees. The average in the United States is now about 
 2.3 lives for each 1000 men employed, or i for every 435. During 
 the year 1902 the total mineral output of the United States re- 
 quired nearly 1,000,000 employees, of whom at least 2500 lost their 
 lives. In Great Britain, France, Belgium, and Prussia, during 
 the same year, coal-mining, in producing 478,459,088 tons of coal, 
 killed 1937 below and 306 at the surface. Metal- mines are more 
 hazardous than coal-mines, but in either the risk may be said 
 to be less than in other of the dangerous occupations. The annual 
 rate of death among railway brakemen during the past decade 
 was 12.5 deaths per thousand; that among trainmen, 9.8 per 
 1000. The highest rate in any coal district to date is 8.5 per 
 1000 ; the anthracite coal-miners is 5.6, and among the anthracite 
 laborers 4.6 per 1000. In the United States and Canada, for 
 the decade ending in 1902, there were -1,150 deaths, a rate of 2.Cj
 
 486 MANUAL OF MINING. 
 
 for each 1000 employees. In the entire railroad service during 
 the same time the average rate of death was 2.64 for each 1000 
 employees. Among the bituminous coal-miners the rate aver- 
 aged 2.2. The accident liability inside the mine is, of course, 
 greater than among the laborers at the surface. 
 
 Statutory Provisions Properly Governing Underground Opera- 
 tions, to provide safety to miners, require: A safe ingress and 
 exit, with ample means of communication between the bottom 
 and the top; the examination of the working places by a fore- 
 man prior to the entry of the men; copious volumes of air in 
 well-directed currents to dilute the accumulations of gas; secure 
 timbering and effective illumination of the working places; the 
 posting of copies of rules, advice, and precautions, printed in 
 several languages; official inspection ; an ample supply of timbers ; 
 large pillars to prevent caves of roof; supervision in the use of 
 explosives in each working place; boundary pillars between 
 gaseous sections of a mine and drowned or abandoned mines. 
 Faithfully executed, these provisions are ample to protect the 
 men from the extreme underground dangers. 
 
 It cannot be denied that the American death-rate is higher 
 than in European countries, nor does it show the strong down- 
 ward tendency that is apparent in their statistics. A fair ex- 
 ample may be the record of Great Britain for the semi-decade 
 since 1857. The annual death-list showed nearly icoo lives 
 sacrificed in 1852, and it is a striking fact that not many more 
 were lost in 1902, but the number of employees have increased 
 during that time from about 200,000 to 850,000. The improve- 
 ment in the conditions of English mines is also manifest by the 
 following summary, showing the tons of coal produced for each 
 life lost: 
 
 1852 59,8oo tons 
 
 1862 82,500 " 
 
 1872 112,000 " 
 
 1882 145,000 " 
 
 1892 200,000 " 
 
 1902 227,280 "
 
 ACCIDENTS IN MINES. 
 
 487 
 
 It is true the statistics of the States include only the coal- 
 mines, though it is improbable that the inclusion of the metal- 
 mines would bring the rate even to that of Great Britain, particu- 
 larly as there are many small coal-mines not subjected to official 
 inspection whose records are never kept. America is behind 
 the other nations in the efficiency of its safeguards, or, at least, 
 in the results. 
 
 AVERAGE NUMBER OF ACCIDENTS IN MINES IN THE UNITED KINGDOM IN 
 SEMI-DECADES. 
 
 
 Number of 
 Employees. 
 
 Total 
 Deaths. 
 
 Annual Death-rate per 1000 Employees. 
 
 Periods. 
 
 
 
 
 
 1 
 
 d 
 
 M 
 
 ' 
 
 5fii 
 
 it 
 
 II 
 
 
 Below. 
 
 Above. 
 
 gf 
 
 i 
 
 I 
 
 "o 3 
 
 1 
 
 
 " " ^"3 
 
 ^ 
 
 ~S 
 
 
 
 
 1 
 
 
 
 a 
 
 r*> 
 
 <fi 
 
 is 
 
 
 4%f* 
 
 -M d 
 
 
 
 
 
 
 W 
 
 & 
 
 a 
 
 i 
 
 H U 
 
 H 
 
 H 
 
 1851-55 
 
 182,427 
 
 47, 47 
 
 937 
 
 47 
 
 .280 
 
 .016 
 
 i. 206 
 
 0-556 
 
 5.149 
 
 .012 
 
 1-301 
 
 1856-60 
 
 208,763 
 
 53,832 
 
 964 
 
 53 
 
 234 
 
 .8 4 6 
 
 O.SQQ 
 
 0.648 
 
 4.628 
 
 0.994 
 
 3.883 
 
 i86i-6s 
 
 237,779 
 
 
 8 Q 8 
 
 68 
 
 .618 
 
 .714 
 
 0.668 
 
 0.790 
 
 3-791 
 
 .105 
 
 3.240 
 
 1866-70 
 
 269,813 
 
 69,574 
 
 1071 
 
 87 
 
 .158 
 
 -578 
 
 0.528 
 
 0.730 
 
 3-995 
 
 .256 
 
 3-433 
 
 1872-75 
 
 399,397 
 
 n,584 
 
 1066 
 
 09 
 
 .516 
 
 .210 
 
 o-437 
 
 0.572 
 
 2.736 
 
 0.800 
 
 2.342 
 
 1876-80 
 
 424,586 
 
 17,876 
 
 1147 
 
 100 
 
 .811 
 
 .132 
 
 o-3i7 
 
 0.440 
 
 2.709 
 
 0.847 
 
 2 . 306 
 
 1881-85 
 
 443,5 02 
 
 16,688 
 
 1025 
 
 99 
 
 .408 
 
 .108 
 
 o. 263 
 
 0.532 
 
 2.312 
 
 0.848 
 
 2 .007 
 
 1886-90 
 
 477> 6 33 
 
 26,654 
 
 975 
 
 117 
 
 0.312 
 
 I.OI5 
 
 o. 196 
 
 0-517 
 
 2.042 
 
 0.913 
 
 I. 806 
 
 1891-95 
 
 57^463 
 
 ^0,804 
 
 970 
 
 123 
 
 0.281 
 
 0.806 
 
 o . 194 
 
 0.934 
 
 1.706 
 
 0.822 
 
 L53 1 
 
 1896-03 
 
 618,507 
 
 50,404 
 
 896 
 
 125 
 
 0.17 
 
 0.764 
 
 0.144 
 
 o-457 
 
 1.451 
 
 O.SOI 
 
 1.381 
 
 Each American State reveals continuous improvement in 
 the underground hazard; but for the entire United States the 
 death-rate has not been reduced. The reason for this is that 
 new coal-fields and new mines are opened from time to time. 
 The new difficulties and the increase of unskilled labor naturally 
 increase the rate for the district. This is apparent in the new 
 Western coal regions, where the rate is greater than in the well- 
 explored Eastern districts. Indeed it is inordinately high. 
 
 Discipline in Mines. Unquestionably a more rigid discipline 
 in the operations of mines, a general education of the men, and a 
 wiser discrimination in the selection of workers is required, and 
 will contribute greatly to the safety of mining as an occupation. 
 It is true that many causes of accidents are unforeseen and un-
 
 4 S8 
 
 MANUAL OF MINING. 
 
 avoidable, but it is also true that the largest percentage of the 
 causes of accidents are those of a preventable nature, largely in 
 the hands of the miners themselves. If a code of rules were 
 officially issued dealing with the severe forms of mine disaster, 
 the first step would be taken toward greater security. Again, 
 the present laws err in not placing direct responsibility for defec- 
 tive conditions upon some one. If this were remedied, better 
 conditions would prevail. Not only should the employers be 
 
 ACCIDENT STATISTICS OF NORTH AMERICA, 1893-1902. 
 
 State. 
 
 Deaths. 
 
 Rate per 
 
 1000. 
 
 Tons Pro- 
 duced per 
 Death. 
 
 Tonnage 
 per 
 Employee. 
 
 
 353 
 4i3 
 8i3 
 J 45 
 239 
 
 212 
 177 
 
 118 
 
 71 
 135 
 I5 
 
 540 
 4344 
 2218 
 422 
 237 
 257 
 911 
 321 
 119 
 
 3.00 
 
 5-59 
 
 2.21 
 
 2-57 
 6.22 
 
 2-34 
 1.82 
 
 i-35 
 1.56 
 1.76 
 7.61 
 1.76 
 3.01 
 2.30 
 6.01 
 24-75 
 7-43 
 3-73 
 9-83 
 1.96 
 
 194,870 
 
 110,000 
 
 276,347 
 
 365,518 
 
 67,700 
 221,849 
 276,271 
 360,000 
 
 '61,366 
 
 218,000 
 
 83,362 
 
 292,596 
 126,795 
 287,601 
 71,118 
 32,236 
 69,232 
 184,413 
 
 68,138 
 
 219,266 
 
 500 
 
 592 
 607 
 
 938 
 421 
 5i8 
 378 
 485 
 957 
 383 
 634 
 5H 
 38i 
 660 
 437 
 764 
 5M 
 687 
 669 
 433 
 
 
 Illinois 
 
 
 
 
 
 Kentucky 
 
 Maryland 
 
 Missouri 
 
 New Mexico 
 
 Ohio 
 
 
 bituminous. ....... 
 
 Utah 
 
 
 West Virginia 
 
 British Columbia 
 
 Nova Scotia 
 
 Total North America 
 
 12,150 
 
 2.89 
 
 187,384 
 
 54i 
 
 fined for infringing or ignoring the laws, but the employees 
 should be penalized for hazarding life and property of others. 
 
 The Causes of Accident. The following prominent causes 
 of underground accidents are classified according to frequency 
 and number of lives involved. Falls of roof or sides of rock or 
 coal; accidents in the haulage- ways; injuries in shafts; prema- 
 ture explosions of powder and explosions of dust and gas. The 
 first three classes of accidents rarely cause more than one death
 
 ACCIDENTS IN MINES. 
 
 489 
 
 or a single injury at a time. Explosions of gas or eruptions 
 of water, fortunately rare, have involved hundreds. 
 
 Falls of Ground. This form of accident arises from the fall 
 of roof or of coal in working places and in roadways. About 
 70 per cent of all accidents arise in the rooms, and about 30 per 
 cent of the deaths occur in roadways. Each accident that occurs 
 involves but a single fatality or injury. Yet it takes the lead as 
 the most prolific cause of underground fatalities. The grave 
 danger lies in the fact that the accident is local and of continual 
 
 FATAL ACCIDENTS IN THE PRINCIPAL COAL-FIELDS OF NORTH AMERICA. 
 
 States: 1893 to 1902. 
 
 Aggregate 
 Number of 
 Persons. 
 
 Total Num- 
 ber of Lives 
 Lost. 
 
 Mortality 
 per 1000. 
 
 Tons of 
 Coal per 
 Death. 
 
 Pennsylvania, bituminous; Ohio, 
 Maryland 
 Indiana, Illinois, West Kentucky. . 
 Missouri, Iowa, Kansas, Indian 
 Territory 
 Tennessee, West Virignia, East 
 Kentucky, \labama. 
 
 1,316,608 
 456,341 
 
 303.I47 
 
 2,829 
 1,017 
 
 763 
 
 2lI I 
 2.26 
 
 2.52 
 
 i 70 
 
 296,658 
 293,874 
 
 170,000 
 16? 636 
 
 Colorado, New Mexico, Utah. . . . 
 Washington, British Columbia.. . . 
 Novia Scotia, 1887-96 
 
 97,296 
 67,263 
 60,716 
 
 755 
 578 
 119 
 
 7-77 
 8-59 
 1.96 
 
 79,536 
 68,245 
 219,266 
 
 Total North America, bituminous . 
 
 2,758,756 
 
 7,806 
 
 2-73 
 
 228,985 
 
 Pennsylvania, anthracite. ....... 
 
 1,443,110 
 
 4,344 
 
 3.01 
 
 126,795 
 
 Total North America 
 
 4,201,866 
 
 12,150 
 
 2.89 
 
 187,384 
 
 occurrence at a place where the men are otherwise engaged at 
 work, and ignore or fail to notice the warning which usually 
 precedes disaster. The weak spots in the roof, the various fossil 
 tree-trunks, horses, balls of ironstone, rocks, naturally creviced, 
 or shaken by the vibration of a neighboring blast, are liberated 
 with little warning. To these, with the coal scaling from the 
 sides or roof, the unnoticed yielding of pillars which are too thin 
 to support the rock, and the underholed coal, the accidents are 
 due. 
 
 Of the total fatalities in the anthracite region of the United 
 States during the past twenty years, 3521 were caused by fall
 
 49 
 
 MANUAL OF MINING. 
 
 of rock and roof in the working, or over 50 per cent of the total 
 deaths during that time. In bituminous regions the proportion 
 is equally large during the same years. The mine inspector for 
 Illinois reports for the twenty years prior to 1903, a death-list of 
 
 CLASSIFICATION or DEATHS FOR 1902. PRINCIPAL CAUSES. 
 
 Deaths from 
 
 Pennsylvania. 
 
 Illinois. 
 
 Ohio. 
 
 Ken- 
 tucky. 
 
 Bituminous. 
 
 Anthracite. 
 
 Deaths. 
 
 In- 
 juries. 
 
 Deaths. 
 
 juries. 
 
 Deaths. 
 
 . In- 
 juries. 
 
 Deaths. 
 
 . In - 
 
 juries. 
 
 Deaths. 
 
 Falls 
 Cars 
 
 223 
 47 
 126 
 8 
 14 
 
 437 
 239 
 
 20 
 41 
 36 
 
 Il6 
 42 
 
 20 
 
 55 
 
 2 5 6 
 
 *33 
 
 1? 
 
 i3 
 
 59 
 15 
 4 
 13 
 
 203 
 125 
 15 
 
 22 
 
 54 
 
 10 
 
 3 
 5 
 
 162 
 
 83 
 
 3 
 *3 
 
 5 
 5 
 
 6 
 
 Gas 
 Powder 
 
 'Surface 
 Totals. ... 
 
 456 
 
 861 
 
 300 
 
 641 
 
 99 
 
 4 06 
 
 81 
 
 298 
 
 19 
 
 THE STATISTICS OF ACCIDENTS IN THE PROMINENT COAL-PRODUCING COUNTRIES 
 OF EUROPE, AND Two AMERICAN STATES. 
 
 
 
 
 
 Death-rate of Persons. 
 
 fi 
 
 : 
 
 Countries. 
 
 Periods. 
 
 Persons 
 
 Tons of Coal. 
 
 
 II 
 
 1^1 
 
 
 
 ( 
 
 
 
 
 
 
 
 
 Q 2 
 
 Pow- 
 
 -j-a 
 
 ~" In ** 
 
 
 
 
 
 Palls. 
 
 Cars. 
 
 o. o 
 
 der. 
 
 I s 
 
 H 
 
 Austria 
 
 1894-03 
 
 125,303 
 
 367,088,336 
 
 0.282 
 
 .586 
 
 .310 
 
 .044 
 
 1.718 
 
 ',937 
 
 
 1896-03 
 
 136 880 
 
 
 
 
 
 
 
 i 786 
 
 France 
 
 1896-03 
 
 i68,6oq 
 
 243,671,927 
 
 
 
 
 
 
 1335 
 
 Great Britain 
 
 1896-04 
 
 774,9" 
 
 1,039,879,428 
 
 o 713 
 
 .I2S 
 
 .114 
 
 .042 
 
 3i 
 
 8,100 
 
 Prussia 
 
 , 896^3 
 
 443,995 
 
 1,049,557,783 
 
 0.98 
 
 .68 
 
 *3 
 
 . II 
 
 .16 
 
 6,948 
 
 Penn., anth j 
 
 1882-91 
 1892-01 
 
 1,059,526 
 1,423,607 
 
 351,058,927 
 455-941,943 
 
 '37 
 '4.1 
 
 .67 
 S9 
 
 .29 
 .27 
 
 .27 
 
 30 
 
 .22 
 . II 
 
 4,427 
 
 Penn., bit 
 
 1893-02 
 
 963,767 
 
 638,213,809 
 
 1.42 
 
 34 
 
 .24 
 
 .06 
 
 .26 
 
 2,184 
 
 Illinois 
 
 1893-02 
 
 368,607 
 
 223,645,612 
 
 1.28 
 
 .21 
 
 -38 
 
 .08 
 
 25 
 
 813 
 
 1392, of which nearly 50 per cent is due to falls. Doubtless all 
 over the United States the number of lives lost from this cause 
 bears a similar proportion to the total. The death-rate from falls 
 of roof and coal alone in the States is far more than is shown by 
 the European statistics for the five years ending with the year 
 1900: France, 0.58; Great Britain, 0.78; Prussia, 1.22; Illinois,
 
 ACCIDENTS IN MINES. 491 
 
 1.34; Pennsylvania (bituminous), 1.84; and Pennsylvania (an- 
 thracite), 2.1 1. In Great Britain 22,190 lives, or 46.5 per cent/ 
 were lost by falls since 1853. 
 
 The death-rate is large and is not decreasing proportionately 
 with other causes, for the conditions in the working places 
 depend entirely upon the miner himself, and here improvement 
 is not manifested. It is true that the pressure of the roof, and 
 the dislodgement of the fragments therefrom, cannot be indefi- 
 nitely resisted or avoided, but its movement can be detected and 
 checked by systematic timbering. The best means of detection 
 is the use, near the face, of wooden props, not 'iron, whose bend- 
 ing will give warning of disaster by buckling. A better illumina- 
 tion will reveal the condition of the roof, continual testing would 
 disclose its loose fragments, and a systematic timbering would 
 materially diminish the risk of its fall. The use of coal-cutting 
 machines would remove, the dangers of underholing coal; and 
 so would the enforcement of a specific order for the dismissal of 
 any employee failing to utilize the props. An accident is the 
 result of deliberate neglect, and the delinquent is more frequently 
 the old hand than the newcomer. 
 
 Rigid supervision and the appointment of a suitable timber- 
 boss to visit all the working places and regulate the timbering, 
 with power to punish carelessness, would be a most effective 
 remedial measure. In certain districts in France the rigid en- 
 forcement of regulations requiring systematic timbering imme- 
 diately upon advancing beyond the distance fixed by the regu- 
 lation has reduced the loss of life to about | man per 1000 
 from the previous rate of about i in 1000. In 1870-1879 it was 
 0.90 per 1000; to 1886, 0.24 per 1000; to 1890, 0.15 per 1000; 
 to 1892, 0.13 per 1000. In Great Britain the rate of decrease 
 during the same period from 1.20 to 0.764, though a decided 
 improvement, was not so marked. 
 
 Accidents Due to Cars, Shafts, etc. Second on the list of 
 accidents in all classes of mines are those occurring in the trav- 
 elling-ways. The men fall from, or are run over, by cars. Such 
 accidents are more numerous in coal-mines than in metal-mines,
 
 -492 MANUAL OF MIXING. 
 
 because the speed of the cars, or trains, and the number of men 
 employed are greater. Many casualties arise from jumping on 
 or off the trains while in motion. These are manifestly invited 
 by the victims themselves; other than these, the accidents along 
 the travelling-ways are those arising from runaway cars, or 
 the lack of clearance in haulage-ways. Numerous safety-niches 
 .along the line will afford the men some protection. Those 
 occurring in the shaft are largely averted by a security of the 
 hoisting appliances, an efficient system of signals, and the use 
 of overwinding and safety attachments. An average of 5 deaths 
 from overwinding occurred annually in Great Britain. Cases 
 in which men are caught between the cage and the shaft-timbers 
 -or are struck by material falling from above are purely accidental 
 and may be classed as inherent to the occupation. Those occur- 
 ring from the fainting of overheated men while being hoisted 
 from the shaft are not numerous. These can be prevented by 
 gates at the sides of the cages and hoods overhead. The Euro- 
 pean shafts are more fruitful of accident than the shallow Amer- 
 ican shafts. 
 
 Accidents from Use of Explosives. In metal-mines the per- 
 centage of injuries due to premature blasts is larger than in coal- 
 mines. They are more .frequently the result of carelessness 
 .arising from the use of iron bars while loading a charge, from 
 tampering with the metallic caps, handling powder near an 
 exposed flame, thawing out frozen dynamite, drilling into unex- 
 ploded cartridges, attempting to ignite too many blasts at one 
 time, and returning to the work too soon. These accidents are 
 manifestly local in their nature and occur either at the working- 
 face, or at the powder-house where the explosive is stored. The 
 victims themselves are the active cause. The comparatively 
 high death-rate in the anthracite coal-mines is attributed by 
 some inspectors to the great consumption of high explosives. 
 
 The use of electric cartridges and the prohibition of loose 
 powder for firing in dusty collieries would materially diminish 
 the number of casualties. Greater care in the selection of the 
 explosive and its use can readily be exercised by both the em-
 
 ACCIDENTS IN MINES. 493 
 
 ployers and the employees. In some mines blasting is assigned 
 to a special class of men. 
 
 Gas Explosions. No coal-mine is free from gas, and no col- 
 liery can be operated without illuminating and blasting ajents. 
 Every mine, therefore, is subject to ignition of a greater or less 
 volume of gas. The result may be a fire or an explosion, accord- 
 ing as the amount of air present be excessive or at the critical 
 percentage, and its effect will be proportional to the volume of gas 
 or the quantity of coal-dust present. All dry mines should be 
 ranked as fiery whether gaseous to a dangerous degree or not. 
 
 A steady flow of escaping gas into the regular working places 
 rarely results in dangerous conditions, for the volume of the air- 
 current there dilutes it to an innocuous degree. But when some 
 internal reservoir of gas bursts into the entries or abandoned 
 rooms, where the circulation is sluggish, or the air comparatively 
 small in volume, the mixture of gas and air may easily become 
 critical. This danger is imminent in the driving of galleries 
 far in advance of faces in the rooms or through a seam beyond a 
 fault where heavy volumes of gas may be liberated, particularly 
 if the ground slip or be soft. A blast or a flame in contact 
 with the mixture may ignite it and cause disaster. One gas 
 explosion involving 62 lives occurred within twenty-four hours 
 after the visit to a mine regarded by the chief inspector as "the 
 safest and best-conducted mine in the State," but caused by 
 striking a clay-seam. The ignition came from either a defect- 
 ive lamp or an open lamp. This is usually the direct cause. 
 Men will disregard the prohibitory notice of the fire-boss barring 
 them from a given working place, and enter. For one cause or 
 another they will uncover the flame of their lamp. The reports 
 for Great Britain show that out of the total of gas or dust explo- 
 sions about 70 per cent owe their injuries to the use of naked lights 
 or imperfect lamps; 18 per cent to shot-firing, and 14 per cent 
 to the accidental, or spontaneous, ignition of the mineral. This 
 form of accidents during the last twenty years has caused 1451 
 deaths, or 0.35 out of every 1000 coal-mines in Pennsylvania. 
 It was 9 per cent of the total fatalities. Explosions killed over
 
 494 MANUAL OF MINING. 
 
 10,000 men during the past fifty years, but the rate is decreasing 
 very much, as shown in Great Britain for the successive semi- 
 decades between 1852 and 1003. 
 
 R2medial Measures. The measures which will safeguard the 
 men from the consequence of sudden outbursts of gas include, 
 besides the dilution of the gas in wide galleries, the use of 
 numerous bore-holes to indicate an approach to a gaseous 
 mass. This practice is sharply criticised by many because of 
 the difficulty of boring in shattered, and consequently infested, 
 portions of the seam. Some mines dispose of the gas through 
 bore-holes, by which it is discharged to the surface and burned. 
 An increased volume of air, its better distribution to working- 
 faces, and the discontinuance of heavy blasts of explosives w ? hile 
 the men remain in the mine are diminishing the number of in- 
 juries and fatalities from explosions. There remains much to 
 be done by the universal employment of some bonneted lamp 
 which is safe and at the same time furnishes good light. 
 
 A number of deaths charged to explosions occur in the exits 
 of the mine where the fleeing victims have been overwhelmed by 
 after-damp. Rescuing parties find them generally in the intake. 
 Here it is that the greatest force of the concussion is exercised, 
 and along it too, the gases make their exit. The return airway 
 is unquestionably far safer than is the intake airway after an 
 explosion, particularly if solid air-crossings exist in the mine. 
 Many cou d have made their escape through that w r ay had they 
 been accustomed to this outlet. 
 
 The Theory of Coal-dust Explosions. The circumstances of 
 many explosions, particularly of those on a large scale, cannot 
 be explained fully by reference to gas alone. Those mines in 
 which explosions occur, and are known as fiery, are invariably 
 dry. These dry mines are dusty, containing a large quantity of 
 the mineral charcoal, "mother of coal," which is continually 
 afloat in the atmosphere. The explosions being more frequent 
 in dry mines and deep mines indicate the influence of the coal- 
 dust in extending and aggravating the danger. 
 
 Different fine dusts are inflammable and consequently dan-
 
 ACCIDENTS IN MINES. 495 
 
 gerous, according to their degree of fineness and chemical 
 composition. Lycopodium-powder, which is a modern vege- 
 table product resembling mineral charcoal, is highly explosive. 
 Hence it may be a cause, and even the sole cause, of the 
 explosion. It cannot be ignited except by direct contact with 
 an intensely hot flame. There is no probability that the ordinary 
 flame of a lamp has produced explosion from coal-dust alone, 
 nor has it been shown that the ord nary blown-out shot has 
 ignited coal-dust without the presence of some gas. The danger 
 of explosion does exist when both gas and coal-dust are present. 
 It is able to extend indefinitely the transmission of an explosive 
 flame and thus intensify the shock. Experiments have proven 
 that, without any floating dust, the flame from a blown-out shot 
 would not travel more than 25 feet, but that soot would convey 
 the flame to 200 feet The conclus on may, therefore, be drawn 
 that, though coal-dust alone may not be dangerous, in the pres- 
 ence of gas, even in small quantities, it becomes highly so. 
 
 A preventative remedy consists in laying the dust by sprink- 
 ling. A spray can be delivered into the air- current from pipes 
 i inch or 2 inches in diameter under pressure of 50 Ibs. per 
 square inch to moisten the air and materially assist in the decom- 
 position of the dust 
 
 Ankylostomiasis. This is a disease attacking miners in the 
 wet collieries, and though but recently discovered in Germany, 
 is receiving considerable attention from other governments 
 because of its infectious character, and energetic measures are 
 being taken to eradicate it Imported into Westphalia from 
 Hungary, in one district alone it has increased from 107 cases 
 in 1896 to 1400 cases in 1902, and during the year 1902, in another 
 district of Germany, 5.29 miners out of every 1000 were afflicted. 
 
 The disease resembles cholera or typhoid, the source of con- 
 tagion being faecal matter. It is treated by the medical men 
 as dangerous and the patients are quarantined. It is promoted 
 by the sprinkling of the mine with pit-water, and by the high 
 underground temperature. Inasmuch as dry mines of the 
 infected districts do not show so large a number of cases, un-
 
 496 MANUAL OF MINING. 
 
 doubtedly the discontinuance of watering for a time would pre- 
 vent the development of the ova. The obstacle to this method 
 is the increased danger from explosions from gas or dust. 
 The nature of the disease is not perfectly understood as yet, 
 and many miners have been treated two or three times without 
 the removal of the parasite. Its destruction in human excre- 
 ment is best effected by the use of fern extract (Extractum Filicis), 
 which proves a better remedy than the thymol used in England, 
 Dry-earth pails, one to every four workers, are distributed to 
 prevent dissemination. The mine owners are aware of the 
 gravity of the situation, and in many districts have discontinued 
 the watering to reduce the plague. That it will in time develop 
 in our American mines, among the foreign element coming from 
 the infected regions, there is little doubt, and careful watch should 
 be kept for its appearance. 
 
 Underground Fires. The most extensive conflagrations in 
 mines are those initiated by gas or powder explosions, com- 
 municated subsequently to timbers, canvas, or floating coal-dust. 
 Fires differ from explosions only in the degree of combustion, and 
 either may result in the other. The flame from a miner's lamp 
 or a spark from powder, in contact with any combustible, 
 are the elements producing explosions or fire, and the remedy 
 is to isolate one or all of them The exercise of care in 
 stables, pump-rooms, and storage-bins for oil, powder, and waste 
 would eliminate a grave danger. In these quarters electric 
 lights can well be provided and the floors sprinkled with clean 
 gravel or sand. In rooms or travelling- ways fires come from 
 defective lamps or exposed lights. 
 
 Fires from Blasting Agents. Powder is not liable to ignite fire- 
 damp because the temperature of its detonation is not over 1000 
 F. An explosive not properly confined to the hole in which the 
 execution is to be done will be blown out in the room and ignite 
 any combustible gas, if sufficient air be present. Blasting 
 agents which are not chemically perfect may also produce fire 
 and explosion. The unconsumed gases of the detonator are 
 projected into the workings, where they combine with the requisite
 
 ACCIDENTS IN MINES. 49 T 
 
 amount of oxygen to cause a second blast, under conditions- 
 dangerous to life and property. 
 
 The . employment of a better grade of explosive, the use of- 
 a powder free from flame, the ignition of the fuse without spark,, 
 the use of water cartridges for extinguishing the flame, the em- 
 ployment of specialists to whom the blast is entrusted, and the- 
 substitution of electric firing when the employees have left the 
 works, will lessen the risk. 
 
 Spontaneous Combustion. This is the slow process of burn- 
 ing where heat, combustibles, and the minimum amount of air' 
 are the elements involved, .and it frequently occurs in a goaf 
 and in abandoned rooms containing broken coal, dust, and rotting' 
 timber. Here the roof pressure compacts the material, develop- 
 ing an amount of heat sufficient to partially decompose the ever- 
 present pyrites. This incipient conflagration is furthered by 
 ingress of air. The amount of combustible consumed depends 
 upon the amount of air gaining access to it. Once started, a. 
 pressure ensues within the room or goaf, and "breathing" follows, 
 forcing gas out through the trenches in the bottom, at the roof, 
 or from cracks in the side walls and admitting air. At this- 
 point the sulphurous fumes give warning, and active measures, 
 must be taken to prevent fire or explosion, or both. But a 
 complete isolation from the mine is difficult and a perfectly 
 impermeable enclosure impossible, hence the failures attending; 
 the employment of these measures. The safer method of pro- 
 cedure would be the construction of a gas-tight bulkhead, or 
 dam, against the goaves, if they can be perfectly walled. If not, 
 their contents should be removed, though this can be done only 
 at great risk. 
 
 Spontaneous combustion can be avoided by keeping cool the- 
 combustible accumulations in the rooms or abandoned work- 
 ings by an active air- current. 
 
 Extinguishing Fires. Besides the remedies already sug- 
 gested for preventing their inception, the usual method of pre- 
 venting the spread of fire is to beat out the flames or drown 
 the fire with a flow of water, steam, or carbonic acid. Resort:
 
 49 8 MANUAL OF MINING. 
 
 must be had to the last method when the fire cannot be extin- 
 guished by cruder methods, or when it has attained such head- 
 way as not to be overcome by simple means. To be effective, 
 however, the heading must be absolutely gas-tight, not only in 
 the bulkheads but also in the upper strata. If they are porous or 
 the mine is shallow, air will enter from above to feed the flame. 
 The mine in that event must be abandoned. A fire at the Calu- 
 met and Hecla copper-mines, said to have been communicated 
 to the shaft-timbers by the friction of the hoist-rope rollers, was 
 extinguished by the liberal use of carbonic acid. After the surface 
 had been frozen to stop leaks in the shaft, carbonic acid was 
 injected and the fire brought under control. For the manufacture 
 of 3000 cubic feet of carbonic acid there were used 1200 gallons 
 of sulphuric acid and 4500 Ibs. of limestone. 
 
 Water is the simplest quencher of flames, provided it can be 
 made to reach all portions which are on fire. But it has happened 
 that into certain rooms the water above the foot of the shaft 
 has not reached the face because of the compression of the air 
 locked in them, which cannot escape. The fire, therefore, raged 
 above the water level until this air was consumed, and in some 
 cases broke out again as soon as the water was removed. 
 
 Steam has been injected into some burning rooms for quench- 
 ing fires, but it has not proven very acceptable. 
 
 Protective Measures. As a prevention against the excite- 
 ment and confusion arising at the time of an explosion or a fire, 
 the following disciplinary precautions may be exercised before 
 the accident: 
 
 i. Consider the quickest and safest mode of descending into 
 the mine when the usual winding arrangements are useless. 
 2. Plan for the installation of a special winding-engine. 3. Con- 
 nect water-pipes from the surface to the mine with branches 
 for use in case of fire. 4. Arrange for the fitting of an extra 
 engine and fan for the emergency. 5. Keep the tracings of 
 all working plans to within three months of work, showing all 
 roads then open, the position of overcast, doors, and brattices. 
 6. Accustom the men periodically to travel certain roads which
 
 ACCIDENTS IN MINES, 499 
 
 they are not in the habit of taking, having finger-boards showing 
 the direction of the upcast pit. 7. Keep on hand all safety 
 apparatus, including the Fleuss machine, a quantity of light air- 
 pipes, and "first-aid" appliances. 8. Appoint, during ordinary 
 working of the collieries, some of the leading officials to act as 
 emergency officers at the time of accident, drilling them in their 
 duties. 
 
 Rules for Guidance after Explosions. The following mode 
 of procedure would facilitate the work of the rescuing parties 
 as presented by Mr. Garforth. Instructions similar to these 
 should be posted about the mine, in various languages, for the 
 guidance of the men and the emergency officers. Each colliery 
 and mine should have a complete map of its workings and a 
 perfect outline indicating correctly the path or paths pursued 
 by the ventilating current in its circuit through the mine. It 
 is necessary when accident occurs as a guide for the rescuing 
 parties, and, although the mining laws do not require this pro- 
 vision, it would be well if maps were quickly available. 
 
 i. Send for the emergency officers, assign to each his duty, 
 and appoint some one as deputy in the event of serious accident 
 to the manager. 2. Examine all old connections with the shaft 
 and arrange to repair broken stoppings. Prepare stretchers and 
 stimulants and arrange for a hospital. 3. Provide exploration 
 parties of five with leaders, supplied with safety-lamps, mine- 
 plans, restoratives, cylinders of oxygen, and a stout cord. On 
 no account must any one enter alone, even on the shortest jour- 
 ney. 4. In advancing, the party should move in single file, the 
 leader of each search-party alone testing for gas. Do not let a 
 safety-lamp be the final guide as to the absence of after-damp. 
 
 5. Loss of life to explorers may, perhaps, be avoided by remem- 
 bering the dangers: (a) after-damp; (6) falls of roofs and sides; 
 (c) underground fires and consequent risk of a second explosion. 
 
 6. If the force of the explosion, has blown out the separation- 
 doors and overcast, they should not be restored because of the 
 possibility of undiscovered fire. 7. Main intake airways, blocked 
 by falls, must not be traversed unless the men carry with them
 
 5 oo 
 
 MANUAL OF :.II:-:I::G. 
 
 an unrolled brattice-cloth, which will admit of a double current of 
 air' through the airway. The brattice should be non-inflam- 
 mable. 8. To discover the existence of fire, restore the ventila- 
 tion and examine the return airways every hour for (a) fire- 
 
 FIG. 171. A Diving Knapsack. 
 
 stink and (ft) a rise in temperature. If the former be noticed, 
 that section from which it comes should be explored first, and 
 its fire extinguished if possible, t>r it should be closed off by stop- 
 pings, or, in extreme cases, the pit entirely closed. 9. Parties 
 should be careful not to go too far at once, even when taking air 
 with them, as the force of the explosion will have forced the
 
 ACCIDENTS IN MINES. 501 
 
 after-damp into the interstices of the goaves, whence it will 
 gradually exude. 
 
 Aerophores. For penetrating a very impure atmosphere 
 aerophores of different makes are to be had. They consist of a 
 portable bag or cylinder carrying enough compressed air or oxy- 
 gen for the respiration of a miner, and his lamp, while making 
 repairs or exploring. The oxygen is inhaled by one tube, while 
 through an exhaler is ejected the CO 2 , which is absorbed by 
 caustic soda, leaving the N only to return to the bag. Fleuss' 
 apparatus looks like a knapsack, weighs 28 Ibs., contains a four- 
 hours' supply of oxygen, and has besides a self-contained illumi- 
 nator an acetylene lamp does not depend upon the oxygen of 
 the air for its burning. A lamp burning methylated spirit heats 
 a plug of lime and renders it incandescent. 
 
 The Fleuss diving knapsack (Fig. 171) consists of a cylinder 
 and a cell in four compartments with a perforated false bottom. 
 The cylinder contains oxygen at 240 Ibs. pressure, and delivers 
 the gas to the nostrils by a tube. The carbonic acid gas is ex- 
 haled by the diver into the cell, where it is absorbed by caustic 
 soda. The entire combination carries a four-hour supply, and 
 has done excellent service to rescuing-parties after accidents 
 arising from fire and inbursts of water or floods of gas. 
 
 REFERENCES. 
 
 Safeguards against Danger from Mine Gases, Am. Mfr., Jan. 15, 1897, 83; 
 Miners' Phthisis, William Cullen, Jour. Chem. Met. Soc. of S. Africa, March 
 1903 and Feb. 1903; Gases in Metalliferous Mines, M. & M., Vol. XXIV, 
 478. 
 
 First Aid the Injured, M. & M., Vol. XXIII, 151. 
 
 Accident and Skeleton Maps, M. & M., Vol. XX, 115. 
 
 Damping the Air of Coal Mines as a Safeguard against Explosions, James 
 Ashworth, Ir. & Coal Trds. Rev., Sept. 26, 1902; Spraying Dust Mines, 
 M. & M., Vol. XXIV, 306. 
 
 Underground Fires, Lecture, Coll. Mgr., 1894, 35; Gob Fires, 111. Min. 
 Inst., II, 1896; Fire, Min. Insp., Pa., 1886, Packet; Shaft Fire and its Lesson: 
 Coll. Guard., Oct. 30, 1896, 838; Fires in Coal Mines, J. Abadie, Genie 
 Civil, Oct. 4, 1902; Mine Fires, M. & M., Vol. XXIII, 305; The Handling
 
 502 MANUAL OF MINING. 
 
 of Underground Fires with Milk of Lime, Wolfgang Kummer, Oesterr. 
 Zeitschr. f. Berg. u. Huttenwesen, April 9, 1904. 
 
 Explosions, the Physics of, Coll. Mgr., Jan. 18, 1895, 14; On Explosion 
 to Managers, Precautions Before, Coll. Guard., June 1897, 1084; Dust Ex- 
 plosions, 111. Min. Inst., II, 40. 
 
 Abandoned Mines, Protection of, Coll. Mgr., Dec. 18, 1896, 662. 
 
 Falls, Accidents: Scope of Committee Inquiry, Coll. Guard., June 18, 1897, 
 1186; Recklessness of Men with Lamps, C. Le Neve Foster, Coll. Guard., 
 Dec. n, 1896, 1186; Falls of Roof and Sides: Accidents and Their Pre- 
 ventions, H. M. Mine Insp., Coll. Mgr., 1896, 22; An Analysis of the 
 Casualties in the Anthracite-coal Mines from 1871 to 1880, H. M. Chance, 
 Amer. Inst. M. E., X, 67; The Prevention cf Accidents in Mines, Austin 
 Kirkup, Trans, of the N. of Eng. Inst. M. & M. Eng.,XLV, Part I, 2; Com- 
 pilation of the Fatal Disabling Accidents in the Breslau Mining District, Ge.- 
 many, Gluckauf, April 7, 1900. 
 
 Wages and Profit in Coal Mining, Ir. & Coal Trds. Rev., June 19, 1903. 
 
 Prevention of Shaft Accidents, Coll. Gua d., LXXIX, no; Fire Damp and 
 Relation to Internal Meteorology, Coll. Guard., Vol. LXXIX, 612; Rescue 
 Work, Coll. Guard., Vol. LXXXII, 1159; Diagram of Ventilation, Coll. 
 Guard., Vol. LXXXII, 645; The Nature of Explosions, Coll. Guard., Vol. 
 LXXXII, 1267; Watering Rooms, Coll. Guard., Vol. LXXXIII, 1045; Elec- 
 tric Accidents in American Mines, Coll. Guard., Vol. LXXXIV, 150; Anky- 
 lostomiasis, Coll. Guard., Vol. LXXXVI, 774; Ambulances, Coll. Guard., 
 Vol. LXXXVI, 1346, and LXXXV, 464; Watering Roads, Coll. Guard., Vol. 
 LXXXVI, 893.
 
 II. 
 
 PRACTICAL MINING. 
 
 CHAPTER I. 
 
 SHAFTS. 
 
 Shafts may be sunk for permanent or temporary purposes, 
 and they may be intended for one especial purpose only of 
 hoisting, travelling, or ventilation; or their size may be suffi- 
 ciently large to warrant division into a number of compartments, 
 one for the pumping and ladderway and the remainder for 
 hoists, according to the output. Collieries require additional 
 communication with the surface for ventilation. The large area 
 required for, and the foulness of, the return air demand a sepa- 
 rate outlet for the upcast, as also for the intake, which should 
 never be interfered with by hoisting. 
 
 The numerous drawbacks to the single-entry compartment 
 shaft or slope are so fully recited, page 20, that only in vein-mines 
 should the development be thus risked. Certainly the ventilat- 
 ing- ways should not be in adjoining compartments, because the 
 bratticing could never be kept tight enough to prevent a leakage 
 of fresh air into the upcast. Only unusually hard rock, or ex- 
 ceptional difficulties in soft or watery ground, warrant a single 
 entry. Where a prospecting drill-hole has been used to test the 
 ground the shaft should not be carried down along on it. It 
 could eventually be of greater service as a ventilator and a 
 ropeway than it could be capable of during sinking. 
 
 503
 
 504 MANUAL OF MINING. 
 
 When it is desired to remove the mineral quickly, several 
 shafts are sunk, their positions being a matter of indifference. 
 Ordinarily, however, the location of a shaft and its equipment 
 is a matter of vital import. The configuration or nature of the 
 surface affecting transportation may govern the selection of a 
 site; but, cateris paribus, the principal shaft should be so located 
 as to reach the lowest point of the workings. This is not at the 
 outset always possible to do so we are accustomed to see one 
 shaft after another abandoned or relegated to secondary uses. 
 Instance the numerous illustrations from the Lake Superior 
 region. The Calumet and Hecla has eight shafts, each over 
 i coo feet deep, with a complete plant over each one. Nor is 
 this an exceptional case. 
 
 How Deep can We Mine? This question is not now the 
 serious one that it was twenty years ago. The natural limits 
 as determined by physical conditions vary Undoubtedly the 
 maximum depth in the copper-mines of Lake Superior will 
 always be greater than in porphyry districts. The question of 
 ultimate depth would be a limit approachable by the possibility 
 of installing at the surface sufficiently large hoisting-engines and 
 stout enough wire ropes to lift the live load and the weight of 
 the rope itself. These mechanical difficulties have a positive 
 limit beyond which no mining can be carried. 
 
 Again, the possible reward of mining compared with the 
 expense of operations would place a limit of depth which is inde- 
 pendent of the mechanical question. Judging from the cost of 
 mining at the increasing average depth in copper-mines, it must 
 be granted that there appears to be no fixed relation between 
 depth and running expense. The main expenses of mining, 
 such as stoping, tramming, and superintendence, are to some 
 degree fixed, but when the heat begins to be troublesome with 
 increased depth, the cost of the working is increased. Pumping 
 need not be a serious element in determining the question of 
 cost, for experience shows in the deep mines a proportionately 
 smaller quantity of water, and a lower price per ton of output, 
 than existed during the earlier history of the mines Increased
 
 SHAFTS. 505 
 
 depth means increased rock pressure, and necessarily stronger 
 braces in shafts and all openings. 
 
 On the other hand, the increasing output and the general 
 improvement in mechanical and engineering appliances have 
 resulted in a saving much greater than the increased cost due 
 to increased depth. Undoubtedly, at the present time, with 
 an estimated depth of mine twice that of fifteen years ago, the 
 total cost of construction is less per ton of output. Again, the 
 increased depth results in increased capacity of the mine, which 
 should be followed by a corresponding increase in the capacity of 
 the shaft and stoping, and by a change in the methods of hoisting. 
 Instead of a single compartment with the intermittent hoist of a 
 single cage, there must be a number of skips or cages in each of 
 several compartments. This condition must limit the depth to 
 6000 or 7000 feet. 
 
 The results of the general discussion indicate that the limit 
 might be placed in the Transvaal or our Michigan copper- 
 mines at about 8000 feet, unless a discovery of exceedingly rich 
 ore-deposits below this depth should warrant exceptional appli- 
 ances for their recovery. 
 
 Some of the deep mines and shafts of the world are men- 
 tioned below: 
 
 In Austria-Hungary 7 are deeper than 1500 feet; in Bel- 
 gium 12 exceed 2000 feet in depth; in France 5 are over 2000 
 feet; Germany has 10 over 2000 feet; in Great Britain are 
 probably 100 exceeding 1500 feet and 20 more than 2000 feet; 
 Norway has but one deep shaft, 1900 feet; in South Africa are 
 10 exceeding 1200 feet in depth; and in the United States are 
 dozens over 2000 feet deep. Of these, the deepest shafts in the 
 several nations are indicated below: 
 
 Adalbert Przibram, Bohemia, Austria-Hungary 3672 
 
 Maria Przibram, Bohemia, Austria-Hungary 3360 
 
 Produits Colliery, Mons, Belgium 3937 
 
 Viviers Shaft, Gilly, Belgium 375 
 
 Montchanin Colliery, Le Creuzot, France 2300
 
 506 MANUAL OF MINING. 
 
 Kaiser William II, Clausthal, Harz, Germany 2900 
 
 Einigkeit, Lagau, Saxony, Germany 2620 
 
 Pendleton, Manchester, Great Britain 3474 
 
 Ashton Moss, Manchester, Great Britain 3360 
 
 Robinson Deep, South Africa 1991 
 
 Red Jacket, Calumet and Hecla, Lake Superior, U. S. A. . . 4900 
 
 Tamarack, Lake Superior, U. S. A. . . 4450 
 
 Lansell's Bendigo, Victoria 3302 
 
 Shafts sunk to facilitate the execution of long tunnels are 
 best located with their axes in the p'ane of the tunnel, afford- 
 ing better alignment, and only because of the difficulty of sup- 
 porting the shafts at the tunnel level is it the common practice 
 to place them at the side. Shafts are, however, losing their 
 importance for this work, since the introduction of the rapid, 
 ventilating, drilling-machines. 
 
 As regards form, the rectangular is the most common (Fig. 
 173). Its timbering is easily accomplished, and the best adapted 
 to loose ground. Where brick or stone is used instead of wood 
 for lining, the sides are arched to give great strength, and this 
 perhaps led to the round or elliptical shapes, which are such 
 favorites in Europe on account of their greater resistance, and 
 particularly because of the loose soils and watery strata encoun- 
 tered. That their entire area cannot be utilized is, however, 
 an objection (Fig. 172). The timbering of the polygonal (12 to 
 16 sides), used in Belgium and the north of France, is not so 
 easy to fit as is that in the hexagonal or octagonal shafts. 
 
 The dimensions of the shafts, governed by the number of 
 compartments, should be carefully studied to meet all require- 
 ments of strength, output and escapement for a prolonged 
 period. The size increases as the depth and output increase. 
 Outputs of ico tons were regarded as large not so long ago; 
 but now many hundreds of shafts have a capacity of 1000 tons 
 daily. Colliery shafts are built of a greater area than those in 
 metal-mines, which latter have less traffic, besides being restricted 
 generally by the distance between the walls. The size of the
 
 SHAFTS. 
 
 507 
 
 compartment is determined by that of the bucket, skip, or cage,, 
 its length being the width of the shaft, the length of which is 
 governed by the number of divisions (see pages 23 and 203), 
 
 FIG. 173. 
 
 FIG. 174. 
 
 FIG. 172. Divisions of a Circular Shaft. 
 
 Compartments placed side by side make a stronger shape than 
 
 if arranged in a more compact form (Figs. 173 and 174). The 
 
 compartments for metalliferous 
 
 cars are about 4' X 5' ; those for 
 
 the coal-cars, from 6 feet to 8 
 
 feet wide, by from 10 feet to 12 
 
 feet long, measured inside of the 
 
 timbers. The common sizes for 
 
 COal-shaftS are IC/X 3 8', and 12 The F r-compartment Shaft. 
 
 X 24', with wall-plates of some even 50 feet long. In the Lake 
 Superior iron region the shaft dimensions are about 9 feet long 
 and 20 feet wide. In Montana and Nevada smaller sizes prevail, 
 while in Colorado a single compartment suffices for the small out- 
 puts of high-grade mineral. The largest shaft yet begun is a nine- 
 compartment shaft 38^42' in the clear. Circular shafts for 
 buckets holding about 1500 Ibs. are 8 feet in diameter; for cages 
 13 feet. The sizes of the ventilating shafts are a matter of in-
 
 508 
 
 MANUAL OF MINING. 

 
 SHAFTS. 509 
 
 difference, so that they transmit the necessary volume of air with 
 the minimum resistance, and at a current velocity not exceeding 
 1000 feet per minute. The upcast shaft is therefore usually round, 
 and the downcast a walled rectangular. Neither should be 
 housed, though the former for a furnace ventilator may be pro- 
 vided with a chimney high enough to prevent the distraction of 
 the current by surrounding buildings; or with traps closing 
 tightly and quickly if a fan is used. An area of i sq. ft. for every 
 eight men employed is a good basis for the upcast of a moderate- 
 sized mine. 
 
 The features governing the selection of a site have already 
 been examined on page 20 ; so there remains to consider the proc- 
 ess of sinking. In a soft-ore lode the shaft section should reach 
 from wall to wall, and massive shaft pillars be maintained, else 
 it is sure to succumb. In hard-rock lodes the shaft should prefer- 
 ably be on the foot-wall;' on the hanging- wall heavy supports 
 are necessary, especially if the country rock is porphyry. 
 
 Sinking a Shaft. This process is slow because of the diffi- 
 culty of putting long angling shot-holes. Small shafts are sunk 
 by hand cheaper than by power-drills, and almost as expeditiously, 
 unless perhaps the continuous system (see Fig. 286) is used; and 
 the loss of time in removing all the implements for each shot 
 bears a large ratio to the total. Even in drifting, the actual 
 drilling heat is not more than half of the whole time. The 
 number of men depends upon the size of the shaft opened; 
 only two miners can drill to advantage on an area of 20 sq. ft. 
 A larger size gives more room proportionately to each miner, and 
 permits faster work, and in a shaft ic/Xn' there is room for 
 three pairs of miners. This space will accommodate two machine- 
 drills, which in ordinary rock can make 5 feet of advance per 
 day (divided into three shifts of eight hours each). A shaft 
 long in proportion to its width, sunk by two or four machines, 
 has two centre-cut ranges of holes (Fig. 309), which are inde- 
 pendently fired. The cost of sinking is from $5 to $18 per cubic 
 yard. Below 100 feet the rate increases each 100 feet almost as 
 the square root of the depth. Rziha says that in Europe the
 
 510 MANUAL OF MINING. 
 
 cost of excavating shafts is from 50 to 100 per cent higher in 
 wages, and the cost of putting in timber 15 to 30 per cent higher 
 in wages, than the estimate for the same amount of tunnel-work. 
 In the Lake Superior region one lineal foot of average shaft 
 costs as much as a lineal yard of gangway and a cubic fathom 
 (216 cu. ft.) of stoping. 
 
 No figures can be given for calculating the cost of any kind 
 of rockwork. Local conditions of labor and supplies vary too 
 much. 
 
 A temporary frame is erected just inside of the permanent 
 posts of the head-gear and over the end of the shaft opposite 
 the pumpway. Here the hoisting machinery is located and 
 the conveniences for loading of buckets. The pumpway is 
 partitioned air-tight, to serve for a ventilating- way as well as to 
 contain the various pipes. Precautions are taken to prevent the 
 material falling down the shaft by providing a fencing around 
 the opening and spanning it by a wide-gauge track upon which 
 a platform car travels to receive the loaded bucket and exchange 
 it for an empty one. Often a hood is provided for the protection 
 of the miners against fallen rocks; also trap-doors at the surface,, 
 unless the ventilation is very poor. 
 
 The work of excavation is begun within the frame, which 
 latter serves as the template for the remainder of the timber- 
 ing. At 6 or 8 feet down, a temporary scaffold is erected across 
 one end of the shaft to receive the material thrown upon it, whence 
 it is lifted to the surface. When it has progressed 15 feet below 
 the surface, the first permanent curbing is put into place. Through 
 the first score of feet the progress is quite rapid. Beyond this, 
 progress depends on the windlass or engine. For 90 feet a 
 windlass will suffice, but deeper than this an engine must ulti- 
 mately be used. Except in small operations the engine is placed 
 at the start. The entire bottom is attacked at once, a small 
 corner sump being carried in advance for drainage and for 
 "bearing in" while shooting. 
 
 Extending Shafts by Pentice. Shafts are prolonged with- 
 out interference to the regular mining operations, and without
 
 SHAFTS. 
 
 danger to the shaftmen, by opening only that portion of the shaft 
 area not under the hoist way for a distance of 12 or 15 feet, and 
 then widening it out to the entire size of the main shaft. This 
 
 FIG. 176. A Pentice. 
 
 leaves a roof of rock ("pentice") (Fig. 176), that shields the 
 men. When another lift has been started the. pentice is cut 
 away, and another started for the next drop.
 
 512 MANUAL OF MINING 
 
 The Service of Shaft-timbers. There is no safety nor econ- 
 omy in the practice of leaving a shaft untimbered, even if the 
 two walls are hard and self-sustaining and can be dressed 
 smooth. There is a thrust from the walls in the country rock 
 which results in the release of fragments that may injure the 
 men by their fall; a tendency to movement exists which may 
 completely close the openings, and the vibration produced by 
 rapid hoisting tends to loosen material; hence timbering will be 
 required, to furnish a close lining to the shaft and a rigid support 
 to the cage-guides. The timbers, therefore, are to serve mainly 
 as a means of preventing movement in the walls, which, having 
 once begun, can be checked. 
 
 The Character of the Timbering of Shafts, whether vertical 
 or inclined, depends upon the speed desired for hoisting. It is 
 more substantial for cage-hoisting, or mechanical haulage, than 
 it would be for a bucket-hoist or animal haulage. The varying 
 conditions as to the character of the rock, the area of the shaft, 
 and the firmness of the enclosing strata, determine the choice of 
 system to be employed. An opening, for the shaft made any 
 greater than is necessary to accomplish the desired end, not 
 only increases the expense of opening, but also the cost of sub- 
 sequent maintenance The timbers become needlessly long; 
 the cost of breaking, high; and the material to be transported, 
 larger than necessary. Moreover, the cost of back-filling the 
 useless space would be materially increased and the size of the 
 timbers in a horizontal direction must be proportionally larger 
 for a given rate of pressure. 
 
 Timbering a Shaft. The timbering of a shaft may be 
 done simultaneously with or subsequent to the process of sinking, 
 according to the firmness of the ground through which the work 
 is carried. Through loose ground, or for the first 30 or 40 feet 
 of depth through the alluvium, the timbering is placed as promptly 
 as possible and of a thickness, compactness, and strength de- 
 pending upon the nature of the ground, whether loose and rocky 
 or wet and sandy. This continues until hard rock is reached,, 
 after which some form of timber-frame is introduced. If any
 
 SHAFTS. 
 
 water is encountered, water-rings or coffer-dams are built after 
 the usual plan (Fig. 177) At the line of union between the 
 
 FIG. 177. A Clay -puddled Shaft-crib. 
 
 FIG. 178. The Timber-frame of a Shaft. 
 
 alluvium and the bed-rock a water-tight joint is made. The 
 shelf of rock is dressed level at a sufficient distance below the 
 upper surface of the stratum to allow for the packing, and on it
 
 514 MANUAL OF MINING. 
 
 is built a wedging curb of segments of heavy timbers. This 
 curb is tightly wedged and packed with puddled clay. 
 
 The standard frame consists of four pieces (Fig. 178), each 
 timber of the set being known as the plate. At the sides, RR, 
 are wall-plates, and the short plates at the ends are end-plates. 
 If the shaft is to be divided into compartments it is done by the 
 use of buntons, BB, or girts bolted to or gained into the wall- 
 plate. The greater the length of the wall-plates the stouter are 
 
 FIG. 179. Boxed Joint. FIG. 180. Boxed Shoulder. 
 
 FIG. 181. The Boxed-shoulder Crib. 
 
 the buntons or girts. On the inside of the end-plates and the 
 girts are spiked the guides for the hoisting-cages. If any com- 
 partment is to serve as the travelling- or the pumping-way, or 
 even for ventilation, it is partitioned off by planks nearly vertical 
 to branches, as in ihe right-hand compartment (Fig. 175). 
 
 Timber Joints. The joints in the frames are of several forms, 
 regard being had to the expense of framing and the convenience of 
 handling the pieces as well as the efficiency of the joint for resist- 
 ing pressure The timbers may be rough-hewn, particularly if 
 solid cribbing work is to be erected, but preferably with their 
 lengths dressed to a particular template. In firm non-decom-
 
 SHAFTS. 515 
 
 posing ground the timbers experience little pressure, and there- 
 fore stability rather than strength is sought, the latter being 
 secured by ample shaft pillars. Under such circumstances a 
 line of stiff guide-planks will be sufficient, placed in close con- 
 tact, or in a frame, according to the liability of the material 
 from walls of the rock to scale off and fall into the shaft way. 
 
 The simplest form of joint for such a lining consists of 2-inch 
 planks cut to a square butt-joint at the ends to a correct length 
 and placed with their longer dimensions vertical, so that the alter- 
 nating sets break joints (Fig. 179); in other words, the wall- 
 plates of one set overlap and rest upon the end-plates of the set 
 below, and in turn support the ends of the end-plates of the set 
 above. The end-plates are correspondingly cut. It is efficient 
 and cheap for curbing shallow shafts. It is finished by spiking 
 firmly from top to bottom four triangular corner-pieces. 
 
 Shaft-linings. In Figs. 180 and 181 is illustrated a form of 
 curbing adopted for somewhat heavier pressure, in which the 
 four pieces in the set are cut to the square-box joint to ex- 
 actly match. These timbers are 4 inches wide by 6 or 8 inches 
 high. In Fig. 182 is illustrated also a somewhat similar system 
 for slightly heavier timbers, in which the framing of the boxed 
 shoulders of the four pieces is vertical instead of horizontal. 
 If these are cut to template, it is not necessary that the timbers 
 should match in height for each set. Corner-pieces are used 
 inside of this frame. 
 
 This form of casing is built in sections of 30 feet each in 
 height by placing a pair of lo-inch stulls in a securely horizontal 
 position and building up from them the sets as indicated. 
 During the progress of this timbering the corners are plumbed 
 continually, and waste rock is packed closely with its upward 
 progress. Two men can complete one section of a shaft 5 feet by 
 9 feet in dimensions in four days' time with one or two helpers 
 at the packing. The men are supported on a cradle suspended 
 by a rope from the upper stulls. 
 
 Two- compartment shafts can also be lined by this form of 
 casing, when buckets are used for hoisting, but for cage use the
 
 MANUAL OF MINING. 
 
 timbers are not secure. Cribbing is used for small shafts, as it 
 is cheap and simple, but would be too expensive and cumber- 
 some in shafts of two or three compartments. 
 
 It is not necessary <hat these timbers should be in close con- 
 tact, but the framing may be made such that a slight open space 
 is left between each set of timbers by leaving the tenons larger 
 than is necessary for the shoulders on the ends. 
 
 A more efficient joint is shown in Fig. 183, when the pressure 
 from the wall-rock is great enough to require the lining of con- 
 
 FlG. 182. The Halved Joint. 
 
 FIG. 183. The Bevel Hitch. 
 
 FIG. 184. The Boxed Joint. 
 
 siderable thickness. In this casing a half-shoulder is cut at each 
 end to a true fit, and the pieces laid for smaller timbers. The 
 bevel hitch at the ends weakens the timber unless its full strength 
 be obtained by bringing it to a close bearing with the bevel 
 face of its mate. 
 
 This character of casing may be employed when passing 
 through dry quicksand or running strata, in which case the 
 planks would be spiked upward from below as fast as they 
 were inserted. This curbing is arranged with the layers in
 
 SHAFTS. 
 
 5*7 
 
 alternate thicknesses, measured in a horizontal direction, so 
 as to present the uneven surface at the back of the lining 
 against and upon which the sand may press. For example, the 
 
 FIG. 185. Shaft-timbering at a 
 Landing. 
 
 FIG. 186. Solid Shaft-lining with 
 Bevel Hitch. 
 
 planks may be in sets 2 inches high and 8 inches thick all around, 
 alternating with planks 2 inches high and TO inches thick.
 
 5l8 MANUAL OF MINING. 
 
 In exceptional cases this form of timbering may stand safely 
 without corner-pieces, but it is wiser to employ them, utilizing also 
 waste packing to keep the joints in position. This may be also 
 arranged in such a manner as to leave intervals between the 
 sets through which some material can be removed to relieve 
 the unusual pressure. 
 
 Another method of open framing of wall-plates resting in 
 the end-pieces, much like the log-cabin style, but without any 
 notches at the ends of the pieces is in frequent use in firm 
 ground. 
 
 Where the strata are self-supporting, the planks need not be 
 more than 4 inches thick, set on the edge for a depth of 150 feet, 
 or 3 inches thick for a depth of 120 feet; but if the ground is 
 soft, or crumbles, or there is a flow of water, other provisions 
 must be made to increase the strength of the set or to provide a 
 supplementary dam to relieve it. 
 
 It is difficult to determine the requisite size of the timbers 
 in a shaft, for the amount of pressure from the walls cannot be 
 measured. Experience with the character of the rock, its strength 
 and freedom from cleavage planes, can alone furnish any guide 
 to the engineer. The dimensions of the timbers are there- 
 fore fixed by other conditions, allowance being made for emer- 
 gencies. 
 
 Square-set Timbering. Shafts are more commonly framed 
 in square sets (Fig. 187). Square-set timbers for shafts are 8 
 inches square and consist of side wall-plates, end-plates, and 
 four posts supporting the horizontal frame. A set consists of 
 four plates and as many girts, buntons, or inner struts as there 
 -are hoist compartments. A three -compartment shaft is 
 shown in Fig. 188 with lagging for lining. The plate-joints 
 are halved with a hitch, or square shoulder-notch, i inch deep, 
 cut into the tenon for a support to the post. A centre line is 
 marked on the inside face of the plates, from which all measure- 
 ments are taken for tenons, mortises, and mitres, as well as for 
 plumbing. 
 
 The timbers are prepared by boring four i-inch holes, by a tern-
 
 SHAFTS. 
 
 U[ 
 
 id 
 
 n 
 
 
 n 
 
 g 
 3 
 
 n 
 
 Qsic 
 
 FlG. 187. A Four-compartment Square Shaft.
 
 520 
 
 MANUAL OF MINING. 
 
 plate, near the ends of the wall-plates. Through these are forced 
 iron rods with a long screw-thread at one end and a hook at the 
 other. The length of these rods is something more than the 
 half distance, out to out vertically, of the frames. Some of 
 
 FIG. 188. A Lagged Lining. 
 
 the rods have a ring at one end instead of a hook, and other 
 rods in pairs are made of a length of the odd multiples of the 
 half distance. A set of liberal-sized washers and nuts is sup- 
 plied for each pair of wall-plates. 
 
 Setting the Timbers. If the shaft is started on the hillside, 
 affording height for the convenient disposal of the debris, a 
 supporting frame, from which is to be suspended the first section 
 of timbers, is laid on the surface. If started on level ground, 
 where height must be obtained artificially, a top collar is laid 
 on the timbering above the surface, which timbers are heavy, 
 and in a frame of four pieces with the shaft -opening in the middle. 
 This frame is then reinforced by an outer crib, between which is 
 a filling of rock. The whole structure must be blocked firmly. 
 If the ground is loose, the side -plates are extended far beyond
 
 SHAFTS. 521 
 
 me limits of the shaft and bolted to heavy cross-sills for support 
 till the firm ground has been reached. 
 
 The upper set having been placed level with two of the sus- 
 pension-rods in each wall-plate with hooks down, two wall- 
 plates of the next lower set are attached with their hooks up. 
 The end-plates are then laid on the wall-plates, the posts are 
 cct in their hitches, and the nuts are screwed down tightly. 
 The posts are plumbed, blocks inserted, and wedges driven. 
 Spiked to the back of each plate is a 2" X4" strip, on which 
 rests the lagging or lining. After the filling has been stowed the 
 process is repeated for the next lower set. 
 
 Two-compartment shafts are made in two divisions of the 
 same size, framed identically alike, the longer wall-plates being 
 strengthened by girts and centre-posts. 
 
 In the three-compartment shafts the one assigned to pump- 
 ing is larger than those intended for hoisting, and its girts and 
 posts are more rigid than at the other end. The process is the 
 same as for the single-compartment, the buntons being fitted 
 into the mortises in the wall-plates after the wall and end-pieces 
 have been laid. The wall-plates should be in one piece unless 
 inconvenient. But if in two pieces, the short one should be 
 at the pumpway, with the divisional girt located at the joint. 
 The eight posts are placed, bound, blocked, wedged, and lined up. 
 
 The four-cornpartment shafts with three hoistways are in single 
 line, the fourth being used for lowering timbers and machinery, 
 and the framing is identical with that for three compartments. 
 Those constructed with the compartments in two rows (Fig. 174) 
 have their errd-plates longer to enclose a double width (Fig. 187). 
 In this event the stations are in the corners and the direction of 
 the shaft-frame is determined by the subterranean requirements. 
 The excavation is square, the timbering compact and rigid and 
 not excessively long, the shaft has greater capacity as compared 
 with the straight line, and there is but 5 cu. ft. more of ground 
 removed per foot of shaft than in the latter. The middle fram- 
 ing in the direction of the hoist is often braced with diagonal 
 struts to solidify the sets.
 
 5 2 -' MANUAL OF MINING. 
 
 Timbering in Firm Ground. If the rock is firm enough to 
 admit of sinking 50 feet or so before commencing the timber- 
 ing, the process can be more cheaply conducted by building 
 upward from reachers or stulls set into both walls. On these, 
 four pieces are framed to the studdles or struts at the corners 
 and at the compartment partitions. On these struts a similar 
 set is framed 6 feet above, to support another parallelepiped, 
 and so on up. Planks ("lagging") are driven in around these 
 frames, and the spaces to the rock filled with broken waste. The 
 joints of each timber are of the pattern shown in Figs. 189 and 
 
 FIG. 189. FIG. 190. 
 
 I 
 
 c 
 
 Joints of Square-set Timbers. 
 
 190. The end-plates and struts are usually 8 inches square, 
 while the wall-plates are laid 8 inches vertically and 10 or 12 
 inches horizontally. Fig. 177 illustrates another form of tim- 
 bering rectangular shafts with vertical corner-plates and hori- 
 zontal lagging. 
 
 Shafts such as the Comstock, 6^X24', for continuous heavy 
 hoisting, are fitted with timber as much as 14 inches squaoe, and 
 lagged with 3-inch plank. Where friable rock is penetrated, 
 the frames are braced by inclined struts that prevent settlement. 
 When the ground is friable, marly, or wet, the methods assume 
 a caisson character. Another plan comprises a stout framing 
 (C, Fig. 177), inside of which is another strong planked cribbing, 
 between which clay is puddled to exclude the water. The B. & O. 
 shaft at Taylorsville, Ind., was thus successfully carried through 
 quicksand; the outside crib was of 12-inch, the inside of ic-inch 
 timbers, with a 4-inch puddled wall. The famous Hollenback 
 shaft,45 / 4"Xn'6" inside, has a 1 2-inch clay wall for 31 feet of 
 depth (Fig. 191). It was designed for a daily output of 2500 tons 
 of coal.
 
 SHAFTS. 
 
 S23
 
 $24 MANUAL OF MINING. 
 
 Repairing Timbers. If the timber shows signs of giving 
 way, other means of securing the shaft must be invoked, \\ith 
 expert timbermen the joints may be strengthened or the frame 
 replaced, but it is preferable to reinforce them by intermediate sets. 
 
 After supporting the several sets on either side of those to be 
 repaired, the faulty pieces can be removed one by one, excavating 
 enough ground from behind each post to allow of its being driven 
 back from the shaft clear of the timbers and then replacing each 
 piece singly, wedging and driving it into place from behind. It 
 may be possible also to chop away the piece if it is not desirable 
 to open up more ground behind the set. The new post is wedged 
 into position before the next post is inserted. When plates are 
 to be removed and new ones inserted, the posts, the lagging, and 
 the blocks supporting that side must be removed and one piece 
 inserted at a time. If it bends with the ground and will not 
 stand during the pressure, false timbers must be used to hold 
 the walls in place. 
 
 Building Landing-stations. The stations are prepared at 
 various depths of about 100 feet each, from which the levels are 
 run into inner workings of the mine. These stations must be in 
 excavations large enough to serve as centres of distribution of 
 cars into the rooms and of storage of the loaded cars preliminary 
 to hoisting. Their height is necessarily greater than in the level 
 and their width is at least that of the longer dimension of the 
 shaft. The timbering and lining of such working stations is a 
 component part of the shaft system, with such modifications as 
 are necessary to connect it with the balance of the works, besides 
 affording a full, clear, unobstructed width for free movement of 
 cars and men. The timbers of the stations usually consist of 
 the four-piece level set, the caps, however, being very heavy, 
 frequently of iron. The lagging at the roof is of 3-inch plank. 
 A height equal to that of two sets of shaft-timbers will afford 
 ample height for handling timbers, pump-pipe lengths, etc. 
 
 In opening a station, four-hitch timbers are wedged under 
 the last set of the shaft crosswise, and the shaft continued some 
 distance below. A three-post cap and sill-set is placed against
 
 SHAFTS. 525 
 
 the shaft-timbers and finished. Portions of the wall-plates 
 lying between the shaft and the station, breast high, are then 
 sawed out, after which the timbers may be continued backward, 
 slanting downward as far as required to usual dimensions of the 
 drift (Fig. 185). 
 
 For inclines the stations are of two kinds: one arranged so 
 that the ore-cars will dump directly into the hoisting-skip; the 
 other is a large ore-bin built below the station at one side, from 
 which a discharge can be made into the skip at proper intervals. 
 The latter is far more economical in the long run. The skip- 
 loading, moreover, is independent of the haulage system. But 
 the excavation must be larger than for the former. 
 
 The sets are lined up by straight-edge and plumb-bobs, the 
 
 FIG. 192. Station-cribbing on a Slope. 
 
 former being long enough to overreach the distance between two 
 sets. A frame can be built in such a manner that a plumb-bob 
 will swing free within small limits from the upper end of the 
 straight-edge, the lower member resting upon the bottom sill. 
 
 In Fig. 192 is shown a station built in cribwork. 
 
 If the expense would warrant it and the diminished area is 
 not objected to, a second lining might be inserted and prove 
 secure. In the Lake Superior region mine caissons were in^
 
 526 MANUAL OF MINING. 
 
 voked after various futile experiments had been made with other 
 methods of reinforcing the shaft-timbers. For example, iron 
 caissons were forced down inside, or outside, of the old timbers, 
 according as the ground was stiff or soft. In the latter it was 
 even possible to drive them by their own weight, with very little 
 supplementary power. These cylinders, in segments and sec- 
 tions, were added one by one at the surface, keeping pace with 
 the progress of sinking. In one instance a cast-iron cylinder 
 15 feet in diameter and an inch thick was forced down at a rate 
 of 2 feet per hour through 80 feet of morainal matter 
 
 Forepoling. In very loose ground the timbering must be 
 conducted promptly by some method of shaft-lining. The usual 
 plan is that of spiling (Fig. 193). A set of planks is driven nearly 
 vertically, blocked and wedged into position. Each piece bears 
 against the walls and is fixed as the portion of a complete lining to 
 the entire shaft. In soft or running ground which cannot be tim- 
 
 FIG. 193. The Forepoling of a Shaft. 
 
 bered or requires timbers excessively heavy for lining, the method 
 consists in holding back the dangerous ground below the last 
 set in position by an enclosing and protecting shell of forepoling, 
 as described in Chapter III for levels (Fig. 271). The forepole is 
 a stout plank sharpened at its foot, often shod with iron, and has 
 a length greater than that of the distance across two sets. It is 
 driven by a sledge, a single plank at a time, from the back of. the 
 plates in the last and lowest set. Little by little the material 
 is removed to allow of it being driven a short distance. Each
 
 SHAFTS. 
 
 5 2 7 
 
 plank around the shaft is driven in turn until the entire shell 
 has been advanced. This process is repeated until a new set 
 can be inserted, behind which and in front of the previous lining 
 of four poles is driven another set in a similar manner. The 
 forepoling is started at a considerable angle outward, but in its 
 downward progress is forced toward the vertical until, by the 
 time an advance has been made for another set to be placed, 
 the planks are nearly in their desired position, giving security 
 to the frame and affording safety to the men. When the second 
 set of spiles has been started those constituting the first set are 
 driven their full length. 
 
 Forepoling has proven very successful, though it requires con- 
 siderable timbering. It is a perfectly safe method and can be 
 
 FIG. 194. Forepoling. 
 
 employed in almost any variety of dangerous ground except 
 quicksand. Frequently, when the ground proves treacherous, 
 there is perhaps too rapid a rise of the bottom of the shaft. In 
 this event the simplest plan is to brace the bottom sets and floor 
 the foot of the shaft over its entire area, leaving only a small 
 opening, advanced first by forepoling (Fig. 194), and subse-
 
 528 MANUAL OF MINING. 
 
 quently enlarged to the full dimensions of the shaft. This plan 
 resembles somewhat that of the Anderson pilot tube, as illustrated 
 in Chapter IV. 
 
 Wood Tubbing for Circular Shafts. For circular shafts the 
 false work descends with the shaft, but the cribbing is built 
 upward in sections from stout reachcrs of timber, bedded when- 
 ever suitable foundation offers, or from a properly dressed ledge 
 of the rock, and firmly wedged against the sides. The timbers, 
 assuming the character of voussoirs, hooped with iron, may be 
 mere short blocks or wedges, or they may be longer timbers form- 
 ing a polygon with less number of sides. In the latter case they 
 are united by iron dogs. In ordinary ground the sets are held 
 apart by vertical props with a solid packing behind them. The 
 solid walling may also be suspended from a heavy frame at the 
 surface by iron rods (Fig. 195). In any event the joints and 
 fitting receive the greatest care, and many of the old shafts are 
 high types of the carpenter's art. 
 
 The increasing scarcity and cost of large timbers, the ex- 
 pense of fitting and maintenance, their short life, and, finally, 
 the corrosion of spikes and splice-plates, with the consequent leak- 
 ages, have caused the abandonment of wood tubbing, and the 
 adoption of iron and masonry for all permanent ways. The 
 effect of the hot atmosphere of the mine upon timbers is a decom- 
 position that is not always detected on the surface, but once 
 begun, only better ventilation can delay ultimate destruction. 
 Dry timbers should be frequently probed; alternations of wet 
 and dry are exceedingly destructive; wet timber will last longer 
 than dry. Preservatives have been attempted, with much suc- 
 cess. In salt-mines steeping in brine gives great endurance. 
 The sulphates and chlorides of zinc have proven excellent anti- 
 septics; and a grand opening offers to the discoverer of a means 
 of freeing the lead ores of the Western States of the obnoxious 
 zinc, and at the same time utilizing it as a preservative. 
 
 Masonry Walling of Shafts. The use of masonry involves 
 but one disadvantage: it presupposes ground that will stand 
 safely for a couple of weeks without much support. Before the
 
 SHAFTS. 
 
 5 2 9 
 
 permanent structure can be introduced, a considerable depth 
 must be reached, to obtain a sure foundation upon reachers, or 
 upon a ledge from which the masonry is erected, the temporary 
 timbering and bracing being gradually removed as the construe- 
 
 I < 1 
 
 # { I 14 V 
 
 FIG. 195. Building the Shaft Walling. 
 
 tion proceeds. When a very secure ledge or base cannot be 
 had, a wedge-shaped chamber is built for some distance back 
 into the rock from which the solid crib supports the walling. 
 
 If the pressure from the walls is not very great, walls are 
 built up of brick or rock, packed securely behind by waste. Often
 
 533 MANUAL OF MINING. 
 
 the mine-water carries matter in solution that cements the whole 
 into one solid mass. When great pressure is expected, the sides 
 are arched toward it; and in very bad ground all four sides are 
 curved, or the circular form is adopted. The arc should be such 
 that its chord is perpendicular to the direction of pressure. In 
 such event, the foundations for the sections are curbs of over- 
 lapping timbers patterned to the curve, or of late years of cast 
 iron, with slabs of wood at the joints. The packing behind is 
 carried up with the brick or masonry until the ledge of the upper 
 section is reached, when it is removed gradually and the two 
 sections united. In some instances the masonry compartments 
 are built at the surface and lowered into place. Brick is well 
 adapted for quick arch work. The wall of a shaft 13 feet in 
 diameter is four half-bricks thick; the labor of laying it from 
 a staging is one and one-half days per M. The masonry is 
 supported by rods, b (Fig. 195), from beams a, a, buried firmly 
 in the walls. 
 
 Masonry is heavy to support, and not any cheaper now than 
 iron, with which many shafts are successfully curbed. Rings 
 of I beams or channel-bars form the curbs, upheld at proper 
 distances apart, by struts of wood or iron, and backed by heavy 
 planks or T 3 j sheeting (Fig. 195). English engineers use old 
 railroad iron similarly. Prepared at the surface, the curbs may 
 be lowered into place and quickly set, with little labor. A pack- 
 ing of concrete is used at Saarbruck, giving increased strength 
 and durability. It is estimated that the initial cost of iron lining 
 in place is twice that of wood and equal that of masonry, but 
 the cost of maintenance is one third that of wood and nearly 
 the same as with masonry, if the shaft is dry. 
 
 Shaft-bottoms. A very important feature in shaft-mines is 
 the arrangement at the bottom of the shaft for receiving and 
 disposing of the cars. The loaded and the empty cars should 
 go on and off at opposite sides of the cage, as on the surface. 
 Hence the heading and the position of the engine will be parallel. 
 As on the surface also, the grades to the cage and from it should
 
 SHAFTS. 531 
 
 TDC downward. The loading can then be done expeditiously and 
 economically, for not only is the cost of handling less, but the 
 capacity of the shaft is increased thereby. By this means the 
 work of caging can be performed by fewer men than if the tracks 
 were on a level or up grade for the empties. The latter are run 
 down to some convenient point where they are picked up by 
 the locomotive or mules. 
 
 If it happens that, in order to accomplish this, the shaft- 
 bottom must be lowered and the main-entry bottom also taken 
 up, the subsequent saving will warrant the initial expense. Occa- 
 sionally it is found desirable to build a special roadway, passing 
 the sides of the cage, to return the empties to their appropriate 
 entries. 
 
 The headings at the shaft-bottom are usually arched with 
 brick, with considerable packing back of the arch, and further 
 protection of longitudinal layers of 4-inch plank. Care should 
 be taken that the space between the rock and the back of the 
 arch be completely filled to avoid movement or side pressure of 
 the roof. The supporting walls to the springing line should 
 not be over 4 feet high, and the curve of the arch as low as will 
 give sufficient height for the work to be done at the shaft. Not 
 infrequently an inverted arch is built for the floor. 
 
 Shaft Pillars. The dimensions of the shaft pillars should be 
 such as to ensure absolute security to the works. Though many 
 local conditions affect their size, the following rule may be given 
 for flat seams, where the depth does not exceed 1000 feet. With 
 D to represent the depth of the shaft, t the thickness of the seam, 
 and R the radius of the pillar, all dimensions being in feet, then 
 
 In inclined seams it Is evident that the larger portion of the 
 pillar should be to the rise rather than to the dip. Its length 
 then is greater than the dip portion by three fourths of d, in 
 which d is the distance in feet along the seam from the foot of
 
 532 MANUAL OF MINING. 
 
 the shaft to the foot of a perpendicular drawn from the top of 
 the shaft to the seam. Then, if R is the length of the pillar to 
 the dip, 
 
 R = 6o+o.o$D\/T. 
 
 The line at right angles to the dip of the seam is the ultimate 
 breaking line. 
 
 In a seam 6 feet thick, with an inclination of one in three, in 
 a shaft 1200 feet deep, the dimensions of the dip pillar will be 207 
 feet. The value for d being 346, the length of the pillar to the 
 rise will be 406.5 feet. The total length of the pillar becomes 
 then 673.5 f eet an d its width for one shaft twice that of the dip 
 pillar, or 414 feet; for two shafts, 180 feet apart, 594 feet. 
 
 In the case of a fresh opening in another seam of coal above 
 or below the bottom already existing, care should be taken that 
 the headings in both seams be not placed immediately above 
 one another and parallel, unless the distance is very great. 
 
 The underground s ables should be in a position as con- 
 venisnt as possible, and well ventilated with splits in each sepa- 
 rate stable for 24 animals. The archway overhead should be 
 12 feet by 16 feet. 
 
 The following references are cited: 
 
 REFERENCES. 
 
 Earth Pressures in Deep Mines, Coll. Guard., Jan. 2, 1903. 
 
 The Framing of Rectangular Shaft Sets, Wilbur E. Sanders, Eng. & Min. 
 Jour., Mar. 10, 1904; Timbering Shaft Coal Mines, Wm. Bradford, Coll. 
 Mgr., Nov. 1896; Shaft Sinking in Germany, H. Huhn, Coll. Guard., LXXII, 
 932; Shaft Linings, Coll. Guard., Mar. 27, 1003; Plumbing Deep Shafts, 
 M. & M., Vol. XXII, 247; Shaft Sinking Contracts, M. & M., Vol. XXIII, 
 260; Shaft Sinkin- Progress: Four Shafts Witwatersrand, E. & M. J., May 
 22, 1897, 507; Shaf Timberi-g, Leith Coal Shaft, Coll. Eng., Aug. 1896, 3. 
 
 The Safety Aspect of Inclines, Henri Ghysen, Coll. Guard., Mar. 30, 1000: 
 Dipping Shaft, M. & M., Vol. XX, 15. 
 
 Automatic Shaft Doors, H. Siede, Gluckauf, Oct. 31, 1003; Methods of 
 Closing the Tops of Upcast Winding Shafts, A. Ried, Fed. Inst. M. E., X;
 
 SHAFTS. 533 
 
 Some Arrangements for Preventing Accidents at Level Landings in Cage 
 Dips and Shafts, A. R. Sawyer, N. Staff. Inst., VIII, 204. 
 
 Deep, Level Shafts on the Witwatersrand, with Remarks on the Method 
 of Working the Greatest Number of the Deep Level Mines with the Fewest 
 Possible Shafts, I. H. Leggitt, A. I. M. E., Vol. XXX, 947; Cost of Mining 
 and Depth, A. C. Lane, Mineral Industry, IV, 777; How Deep can We 
 Mine? A. C. Lane, Mineral Industry, IV, 767; Deep Shafts of the World, 
 B. H. Brough, Coll. Guard., Dec. 1896, 1170; Difficulties in Deep Mines, 
 M. & M., Vol. XXII, 58. 
 
 Laying Out Shaft Bottoms, W. Stewart, Coll. Eng., Dec. 1896, 188; Shaft 
 Pillars, W. Stewart, Coll. Eng., Dec. 1896, 189; Shaft Bottom, M. & M., 
 Vol. XX, 406, 454. 
 
 Renewing Old Shafts, Coll. Guard., Vol. LXXXI, 1991.
 
 CHAPTER II. 
 
 SINKING IN RUNNING GROUND. 
 
 Excluding Watery Strata. In he search for minerals, mining 
 is conducted to greater depths and into more treacherous ground, 
 as time advances, and the highest type of engineering skill is 
 called into play as water-bearing strata are encountered. Not 
 only is the shaft to be sunk through them, but their underground 
 currents must be excluded from the mine by water-tight casing, 
 otherwise an elaborate sy tern of pumps must be continually 
 maintained during sinking and mining. Not infrequently the 
 expense of removing seepage-water while sinking becomes so 
 large an item of dead work as to compel the abandonment of 
 the works. A system of tubbing is, under those circumstances, 
 advisable. 
 
 Two varieties of cases present themselves one in which the 
 ground penetrated is quite firm but porous, and the other includes 
 running soil, marl, quicksand, etc. In the first case, the pumping 
 facilities must be ample, or the water kept back during mining; 
 in the second, the possibilities of excavation are greater than the 
 facilities for its hoisting, and the laborers are in danger of being 
 overwhelmed with soil. In either case the ground traversed 
 must be insulated by a tubbing, hermetically sealed above as 
 well as below the soft measures. 
 
 This not only renders sinking possible, but it excludes water 
 and silt from the mine, and permanently dispenses with much 
 of the pumping arrangements. It very seldom comes into play 
 in vein-mines, where, with the vertically of the lodes, water 
 cannot be prevented from percolating into the mine. It is in 
 
 534
 
 SINKING IN RUNNING GROUND. S3S 
 
 the stratified regions that the use of the crib is of the highest 
 importance. Beds of gravel, sand, and clay, or porous strata, 
 percolating large quantities of water, are not easily traversed or 
 held up without a strong water-tight lining, for the pressure of 
 the moving material tends to make the bottom rise, as well as 
 threaten the sides. A deep shaft in such a region may encounter 
 several occasions for such tubs, which under suitable conditions 
 may be introduced in lengths as required, and only to the extent 
 of the soft ground. Still, it would give more substantiality to the 
 work to form one continuous length of tubbing, even across the 
 good ground. It is not uncommon to find in Germany shafts 
 with three sections of iron tubbing, united by lengths of brick or 
 wood lining. 
 
 Tubbing. This process consists in confining the seepage 
 area to the bottom of the shaft only, by building a water-tight 
 cylinder lining to the shaft, and carrying it down with the sink- 
 ing beyond the wet stratum. In England, a bed of sand called 
 the lower red sandstone, which is almost fluid, has several 
 shafts tubbed through it. In Britain, Belgium, and the North 
 of France, several mines are reached by tubbing through the 
 fissured chalks and marls of the Cretaceous. The Thonmer- 
 gel of Germany is frequently tubbed to the Bohn Erz, below, 
 dry enough for work. While sinking the Murton pits 4000 
 gallons were pumped per minute, and the "come in" of water 
 for the Exhall shaft was 1650 gallons. Still, the inefficiency of 
 this plan, sometimes called the English system, is recognized, 
 and several methods bet er applicable to loose and watery beds 
 have been applied with more or less success. Excepting the 
 Poetsch method of freezing the ground to be penetrated, they 
 are modifications of the diving-bell, or pneumatic pile. 
 
 Wrought-iron tubes in segments are bolted on at the surface 
 as fast as the lowering proceeds, until the secure, impermeable 
 bed is reached. Here a smooth base is prepared for one or more 
 wedged curbings, behind which moss or concrete is rammed. 
 The tubbing is backed with rock or concrete all the way up, and 
 connected with the next upper section. The holes in the seg-
 
 53 6 MANUAL OF MINING. 
 
 ments, for convenience in handling, and to relieve the tubbing 
 of pressure till the work is completed, are plugged up. The 
 early practice of bolting the segments together through the inside 
 flanges was soon abandoned, and now the flanges are outside, 
 wedging, pressure, and friction keeping them. On account of 
 the curious accidents occurring from the pressure of air locked 
 behind the tubes, it is advisable to lay a pipe to the surface for 
 the gas to escape, and, similarly, to relieve the water. A shaft 
 of 1 6 feet diameter was sunk at a monthly average of 104 feet 
 with four shifts of 8 and 10 men each. Several Canadian salt- 
 mines, having shafts 10 feet 6 inches diameter and reached at a 
 depth of 1150 feet, are tubbed through 260 feet of water-bearing 
 strata, in sections 2 feet high and f inch to ^| inch thick. 
 The columns rested on iron curbs with firm base. The joints 
 are calked. 
 
 Masonry Curbing. This is a much cheaper and more effi- 
 cient wall in shafts than is iron tubing. With good hoisting 
 machinery, three masons in four-hour shifts finish 10 feet per 
 day of a 1 6-foot shaft. Towers of masonry, resting on an iron 
 curb with a cutting edge, were built on at the surface; while, to 
 facilitate the sinking, digging was being carried on below, or, if 
 the material was wet, a process of "bagging" was employed. 
 When abundant in size and quality, wood gives great satisfac- 
 tion, being elastic, easily laid and repaired qualifications not 
 possessed by masonry. Iron offers the advantages of strength, 
 combined with a facility of handling, which recommend it for 
 large shafts and enormous flow. Though it is not possible to 
 presage or measure the pressure, and thus determine the kind 
 and strength of tubs, a thickness of 1 2-inch wood, y-inch masonry, 
 and ^-inch iron may be suggested as common. As a matter of 
 fact, the tubbing should taper off toward the top. In many 
 cases, however, the use of 12 inch staves hooped with iron did 
 not prove adequate, nor did the backing of 12 inches more of 
 concrete help matters; where the shafts were not abandoned, 
 A -inch sheeting met the emergency. 
 
 Sinking through Running Ground. When a bed is encoun-
 
 SINKING IN RUNNING GROUND. 537 
 
 tered of a material so soft as to behave like a fluid and be pump- 
 able, the area which can be opened with safety is very small 
 indeed. In this event some variation of the "spilling" proc- 
 esses, a pneumatic pile or some process of congelation, should be 
 employed. 
 
 Triger's Method by Pneumatic Tube. M. Triger employed, 
 in 1839, the principle of the pneumatic pile, in which an iron tub- 
 bing was built down to an air-lock which communicated with a 
 diving-bell at the bottom. The atmosphere of the caisson was 
 maintained at a pressure of not over 60 Ibs. per square inch, 
 and checked the influx of the sand, which the miner shovelled 
 to a sump, whence it was aspirated to the surface. Meanwhile 
 the tubbing was being forced down from the surface as rapidly 
 as possible. The tubbing was divided into three chambers by 
 partitions with hinged doors, the middle one constituting an air- 
 lock. The natural limitations of this process are the outside 
 earth friction and the physiological difficulties of men working 
 in compressed air. 
 
 The Kind and Chaudron Process of Boring Shafts. In 
 France and Germany the loose, watery marls presented diffi- 
 culties which the methods described failed to overcome. What 
 with pebbles and fine rock interfering with aspiration, water 
 completely inundating the shafts, and the difficulty in establish- 
 ing water-tight joints, the operators were routed. In 1850 Herr 
 Kind devised a scheme for mechanically sinking shafts, just as 
 one does a bore-hole, and still further conquered difficulties 
 hitherto insurmountable by a variation in the mode of lowering 
 the tubbing, and by a device for regulating the influx of water. 
 When M. Chaudron added the sliding bottom-piece to form a 
 perfect joint, after the Kind boring-tool had prepared the base, 
 the acme of shaft-sinking was reached. Since 1862, when the 
 first shaft was sunk, 6 feet wide, 480 feet deep, at a cost of $450 
 per foot, not a single fatal accident is recorded against the process, 
 which owes much of its success to the fact that the sinking and 
 lining are completed before a man enters the shaft. Two aban- 
 doned shafts, through soil feeding 11,000 gallons per minute,
 
 538 
 
 MANUAL OF MINING. 
 
 were carried down 267 and 216 feet, respectively, in 23 and 20 
 months, with a cost of $280 and $340 per foot. In the latest 
 application 569 feet were sunk, 16 feet diameter, at an average 
 cost of $143 per foot for both shaft and lining, which latter cost 
 $70 per foot. With a guaranteed success at so low a rate, it is 
 surprising that American engineers, usually so progressive, have 
 not employed this method before acknowledging failures; but 
 no attempt to introduce this plan here is as yet recorded. 
 
 The Trepans. The ordinary procedure comprises first drilling 
 a guiding bore-pit about 50 feet deep and 4.5 feet across; then 
 widening the shaft by a reamer to the required diameter, and 
 alternating these drills with every 50 feet of drilling. These 
 drills, or "trepans," are operated from a walking-beam by a 
 surface engine. 
 
 The small trepan (Fig. 196) consists of a blade of forged 
 iron, into the lower side of which are keyed a number of pointed 
 
 steel teeth, and a stem connecting the blade to 
 the suspension- rods by means of a sliding-box. 
 This last partially corresponds to the "jars" 
 (Fig. 277) of oil-well outfits, takes up the jar, 
 and is an essential element of the tool. The 
 trepans are massive for hard rock, weighing 
 from 8 tons up and are raised 6 inches or so, 
 turned slightly for each blow, and dropped; 
 their concussion disintegrates the rock along a 
 diameter of the circle. The progress in flint is 
 3 inches per day; in chalk, 3 feet; in sandstone, 
 i foot, and in coal measures, 16 inches Most of 
 the material is broken quite fine, though 2-inch and 3-inch stuff 
 is not unusual. 
 
 The larger trepan replaces the smaller one after 50 feet of 
 drilling. It is similar (Fig. 197) to the smaller one, but the blade 
 is deeper at the centre than at the ends, so that its teeth cut a 
 base sloping to the centre. The central, toothless portion of 
 the tool has a U-shaped guide that fits the smaller hole. This tool, 
 often weighing as much as 16 tons, cuts the shaft to full width , 
 
 FIG. 196. 
 A Trepan.
 
 SINKING IN RUNNING GROUND. 
 
 539 
 
 or it may be succeeded by another similar reamer, the detritus 
 falling into the smaller hole, from which it is hoisted by the 
 sludger (Fig. 199). In alternate stages the drilling and widening 
 progresses, while the tubbing is subsequently lowered by a sepa- 
 rate engine. All these operations being conducted under water, 
 the trepan requires to be automatically kept vertical. Two 
 guides, carrying at the extremities horizontal and vertical cutters, 
 accomplish this marvellously well. The record of the preliminary 
 shaft, 4! feet in diameter, showed for 508 feet an average progress 
 of 3.3 feet per 24 hours, divided up as follows: 51 per cent o 
 
 FIG. 197. Fir,. 199. 
 
 The Kind-Chaudron Trepans. 
 
 FIG. 200.. 
 
 the time was occupied in drilling, 19 per cent in raising and lower- 
 ing the tools, 20 per cent in dredging, and 10 per cent in repairs 
 and delays. Widening the shaft to 14 feet and down 460 feet 
 took ten months; reaming, 46 per cent of the time; altering 
 and operating the tools, etc., 14 per cent; dredging, 22 per cent;, 
 delays and repairs, 18 per cent.
 
 540 
 
 MANUAL OF MINING. 
 
 The widening trepan reamer and sand-bucket dredge are 
 shown in Figs. 197 and 198. 
 
 There are no screws or nuts to loosen; all the parts are keyed 
 to place; but special tools are provided for grappling broken 
 rods, stems, trepans, teeth, etc. 
 
 At the surface the operations of boring the pits, building 
 and lowering the tubbing, puddling and sealing the base, are 
 conducted with engines and capstans from a tall derrick, to 
 which extra lengths of rods may be attached with the progress 
 of the drilling. Nine men are employed about the works, only 
 three of whom are skilled laborers. The cost of the installation 
 of the machines, tools, etc., all of which are portable, is about 
 $13,000 to $20,000. 
 
 The Tubbing and the False Bottom. This element of the 
 system is of iron sheeting, built on in 6-foot sections, with leaded 
 joints, and suspended by rods. The flanges are on the inside of 
 the tubbing bb, leaving a perfectly smooth exterior, and the joints 
 are true and bolted together. Two sections are lowered daily. 
 As an example, a tubbing 12 feet 7 inches 
 in internal diameter and 280 feet high is 
 quoted; it was i inch thick at the top, 
 if inches at the bottom, and weighed 400 
 tons. The sections were 5 feet high, the 
 flanges 3^ inches wide, 2 inches thick, 
 having leaden wedges between, 4^ inches 
 wide and inch thick, and 20 bolts 
 i A inches in diameter. 
 
 At the bottom of the iron cylinder 
 are attached two very ingenious appli- 
 ances, which, operating automatically, 
 FIG. 201. -The False Bottom. have established the process as a success 
 beyond all cavil; the first is a moss-box, a, for hermetically seal- 
 ing the lower end of the tubbing against any influx of water; 
 and the second is the introduction of a false bottom, /, by which 
 the sinking of the tubbing is cleverly controlled. These are both 
 adapted to the bottom of the tubbing, as is illustrated in Fig. 201. 
 All the flanges of the tubbing turn inward except the lower one
 
 SINKING IN RUNNING GROUND. 541 
 
 of the bottom section, bb, which is outside, and may act as an an- 
 nular piston to a lower section, aa, of smaller diameter, the upper 
 flange of which turns inward, and the lower one outward. Be- 
 tween these flanges, the moss-box, and the rock, the annular 
 space is filled with moss, which is not, however, under com- 
 pression so long as the screw-bolts 55 support it from the tubbing. 
 It operates like the seed-bag of oil-well diggings. 
 
 The false bottom // is attached to the tubbing, with the lower 
 sections of which it forms a diving-bell that floats the whole 
 system. The greater the head of water encountered, the more 
 complete the balance, and the greater is the relief to the rods, dd, 
 supporting the hundreds of tons of iron. The safety-pipe, g, 
 with cocks and plugs operated from above, is an equilibrium 
 column that permits sinking, or rather regulates its speed. 
 Opened at the top, sinking proceeds rapidly, as the compressed 
 air and water find vent; closed, the whole structure is upheld 
 against gravity. When the plugs are opened they discharge 
 into the tubbing, weight it with water, and at the same time 
 release the pressure below. By proper manipulation, perfect 
 control is had over the lowering of the casing. 
 
 Making a Water-tight Joint at the Bottom. When the tubbing 
 has traversed to the required depth through the water-bearing 
 measures, a seat having been scraped for the moss-box, the entire 
 weight is allowed to fall on the annular piston, b, by opening 
 g at the surface. The moss is compressed to a 
 hard, water-tight mass, the rods 5 gliding in 
 their bearings (Fig. 202) Up to this time the 
 shaft is more or less full of water (the process is 
 independent of the amount of water encoun- 
 tered). This may now be pumped, but usually 
 is not until a cement backing has been inserted 
 and hardened to insure solidity; after that, if 
 the joints of the metallic column are well made, 
 the shaft is perfectly tight, and the mine is in- 
 sulated from the subterranean current. The FlG - 202 . 
 
 The Moss Joint, 
 introduction of the cement is effected by a 
 
 closed spoon holding a barrel or so, curved to suit the space.
 
 542 MANUAL OF MINING. 
 
 Three sets of six men each do this work, burying 400 cu. ft. per 
 day, at a cost of about forty cents per square foot area of lining. 
 A solid foundation of wedged iron curbing is subsequently built 
 on a stout ledge, to take the weight of the cribbing after the other 
 work has been completed. 
 
 This method is generally applicable to conditions of soft 
 ground, and especially in watery ground, which can be pierced 
 without recourse to ponderous pumping machinery. Though 
 the pressure of the water is not essential to success, it materially 
 facilitates operations. In a few cases, where the ground was 
 merely wet, not running, tubbing was cheaper than this method. 
 But the facts should not be lost sight of, that none of the delays, 
 perils, and discomfitures of the ordinary methods are here experi- 
 enced. Its progress is greater, and initially its maintenance 
 is cheaper than other schemes for wet ground, besides never 
 having had a failure, though no shaft of over 14 feet in diameter 
 has yet been sunk by this method. 
 
 An objection to the Kind and Chaudron method is that, 
 unless an exact geological section is at hand, there are no means 
 of knowing when the water-bearing stratum has been penetrated. 
 
 A short while ago a pair of shafts were sunk in Samlund, East- 
 ern Prussia, for amber, through 147 feet of clay and sand, by a 
 variation of this method. The drill- tools, weighing 1700 Ibs., 
 cut a 4-foot 6-inch space, though they had little to do except in 
 the shale-beds. No moss was necessary, as the ground was not 
 wet. Four-feet lengths of tubbing were forced down by jack- 
 screws, each shift with 27 men. The total weight of tubbing in 
 each shaft was 45 tons, and the total cost $17,500. 
 
 Lippmann's Drill. This is a drill of a double V-shape, 
 instead of a straight trepan. It does faster work, as it cuts 
 equally at the periphery with the centre of the circle. With 
 Kind's trepan the blows fell too far apart at the outer edge of 
 the shaft, and too near together at the centre. 
 
 Haase's System. This employs sheeting-piles, in small 
 round iron cylinders driven close together to form a cribbing 
 for the intended shaft. The tubes were about 15 feet long,
 
 SINKING IN RUNNING GROUND. 543 
 
 ^ inch thick, and 4 inches in diameter, and enclosed an area 
 10X7, which was then timbered. These were driven through 
 90 feet of quicksand in five months' time, at a cost of $135 
 per foot. While standing, they gave good drainage, and d'd not 
 yield when the excavation of the shaft began. 
 
 The Freezing System. For loose wet alluvium, Mr. C. 
 Poetsch has originated the novel idea of freezing the mass to a 
 solid by boring a great number of holes through the alluvium 
 about 3 feet apart, and lining with copper tubes, inside which are 
 smaller tubes. A concentrated solution of the chlorides of mag- 
 nesium and calcium circulates through the tubes and freezes 
 the ground, after which the pit can be excavated in the centre 
 of the mass in the ordinary manner, and the tubbing put in. 
 This refrigeration is continued till solid rock is reached. A 
 shaft 7'Xn' was sunk through .26 feet of quicksand, the frozen 
 wall enveloping it being 7 feet thick, at a cost of $190 per 
 foot. 
 
 Conducting the Seepage through the Walls. Water-traps are 
 frequently provided in shafts which are lined with masonry or 
 timber. These consist of grooves cut around the entire periph- 
 ery of the shaft into a recess in the walling of the shaft and 
 provided with pipes and spouts through which the water can be 
 carried to some more convenient point rather than permit it 
 to give annoyance in the shaft. 
 
 If the waters are muddy or contain salts in solution which 
 would clog the spouts, considerable annoyance is given and 
 expensive arrangements must be resorted to. 
 
 REFERENCES. 
 
 Artificial Foundations and Methods of Sinking through Quicksand, W. 
 E. Garforth, Mid. Inst, XI, 407; Dealing with Water during Sinking, 
 Institution of Civil Engineers, Coll. Guard., May 28, 1897, 994; Sinking 
 through Quicksand, Peter Jeffrey, 111. Min. Inst., II, 90, 230, 240; A New 
 Method of Shaft Sinking, G. C. McFarlane, Eng. & Min. Jour., April 7, 1900. 
 
 Freezing Process Shafting, Gobert, Coll. Eng., XVII, 171; Freezing 
 by Compressed Air, M. & M., Vol. XXII, 411; Freezing Water an Aid to
 
 544 MANUAL OF MINING. 
 
 Mining, Coll. Eng., XV, 86; Sinking of Shaft, B. Barnum Mine, Ishpeming, 
 Mich., R. H. Vondy, The School of Mines Quarterly, May 1882, 277; Sink- 
 ing a Shaft by Compressed Air, Bergassessor Luthgen, Coll. Guard., Jan. 9, 
 1903; Sinking by the Freezing Method at Washington, County Durham, 
 Mark Ford, Ir. & Coal Trds. Rev., Dec. 19, 1902; Difficulties with the 
 Freezing Process, Coll. Guard., Vol. LXXXI, 201. 
 
 Kind-Chaudron Process, A. I. M. E., V, 117. 
 
 Cleaning Shaft of Water during Sinking, Coll. Guard., Vol. LXXXII,
 
 CHAPTER III. 
 
 TIMBERING ROOMS AND GALLERIES. 
 
 The Service of Timbering. The accident-tables discussed in 
 Chapter XIII show how manifest is the neglect of a few simple 
 rules, endangering life and property; and in no respect is this more 
 painfully impressed than by the mortality record of unpropped 
 rock. Excavations, even in the "rock of ages," cannot be left 
 open any great length of time without support, which, if intro- 
 duced in time, will prevent disastrous results. Successful super- 
 intendents personally watch the timbering and the face-rock- 
 diligently, and guard against any springing of the walls. All 
 the effects of pressure are intensified by neglect, and the secret 
 of success is to place timbers before movement begins. Supports- 
 are not for bad roofs only; while "awaiting a weak spot, the 
 good roof, so called, catches him," and his stope or room is 
 lost. The eagerness to quickly win the face, while pardonable,, 
 promotes avarice, parsimony, want, and then provokes collapse. 
 
 Though the conditions underground are such that very simple 
 timbering is required compared with that on the surface, the 
 tendency of the time is toward the employment of special timber- 
 men to make and place the supports. In rooms, driving in 
 soft ground, and the like, the miners must at once prop the exca- 
 vations. For this reason and the proscribed space the char- 
 acter of the timbering should be simple and the sticks light. 
 Fortunately, the pressure of the country rock is inward and to- 
 ward the openings, and compression, not tension, as on the 
 surface, is to be combated with. This tends to hold the sets 
 together. Tenons and framing may therefore be dispensed with,. 
 
 545
 
 546 MANUAL OF MINING. 
 
 except in loose ground, where they are essential for maintaining the 
 integrity of the timbers. 
 
 The relative merits of the different varieties of wood need 
 not be discussed here. Oak is undoubtedly the most preferable, 
 but the mines take what can be had in the vicinity. Above 
 "timber-line" we are content with "scrubs." Sawn timber is 
 better than hewn, on account of its greater resistance to decay; 
 and durability is of prime importance to strength. Again, green- 
 wood is heavy; the ordinary lo-inch stick, say 6 feet long, is 
 as much as three men can well handle. Lightness is an essential 
 feature in this most onerous of underground work. 
 
 The Life of Timber. This varies with the condition of the 
 atmosphere and care in dressing. It is rarely as great as that 
 of railroad-ties (twelve years). In many mines head-pieces 
 crumble after two years' standing. Wood rots faster, and shows 
 it less on the surface in dry, vitiated air than in moist air. Alter- 
 nations of temperature or moisture are very destructive. A 
 cotton-fungus mould is a sure indication of bad air, and, being 
 contagious, requires attention at once. 
 
 The decay results from the fermenting of the albuminoids of 
 the sap, the admission of water, and the attack of insects, to 
 which several causes contribute, bad air, damp air, standing 
 water, and oxidation, causes all of which are mitigated by an 
 active circulation, and materially remedied by saturation of the 
 pores with some antiseptic. Creosote, Kyanizing, or Burnet- 
 izing will give greater life to timber. The timber is placed in 
 a wrought-iron cylinder through end-doors, after closing which the 
 air is exhausted and creosote forced in. Pine absorbs from 10 
 to 12 Ibs. of oil per cubic foot, and the hard woods less. The 
 pressure during the operation is 100 Ibs. per square inch. 
 
 Timber Consumption. While it cannot be accurately stated 
 for the average mine, it has been estimated that the timber con- 
 struction in anthracite mines is i cu. ft. per ton product; in the 
 L. S. copper-mines, if; in Leadville and L. S. iron-mines, 3; 
 .and in Nevada, 4^. 
 
 Every 100 cu. ft. of coal extracted consumes 3.4 cu. ft. of tim-
 
 TIMBERING ROOMS AND GALLERIES. 547 
 
 ber; every ton of excavation in running ground requires about 5 
 cu. ft. to support the balance. In the Anaconda mines, Mont., 
 80,000 cu. ft. of timber are used daily; 2000 of such mines would 
 consume the entire forest-area growth. The West Vulcan iron- 
 mines, in L. S., annually consume 2,000,000 feet of lumber and 
 60,000 pieces of lagging, at a cost of 37 cents per ton of ore mined. 
 In the copper- mines of L. S. this item amounts to from 15 to 31 
 cents per ton of rock hoisted. So important in the economy 
 of mining and to the safety of lives, the selection and placing 
 of timbers should therefore receive skilled attention; adequate 
 ventilation is equally urgent for their preservation. 
 
 Owing to its destructibility and the increasing scarcity of 
 supply that is bringing about the employment of iron and 
 masonry as the certainties of future support. Fortunately, 
 the metal-mines above "timber-line" require but a moderate 
 supply. 
 
 Elements of Timbering. Though the question as to whether 
 the timbering is to be done in a substantial manner at once, or to 
 be considered as provisional, is only answered according to the 
 importance of the gangway; the practice is to assume it per- 
 manent. Rooms and stopes are only temporary, and treated 
 as such. Gangways may be subsequently cribbed or masonried; 
 but all pump-rooms, machine-rooms, and stables are very sub- 
 stantially lined. 
 
 Each piece should be placed conformably to the principles 
 of the strength of materials, and laid in such direction as will 
 best withstand the pressure whose direction is known. The 
 crushing force is better resisted than is a bending force. They 
 should be placed in the line of the pressure wherever practicable, 
 or in such manner as to act like or take the part of an arch ; then, 
 when any movement takes place r its effect will be to tighten the 
 timber in place. Joints should bear the pressure uniformly and 
 their planes be perpendicular to its direction. 
 
 Rooms and levels are timbered in i-, 2-, 3-, or 4-piece sets. 
 The vertical posts, horizontal piece, and top and bottom sill con- 
 stitute the usual full set, and the inclined stull or cap represents
 
 548 
 
 MANUAL OF MIXIXC. 
 
 the one-piece or "quarter" set. Usually the system of room- 
 timbers differs from that in the levels. 
 
 Props and Stulls. Single sticks are used to support the 
 roof back of the men. In long-wall working a large number is 
 used, resting on the floor or on a plate, and hammered into place 
 with a wedge-plate at the roof. They are from 6 to 8 inches 
 diameter, and stand 3 feet apart, in two or three rows, beyond 
 which the roof caves in on the gob. They remain only a few 
 days, are removed by rows to let the roof cave, and are replaced 
 nearer the face (Fig. 3). An average of 70 per cent is recovered; 
 some of the balance cannot be removed ; others \vould endanger 
 the timbermen. Flat caps on top, 20" X io"X2", are ample 
 
 FIG. 203. A Stull-piece. 
 
 for most bad roofs. Slate requires a large plate. It is a poor 
 roof, because it crumbles from the presence of pyrites; sand- 
 stone or conglomerate makes a good roof; soapstone is bad; 
 but the most dangerous is fire-clay, which runs when exposed 
 to the moist air. Props come into play (Fig. 210), 12 inches 
 long, 3 inches or 4 inches in diameter, for holding up holed coal. 
 The props do not support the strata above the coal this the 
 pillars do. They support only a portion of the immediately 
 overlying seams which constitute the roof. The condition of the 
 roof and the method of mining determines the number and 
 distance apart of the props. Every loose block should be 
 removed or propped up. 
 
 The presence of seams and cleavages traversing one another 
 in the rock materially affects the selection of the modes of tim-
 
 TIMBERING ROOMS AND GALLERIES. 
 
 549 
 
 bering. The parallel joints are not troublesome. Horses in the 
 vein usually require special attention, as do evidences of sigil- 
 lariae. The latter occur like truncated cones, base down, and the 
 circular layer in the roof should be propped as soon as observed. 
 It must be remembered that the prop, as the support for 
 the roof, is employed, not as a means of security, but as a warning 
 of excessive pressure. Its elasticity is of greater importance than 
 its strength. By its bending will be indicated the great pressure 
 
 FIG. 206. FIG. 207. 
 
 Single -stick Timbering in Veins. 
 
 to which it is subject, and therefore the necessity for extra pre- 
 cautions, if the room is still to be kept open, or for protecting 
 the miners, if the roof is to be permitted to fall. For this reason 
 props of cast iron or of steel in working places are not recom- 
 mended, however well they may serve as a secure support for 
 rock in the main travelling-ways. 
 
 States vary in their statutory requirements as to the nearness 
 of the props to the face, but 15 eet is the farthest al owed in
 
 55 
 
 MANUAL OF MINTSG. 
 
 any coal region. A distance of 5 or 6 feet is ample space for 
 the men, but not for machine cutters. 
 
 In metal-mines the prop is used as a stull (Figs. 203, 205, 206), 
 resting in a notch ("hitch"), generally on the foot-wall, unless 
 the hanging-wall is much softer, and driven into place with 
 a wedge-piece, by mallets. In veins of small inclination the 
 stulls are normal; otherwise they stand between normal and 
 the vertical, because of the combined action of the wall-thrust 
 and gravity. Its angle toward the vertical from the normal is 
 about one fourth the pitch of the vein. They are round or 
 dressed, and of a size and distance apart dependent upon the 
 weight of waste stope-rock to be upheld. It is better to increase 
 
 FIG. 208. Braced Stulls. 
 
 FlG. 209. Vein Timbering. 
 
 the number than the size, though their strength is directly as 
 the cube of their diameter. d 3 = o.o$hw 2 m is the formula for. 
 calculating the size of any round stull held at two ends, h is 
 the height of rock along the vein in feet; m the distance between 
 the stulls in feet; iv the width of the vein in feet; d the diameter 
 
 FIG. 210. Coal Chocks. 
 
 of the stull in inches. To support 60 feet of stull-dirt the timbers 
 are f long, 12" diameter and 30 inches apart. If reliance is 
 to be placed upon the stull, it must be of ample proportions
 
 TIMBERING ROOMS AND GALLERIES. 5$ I 
 
 and have sufficient bearing to receive the entire pressure uni- 
 formly. 
 
 Mill-holes. Steep deposits with good walls will require stulls 
 at distances apart convenient for the men to go to and from their 
 places of work. If there is sufficient waste to furnish convenient 
 footing for the men, no timbering will be required except that 
 used in lining the mill hole and the stulls over the drift to sup- 
 port the waste. Under each mill-hole is a plat, framed for the 
 gate of the chute, with loose boards over the car-track laid upon 
 three horizontal spreaders 7 feet above the track, to afford a 
 loading chute. In wide veins the mill-holes are cribbed and 
 frequently have two compartments, one being timbered for the 
 storage and the other for the passageway. 
 
 If either wall is soft, a broad slab or post is laid against it to 
 take the thrust (Fig. 207). 
 
 FIG. 211. FIG. 212. 
 
 Underhand Timbering. 
 
 Braced Stulls. When the distance between the walls is too 
 great for a single convenient-sized stull-piece, an arrangement 
 of shorter sticks, indicated in Fig. 208, is common. A wedge or 
 plank is braced against the walls and extends longitudinally 
 with the drift to be covered. This is not so good as Fig. 209, 
 wherein one or two struts relieve the stull. Figs. 211 and 
 212 are for flooring in underhand work, and give support also 
 to the stull-dirt overhead. In Fig. 209 the saddleback, or strain- 
 ing-beam, carries the load. Not infrequently the caps may 
 be supported by struts in flat seams, like Fig. 213, or by a single 
 centre-prop in double-track gangways and slopes (Figs. 214
 
 552 
 
 MANUAL OF MINING. 
 
 and 215); but, besides taking up room, they are the cause of 
 too many accidents. In fact, much depends upon the clever- 
 ness of the men in setting the timbers to the best advantage. 
 For example, a curved stick is beneficially placed if used as 
 shown in Fig. 204. It then becomes an arch. The temptation 
 
 FIG. 213. FIG. 214. FIG. 215. 
 
 The Use of a Centre-prop. 
 
 to cut and notch and spike should be restrained, or the continuity 
 of the fibres will be destroyed. The use of wedges should be 
 avoided unless care be taken to distribute the pressure over the 
 entire area of the post or stull. 
 
 FlC. 216. Bevelled Joints. 
 
 Framed Sets. The props simply support the roof or the 
 stull-dirt and receive longitudinal pressure only. When the stress 
 is also from the sides, frames are made in sets composed of 
 a cap or collar resting on two posts or legs studded on a sill 
 or sleeper. The trapezoidal form is stronger than the rectangular, 
 and is equally serviceable for carway. This form is susceptible 
 of many varied modifications of shape, frame, and joints. 
 
 Joints. The joints employed in gangway sets where no move- 
 ment is expected are the flush or butt (Figs. 218 and 219), cut
 
 TIMBERING ROOMS AND GALLERIES. 
 
 553 
 
 with precision. Whether the sticks be round or square, the 
 joints should be flat. Never should a round cap be made to rest 
 in the hollow of the post (Fig. 217), for the fit cannot be made 
 perfect nor the splitting of the post averted. The cap should 
 be shouldered to bear flat on the leg (Figs. 218 and 219). When 
 the cap receives vertical pressure only, its entire width bears on the 
 FIG. 217. FIG. 218. FIG. 219. 
 
 Flush-joint 
 
 legs, as in Figs. 218 and 221; if the pressure is partly from the 
 sides, the joint is dressed to form as in Fig. 219; for Fig. 220 the 
 lagging and backing must be firm. The prop and collar-joint 
 (Figs 222 to 225) are simple and effective; the bevel-joint is not 
 uncommon in mining- work; Fig. 267 is an elaboration of it, seen 
 
 FIG 220. 
 
 FIG. 221. 
 
 FIG. 222. 
 
 Shoulder-joints. 
 
 in large tunnel- work, but is a very injudicious concentration of 
 pressure at one point, the avoidance of which is the very 
 design of framing. The mortise and tenon joint is rare in un- 
 derground work, except in pump-rooms and the like. So, also, 
 there is little use for the scarf-joints, unless perhaps in building 
 beams for arch centres. Wedges and head-blocks are essen- 
 tials in the tightening of frames and to lengthen timbers. Their
 
 554 
 
 MANUAL OF MINING. 
 
 removal eases the work of reclaiming the sticks whole. The 
 joints ought to be tarred for effective preservation. 
 
 Except as clamps, dogs, staples, bands, and spikes, little 
 iron is used in underground work perhaps i Ib. for every 100 
 cu. ft. of lumber placed. Timbers near to blasting operations 
 are often fastened, and bands used around them to prevent split- 
 ting or derangement; otherwise the use of iron is not commended, 
 
 FIG. 225. 
 
 FIG. 226. 
 
 Two-piece Sets. 
 
 except as auxiliary fasteners, for centres, etc. Iron rusts rapidly 
 when in contact with ligneous matter. 
 
 The Dimensions of Sets. Sets are of a clear height of 5 feet 
 6 inches or 6 feet 6 inches, and a width dependent upon the num- 
 ber of compartments. A single way is about 4 feet wide, though 
 this leaves little spare room. The narrowest heading in which 
 miners can work conveniently is 3 feet wide and 5 feet high. A 
 width of 5 feet at the bottom, and at the top of 4 feet, is ample 
 for all purposes of a single way. A double way should be 9
 
 TIMBERING ROOMS AND GALLERIES. 555 
 
 feet wide for men and cars, and a three-compartment way a 
 little wider. 
 
 Though the total cost per lineal foot may be somewhat 
 greater f6r a wider drift than for one smaller, the cost per cubic 
 yard of broken rock is less, and the difference in the cost of tim- 
 bering is slight, but the gain in rapidity of driving markedly 
 favors large tunnels. Again, as fully 12 per cent of the mine 
 fatalities are from crushing by cars, not only should ample room 
 be provided for passageway for the men, but numerous niches 
 for retreat. 
 
 In the standard coal-seams a great height is not possible, 
 
 FIG. 227. Reinforced Two-piece Set for Wide Vein. 
 
 though not infrequently the roof is ripped to secure mule height. 
 When the gangway carries a ventilator-box or gutterway, greater 
 height is required, and the timbering for the purpose is illustrated 
 in Figs. 228, 230, and 231. The diameter of he timbers used 
 is as small as will serve the purpose, though it is often 1.4 
 inches. Large timbers are often needed, but not placed, because 
 inconvenient to handle. Instead of the building of thicker pieces, 
 the sets are placed nearer than the average 3 feet 6 inches; 
 skin to skin is not unusual in shattered ground. The height of 
 slopes depends upon the mode of haulage The use of carriage 
 requires great height; with a dip of less than 40, the height is' 
 about 7 feet. For skips, the normal height of 6 feet will do
 
 556 
 
 MANUAL OF MINING. 
 
 Gangway Sets. Four pieces usually constitute the frame, 
 but the sill may be dispensed with where the floor is good. In 
 slopes sills are essential, wedged into place to secure the track- 
 way. In laying the sills the trench is dug lower in the centre 
 than at the ends, so they will not break. Upon their ends the 
 
 FIG. 230. 
 
 FIG. 231. FIG. 232. 
 
 Three-piece and Four-piece Sets. 
 
 posts rest to support a cap, which is wedged into bearing by 
 blocks and packing. 
 
 Where a level road is driven through firm material, only the 
 roof of which needs support, the post and collar form of tim- 
 bering (Fig. 236 or 223) will suffice. In pitching seams a variety 
 is employed, iis in Fig. 232, when the vein and country rock are 
 sound and the hanging-wall soft. With a poor roof and firm 
 vein- matter one leg is floored (Fig. 227), and the other rests on 
 the vein. This is also seen in coal regions. Fig. 233 is a strong 
 form for pressure from sides and top. For such conditions the 
 arched form of short timbers may be utilized as in Fig. 204. 
 
 Sets are laid close together, or the distance between them is
 
 TIMBERING ROOMS AND GALLERIES. 
 
 557 
 
 Timbering Loose Ground. Should the vein-matter be too 
 loose to stand up, the gangway is lagged against a long brace 
 FIG. 233. FIG. 234. 
 
 Wide-vein Timbering. 
 
 (Fig. 227). In wide, soft veins a similar idea is employed, as 
 in Fig. 234. 
 
 Gob-roads are timbered only at the roof, by the lagging of 
 caps, resting on dry pack-walls. Occasional chocks will give 
 
 FIG. 235. Timbering a Gob-road. 
 
 greater consistency to the whole (Fig. 235). But as the sub- 
 sidence of the roof cannot be prevented the road cannot be kept 
 open without repairs until the gob is pressed solid. Meanwhile 
 the maintenance is a serious item. The plan of keeping a road 
 open along the face of coal through waste is never to be com- 
 mended. The mode of timbering alluvial gold-mines, called 
 "blocking," is said by the inspector of mines at Sandhurst to be 
 perfectly safe. The prop and collar system is used while pass- 
 ing the chain pillar on either side of the gangway; beyond, a 
 modification, in which each cap is, in turn, made to rest upon 
 the caps of the preceding sets, already built, and upon the collar
 
 558 
 
 MANUAL OF MINING. 
 
 of the next. The timbers are 8 inches in diameter. In very 
 soft ground each cap has two posts for its support. 
 
 Timbering Roads in Large Deposits. This is difficult. The 
 gangway takes only a small portion of the width of the vein; 
 the rest, if firm, is left to stand, or if loose, packed with waste. 
 FIG. 236. FIG. 237. 
 
 Special Methods of Frames. 
 
 In the Great Devon Consols mine the compartments are used 
 one for "attle" (waste), and the other for travelling (Fig. 237). 
 The vein is 22 feet wide; the stulls are 20 inches and 18 inches. 
 In a 24-foot vein a sole-piece of 24-inch timber and struts of 
 18 inches with a longitudinal piece at the apex, made a very 
 FIG. 238. FIG. 239. 
 
 Reinforced Timber Sets. 
 
 strong frame (Fig. 234). With 3-inch plank lagging, from 10 to 50 
 fathoms of attle are carried. The hitches are cut 18 inches 
 deep. In Southern France, with great pressure from the roof 
 and for supporting heavy waste, Figs. 238 and 239 are much 
 seen. In many cases an arch of vein-matter 10 feet thick remains 
 untouched from the stope below. With a system of filling this 
 arch is subsequently recovered. In fact, without filling, no large 
 deep mine can be held by timbers.
 
 TIMBERING ROOMS AND GALLERIES. 
 
 559 
 
 FIG. 240. Square-set Timbering.
 
 5 6 
 
 MANUAL OF MINING. 
 
 If the vein-matter is very soft, the character of the framing 
 must be entirely altered. In the Austrian salt-mines the prob- 
 lem is very difficult, because the material assumes the nature of 
 a fluid. In rock that decomposes upon exposure to the air the 
 timbering involves some elaborate form of framing. The gang- 
 ways in a vein or bed may then be a component part of the square- 
 set system (Fig. 240). Where ground has a tendency to swell, 
 the only way to save the timbers is to ease up the ground behind 
 
 FlG. 241. Heavy Timbering for Galleries. 
 
 the timbers from time to time until the ground settles to its nat- 
 ural state. The swelling can neither be prevented nor resisted.. 
 For shifting ground, the style of Sutro-tunnel timbering (Fig. 
 241) is preferable to that illustrated in Fig. 268, which has proven 
 insecure. 
 
 In relieving timbering and tunnel linings of the result of con- 
 tinual movement an admirable arrangement consists in driving; 
 a lateral tunnel or two with heavy timbering, but having an open
 
 TIMBERING ROOMS AND GALLERIES. 561 
 
 face to serve as a safety-valve through which the excessive 
 pressure will be expended. 
 
 Few cross-cut tunnels require timbering further than 50 
 feet or so from the mouth, for the ground stands well. In pitch- 
 ing rock and porphyry the roof should be heavily braced. 
 
 Lagging. If the spaces between the sets cannot be left open, 
 the sides and top are lagged with plank or "slabs" from the 
 
 FIG. 242. Iron Lining for Double-track Gangway. 
 
 saw-mills. These are driven close with the flat side out, though 
 they are better reversed. Lagging shou d never be very strong; 
 the slabs are always weaker than the members of the frames, 
 and serve merely to prevent the fine rock from sifting into the 
 gallery and leaving an open space that gives opportunity for 
 movement, which if once begun can never be resisted. A 
 very slight movement produces sufficient pressure to break the
 
 5$ 2 MANUAL OF MINING. 
 
 lagging, and thus relieve the costlier work. These are readily 
 replaced on occasion. In many cases round poles are used, 3 
 or 4 inches in diameter, overlapping two sets. If the roof 
 does not run, a few logs suffice (Fig. 227). In soft ground or 
 under poor roof the men are protected and advance made by 
 "poling" ahead of the face (Fig. 271). The open space between 
 the lagging and the rock is packed with waste or wedged per- 
 fectly. Brush piled back of the lagging holds up the "smalls" 
 well. 
 
 When the capping over the firm ore of a flat-bed is not good, 
 timbering may be saved by not stripping the entire height of 
 the vein, but leaving a layer of it for roof to be removed later. 
 
 The Use of Iron Underground. Of the relative merits of 
 iron, wood, and mason y much is heard. Suffice it to say, that 
 in Europe, w r here the utmost care is taken to preserve the timber 
 by replanting for each tree cut down iron and masonry 
 are put to extensive service. As to props of metal or of wood, the 
 number would be the same more for a brittle, less for a flexible 
 roof; and whatever the condition of the roof, the size of the 
 metal props would be nearly the same Iron props are of the 
 -f or O cross section, set on the thill or upon a foot-block, to 
 be drawn by lever or bar and chain Jack-screws have been 
 used, but their expense is too great for their general use. Cast- 
 iron props, auxiliary to pack-walls, 9 feet apart, 5 feet long, 
 4 inches outside diameter, weighing 150 Ibs., have been employed 
 in collieries; their use is emphatically stated to be cheaper than 
 wood. 
 
 Levels are not infrequently lined with iron tubbing similar 
 to that adopted for shafts 
 
 The increased price of timber for underground work has in- 
 duced the use of steel. The common structural sizes of steel 
 are used in the galleries. The sets are usually of the three-piece 
 lype and the structure is of the strap type of connection. I beams 
 or channels are used for the caps and posts, the latter being 
 bevelled for the batir desired. The connections are made by 
 angle straps, bolted or riveted into place. The foot of the
 
 TIMBERING ROOMS AND GALLERIES. 
 
 563 
 
 posts is usually set into a cast shoe upon a cast base. In some 
 mines the pieces are connected by pins instead of angle straps, 
 the latter being cheaper but more difficult in connection. In 
 choosing the size of the members the heaviest shapes are usually 
 taken in order to obtain the thickest possible web and thus a 
 small size of the pins. When channels are used in pairs with 
 gas-pipe separators, they have a special socket for the foot. 
 
 Masonry Lining for Roads. The use of masonry for drifts 
 and tunnels is very common. Where timber is rapidly destroyed, 
 or where the pressure is too great for rectangular frames to be 
 economically employed, the arch is pressed into service. On the 
 other hand, good stone must be plenty and cheap. It is laid 
 dry or cemented, with the walls straight or curved in the top 
 arch. The principle usually followed is to turn the arch out- 
 FIG. 243. FIG. 244. FIG. 245. 
 
 The Segmental Arch as a Stull. 
 
 ward against the direction of pressure. The top and the floor 
 may be arched and the sides in four separate curves, or they are 
 combined into one geometrical figure, either the egg shape, ellipse, 
 or circle. In some mines the masonry is built exclusively in all 
 the openings, dispensing entirely with the use of timber In others 
 only the main underground arteries are walled. Generally the 
 masonry is built after preliminary framing has served for the 
 exploration work. The walls may be continuous along the 
 entire length of the gallery or may be in sections where the wall- 
 rock is not strong enough to dispense with auxiliary support. 
 The sides of air-crossings require walling, as also the spaces 
 where ventilating-doors are fixed between their posts and the sides 
 of the roads. Whatever the form of the masonry or lining, it is
 
 504 
 
 MANUAL OF MINING. 
 
 indispensable that the spaces back of them shall be filled with 
 waste rock and contain no decomposable material. 
 
 Where the roof is firm and the sides only are weak, straight 
 walls are built of stone or brick irom a foundation a few inches 
 
 FIG. 248. 
 Masonry Arches on One Soft Wall. 
 
 below the floor line to the roof on either side (Fig. 252). If 
 there is a slight pressure from the sides the walls may be inclined 
 or receive a batir to give a secure base. Upon these walls timber 
 caps or struts covered with lagging are frequently employed 
 to support the roof instead of leaving the roof undisturbed.
 
 TIMBERING ROOMS AND GALLERIES. 
 
 565 
 
 The arch is preferred for all permanent ways. It may be 
 segmental or semicircular for the support of waste in stopes, 
 according to the weight to be upheld. They are seldom more 
 than two bricks thick, and when properly laid should be able to 
 withstand all the pressure which emergencies require. 
 
 In Figs. 246 to 248 are illustrated arches when the vein-matter 
 and the hanging-wall are too weak to be self-supporting. Firm 
 
 FIG. 249. Arch Centre. 
 
 FIGS. 250 and 251. Arch Failures. 
 
 seats for the arch must be provided or imposts are built, as in 
 Fig. 247. 
 
 Arch Centres. Whatever the procedure, arches are built on 
 centres and by template, for invert and walls. The centers 
 should be made of light, small, easily framed sticks, that are 
 not so close as to interfere with work, yet strong enough to sup- 
 
 FIG. 252. Walling Gob-roads. 
 
 port the thrust that may fall on them when the tunnel-timbering 
 is removed. Its shape may be whatever is the most convenient 
 for the traffic. The elliptical linear arch is, however, the form 
 most commonly adopted, the side and roof comprising the upper 
 part of the ellipse, which is closed below by a segmental invert 
 arch, with the springing lines on horizontal faces. In stratified 
 rocks, the strongest form for the roof is that of a pointed arch.
 
 566 MANUAL OF MINING. 
 
 Sometimes in solid rock the horseshoe form is used for the top 
 and sides, the floor being level. 
 
 The two accompanying figures, 250 and 251, may be in- 
 teresting as suggesting the places of weakness with the given con- 
 ditions of pressure. If the excessive pressure is from the top 
 and any opportunity for bulging is given, the collapse will take 
 place as shown in Fig. 251. If the side pressure is very great 
 and the roof resistance small, the break occurs at the keystone. 
 
 Tubular Walling of Galleries. In German mines will be 
 seen examples of tubular walling of the elliptical (Figs. 253 and 
 254) or inverted oval form, the choice between them for greatest 
 strength being still an unsettled matter. The latter gives greater 
 width at the bottom and smaller area for pres- 
 sure at the top. Masonry cannot be built unless 
 the ground is previously timbered, or firm 
 enough to stand while the mortar is drying. 
 In soft ground the level is driven by spilling, 
 which can only be replaced by the masonry 
 
 FIG. 253. retreating toward the shaft. When the tempo- 
 Masonry Walling. 
 
 rary service of the timber has been accomplished, 
 
 and masonry is to be substituted, the uprights are cut at the foot, 
 the sills and spilling laths (Fig. 273) removed, the bottom arch 
 is made first, side walls next replace the posts, the cap being tem- 
 porarily supported on props, the centre set up, and the top laid. 
 Figs. 255 and 256 represent a masonry walling employed where 
 the floor is sound. 
 
 Dams. Dams for keeping back water are either straight- 
 backed (Fig. 257) or arched. 
 
 The use of iron is advocated and receiving ready acceptance 
 in mining- as well as tunnel-work. The life of timber is short, 
 its resistance low, and the component parts of the frame must 
 be rigidly connected. The ordinary constructive forms of iron 
 are applied in the ordinary way for columns, caps, or arches. 
 
 Timbering Soft Ground. In soft ground, which is liable to 
 run, some form of stout framing is indispensable. Indeed, ore- 
 bodies of large, irregular dimensions arc peculiarly adapted to
 
 TIMBERING ROOMS AND GALLERIES. 
 
 567 
 
 some form of cribbing or square sets. The latter is a natural 
 extension of the system of running contiguous parallel drifts 
 in the ore in two or more stories. In the creviced matter of the 
 Comstock mines, in the very poor ground of the Lake Superior 
 region, in iron-mines where pillars cannot be trusted, in the 
 FIG. 254. FIG. 255. FIG. 256. 
 
 Stone Walling for Gangways. 
 
 rotten lead-ores of the Leadville beds, and in the deposits at 
 Butte, Montana, an extensive system of framing is in use and 
 has found ready acceptance in various sections of the world. As 
 a distinctive method of mining, it was referred to in page 47, 
 and though many properties are substituting a filling method, 
 it still retains a hold on the mining fraternity that qualifies it 
 
 FIG. 257. Dams. 
 
 for a place here. Still, several causes hasten the decline of its 
 popularity. 
 
 The plan is devised for soft ground, and is equally well adapted 
 to large openings in rock with a small tendency to cave, slide, or 
 swell. The system is a simple and natural extension and refine- 
 ment of the practice of running contiguous parallel drifts in the 
 ore and in two or more stories when the thickness of the ore 
 required it.
 
 5 68 MANUAL OF MINING. 
 
 The Square Set. The square set in stoping requires vast 
 quantities of timber, and its framing is also expensive. Never- 
 theless, its adaptability to all forms of excavation commends it. 
 The timbering of shaft, slope, and slopes may all be rigidly con- 
 nected in one system with facility. Though there are objections 
 to it, nevertheless it frequently proves in the. end to be cheaper 
 and more economical. The open cells of timber may be inserted 
 in the event of an unexpected thrust. 
 
 The set is made up of posts, cap, girt, and frequently a brace, 
 which may be the middle number of a set of three pieces form- 
 ing the letter N inside of the main set. The cap usually is 
 placed across the deposit, while the girt, also resting on the post, 
 extends longitudinally with the roof of the ore-body. They are 
 usually framed for 7 feet of clear height to allow of reinforce- 
 ment sets being placed later, and still to leave ample passageway. 
 The posts are 5 feet apart in the clear. 
 
 The usual plan consists in driving the heading along the 
 level and near the centre of the deposit, from which is extended 
 the timbering. Two sill-pieces and girts are placed, then the 
 posts; after which a cap across the drift and girts connects this 
 set with the one last placed in position. The frame is block?d, 
 plumbed, and wedged against the back of the drift. As fast 
 as excavation proceeds on either side of the heading, one- post 
 sills are laid on the floor and penned to the two-post sills, a stull- 
 girt, then a post fitted on that, the cap and girt holding its top 
 in position. As soon as wedged against the top and the side, two- 
 inch plank roofing is laid from cap to cap on top as a floor for 
 the upper set. 
 
 This continues to the foot-wall, along which the sets are car- 
 ried by means of a cap-sill, which combines the functions of the 
 two pieces as shown in the illustration. From this a new line of 
 sets is carried up indefinitely. 
 
 Thus each one of the full square sets encloses a rectangular 
 volume or cell. Each post supports two cross-sills and two caps 
 and rests upon four sills (Fig. 258). Note also in Chapter III, 
 Part I, the method of square-set timbering.
 
 TIMBERING ROOMS AND GALLERIES. 
 
 5 6 9 
 
 The frames are built up as fast as the work is opened, unless 
 it happen that the ground will stand a while, until the timbermen 
 can attend to the chamber. In ore that will not remain in place 
 long enough to advance a set, an intermediate false set of cap 
 and uprights supports lagging overhead until the men reach the 
 full length of a cut. If the ground will not allow of this advance, 
 it runs, and only caving or filling is admissible. 
 
 FIG. 258. Sc 
 
 The method of procedure varies in different regions, or per- 
 haps with the nature of the ore. The lower floor is worked out 
 first, the bents being added on at the right, left, and ahead; 
 after which the next tier is set in the same manner directly over 
 the first. A species of overhand stoping is sometimes employed 
 whereby the floor tier is progressed only a set or two before the 
 next upper is advanced, to be followed later by another set above, 
 and so on up. The order is a matter of indifference provided a
 
 57 MANUAL OF MINING. 
 
 perfect alignment is secured. Where this method is used in 
 steep veins, very careful surveying is necessary to carry up the 
 tiers of the lower level to a line with those in the upper stope. 
 
 When a sill-floor is to be laid upon a large body of ore to be 
 mined from the next lower level, the sills should cover several 
 sets to render less dangerous the work of connecting the lower nest 
 of timber with the upper. The plates against the walls should 
 also cover several sets, if the walls are bad or crumbly. 
 
 Sticks 10 inches in diameter or square are used for 7 -foot 
 sets, and some as large as 14 inches are not uncommon with 
 9-foot posts. They are dressed to a faultless fit, being cut to 
 template by some saw like the Hendey (Fig. 260). 
 
 When the permanent roof is reached, lagging is laid, wedged, 
 and packed. 
 
 Reinforcing Square Sets. Where the posts cannot be placed 
 in the direction of the greatest pressure, reinforcing sets are used 
 in heavy ground, with the planes of the inclined piece in the 
 direction of the greatest pressure. The sets then assume the form 
 of X or N. 
 
 Objections to the System. All the sets are dependent upon 
 one another, and therefore an absolutely perfect fit is essential 
 to transmit the pressures equally and to maintain the framing 
 intact. If the joints are not true, each one will have a slight 
 play and an opportunity for movement will be afforded. The 
 caps may slip off the posts. The sets should be "plumbed" 
 occasionally to watch for incipient displacement. Sometimes, to 
 prevent this form of failure, solid cribs of timber, two sets wide> 
 are built across the ore-body from one w r all to the other. 
 
 The sets may crush when the safety-valves of lagging have 
 been omitted. These consist of frames or lagging, slightly 
 weaker than the timbering, arranged to be easily replaced. Fel- 
 site or trachyte tends to swell and crush timbers, and in such 
 ground should be closely watched. At the first sign of fracture 
 of any member of the system it is advisable to at once withdraw 
 the men from that locality; for ruin spreads very rapidly after 
 the destruction of a single member of a framing so loosely con-
 
 TIMBERING ROOMS AND GALLERIES. 
 
 57*
 
 572 
 
 MANUAL OF MINING. 
 
 nected as is the square set. A chamber 90 feet high, with 13 
 tiers of timbers, was in complete ruin within thirty minutes after 
 the first sign of break of a sill. 
 
 Cribs. Rooms and abandoned large stopes are supported 
 by massive columns of heavy timbers carried to the roof and 
 filled with waste. These "cribs" are also built as an abut- 
 ment in a chamber having the wall as a back. A good founda- 
 tion is prepared, of the proper size and shape; two logs are 
 laid parallel, and upon them, in the notches at the ends and at 
 
 FIG. 260. A Crib. 
 
 the middle, three cross-sills are laid; upon them again rest a 
 pair of sills slightly inside and above the others; upon these, 
 in turn, another layer, etc. (Fig. 260). Inside of this space, as 
 fast as enclosed, waste is piled. The logs are 10 feet and 14 feet 
 long, and 12 inches to 20 inches in diameter at the butt. Crib- 
 bing may also be built with the removal of the ore in overhand 
 stoping, the waste being utilized to fill the crib. As the height of
 
 TIMBERING ROOMS AND GALLERIES. S73 
 
 these cribs increases, the area of their base is proportionately 
 increased. 
 
 These pens, or cribs, are also used for construction of dams or 
 abutments at mine-dumps, the front face consisting of long 
 face-sills parallel with face of the hill, with cross-sills projecting 
 from the hillside and resting on the face- sills. They may, or, 
 may not, be notched. The structure continues as high as desired 
 the framing being carried up as fast as the waste rock accumu- 
 lates upon those already laid in position. 
 
 Mill-holes are built of cribs 30 inches square in one or two 
 compartments. 
 
 It is questionable if there is any choice between the square 
 set and cribbing in large rooms. The former is well suited to 
 turning off into small cells at the ends of the room, but it is danger- 
 ous if side pressure exists. Cribbing will never do in very soft 
 ore under brittle roof. This arrangement constitutes a very 
 strong "made" dump, where the rock is not permitted to roll 
 away freely, as, for instance, on account of contiguous buildings 
 down the hill. 
 
 Protecting Underground Chambers. Underground chambers 
 intended for steam-pumps, hoisters, stable, etc., are built in any 
 suitable shape that provides sufficient room, and, being large, 
 require great skill to utilize framing or walling materials to the 
 "best advantage. Undoubtedly an arched roof will give the 
 greatest resistance and strength, and masonry is therefore sug- 
 gested. Besides, the hot, damp atmosphere of steam would 
 rapidly destroy timber. Still, the latter is more convenient than 
 masonry, in heavy sets of three-piece or five-piece arches, and 
 spliced. They are laid close together, lagged over, and packed 
 to prevent wedging apart. The ventilation of these rooms 
 should receive special attention. For air- compressors, engines, 
 and coal-cutters an enlarged level will do, with some stouter caps, 
 railroad rails, and I beams on the wall-posts. 
 
 Timbering Landings. The timbering plats and landings 
 must vary with the character of the intersections. The frames 
 must support one another as well as the country rock, and should
 
 574 MANUAL OF MIX IXC. 
 
 afford firm fastenings for the plats and doors used for the landing 
 of the cars and buckets. The level or gallery should be widened 
 near the shaft, to give room for sidings, storage closet for powder 
 and steel. 
 
 The masonry lining of the shaft is supported by an arch to, 
 or by lintel on, two posts or walls at the sides of the gallery land- 
 ing. The former gives a high opening for landing the buckets. 
 
 Mill-holes carried up with the waste are solid cribs of 30 inches 
 square for a manhole, and may also have a compartment for 
 sliding ore (Fig. 87). Either cordwood or sawed blocks are 
 used for the lining. The latter plan gives them greater dura- 
 bility, for the abrasion is along their cut faces. 
 
 The timbering of slopes is similar to that of galleries, except 
 that greater care is taken in cutting. 
 
 Timber-cutting Tools. The timberman's tools are few. They 
 need be only hatchet, hammer, and wedge, with a bar and chain 
 for casual work. The timber should be delivered below ready 
 for insertion. A sawmill at the surface is now as much a com- 
 ponent of the surface improvements as is the boiler. 
 
 Mine-timbsr Framing-machine. In Fig. 29 is illustrated a 
 machine which is designed to saw tenons upon timbers of stand- 
 ard size. It consists of a number of adjustable saws which 
 may be arranged at any distance apart to produce the desired 
 dimensions of the head on the timbers. By a machine set to a 
 given pattern, all timbers are finished to an exact shape, which 
 when placed in position ensures an alignment of the framing, 
 the joints being accurate and the contact perfect. Security is 
 thus attained and the life of the frame is prolonged. 
 
 Though the work of timbermen cannot accurately be stated, 
 the following is given as the experience of an "old miner" for 
 one day's labor: 
 
 Making about TOO feet of wooden ladders; 
 Placing 65 sq. ft. of close shaft-lining down to a depth of 70 ft.; 
 
 40 " " " " r " " " " " "200 " 
 
 Cutting and dressing mine frames, 50 running feet; 
 Framing mine-sets, 40 ' '
 
 TIMBERING ROOMS AND GALLERIES. 575 
 
 Dressing sills, 30 running feet; 
 Making 60 poling planks; 
 " 100 " wedges; 
 Setting up 20 pairs of props and head-blocks. 
 
 REFERENCES. 
 
 Timbering, M. & M., Vol. XXIII, 117; Theory of Timbering, M. & M., 
 Vol. XX, 495; XXII, 566; Min. Inst., X, 726; Dams, M. & M., Vol. XX, 
 445; Structural Timber, Comparative Tests of Bracing, Prof. E. Kidwell, 
 L. Sup. Min. Inst., IV, 1896, 34; Timbering of Silver and Gold Mines, J. T. 
 Freeland, Coll. Eng., 1895, 100; Timbering: Economical Use of Timber in 
 Mines, H. W. Halbaum, Coll. Guard., Aug. 1896, 407; Mine Timbering, J. 
 Clark Jefferson, Chest. Inst., VII, 270; Mine Timbering, C. W. Swallow, 
 Sci. Am. Sup., Sept. 8, 1894; Timber in Mining Regions about Prescott, 
 Arizona, J. F. Blandy, Amer. Inst. M. E., XI, 291; Timbering in Comstock 
 Mines, R. W. Raymond, Amer. Inst. M. E., VIII, 91; Mine Timbering in 
 the Old Ironsides and Knob Hill Mines, H. P. DePencier, Can. Min. Rev., 
 July 31, 1902; Drift Timbering, W. H. Storms, Min. & Sci. Press, Oct. 10, 
 
 i93- 
 
 Timbering in German Collieries, Coll. Guard., Oct. 24, 1902; Coal-mine 
 Timbering, H. W. Halbaum, Coll. Eng., Feb. 1897, 303; Drawing of Timber, 
 E. B. Wain, M. & M., XVIII, 159; Colliery Timber and Timber Calculations, 
 George Johnson, Coll. Guard., Dec. 8, 1899; The Different Methods of Sup- 
 porting the Roof and Sides and of Resisting the Thrust of the Floor in Coal 
 Mines, a Prize Essay, Wm. Bradford, Coll. Mgr., 1896, 564; Timbering 
 Props in Coal Mines, Prize Essay, Coll. Mgr., Nov. 20, 1896, 564; Falls of 
 Roof and Sides, Accidents, Prevention, Mine Insp. Coll. Mgr., 1894, 22. 
 
 Creosoting of Timber Sables d'Olonne, Coll. Guard., LXXI, 886; Wood 
 Preservation, H. Flad, Bull. No. i, U. S. Department of Agriculture, 66; 
 Behavior and Decay of Wood, Bull. No. i, U. S. Department of Agriculture; 
 The Best Woods for Mines, B. E. Fernow, Am. Inst. M. E., XVII, 269; The 
 Timber-preserving Plant of the Great Northern R. R., Gaz., May 30, 1002; 
 Timbering Supporting Roof and Sides (Prize Essay), Coll. Mgr., Nov. 20, 1896, 
 564; Timbering: Economical Use of Timber in Mines, Coll. Guard., Aug. 
 28, 1896, 407; Preservation Method, Calif. Bureau, i3th Report; Preserva- 
 tion of Timber, Calif. Mineralogist, i3th Report, 1896, 3d Biennial, 647; 
 Methods of Preservation, Illustrated, and Table of Solutions Used, Amts., 
 etc., The Transit, I, 63; Treatment of Timber for Mine Use, R. Martin, 
 M. E., X. 
 
 The Standard Chute Discharge for Mill Holes, M. & M., Vol. XXIV, 
 
 254-
 
 5?6 MANUAL OF MINING. 
 
 Steel Shaft-lining, M. & M., Vol. XXIII, 128, 155; The Use of Iron Sup- 
 ports in Main Roads of Mines, instead of Masonry or Timbering, G. Meyer, 
 Vol. XXXVII, 221. 
 
 Systematic Timbering in French Mines, Coll. Guard., Vol. LXXXIV, 
 1174; Setting of Pit Props, Coll. Guard., Vol. LXXIX, 500; Systematic 
 Timbering, Coll. Guard., Vol. LXXIX, 1075; Erecting Dams, Coll. Guard., 
 Vol. LXXXI, 569; Timbering German Mines, Coll. Guard., Vol. LXXXI, 
 693; Cribs in Collieries, Coll. Guard., Vol. LXXXI, 570.
 
 CHAPTER IV. 
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 
 
 Levels. These horizontal openings in metal-mines receive 
 an amount of care commensurate with their importance and 
 service. Like the levels, which also are driven through hard 
 rock, they are classed as "dead work." The difficulties of driving 
 are not so great as for shafts, as their area is smaller and seepage 
 water gives less annoyance. Their location is fixed by condi- 
 tions discussed in Chapter I. The drifts and cross-cuts of a 
 given mine are of the same dimensions. They serve for one track, 
 and only at the landings are wider to accommodate a second 
 track. They are rarely timbered, for the country rock will stand 
 without support for the entire period of their utility. If they 
 require protection at all, a masonry or iron lining is used. 
 
 Adits and 'Levels in the lode are secured as shown in the 
 previous chapter. Usually the stull and lagging suffice for the 
 average length of time the drift is kept open. Gangways and 
 galleries in coal-mines have a larger area than levels, and being 
 exposed to more treacherous conditions, possess better examples 
 of the timberman's art. Tunnels for railway or drainage pur- 
 poses must be walled. 
 
 Before locating a tunnel of any importance a careful study 
 of the ground is requisite. All the data obtainable from geolog- 
 ical reports, borings, etc., should be availed of. The character 
 of the strata, their pitch, and the direction of the subterranean: 
 drainage should be known before the character of timbering and 
 the dimensions and amount of timber required can be determined- 
 
 577
 
 S7 MANUAL OF MINING. 
 
 In Drinker's "Tunnelling" will be found a discussion of the 
 geological conditions affecting tunnel locations. 
 
 The Alignment of a Tunnel. The axis of a tunnel is kept in 
 a vertical plane by the use of three plumb-bobs (Fig. 261), and 
 its horizontaiity is tested by a long spirit-level. A grade of two 
 feet in a hundred is given to the floor for purposes of drainage. In 
 collieries the grade is involved with the proposed system of mining. 
 
 The alignment of a gentle incline is similarly conducted, 
 though four adjustments are necessary. A saw-cut is made 
 across each sill and 8 inches from the end. Two hubs are set 
 in the sills on this line by transit within 50 feet of the face of the 
 drift or incline. They are lined by a string which must pass 
 through the saw-cut notches. The grade stick is now placed in 
 position to depress or raise the sills to the required inclination, 
 after which they must be laid level and across the drift at right 
 angles with the line. 
 
 FIG. 261. Aligning a Tunnel Drift. 
 
 Rock Characteristics. Hard rock presents no serious dif- 
 ficulties to the miner beyond the time and cost of drilling, for 
 it usually affords a good roof and if uncre viced is dry. Granite 
 dolomite and gneiss are of this character. Slates and shales 
 are bad and require arching or block timbering. Porphyry is 
 treacherous, because, though hard when first opened into, it 
 soon decrepitates on exposure. Some of the limestones and 
 sandstones are porous and wet. Clay seams are bad primarily, 
 besides being watercourses for the upper porous strata. 
 
 Driving in Creviced Rock. In uncreviced rock the order 
 of the breaking is immaterial, so long as a good bench is had 
 to shoot from and a favorable working-face is obtained for the
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 
 
 579 
 
 next shot. The crevices and planesi n stratified ground materi- 
 ally assist the shooting. If the seams pitch toward the face 
 (Fig. 262), the upper holes are fired first, or in high galleries 
 the central holes will make a good bench to shoot to from top 
 to bottom. If the cleavage is not marked, or away from the 
 face (Fig. 263), the bottom holes are fired before the upper ones. 
 If electric firing is adopted, the sequence of shots is of little im- 
 portance. Three men form a gang in driving, and can manage 
 the ordinary-sized drift. In the gangway of 7X7 two machines 
 or two pairs of cutters have ample room. Their progress cannot 
 
 FIG. 262 
 
 Placing Holes in Creviced ROCK. 
 
 generally be stated. In hard rock a foot a shift is fair, while 
 soft rock can be penetrated at the rate of three feet a day. In 
 shaly ground greater progress might be made did it not require 
 good timbering. In soft ground the advance depends upon 
 the skill of the timberman. 
 
 Driving by Air-drills. It is a matter of indifference where the 
 holes are placed for explosives if machine-drills and simultane- 
 eous firing are employed. A drift of ordinary size will easily 
 accommodate one drill, wbile two can advantageously work in 
 a heading 10 feet wide, obtain more angling holes, and advance 
 more rapidly. In railroad tunnels four are simultaneously drilling. 
 Hand-work is much more depended upon for the driving of 
 gangways, but machine-drills are becoming popular for tunnels. 
 Undoubtedly the progress by machine is greater than by hand, 
 while the cost is nearly the same per lineal foot. Where the drills 
 and air-compressors can be put to use after the mine i? opened,
 
 580 MANUAL OF MINING. 
 
 it certainly would be advantageous to employ them in the pre- 
 liminary and development work. 
 
 The consumption of steel in medium ground is about 25 
 cents per cubic yard removed, though pink quartz will dull 150 
 bits to the hole, and the blacksmith will consume more steel 
 than the rock. The consumption of powder is about $i per 
 cubic yard of medium-tough rock removed. This amount will 
 ry with the area of the face and whether the breaking is done 
 
 simultaneous or single shots. 
 
 Driving Slopes. The operations of sinking and the timbering 
 of a slope are similar to those of level driving, except that the 
 sill is indispensable to the set. It must be well bedded and 
 let into the rocks on each side to prevent the roadway from slip- 
 ping down hill. Ofttimes it is stayed by plugs driven into the 
 floor. The sets are also braced against one another by longitu- 
 dinal studs between the posts at their head. 
 
 Driving Tunnels. Levels and drifts are worked over their 
 entire face. But when tunnels are to be driven of a height greater 
 than 8 feet, they must be broken in benches. Railroad tunnels are 
 of this description, and in hard rock may be driven without any 
 temporary timbering, the benches being attacked like stopes 
 (Fig. 27), where the numbers express the sequence of openings. 
 Frequently, in driving long tunnels, the drift No. i is pushed as 
 fast as possible in order to make connection with a shaft or a 
 similar drift approaching from the opposite direction. This com- 
 munication is for ventilation and haulage purposes. 
 
 Numerous tunnels have been driven more than five miles 
 for mining purposes. The Freiberg is 24 miles long; at Claus- 
 thal is one nearly n; the Joseph II. is 9^; the Ernest August, 
 6 miles; and the Sutro, 5. These have, moreover, lateral 
 branches that enable them to subserve a great territory. The 
 Gwennap adit, in Cornwall, is said, with its branches, to attain 
 a length of 30 miles. The greatest depth below the surface is 
 not over 900 feet for any one of these. 
 
 The Use of Supplementary Shafts. Auxiliary shafts are sunk 
 at convenient points along the line of a long tunnel to its level;
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 58* 
 
 and from them the excavations are begun contemporaneously 
 with those at the mouths of the tunnel. Sometimes the shafts 
 are to one side of the tunnel line, to keep them free and clear 
 of the tunnel-work. The amount of time allowed for the com- 
 pletion of the tunnel determines the number of points of attack- 
 and the number of shafts to be sunk. The latter is also depend- 
 ent upon the cost of sinking and the hoisting through them rela- 
 tive to that of the long haul in the tunnel. These relations car* 
 be mathematically expressed when the several aspects of the case- 
 are given. From Foster's Gallon's "Lectures on Mining" the 
 following formula is taken: 
 
 wherein 17 = the total known cost of shaft, x= number of shafts 
 to be opened over a length I, 5 = the area excavated, S' the 
 section of the lining, Z) = the distance between the two adjoining 
 shafts, P and P' = the cost of haulage per lineal yard for each 
 cubic yard of rubbish and of walling material. 
 
 The St. Gothard Tunnel, 48,840 feet long, was run from the 
 two ends only; so the Mont Cenis Tunnel, 39,840; the Hoosac 
 and the Sutro tunnels with the assistance of two shafts. The 
 Washington Tunnel, 20,715 feet, had four working shafts. The 
 Rothschenberger mining adit had 18 points of attack along its 
 24 miles of length. Rziha estimates that the "additional cost 
 of running headings from a shaft is from 5 to 10 per cent higher 
 than running from portals." He also gives a unique table to 
 show that the rate of progress per month is apt to be greater in 
 long than in short tunnels. Of those less than 100 m. long, 
 29 feet advance was the average of 3; 13 between 400 and 600 m. 
 showed average of 57; 8, up to 1000 m., 114 feet, while 6 between 
 3000 and 4000 m. had an average of 219 feet, and 4 over 4000 m, 
 progressed 259 feet per month.
 
 58' 
 
 MANUAL OF MINING. 
 
 Provision for traffic is not a simple matter when it is recalled 
 that, besides the actual mining of an area (say, for example, 
 that of the St. Gothard Tunnel), 26X20, at a rate of 18 feet per 
 day, and the placing of 1000 cu. ft. of timber and 300 tons of 
 masonry per day, 750 tons of broken rock, 15 tons of lumber, 
 and 300 tons of masonry must be handled, loaded, transported, 
 .and unloaded at the same time. 
 
 As the hoisting through shafts is about half as fast as haul- 
 age through a heading, the use of auxiliary shafts is not so eco- 
 nomical as it was before the present development of machine- 
 drilling and blasting methods. 
 
 Driving a Tunnel. There are several methods of opening 
 a tunnel, one of which is depicted in Fig. 264. The two side 
 
 FTG. 264. The Order of Attack on a Tunnel-face. 
 
 drifts at the bottom may precede the main excavation; in these 
 are built the walls for the roof arch which ultimately lines the 
 tunnel. While the small drift, i, is being driven to the connec- 
 tion, the sides 2, 2 are following, and in firm ground they may 
 precede 5, 5 at the bottom by 150 feet. At this point the tunnel 
 is being lined with timber or masonry, while gangs are also break- 
 ing ground on benches 3 and 4. 
 
 If the roof needs support, the plan of work contemplates 
 one similar to that of Fig. 265, with the exception that the central
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 583 
 
 core 3, 4 remains as a support to the timbers arranged as shown 
 below. Inside of this frame the masonry may be completed, 
 after which the inverted arch, if required. 
 
 In soft ground, and especially in inclined strata, the order 
 of driving the benches and the amount and character of tim- 
 bering varies according to the dip. Circumstances and diffi- 
 culties are so diversified that no uniform infallible rule has been 
 established for the guidance of the engineer probably because 
 no system has a superiority over all others for any and all con- 
 ditions. Safety is an clement as important as time and money > 
 and often is a determinant. 
 
 FIG. 265. 
 
 Tunnelling Systems. A brief review of the English, Belgian,, 
 German, and Austrian is taken from Drinker's ''Tunnelling." 
 
 The English system is developed from the experiences in 
 the Thames River tunnel and consists in taking out the full area 
 at once, after the preliminary top-heading has been made, and 
 in supporting the roof by longitudinal top bars while removing 
 the lower section. This gives a full clear area for putting in. 
 the maSonry, after which, in material having any tendency to
 
 584 MANUAL OF MINING. 
 
 swell, the space behind the side walls is securely grouted. Though 
 built quicker and applicable in 90 per cent of cases, it is unsatis- 
 factory in very bad ground; in heavy ground it requires more 
 limber than does the Austrian, its strongest competitor. 
 
 The Belgian method, introduced after the iron shield had 
 t>een tried ineffectually in quicksand, builds the tunnel as an 
 open cut down to the springing-line of the arch. The arch is 
 ihen laid, recovered, and underpinned, the bottom removed in 
 benches, working downwards, and the abutments built. Some- 
 times the centre-core is left, though that is only done in. the 
 Trench and German modifications. The entire area is not 
 attacked at once, but divided into several benches, each being 
 worked separately. The underpinning of the arch may be safe 
 enough in hard ground, but it certainly affords a doubtful security 
 in loose material. 
 
 The German system gave rise to what may be called the 
 centre-core system. The work of excavation begins at the foot 
 of each abutment of the arch, where a small heading (Fig. 265) 
 with timber sets is first driven. In this the foundation is laid; 
 above it a second heading, large enough to build another height 
 of wall; above this another, in which the masonry is carried 
 up. The top is then excavated across, and a connection effected 
 between the two sides. In this the arch is completed without the 
 use of centres, while the roof is being supported by stulls or props. 
 The core is then removed. 
 
 In hard ground it gives cheap working, from the fact that 
 the core has several faces of attack. In soft ground it is safer, 
 because of the small exposure of roof and face; and the centre- 
 core saves timber. Its ventilation is bad, and the cost of lay- 
 ing masonry is larger than where the masons have elbow-room; 
 it is hard to securely timber, and several prominent engineers 
 have decided against it. Certainly, in soft, treacherous ground, 
 like shales and clays, its defects disqualify it. 
 
 The Austrian system admits of mining the whole area in small 
 sections. First a bottom heading is driven and afterwards con- 
 .nected to a top heading, which is finally widened to full width
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 585 
 
 for a bar-timbering which is carried down to the springing-line. 
 Sides are excavated for the walls, after which foundations are 
 prepared and the side walls built; finally, the invert arch. The 
 cross-rafter timbering (Figs. 267 and 296) used for support ad- 
 mits of the transfer of pressure, and there is no such undue con- 
 centration as in Fig. 265. This disposition of the timber affords 
 
 FIG. 266. 
 
 a greater strength, and is the leading feature. Additional braces 
 are sometimes added and the sets connected every which way; 
 but the design is to arrange the timber of each section in such 
 manner as to form an integral part of the completed system. 
 After the roof-timbers are in place, plenty of space is left below 
 for masons, and good ventilation is had.
 
 586 MANUAL OF MINING. 
 
 Comparison of the Four Systems. Each system has dis- 
 tinctive features commending it to favor under given conditions, 
 and a brief comparison may be given. 
 
 The centre-core system is emphatically inapplicable for 
 loose rock where other methods are better. It is quite expen- 
 sive in all of its operations, and is not safe. 
 
 The English system is safe enough for most tunnels that 
 are likely to be driven, and offers ample space for the miners 
 and masons to work, but it is not adapted to very bad ground 
 or where heavy pressure is encountered. 
 
 The Belgian system seems to be preferred for moderately 
 firm ground, in which it is quicker than either of the two above 
 mentioned; but it is hardly suited to running ground, because 
 of the great care which must be taken with the underpinning. 
 The cost of transportation while building is higher than in the 
 others because of the frequent interruptions caused by the trans- 
 portation scaffolding. In very loose ground it is positively 
 dangerous. 
 
 The Austrian system has more to commend it and less to 
 condemn it than the others. All the driving operations are 
 cheaply and easily conducted. The driving is least interrupted, 
 and safety is assured if the middle timbering and the longitu- 
 dinal bracing be well apportioned. The forepoling, if em- 
 ployed, would be in the direction of the tunnel axis, and the 
 masonry arching is not delayed. 
 
 Block Timbering. In Figs. 266 and 268 are illustrated the 
 timbering permanently placed in the tunnel. Each of the nine 
 blocks is io"Xio", the legs being 12 inches square. From the 
 three- rafter system of roofing the tunnel American engineers 
 developed the block-timber arching for the support of the roof 
 rock which is very loose and requires close bracing. The wooden 
 voussoirs were first used for temporary support only, until they 
 were replaced by brick or stone arching, but latterly they have 
 served as a permanent arching lined with a fire-proof sheeting 
 for railroad use. 
 
 We cannot claim any system of tunnelling as our own, for
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 
 
 587 
 
 neither the number of tunnels nor the difficulties encountered 
 are as great as in the Old World. The Austrian method is the 
 nearest approach to ours, or, rather, is the one which our en- 
 gineers have adopted with a modified framing. The mode of 
 driving and timbering is illustrated in Fig. 264. The upper 
 heading, i, with its enlargement, 2, 2, precedes the work upon 
 the "bench." With a tunnel area of 2i"X27" the heading is 
 about 8 feet high, and the bench, the remainder, attacked in 
 one section. The upper heading is timbered, rafter style, three 
 or nine voussoir bents forming a block-arch on heavy perma- 
 nent pole standards, or on the side rock if sufficiently sound. In 
 loose rock the number of bents increase and the timbering is 
 heavier. Inside of this frame arch the masonry is built. A 
 segmental arch 21 feet high by 18 wide in the clear had nine 
 voussoir-blocks 25 inches long and io"Xio" section; two 
 wall-plates, 6 feet 4 inches long, 12 inches square; two posts, 
 14' i"Xio"Xi2"; and 45 lagging, 6' 6"X6". 
 
 Fig. 268 illustrates the style of bar-timbering, which, how- 
 ever, will not permanently withstand much lateral pressure. 
 
 FIGS. 267 and 268 The Block Arching of Tunnels. 
 
 Tunnelling Soft Ground. In running ground the area at- 
 tacked is limited by the rapidity with which permanent support 
 can follow the excavation, during which provision must also be 
 made against the pressure coming from all sides. In no region, 
 not even in the Rocky Mountains, is the engineer free from the 
 liability of striking ground that may overwhelm the miner before 
 timbers can be inserted; so that the ground has to be restrained 
 up to the face as well as behind the frames, with which close
 
 S 88 
 
 MANUAL OF MINING. 
 
 lagging may suffice to prevent movement. If it is not in new 
 ground, the consistency of which will determine the details of 
 its penetration, it may be where timbers have rotted or given 
 way that a cave or a run may occur. 
 
 In alluvial or sand the method of spilling or a pneumatic 
 system is indispensable. These have been employed under the 
 waters of the Thames, the Severn, and the Mersey Rivers in 
 Great Britain and are universally employed for driving through 
 similar ground in America. 
 
 Spilling by laths is a method equally applicable to drifts as 
 to shafts. In front of one stick of a set, and behind its mate in the 
 next advancing set, pointed heavy planks 
 are driven, one set in advance of the 
 face, close together on one or all of the 
 sides from which the pressure is exerted. 
 The fore end of the plank is forced 
 down upon its set, while the rear end 
 is held against the lower side of its cap, 
 being protected from pressure by the 
 previously driven upper lath. This is the 
 method of forepoling illustrated in Fig. 269 
 The progress in soft, heavy timbered rock 
 is three times as fast as in very solid rock. 
 The "spilling" protects the fore- 
 breast by horizontal laths, as long as the 
 breast is wide, held against it and braced 
 to the nearest set. Each lath, a (Fig. 
 270), is removed in descending order to 
 permit some ground to run off; it is 
 
 advanced a short distance and braced again, b. The progress 
 depends upon the speed with which the spaces may be opened 
 and closed; as these are small, the movement is controllable, 
 and there need be no fear of sudden shock to the timbers. The 
 method is simple, and has been eminently successful under many 
 circumstances; once, with masonry lining following the fore- 
 pole, a tunnel was executed within 18 feet of a river-bed. 
 
 I it HUM 
 
 FIG. 269. Forepoling.
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 589 
 
 FIG. 270. Method of Spilling for Running Ground.
 
 59 
 
 MANUAL OF MINING. 
 
 An entirely different principle is that employed by Durieux 
 and others in Westphalia, whereby the ground was forced ahead 
 of the work and was not removed at all (Fig. 271). The walls 
 
 FIG. 271. A Picketing Method for Sandy Ground. 
 
 and roof were forepoled, but the breast and floor were checkered 
 by pyramidal pickets, completely covering the exposure. Those 
 on the face were square faced and larger than the floor-wedges, 
 which were 12 inches long and 4 inches diameter, driven by 
 mallets. The floor-pickets remain, in permanently, while those 
 at the face force the soft material ahead and advance in this 
 manner. The battering they receive renders them useless in a 
 very short while, when they are replaced. There is no material 
 to be hauled except that used in the construction, and rather 
 bad ground has been thus traversed. On occasion a short 
 lateral drift is similarly pushed to relieve the main work. Four 
 men in a drift s'X6' in morainal matter will advance 4 feet a 
 shift. 
 
 The finest piece of tunnelling is the construction of the new 
 Croton Aqueduct, where water, mud, quicksand, and all varie- 
 ties of loose soil were penetrated in an area of 676 sq. ft. Fore- 
 poling for the roof and sides, picket- spill ing for the floor, and 
 the American system with block-arching were adopted at various
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 59 1 
 
 FIG. 272. The Brunei Shield.
 
 592 MANUAL OF MINING. 
 
 points. The ground in extended places was so bad that 24- 
 inch timbers were crushed by the pressure. 
 
 Iron Shields at the Tunnel Face. Iron is a safer constructive 
 material for loose soil; and in 1825 Brunei put it to use for exca- 
 vating an area 38X22 under the Thames River. As in the spill- 
 ing method, the face was covered by laths about 3"X3"X6", 
 and for rapidity of work was subdivided into three tiers of n 
 breasts each (Fig. 272), each being protected by a cast-iron shield, 
 from which struts hold the laths d in place. One man to a cell 
 operates by taking away a lath and replacing it 3 inches in ad- 
 vance. The laths are removed successively downwards until 3 
 inches of progress has been made, after which the alternate frames 
 are carried ahead 6 inches and the performance repeated. In 
 each cell is a similar performance. The masons follow the miners 
 very closely at G. 
 
 The second Thames Tunnel was completed by the use of a 
 shield, which with pointed shoes was forced into a stiff clay at 
 the rate of 9 feet per day by jack-screws exerting a pressure of 
 60 tons on it. Three men crawled through a door in its face 
 and excavated some earth preparatory to the next move. 
 
 The Hydraulic Shield. The present practice for subaqueous 
 tunnels involves some variation of the pneumatic system. A 
 pilot-tube or caisson penetrates the soil, which is held up by 
 compressed air. The masonry, perforce, is built as fast as the 
 shield, tube, or caisson advances. In the Greathead system a 
 cylindrical iron shield, 21 feet in diameter, is thrust from the 
 masonry by hydraulic pumps under a pressure of 3000 Ibs. per 
 square inch. The front edge of the shield is a heavy ring shoe 
 to facilitate its advance into the silt. The rear end of the shield 
 encloses the masonry, which follows the progress of the shield. 
 Doors in its face are opened to allow the soft earth to enter. 
 This is promptly removed (Fig. 273). 
 
 In preparing to tunnel silt, both the weight and the vertical 
 pressure of the overlying material and the lateral movement 
 of the loose paste are to be resisted. The first is a matter of 
 determination, and the ability of the completed structure to
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 593 
 
 FIG. 273. The Hydraulic Shield,
 
 594 
 
 MANUAL OF MINING. 
 
 withstand this is also a matter of mathematical calculation; 
 but the second is the difficulty to be apprehended. The Hudson 
 River Tunnel engineers, however, satisfied themselves that the 
 tendency of the gravel to pour over the top of the tube would 
 reduce the lateral stress to the resistance of the tube. Certainly 
 their experience in subaqueous driving goes to show that the 
 tube serves better than the timbering systems as regards the pre- 
 vention of overhead settlement. The liability to settlement in 
 front of the shield may be overcome by grouting under pressure 
 at the rear of the shield between the lining and the roof. 
 
 The Anderson Pilot-tube. This is a segmental tube of about 
 6 feet diameter, strengthened inside by radial timbers, which 
 precede the main work. It is of -J-inch plates, I2"X24", riveted 
 together by means of flanges; and when a cut has been excavated 
 into the heading large enough, one of the plates is placed and 
 
 RTClN^L^^ 
 
 FIG. 274. The Anderson Pilot -tube. 
 
 held by props (often the plates arc held by compressed air during 
 the work) ; on each side other cuts are made for two more plates, 
 which are riveted to it. Rings of the pilot are thus successively 
 completed. 
 
 Around this, in small terraces, and considerably behind the 
 pilot, the main shell, 17 feet in diameter, is finished in a similar 
 manner, the plates being propped from the pilot-tube, which is 
 always braced from the masonry' that lines the shell. With its 
 progress the rear rings of the pilot-tube are removed and their 
 plates shifted to the front end. The masonry consists of six 
 courses of brick laid in cement. To reduce the volume of the
 
 DRIVING DRIFTS, TUNNELS, AND GANGWAYS. 595 
 
 tunnel that is kept under the compressed air, brick bulkheads, 
 4 feet thick, provided with two air-locks, are built every 400 
 or 500 feet. Only the two nearest the work are maintained. 
 The following references are cited: 
 
 REFERENCES. 
 
 Tunnel Shields for Driving, Eng. News, Jan. 9, 1892, 26; Shields, Eng. 
 Record, Mar. 25, 1893, 337; Lining, Jour. Assn. Eng. Soc., July 1893, 331; 
 Lining, Eng. News, Oct. 12, 1893, 289; Tunnelling with Steam Shovels, 
 M. & M., Vol. XX, 128; Alignment of Tunnels, E. R. Durham, Am. Soc. 
 C. E., Oct. 12, 1893, 289; Methods in Tunnelling in Soft Grounds, W. 
 Beaham, A. Soc. C. E., XXVI, 490; Lining Tunnels, E. R. McNeill, Engi- 
 neering News, Oct. 12, 1893, 289; Timber Lining of Tunnels, S. B. Fisher, 
 Engineering News, XXX, 268; Lining Galleries Exposed to Heavy Rock 
 Pressure, Bergassessor Jacob, Coll. Guard., Dec. 12, 1902. 
 
 The Ventilation of the St. Gothard Tunnel, Rev. Gdn. des Chemins de 
 Fer, Nov. 1899; The Cripple Creek, Colo., Drainage Tunnel, W. B. Wilson, 
 Min. & Sci. Press, Jan. 17, 1903; The Ventilation of the Simplon Tunnel, C. 
 I. Wagner, Oesterr. Wochenschr. f. d. Oeffent. Baudienst, Feb. 15, 1902; 
 Methods of Driving Tunnels in Soft Ground, Anon. Eng. News, Jan. 16, 
 1892, 64; Driving Shields, M. P. Paret, Eng. News, Jan. 9, 1892, 26; Tim- 
 ber Lining of Tunnels, S. B. Fisher, XXX, 268; and The Ventilation of 
 Tunnels, Heating and Ventilating, Serial, 1904.
 
 CHAPTER V. 
 
 DRILLING- AND BORING-MACHINES FOR EXPLORATIONS. 
 
 Classes of Machines. The penetration of rock or of loose 
 material may be accomplished by either of two operations repre- 
 sented by corresponding types of machines. By a repeated 
 reciprocating blow or concussion upon the rock, fragments may 
 be broken therefrom and a hole drilled to any depth and at any 
 rate, depending upon the strength of the machine and the resist- 
 ance of the rock to abrasion. Machines operating upon the 
 rock in this way are known as percussive machines or punch- 
 drills. The second method of penetration consists in employing 
 a rotary cutter which, with sufficient pressure or weight upon it, 
 will abrade the rock at a rate depending upon its relative soft- 
 ness and the motor power. 
 
 Percussive drills are employed to advantage upon hard uni- 
 form rock only. Rotary borers are constructed of various types 
 as to be equally applicable to either soft rock, loose ground, or 
 hard rock. 
 
 These drills or borers may be operated by directly connected 
 engines upon the same frame to drill holes of thousands of feet 
 in depth for prospecting purposes. When these machines are to 
 be used underground they are of a more portable type and oper- 
 ated by electricity or compressed air, conveyed to it from a dis- 
 tinct source. The depth of the holes made possible with these 
 is limited to a few feet. Hence their utility is confined to drill- 
 ing holes for blasting purposes and not for exploration. 
 
 596
 
 DRILLING- AND BORING-MACHINES. 597 
 
 Prospecting-machines. The machines employed for pros- 
 pecting are either of the punch-drill type represented by the 
 American oil-well rig, or the rotary types represented by the 
 "diamond" drill and other forms of tubular drills. 
 
 Bore-holes are driven for the testing of the strata; to prospect 
 for the gas; to afford an outlet for water; to pump brine to the 
 surface; or to subserve some precautionary measures in mining, 
 such as special ventilating openings, deep sump-holes to drain 
 the mine into a lower porous stratum or conduits, for tail-haulage 
 rope. These machines are also employed for drilling a series of 
 holes in connection with the Poetsch method and the long hole 
 process for shaft-sinking. 
 
 Percussive Drills. Reciprocating motion is obtained from 
 a steam- or air-piston either directly attached to the drill or indi- 
 rectly connected with it by rope. The blow may be produced 
 by the pressure of the motor fluid in the first case or by the weight 
 of the drill allowed to fall freely after having been raised by the 
 rope. In the first method the piston pressure is great compared 
 with the weight of the tool, while in the latter the force of the 
 concussion is solely due to the heavy weight of the drill. The 
 early type of punch-drill consisted of a continuous line of iron 
 rods, lengthened from time to time with the progress of the hole, 
 which was raised and lowered at the surface by a walking-beam 
 operated by a cam or other lifting device. The shock transmitted 
 through the long line of metal caused breakage of the tools, and 
 other difficulties which led to the later forms of machines in 
 which heavy drill lengths were suspended from the surface by a 
 flat or round rope and at a much more rapid rate than the 
 line of rods. Later the solid rods were replaced by several lengths 
 of hollow rods which gave greater progress with the same power, 
 and overcame the difficulties of feeding water and disposing of 
 the slimes produced during the drilling. The rods were screwed 
 together or bayonet-hooked, the former being preferred. The 
 engine may be single-acting with a reversible link motion pro- 
 vided with a sectional fly-wheel, whose weight was increased as 
 the hole deepened. It communicates motion to a walking-
 
 598 
 
 MANUAL OF MINING. 
 
 beam which is pivoted at its centre on top of two sampson- posts 
 and fastened at the derrick end to the drill-rope attachments. 
 
 Oil-well Rig. The typical American drill rig is illustrated in 
 Fig- 275. 
 
 This consists of a derrick 75 feet, or more, in height, having 
 at its crown a sheave over which the drill- rope passes to the "bull- 
 wheel," or large drum. A high derrick is required to facilitate 
 the attachment and the removal of an 
 entire string of tools. 
 
 A second rope pulley, at one side of 
 the derrick near the top, guides the rope 
 attached to the sand-pump, which rope is 
 connected with its reel. The several bents 
 of the derrick are fastened together by 
 wedges, if of .wood, with the exception of 
 the derrick floor-planks, which are bored 
 or spiked. Its construction permits of easy 
 dissection when a hole is to be abandoned. 
 
 The Jars. The concussion produced 
 by the sudden arrest of a fall of a mass of 
 iron such as was in the drill-rods when 
 transmitted to the latter caused such diffi- 
 
 FIG. 275. An American Drill Rig. 
 
 culties as were sough to be removed by various appliances. 
 Oennhausen's chisel had at the top a four-armed projection 
 which played freely in corresponding slots of a cylinder attached 
 at one end of the rope. As the rope was lowered with its drill 
 it was turned slightly and the ledges of the cylinder slipped
 
 DRILLING- AND BORING-MACHINES. 59<T 
 
 under the cross-arms of the chisel. This was then caught on the 
 up- stroke of the rope and its cylinder. When^it was raised to the 
 proper height a slight jar or jerk released the tool, whose cross- 
 arms slid in the slots and permitted it to drop a distance equal! 
 to the length of the slots. The invention of M. Hind and 1 . 
 Chandron consisted of a pair of bent levers acting like pincers- 
 which grasped the tool and raised it to the proper height, after 
 which it was allowed a free fall to do its work without recoil 
 upon the apparatus. 
 
 The jar universally used is illustrated in Fig. 276. It is a. 
 pair of open links allowing of i to 3 feet of play. One link is- 
 attached to the upper rod length and its rope, the lower to the- 
 drill-bit. These aggregate 1200 pounds in weight for an 8" hole 
 and a total drilling string of 60 feet. 
 
 The Drill-tools. In Fig. 276 is illustrated a full string of 
 tools. Beginning at the left they are: the sinker-bar, wrench- 
 bar, wrench, jars, temper-screw, drill-bit, spudding-bit and rope- 
 socket, and auger-stem. Above is the floor-circle. 
 
 The temper-screw is suspended from the walking-beam and 
 regulates the feed of the drill and the rope. The swivel below 
 is turned a half-revolution after each blow, paying out the rope 
 and rotating the drill-bit a small amount. When the full speed 
 of 4 feet has been paid out the screw is undamped, while the 
 drill- rope is otherwise supported, and its feed raised for the. 
 next stage or drilling. 
 
 The stem, drill-rods, jars, and bit are about ij" in diam- 
 eter, weighing, for a 60- foot }sring, 1200 pounds. Each mem- 
 ber has a screw-pin at its upper end, fitting accurately into- 
 the box at the lower end, leaving the outside of the flush. A. 
 pair of wrenches is employed for tightening, one below and 
 the other above the joint. The upper one is further aided by 
 the wrench-bar inserted into a hole in the floor-circle bolted to* 
 the frame. 
 
 Drilling the Hole. The hole is started by the spudding-bit 
 and stem, which are attached to the rope passing over the sheave 
 and once or twice around the bull-wheel. The engine in opera-
 
 boo 
 
 MANUAL OF MINING. 
 
 FIG. 276. A String of DriU-tools.
 
 DRILLING- AND BORING-MACHINES. 6ox 
 
 tion revolves the bull-wheel continuously. The free end of the 
 rope, being held by a man, the bit and rope are raised a slight 
 distance, after which the rope is released, slipping on the drum, 
 and allows the tool to fall. This is repeated till the surface-soil 
 has been penetrated. The verticality of the hole is maintained 
 meanwhile by an upright plank box leader. Formerly the 
 tool was manipulated by a spring-pole with a foot-stirrup for av 
 moderate depth and was known as "kicking down a hole." 
 
 When bed rock has been reached the spudding- bit is replaced 
 by the drill-bit, stem and jars, and their rope clamped to the 
 temper-screw which communicates the reciprocating motion from 
 the walking-beam. The tool is raised from 2 to 6 feet and allowed 
 to fall at a rate varying with circumstances, the average being 
 six times a minute. After each blow the swivel is turned, for a 
 uniform rotation of the drill is essential and a constant watch- 
 fulness necessary, otherwise the drill may work out of line.. This- 
 is particularly essential if the strata are inclined or the material 
 not homogeneous. 
 
 Frequently, to maintain verticality and sometimes to straighten 
 a curved hole, a winged tool is brought into play. The drill-rod 
 is provided with projections extending on the four sides of the 
 chisel for the same distance along it and nearly fitting the hole, 
 and sometimes a hollow cylindrical reamer is resorted to for the 
 purpose. 
 
 In the Mather and Platt system the partial revolution of the 
 tool and its 'flat rope was obtained after each blow by a movable 
 collar rotating with the tool. It was cut with inclined teeth at 
 both ends and was capable of a motion of 2" or 3" vertically inside 
 of an iron bow. Above and below the collar were sets of inclined 
 teeth into which the collar ends geared. With each drop the 
 collar engaged the teeth of the lower set, turning it with its chisel 
 one half a tooth. At the end of the upper stroke the collar strikes 
 the upper set and is again turned one-half of a tooth. 
 
 When the temper-screw has reached its limit it is returned to 
 the starting-point, fresh tools meanwhile being attached; or the 
 debris is pumped out preparatory to the next run. If no under-
 
 <602 MANUAL OF MINING. 
 
 ground stream has been encountered to supply water for cooling 
 ithe bits and facilitating the removal of the debris, it must be fur- 
 .nished at the surface. 
 
 The Sand-pump. This consists of a slender cylinder 1 with a 
 valve at its foot which is attached to a rope having its own pulley 
 at the derrick head, and its own wheel at the engine. It is lowered 
 into the hole after each run and pumped up and down. The 
 sludge produced by the drill and the water is drawn into the 
 pump and raised to the surface. 
 
 Recovering Lost Tools. If the sand-pump or reamer breaks 
 loose, it can be fished out by some form of grapnel, or it is chopped 
 up. If pebbles fall into the hole above the tool and wedge it 
 fast, it may be jarred loose or freed by a spear. Occasionally 
 the hole is reamed out to a larger size. 
 
 The Progress. The rate of progress varies from 10 feet a 
 day in magncsian limestone to 70 feet in soft ground. Much 
 depends on the skill of the driller, who may determine by the 
 "feel" of the concussion transmitted through the rope whether 
 the blow has been effected or is cushioned. The feed is regu- 
 lated accordingly. Two men are sufficient to operate a drill 
 rig, and 200 pounds fuel are consumed per day. 
 
 The cost of drilling wells is 50 and 60 cents per foot, though 
 some very deep holes average $1.14. Reaming is done at a cost 
 of about 40 cents; and the royalty for the rent of a machine is 
 nearly 10 cents. 
 
 Hand-boring of Deep Holes. A unique mode of drilling 
 artesian wells is in vogue in San Luis Valley, Col., where 300 feet 
 of 5" hole can be made in twenty-four hours. The soil of that 
 .valley is very porous, and for 900 feet down will hardly hold 
 water except in the clay seams. Ditches have failed of their 
 irrigating purposes, and each ranch is provided with one or more 
 .of these spouting wells. From a tripod a 5" tube is held ver- 
 tically and by a pair of blacksmith's tongs is turned by hand. 
 At the foot of the pipe a bit projects outward inch or so. By 
 pressure and rotation a hole is cut spirally into the gravel, lengths 
 i>eing added with the progress. A barrel or two of water is
 
 DRILLING- AND BORING-MACHINES. 603 
 
 poured down the pipe to wash the detritus out through the 
 annular space until a ground-current is encountered which will 
 supply the necessary water. Beyond 300 feet horse- or engine- 
 power will be required to turn the borer. 
 
 Casing a Bore-hole. When a loose or fragile stratum is 
 encountered its caving interferes with the progress of the drill. 
 A pipe-tubing is forced down to support the sides of the bore. 
 It may be driven while the drilling is in progress below or the 
 operations of drilling and casing may alternate. Though wood 
 is the best tubing material, the objections to it are so apparent 
 that wrought-iron pipe is used. The pipes may be screwed 
 together, joined by wrought-iron couplings or telescoped. The 
 first length, having been sharpened, is driven by percussion until 
 its full length is in the guide tube and the successive lengths 
 are screwed on and hammered. When the depth has so far 
 increased that the pipe can no longer be rammed, the bore-hole 
 is enlarged for the same size of pipe to be lowered, or the same 
 diameter of bore is maintained for a smaller pipe. The pipe is 
 driven by repeated blows from a drop-weight or by the steady 
 pressure from jack-screws, a block of wood at the top of the pipe- 
 line receiving the impact. 
 
 A deep bore-hole usually encounters water-currents at differ- 
 ent depths before the object for which the bore has been drilled 
 is attained. These are isolated by casing. A salt-water current 
 may be met before the artesian flow is reached or either of these 
 may be intercepted before the oil-sands are struck. A pipe of 
 large diameter may be carried to the first flow and sealed at its 
 foot. A second smaller pipe, lowered to the next flow and sealed, 
 would afford an escape for the upper current in the annular 
 space and a discharge for the lower one inside the -smaller pipe. 
 Three separate artesian flows are thus obtained by casing each 
 current and discharging them at different surface elevations. An 
 undesirable current may be dammed back in like manner. This 
 may be necessary when drilling for oil, as the accumulation of 
 the flowing water reduces the efficacy of the blow and also exerts 
 such pressure, at the rate of 14.7 Ibs. per square inch for every
 
 604 MANUAL OF MINING. 
 
 34 feet in depth, that when the oil stratum is reached neither 
 oil nor gas can escape. 
 
 The casing is sealed by hydraulic cement or by the use of a 
 seed-bag. A stout leather tube several feet long is attached 
 outside of the piping at the lower end. The annular space is 
 filled with flaxseed bran or other absorbent material and the 
 top lashed. When lowered into place the seeds swell and fur- 
 nish a tight joint somewhat like that of the moss-box (Fig. 203). 
 
 FIG. 277. The Cyclone Prospecting-machine. 
 
 Shooting the Well. It frequently occurs that the oil does 
 not flow freely, or even at all, when the sands are encountered. A 
 practice prevails of opening communication with an underground 
 oil-chamber by exploding a torpedo at the bottom of the well. 
 This is a charge of nitroglycerine carefully lowered into position, 
 after which a heavy weight is dropped upon if. This heroic 
 treatment is not resorted to in shaly ground, unless provision 
 is made for immediately casing the shattered portion of the 
 hole, as it often results disastrously. The charge varies from 
 10 Ibs. to 100 Ibs. in desperate cases. 
 
 Removing the Casing. When the object for which the hole 
 was drilled or lined has been attained the tubing may be recovered 
 by the use of an ovoid screw-plug of oak, which is attached 
 to the tool rod and lowered to the bottom. A little sand is 
 poured in and wedges the plug, which can be hoisted with- 
 out trouble. To loosen its hold on the tube it is only neces- 
 sary to drop it slightly and let the sand run out. A three-
 
 DRILLING- AND BORING-MACHINES. 
 
 005 
 
 FIG 
 
 FIG. 279. 
 
 FIG. 278. The Keystone Sand-pump. 
 FIG. 279. Casing a Well in the Keystone Process. 
 FIG. 280. Keystone Pipe-puller. 
 
 FIG. 280.
 
 606 MANUAL OF MINING. 
 
 pronged expansible hook lowered down under the tubing is also 
 called into service. 
 
 The Keystone Prospecting-machine, shown in Fig. 279, is 
 a type of percussion drill, giving service in loose placer-ground 
 and the zinc-lead fields of Missouri. The string of tools is iden- 
 tical in character with those of the oil-well rig. The hole is 
 started with a fluted spudding-bit, stem, and rope-socket and 
 alternated with the sand-pump (Fig. 278). When 50 feet has 
 been drilled the jars are inserted in the line and 
 the hole continued. 
 
 The drive-pipe casing is of merchant pipe, 
 with straight threads and squared ends, in 
 lengths of 6 feet, coupled together. At the foot 
 is a heavy drive-shoe with a cutting-edge and at 
 the top a cap on which the driving blows are 
 delivered. The casing follows the drilling closely 
 as shown in Fig. 279. 
 
 The drive-pipe is withdrawn by the pipe- 
 pullers (Figs. 280 and 281). 
 The Cyclone Drilling-machine resembles the other portable 
 drill-rigs, though the detail of operation is somewhat different. 
 The drill-bit is not solid but is hollow-stemmed. Holes in the 
 side of the bit afford communication from inside to outside. 
 Hollow rods support the drill-bit instead of a rope and at their 
 junction is a ball valve. When the line of tools falls, the valve 
 opens and permits the drillings to enter the drill-stem and rise 
 in the hollow rods. On the up stroke it closes. With each down 
 stroke additional material enters, to be finally discharged at the 
 surface into a proper receptacle. Extra-heavy rod lengths are 
 employed near the bottom and lighter rods above. For holes 
 exceeding 4 inches in diameter the rods are of 2-inch and i^- 
 inch pipe (Fig. 277). 
 
 Rotary Drills. M. Leschot has the credit for the first appli- 
 cation of rotary diamond-drills to the miner's art, since which 
 time it has gained in favor and increased in range of utility. 
 The lower end of a long line of tubing is supplied with a cutting-
 
 DRILLING- AND BORING-MACHINES. 607 
 
 surface and is caused to rotate by appropriate machinery at the 
 surface. A continuous pressure is produced at the cutting-edge 
 by the weight of tool above. The cutting-medium may be chilled- 
 steel fragments, diamonds inserted into the rim of the boring- 
 tool or a steel tube may be dressed to a hard cutting-edge for 
 the purpose, each having its special adaptability to rock of a 
 definite character. Kl 
 
 The Diamond-bit. A cutting-tool is constructed of diamonds 
 which are forced by hydraulic pressure into sockets on the end of 
 a steel tube. Recesses are accurately prepared into which they 
 are secured by metal hammered up around them. In some 
 cases a firm setting is obtained by forcing the stones forward 
 through small holes in the metal by means of a screw or by hydraulic 
 pressure. A later method consists in forcing the stones nearly 
 through the metal and subsequently grinding the steel down 
 until the stones are exposed. 
 
 The tube may be closed by a concave surface (Fig. 284) or 
 by a convex face (Fig. 283) and the entire area studded in such 
 manner that no concentric circle can be drawn failing to touch 
 one or more of the diamonds. Some also project beyond the 
 
 FIG. 282. FIG. 283. FIG. 284. 
 
 The Annular, Convex, and Concave Diamond-drill Bits. 
 
 tube. The annular end of a tube may be charged with diamonds, 
 as in Fig. 284. For drilling a small hole the first-named types of 
 bit are used, the concave being preferred to the convex, but 
 for large holes or where a central core is sought and withdrawn 
 for inspection the annular bit is used. 
 
 The diamonds are of two varieties, the black "carbonadors" 
 and the deep-red "borts." The former lack the crystallization
 
 608 
 
 MANUAL OF MINING. 
 
 making them durable, while the borts are rather brittle. Theo- 
 retically too many carbons cannot be put in; "there should 
 be never less than 12, and as many as 20 may be mounted 
 on a bit." 
 
 The Diamond-drill. The bit is coupled to a line of tubing 
 and guided by a short or long length of tube 
 (Figs. 285 and 286). The tubes are in 8-foot 
 lengths and of an outer diameter of rarely over 
 8 inches. Holes of great depths are thus drilled 
 of a uniform diameter to the bottom or in sections 
 of diminishing diameters. 
 
 The engine power varies with the length and 
 size of drill-tube to be manipulated. For 1000 
 * eet f k re an 8-horse-power engine will suffice. 
 This with its running gear is framed and suitably 
 braced to carry the full line of rods. A derrick is often erected 
 
 FlG. 286. The Core-barrel. 
 
 to facilitate the addition or disjointing of tubes. 
 
 Regulating the Feed of the Drill. The changes in the struc- 
 ture and toughness of the rock during drilling are so rapid and 
 pronounced that the drill cannot be fed uniformly. A variable 
 feed is necessary commensurate with the progress of the drill 
 and two methods of avoiding a positive feed are in vogue. One 
 is a spur-wheel feed, the other an hydraulic. The former is 
 so adjusted by differential gear that its friction shall equal a 
 desired resistance, and when this is exceeded because of undue 
 strain below, a regulation is obtained. This is inferior to the 
 hydraulic feed. In Fig. 288 is shown a simple motor which by 
 means of hydraulic pressure on the piston produces a pressure 
 which is maintained constant. Both ends of the cylinder are 
 connected with the pump and the suitable cocks admit of a per-
 
 DRILLING- AND BORING-MACHINES. 
 
 609 
 
 feet control by the operator, v/ho gives any variation or reversal 
 of speed within the limit of the pump and the piston area. 
 When the pump pressure increases, the piston B and its rods 
 
 FIG. 287. A Prospecting Diamond -drill 
 
 are raised. A decrease of "pressure results in its lowering. 
 The rate of feed is fixed by the hardness of the rock. In Fig. 
 287 the feed cylinder is shown in position above the drill-tubes. 
 The pressure exerted by the feed is just sufficient to produce 
 abrasion, but not to cut the rock.
 
 6io 
 
 MANUAL OF MINING. 
 
 FIG. 288. The Hydraulic Feed. 
 
 A chuck intervenes between the 
 tube and the feed, which may be 
 loosened at the end of the feed-travel 
 and run up to the top for a fresh 
 start. The tube line is partially sus- 
 pended by friction-rollers at the sur- 
 face, so that it is subjected to very 
 little tension. But a high degree of 
 torsion is developed by the rotary 
 effort against the abrasive resistance, 
 and the torsional strength of the tubes 
 places a limit on the possible depth of 
 drilling. The regulating power of the 
 feed limits the capacity of the ma- 
 chine. 
 
 Removing the Cuttings and the 
 Core. The tube is revolved at a rate 
 of from 400 to 800 turns per minute 
 and continues without interruption 
 from five to eight hours, cutting mean- 
 while from 13" to 2' each hour. A 
 stream of water is fed inside of the 
 pipes and washes the rock face and 
 the annular bit. It carries the cuttings 
 up through the annular space outside. 
 The grooves on spiral the outside of 
 the pipe (Fig. 286) facilitate the escape 
 of the water with the solid bits. The 
 water is fed outside the pipes and 
 escapes with the debris through holes 
 into the interior. A guide-bit (Figs. 
 285 and 286) and core-barrel main- 
 tain the bit in the direction in 
 which it is started. The water-supply 
 is forced by means of a pump. 
 
 The progress of the drill is carefully
 
 DRILLING- AND BORING-MACHINES. 6ll 
 
 watched and the character of the cuttings supplied by the over- 
 flowing water-current is examined. Samples are taken and pre- 
 served for reference. With a solid bit is produced only a fine 
 sludge for inspection, but the annular bit furnishes in addition, 
 a core. 
 
 With every 10 to 15 feet of advance the tubes are raised 
 for examination and for the purpose of extract- 
 ing the core. In the inside of the drill-tube is 
 a lifter (Fig. 289). This is a spring ring which 
 grasps the core only when it is raised It wrenches F IG . 2 8g. The 
 the latter from its place, and will retain hold Core-lifter. 
 on it until released. This can be done by lowering the 
 tube. 
 
 The Core. The core furnishes a guide also as to the direction 
 of inclination of the strata if care be taken that the tubes are not 
 turned while being lifted. Flat ropes are therefore used for 
 hoisting as less likely to untwist and turn the drills. False informa- 
 tion may be given by the core if a soft layer has been penetrated 
 between two hard strata, for the former may wear away or even 
 grind off loose in the core. Hence the core may not be an absolute 
 guide. Many causes combine to make this examination unre- 
 liable. A careful measurement, an allowance for wear, and 
 frequent inspections are the only means of checking results. 
 The progress in homogeneous rocks is uniform, but a composite 
 rock with constituents of unequal hardness is difficult to pene- 
 trate in perfect alignment. In clay a thick pulp is formed which 
 chokes the tubes and requires the full pressure of the pump to 
 remove it. 
 
 Deflection of the Drill-hole. Changes in the inclination of 
 the strata or variation in hardness of the material tend to deflect 
 the drill-tubes and to buckle the rods under torsion. The correct- 
 ness of the alignment must be maintained, and this may be tested 
 by a phial of gelatine lowered to a certain depth "and allowed to 
 harden. When raised to the surface, it is inclined in the same 
 plane it occupied till the gelatine level is horizontal, whence the 
 dip of the hole is made known. The line etched by hydro-
 
 612 
 
 MANUAL OF MINING. 
 
 fluoric acid on glass is another method of test. Neither of these 
 are entirely satisfactory. 
 
 Accidents are Rare. Occasionally a diamond may fall out, 
 and if it cannot be recovered, is chopped up at once, or the direc- 
 
 FIG. 290. An Underground Diamond Drill. 
 
 tion of the water feed is reversed and its pressure increased to 
 wash the stone up the tube. Hard nodules of rock that retard 
 a fair progress are chopped up by a special bit. A diamond has 
 also been recovered by inserting a lump of wax at the bottom of 
 a tube, which is lowered into the hole.
 
 DRILLING- AND BORING-MACHINES. 
 
 613 
 
 The limiting depth of vertical holes is determined by the 
 torsonial strength of the tubes and the power of the machine. 
 That of inclined holes depends upon the amount of friction 
 between the tube and the rock. 
 
 For underground work a three horse-power electric motor 
 like the "Little Beauty" (Fig. 290) is mounted on a truck, with 
 
 FIG. 291. Exploration by Diamond-drill Holes. 
 
 drum-drill and a pump, and permits core-drilling to advantage 
 in small spaces. In many mines i-inch cores in sections of 5 to 
 20 inches are cut for 80 feet depth, and a great deal of prospecting 
 has been done with this compact machine, which often makes 
 2 feet per hour at a cost of 68 cents to $1.06 per foot. In Fig. 
 291 are exhibited the explorations conducted at the Silver Islet 
 Mine by the use of the underground drill. Doubtless many 
 properties owe their existence to the result of diamond-drill 
 discoveries, and its use has frequently saved expense in various 
 ways. By graphically representing on cross-section paper to
 
 614 MANUAL OF MINING. 
 
 scale the results of underground borings a more intelligent inter- 
 pretation can be made. On the charts are indicated in colors 
 the various rocks penetrated by the several bore-holes. A com- 
 parison of position and thickness enables the engineer to judge 
 of the prospects. Still, the diamond drill is not considered 
 infallible in its indications as to the presence or absence of the 
 ore body sought. 
 
 The Chapman hydraulic pipe-rotating drill and the Davis 
 calyx drill are types of rotary borers serviceable for loose ground 
 such as that of the Beaumont oil-field of Texas. Both systems 
 have a turntable for directing the drills, and a pump for the water- 
 supply. The Chapman drill depends upon "adamantine" for 
 its abrasive material, and the Davis drill uses chilled shot in hard 
 uniform rock, but the normal bit in soft ground. The drillings 
 pass up the tubes and are caught in a calyx above the Davis 
 core-barrel, or at the surface in the Chapman system. A rate of 
 speed has frequently been attained equal to 900 feet in twenty- 
 five hours. In California an experimental hole was drilled 1500 
 feet in seven full working days. Holes of 1 5 inches diameter have 
 been bored by these systems and to a depth of 3000 feet.
 
 CHAPTER VI. 
 
 MINERS' TOOLS. 
 
 The Texture of Rocks. Rocks to be excavated or broken 
 include those which are hard or soft, tough or brittle, and creviced, 
 stratified or massive ; and the difficulties of their removal present 
 such varied and delicate questions that it is difficult to give a 
 systematic account of the principles of breaking ground or the 
 appliances employed. Materials are gauged by their resistance to 
 abrasion being called "hard " or " soft " as they affect drilling opera- 
 tions; and tough or brittle according to their resistance to concus- 
 sion in blasting. These qualities determine the cost of breaking 
 ground. For example, quartz is hard wherever encountered, 
 but not all of it is tough. Hard minerals wear the cutting tools 
 rapidly, but a brittle mineral is easily shattered by the blow of 
 a hammer or the concussion of a drill. A tough rock possesses 
 the tenacity which resists the rupturing agency of an explosive. 
 It will require powerful blasting agents for its removal. Porphyry 
 is of an average degree of consistency and may be easily drilled 
 and readily broken. Metamorphic rocks are frequently hard, 
 but always tough. They drill easily, but do not break freely. 
 Trap-rock, syenite, and granite are hard on the drills, but require 
 little explosive, the grains having slight cohesion. Fire-clay is 
 tough, though very soft. It has an elasticity which prevents 
 much impression being made on it by an explosive. 
 
 The massive rocks are usually homogeneous and present the 
 same resistance throughout their mass. Some stratified rocks 
 and gravels are not, being composed of hard grains or pebbles 
 cemented by a softer material. These are easily ruptured by 
 powder, but are difficult to drill into. Creviced rocks are not. 
 
 615
 
 616 MANUAL OF MINING. 
 
 economically broken by explosives, but are split by wedges and 
 slow rupturing agents, utilizing the weak lines of cleavage. Bitu- 
 minous coal affords examples of a rock split up by numerous 
 planes into prismatic fragments which would permit the gases 
 from powder explosions to escape. Anthracite being brittle, 
 of short texture and hard, compared with bituminous coal, there- 
 fore requires powder, while bituminous coal is mined by picks. 
 
 The engineer must therefore be familiar with the texture 
 and behavior of the rock before any estimate of the cost of its 
 extraction can be made. In Drinker's "Tunnelling" and Fos- 
 ter's Gallon's "Lectures on Mining" are given elaborate formulae 
 ior estimating the cost of extraction, and they may serve as approxi- 
 mate guides. In each camp a close observation of the results 
 of experiments will supply the practical coefficients which can 
 foe used in the above-mentioned formulae for determining the 
 cost of working. 
 
 Fire-setting System. This is a method applied by the miners 
 of the middle ages to conquer hard rock by the instrumentality 
 of fire. The face of the rock was exposed to the heat of fire and 
 then suddenly cooled by water. The unequal contraction, upon 
 cooling, softened the rock, which then was amenable to the ordi- 
 nary tools. A grate was built inclined toward the face of the 
 rock, and in it were piled billets of wood. When ignited, the 
 flame was directed against the breast by a shield overhead. In 
 some cases a basket was suspended from the roof with the same 
 Tccult. Usually in the mines of the early times the wood was 
 fired on Saturday and allowed to burn out. The men could 
 attack the calcined surface upon their return to work. This 
 method may profitably be applied upon hard rock where fuel 
 Is cheap and the ventilation is good. 
 
 Miners' Tools. The operations to be performed by the miner 
 consist in breaking down the mineral or rock, removing the 
 1:roken material, and illuminating the place of work. These 
 require groups of tools the nature of which varies with the char- 
 acter of the mine in which the work is performed. The mineral 
 may be broken down by picks, drills, striking-hammers, sledges,
 
 MINERS' TOOLS. 617 
 
 augers, gads, moils, wedges, and feathers. Auxiliary tools are 
 also employed, such as the swab, spoon, gun, etc., when the mate- 
 rial is tough enough to require blasting. In the latter case, in 
 addition to the striking- hammers and drills, will also be required 
 the tools for making powder cartridges, cans for carrying the 
 same, and needles, squibs, fuses, etc., to ignite the powder. The 
 mineral is removed by the aid of shovels, picks, and spades; 
 occasionally sledges and crowbars are required for loosening or 
 breaking fragments too large for convenient handling. The 
 illumination of the place of work will require a candle and its 
 stick, or oil-lamp and an oil supply. 
 
 All tools loaned or handed to the miners are entered on an 
 account against them, as otherwise many would be lost in the waste, 
 or otherwise disappear. A check is made of the condition of all 
 tools returned by the men. Steel should be measured or weighed 
 out on delivery at the beginning of the day, month, or contract. 
 
 Rotary Drills. The second class of drills cut away the rock 
 by the rotary motion, and according to the character of the cut- 
 ting bit they are breast-augers, roof-augers, and machine-power 
 augers, worked by compressed air or electricity. 
 
 Hand Boring-machines consist of a drill-piece of twisted steel r 
 W 7 elded to an iron stem and fixed to the end of a long screw pass- 
 ing through a nut which is supported on an upright bar by a ball- 
 and-socket or universal joint, permitting any direction of bore. 
 The rotary motion is ccmmunicated to the drill-screw by two 
 bevel-gear wheels, or directly by a revolving handle at the rear 
 of the screw. The bit may be a wedge-shaped bit with a V edge, 
 or it may be twisted into the common auger form. 
 
 The form illustrated in Fig. 292 is a special bit used by bitu- 
 minous miners. The shank to which the twisted steel bit is 
 welded is bent to form a handle and terminates in a breastplate 
 (Fig. 293). The boring of the holes is accomplished by hand, 
 the pressure being applied by the weight of the body against the 
 plate. The bits of these augers are from ij to i\ inches in diam- 
 eter. Sectional augers are more convenient for delivery to the 
 blacksmith, therefore they are made removable from the shank.
 
 6i8 
 
 MANUAL OF MINING. 
 
 Of the hand- and roof-power augers there are several styles, 
 the two general classes being the grip-drill and the post-drill. 
 The grip-drill (Fig. 294) has a metal rest a for the feed-screw 
 nut &, in which the feed-screw c turns while the hole is being 
 bored. A small recess in the coal is provided into which the 
 grip-bar d is driven and wedged and thus gives stability to the 
 machine. The boring can be done from the side or end. 
 
 FIG. 292. A Bituminous-coal Bit. 
 
 osoooc=i/ \^ 
 
 L 
 
 FIG. 294. A Grip-drill. 
 
 The post- drill consists of a suitable borer attached to the 
 post, which is usually placed upright, having a jack-screw at one 
 end and firmly wedged between the floor and the roof. The post 
 is split or may be obtained solid. The drill is attached to the 
 feed-screw which advances through the feed end by a handle 
 and crank operating the gearing shown. The universal joint sup- 
 porting the drill-gear permits holes to be drilled in any direction.
 
 MINERS' TOOLS. 
 
 619 
 
 Portable Power Borers. The Jeffrey rotary electric coal- 
 drill (Fig. 295) has a small enclosed motor on the end of whose 
 
 FlG. 295. The Jeffrey Rotary Electric Coal-drilL 
 
 armature-shaft is a small pinion-gearing with a larger wheel. 
 Two drills are used for each hole, one being 3 feet and the other 
 6 feet. The time required to drill a 6-foot hole in ordinary 
 bituminous coal, with either type, is from twelve to fourteen 
 minutes. In addition to gearing for hand-power drills there 
 are ratchet attachments for driving them in 
 such places where the holes must be drilled too 
 close to reach the roof or the floor for the crank- 
 motor to be used. It requires from | to i horse- 
 power to drive them. The power is communi- 
 cated through gears and a small pinion by which 
 rotation is produced. A feather on the larger 
 gear-wheel slides along a longitudinal slot in the 
 feed-screw and permits of an advance of the bit 
 with the progress of the boring. These force the 
 drill forward at any determined rate without un- 
 duly crowding the bit. 
 
 A similar pattern of drill (Fig. 296) is 
 operated by a small air-engine or by hand. In 
 a hand-borer this drill is serviceable where electric or air coal-
 
 620 MANUAL OF MINING. 
 
 
 
 cutters are also employed. Unless, however, the power-plants are 
 to be installed for other purposes, it is hardly necessary to install 
 one safely for the use of these drills, so long as equally serviceable 
 hand-power drills can be obtained. 
 
 In loose material and shattered rock, shovels and spades are 
 the only tools needed; clays and soft rocks require picks, crow- 
 bars, and shovels; ground that is scaly, brittle, and seamy is 
 split by wedges in different ways. Massive rock cannot be 
 broken without the hammer and drill. Powder and explosives 
 arc used in the latter case, and occasionally in the others, as 
 auxiliary aid. For economy and speed, there are machines which 
 are driven by steam, electric, or pneumatic power, imitating the 
 operations of the above hand tools. 
 
 Underholing Bituminous Coal. The operation called under- 
 cutting, underholing, bearing in, or kirving, consists in cutting a 
 groove underneath a mass of coal in the soft floor, or in the coal 
 itself, to a depth about equal to the thickness of the bed. The 
 groove is 9 inches high at the face and about 2 inches at the back. 
 The coal is propped by sprags while the miner is engaged in 
 digging a deep groove. The coal then freed below may be 
 broken off by the pressure of the roof overnight, or its fall may be 
 furthered by a few drill-holes loaded with powder. The block 
 may also be sheared by grooves along the sides of the rooms or 
 stalls, cut vertically from roof to floor. Sometimes all three 
 methods are practised to break the coal down. Shearing is 
 more expensive than undercutting, hence is used only for shallow, 
 firm coal or under a very strong roof. The cleats in the coal 
 materially assist in its breaking. Anthracite is too firm to be 
 treated in this way; it must be blasted off the solid. 
 
 Shooting Off the Solid. This term is employed to designate the 
 character of blasting, which is in practice in the anthracite-coal 
 mines wherever there is but one face exposed toward which a 
 blasting agent may operate. In anthracite mines this practice 
 is universal, the miner exploding the holes so that each succeeding 
 one will have a free face for blasting. In flat beds he makes 
 centre cuts (Chapter VII) on the face by blasting wedged-shaped
 
 MINERS' TOOLS. 621 
 
 masses of coal. In pitching-beds he blasts in the coal near the 
 floor and undermines the bed. By systematic work a keg of 
 25 Ibs. of powder can thus blow down from 30 to 50 cars of coal. 
 This practice is warranted in bituminous mines when the coal 
 is to be delivered to the coke-ovens, but not when it is tc/ be 
 marketed for domestic use. 
 
 Shovels. The ordinary shovels are made of iron plate rolled 
 under a welding heat with an edge of steel, and ears drawn out 
 for the handle. A concavity to the blade imparts stiffness and 
 carrying capacity to the tool. Such a scoop-shaped shovel is 
 preferred by the anthracite miner. The end of the shovel is 
 square or pointed, according to its uses. The pointed blade 
 with the long handle is the most used. That of the "diamond 
 label" is the most popular; the width of blade in inches desig- 
 nates the commercial number or size of the shovel. 
 
 Picks. The pick is variously known as a pike, mandril, 
 slitter, hack, and mattock, according to its shape and length. 
 It is of iron with steel tips or all steel. It has an opening or an 
 eye in the centre, the ends being pointed to a pyramidal shape 
 or a chiselled edge. The points or tips are hardened to suit the 
 nature of the work to be done. They are the wearing parts and 
 should be replacable. On this account the all-steel pick is not 
 serviceable. Each sharpening after use consumes an amount of 
 metal, which, in time, reduces it to an inconvenient length. An 
 iron pick 14 inches long with two steel tips wedged into the ends 
 gives longer service, for the points when worn out may be replaced. 
 Removable points are sometimes supplied with picks. "Pick 
 steel" is a special steel in bars i"Xf" or i"Xf" which are 
 cut into suitable lengths. 
 
 The picks are generally made at the mine by welding into the 
 iron lengths of steel wedged and pointed. When finished they 
 are about 22 inches long, though some are 29 inches long, and 
 in the hands of the "box" cutters are doing remarkable "jadding" 
 (cutting the top). The weight of picks varies between 2 and 
 9 Ibs., with 3 1 or 4 Ibs. as an average, the heavier weights being 
 used for downward cutting.
 
 622 MANUAL OF MINING. 
 
 The eye of the pick is oval. It is formed by gashing the red- 
 hot bar in the middle, upsetting it, cutting it open by a drift, and 
 hammering out stout cheeks with abundant metal at the sides 
 and a bearing surface for the handle as long as possible. All 
 of the strain of prying by the points and handle falls upon the eye, 
 which therefore must be stout. 
 
 The shape and weight of the pick depends on the service it 
 is to give. There is the straight or the curved pick, the anchor- 
 or the poll-pick, each having favor for certain work. Indeed, 
 in the same mine several forms are to be seen, perhaps as a matter 
 of individual prejudice. The straight pick assists the reach, the 
 curved pick enables a more effective blow to be struck. So, 
 for overhand work, for underholing, and for getting into corners, 
 a straight-head pick is used, but does not strike as well in down- 
 ward work as the curved. The curve of the latter should be an 
 arc of a radius equal to the combined lengths of the arm and 
 handle. 
 
 Usually the bituminous-coal miner has a light pick of 2 or 
 3 Ibs. weight for the clay and a heavier one for cutting bony 
 coal, slate, etc. The light-weight picks have one square chiselled 
 edge and a pointed one. The anthracite picks are heavier. The 
 poll-pick used for prying loose coal from the face or rock from 
 the roof is as heavy as can be conveniently handled. 
 
 The pick is single- or double-pointed. In the first variety the 
 other end is forged into a hammer-head, known as a "poll," and is 
 used for breaking fragments or driving a wedge. The tapered or 
 chiselled end of the pike is used to pry into and slit the rock with 
 the leverage obtained from the length of handle. For soft ground 
 the pointed end is long and slender. 
 
 The handle or helve of hickory or ash is a straight-grained, 
 firm, well-trimmed stick. It is trimmed to the shape of the eye 
 and wedged tight by a pair of iron feathers. The handle should 
 be at right angles to the pick. When so fastened, the arcs drawn 
 by the tips, using the handle-ends as centres, will have equal radii. 
 The length of American handles is about 28 inches; those of the 
 English are 30 to 35 inches.
 
 MINERS' TOOLS. 623 
 
 Breaking Dimension Stone. Quarrying at the surface involves 
 the procural of blocks of stone of definite shape and size suitable 
 for dressing by the stone-mason to desired dimensions. The 
 process must be such as to ensure with some degree of accuracy 
 the shape desired without injury, and the rupturing agents must 
 be less violent than powder. Wedges were therefore used, though 
 they are rapidly becoming obsolete. The primitive wedges were 
 of dry or wet wood. Now they are of steel or iron. They were 
 forced or hammered into the planes of cleavage of the rock, splitting 
 the latter along the given directions. 
 
 In Fig. 306 is illustrated a block being split by a number of 
 plugs. The wedges are driven by a sledge in succession, begin- 
 ning at either end and working toward the centre, each one receiv- 
 ing a few taps in turn. The holes are put alternately long and 
 short, and of a depth 12 to 20 inches, according to the thickness 
 of the stone. 
 
 The rock may also be broken, by forming a rift along the 
 face, by a sledge. The stone is then pounded along a given line 
 rapidly, until the continual concussion overcomes the tenacity of 
 the stone, and starts a crevice along the line into which a wedge 
 may be inserted and driven. 
 
 The Plug and Feather are more efficient than the simple 
 wedge. The plug is a flattened wedge and the feathers have two 
 wedges each, with one face flat and the other curved. They are 
 introduced into a drilled hole with their points up. The plug 
 is hammered between them and the design is such that the feathers 
 are forced apart at the bottom of the hole with greater power 
 than at the top. 
 
 Hydraulic Wedges. Hydraulic wedges have been used with 
 good success at the collieries near Saarbruecken, the position of 
 the driving- wedge being reversed from that of the plug, the thicker 
 end being placed in the bottom of the hole and the edge near the 
 top. Fitting between the two half-round wedged cheeks which 
 point down, it is driven from below upward by the hydraulic 
 pressure, or the force from the explosion of powder. This snaps 
 the rock in the plane of the thin edge. Dimension store is
 
 624 
 
 MANUAL OF MIX IXC. 
 
 obtained in some quarries by this process, but the risk of huffing 
 off the rock at the top of the hole is great. 
 
 Gads or Moils are very useful accompaniments to every 
 miner's equipment, differing from the early wedges only in that 
 they have pointed tips, not chiselled edges. A piece of steel that 
 has done service for drilling, has been dressed and smithed until 
 it is less than 10 inches long, and can no longer be used for 
 starting a hole, is converted into moil by tapering off the bit- 
 point. It is used for chipping or trimming the rock, for a smooth 
 bearing to timbers, etc. It is used in the Lake Superior copper- 
 mines for "blockholing" or splitting the large masses of copper. 
 It is almost indispensable in hard rock, an'd a number of them, 
 sharpened and hardened, should always be on hand for the use of 
 timbermen and shaftsmen. 
 
 A crowbar is a tool of occasional service underground. It is 
 a long, unhardened, steel bar with a point at one end, and serves 
 as a pry. 
 
 Jumpers and Drill-rods. Percussion on the rock is accom- 
 plished by a jumper for vertical holes only or by a drill for holes 
 
 (b) 
 
 FIG. 297. Forms of Bits. 
 
 of any angle. These are heavy steel bars provided with cutting 
 vH<*es, which may be straight, concave, or convex, and acute 
 (slim) or "bluff" (Fig. 297), the shape being somewhat a matter 
 of individual preference and skill. The bit is shaped by
 
 MIXERS' TOOLS. 625 
 
 the blacksmith, who hammers the red-hot end of a steel drill- 
 rod to a wedge-shape, flaring it out to a width somewhat 
 greater than that of the rod. The edge is worked to a straight 
 or curved form, and then hardened by reheating the end to 
 the proper degree and plunging it into a cooling fluid. The 
 convex bit is stronger than the straight bit, and transmits the 
 blow better with less risk of fracture at the corners. The 
 cutting edge of |-inch drills is i| inches wide, and other sizes 
 are correspondingly flared. The other end of the drill-rod is 
 not dressed after being cut off. It receives the blow of the 
 striking hammer. When both ends of the drill-rod are worked 
 with cutting edges the tool is known as a jumper. 
 
 The Jumper is a heavy 5- or 6- foot round iron bar, heavier 
 near the middle and hammered to an edge at each end. A hole 
 is churned down into the rock by lifting the jumper and allowing 
 it to fall, after which it is turned, slightly raised, and dropped a 
 foot or so. When one bit is dulled the jumper is reversed and 
 the other end continues the operation. In this way a hole is 
 "jumped down." The debris is cleaned out at intervals by a 
 scraper or a spoon. Coal and limestone are churned in this way 
 at a rate of 40 or 50 feet of hole in a shift; granite, at 15 feet. 
 The jumper, with one cutter and a head for hammering, has lost 
 the significance of its name and is properly a drill. 
 
 The Drill. The steel rods for hand- or power-drills are pro- 
 cured in lengths of 14 feet and of any diameter. Though there 
 are many grades, the English Jessop and the American brand 
 "Black Diamond" have pronounced adherents. The steel is 
 round or octagonal, the diameters usually being between f and 
 2 inches. The length of the bar depends upon the depth 
 of the hole to be drilled. The rods of less than i inch are used 
 only for short holes and light hammering, as when a miner 
 both holds the drill and strikes it with a hammer. In "double- 
 hand work," where a companion strikes, the diameter of the 
 drill is between i and i| inches. For "three-hand" or "double- 
 hammer" work the steel may be as large as 2 inches, though 
 it is doubtful if a hand-drilled hole greater than i| inches
 
 626 MANUAL OF MINIXG. 
 
 diameter is justifiable. Generally speaking it is preferable to 
 have the drill as large as possible to transmit the blow, and also 
 to produce a hole as broad as convenient. The drill of small 
 diameter is rather light and yields under the blow without trans- 
 mitting the effect. 
 
 Drilling Operations. The hole is drilled by the repeated 
 percussive blows upon the drill by the hammer. The rock is 
 chipped by the concussion of the bit; after each stroke the drill 
 is turned slightly and the blow repeated. Though the hole 
 may be drilled without the addition of any water, it is never- 
 theless the practice to occasionally introduce some water into 
 the hole to preserve the temper of the tool and to facilitate the 
 removal of the debris. 
 
 The lengths of drills employed increase with the depth of 
 the hole and its progress, the starting drill being a short, stout 
 drill of comparatively large diameter, and the finishing drill, 30 
 to 36 inches in diameter, being of smaller diameter. The several 
 lengths of drills constitute a set, the bits of a set being flared less 
 with each increasing length of drill. 
 
 The depth of holes varies with the character of the work. It 
 rarely exceeds 36 inches with double-hand work and averages 
 25 inches in single-hand work. In hard, brittle rock the holes 
 are made deep and narrow, while in the tough and fissured mate- 
 rial they are short and broad. Deep holes are more economical, 
 but the miner usually refrains from drilling to the limit, except 
 upon rare occasions, because of the difficulty arising from leaving 
 the ground in poor shape for subsequent blasting, as will be 
 seen in the Chapter on Blasting. The miner places the direc- 
 tion of the holes and carries them to such a depth as will enable 
 him not only to produce a maximum effect, but also to leave a 
 suitable free face for attack by the subsequent shots. As to 
 diameter, it is cheaper to drill a small hole and to increase 
 the strength of powder than to have a hole of large diameter 
 with a large quantity of weak powder. The work of drilling 
 a ij-inch hole is nearly three times that of putting down 
 a ^-inch hole of the same length; the relative volumes of
 
 MINERS' TOOLS. 627 
 
 the rock pulverized are as the square of the diameters; or, 
 in other words, all else being equal, 25 lineal feet of small 
 holes can be put down while 9 feet of the i|-inch hole are 
 being drilled. 
 
 Twelve feet, in the aggregate, constitute a stent with double- 
 hammer work, and 9 feet of holes with the single hammer and 
 one striker. In the case of single-hand work the bits are smaller 
 in diameter than in the double-hand work for a gang of two 
 strikers. About 30 inches of holes can be drilled per shift by 
 single-hand work in rock of a medium character. Jn rare excep- 
 tions the progress may be as great as 5 feet per man. 
 
 The consumption of steel varies with the tool, but an average 
 may be given as about i Ib. for 15 tons of rock blasted. 
 
 Removing the Drillings. During the progress of drilling 
 the debris is removed by a squib, stick, spoon, or gun, accord- 
 ing to the circumstances. The spoon is a round f-inch iron 
 bar, 40 inches long, having a handle at one end and a spoon at 
 the other, which latter is made by curving slightly the lower 5 or 
 6 inches of its length and bending up a spoon at the end. The 
 gun is a syringe, made of a length of gas-pipe and fitted with a 
 suction-piston having a handle on its rod. By these appliances 
 the load may be removed from the hole more effectually than 
 can be done by the spoon. The latter is employed for scraping 
 the sand from a dry hole. The swab-stick is sometimes used 
 to remove the last remnant of moisture in the hole prior to 
 charging it with powder. 
 
 Hammers are employed for breaking large fragments of 
 rock and for striking drills. In weight and finish of their faces 
 they differ. The face is usually flat and very hard. The convex 
 striking-face is rarely used. Striking-hammers are not used for 
 rock-breaking, which duty is left to the sledges. 
 
 For single-hand work the hammer-head weighs 3^ to 5 Ibs.; 
 for double-hand work, 5 to 7 Ibs., except for wedge- driving; it is 
 short with a broad face, with a handle correspondingly short or long. 
 
 Single-hand and Double-hand Work. By single-hand work 
 is understood the method of drilling holes by a single individual,
 
 628 MANUAL OF MINING. 
 
 who holds the drill in one hand and strikes it with a hammer in 
 the other. In double-hand work the operations of holding and 
 turning the drill and of striking the tool are performed by two 
 men. In double-hammer work the gang is composed of three 
 men two striking and one holding. Except in very hard rock, 
 in which satisfactory progress is made by double-hammer work, 
 single-hand work is the usual practice in narrow places and 
 double-hand work where the room is sufficient to allow of swing- 
 ing the sledge for a blow. For many reasons the double-hand 
 work is almost universal. Though many Abjections obtain to 
 its practice, it has an advantage in enabling the men to alter- 
 nate the work of striking and holding and thus relieve one another. 
 But the benefits derived from the single-hand work are so great 
 that where it is possible of introduction it is employed. 
 
 Drinker says that in the point of economy of time and money 
 "one-hand drilling" is from 30 per cent in soft schist to 20 per 
 cent in soft sandstone cheaper than "two-hand drilling." In 
 hard rock "one-hand drilling" gives the more rapid advance. 
 Dr. A. Serlow, in his "Leitfaden zur Bergbaulunde," believes 
 that except in shaftw'ork all other forms of drilling may be exe- 
 cuted more accurately by the " single " than by the " double-hand, " 
 and perhaps more cheaply. 
 
 Itemizing the Drilling Expenses. There are various methods 
 of apportioning the cost of broken ground. One simple method 
 consists in keeping the supplies, such as drill steel, tools, etc., 
 under one heading; the power, fuse, caps, oil, and candles as 
 sundries; the labor by itself; while the fixed charges, such as office 
 and shipment expenditures, are given separate accounts. Occa- 
 sionally the labor item is subdivided and distributed as a guide 
 for future economies. It may be possible thereby also to ascer- 
 tain the number of tons or cars broken per miner, from which 
 may be fixed a standard of settlement. Again, the cost-sheet may 
 be divided into accounts as follows: Under labor are miners, 
 drilling, blasting, trammers, blacksmiths, engineers, timbermen. 
 trackmen, engineering, and superintendence. The minerals- 
 account contains items for powder, caps, and fuse, oils, waste,
 
 MINERS' TOOLS. 629 
 
 candle, coal, stable repairs, track tools, and timber; and the mis- 
 cellaneous account includes office expenses, taxes, insurance, 
 sinking-fund, interest, and shipments. These different items are 
 reduced to cost per ton of product as well as the aggregate cost 
 per week or month. 
 
 The Blacksmith Shop. The duties and work of the blacksmith 
 may not seem relevant to the engineer, but no manager can afford 
 to be ignorant of any element connected \vith the economy of his 
 work. As a matter of fact, it is highly essential that the latter 
 be capable of judging the performance of the blacksmith, who 
 is the butt between the complaints of a miner's inefficiency and 
 that man's retorts in pleading bad tools. A smith who can 
 sharpen tools suitably for hard ground is held in high esteem by 
 miners. The equipment of the shop comprises a full kit of 
 tools, costing perhaps $30.00. a good bellows and tuyeres, Peter's 
 anvil, vise-taps and dies, twist-drills and hoop iron, an assort 
 ment of carriage- and machine-bolts, screws, spikes, nails, a few 
 horseshoeing tools, benches, etc. The building is about 14' Xi2 f 
 with hinged door-openings, near or over the fire and in the two 
 walls, for working long bars. If machine-drills are used in the 
 mine, a set of special swages are necessary for finishing the bits 
 to the X, Z, or + form, as required. 
 
 The miners, who usually must pay for the wasted steel or for 
 the sharpening of bits, are entitled to the services of a competent 
 blacksmith and suitable drill steel. The charge is usually $i 
 per month in coal-mines and 3 cents per bit in metalliferous 
 mines. The sharpening is often done, even for contractors, 
 at the expense of the mine. 
 
 For a small mine employing 20 men in all, i blacksmith 
 will suffice, though, of course, it depends upon what he must 
 do. A good sharpener can dress tools for 20 men on medium 
 rock, or swage the I or X bits for 7 machine-drills. Excepting 
 the pointing of picks, the cutting of steel, and the handling of 
 large pieces, he will need no striker. With this help he can 
 make 12 heavy picks, 20 light ones, or weld 40 pick-stems in a 
 shift ; or he can finish 2 sets of colliers' tools of 5 coal-picks }
 
 630 MANUAL OF MINING. 
 
 2 wedges, a hammer, and 2 bottom-picks. Alone, he can dress 
 40 bits an hour; with help, he can forge 25 double hand-bits, 
 or draw out and temper 50 pick-points per hour. 
 
 Forge Fuel. One important element of success in forging 
 iron is a clean, pure fuel. This may be a slightly caking coal 
 that gives flame and a high heat, or coke which is hotter but 
 more difficult to keep alight. The coal should be clear of shale 
 and slate, for they fuse and make a pasty cinder that adheres to 
 the iron. It must be free of sulphur, which makes the iron "hot 
 short" and also tends to produce scales while forging. For the 
 latter reason white-ash coal is preferred to red-ash coal. 
 
 Welding. A very useful property of wrought iron is its 
 capability of welding, by which two short lengths may be united 
 to form a bar of a serviceable length. The process consists of 
 wedge-tapering an end of each bar, heating them to red, and sub- 
 sequently hammering the softened parts together. A more 
 difficult joint, known as the split, is described further on in the 
 steeling of picks. If the welding has been well done, the point 
 of union is as strong as any other part of the bar. Precaution 
 must be taken to keep the surfaces clean and free from scales, 
 which would interfere with perfect welding contact and are so 
 apt to form in a thin fire of the forge. Scales are due to the 
 oxidation of the iron, which while red-hot is not sufficiently sur- 
 rounded by ignited carbon to consume the free oxygen of the air. 
 When the layer of fuel is thin, or where too much blast is given, 
 the nascent iron absorbs the oxygen. Once formed, the scales 
 cannot be melted or fused off, but their formation may be pre- 
 vented by a liberal covering of fuel over the iron or by sprinkling 
 borax over the surfaces during the heat. Sometimes sand may 
 do instead of borax, though it requires a higher temperature for 
 its fusion. With proper regulation of fire, steel may also be 
 welded to iron with advantage; as, for example, the steel ends 
 of a pick, which receive the wear of the rock, are welded into the 
 iron head of the pick-eye. 
 
 Steel is a compound of carbon with iron in varying propor- 
 tions. H. M. Howe, in his "Metallurgy of Steel," says that steel,
 
 MINERS' TOOLS. 631 
 
 in its specific sense, is "a compound of iron possessing or capable 
 of possessing decided hardness simultaneously with a valuable 
 degree of toughness when hot or when cold, or both. It includes, 
 primarily, compounds of iron combined with from, say, 0.3 to 
 2 per cent carbon, which can be rendered decidedly soft and tough 
 or intensely hard by slow and rapid cooling, respectively; and, 
 secondarily, compounds of iron with chromium, tungsten, man- 
 ganese, titanium, and other elementary compounds, which, like 
 carbon steel, possess intense hardness with decided toughness." 
 "The terms 'iron' and 'steel' are employed so ambiguously 
 and inconsiderately that it is to-day impossible to arrange all 
 varieties under a simple classification." The various adjectives 
 qualifying the term "spring/' "shear," etc., apply to the uses 
 to which the steel is put, and imply a certain percentage of car- 
 bon constituency. 
 
 Hardening Steel. The homogeneity of steel and the presence 
 of carbon imparts to it a capability of hardening and tempering 
 to a degree depending on the temperatures of the heating and 
 the subsequent cooling. As the amount of carbon increases, 
 the melting-point of the iron decreases; and this greater fusi- 
 bility reduces its welding quality. 
 
 A steel is said to be "hardened" if, when red-hot, it is suddenly 
 cooled. The reason for this change is not readily understood, 
 though it is in some degree owing to the presence of the carbon; 
 for pure malleable iron is not in the least affected by the opera- 
 tion, whiile both steel and cast iron are hardened to a marked 
 degree. That attained by cast iron by plunging it into a cooling 
 fluid is not so great as with steel. The greater the difference 
 in temperature between the steel and the cooling fluid and the 
 shorter the time of hardening, the harder it becomes. 
 
 The fluids used for the cooling bath, are water, oil, mercury, 
 and lead, according to the care to be exercised in the process. 
 Water, having the highest specific heat, performs the operation 
 of cooling most quickly. The steel, however, scales somewhat 
 and becomes excessively brittle. The oil is of slower action 
 anu is believed also to supply from its decomposition some carbon,
 
 632 MANUAL OF MIXING. 
 
 which carburizes with the steel. Lead and mercury are used 
 in large establishments. 
 
 Tempering. This process follows hardening. The use to 
 which the tool is to be put determines the degree of hardness 
 desired, but the latter cannot be secured accurately without skill 
 in one operation when water is used. It may be attained by a 
 bath of molten lead for certain uses, but usually the steel is sub- 
 jected to a second operation of annealing. 
 
 The hardened steel is again heated to a red heat and thus is 
 softened slightly. Its point is then plunged for an instant into 
 a cooling fluid to harden the lower end. Withdrawing the article 
 and rubbing off the scales, the heat from above the end will be 
 conducted to the edge. A series of colors plays over the surface 
 in succession as the temperature increases. Beginning \vith a 
 light straw, passing through the shades of yellow, brown, purple, 
 and blue, the effects of the chill are partially removed, till, if the 
 end should become red, all its hardness will have been removed. 
 When the steel point assumes the color desired, leaving a given 
 degree of hardness, the tool is finally cooled by immersion. A 
 tool plunged at a straw is very hard, while one allowed to anneal 
 by reheating to a blue is quite soft. 
 
 Caution is urged that the plunged tool while tempering be 
 not held too long a time at a certain color-line, for it has a strong 
 tendency to break there when in use. The tool should be slightly 
 waved in the water. Pieces which are to be tempered throughout 
 must be allowed to "soak"; i.e., become uniformly hot before 
 plunging. 
 
 The Correct Temper for Tools. The proper color for a 
 given ground is only ascertained by experience. Generally 
 speaking, the picks and drills are stopped at a straw if intended 
 for hard rock and carried nearly to a blue for mild ground. It 
 is always desirable to preserve the toughness of the steel as far 
 as possible; therefore the lowest color is selected which is com- 
 patible with the sendee to be performed. A high-carbon steel 
 is given a lighter color than steel of low carbon. 
 
 Metal- working tools are given a pale straw-yellow; wood-
 
 MINERS' TOOLS. 633 
 
 working tools, a brownish tint; hatchets, saws, etc., a light purple; 
 picks, to a rose; cold-chisels, to an orange-rose; keydrifts, orange; 
 rock-drills, yellow-orange; screw-cutting dies, light yellow, and 
 hammer faces, a pale straw. Few, if any, miners' tools are carried 
 to a blue temper. They would be too soft. Crowbars should 
 never be hardened, as they would be too brittle under severe 
 strain. 
 
 Bits which have been properly tempered will wear down 
 uniformly till too blunt for further service. Those which are too 
 brittle will be found to have broken or cracked off. The former 
 might be again tempered at a lighter color and the latter should 
 have been softened further during the sharpening. 
 
 REFERENCES. 
 
 The Testing of Explosives in Belgium, Coll. Guard., Jan. 9, 1903; The 
 Theory of Misfires and some Conclusions of Practical Value, E. H. Weis- 
 kopf, Jour. Chem. and Met. Soc. of S. Africa, Nov. 1902; Theory of 
 Misfires, Coll. Guard., Vol. LXXXIV, 689; The Use of Explosives in 
 Mines, Coll. Guard., Vol. LXXX, 738; Explosives in Building, Coll. Guard., 
 Vol. LXXIX, 32; Safety Explosives, Coll. Guard., Vol. LXXIX, 832. 
 
 Shot-firing, M. & M., Vol. XX, 409, 426; Blasting Agencies, M. & M., 
 Vol. XXII, 301; Blasting Rules, M. & M., Vol. XXII, 501; Testing Blast- 
 ing Caps, M. & M., Vol. XXIV, 302; The Testing of Explosives in Bel- 
 gium, Coll. Guard., Jan. 9, 1903; The Theory of Misfires and some Con- 
 clusions of Practical Value, E. H. Weiskopf, Jour. Chem. and Met. Soc. of 
 S. Africa, Nov. 1902; Modern Explosives, J. S. S. Erame, Nature, April 5, 
 1900; The Question of Underground Magazines for Explosives, Franz 
 Prispisle, Oesterr. Zeitschr. f. Berg.- u. Hiittenswesen, March 17 and 24, 1900; 
 The Theory of Safety Explosives, H. Heise, Gluckauf, March 24, 1900; 
 Explosives on Witwatersrand, T. Lane Carter, Eng. & Min. Jour., Nov. 7, 
 1903; Testing Mine Explosives, A. W. Warwick, Mines & Min., Oct. 1902; 
 Some Tests of the Relative Strength of Nitroglycerine and Other Explo- 
 sives, F. N. Clark, Vol. XVIII, 515; Dynamite Accidents and Preventions, 
 James Ashworth, Coll. Guard., Nov. 1896, 927; List of Flameless Explo- 
 sives, Copy of Orders, Coll. Guard., June 1897, 1 129; Preventions of Accidents 
 from High Explosives, James Ashworth, Coll. Guard., Nov. 1896, 927;, 
 Smokeless Powder, E. & M. Jour., Vol. LVI, 117; Testing Explosives for- 
 Coal-mines, E. & M. Jour., Vol. LXI, 567; Manufacture and Use of. Dyna- 
 mite, Harry Lee, E. & M. Jour., Vol. LXI, 182.
 
 CHAPTER VII. 
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 
 
 Machine Rock-cutters. The successful substitution of ma- 
 chinery for hand labor has proven a most important advance in 
 mining engineering. Hard rock is no longer an obstacle to 
 extraction of fuel, ore, or rock, and very long and large tunnels 
 are now rendered possible in a comparatively short time. The 
 opening of mines is accomplished in such short time and with 
 more prompt returns for the investment that machine drills are 
 employed to the exclusion of hand labor for driving shafts and 
 tunnels. Every form of hand-labor tool has been successfully 
 imitated and extensively introduced. The quarry methods of 
 lewising the jumper, saw, chisel, pick, and auger find their counter- 
 parts in the channeler, percussive drill, coal-cutter, and diamond- 
 drill. 
 
 Quarrying. The quarrying of dimension stone was formerly 
 accomplished by the trenching along lines decided upon. Carried 
 often to a lo-foot depth and wide enough for a man to operate 
 his pick, these trenches wasted much good material. These 
 trenches are now replaced by channelers and gadders, which 
 dig as deep as desired, but only 2 or 3 inches wide. These 
 machines are mounted in different styles and cut perfectly true 
 lines at any angle with or across the strata. 
 
 Channelers. For extensive quarries these machines are 
 mounted on a portable slining carriage, with boiler, rails, etc., 
 and a feed which automatically moves it with the progress of 
 its channel. A set (gang) of five cutters receives a reciprocating 
 
 634
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 635 
 
 FIG. 298. A Single-gang Channeler.
 
 6 3 6 
 
 MANUAL OF MINING. 
 
 motion from a steam-piston, through a connecting-rod, or through 
 some yielding contrivance from the cross-head of the engine. 
 The latter gives an elastic blow to the cutters. Automatic con- 
 trivances keep the cutters to their work. Machines are also 
 supplied for cutting two channels at a desired distance 
 apart; these are known as "double-gang machines" and cost 
 
 FIG. 299. A Broach-bit. 
 
 from $1200 to $2000 complete. With three men and 400 Ibs. of 
 coal, at 150 strokes per minute, they cut from 75 to 400 sq. ft. 
 of stone the former in marble, the latter in soft lime and replace 
 fifty men (Fig. 298). 
 
 The Broach-bit. Many quarries employ a small machine 
 similar to the channeler mounted on a transversing and long 
 stout bar which maintains the alignment of the work. Its frame 
 (Fig. 299) is comparatively light and is adjustable to a high or 
 low position. Its bit may be a broad channeler bit and drive 
 rectangular holes in any direction and to any desired depth, or, by 
 permitting the rotation of the bit with each blow, the ordinary
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 637 
 
 round hole may be drilled by the same machine. The broach- 
 bit proper is a channeler-bit which breaks down the partitions 
 between the holes drilled by a previous bit, as in Fig. 299. By 
 this machine the channeler is cut to any length and depth desired, 
 the line between the holes being directed by the stout bar of the 
 machine. This machine is also employed for drilling holes for 
 the plugger and feather work. By it 300 linear feet of 2 -foot 
 holes are put in ten hours, or 70 square feet of channeling in 
 sandstone and 28 square feet in granite. 
 
 Lewising is accomplished by machine by fitting a tripod with 
 a slot movement for the drill body such that three parallel holes 
 can be drilled with a single setting. 
 
 Power Drills. Reciprocating drills employing steam or air 
 as a motor fluid are of common use for piercing the rock. The 
 construction of the drill is the same for either fluid, though the 
 name given to the drill varies with the fluid used. A cylinder 
 of cast iron is capable of sliding in a guide bed-plate (Fig. 300) 
 which is mounted on a tripod or a column. The cutting-tool 
 is clamped to an extension of the piston-rod and the balance of 
 the mechanism is such as is necessary to obtain a return of the 
 drill with its progress and reversal of the motion of the piston. 
 The essentials for a successful drill are simplicity in construction; 
 as few moving parts as possible; a strength commensurate with 
 light weight and easy transportation ; variable stroke to the piston, 
 whose blow shall be cushioned; and an automatic return and 
 revolution of the drill. The weight of the machine must be 
 sufficient to receive securely a shock of the blow. Thus a 
 piston having a diameter of 2^ inches operating under an 
 air pressure of 60 Ibs. per square inch would receive a con- 
 stant accelerating force of 294 Ibs. with every stroke. To 
 counteract this recoil its mass must exceed this in amount. As 
 the steam- or air- drill has its largest sphere of usefulness in 
 remote mining districts it should be durable and require a 
 minimum of repairs. 
 
 The motor fluid underground is usually compressed air and 
 on the surface steam, with an ordinary pressure of 50 to 80 Ibs.
 
 6 3 8 
 
 MANUAL OF MINING. 
 
 per square inch. The horse-power of the drill is estimated as a 
 simple steam-engine, with the important difference that the ratio 
 of the area of piston-rod to piston is larger. Again, the steam- 
 
 FIG. 300. A Steam- or Air-drill. 
 
 engine does its work throughout the entire stroke, but the drill- 
 engine only at the end of its stroke. Hence it can never work 
 expansively. The air enters the cylinder and propels the piston
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 639 
 
 to the end of its stroke, when the attached drill strikes the rock. 
 At that moment the piston automatically reverses its own valve, 
 which admits air at the lower end of the cylinder, while a ratchet 
 and spiral device slightly turn the tool, which is being drawn back 
 for the next blow. As the work to be done on the return-stroke 
 is merely to lift the tool, the annular area of the piston is much 
 less than that on the other side, and little power is consumed. 
 At the proper point in the up-stroke the valves are again reversed 
 and the operation repeated. 
 
 The Automatic Steam-valve. The main difference between 
 the drills upon the market is in construction of the valve and the 
 means by which it gives a reversal of motion to the piston. It is 
 desirable for rapid penetration that the number of blows per 
 minute should be a maximum, but structural difficulties place a 
 limit to the length of time between each blow. It is essential that 
 the full air pressure be exerted at the instant of the impact and 
 that the valve should not reverse until the blow has been struck. 
 Then it should reverse instantly, permitting only sufficient steam 
 to remain in the clearance space under the piston to safely cushion 
 the blow against the front head of the cylinder. The average 
 speed is about 200 blows per minute. The frequency of the 
 blow varies with the ability of the machine and is altered to 
 suit the hardness of the rock. A short stroke, light blow, and 
 rapid rate give the best progress in hard rock, and a hard blow 
 is best in soft rock, provided the drill does not "stick" in the 
 hole. 
 
 There are two systems of moving the valves. The first 
 requires reversing rod and tappets, while the second is a duplex 
 system requiring a fluid to reverse the motion. 
 
 The Tappet-valve. In the earlier forms of drills the valve 
 was operated by means of an external rod with an exposed three- 
 armed tappet moved by a projection on the piston-rod, as in the 
 steam-pump. The rate of speed of a drill, however, was too 
 high for the mechanism to withstand the numerous and violent 
 shocks ensuing at the time of the reversal of the valve.
 
 MANUAL OF MINING. 
 
 m 
 
 1 
 
 The later forms of the tappet are 
 concealed. The arc of its motion 
 is reduced and a more compact 
 machine is the result (Fig. 301). 
 
 The valve motion was positive, 
 and though there are many disad- 
 vantages inseparately connected with 
 it, it is safer in the hands of unskilled 
 labor. This accounts for its reten- 
 tion in our present successful forms. 
 
 The little Rand giant drill (Fig. 
 300) is a tappet-moved valve and 
 requires less dead space and con- 
 sumes less steam than the Burleigh 
 pattern, which it resembles, whose 
 piston operates two rockers to oscil- 
 late its valve. Centrally in the cylinder 
 a three-toed rocker is located, the 
 upper toe being fitted into a recess 
 in the valve which it moves and the 
 two lower toes being alternately rocked 
 by the piston, to produce the reversal 
 of motion. By thus separating the 
 spindles from the valves and the 
 tappets which they connect, a greater 
 durability was obtained than in the 
 earlier types of drills. In the Ser- 
 geant tappet-drill the valve and the 
 rocker are also in a single three-toed 
 piece, driven alternately by either 
 end of the piston. 
 
 In the Sergeant (Fig. 303) the 
 piston-valve is moved by exhaust 
 steam from the opposite ends. An 
 auxiliary slide-valve moves over the 
 arc of a circle by shoulders on the
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 641 
 
 piston, opens and closes the ports, and is a trigger regulating 
 the movement of the main valve. There are no openings in 
 the side of the cylinder and no ports for the piston to close; the 
 exhaust remains open at one end till the blow is struck, when the 
 valve reverses immediately. 
 
 The Fluid-driven Valve. This is a piston-valve which reverses 
 the motion of the piston by opening the appropriate ports for 
 admission and exhaust of steam or compressed air, being itself 
 cause to oscillate in its valve- chest by the motor fluid which is 
 automatically admitted to one head or the other of its piston. 
 It is adopted in the Rand "Slugger" (Fig. 302), the Ingersoll 
 "Eclipse" (Fig. 303), the Schram drill, and others. The first 
 two named are of similar construction and action, and are shown 
 in section. Two port- holes connect the annular groove in the 
 piston with each opposite end of the valve-chest and are opened 
 or closed by the piston passing over them; the supply for one 
 end and the exhaust to the other end of the valve-chest are 
 simultaneously opened. The annular groove, therefore, is a gen- 
 eral exhaust outlet for the valve steam, while the motor steam 
 is exhausted by the valve connecting the inlet passage with the 
 exhaust-pipe. 
 
 There are no means on the piston in Fig. 302 for maintain- 
 ing a steam-tight or air-tight separation of the two ends. When 
 wear occurs the steam pressure is lost and leakage ensues; the 
 exhaust becomes imperfect and the valves may fail to act properly. 
 When in practice the machine is found not to reciprocate properly 
 the fault is usually found to be due to wear. 
 
 The Schram Drill. An account of the Schram and the Darling- 
 ton drills is to be found in Andre's "Mining Machinery," from 
 which the following is taken : " Schram's consists of a slide-valve 
 and a slide-rod that admit steam to the cylinder for raising the 
 piston and drill. When the piston passes a certain front port- 
 hole, steam enters through it into the back of the valve-chest 
 at the same time that the front valve-chest, through the other 
 port and the hollow circular groove of the piston, communicates 
 with the exhaust-pipe. Steam then works full pressure on the
 
 6 4 2 
 
 MANUAL OF MINING.
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 643 
 
 slide cylindrical rod, which, with the slide-valve, is forced towards 
 the front valve-chest, so that the back steam passage is open 
 to the cylinder, and the front steam passage connects with the 
 exhaust-pipe. The piston moves forward, and, when it passes 
 the back port, allows the steam to enter the front valve-chest 
 at the same time that the back valve-chest, through its back port 
 and the circular groove of the piston, communicates with the 
 exhaust. The slide-rod is forced back, the front steam passage 
 opens, and the back passage communicates v/ith the exhaust. 
 The slide is in the form of two spindle-valves, so that it remains 
 in position without recoil, and the annular groove of the piston 
 is always in communication with the exhaust. 
 
 " The Darlington Drill has only two working parts, an extreme 
 of simplicity: a cylinder and its cover, and a piston and its rod. 
 The piston is made to operate as a valve. The inlet-pipe, having 
 open connection with the cylinder, always furnishes the pressure 
 to lift the drill, which rises whenever there is no pressure on 
 the back. On its way up, the piston first covers the exhaust 
 (above the inlet), and then uncovers an equilibrium passage, 
 by means of which communication is established between the 
 front and back ends of the cylinder. Then air or steam enters 
 and operates over the greater area, at the back, and first checks 
 the upward movement, soon overcomes it, and finally produces 
 a forward motion. The propelling force, now, is dependent 
 upon the difference of area between the back and front of the 
 piston. On its way down it soon cuts off the equilibrium passage 
 and the air can only enter at the inlet; the steam operates by 
 expansion for a short space, till the piston has passed and uncovered 
 the exhaust-port, when a discharge takes place as the blow is 
 being struck. One fact is noticeable, that the amount of steam 
 used is only that necessary for the down stroke; for that used 
 to raise the drill escapes by the equilibrium passage to the top." 
 
 The Drill-rods. The steel drill tool is of a diameter depend- 
 ing on the size of piston, and from f to i^ inches, according 
 to the intensity of the blow to be struck. The smallest size 
 is attached to a 2 -inch piston and the largest size mentioned
 
 644 MANUAL OF MINING. 
 
 to a 5-inch piston, corresponding to blows of 200 Ibs. and 1200 
 Ibs. respectively. The average size of the mining drill is a 3-inch 
 piston with i- or i^-inch steel. The drill-rod is clamped by a 
 heavy split chuck locking into the enlarged end of the piston-rod, 
 and reciprocates with it. The end of the drill or shank varies 
 in dimensions according to the piston size. For a 2-inch-diameter 
 piston the shank is f"X-g", for a 2-inch piston i$"X7", for a 
 piston of 4^-inch diameter and proportionate sizes for the inter- 
 mediate diameters. It reciprocates through a split or solid spool 
 front head, which is provided with a stuffing-box and a gland 
 to prevent the escape of steam. The latter head is adapted to 
 air-power, while the split fronts are preferred for steam-drills. 
 
 The bit is usually forged with a bluff cutter edge for the 
 strength and is provided with the usual flare. For rocks which 
 do not crush, but chip in large fragments, a sharper edge will do 
 better execution. The cutter is dressed by special tools to the 
 X, I, Z, or S form, each having its specific value, the S form 
 being more likely to maintain a round hole. 
 
 A set of tools accompanying each drill is graded according 
 to the amount of feed which the drill is capable of. The shortest 
 and the stoutest drill of the set is about 15 inches in length and 
 the other lengths are successively longer by the amount of feed 
 of the machine. The longest drill in the set corresponds to 
 the maximum depth of the hole. A machine feeding 15 inches 
 and drilling a hole 10 feet deep will have 8 drills in a set, the 
 difference between their lengths being 15 inches. For a 5-foot 
 hole, 3 or 4 drills constitute the set. The bits should be hardened 
 in such a manner as to have durability corresponding to the 
 length of the feed. A bit usually will require sharpening after 
 4 inches of progress. 
 
 During drilling the tool must be turned uniformly, as other- 
 wise the hole may become rifled. By rifling is understood the 
 tendency of a drill to cut a triangular hole instead of circular 
 hole, which it strikes successively at one point or along a given 
 line. With the straight-edge cutter the tendency to rifling is 
 stronger than with any other form. ' In rock of an unhomogeneous
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 645 
 
 character rifling is common. With deep holes this may also 
 result in a deflection of the hole from a straight line. 
 
 To prevent rifling and to produce rotation of the drill through 
 a small arc during each stroke, the fluted bar-nut constituting 
 the long thread ratchet is a 
 usual appliance. In the 
 Burleigh and the Darlington 
 drills the device is a spiral 
 feather on the piston-rod, re- 
 cessed into a grooved piece 
 in the cylinder-head. It is 
 toothed and held by a detent, FlG - 304--The Rifle-bar. 
 
 which permits it to turn on 
 
 the forward stroke, but prevents turning during the up stroke 
 of the engine. In the Ingersoll drill a grooved bar fitting into 
 the back of the piston turns it on the back stroke and is itself 
 allowed to rotate on the down stroke (Fig. 304). 
 
 In the Darlington drill the ratchet turns the piston and drill 
 on the up stroke, and itself turns during the down stroke. 
 
 In the Schram pattern an auxiliary piston turns the drill, the 
 former being driven by the fluid in a manner similar to that of the 
 valve. 
 
 The Feed. Whatever may be the pattern of the drill, the 
 method of feeding the cylinder and its piston to the work is the 
 same. It is always under control of the drill-runner and is capable 
 of adjustment to the progress in the rock. The cylinder and 
 its tool is allowed to move in a guideway which is rigidly 
 mounted in the frame by the means of a square thread feed-screw 
 which is turned by a crank at the rear end of the machine. The 
 drill-runner usually stands at work with one hand on the feed- 
 handle crank and the other on the steam-throttle valve. 
 
 Cushioning the Blow. The maximum percussive effect is 
 obtained when the blow is struck under the full head of steam 
 or air. Such a dead blow is highly desirable; but on account 
 of the shock to the machine, for the subsequent repairs this is 
 unadvisable. A clearance space is provided in which is enclosed
 
 6 4 6 
 
 MANUAL OF MINING. 
 
 a small amount of air to act as the cushion or elastic buffers are 
 inserted to terminate the stroke. The former is less expensive 
 than the latter. 
 
 The Drill Supports. A rigid support is a rigid adjunct to 
 the drill, and several types of mountings are provided, each 
 having a special end in view, though a machine can be shifted 
 from one style to another. In tunnels and shafts where the 
 ranges of holes have approximately parallel directions, it is 
 clamped directly upon a stout, hollow, cylindrical column (Fig. 
 305), or upon an arm projecting from it. 
 
 Ho. 305. The Column Support for Drills. 
 
 This admits of drilling several holes from one position of 
 support. Jack-screws at one end clamp the column, while 
 claws at the other end bear against blocks resting upon the rock. 
 The column can be had in various lengths. It weighs about 
 30 Ibs. per foot. When placed horizontally, as in shafts (Fig. 
 305), it is known as a shaft-bar. [When used in the vertical 
 position it is spoken of as a column. 
 
 The tripod is the more common drill support, having a wider 
 range of position than the column because of the universal joint
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 647 
 
 at its head (Fig. 306). Its stability depends upon its weight, 
 and additional dead weights are, therefore, clamped on each 
 leg as in Fig. 300. 
 
 FIG. 306. Blocking out Dimension Stone. 
 
 The drill has no intricate mechanism and can be intrusted 
 even to unskilled labor. Care should be taken in beginning 
 operations that the drill strikes squarely upon the face of the rock, 
 and with short, light blows. It can do more work with the con^ 
 sumption of less pow r der, steel and blacksmith w r ork than can hand- 
 drills. It can be set up in any place capable of accommodating 
 a double-hand gang of men, and gives little trouble in placing 
 into position. It cannot be adopted in small veins or where the 
 mineral is friable or occurs in thin streaks. 
 
 The comparative merits of the several types of machines can- 
 not be stated, for in one mining-camp the Rand drill, and in 
 another the Ingersoll drill, are exclusively used. The Waring 
 drill, the National, and the Burleigh, are employed for the same 
 operations in the different mines of the same camp with appar- 
 ently equal success. One range prefers the little Giant, but in
 
 648 MANUAL OF MINING. 
 
 another mine it is discarded for Slugger pattern of the same 
 company. In like manner a preference may be displayed for 
 the Eclipse- or Sergeant- drill made by the same company. They 
 are all highly commended, and their improvement in the inland 
 community may be a matter of accident or of natural selection 
 after periods of test. 
 
 The fluid- moved valve- drill style seems the favorite pattern 
 for hard rock; but whether under all circumstances it is the 
 best, one would not dare to aver. The nature of the rock, the 
 proper air pressure, the rate of speed, and the proportion of 
 rotary motion necessary for a maximum effect vary so largely that 
 a comparison of the numerous published results cannot be made. 
 
 Progress and Cost by Machine-drills. -An average of nine 
 neighboring mines in conglomerated rock shows machine-driving 
 and stoping to be, respectively, 22 and 36 per cent cheaper than 
 hand, and sinking 4 per cent dearer, with the progress respectively 
 60, 54, and 38 per cent more rapid. The gain in time during 
 sinking more than compensates for the cost. In iron- mines the 
 product of the machine labor costs about one fourth as much 
 as that of hand labor. The consumption of fuel or air per drill 
 may be calculated in the same manner as for the ordinary steam- 
 engine. The cost of the average mining drill is about $325, 
 and a complete plane with its drills and suitable compressor, etc., 
 is $7000. 
 
 Systems of Machine-drilling. With the advance made in 
 the use of the power-drill, the systems of drilling and blasting 
 rock have undergone corresponding changes. In hand-work the 
 object sought when placing a hole for drilling and blasting is 
 as much to secure a good bench for the succeeding shot as to 
 break a maximum of ground. With machine-drilling, simul- 
 taneous shooting is practised and numerous holes are drilled 
 over the entire area of shaft or drift, and fired in volleys with- 
 out regard to the probable condition of the face after the blast. 
 The removal and the resetting of the machine after each blast 
 occupy so much time that a maximum number of holes is drilled 
 with each set-up, and the time thus saved more than compensates
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 649 
 
 for the value of the powder wasted by not conforming to the 
 fundamental principles of blasting, as in single shooting. 
 
 
 Systems of Drilling Holes. There are two general systems in 
 practice for the drilling of holes, the first being known as the 
 centre-cut system, employed for tunnels and shafts of an area
 
 650 MANUAL OF MINING. 
 
 larger than fX$ f . The second system is known as the Brain 
 radial system, which is employed in headings only large enough 
 to accommodate one machine. In either system as many machines 
 as possible are set up against the face. Three are employed in a 
 single-track tunnel; two in a n-foot heading; and as many as 
 six are arranged in a double- track tunnel 27 feet wide. It is not 
 unusual to mount two drills on a single column. The holes 
 are drilled as determined upon, and after loading are blasted 
 for an advance of 6 to 10 feet with each set-up. 
 
 The Centre-cut System is of almost universal acceptance in 
 America. Vertical tiers of holes are drilled over the face, as in 
 Fig. 307. 
 
 The method of placing the holes is shown in Fig. 308, the 
 
 arrows indicating approximately the direction in which they 
 
 are drilled. The top or back holes 
 
 o> a . . 
 
 | aa are pointed a little upward in 
 
 t<* order to obtain the top height and to 
 
 b\ ^^ | ^^ Jb break down the rock left after the 
 
 _c j other holes have been fired. The 
 
 side holes bb are pointed a little 
 
 *" "* * upward and outward to obtain the 
 
 b\ ^* ^-^. c tb full width of tunnel after the centre 
 
 FIG. 3 o8.-Direction of Holes holes are fired ' The Centre tiers of 
 for the Centre-cut System for holes cc are pointed centreward, the 
 
 Wide Gangways. * . 
 
 bottom ones of the tier being given an 
 
 upward slant, the top ones a downward slant. Sometimes a 
 plunger- hole d is drilled and fired with the centre holes. The holes 
 are fired in rounds. Those at the centre, having the hardest 
 work, are charged with a higher grade of powder and throw out 
 a wedged-shaped mass; the side tiers b, being charged with a 
 lower grade of dynamite, are fired later together, after which the 
 back and trimming-holes are exploded together to complete 
 the cut. 
 
 In the use of the centre-cut system for wide tunnels, in con- 
 nection with the American methods of tunnelling, as illustrated 
 in Figs. 309 and 264, the centre holes n (Fig 310), fired first,
 
 CHANNELERS, DRILLS, AND-COAL CUTTERS. 
 
 05* 
 
 facilitate the work of the next two rows 2 2, after which the holes 
 3 3 are fired in volleys, succeeded by the trimming-holes 4 4. 
 
 The profile should be finished with the advance of the face, 
 regularly checked up by the surveyors, for it seems difficult to 
 maintain the proper floor level, the tendency invariably being 
 to gradually raise the floor wiih the progress of the work in an 
 unconscious effort to insure drainage. 
 
 FlG. 309. Drilling Bench-holes in the American System of Tunnelling. 
 
 In driving such a tunnel the work is conducted in three shifts 
 of eight hours each. During the first shift the drill-runners and 
 their helpers are engaged in operating the machines and drilling 
 the holes to their proper depths. The powder-men follow; they 
 remove the machines from the face to a safe distance, and charge 
 the holes with explosive and ignite them. In the third shift the 
 broken material is removed by loaders, who tram the car sfrom 
 the face to the point of discharge and also deliver the drill and 
 appliances to the face for the drillers.
 
 6 S 2 
 
 MANUAL OF MINING. 
 
 Drilling in Benches. The bench of the tunnels may be 
 a. tacked in two sections, as in Fig. 309, A and B, or in one only, 
 as in Fig. 310. In the former, two wall-holes, one or two trans- 
 verse rows of four top holes downward, and half a dozen bottom 
 holes, lift each bench with every other shift. Fifty-four feet of 
 progress a week is the record on a very hard sandstone bench 
 I4'X27'. This bench- work is accomplished, not only more 
 
 FIG. 310. The Centre-cut System of Drilling Wide Tunnels. 
 
 rapidly, but also with a powder consumption per cubic yard of 
 rock of about one half that in the heading above it. 
 
 The Centre-cut System in Shafts. The method of placing 
 holes in shafts does not materially differ from that in the head- 
 ings. In Fig. 311 is the elevation of a 6-foot cut in a shaft whose 
 area is io'X2o'. The centre holes are placed 10 feet apart 
 approaching, but not intersecting. Six pairs of holes, spaced 
 2 feet apart, extend across under the shaft. The holes bb are 
 in two pairs of rows of five each, the first being given a greater 
 inclination than the second row. They are fired in separate tiers 
 of ten holes each. Occasionally all of the twenty holes are fired
 
 CHANNELERS, DRILLS,. AND COAL-CUTTERS, 
 
 653 
 
 together, followed by the squaring-up holes cc. For a deeper 
 cut the centre holes (Fig. 312) are in two tiers, those in tier a nearly 
 meeting at 6 feet from the shaft floor and the other bb at u feet. 
 The other holes are driven as before. In this case the 6-foot 
 centre is opened first, followed in sections to the sides by the 
 others. 
 
 
 X 
 
 x & 
 
 X. K 
 
 a x x . 
 
 b x. 
 
 x 
 
 x y 
 
 x x 
 
 & X 
 
 * b: 
 
 x xa 
 
 x x x 
 
 
 FIG. 311. The Six-foot Centre Cut 
 for Shafts. 
 
 FIG. 312. The Ten-foot Centre Cut 
 for Shaft-sinking. 
 
 Brain's Radial System. This is designed for small faces 
 which can accommodate but one machine at a time. Like the 
 centre-cut system, it is equally applicable to shafts. All the 
 holes to be drilled are put from one position of the machine, which 
 is removed only when firing is to be done. This minimizes the 
 time lost in shifting. In a certain case the machine, from a position 
 4 feet 8 inches from the bottom, 2 feet from the top, and 2 feet 
 6 inches back from the face, put twenty-nine holes with a total 
 length 70 feet, advancing 3 feet with an average of 2.4 cubic feet 
 broken rock per lineal foot of hole. 
 
 The holes are comparatively shallow, though they vary greatly 
 in length, those making the smallest angle with the face being 
 the longest. Four ranges of holes are drilled. Sometimes a few 
 extra squaring-up and lifting holes are necessary to trim the
 
 654 MANUAL OF MINING. 
 
 periphery of the drift, but ordinarily the firing of the most angling 
 holes first breaks out the rock to daylight and opens a face for 
 the other successive rounds. The advance cannot be large, 
 for neither deep nor angling holes are possible in a narrow drift. 
 In a drift 8 feet wide, two settings of the machine are sometimes 
 made drilling from near each wall and thus forming a modified 
 centre-cut plan. In some mines a practice prevails of cutting 
 a horizontal range of bottom holes, two ranges of holes looking 
 downward, and a top row to break out horizontal instead of 
 vertical wedges; this plan requires a bar- mounting for the drill, 
 in a drift of say fx&. 
 
 The Continuous Process of Diamond-drilling. Gen. Henry 
 Pleasant's method of shaft-sinking is a novel and eminently suc- 
 cessful application of the diamond-drill. One or more diamond 
 drilling-machines are set up over the site of the shaft, and bore 
 vertical holes as deep as the shaft is to be carried. The machines 
 are moved to new positions and additional long holes bored. 
 The operation is continued until the entire area of the shaft is 
 pierced by holes at suitable distances apart. One shaft had 
 thirty-five holes bored over its area of 25' 8"Xi3' 10" to a depth 
 of 300 feet in six weeks by three machines. 
 
 When the "continuous process" is completed, the machines 
 are removed for the blasting. The holes are filled with sand or 
 water for the full length, except in the upper 3 or 4 feet, w r hich 
 are treated like short holes, charged with dualin and fired, the 
 central ones first. When the debris has been cleared away, the 
 shaft will have advanced 3 or 4 feet. A few feet of each hole 
 are again cleaned out, loaded, and fired. Thus each section 
 advances with an uninterrupted alternation of shooting and 
 hoisting. Though it is not always cheaper per cubic foot, it 
 effects a great saving in time, and quick access under ground 
 may prove the element essential to the success of the undertaking. 
 
 Boring Headings. The diamond-drill cannot be used advan- 
 tageously for tunnel work or driving headings. Percussion-drills 
 are there used almost exclusively, though rotary machines have 
 been employed for boring out a heading almost to full size. The
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. . 655 
 
 Stanley header represents this class of machine. At one opera- 
 tion a series of cutters on a rotating boring-head grinds away 
 the whole face for the core of a heading some 7 feet in diameter. 
 One was used in the Mersey subaqueous tunnel. It travelled at 
 the rate of 39 inches per hour and executed its work satisfac- 
 torily in the argillaceous chalk. 
 
 The approximate cost of sinking and tunnelling work varies 
 with the character of the rock, the depth, and excavation, of dis- 
 tance from the surface; the efficiency of the workmen also deter- 
 mines largely the economy. In the New house tunnel, with an 
 area of 144 sq. ft., the cost of driving was $28.80 per foot, in the 
 Wooster tunnel it was about $40 per foot, and in the Aspen tunnel 
 $19 per foot. The cost of driving is usually greater through 
 soft ground than through hard ground, on account of the increased 
 quantity of timbers required and the difficulties of draining the 
 water. The cost of driving a rock tunnel 5' X f on contract 
 without timbering is $8 per foot. An upraise can be driven at a 
 cost of between $5 and $10 per foot, varying with the size and 
 the character of the timbering needed. A two- compartment 
 shaft with the timbers can be sunk 6' X 12' in the clear at an 
 expense of $25 per foot. 
 
 Electric Rock-drills. The various attempts to introduce an 
 electric percussion-drill to replaced compressed air-drills have 
 resulted in only a partial success. The advantages which elec- 
 tricity present in transmission over compressed air are sufficient 
 to suggest an installation of drills. The difficulty, however, lies 
 in the inability to produce a reciprocating motion without the 
 use of solenoid. 
 
 The Solenoid Drills. Two coils of wire are wound in the 
 form of a spiral through which a current is alternately passed in 
 the one direction and the reverse. If a steel core is passed through 
 the adjacent coils and a current of electricity be passed through 
 one of them the coil will be repelled through the other coil. The 
 latter being next charged, will give a reverse action to the coil 
 and repel it toward the first. This sets up an oscillatory move- 
 ment which, if the machine be perfect enough and the coils carry
 
 656 MANUAL OF MINING. 
 
 sufficient current, will accomplish work similar to the percussive 
 machine (Fig. 313). 
 
 ki^. jij. A Solenoid Drill. FIG. 314. A Gardner Drill. 
 
 The Siemens, Halske, and Gardner Electric Drills are based 
 upon the attempt to produce indirectly a reciprocating motion. 
 The current does not enter the drills, but drives gearing housed 
 in the drill cylinder by means of a motor and a flexible shaft. 
 The drill is mounted on an adjustable column arm or on a tri- 
 pod (Fig. 314). This machine produces a rapid succession of 
 light blows rather than a small number of heavy blows. It 
 requires 2 to 3 horse-power to operate it. It is also quite heavy. 
 
 The Box Electric Drill dispenses with the flexible shaft. In 
 the rear of the drill is a motor, geared by a pinion to a wheel 
 on the crank-shaft. At the other end of the crank is a pinion 
 which operates a train of gearing, turning the drill after each 
 blow. Between the crank-shaft and the small cylinder is a 
 connecting-rod. 
 
 Reciprocating motion is imparted to the cylinder by the 
 crank. The heavy piston with a thick piston-rod constitutes 
 the hammer whose weight determines the blow. 
 
 Each end of the moving cylinder is filled with air at atmos- 
 pheric pressure, which being alternately compressed and rarified 
 cushions the blow. Two ports located somewhat centrallv at the
 
 CHANNELERS, DRILLS, AND COAL-CUTl ERS. 657 
 
 side of the moving cylinder are provided for admitting air when- 
 ever the piston uncovers them. The weight of the drill with its 
 motor is about 350 Ibs., the latter alone being 100 Ibs. 
 
 The Adams All-steel Drill is simple in mechanism, has few 
 parts easily gotten at, and all wear taken up, since every 
 bearing is adjustable. The drill is actuated by a loose rod run- 
 ning through the gear-case and motor, and this imparts motion 
 through a pair of bevel gears and crank-shaft to the draw-bar 
 and piston, the latter being cushioned with gangs of helical 
 springs which serve to absorb the shock of the blow to protect 
 the parts from injury and also to store up energy on the back 
 stroke, which it can expend in forcing the steel into the rock. 
 These springs prove extremely useful when reaming or for 
 pulling the drill out of a fissured hole. The drill, striking from 
 575 to 600 blows per minute, will consume less power for its 
 work than an air-drill. It consumes three horse-power at the 
 generator to operate it. 
 
 It is claimed that in point of first cost the electric drill has 
 the advantage over an air-drill, for a one-drill plant complete, 
 operated electrically, costs one half of an air compressor and 
 drill. The simple and smooth-running dynamo is less noisy 
 than the complicated air-compressor as a source of power. The 
 objection to the electric drill, that it provides no means for ven- 
 tilating the workings as does an air-drill, is met by the equally 
 valuable advantage that light can be obtained from the same 
 wire that supplies the power. The comparative efficiencies of 
 the rotary electric drills, reciprocating electric drill with flexible 
 shaft, the solenoid reciprocating drill, and the air reciprocating 
 drill are respectively as i : 1.66 : 4.5 : 10. 
 
 Coal-cutting Machines. Coal is mined by machine in the 
 same manner as by hand labor. Holes may be bored by 
 machinery and the powder charges in them exploded, or the 
 coal may be undercut or sheared by machine and broken down 
 with or without the aid of powder. For boring holes in the coal 
 various types of portable augers are employed, such as shown 
 in the following chapter. For the undermining of coal there
 
 658 MANUAL OF MINING. 
 
 are two types of machines, one depending on percussion and 
 the other on abrasion, as produced by the chisel-edged teeth. 
 Shearing-machines are of a similar type of construction to the 
 undermining machines. The percussive machines are identical 
 in appearance and construction to the air- or steam-drills em- 
 ployed in rock with the exception of the character of the bit. 
 Of abrasive machines there are three classes using sharpened 
 teeth attached around a bar which advances broadside on the 
 coal face, on a link chain advancing normally against the face, 
 and on the edge of a rotary wheel at the side of the machine 
 advancing parallel to its face. The first two classes are breast 
 machines and the one last named is a longwall machine. The 
 motor power for any of these types may be electricity or air. 
 
 The same character of labor that the miner performs Li 
 removing and cutting a groove is imitated by the undermining 
 machines in the floor under the coal or on the bottom layer of 
 the coal itself. This groove is carried as far back from the face 
 as possible and over the entire width of the room. The shearing- 
 machine cuts a vertical groove along either side of the rib into 
 the coal face and to the depth the machine can reach, beginning 
 at the roof and extending to the floor. The use of either of these 
 machines removes from the miner's toil the most laborious por- 
 tion of his work 'as well as the most hazardous phase of his occu - 
 pation. The normal effort of the digger exerted under the 
 unfavorable conditions existing in the room is most wasteful U 
 applied and produces in addition an excessive amount of fip<i 
 coal, which is lost. The machine produces more coal with less 
 waste in a much shorter time. It reduces the labor of the minei 
 and results in subdividing the work formerly imposed upon 
 one man and increases the efficiency of each man by specializing 
 each branch of the work. From four to eight loaders follow each 
 machine and a total of twelve men, on the average, are employed 
 in a mine for each machine in service. Its introduction results 
 in an increase in the earning capacity of each miner besides 
 improving the conditions under which his labor is applied. 
 
 The introduction of machines decreases the number of delays
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 659 
 
 of standing shots. It gives a steadier output and concentrates 
 the operations of the mine because less territory must be kept 
 open for the desired capacity. It requires, however, a more 
 systematic development of the mine. 
 
 In Illinois the Legg machine is used in driving-rooms, and 
 elsewhere the Harrison, Jeffrey, Lechner, and Sullivan machines. 
 In Europe the machines in vogue are known as the Marshall 
 and Frith. 
 
 The Requirements of the Machine. It should be light and 
 capable of being handled by two men and occupy small floor 
 space or height, to admit of being moved around and between 
 the roof-props. It should be capable of starting in a corner of a 
 pillar or a loose end, and of cutting a groove, from wall to wall 
 of the room, to any height, right handed or left handed. 
 
 The Percussive Machines. The reciprocating-drill is largely 
 employed for underholing the coal in the Mississippi Valley 
 States. It is often known as the punch- drill, and has as its chief 
 representative the Harrison machine (Fig. 315), which requires 
 little explanation. It is a modification of the compressed-air 
 drill, having for a valve-motor a single rotary device. It is 
 mounted on low wheels on a platform inclined toward the face 
 to enable it to move freely while making a cut; it is guided to its 
 work by handles on either side and is prevented from recoiling 
 by wooden blocks back of the wheels. In addition to the drill- 
 runner, one helper is necessary to remove the chippings from 
 the face of the work. An air pressure of about 60 Ibs. to the 
 square inch is used expansively, the point of cut-off being fixed 
 by the runner according to the hardness of the coal. About 200 
 strokes are delivered per minute. Usually three bits are supplied 
 with each machine in lengths of 2, 4, and 6 feet. A considerable 
 slack coal is produced while digging a deep undercut of 5 or 
 6 feet, its height being 16 or 18 inches at the face and 3 inches 
 at the rear. The lateral dimensions and shape of the groove 
 are determined by the reach of the machine, but the entire face 
 of a room is entirely undercut by a succession of such grooves, for 
 which purpose the platform and machine are shiftsd sidewise.
 
 66o 
 
 MANUAL OF MINING.
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 66l 
 
 The average work of the machine is 80 lineal feet of face in ten 
 hours, which is equivalent to from 50 to 100 tons of coal per hour, 
 according to the height of the seam. The machine is compact 
 and can be employed where the roof is weak and the timbers are 
 close to the face of attack. Its weight is from 500 to 700 Ibs. 
 and its height rarely exceeds 17 inches. 
 
 The Sergeant Drill Company has an acceptable adaptation 
 of its drill for coal- work which is largely used in the Southern 
 States. 
 
 A variety of pick-machine is in use by which a 75-lb. pick 
 is swung by a bell- crank lever at a weight of 70 blows per 
 minute, cutting a hundred square feet of ground 2 inches high 
 and 42 inches deep. 
 
 The Breast-machine. In this type the coal-cutter with its 
 motor is mounted on a frame which is caused to slide, inside of a 
 substantial base frame, toward the face or breast of the coal, and 
 under it to the limit of the slide. The cutter consists of a num- 
 ber of hardened steel bits on an edge of a wheel or over a rotating- 
 bar in front of the machine ; or it may be a series of bits inserted 
 into an endless link-belt. The wheel or bar is forced forward, 
 cutting a width of grooves 24 to 44 inches to a depth of about 
 6 feet. The groove is from 2\ to 5 inches high. The chain is 
 drawn horizontally and uniformly around four wheels at the 
 corners of the sliding-frame, and as its cutters pass in front of 
 the machine the coal is cut to about the same width and depth 
 as above mentioned. After each cut the motor and its cutter 
 are withdrawn, the standard frame is moved along the face a dis- 
 tance equal to the width of its groove, and from the new position 
 the next advance is made. The power employed may be either 
 air or electricity, the former being preferred, unless the mine 
 contains considerable gas. 
 
 In Fig. 316 is shown the Jeffrey bar-cutter. This is a bar 
 fitted with numerous cutting-chisels, rotated by a link-chain 
 shown at the side of the machine. The bar-cutters are kept to 
 their work by a feed at the rear; their frame is shown advanced 
 slightly from its starting position. A pair of jacks, one at the
 
 662 MANUAL OF MINING. 
 
 front and another at the rear, braced by a screw against roof 
 and coal, gives stability to the machine. Moreover, the jack 
 at the front in a measure assists in holding up the coal from bear- 
 ing excessively upon the cutter-rod. This machine is still ser- 
 viceable in some coals, though its cutters have a tendency to 
 climb up into the coal. This is a great annoyance to the operator. 
 It also acts in its abrasion across the bedding-plane of the coal 
 instead of parallel to it, as does the chain-machine, and in this 
 respect is not so efficient. 
 
 The more modern type of breast- machine manufactured by 
 the Jeffrey Machine Company is illustrated in Fig. 317. This 
 
 p IG 3I7> A Jeffrey Breast-machine. 
 
 chain-cutter is the logical outcome of the bar- cutter of the past. 
 An endless chain carries a number of bits made of f-inch steel, 
 having a chisel cutting edge f inch wide. The chain, revolved 
 by the motor and suitable gearing, is advanced automatically. 
 The bits project from the chain sufficiently to make a cut 4$ 
 inches on either side of it (Fig. 318). The machine weighs about 
 2000 Ibs., but is easily removed to a new position by releasing 
 the jacks and skidding the frame, after which the jacks are reset. 
 The voltage of the electric current is 500, and 35 amperes are 
 consumed. Excellent results are obtained in gaseous coal, splint 
 coal, or block coal.
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 
 
 663 
 
 Longwall Machines. By inserting the chisel-bits into the 
 circumference of a rotary disc which is operated at high speed by 
 the electric motor, projecting it from the side of the machine, 
 
 a longwall cutter is produced by which the machine may be caused 
 to move parallel to the face of the coal, while the wheel or toothed 
 disc undercuts the coal for a depth equal to a little over half its 
 diameter.
 
 C6 4 
 
 MANUAL OF MINING. 
 
 The Winstanley machine consists of a rotary toothed disc, 
 capable of being tucked away under the carriage or turned out 
 against the face and revolved by two oscillating cylinders. With 
 an air pressure of 30 Ibs. per sq. in. and a machine weighing 
 1500 Ibs., mounted on a carriage moving along the track, the 
 progress of the cutter is 70 sq. ft. of underholing in an hour. 
 
 The longwall machine placed on the market by the Jeffrey 
 Machine Company is represented by Fig. 319. This is a 
 one-rail machine, balanced in such a manner that the cutter- 
 wheel may pass over all irregularities in the floor. A series of 
 
 FIG. 319. A Longwall Machine. 
 
 gears are arranged, winding a rope on a drum to move the machine 
 along the rail at a rate fixed by the operator between 8 and 25 
 inches per minute. The cutter-wheels are from 3 to 6 feet in 
 diameter, according to the thickness of the coal-bed. The bits 
 inserted on the circumference of the wheels make a cut of about 
 4 inches wide. On the top of the cutter-wheel are gear-teeth 
 engaging with the pinion and the engine-shaft, by which the 
 rate of rotation of the wheel can be regulated to that which is 
 most suitable for the coal. This machine is said to undercut an 
 average of 600 lineal feet of coal-face per day. 
 
 Shearing Coal-cutters. The shearing of coal by machine may 
 be accomplished by a percussive machine, mounted on high
 
 CHANNELERS, DRILLS, AND COAL CUTTERS. 665 
 
 wheels or by a chain-cutter turned in position to a vertical plane. 
 The reciprocating-machine, when shearing coal, is mounted on 
 wheels of a diameter of 3 feet, or even more, enabling it to assume 
 a wide range of position in making the vertical cut from a high 
 angle pointing upward to a steep angle pointing downward. 
 
 One variety of the shearing-chain cutter is shown in Fig. 320. 
 It is mounted on four posts, each supplied with a jack-screw, to 
 secure it to the roof and floor with suitable clamps that permit 
 of a rotary movement about the column and an adjustment to 
 any height. The armature-shaft of the motor is parallel to 
 the centre rail. The cut is commenced at the top of the coal 
 and a groove as wide as the machine will allow and as deep as 
 it will reach is finished, after which it is lowered for successive 
 
 FIG. 320. A Coal-shearing Machine. 
 
 and adjoining cuts. The cutters travel downward along the 
 coal, thus always drawing the chippings down and out. 
 
 Many of these coal-cutters are made adjustable in position 
 for undercutting or shearing. The latter is not an economical 
 method of procedure as compared with undercutting, but is 
 serviceable in firm coals and gaseous mines where it is advisable 
 to use powder. 
 
 Another shearing-machine, made by the "Sullivan Machine 
 Company," is of the pick type, mounted on a truck and pivoted 
 to swing up and down to the point. It" usefulness, however, 
 is confined to the centre of headings and it cannot be employed 
 for shearing ribs. The machine makes a cut 6 inches wide 
 and from 7 to 8 feet deep, averaging 30 lineal feet in a ten- 
 hour shift.
 
 666 MANUAL OF MINIXG. 
 
 The Comparative Advantages of Coal-cutting Machines. 
 Coal-cutting machines can be employed to advantage underground 
 when the coal is firm and in seams which are of standard dimensions, 
 free from eccentricities, and under a strong roof with a good floor. 
 They are employed when it is necessary to develop the colliery 
 rapidly and where it is essential to obtain a maximum proportion 
 of large coal. Th?y have an advantage over handwork in concen- 
 trating the field of operations. They require less men and are 
 especially advantageous where shooting off the solid rock is 
 practised and where there is no soft underlying layer in which 
 the miner can underhole the coal. They cannot be employed 
 to advantage in wet mines, where the work is intermittent, where 
 the pressure of the roof bears too heavy on the coal face, or where 
 the layer under the coal is pyritiferous, or in the anthracite coal, 
 which is full of slip- joints. 
 
 The statistics furnish ample evidence that the powder con- 
 sumption in machine mines is less per ton of coal than in hand 
 mines. Likewise the number of fatal and serious accidents 
 occurring in machine mines is much less. It usually requires 
 from three to five additional machines to maintain an uninter- 
 rupted work for seven machines. 
 
 A machine can cut a groove of 6 feet depth and 44 inches 
 width in from 3^ to 5 minutes, and with the time occupied in making 
 six changes the undercutting of a room 20 feet wide can be effected 
 in two hours' time. Assuming thirty minutes as necessary to 
 shift the machine to the 'next room and to reset it, it will be 
 possible to underhole four rooms in a day w r ith the employment 
 of two men. This in a standard seam of 4 feet thickness corre- 
 sponds to 144 tons. By handwork with two men at a face, eleven 
 rooms would have to be kept open to equal the supply of this 
 one machine, which operates in four rooms. A mine producing 
 looo tons of screened coal, 1300 tons of run-of-mine coal, can 
 obtain it by nine machines in thirty-six rooms or by 200 
 miners in TOO rooms. The operations therefore are more 
 concentrated, as less territory is kept open than in the hand- 
 work, and this is an important feature when there is taken
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 667 
 
 into account the maintenance of roads, ventilation, and supply 
 of timber. 
 
 One marked objection to coal-cutting machines lies in the 
 quality of coal which is produced. This arises from the fact that 
 the coal is broken in large lumps, which contain an undue pro- 
 portion of impurities. They require less explosive, and indeed 
 necessitate the use of light charges ; since heavy shot-firing weakens 
 the pillars and wastes the coal. " Cutting-machines require broad 
 rooms to obtain the maximum quantity of coal at the cheapest 
 rate, and this results in the reduction in the size of the pillars, 
 with the consequent sacrifice of much coal in the ribs and stumps. 
 Moreover its accumulation in the gob tends to develop spontane- 
 ous combustion. 
 
 Percussion Machines versus Chain Machines. The former 
 class of undercutting machine is especially advantageous over 
 the other types when the roof is so weak as to require timbering 
 close to the face and where the bearing- in layer under the coal 
 is too stony or too full of pyrites to be cut with continuous-motion 
 cutters. It has also an advantage where there is excessive pres- 
 sure from the roof upon the face of coal. It reduces the cost of 
 mining to quite a considerable degree even when the cost of 
 operation, interest, and maintenance are included. It is, however, 
 slow and produces a great deal of slack coal, besides doing less 
 work in a given time than can the breast machines. The 
 latter can undercut in fire-clay or in soft under layer quicker 
 than the punching-machine. In fire-clay the chain machine 
 is preferred, while for soft coal or slate the stronger disc or wheel 
 must be used. 
 
 The breast and longwall machines, employing chains, wheels, 
 or rotating bars, can be operated by electricity or by air, whereas 
 the reciprocating machine must be operated by compressed air. 
 The cost of installing their power plants is about the same for 
 an equipment of equal capacity and a distance of one mile 
 from the power-house. For greater distances electricity ad- 
 mitted is far cheaper than compressed air. In comparing the 
 cost accounts of producing coal by either type of machine not
 
 663 
 
 MANUAL OF MINING. 
 
 only are wages, materials, interest, and depreciation to be con- 
 sidered, but also the profit on the amount of coal lost by each 
 method. 
 
 Comparison of Hand-mining with Machine- work. The 
 amount of slack dirt and waste which must be allowed for in 
 hand-mining is as a rule 50 per cent more than with machine. 
 
 co/v- 
 
 P! G 221. Showing Position of Sullivan Coal-cutter in a Room. 
 
 When the net profit it represents is included the saving of machine 
 coal over hand-mined coal is still greater. The machine-cut 
 groove 6 feet deep saves three cubic feet of coal per yard of face 
 in the reduced slack. A machine cannot be employed to advan- 
 tage in seams as thin as those which can be mined by hand. 
 Neither can the breast or longwall machine be employed in as
 
 CHANNELERS, DRILLS, AND COAL-CUTTERS. 66) 
 
 narrow and confined quarters as can a percussive-machine cut- 
 ter. As to output per year 13,000 tons may be cited as an aver- 
 age production for either class of machine. Hand labor as com- 
 pared with machine is expensive, and the tendency therefore is 
 to confine the former to narrow work and the extraction of coal 
 from the ribs and pillars. At the present time more than one 
 quarter of the soft-coal product of the United States is being 
 machined, with a rapid increase in its adoption with time. When 
 hand labor can no longer be employed profitably the ribs and 
 pillars must be recovered in the second or room working by some 
 other type of machine than those employed at the present time. 
 However rapid the present machine may undercut or shear the 
 coal it is not sufficiently rapid for the extraction of the rib coal 
 in its entirety without considerable loss and risk. The longwall 
 type of machine will probably be called into requisition even 
 though it may differ from the character of the machine employed 
 in the rooms and workings of the mine. In this regard the 
 Sullivan coal-cutter, Fig. 318, meets the requirement, as it may 
 be moved parallel to the face by a feed-chain, as shown in 
 Fig. 321. 
 
 REFERENCES. 
 
 Coal-mining Machines, R. M. Hazelton, Ohio Mine Inspector, 1985, 22; 
 Discussion of Present Form of Machine Cutters, Cyrus Robinson, Amer. 
 Mfr., 1897, 121 ; Efficiency of Modern Mining Machinery, Cyrus Robinson. 
 Amer. Mfr., 1897, 588; Machine Mining and the Labor Question, W. E, 
 Garforth, Coll. Guard., 1897, 480; Coal-cutting by Machine in Iowa, Foster 
 Bain, Coll. Guard., June 1897, 1085; Coal-cutting Machines for Pillar and 
 Stall or Narrow Work John Davis, Coll. Guard., May 14, 1897, 918; Coal- 
 cutting Machinery, Chas. Latham, Coll. Guard., 1897, 133; Notes on Coal- 
 getting by Machinery, T. H. Woodsworth, Fed. Inst. M. E., Vols. VI and 
 VII; Blakemore, Vol. XI, 179; Machine Mining in Ohio, Amer. Mfr., 193. 
 
 Relative Advantages of Coal-cutters, Anon., Coll. Eng., Feb. 1897, 313; 
 Coal-cutting Machines in Longwall, England, T. B. A. Clarke, Coll. Guard., 
 Dec. 1896, 1078; Electricity in Bituminous-coal Mining, Robert M. Hasel- 
 tine, Mine Inspector, Ohio, 1894, 18; Coal-cutting by Machinery, W. Blake- 
 more, Trans, of the N. of Eng. Inst. of M. & M. Eng., Vol. XLV, 177; Coal- 
 cutting Machinery, John B. Atkinson, Engineering Association of the South,
 
 6 70 MANUAL OF MINING. 
 
 Pub. No. 4; Coal-mining Machines, A. Dick, Mineral Industry, Vol. II, 
 230. 
 
 Electrical Coal-cutting, J. T. Burchell, Eng. News, April 6, 1893, 334; 
 Tunnels in Coal-mines: Cost of Driving, With and Without Explosives, 
 M. Elce, Coll. Man., Dec. 1893, 222. 
 
 Diamond Drill Prospecting, J. Parke Channing, Eng. Mag., March 
 1896; Exploring by Diamond Drill: Cost, etc., Archibald Blue, 221; 4th 
 Report, 1893, 164; The Diamond Drill for Deep Boring Compared with 
 other Systems of Boring, Oswald J. Heinrich, Amer. Inst. M. E., Vol. II, 
 241; Cost and Results of Geological Explorations with the Diamond Drill 
 in the Anthracite Regions of Pennsylvania, Lewis A. Riley, Amer. Inst. 
 M. E., Vol. V, 303. 
 
 Machine Mining and the Labor Question, W. E. Garforth, Coll. Guard., 
 1897, 1085; Coal-cutting Machines for Pillar and Stall or Narrow Work, 
 John Davis, Coll. Guard., May 14, 1897, 918; Coal-cutting by Machinery 
 in Iowa, Foster Bain, Coll. Guard., June 1897; Diamond Drill Prospecting, 
 J. Parke Channing, Eng. Mag., March 1896; Exploring by Diamond Drill: 
 Cost, etc., Archibald Blue, Ontario Bureau of Mines, 1895, 221; Diamond 
 Drill Prospecting, Rich. A. Parker, S. of M. Quart., Vol. XVI, 31; Curva- 
 ture of Diamond Drill Holes, J. Parke Channing, E. & M. Jour., Vol. LVIII, 
 268; Machine Lining, M. &M., Vol. XXII, 173; Morgan Gardner Machine, 
 M. &M., Vol. XXII, 237; Longwall Machines, M. & M., Vol. XXIV, 25; 
 Pick and Shaft Machines, M. & M., Vol. XXIII, 32 and 510. 
 
 Coal-cutting by Machinery, Blakesmore, Trans. Fed. Inst., Vol. XI, 179; 
 Coal-cutting Machines, Jeffrey, Horse-power, Coll. Eng., Nov. 1895, 75; 
 Machines in Longwall, England, Coll. Guard., Dec. 1896, 1078; Coal- 
 mining Machines, Mineral Industry, Vol. II, 230; Coal-mining and Methods, 
 Vol. IV; Electric Coal-cutters, Am. Mfr., Jan. i, 1897, 9; Cutters and Drills 
 in Alaska, Coll. Guard. Eng., Feb. 1897, 314; Steam-shovel in Mining, De- 
 scriptives, L. Sup. Mm. Inst., Vol. IV, 1896, 60; Diagram of Forces, Cas- 
 sier's Mag., Jan. 1895, 195; In Lake Superior, Coll. Guard., June 18, 1897, 
 1188; Diamond Drill Prospecting, Channing, Eng. Mag., March 1896; 
 Stanley Header in England, Trans. M. & M. Eng., Vol. XLV, 179.
 
 CHAPTER VIII. 
 
 BLASTING. 
 
 The Principles of Blasting. The principle employed in 
 rupturing rock consists in subjecting the surface of a subfacial 
 cavity regular or irregular to a sudden increase of pressure, acting 
 radially outward. When the agent is sufficiently powerful to 
 produce a high degree of compression upon the surrounding 
 rock it either fractures the material by the formation of a con- 
 geries of crevices or it shatters it. The extent of the destruc- 
 tion depends upon the intensity of the pressure and the cohesion 
 or the toughness of the material. 
 
 Blasting-agents. There are two classes of blasting- agents. 
 The first includes all varied means of causing a prolonged pres- 
 sure of greater or less intensity, while the second develops the 
 force more or less instantaneously. In the first class are mechan- 
 ical devices such as the steady hydraulic pressure obtained from 
 a ram; that of pistons actuated by compressed air; the swelling 
 of slaking lime; and the spreading produced by wedges. The 
 second class comprises various chemical combinations such as 
 black powder, nitroglycerine, ammonite, etc. 
 
 Blasting-agents of the First Class. These are employed 
 for driving preparatory works in the fiery mines and are safer 
 and less expensive than powder. Lime is a safe effective blasting- 
 agent in coal, and can be made up into cartridges, which after 
 insertion into a deep hole are packed tight and moistened. In a 
 short time the lime slakes and expands to break the coal away. 
 Compressed air forced into a cylindrical case by a powerful air- 
 pump is capable of exercising such a pressure as may be sufficient 
 
 671
 
 6 7 2 
 
 MANUAL OF MINING. 
 
 to break the coal. This is employed in some coal regions and 
 is a simple, safe, cheap method. Wedges driven by any of these 
 agents are similarly employed by being inserted in the line of 
 the cleavage of the rock. 
 
 Explosives. An explosive, according to Andre, is a mixture 
 "capable of being suddenly transformed into gases by the appli- 
 cation of heat." In this sudden evolution of gas, in a space 
 formerly occupied by a solid, a pressure is produced upon the 
 confining surface to the volume of the evolved gas to that of the 
 explosive. The expansive force of the gas is greater as the tem- 
 perature of ignition increases and with the rapidity of evolution of 
 the gas. If the latter was instantaneous, the maximum pressure 
 is imparted at the moment of explosion; if the combustion is slow 
 and transmitted from grain to grain, its strength is dissipated 
 over a longer period of time, and the pressure is less. Thus 
 the strength of an explosive is measured by its specific volume, 
 the amount of gas it produces, the temperature, and the rapidity 
 of the evolution. 
 
 Below is a table of the relative volumes and pressures of gas 
 produced by the perfect combustion of i Ib. of explosive occupy- 
 ing about 0.016 cubic foot of volume. 
 
 
 Volume of Gas. 
 
 Heat-units. 
 
 Relative Force 
 
 By Single 
 Explosion. 
 
 ^ By 
 
 Detonation. 
 
 Blasting-powder 
 
 2.38 cu. ft. 
 5-09 " 
 
 II. 01 
 
 10.72 " 
 9.76 " 
 
 900,000 
 570,000 
 1,050,000 
 1,240,000 
 2,375ooo 
 
 I .OO 
 1. 08 
 3-06 
 3-15 
 
 4-55 
 
 4-34 
 3.61 
 6.46 
 6.00 
 
 10.00 
 
 Chloride of nitrogen 
 
 Guncotton 
 
 Picric acid 
 Nitroglycerine 
 
 When an explosive is ignited the heat developed in its com- 
 bustion is rapidly communicated from grain to grain and the 
 particles decomposed with the liberation of gases. According 
 as these two phenomena follow each other slowly or quickly 
 we have rending or shattering powders, of which the representa-
 
 BLASTIXG. 673 
 
 live types are black powder and nitroglycerine. In the first 
 the gas is evolved so slowly as to give time for a concentration 
 of pressure along a line or lines of least resistance. This is the 
 quality desired for the artilleryman or for a sporting powder 
 ability to project. The slow combustion operates upon the 
 small mass of the bullet, which can instantly take up a very high 
 velocity and thereby give a rapidly increasing space for the evolved 
 gases to escape. If the bullet or plug were tight and moved 
 slowly it would burst the breech or muzzle. 
 
 The miner, however, desires to break, and this property is 
 obtained from such agents as rapidly produce gases at high 
 initial temperature and pressure. Very little plugging is needed, 
 for the concussion produced by the gases of the quick powder 
 is practically instantaneous and the wave of pressure extends 
 in all directions, which, being resisted by the rock, spends its 
 force there and shatters it. The more sudden the action the 
 more local is the effect. 
 
 Igniting and Detonating Explosives. Without attempting to 
 follow the history of blasting, for which the student is referred 
 to Rziha's "Lehrbuch dcr Gesammten Tunnelbaukunst," an 
 enumeration of the several simple and compound substances 
 used or suggested at various times to produce concussion may 
 be here given in chronological order: Common black powder, 
 picric acid, guncotton, terchloride of nitrogen, nitroglycerine, 
 and ammonite. This list may seem brief, but a longer list would 
 be merely an enumeration of the varieties obtained by the sub- 
 stitution of a single constituent. We have the artillery, sport- 
 ing, and blasting-powders, composed of charcoal, sulphur, and 
 saltpetre in varying proportions; picric acid and picrates with 
 saltpetre or chlorate of potash; guncotton, combined with other 
 explosives; nitroglycerine, with admixture of absorbents and 
 dilutants. The result is that we have various grades of explosive 
 compounds, from those which may be ignited by heating to a 
 temperature of about 300 C. to the nitroglycerine, which requires 
 a shock. In other words, we have igniting or detonating com- 
 pounds.
 
 674 MAX UAL Ol' MINING. 
 
 Black Powder. The term black powder, or gunpowder, 
 embraces mechanical mixtures of carbon, sulphur, saltpetre 
 varying from 12, 8, and 80 per cent respectively for sporting 
 purposes to 20, 16, and 64 per cent for open-air blasting and 
 11.5, 17.5, 71 for hard-rock underground. It is black or brown 
 according to the per cent of carbon. Its constituents are pul- 
 verized, compressed into a cake, granulated, sifted, glazed, and 
 dried. The size of the grains depends upon the use to which the 
 povcder is to be put. The smaller- grained powders are pre- 
 ferred for mining purposes. The color of the powder varies 
 from black to dark brown, the former having a larger percentage 
 of carbon than the latter. 
 
 Methods of Charging the Hole with Powder. When the hole 
 has been drilled to the required depth, the powder is either poured 
 into it from the can or is inserted as a cartridge. The cartridge 
 is a tube of oiled paper which is closed at the bottom and also 
 at the top after the black powder and fuse have been inserted. 
 
 Black powder is loaded by the barrel method or the needle 
 method. In the latter method, which is more common, the 
 cartridge is inserted into position with a needle projecting from 
 it out to the daylight. Above the powder, around the needle, 
 and filling the hole for some distance is driven a very soft-clay 
 packing material, which is rammed tight into position. The 
 needle is then withdrawn and replaced by a fuse for ignition. 
 The needle is a slender-pointed rod of copper with a ring at the 
 upper end. 
 
 In the barrel method the powder cartridge is pierced by 
 a wire which leads up through a half- inch copper tubing, or 
 "barrel," that extends the entire length of the hole. Around the 
 barrel is tamped the packing, after which the wire is pulled out 
 and replaced by a fuse. A comparison of the methods shows 
 a preference in favor of the barrel because less and poorer quality 
 of tamping may be used; it is twice as fast; the cartridge has 
 less opportunity to soak water; and the cheap barrels are 
 recovered after shooting. Some anthracite miners use a f-inch 
 \vrought-iron pipe for the blasting bale.
 
 BLASTING. 675 
 
 A Blast-hole Loader on the Barrel System is in use for 
 
 loading and tamping large charges of powder without waste or 
 danger of premature explosion. Tubes in 5 -foot lengths are 
 screwed together with a funnel at the top, and a short bronze 
 discharge-tube is at the lower end. A hickory tamping-bar com- 
 pletes the outfit. A system so complete and compact saves time 
 and powder (Fig. 340). 
 
 Tamping and Ignition of the Powder. To confine the large 
 volumes of gases within the small volume occupied by the powder 
 a tight cover of melted clay is rammed on top of the charge and 
 packed tight. The process is known as tamping. 
 
 As a precaution to prevent a premature explosion the mate- 
 rial used must be free of any hard substances which may pro- 
 duce sparks by accidental contact with the tools used in the 
 process. The rammer should not be of metal. A stout hickory 
 bar is usually the tamping-rod. Frequently tamping-bags of 
 stout paper are filled with the tamping material, when the latter 
 is of low quality and contains grit. This prevents contact of the 
 gritty substance with the rock surrounding the hole. Anthracite 
 miners use a half-inch iron bar clubbed at the end to i J inches 
 for a tamper. This has a groove at the side to allow for the 
 needle. 
 
 For the same reason the needle mentioned above must not be 
 of iron or steel, but of copper. Sometimes it is merely a copper- 
 tipped iron rod. In either case it is sharply pointed and when 
 withdrawn leaves a cavity for the insertion of the fuse. Black 
 powder is ignited by intense heat or by a spark or by de-tonation. 
 The last named is produced by the explosion of a small quantity of 
 fulminate of mercury in a cap by conducting to it sufficient amount 
 of heat through a fuse ignited at the surface. The detonator 
 may be dispensed with and the heat from the fuse alone may 
 be sufficient to ignite the powder. The result is not so bene- 
 ficial as when obtained by detonation, as shown by the table 
 on page 672. An electric spark or detonator may also be 
 employed. 
 
 When the hole to be fired is in creviced rock, which is liable
 
 676 MANUAL OF MIX IXC. 
 
 to dissipate the force of the gases, a "bulling-bar" or "clay iron" 
 is used before charging the hole to stuff soft clay into the crevices. 
 
 The Fuse. The fuse, also called a squib, is a thread of 
 powder wrapped in tarred hemp or in cotton and water-proofed 
 outside. When properly made it burns uniformly at a rate 
 definitely known to the miner. One variety, for example, burns 
 at about 20 inches per minute. A knowledge of the rate enables 
 the miner to determine how short a length of fuse may be used 
 and still provide him with safe escape. The fuse is supplied 
 in rolls containing 24 to 40 feet and is cut off in lengths as desired. 
 It is inserted into the cartridge and lowered into the hole with it 
 or it may be used in connection with the needle as previously 
 described. It furnishes a means of communicating the necessary 
 spark and heat for the ignition of the powder and frequently is 
 fitted with a fulminating cap at its end. 
 
 The Cap. A fulminating cap is of copper f inch long 
 and | inch diameter, having a small quantity of fulminate of 
 mercury deposited in it. This chemical explodes with heat with 
 violent fuse and ensures by its detonation the prompt ignition 
 of the powder in the cartridge into which it is inserted. The cap 
 is slipped over the end of the fuse, the top of it being greased 
 with a little cartridge soap. Numerous accidents have happened 
 from careless handling of the caps which contain the violent 
 explosive, and care should be taken that they do not fall into 
 the hands of the careless miner without suitable precaution 
 being exercised. The detonation of the explosive in the cap 
 increases the effect of the black powder fourfold beyond that 
 obtained by a mere ignition with the use of the simple fuse, as 
 may be noticed by the table given above. 
 
 Precautions in the Handling of Black Powder may be sum- 
 marized as follows: 
 
 1. That black powder and high explosives of any kind are 
 not safe to use together in the same hole. 
 
 2. The cap is the only exploder that is safe to use in firing 
 off high explosives, which does away entirely with tamping. 
 
 3. The only tamping necessary on high explosives is a small
 
 BLASTING. 677 
 
 piece of miners' clay, which is easily put down on the charge 
 with a wooden bar. 
 
 4. The practice of picking and boring out missed holes that 
 have been charged with either black powder or high explosives 
 should be strictly prohibited. 
 
 5. Any miner on the mine known to use black powder and high 
 explosives together, thereby necessitating tamping, will be dis- 
 charged. 
 
 The gases produced by the explosion of the fulminating 
 powder have the following percentages: CO 2 , 43; N, 35; CO, 12; 
 H, 6; carbohydrates, 4. 
 
 Black powder is employed for all varieties of rocks and is 
 very serviceable in coal. Its service in the Galena veins and soft 
 rock is better than obtained from the more violent shattering 
 explosives, since it does not pulverize the mineral so finely. 
 
 L^wising. Granite is often quarried by a method known as 
 le wising. Several holes are drilled close together and the parti- 
 tions between them broken down with a flat steel bar, or broach-bit 
 (Fig. 299). This extensive hole fixes the direction of the fracture, 
 which is usually selected as parallel to the "rift" or cleavage. 
 Three drill-holes make a ''complex" lewis-hole. The benefits 
 of this lewising may also be secured by the Knox system, which 
 is meeting favor for dimension work. The hole having been 
 drilled, a reamer cuts V-shaped grooves in its opposite sides to 
 determine the line of break. The tamping is not driven down 
 on the powder, but an air space is left between them. This 
 scheme permits expansion of the gases and gives time to effect 
 iiipture along the plane desired. 
 
 Powder Consumption. The amount of powder consumed 
 for a given work varies too greatly for any general statement. 
 One foot of a i-inch hole is capable of containing 5 ounces of 
 black powder or 38 inches for a pound. The powder, to be 
 economically used, should not fill more than one third of the hole^ 
 In anthracite mines a keg of powder weighing 25 Ibs. is con- 
 sumed for every 40 tons of coal. The bituminous miner breaks 
 on the average 300 tons with a keg. The amount of powder
 
 078 MANUAL OF MIX IXC. 
 
 consumed per ton of coal worked by hand and that by machine 
 is about the same. In longwall mining the consumption of 
 powder is very small. 
 
 Nitroglycerine. Chemically known as trinitroglycerine, or 
 glonoin oil, C B H 5 N 3 O 18 . Made by treating glycerine to nitric and 
 sulphuric acids at a low temperature. The temperature at which 
 it will fire is 360. It is exceedingly sensitive to shock and will 
 explode under the influence of a neighboring blast. It will not 
 take fire when touched by a red-hot body, or if it does, it burns 
 quietly without smoke. If a thin layer of it be struck by a severe 
 blow it will instantly explode. If heated to the temperature of 
 combustion or detonated or even jarred by a wave of concussion 
 its liquid mass will be converted at once into gas and if confined 
 will burst its casement. The process of decomposition will be 
 complete and a perfect combustion of the chemical elements 
 will ensue. A small amount of it explodes by detonation, even 
 if not buried in the rock, but when lying on the face of a boulder 
 in the open air, the surrounding air will not be able to move 
 aside, or to take up the rate of vibration of the shock quickly 
 enough to allow it to expend its energy on the air and the rock 
 underneath will be badly shattered. 
 
 The Manufacture of Nitroglycerine. Several jars of one- 
 gallon size are placed in a cool stream. Into them are poured 
 nitric acid and sulphuric acid in due proportions. A fine stream 
 of pure colorless glycerine is then poured on while stirring the 
 mass. When boiling begins thick red vapors of nitrous gases 
 are emitted. If the glycerine is poured in too fast, or the water is 
 not sufficiently cool, the mixture will take fire with a hissing noise, 
 and if it be not stirred enough the blaze will shoot up several 
 feet. Stirring is continued until the action becomes less intense. 
 Each jar in turn receives a small charge of glycerine, after which 
 the process is repeated. When no more fumes are given off and 
 the addition of glycerine produces no effect, the decomposition is 
 complete and a heavy milky, oily fluid will be noticed on the 
 bottom. The acids are decanted and washed thoroughly with 
 water. Into an old-fashioned wooden churn the nitroglycerine
 
 BLASTING. 679 
 
 is then poured with an abundant supply of water and churned 
 actively until litmus paper shows no acid reaction. Upon the 
 freedom from all acids depends the safety and the permanence of 
 the nitroglycerine. The product being poured into a wooden 
 bucket with some clear water its milkiness will soon disappear and 
 a colorless compound is the result. If after a few days it dis- 
 colors the acid must be removed. A yellow tinge requires the 
 additional churning with cold water; an orange color requires 
 some alkali; but if it becomes cloudy and deeper in tinge the 
 material is valueless and should be either discarded or replaced 
 at once. 
 
 Dynamite. This explosive was discovered in 1864 by 
 : ':'. Nobel and has received universal adoption in a variety of forms 
 for almost all purposes of blasting. The term is a generic one 
 and includes any mixture of nitroglycerine with some inert or 
 mechanical absorbent material. Earth, sawdust, wood pulp, 
 infucorial earth, and magnesia are used as absorbents, in addi- 
 tion to which are added in some cases explosive bases which 
 increase the strength. Nitroglycerine is the active principle of 
 the explosive and its percentage in the compound determines 
 the power of the explosive. Infusorial earth, which absorbs three 
 times its bulk of nitroglycerine, excels all other bases and fur- 
 nishes a very strong explosive. It is usually graded as No. i. 
 The weaker grades of dynamite are designated as No. 2, No. 3, 
 etc. The strength of No. i is about three quarters that of the 
 pure article and six times that of black powder. The powders 
 known as Atlas, dualine, forcite, etc., differ only in the nature 
 of the material used for absorbent. 
 
 The compound is like moist brown sugar in color. It freezes 
 at 46 F., hardening into a white mass. It is almost as sensi- 
 tive to sudden changes of temperature or pressure as nitrogly- 
 cerine. It is perfectly safe to handle and is harmless so long as 
 it does not exude from its absorbent. 
 
 The constituents of some of the high explosives are indicated 
 below:
 
 680 MANUAL OF MINING. 
 
 Tonite is macerated guncotton 52.5 and baryta nitrate 47.5. 
 
 Gelatine is soluble guncotton 2.5 and nitro 97.5. 
 
 Dualine is nitro 50, sawdust 30, and nitre 20. 
 
 Rendrock is nitro 40, parafPne 7, nitre 40, and wood fibre 13. 
 
 Atlas A is nitro 75, tibreless wood 21, nitre 2, and magnesia 2. 
 
 Hercules No. i is nitro 75, chlorate of potash i, nitre 2, sugar 2, and magnesia 20. 
 
 Giant No. 2 is nitro 40, rosin 6, sulphur 6, absorbent 8, and nitre 40. 
 
 Rackarock is nitrobenzol 22.3 and chlorate potash 77.7. 
 
 Vulcanite is mealed gunpowder and nitro in different proportions. 
 
 Experiments upon the relative efficiency of the various ex- 
 plosives have been made under water and are recorded in General 
 Abbot's "Submarine Mines," but no formula? can be prepared 
 from the results because of the varying influences of the foreign 
 substances added to the mail explosive. 
 
 The Storage of Nitroglycerine. Nitroglycerine and its 
 various compounds may be exploded by sudden increase in 
 the temperature and some by a neighboring shock. All are 
 liable to decompose under a low heat. It will exude from its 
 absorbent when hot or exposed to a wet atmosphere. Precau- 
 tion must therefore be taken in the storage and the handling of 
 the material to avoid these sources of danger. They should 
 be stored in large cool dry caves or sheds well provided with 
 ventilating flues and remote from scenes of blasting. They 
 should not be allowed to freeze, for, while in that state they 
 cannot be fired, thawing will be necessary before they can be of 
 service. The thawing of dynamite introduces an element of 
 danger, unless care be taken to avoid a sudden increase in tem- 
 perature or an excess of heat. Dynamite is thawed out by slow 
 heating in a bucket immersed in another. In a jacket is circtK 
 latcd lukewarm water. If the bucket be kept away from the 
 direct fire or heat an excessive heating is impossible. Dynamite 
 which has been stored a long time should be tested by litmus 
 paper for acid reaction, which indicates decomposition, from 
 which spontaneous combustion may ensue. Less accidents 
 occur from the storage of dynamite or nitroglycerine than from 
 the keeping of black powder. 
 
 Being a more powerful mixture than black powder, the ni'ro- 
 glycerine compounds are more economical by reducing the labor
 
 BLASTING. 68 1 
 
 of the miner; requiring less tamping; breaking the mineral 
 finer; requiring smaller holes; consuming less steel and supplies, 
 and materially shortening' the time of mining. As water does 
 not injure it, it js without rival as an explosive in wet ground, 
 and this becomes more manifest as the rock encountered is harder. 
 Methods of Charging with Nitroglycerine. The holes are 
 loaded with nitroglycerine by pouring from a tin cup upon the 
 fuse with its cap and covering the mass with water. Dynamite 
 may be loaded like black powder, the cartridges into which it is 
 made being obtainable of any desired diameter and of a length 
 of about 8 inches each. As many such cartridges may be inserted 
 as necessary. Safety-fuses or electric wire and cap are buried 
 in the uppermost cartridge, which, after placing in position, is 
 slightly tamped with earth and ready for ignition. 
 
 Black powder is often used in combination with giant powder 
 "to start the hole." The proportion is usually one half pound 
 of black powder to three sticks of giant. This practice is often 
 questioned, but the fact that the entire ground is broken to the 
 very bottom of the hole without leaving any collar at the surface 
 is sufficient evidence of its advantage. 
 
 The Secondary Explosion of Powder. The great disadvantage 
 attending the use of any form of explosive, particularly in coal- 
 mines, lies in the production of smoke or gases more or less 
 noxious. The combustion of carbon hydrogen and sulphur in 
 the presence of the oxygen provided by the various nitrates added 
 to the explosive results in the production of carbonic acid, car- 
 bonic oxide, sulphurous acid, nitrous, and other injurious gases, 
 and water, which necessitate plentiful ventilation. When the 
 combustion is perfect and complete all the carbon is converted 
 to carbonic acid and the sulphur to sulphuric acid. But in 
 the deficiency of oxygen resulting from the addition of adul- 
 terants or an imperfect mechanical mixture of the constituents, 
 carbonic oxide is developed as well as smoke, which, with the 
 other injurious and combustible gases, are injected from the hole 
 into the room or stope. Sometimes this is accompanied with 
 a large volume of sparks. The conditions are therefore favorable
 
 682 MANUAL OF MIXIXC. 
 
 for a secondary explosion in the room, which might be further 
 aggravated by the presence of coal-dust or gas, and thus form a 
 nucleus for a bigger explosion after being brought into contact 
 with neighboring volumes of gas. For this reason common 
 black powder, blasting gelatine, and carbonite are prohibited 
 in many coal-mines inclined to be dusty or fiery. Every com- 
 missioner who has examined into the safety and efficiency of 
 blasting-powder has reported that black powders are dangerous 
 and should be prohibited. Some of them are so highly adulter- 
 ated that not only is the combustion incomplete, but the gases 
 evolved by the combustion afterwards dissociate and sometimes 
 recombine, forming dangerous gases that may incite secondary 
 explosions. Nitroglycerine and dynamite, which are chemical 
 compounds capable of complete and instantaneous combustion, are 
 much safer, besides being more economical, since no carbonic oxide 
 is produced. These latter explosives may be employed to great 
 advantage and with less risk even in the dry, dusty, and fiery 
 mines. 
 
 Flameless Explosives. A variety of explosives have been 
 made which produce no carbonic oxide and hence no flame, 
 and which are known as safely explosives, sometimes also called 
 Sprengel explosives. These are composed of a mixture either of 
 two solids, a solid and a liquid, or two liquids, one of which 
 should be a hydrocarbon and the other a compound rich in 
 oxygen, neither at the same time being sensitive to friction. Prin- 
 cipal among these is a group of explosives ammonite, roburite, 
 carbodynamite, and bellite in which the chemical to supply the 
 oxygen is a nitrate, as in black powder. Nitrate of ammonia or 
 of barium was most commonly employed, while the other com- 
 bustible is nitronaphthalene or dinitrobenzol. Owing to the 
 deliquescent nature of ammonium nitrate the two elements are 
 not mixed until ready for immediate use. When so mixed the 
 explosive is dipped in melted wax to keep it dry. 
 
 A detonating wave such as is produced by the fulmination 
 of mercury is alone capable of firing it. When the two elerrents 
 are intimately mixed in proper proportions, the shock of the
 
 BLASTING, 683 
 
 detonation is communicated to the layers of molecules in the 
 immediate proximity of the cap and the powder, whereby the 
 "molecule edifice" is destroyed. This initial force is augmented 
 to the degree corresponding to the heat evolved in the decom- 
 position until the total is consumed. The decomposition of the 
 ammonia nitrate absorbs such an amount of heat from that of 
 the initial force as to reduce the final temperature of the gases 
 developed and projected into the air below the temperature of 
 ignition of fire-damp. Owing to the impossibility of an incom- 
 plete combustion, an excess of available oxygen is added in an 
 excess of nitrate of ammonia, which, absorbing more heat, effectu- 
 ally prevents the formation of carbonic oxide and nitrogen oxides. 
 Changes of temperature do not affect the mixture, freezing for 
 three months does not injure it, and there is no exudation to 
 endanger it. Unlike gunpowder or any nitroglycerine com- 
 pounds, it will not explode by percussion, fire, or electric spark. 
 If struck with a heavy hammer the portion of the explosive 
 directly hit is decomposed by the heat developed by the blow, 
 but the remainder is not affected. If mixed with gunpowder and 
 fired the latter explodes, scattering the former without affecting it. 
 Detonators. These are required to produce a violent con- 
 cussion and an explosion, if otherwise weak explosives, and are 
 of different grades, depending upon the quantity of fulminate 
 of mercury in them. The strongest contains 23 grains of ful- 
 minate and is known as No. i; No. 2 contains 15 grains, and 
 No. 3, 8 grains. The strength of the detonator used increases 
 with the uncertainties of the two constituents of the powder. 
 Ammonite containing between 87 and 89 parts of ammonium 
 nitrate and n to 13 parts of dinitronaphthalene, and bellite 
 which contains 79 to 81 parts of nitrate with 19 to 21 'parts of 
 metadinitrobenzol, require a No. i detonator. Roburite with 
 nearly the same amount of nitrate, but containing the more 
 sensitive chloronaphthalene and dinitrobenzol, and carbonites 
 containing 30 to 36 parts of barium nitrate, 25 to 27 parts of 
 nitroglycerine, and 37 to 43 parts of wood meal, may be fired by 
 a No. 2 detonator. Ardeer powder, containing a little more
 
 6S4 MANUAL OF MINING. 
 
 nitroglycerine than carbonite, with n to 13 parts of kieselguhr, 
 requires for its explosion a No. 3 detonator. 
 
 These detonators add a new element of danger to mining, 
 for in their deflagration a spluttering stream of sparks frequently 
 results, which is liable to ignite any coal-dust which may hap- 
 pen to be present in the room. This is the main objection to 
 the use of flameless explosives, though the latter are safer than 
 the other forms of powder now in use. Perhaps the use of 
 a detonator of still higher power might reduce the danger by 
 producing total combustion, but it is probable also that the in- 
 creased temperature and the heat evolved might be sufficient to 
 ignite the gaseous mixture in the room. 
 
 Safety-explosives cost nearly twice that of black powder, 
 besides producing a greater proportion of coal-dust. 
 
 Smokeless Powders. Powders which are said to be smokeless 
 are made by mixing in suitable proportions guncotton and 
 nitroglycerine, their combinations being effected by the use of 
 camphor and acetones. Their energy is moderated by the addi- 
 tion of inert materials or by increasing the proportion of gun- 
 cotton, though this results in a lower rapidity of explosion. Such 
 powders are handled in small cubical grains and are safe. They 
 are exploded by detonation, not by percussion. 
 
 " Schneiderite " is a powder, light yellow in color, quite oily 
 to the touch, forming lumps readily when pressure is applied. 
 Considered alone " schneiderite " is a wholly inert substance of 
 perfect stability and containing in itself no explosive .substance 
 whatever. The elements of which it is composed only com- 
 bine to form an explosive at the very moment of the explosion 
 under the influence of a detonating primer. 
 
 When the detonator is not used "schneiderite" may be sub- 
 mitted to the most violent shocks with impunity. It is not influ- 
 enced by fire. Thrust into a fire it burns with difficulty. It is 
 also uninfluenced by the most extreme cold. It is sensitive to 
 but one alteration and that only diminishes its explosive qualities; 
 this is the alteration which may result from its hygroscopicity. 
 To avoid the absorption of moisture it is necessary to make sure
 
 BLASTING. 685 
 
 of the imperviousness of the cartridge papers. It is easy to 
 restore all its properties by drying it in a stove or simply in the 
 sun. 
 
 The handling or the transportation of " schneiderite " is not 
 dangerous under any circumstances or under any conditions of 
 preservation. 
 
 The Theory of Blasting. When an explosive is fired, the 
 tension of its gases acts uniformly in all directions upon its encase- 
 ment. A tendency to rupture is developed along the line offer- 
 ing the least resistance to the action of the powder. In a rifle 
 the bullet finds a comparatively easy line of escape along the 
 barrel. This is the line of least resistance. In rock the line of 
 least resistance will be along one of the systems of cleavage planes, 
 in the direction of the shortest distance from the centre of gravity 
 to the surface or along the bore-hole itself. 
 
 It is the purpose of the miner to ascertain the direction of 
 the line of least resistance and to drill the bore-hole to be loaded 
 with powder in a direction approximately at right angles to it. 
 Except in massive rock presenting a plane surface for attack 
 this line of fracture cannot be arbitrarily selected, for the powder 
 will seek its easiest vent for escape from the explosive chamber. 
 For this reason a hole cannot be drilled along the line of least 
 resistance. It becomes merely a cannon; the powder blows out 
 without other result. 
 
 The Line cf Least Resistance. In Drinker's "Explosive 
 Compound" will be found a table showing the relative resistances 
 of different rocks and coefficients representing their toughness. 
 By its use may be determined the quantity of a given grade of 
 powder necessary to produce the rupturing effect. Let W be 
 the weight in ounces of the blasting-agent, L the distance in 
 feet of its centre of gravity from the surface of the working-face, 
 and C the coefficient of the rock resistance, then W=CL 3 . In 
 a given mine the value of C may be determined by a series of 
 trials, the other two variables being altered in amount, employing 
 as little explosive and placing the hole for as long a line L as 
 possible. Thus if a 27-inch hole breaking rock along a line of
 
 686 
 
 MANUAL OF MINING. 
 
 20 inches length shows an average consumption of 5 ounces dyna- 
 mite, then C becomes 1.09. A subsequent hole with a line of 
 least resistance of 30 inches would require 17 ounces of the same 
 explosive. 
 
 The line of least resistance, which is the line of general throw, 
 extends from a point approximately the centre of the charge 
 to the nearest external point, from a to b in Figs. 322 to 325. 
 FIG. 322. FIG. 323. FIG. 324. FIG. 325. 
 
 
 
 Illustrating the Line of Least Resistance. 
 
 In soft rock \vhose cohesion is comparatively slight, the line of 
 least resistance may be a long one nearly that of the depth of 
 the hole (Fig. 322). In tough rock the line ab is short (Fig. 323). 
 The volume of rock broken down is proportional to the weight 
 of explosive used and to the cube of the line of least resistance. 
 With a given depth of hole and of charge in it the length of line 
 is fixed. The" volume of rock broken is determined. To increase 
 the effectiveness of the powder the line must be increased by 
 varying the direction of holes or the amount of powder diminished. 
 A deep hole loaded with a powerful explosive is Jience more 
 economical than shallow holes or low r^ade of Dowder. 
 
 . 
 
 FIG. 326. Shooting in Tight Ground. 
 
 When the line of least resistance exceeds in length that of 
 the hole, as in Fig. 326, no execution is done, other than that
 
 BLASTING. 687 
 
 indicated by the result at K. The inclination of the drill-hole 
 with a flush working-face cannot ordinarily exceed 45 when black 
 powder is used. It may be deflected to 60 if the rock is soft 
 and the explosive powerful. 
 
 Blasting in Creviced Rocks. The cleavage planes of the 
 stratified rocks furnish numerous lines of least resistance which 
 may be advantageously employed as lines of rupture. Wedges 
 or picks are often inserted to split FIG 
 them. Certain coals break freely 
 along these cleavage planes. In 
 some of the massive rocks the 
 rifts and seams are sufficiently 
 pronounced to afford lines of weak 
 
 Drilling in Creviced Rock. 
 
 resistance to rupture. The miner 
 
 avails himself of these faults in the rock and drills the hole normal 
 to them but not down to the seams in Fig. 327. If the centre 
 of gravity of the powder is well below the crevice b, the block 
 above a will be loosened for a considerable distance on either side 
 of the hole. 
 
 In the case illustrated in Fig. 328 the line of least resistance 
 is across to the free face of rock, but there is likewise a tendency 
 to split along the plane b ; the rock will doubtless break downward 
 to the right. In Fig. 261 the shot m has broken to the bedding 
 plane, and somewhat shattered the rock ahead. The shot n 
 and that at m (Fig. 262) will break to the black clay gouge. That 
 at o will probably blow out without any execution. The shot p 
 will blow out a large block if those near o have been properly 
 fired. By firing the holes in order from above downward greater 
 extension is effected, as the bedding planes assist the miner. The 
 shot g in Fig. 263 is well placed to break the rock at the plane 
 of cleavage. With the stratification dipping down from' the 
 face the shots proceed in order upward. 
 
 In coal and soft or brittle metalliferous minerals a weak explo- 
 sive is used to avoid injury to the product. When a thin vein 
 of the latter adjoins a hard rock the usual practice is to blow 
 down the rock with lightly loaded holes. In shafts or tunnels
 
 688 MANUAL OF MINING. 
 
 no attention is paid to the condition in which the mineral is broken 
 down, as rapid progress is sought. For simultaneous firing no 
 heed is paid to the economy of the line of least resistance. Such 
 holes are machine-drilled and are limited in inclination. 
 
 Blasting in Massive Rocks. In homogeneous massive rocks 
 there is no pronounced crevice or seam to assist the miner in the 
 displacement of large masses. The line of least resistance must 
 be created by properly directing the drill to 
 L blow down the maximum possible volume of 
 
 rock with the minimum of drilling. But an 
 additional problem is presented. The miner 
 must consider the manner in which working- 
 m% face will be left after the shot. Each shot in 
 turn becomes a "bearing-in" shot for its suc- 
 cessor (Fig. 329). Without any planes of cleav- 
 FIG. 329. Bearing- a g e t o which to break porphyry, vein quartz, 
 and granites, the direction of the shooting 
 must be to their working-face. This problem is of special diffi- 
 culty when the working-face is the small area of a tunnel or shaft. 
 Such an explosive is known in miners' parlance as " tight ground." 
 Considerable skill is necessary to accomplish much work under 
 such circumstances. The usual practice consists in so breaking 
 the ground as to avoid plane surfaces, but produce irregularities of 
 prone such as are illustrated in Figs. 324 and 330 to 332. Rupture 
 FIG. 330. FIG. 331. FIG. 332. 
 
 Taking Advantage of Irregularities of Face. 
 
 may then be effected along the dotted lines toward the recesses 
 in the rock. Judgment must be exercised in directing the holes;
 
 BLASTING. 
 
 689 
 
 as, for example, hole 2 in Fig. 301 will displace more rock than 
 No. i with equal powder and work. Indeed it may happen 
 that the powder will simply pulverize the rock along the line of 
 break of No. i, dislodging the bowlder alone. So in Fig. 260 
 is illustrated a poorly directed hole. If drilled to k the blast 
 will take place along ks, but only a small charge is needed for the 
 short line. As drilling is an expensive operation it would 
 have been more economical to have drilled it less vertically. 
 Had the operations stopped near m no more powder would be 
 required to blow out at bm than at sk, though difficulty would 
 have been experienced later in blasting out bmks. 
 
 Expanding-bits. A soft or brittle rock may be blown out in 
 large masses by excavating a large chamber at the bottom of a 
 FIG. 333. FIG. 334. 
 
 An Expanding-bit. 
 
 very deep hole and rilling it with black powder, sufficient for the 
 work. The chamber may be produced by an expanding-bit
 
 690 
 
 MANUAL OF MINING. 
 
 (Fig. 333) or by a small charge of powder. When the desired 
 depth of hole has been reached a pair of cutter- wings are forced 
 out, which when revolved cut out a chamber. Powder accom- 
 plishes the same end by shattering the rock. This method is 
 known as squibbing. 
 
 Simultaneous Firing. It has been shown that increased 
 effectiveness is obtained from an explosive by the use of per- 
 fect detonators. A still further gain results from the ignition 
 of the shots simultaneously, by which they materially assist one 
 another in effecting disruption. The concussion produced by 
 the explosive, though progressing radially in every direction 
 from the centre of the charge, succeeds in rupturing the rock 
 only over a limited volume, as is represented in Fig. 335. 
 
 
 
 b 
 
 / 
 
 IF 
 
 x . *' 
 
 '\ >X 
 
 
 -^t 
 
 
 
 T 
 
 
 
 i 
 
 
 i 
 
 1 
 1 . 
 
 
 FIG. 335- 
 
 Either shot (i or /) drilled into the block would break out 
 only a triangular prism. The hole i would break the prism 
 bounded by the parallel faces CAd and eif, or the hole ; would 
 break the prism bounded by the faces ABD and IFJ. The 
 rock is broken toward the free face, but the force expanding 
 laterally and backward into the solid ground is entirely lost, 
 the amount lost increasing with the intensity of the explosion. 
 If, however, several contiguous holes be so placed that their lateral 
 rupturing tendencies will superpose, they will shatter rock which 
 by the single shooting would only have been fractured. The 
 holes i and j would break the mass of rock included between 
 the trapezoids ABGC and IJHE. The latter mass is mani- 
 festly much in excess of the aggregate product of the single holes.
 
 BLASTING. 
 
 691 
 
 The zone of an explosive effect may be divided into three sec- 
 tions a sphere of pulverization immediately surrounding the 
 bottom of the hole, a cone of rupture with its apex near the 
 centre of gravity of the cartridge and its base of an area dependent 
 upon the relative explosive and resisting powers, and a zone of 
 vibrations extending into the rock where the undulations are too 
 weak to produce more than an occasional fracture. By simul- 
 taneous firing, the explosives may be placed at such distances 
 apart as to permit the overlapping of the zones of vibration,, 
 whose cumulative effect may produce rupture. 
 
 The effect of synchronous firing is about 1.4 times as great 
 as that obtained for the same amount of powder fired in consecu- 
 tive shots. Benjamin Frost says in his report on the Hoosac 
 Tunnel that "greater depths of holes are admissible" and greater 
 advance obtained. 
 
 The sole means of obtaining perfect simultaneity is by elec- 
 tricity; fuses will not burn with sufficient regularity to be relied 
 upon. The holes are charged, the cap is inserted, two wires-
 
 6 9 2 
 
 MANUAL OF MINING. 
 
 take the place of fuse, and the tamping is done. The wires are 
 separated, each one being connected with its neighbor in such 
 manner (Fig. 336) that a continuous-wire circuit E extends 
 throughout the series of holes. A is the fuse, B the cartridge, 
 and C the cap. 
 
 The Electric Fuse. The several holes to be fired simultane- 
 ously are charged with an electric fuse, A (Fig. 337), filled v.ith 
 fulminates, B\ the latter is detonated by the 
 incandescence of a fine platinum thread, E, 
 from which two copper wires, C, lead. The 
 fuse-wires are of standard lengths, usually 4 or 
 6 feet for hand-driven holes and 16 or 18 feet 
 for machine- drilled holes. 
 
 The ends of the fuse-wires are cleaned and 
 twisted on the connecting wires D (Fig. 336), 
 and when all these have been connected, the 
 circuit is completed by the final connections with 
 the machine. The leading wire E is usually 
 wound on reels to pay out a supply as needed 
 for the work. The connecting .wires are of com- 
 mon copper wire and no larger in diameter than 
 is necessary to transmit the current freely. By 
 a judicious division of the mine into districts the 
 amount of wiring may be materially reduced. 
 
 In Fig. 339 is shown a section of the magneto 
 machine, which consists of a weak principal 
 magnet, A, with an armature, B, revolving 
 between its poles; said armature is rotated by 
 the rack-and-pinion, C, operated from the 
 handle on the down stroke. As the rack-bar 
 strikes the spring, D, at the bottom, it breaks the continuous cur- 
 rent between the two platinum bearings, E, and causes it to pass 
 through the outside circuit, the fuses, etc. A machine which will 
 fire 16 shots at one time costs $25; one good for 70 can be bought 
 for $55. 
 
 Much wire is buried in the debris of shooting. In a mine 
 
 FIG. 337- The 
 Electric Fuse.
 
 BLASTING. 
 
 '693 
 
 firing regularly 80 shots at a time, the loss from this source is 
 2 Ibs. per day; in a small double- track mine tunnel, 2 ounces 
 of connecting wire and 7 exploders per lineal foot; and in a 
 large railroad tunnel, 0.32 feet of connecting wire, 0.07 feet 
 leading wire, and 0.7 exploder per cubic yard of rock removed. 
 FIG. 338. FIG. 339. 
 
 
 A Magneto Machine and its Mechanism. 
 
 When the fuses have been connected, the last man to retire 
 raises the handle (Fig. 338) and with one plunge ignites afl 
 the shots. An increased efficiency of the explosive is not the 
 sole advantage of this method of firing. It eliminates the numer- 
 ous accidents due to premature explosions or to delayed shots* 
 There is but one electric current and one igniting source. Fail- 
 ing to fire them, there can be no ignition later. It saves the 
 time otherwise wasted by the men, who with their neighbors must 
 retire periodically to shelter. It permits prompt firing from a
 
 694 
 
 MANUAL OF 3//AV.VC. 
 
 FIG. 340. The Cyclone Powder Loader.
 
 BLASTING. 695 
 
 safe distance. It is equally adapted to black powder and to 
 dynamite and to flameless powders. 
 
 REFERENCES. 
 
 Experiments with Safety Explosives, Bergassessor Winkhaus, Trans. 
 M.&M. Eng., Vol. XL VI, 17; Safety Explosives, Bergassessor Winkhaus, 
 Trans. M. &M., Vol. XLV (2), 141; Flameless Explosives, Report of Com- 
 mittee, A. C. Kayll, Coll. Eng., April 1896, 208; List of Flameless Explo- 
 sives, Copy of Orders, Coll. Guard., June 1897, 1129; Report on Flameless 
 Explosives, A. C. Kayll, E. & M. Jour., Vol. LVIII, 556. 
 
 Blasting in Fiery Mines, Franz Brzezowski, Coll. Eng., April 1896, 209; 
 The Effects of Different Explosives on Coal Dust, Winkhaus, Coll. Eng., 
 1896, 39; Explosives for Coal Mines, Vivian B. Lewis, Coll. Eng., Vol. XVI, 
 150; Water Cartridges in Blasting, L. Jarolijmek, Coll. Guard., Vol. LXXI, 
 162; Elements of Defectiveness in Shot -firing, reprint, Coll. Guard., Dec. 
 1894, 1088; Diminished Use of Explosives in Belgian Collieries, Victor 
 Watteyne, Coll. Guard., Aug. 1895, 299; Use and Value of Explosives, 
 C. J. Thompson, Coll. Mgr., Jan. 1896, 15; Testing Explosives for Coal 
 Mines, E. & M. Jour., Vol. LXI, 567. 
 
 Modern Development of Explosives, V. B. Lewis, Coll. Eng., March 
 1895; Nature and Use of Industrial Explosives, J. Daniell, Coll. Guard., 
 March 15, 1895, 497. 
 
 Electric Ignition of Blasting Explosives, Coll. Guard., Oct. 1896, 792; 
 Electric Ignition of Blasting Explosives, B. Heise, Coll. Guard., March 
 189?) 598> Electric Firing, J. von Lauer, Coll. Guard., Jan. 1897, 161; Shot- 
 firing by Electricity, P. Mehers, Coll. Guard., 1894, 241. 
 
 Dangers of Percussion Fuses, Trans. M. & M. Eng., Vol. XLVI, 41; 
 Percussion Fuses and Their Suitability in Fiery Mines, J. von Lauer, Coll. 
 Guard., 1896, 258. 
 
 Detonators, Coll. Eng., Dec. 1895; Explosives and Detonators, J. House, 
 Coll. Guard., March 22, 1895, 546. 
 
 Relative Cost and Efficiency of Powder, Coll. Eng., May 1897, 459; 
 Influence of Diameter of Holes in Blasting, J. Daniell, Coll. Guard., Dec. 
 1894, 1088; Experiments for Ascertaining the Comparative Effect of Ex- 
 plosives, Bergassessor Winkhaus, Coll. Guard., Sept. 1895, 539- 
 
 Dynamite Accidents and Prevention, James Ashworth, Coll. Guard., Nov. 
 1896, 927; High Explosives, W. J- Orsman, Coll. Guard., Jan. 1897, 168; 
 Explosives, W. T- Orsman, Coll. Guard., Nov. 1804, 835; Explosives in Bel- 
 gium, Victor Wattevne, Coll. Gu~rd., 1807, 208; Prevention of Accidents 
 from High Exolosives, James Ashworth, Coll. Guard., Nov. 1896, 927. 
 
 Cost of Mining Items, Calif. Bureau Mineralogy, loth, 852; Coal-
 
 696 MANUAL OF MINING. 
 
 mine Accounts, Min. Industry, Vol. IV, 207; Mine Accounting, Walter M. 
 Jeffrey, M. & M., Oct. 1903; The Colliery Cost Sheets, J. J. Priest, Fed. 
 Inst. M. E., Vol. IX; Mining Problems in Colorado, Percy A. Leonard, 
 Min. Sci. & Pr., Jan. 30, 1904; Mining Errors, F. Danvers, Iron and Coal 
 Trades Rev., March 30, 1900; The Training of a Mining Engineer, R. A. 
 S. Redmayne, Coll. Guard., Dec. 19, 1903; A Practical System of Mine 
 Accounting, E. Jacobs, Eng. Mag., May 1903. 
 
 Regarding the Output of the Mine, M. & M., Vol. XXIII, 51; Ore in 
 Sight, M. & M., Vol. XXIII, 13; The Leasing System, Prof. Arthur Lakes, 
 M. & M., Nov. 1903. 
 
 Why Some Mines are Abandoned, M. & M., Vol. XXIV, 565; Loss 
 Reduction in the Life of a Mine, M. & M., Vol. XXIII, 66. 
 
 What Constitutes a Mine, M. & M., Vol. XXIII, 319; Mining Engi- 
 neering and Valuation of Mines, G. W. Miller, Min. Sci. Pr., April n, 
 1903; Daily Examination of Mines, James Freer, 111. Min. Inst., Vol. I, 183; 
 Notes on Mine Sampling on the Main Reef Series, D. JT Williams, Jour. 
 Chem. Met. Soc. of S. Africa, Oct. 1902; The Valuation of Iron Ores and 
 other Raw Materials, Bernard Osann, Stahl u. Eisen, Oct. i and 15, 1902. 
 
 Anthracite Mine Surveying Practice, Eng. & Min. Jour., Feb. n, 1904. 
 
 Colliery Costs, Coll. Guard., Vol. LXXIX, 1028; Pumping by Com- 
 pressed Air, Coll. Guard., Vol. LXXXI, 571. 
 
 Explosives, Henry Louis, Coll. Guard., Feb. 1897, 302; Use and Abuse 
 of Dynamite, A. Slaght, Bureau of Mines, Ontario, 1895, 285; Manufacture, 
 Use, and Abuse of Dynamite, Harry A. Lee, E. & M. Jour., LXI, 182.
 
 GLOSSARY OF MINING TERMS. 
 
 Absolute pressure. The pressure reckoned from a vacuum. 
 
 Absolute temperature. The temperature reckoned from 461 
 F. or -2 7 3C. 
 
 Adit. A horizontal passage from daylight into a mine, and 
 en or along the vein. It differs from a tunnel in that it is not 
 through country rock across to the vein; it differs from a, drift, 
 which has neither end in daylight. 
 
 Aerophone. A respirator in the form of a tank, receiving 
 the exhalations from the lungs, which contains chemicals designed 
 to revive the air and render it fit for breathing. It is used by 
 rescuers after mine accidents. 
 
 After-damp. An irrespirable gas remaining after an explo- 
 sion of fire-damp. 
 
 Anemometer. An appliance for measuring the velocity of 
 an air-current. 
 
 Anticlinal. A fold of the rock or strata, convex upward. 
 
 Apex. The edge of a vein nearest to the surface. 
 
 Back pressure. The loss, expressed in pounds per square 
 inch, due to getting the steam out of the cylinder after it has done 
 its work. 
 
 Balance-bob. A heavy triangular truss, the long horizontal 
 arm of which supports a weight at one end ballasted to balance 
 the weight of pump-rods at the other. 
 
 Barrier pillars. Pillars, larger than the ordinary, left at 
 stated intervals to prevent the crushing of the roof from extend- 
 ing beyond the section enclosed by them. 
 
 Bar-timbering. A system of supporting a tunnel roof by 
 
 697
 
 6gB MANUAL OF MINING. 
 
 long top bars while the whole lower tunnel- core is taken out, 
 leaving an open space for the masons to run up the arching. 
 Under certain conditions the bars are withdrawn after the masonry 
 is completed, otherwise they are bricked in and not drawn. 
 
 Battery. The lower platform of a coal-chute. A term also- 
 used to define a timber bulkhead in a gallery. 
 
 Bearing-up stop. A partition of brattice or plank that serves 
 to conduct air to a face. 
 
 Bed. A seam of mineral occurring among the stratified rock, 
 
 Bench. The divisions of a coal-bed caused by seams of clay 
 or slate; also used to express the artificial divisions in the process 
 of mining; also a terrace at the outcrop of a seam. 
 
 Black-damp. Carbonic acid gas. 
 
 Blast. The operation of forcing air by blowing. The opera- 
 tion of exploding powder or other agents. 
 
 Blind drijt. A horizontal passage in the mine not yet con- 
 nected with the other workings. 
 
 Blind lead or blind lode. A vein having no outcrop. 
 
 Block coal. Coal that breaks freely into rectangular blocks. 
 
 Blossom rock. The rock detached from a vein, but which 
 has not been transported. 
 
 Blower. A discharge of gas from coal; also a fan for forcing 
 air into a mine. 
 
 Blow-out. The decomposed mineral exposure of a vein. 
 
 Bluff. Blunt. 
 
 Bob. See Balance-bob. 
 
 Brattice. A canvas or plank partition, nailed to posts longi- 
 tudinally, with a level or shaft to divide the same into two com- 
 partments for the purposes of separating two air-currents. 
 
 Break-through. A narrow passage cut through a pillar con- 
 necting rooms. 
 
 Breast. The face of a gallery or heading 
 
 Buggy. A small mine-wagon for conveying coal from face 
 to gangway. 
 
 Bulling-bar. An iron bar used to pound clay into the crevices 
 crossing a bore-hole, which is thus rendered gas-tight.
 
 GLOSSARY OF MINING TERMS. 699 
 
 Bull-pump. A single-acting direct pump consisting of a 
 steam-cylinder placed over the shaft. The steam drives its 
 piston, to which the pump-rods are attached. By means of the 
 piston attached to the rods the water is lifted by the steam pressure. 
 The down stroke is effected by the weight of the pump-rods. 
 
 Bunions. Timbers placed horizontally across a shaft. They 
 serve to brace the wall-plates of the shaft-lining, and also, by 
 means of planks nailed to them, to form separate compartments 
 for hoisting or ladder ways. 
 
 Butt. The end faces of coal. 
 
 Butt-entry. The gallery driven at right angles with the 
 butt-joint. 
 
 Butty. A miner working on contract. 
 
 Cage. An elevator platform. 
 
 Cap. A term signifying the point at which a vein is con- 
 tracted; it is also the rock covering the ore. 
 
 Cartridge. A paper tube filled with an explosive. 
 
 Casing. The tubing of a well-hole to prevent caving. 
 
 Centre. A temporary support, serving at the same time as- 
 a guide to the masons, placed under an arch during the progress 
 of its construction. 
 
 Chain pillar. An untouched block of mineral on either side 
 of the gangway. 
 
 Choke-damp. Carbonic acid gas. 
 
 Chute. An inclined trough or timbered shaft through which 
 ore is delivered by gravity to a receptacle below. See also Shoot. 
 
 Clack. A pump-valve. 
 
 Clearance. The space between the piston at the end of its 
 stroke and the valve face, or the end of the cylinder. 
 
 Cleat. A joint produced by the natural tendency of coal 
 or rock to cleave or split in a certain direction not parallel to 
 the plane of bedding. 
 
 Coal-measures. The strata embracing the carboniferous coals. 
 
 Column- pipe. The line of pipe through which the mine- 
 water is pumped. 
 
 Compression. In steam practice, the action of the piston in
 
 70 MANUAL CF MINING. 
 
 compressing the steam remaining in the cylinder, after the closure 
 of exhaust-valves, into the clearance space. 
 
 Contact vein. A vein lying between two dissimilar rock 
 masses or strata. 
 
 Counter. A cross vein. 
 
 Counterbalance; Counterpoise. A weight used to balance 
 another weight or the vibrating parts of machinery. 
 
 Counter gangway. One which is driven diagonally to the 
 rise until the workings are reached, when it turns off parallel 
 to the main haulageway. 
 
 Country rock. The main rock of the region through which 
 the veins cut, or that surrounding the veins. 
 
 Crab. An iron windlass for moving heavy weights. 
 
 Creep. The crushing of the overlying rock resulting in the 
 floor rising. 
 
 Crib. A framework built like a log-cabin. It may be a 
 mere pillar afterwards filled with rock, or it may be the lining 
 to a mill-hole or shaft. 
 
 Cross-course. An intersecting vein. 
 
 Cross-cut. A horizontal passage driven across the country 
 rock to a vein. 
 
 Cross-heading. A transverse drift which is driven for pur- 
 poses of ventilation from one gangway to another. 
 
 Culm. The fine waste of coal-mines containing dirt as well 
 as coal-dust. 
 
 Curb. A timber frame intended as a support or foundation 
 for the lining of a shaft. 
 
 Cutter. A term employed in speaking of any coal- or rock- 
 cutting machines, the men operating them, or the men engaged 
 in underholing by pick or drill. 
 
 Dam. A bulkhead. 
 
 Day-shift. The gang of miners working during the day- 
 time. 
 
 Dead. The valueless matter of a vein, also gangue, or waste. 
 It is usually made the stowing or filling of an excavated por- 
 tion of the seam, bed, or vein, and then is called gob, waste, or
 
 GLOSSARY OF MINING TERMS. 701 
 
 stull dirt. Sometimes the term is employed in speaking of a 
 sluggish ventilating current. 
 
 Dead quartz. Quartz carrying no mineral. 
 
 Dead work. Exploratory or prospecting work that is not 
 directly productive. 
 
 Deposit. Irregular ore bodies not veins. 
 
 Dip. The angle, measured by the steepest line in the plane 
 of a layer of rock, from the horizon. 
 
 Dirt- fault. A partial replacement of coal in a seam by clay. 
 Not a true fault. 
 
 Drag. The point of union of two veins which meet with- 
 out intersecting. 
 
 Drift. Any subterranean horizontal passage. See Adit. 
 
 Driving. Excavating drifts, adits, or levels. 
 
 Drum. The cylinder on which a hoisting-cable is wound. 
 
 Dumb drift. A gallery which conducts the air around a 
 ventilating furnace to the up-cast shaft. 
 
 Dump. The pile of rock which has been hoisted to the sur- 
 face and deposited there. It may be said to be a low-grade 
 ore reserve. 
 
 Duty. The unit of measure of the work of a pumping-engine 
 expressed in foot-pounds of work obtained from a bushel or 
 100 Ibs. of fuel. 
 
 Dyke. A fissure filled with igneous matter. 
 
 Empty or empties. An unloaded car or the track along which 
 it travels. 
 
 Exploder. A chemical employed for the instantaneous ex- 
 plosion of powder. 
 
 Exploitation. The working of a mine and similar under- 
 takings; the examination instituted for that purpose. 
 
 Eye. The hole in a pick- or hammer-head for receiving the 
 handle. 
 
 Face. The exposure of rock at which work is being per- 
 formed. 
 
 Fathom. A volume of rock equal to six feet square multi- 
 plied by the thickness of the vein.
 
 7 2 MANUAL OF MINING. 
 
 Fault. The dislocation of a vein. The term is also im- 
 properly used in coal-seams. See also Rock-fault or Dirt-fault. 
 
 Fire-damp. A carburetted hydrogen gas, inflammable and 
 specifically lighter than air. 
 
 Fire-setting. The process of exposing very hard rock to 
 intense heat, rendering it thereby easier of breaking down. 
 
 Fissure vein. Any mineralized crevice in the rock of very 
 great depth. 
 
 Float. Broken and transported particles or bowlders of vein, 
 matter. 
 
 Floor. The stratum below a mineral bed. 
 
 Foot-wall. The face of rock below the vein. 
 
 Forepoling. A mode of timbering. 
 
 Free. A term employed in speaking of loose mineral. 
 
 Fuse. A tube, ribbon, or wire filled or saturated with a 
 combustible compound, used for exploding powder. 
 
 Gad. An iron or steel wedge; a chisel-bit pick. 
 
 Gallery. A horizontal passage. 
 
 Gallows- frame. The frame supporting a pulley, over which 
 the hoisting- rope passes to the engine. 
 
 Gangue. The barren portion of the vein. 
 
 Gangway. The principal level of a coal-mine. 
 
 Gash-vein. A mineralized fissure that extends only a short 
 distance vertically. It may be confined to a single stratum of 
 rock, but is a comparatively shallow vein. 
 
 Gateway. A gangway having ventilating doors. 
 
 Gauge pressure. The pressure shown by an ordinary steam- 
 gauge. It is the absolute pressure plus that of the atmosphere. 
 
 Goaf. The excavated space of a coal-mine, usually filled 
 with the valueless portion (gob} of the seam. 
 
 Gobbing-up. Filling with waste. 
 
 Gouge. The layer of clay or decomposed rock which lies 
 along the wall or walls of a vein. It is not always valueless. 
 
 Guide. The timbers nailed to the timbers of a shaft for the 
 purpose of guiding the cage. 
 
 Gunboat. A skip; a self-dumping box used in slopes.
 
 GLOSSARY OF MINING TERMS. 703 
 
 Hanging-wall. The wall of rock above the vein. 
 
 Head-gear. A derrick. 
 
 Heading. A drift or airway. The section of tunnel driven 
 in advance of the lower section or bench. 
 
 Heave. A dislocation of the strata. 
 
 Helve. A handle. 
 
 Hewer. A coal- miner. 
 
 Hitch. A dislocation of a vein. Also, a shoulder or hollow 
 cut in the rock to support one end of a stull or other timber. 
 
 Hogback. See Horse. 
 
 Holing. The picking of a groove in the lower part of a coal- 
 seam for the purpose of facilitating the breaking down of the 
 upper mass. 
 
 Horse. A mass of country rock lying within a vein. Any 
 irregularity cutting out a portion of the vein. See Dirt-fault 
 and Rock- fault. 
 
 H-piece. The portion of a column pipe containing the 
 valves of the pump. 
 
 Incline. An entry into a mine following the dip of the vein 
 or seam. 
 
 Indicator. That which points or directs. 
 
 Some forms show the position of the cave in the shaft. Others 
 record upon paper the pressure of the steam in an engine-cylinder 
 .at various points in the piston-stroke. 
 
 Inplace. A vein or deposit in its original position. 
 
 Intake. The entry which conducts the incoming air-current 
 to the mine. It is synonymous with downcast. 
 
 Jar. That part of the drilling apparatus which takes up 
 the shock of impact of the falling tools upon the bottom of the 
 hole. 
 
 Kibble. An iron ore-bucket. 
 
 Kirving. See Holing. 
 
 Lagging. The slabs or small timber placed between the 
 main timber sets and the roof or walls to prevent small rock 
 from falling into the drift. 
 
 Lath. A plank laid over a framed centre or used in poling.
 
 7 4 MANUAL OF MINING. 
 
 Laundry-box. The box at the surface receiving the water 
 pumped up from below. 
 
 Level. A horizontal passage in a vein-mine, numbered i, 
 2, 3, etc., consecutively from the surface, or from the cross-cut 
 tunnel, down. 
 
 Lift. All the mine workings connected with, opened from, 
 and mined out at one level; also the length of pump-pipe between 
 stations. 
 
 Live quartz. A variety of quartz usually associated with 
 mineral. 
 
 Location. A mining claim. 
 
 Lode. A mineralized fissure. 
 
 Longwall. The system of mining coal without leaving any 
 pillars. 
 
 Man-engine. A mechanical appliance for raising and lower- 
 ing miners. 
 
 Manhole. A hole or an auxiliary shaft through which a 
 man may pass in going from one level to another, into a stope,. 
 or from one ladder to another. 
 
 Measures. A term embracing the strata of a geological 
 series. 
 
 Mill-hole. An auxiliary shaft connecting a stope or other 
 excavation with the level below. 
 
 Mill-run. A test of the value of a quantity of ore as distin- 
 guished from an assay, which tests "pocket specimens." 
 
 Mineral. Any constituent of the earth's crust that has a 
 definite composition. 
 
 Mineral oil. Petroleum or other liquid obtained from the 
 earth. 
 
 Miner's inch. The unit of measurement of water used by 
 the sluice-miners. It is that amount of water hourly discharged 
 through an opening i inch square under a head of several inches. 
 If the head is 7 inches and the hole is through a plank 2 inches 
 thick, a miner's inch is equal to about 90 cubic feet per hour. 
 
 Mining. In its broad sense embraces all that is concerned 
 with the production of minerals and their complete utilization.
 
 GLOSSARY OF MINING TERMS. 705 
 
 Mining retreating. A process of mining by which the vein 
 is untouched until after all the gangways, etc., are driven, when 
 the mineral extraction begins at the boundary and progresses 
 toward the shaft. 
 
 Moil. A short length of steel rod tapered to a point, used 
 for cutting hitches, etc. 
 
 Narrow work. Working places narrower than the rooms, 
 entries, headings, break-throughs, gangways, etc. 
 
 Nitro. A corrupted abbreviation for nitroglycerine or dyna- 
 mite. 
 
 Ore. A mineral of sufficient value (as to quality and quan- 
 tity) which may be mined with profit. 
 
 Ore-shoot. A large and usually rich aggregation of mineral 
 in a vein. Distinguished from pay-streak in that it is a 
 more or less vertical zone or chimney of rich vein matter 
 extending from wall to wall and having a definite width 
 laterally. 
 
 Outcrop. The exposed portion of a vein on the surface. 
 
 Outlet. An exit passage from the mine. 
 
 Output. The product of a mine. 
 
 Panel. The division of a mine which is isolated from neigh- 
 boring districts and provided with distinctive haulage and mining 
 systems. 
 
 Parting. A joint in the rock, or a crevice in a seam, filled 
 with clay or slate ; a switch or turnout to allow loaded and empty 
 cars to pass one another. 
 
 Pay-streak. The thin layer of a vein which contains the 
 pay-ore. 
 
 Pike. A pick. 
 
 Pinch. A contraction in the vein. 
 
 Pipe. An elongated body of mineral. Also the name given 
 to the fossil trunks of trees found in coal-veins. 
 
 Pit. A shaft. 
 
 Pitch. See Dip. 
 
 Pit-man. The shaft-man who attends to the shaft equip- 
 ments, pumps, etc/
 
 706 MANUAL OF MINING. 
 
 Placer. A surface accumulation of mineral in the wash of 
 streams. 
 
 Plane. An inclined tramway for lowering cars by gravity or 
 raising them by means of a stationary engine. 
 
 Plat. A platform. A swinging or revolving door used in- 
 termittently to connect two trackways. 
 
 Plumb (adjective). Vertical. 
 
 Plummet. A string or fine copper wire attached to a heavy 
 weight; used for determining the vertically of shaft-timbering. 
 
 Plunger. The solid piston of a force-pump. 
 
 Pocket. A rich and large body of ore in the vein. 
 
 Poling. The process of timbering by the use of poles, for 
 timbering in soft ground. 
 
 Poll-pick. A combination pick- and hammer-head. 
 
 Poppet; also puppet. A pulley-frame or the head-gear 
 over a shaft. A valve which lifts bodily from its seat instead 
 of being hinged. 
 
 Post and stall. See Pillar and Room. 
 
 Power-drill. A rock-drill employing steam, air, or electricity 
 as a motor. 
 
 Prerelease. The act of discharging steam or air from the 
 cylinder before the piston has reached the end of its stroke. 
 
 Prop. A piece of timber or metal placed normally to the 
 roof or wall for its support. 
 
 Prospect. The name given to underground workings the 
 value of which has not yet been made manifest. A prospect 
 is to a mine what mineral is to ore. 
 
 Prospecting. The process of seeking pay-ore or the prelim- 
 inary operations of a mine. 
 
 Pump. Any mechanism for raising water out of a mine. 
 
 Quick (adjective). Soft, running ground; an ore or pay- 
 streak is said to be quickening when the associated minerals 
 indicate richer mineral ahead. 
 
 Rafter timbering. That in which the timbers appear like 
 roof-rafters. 
 
 Rgacher. A slim prop reaching from one wall to the other.
 
 GLOSSARY OF MINING TERMS. 707 
 
 Reamer. An enlarging tool. 
 
 Reef. The outcrop of a hard vein projecting above the sur- 
 face. Also applied to auriferous quartz lodes. 
 
 Regulator. A sliding- door to apportion the amount of air 
 to be admitted into a section of the mine. 
 
 Rib. A pillar of vein-matter left to support roof or wall. 
 
 Robbing. The taking of mineral from pillars. 
 
 Rock- fault. A disturbed portion of a vein in which coal is 
 replaced by sandstone. 
 
 Roof. The stratum overhead. 
 
 Room. A working place in a flat mine; corresponds to slope 
 in a steep vein. 
 
 Run. A mode of contract work in which steep parts of coal- 
 seams are driven and paid for by the lineal foot or yard of progress. 
 
 Saddle. The ridge, of a stratum or ore-bed. 
 
 Safety-cage. One supplied with safety appliances. 
 
 Safety-lamp. A lamp in which the flame is protected from 
 immediate contact with the surrounding atmosphere. 
 
 Salting. Placing foreign ore in the crevices of a vein or else- 
 where to fraudulently raise its apparent value. 
 
 Samson-post. An upright supporting the working-beam 
 which communicates oscillatory motion to pump or drill-rod. 
 
 Sand- pump. See Sludger. 
 
 Scale. The incrustation deposited in boilers from evapo- 
 rated waters. 
 
 Scraper. A tool for cleaning out drill-holes. 
 
 Seam. A layer of mineral. 
 
 Seed-bag. A water-tight packing of flaxseed around the 
 tube of a drill-hole to prevent the influx into the hole of water 
 from above. 
 
 Selvage. See Gouge. 
 
 Sett or Set. A frame of timber. 
 
 Shaft. A vertical opening from the surface. 
 
 Shaly. Brittle ground. 
 
 Sheave. A grooved wheel over which a rope is turned. 
 
 Shell- pump. See Sludger.
 
 708 MANUAL OF MINING. 
 
 Shelly. Broken ground. 
 
 Shijt. The duration of day's work from six to ten hours. 
 
 Shoot. To break rock by means of explosives. 
 
 Shute. An inclined boardway through which coal is delivered. 
 
 Sill. The floor-piece of a timber sett, or that on which the 
 track rests. 
 
 Slack. Small dirt or coal. 
 
 Slate. Bony coal and hard clay. 
 
 Slickenside. The polished surface of the vein or its walls. 
 
 Slitter. See Pick. 
 
 Slope. An incline. It is an inside slope when it does not 
 extend to the surface. 
 
 Sludger. A cylinder having an upward opening valve at 
 the bottom, which is lowered into a bore-hole to pump out the 
 sludge or fine rock resulting from drillings. 
 
 Smift. A slow-burning fuse. 
 
 Snore. The hole in the lower part of a sinking or Cornish 
 pump through which water enters. 
 
 Sollar. The plank flooring of a gallery covering a gutter- 
 way beneath. Also the platform in a shaft between two ladders. 
 
 Spears. Pump-rods. 
 
 Spilling. A process of timbering through soft ground. 
 
 Spoon. A slender iron rod with a cup- shaped projection at 
 right angles to the rod, used for scraping drillings out of a bore- 
 hole. 
 
 Sprag. A billet of wood used to block the wheels of a car 
 and check its speed. Sprags are permanently used on self- 
 acting inclines. A very steep line requires a sprag in each of 
 the four wheels, while on a moderate pitch only one may be 
 necessary to block a hind wheel. Also, a short prop. 
 
 Square-sett. A variety of timbering for large excavations. 
 
 Squeeze. The closing of a room by the settling of the roof 
 or the rising of the floor. The thinning away of a seam. 
 
 Squib. A slow fuse used for igniting an explosive. 
 
 Stall. The longwall working- face, or a room. 
 
 Station. An excavation adjoining a shaft for receiving the
 
 GLOSSARY OF MINING TERMS. 709 
 
 pump or V balance-bob, or for landing the hoisting convey- 
 ances. 
 
 Stockwerke. A mass of country rock so impregnated by a 
 congeries of veins as that the whole must be mined together. 
 
 Slope. A step. The excavation of a vein in a series of 
 steps. 
 
 Sloping overhand. Mining a stope upward, the flight of 
 steps being reversed. 
 
 Sloping underhand. Mining a stope downward in such a 
 series that presents the appearance of a flight of steps. 
 
 Slowing. The debris of a vein thrown back of a miner to 
 support the roof or hanging wall of an excavation. 
 
 Strike. The bearing of a horizontal line through the middle 
 of a vein. 
 
 String-rods. A line of surface rods connected rigidly for 
 the transmission of power; used for operating small pumps in 
 adjoining shafts from a central station. 
 
 Slull. A stick of timber or platform for supporting miners 
 or vein-waste temporarily or permanently. 
 
 Slull dirt or slull rock. Material supported upon the stulls. 
 
 Stump. A pillar between the gangway and its parallel air- 
 way. 
 
 Slylhe. Carbonic acid gas. 
 
 Sump. The lowest point of the workings, usually a pro- 
 longation of a shaft, into which the mine water is drained and 
 from which it is pumped. 
 
 Swamp. A trough-shaped basin in a coal-mine. 
 
 Synclinal. The depression of a seam or stratum. 
 
 Tail rope. The secondary rope used for balance, which is 
 attached underneath the cages of a hoisting-plant, or at the tail 
 end of the loaded and empty trains &f cars on a slope for rais- 
 ing the empty cars or skips. 
 
 Tamping. The process of making a bore-hole gas-tight by 
 the use of clay. 
 
 Tempering. The act of reheating and properly cooling a 
 bar of metal to any desired degree of elasticity.
 
 710 MANUAL OF MINING. 
 
 Through* or Thirling. A passage cut through a pillar to 
 connect two rooms. 
 
 Throw. The amount of dislocation of a vein. 
 
 Tram. The pair of parallel lines of rails of a trackway. 
 
 Trammer. One who pushes cars along the track. 
 
 Trap. A door used for cutting off a ventilating current, 
 which is occasionally opened for haulage or passage; guarded 
 by a trapper. 
 
 Trend. The course of a vein. 
 
 Tribute. A system of contract mining by which the miner 
 receives his pay out of the gross value of the ore sold, less a 
 certain deduction for royalty to the mine-owner. 
 
 Trolley. A small carriage truck having no body. A trav- 
 eller making connection between two electric wires. 
 
 Tubbing. An iron o wooden cylindrical lining of shafts. 
 
 Tubing. The tube-lining of bore-holes. 
 
 Tunnel. A horizontal passage; properly speaking, one with 
 both ends open to the surface, but is applied to one opening 
 at daylight and extending across the country rock to the vein 
 or mine. 
 
 Underhand work. Picking or drilling downward. 
 
 Underholing. See Holing and Kirving. 
 
 Upcast. An entry through which the air-current rises. 
 
 Upraise. An auxiliary shaft, a mill-hole, carried from one 
 level up toward another. 
 
 Vein. A mineral deposit filling a fissure or crevice. 
 
 Vug. A cavity in the rock. 
 
 Wagon breast. One from which ore or coal can be carried 
 by wagon. 
 
 Wall. The faces of a fissure; the sides of a gallery. 
 
 Wall-plate. The long horizontal stick in a shaft timbering 
 frame which is parallel to the vein. 
 
 Waste. The debris of an excavation; gob; goaf. 
 
 Wedging. The material, moss or wood, used to render the 
 shaft-lining tight. 
 
 Whim. A hoisting appliance consisting of pulley support-
 
 GLOSSARY OF MINING TERMS. 711 
 
 ing the hoisting-rope which is wound on a drum turned by a 
 beam attached to a horse. 
 
 Whip. A hoisting appliance consisting of a pulley support- 
 ing the hoisting-rope to which the horse is directly attached. 
 
 White-damp. The noxious gas called carbonic oxide gas. 
 
 Winch or Windlass. A hoisting-machine consisting of a 
 horizontal drum operated by crank-arm and manual labor. 
 
 Winning. Recovering or mining. 
 
 Winze. A small auxiliary underground shaft sunk from an 
 upper level. 
 
 Wire-drawing. The operation, accidental or otherwise, of 
 reducing the pressure of steam between the boiler and the cyl- 
 inder. 
 
 Working-barrel. The cylinder in which the pump-piston 
 operates. 
 
 Workings. Any underground development from which ore 
 is being extracted.
 
 SIGNALLING. 
 
 The writer would suggest the following code of signalling: 
 The engineer should signal below when he is ready to hoist by 
 raising the bucket or cage a foot or two and lowering again. 
 This is important, particularly for the safety of the man who 
 may be engaged in igniting the blasts near to the place of hoist. 
 The strokes should be made at regular intervals and a halt of 
 a few seconds between the signals. Thus 
 
 5 bells-4 bells would mean send tools to the fourth level. 
 
 1 bell: To hoist or to stop if in motion. 
 
 2 bells: To lower. 
 
 2-1-1 bells: No more hoisting. 
 
 2-2-2 bells: To change buckets from ore to water. 
 
 3 bells: Man on board; lower or hoist slow. 
 
 4 bells: Stop or start the pump. 
 
 2-2 bells: Stop or start the air-compressor. 
 
 5 bells: Send tools down. 
 
 6 bells: Send timbers down. 
 
 USEFUL INFORMATION. 
 
 The area of a circle is 0.7854 (diameter) 2 . 
 
 Ratio of area to circumference is as its radius is to 2. 
 
 An acre is 43,560 square feet. 
 
 A ton contains 2000 Ibs. or 29,166$ troy ounces. 
 
 A troy pound = 0.82 285 7 avoirdupois pound. 
 
 A troy ounce = 43 7. 5 grains. 
 
 An avoirdupois pound = 7000 grains. 
 
 A troy pound = 5 760 grains. 
 
 A cubic foot of gold = $300,000. 
 
 A cubic foot of silver = $10,000. 
 
 A long ton is 2240 Ibs.; a short ton 2000 Ibs. 
 
 712
 
 USEFUL INFORMATION. 713 
 
 A bushel of coke =40 Ibs. 
 
 A bushel of bituminous coal = 76 Ibs. = 2688 cu. in. 
 
 1000 feet (board measure) of dry white pine = 4000 Ibs. 
 
 1000 feet (board measure) of green white pine = 6000 Ibs. 
 
 One cord of seasoned wood =128 cu. ft. 
 
 A mile of track (rails 16 Ibs. per yd.) weighs 25 tons 320 Ibs. 
 
 A mile of track (rails 25 Ibs. per yd.) weighs 39 tons 640 Ibs. 
 
 A mile of track (rails 35 Ibs. per yd.) weighs 55 tons o Ibs. 
 
 A mile of track requires 9 kegs (1780 Ibs.) of 3-in. spikes. 
 
 A mile of track requires 15 kegs (3110 Ibs.) of 4^-in. spikes. 
 
 A mile of track requires 20 kegs (3960 Ibs.) of 4^-in. spikes. 
 
 A mile of track requires 2640 cross-ties (2 ft. apart). 
 
 A mile of track requires 528 splice-joints (2 bars, 4 bolts and 
 nuts per joint), each weighing 5 to 10 Ibs. 
 
 One miner's inch = 21 59 cu. ft. per 24 hours = 0.025 cu - & 
 per sec. 
 
 The pressure of i atmosphere is represented by a column 
 of mercury at 32 F., of a height equal to 0.760 metre = 29.92 
 inches = 14. 701 Ibs. per sq. in., or by a column of water 33.94 
 ft. high. 
 
 i Ib. air at 32 F., 14.7 Ibs. pressure, occupies 12.387 cu. ft. 
 
 i " " 62 F., " " " " 13.141 " 
 
 i " " i8oF., " " " " 16.106 " 
 
 i " " 212 F., " " " " 16.910 " 
 
 i "water at 62 F., " " " " 0.016 " 
 
 i cu. ft. water at 39 F. = 7.49 gallons, weighs 62.425 Ibs. 
 
 i " " " 62F.= " " " 62.355 " 
 
 i gallon " " 62F. = 23i cu. in., " 8.325 " 
 
 i cu. in. " " 62 F.= 0.00434 gal., " 0.036 " 
 
 i ton " " 62 F. = 240.3 gals., occupies 35.9 cu. ft. 
 
 A col. of water, i sq. in. base, 33.947 ft. high = i atmos. 
 
 " " " " i " " 27.7 in. high = i Ib. press. 
 
 " " " " i " " i ft. high = 0.434 Ib. pr. 
 
 " " " mercury, i " " i in. " =0.4914 " 
 
 A bar wro't iron, i " area, i ft. long =3-33 Ibs. wt. 
 
 " " " " i in. diam., i " " =2.618 "
 
 MANUAL OF MINING. 
 
 TABLE OF WEIGHTS OF VARIOUS SUBSTANCES. 
 
 
 Weight 
 per 
 Cu. Ft. 
 
 Cu. Ft. 
 per 
 Long 
 Ton. 
 
 
 Weight per 
 Cu. Ft 
 
 Cu. Ft. per 
 Long Ton. 
 
 Gold 
 
 
 I 87 
 
 Syenite 
 
 
 
 Lead 
 
 7IO 
 
 1. 1C 
 
 Porphyry 
 
 1 66 to 171 
 
 13.5 to 13 i 
 
 Silver 
 
 62 c 
 
 3-4 2 
 
 Slate 
 
 162 to 178 
 
 13.8 to 12 6 
 
 Rolled iron 
 
 480 
 
 4.68 
 
 Quartz 
 
 16? . 2 
 
 i-? 6 
 
 Galena 
 
 468 
 
 4.82 
 
 Sandstone . 
 
 130 to 157 
 
 
 Nickel glance. . . . 
 
 468 
 
 4.82 
 5 60 
 
 Brick 
 Clav 
 
 125 to 135 
 
 18.1 to 16 
 18 7 
 
 Chalcocite 
 Magnetite 
 Specular iron ore. 
 
 355-7 
 338.6 
 3 2 7-4 
 
 6.30 
 6.63 
 6.84 
 
 Anthracite 
 Bituminous 
 Cannel 
 
 85.4 to 99 
 75 to 83 
 
 75 
 
 26.2 tO 22.6 
 
 29.8 to 26.1 
 29.8 
 
 Pyrites 
 Barytes 
 Chalcopyrite 
 
 312 
 277-5 
 
 262 I 
 
 7-os 
 8.07 
 8 cc 
 
 Lignite 
 Oak 
 
 Ash 
 
 78 to 84 
 73 
 
 28.7 to 27 
 
 30.6 
 42 7 
 
 Zinc blende 
 
 2 CQ 
 
 896 
 
 White pine 
 
 
 CQ C. 
 
 
 
 8 06 
 
 
 ^8 7 
 
 57 8 
 
 Limestone 
 
 168 
 
 13-3 
 
 Wood charcoal. . 
 
 25 to 39 
 
 89.61057.4 
 
 EQUIVALENTS OF FRENCH AND ENGLISH MEASURES. 
 
 ABBREVIATIONS. M. = metre; cm. = centimetre; G. = gramme; L. = litre; 
 ft. = foot; lb. = pound; in. = inch; oz. = ounce; dwt. = penny weight; gr. = grain; 
 
 yd. = yard; gal. = gallon; T. = troy; A. = avoirdupois; sec. = second; sq. = square; 
 cu. = cubic; h.-u. = heat-unit. 
 
 M. 
 
 = 3.28 ft. 
 
 atmosphere 
 
 = 0.760 M. 
 
 M. 
 
 = 39.39 in. 
 
 ft. per sec. 
 
 = 0.305 M. per sec. 
 
 ft. 
 
 = 0.3048 M. 
 
 mile per hour 
 
 = 0.447 M - per sec. 
 
 in. 
 
 = 0.0254 M. 
 
 M. per sec. 
 
 = 3.281 ft. per sec. 
 
 yd. 
 
 = 0.9144 M. 
 
 ooo M. per hour 
 
 = 0.621 mile per hour. 
 
 Gunter's chain = 20.1168 M. 
 
 ooo G. per sq. M. = 0.205 Ib. per sq. ft. 
 
 mile 
 
 = 1609.35 M - 
 
 ooo G. per cu. M. = 0.0624 Ib. per cu. ft. 
 
 sq. M. 
 
 = 1.2 sq. yd. 
 
 Ib. per sq. ft. 
 
 = 4883 G. per sq. M. 
 
 sq. yd. 
 
 = 0.836 sq. M. 
 
 Ib. per sq. in. 
 
 = 703077.0 G. per sq. M. 
 
 sq. in. 
 sq. M. 
 
 = 0.00065 S{ 1- M. 
 = 1555.2 sq. in. 
 
 ton per ft. 
 gal. per sq. ft. 
 
 = 3,333.333 G. per M. 
 = 48.905 L. per sq. M. 
 
 acre 
 
 = 4048 sq. M. 
 
 L. per sq. M. 
 
 = 0.0204 gal. per sq. ft. 
 
 cu. in. 
 
 = 0.0000164 cu - M. 
 
 G. per L. 
 
 = 70.116 gr. per gal. 
 
 cu. ft. 
 
 = 0.02832 cu. M. 
 
 Ib. per cu. ft. 
 
 = 16020 G. per cu. M. 
 
 cu. M. 
 
 = 1.31 cu. yd. 
 
 cu. ft. per Ib. 
 
 = 0.0624 cu. M. per 1000 G. 
 
 G. 
 
 = 15-43 gr- 
 
 degree Fahr. 
 
 = 0.5555 de g- centigrade. 
 
 G. 
 
 = 0.0022 Ib. A. 
 
 degree Cent. 
 
 = 1.8 deg. Fahrenheit. 
 
 T.gr. 
 
 = 0.0648 G. 
 
 Ib. per sq. ft. 
 
 = column of mercury 0.00359 
 
 T. Ib. 
 
 = 576ogr. 
 
 
 M. high. 
 
 T. Ib. 
 
 = 373.242 G. 
 
 L. of normal air 
 
 = 19.955 grains. 
 
 A. Ib. 
 
 = 453-593 G. 
 
 G. M. 
 
 = 0.007233 ft.-lb. 
 
 looo G. 
 
 = 2 Ib., 8 oz., 3 dwt., 
 
 ft.-lb. 
 
 = 138.2 G. M. 
 
 
 0.35 gr. T. 
 
 72 ft.-lbs. 
 
 = 106700 G. M. 
 
 1000 G. 
 
 = 2 Ib., 3 oz., 4 dr., 
 
 calorie 
 
 = 3.968 heat-units. 
 
 
 10.473 gr- A. 
 
 heat-unit 
 
 = 0.252 calorie. 
 
 i fluid oz. 
 
 = 0.02957 L. 
 
 thermal unit 
 
 = 0.4536 calorie. 
 
 i quart 
 i gal. 
 
 = 0.9464 L. 
 = 3.78543 L. 
 
 h.-u. per Ib. =0.5555 calorie per 1000 G. 
 calorie per 1000 G. = 1.8 h.-u. per Ib.
 
 INDEX. 
 
 ABSOLUTE pressure, 308 
 
 temperature, 308 
 Access to workings, 19, 69, 249 
 Accidents, causes, 483, 682 
 
 car, 492 
 
 drilling, 492 
 
 statistics, 488 
 Adiabatic curve, 309 
 Adit, 22, 577 
 
 Adjustable doors, 229, 463 
 Aerial tramways, 299 
 Aerophores, 501 
 Affidavit of labor, 15 
 After-damp, 394 
 Age of veins, 6 
 Air-bridges, 45, 459 
 Air, compressed, 307 
 
 compression of, 307 
 
 -current, 455 
 
 -drills, 579, 637 
 
 expansion of, 333 
 
 friction of, 329, 441, 448 
 
 -pumps, 369 
 
 -receivers, 328 
 
 -supply, 96, 447 
 
 -valves, 326 
 
 velocity, 456 
 
 -ways, 407, 442 
 
 weight, 308. 388, 411, 713 
 Alignment of shafts, 22, 510 
 
 of tunnels, 578 
 
 Alternating electric currents, 172 
 American system of tunelling, 583, 
 
 650 
 
 Ammonite. 682 
 Ampere, 67 
 
 Anderson system of tunelling, 594 
 Anemometers, 455 
 Animal haulage, 268 
 Ankylostomiasis, 495 
 Anthracite mining waste, 63 
 Apex, 14, 1 8 
 
 Arches, in drifts, 565 
 
 in tunnels, 392 
 Armature, 170 
 Asphyxiation, 395 
 Assessment, 15 
 Atlas powder, 680 
 Atmosphere, composition, 388 
 Atmospheric pressure, 309, 411, 713 
 
 tension, 308, 411 
 Attle, see Waste. 
 Auger, stem, 599 
 Austrian method of tunnelling, 584 
 
 BAILING-TANK, 338 
 Balance-bob, 289, 352 
 Balancing the resistance, 448 
 Barometerrelationofexplosions,396 
 Barrier pillar, 447 
 Bar timbering, 587 
 Battery dam, 56, 567 
 
 for blasting, 691 
 Beard-Mackie lamps, 399, 474 
 Bearing in, 42, 620, 655, 686 
 Bedded vein, 4, 7, 37 
 Behr's dumping device, 248 
 Belgian lamp, 476 
 
 tunnel system, 5:84 
 Bellite, 436 
 
 Bending ropes, 152, 201, 291 
 Bit, convex diamond, 607 
 
 of percussion drill, 614, 625, 644 
 
 of rotary drill, 619, 625 
 Bituminous coal-mining, 3 7, 387, 658 
 Black damp, 384 
 Black powder, 674 
 Blacksmithing, 629 
 Blanket vein, 17-20 
 Blasting, 670 
 
 agents, 496 
 
 off the solid, 620 
 
 with electricity, 277, 435, 691 
 
 with lime, 671 
 
 715
 
 7 i6 
 
 INDEX. 
 
 Bleichert tramway, 304 
 Blockholing, 624 
 Blocking out the mine, 24, 70, 76 
 Block system of mining, 586 
 Blowers of gas, 100 
 Blowing- fans, 100, 391, 421 
 Bob, 348 
 Boiler, 88 
 
 evaporation, 89 
 
 scale, 92 
 
 sectional, 99 
 Bonneted lamps, 476 
 Bore-holes, 597 
 
 for prospecting, 597, 613 
 Boring methods, 537 
 Brain's system of drilling, 653 
 Brattice, 457 
 
 Break-through, 47, 57, 465 
 Breasts, 56, 64, 70 
 
 drill, 618, 661 
 Brine evaporation, 29 
 Broach-bit, 636 
 Brown's panel system, 61 
 Buggy roads, 56 
 Bull pump, 347 
 
 wheel, 599 
 Buntons, 513 
 Burleigh air-compressors, 326 
 
 drills, 642 
 Butt headings, 47, 59 
 
 CABLES, 289 
 
 Cage, safety, 213, 226 
 
 Calculating depth of engine service, 
 
 hauling capacity, 158, 276, 280, 
 282 
 
 power transmission by air, 212, 
 256, 312, 315, 328, 334 
 
 pump capacity, 373 
 
 size of engine, 158 
 
 ventilating power, 412, 429 
 
 work of compressing air, 312 
 Calorific of value fuel, 96 
 Calumet and Hecla mine, 498 
 Calix drill, 614 
 Cameron pump, 207 
 Camphausen system, 149 
 Candles, 468 
 Caps, fulminating, 676 
 
 in lamp, 398 
 
 timber, 552 
 Carbonic acid, 388 
 Carbonic oxide, 389, 395 
 Carbonite, 682 
 Carriage, 225 
 Cars, 231, 238, 242 
 Car- wheels, 153 
 Cataract engine, 348 
 
 Caving system, 65, 79 
 Cementing shafts, 513, 535 
 Centre-cut system of drilling, 509, 
 
 650 
 
 Centre props, 552 
 Centrifugal pump, 32, 377 
 
 ventilators, 429 
 Chain pillars, 58 
 Chairs, 230 
 Channelers, 635 
 Chapman drill, 615 
 Chimney draft, 100, 413 
 Chimney, Evasee, 426 
 Chocks, 550 
 Chute, 55, 78, 82 
 Circular mils, 169 
 Claim, mining, 14-16 
 Clanny lamp, 455, 472 
 Clearance, 113, 327 
 Cleat, 27, 407, 520, 687 
 Cleavage, 27, 52 
 Clifford lamp, 473 
 Clip-pulleys, 288, 302 
 Closed running fans, 425 
 Clutches, friction, 138, 228 
 Coal-beds, 38 
 
 -cutters, 655 
 
 -dust, 493 
 Coal in lead veins, 8 
 
 mined per fatality, 306 
 
 mining, 36 
 
 storage, 99 
 
 strength of, 51 
 Column pipe, 359 
 Combustion of explosives, 426 
 
 of fuel, 88-96 
 Compartments, 515 
 Composition of explosives, 674 
 Compound cylinders, 108 
 Compressed air, 275, 408, 424 
 efficiency of, 335 
 loss, 328 
 transmission of power by, 
 
 334 
 
 Comstock mine, 522-567 
 Condensers, 108-112 
 Conductors, 175-177 
 Conical drum, 128 
 
 hoist, 143 
 Consumption of air, 96, 447 
 
 of fuel, 98 
 
 of powder, 677 
 
 of steam, 117, 355 
 
 of timber, 546 
 Continuous system of drilling, 509, 
 
 654 
 
 Cooke fan, 422 
 Cooling air, 317 
 Converters, 179
 
 INDEX. 
 
 717 
 
 Core, 508 
 
 drill, 606 
 
 lifter, 608 
 
 Corliss engine, 106, 114 
 Cornish pump, 348 
 Counterbalance, 146 
 Counterpoise, Koepe, 146 
 County of Durham system, 63 
 Cradle-dump, 248 
 Creep, 390 
 Cribbing, 525, 572 
 Cross-cut, 22 
 Cross-overs, 244 
 Culm, 68 
 
 Cummings' system, 335 
 Curb, 526 
 Current, alternating, 169 
 
 capacity, 167 
 
 continuous, 169 
 
 direct, 169 
 Curvature, 264 
 
 radius of, 201 
 
 Cushier's system of pumping, 355 
 Cut-off, 104, 118 
 
 automatic, no 
 Cyclone prospector, 606 
 
 DAMPS, 391, 394, 477, 493 
 
 Darlington drill, 394, 470, 643 
 
 Davy lamp, 471 
 
 Dead- work, 24 
 
 Deane pump, 363 
 
 Death-rate, 485 
 
 De Laval wheel, 119 
 
 Depression produced by fans, 439 
 
 Depth of holes, 626 
 
 of mine, limiting, 504 
 
 of shafts, 505 
 Derrick, 129, 210 
 Designing hoisters, 150 
 Detaching-hook, 217 
 Detonators, 673, 683 
 Diamond drill, 606 
 Dick lamp, 473 
 Diesel motor, 124, 370 
 Diffusion of gas, 397 
 Dimension-stone, 29, 623, 635 
 Discipline in mines, 487 
 Discovery, 1 5 
 Ditches, 192 
 Divining-rod, 8 
 Doble water-wheel, 190 
 Dog-holes, 465 
 Doors, 229, 462 
 Double entry, 19, 58 
 
 hand-work, 627 
 Draught, 99, 402, 410 
 Drag of mine, 403 
 Draw-bar pull, 270 
 
 Drifts, 26, 258, 571 
 Drilling, 623 
 
 by Diamond drill, 606 
 
 by hand-auger, 619 
 
 by power-drill, 637 
 
 by spring-pole, 597 
 Drill-hole deflections, 6n 
 Drinker, H. W., 583 
 Drums, 130-142, 158 
 Dualine, 680 
 Dumb channel, 402 
 Dump, 248 
 Durability, rope, 203 
 Dust explosions, 494 
 Duty, 365 
 Dynamite, 679 
 Dynamo, 171 
 
 ECONOMIZERS, 92, 93 
 Electric coal-cutter, 664 
 
 drill, 619, 655 
 
 firing, 650, 691 
 
 fuse, 692 
 
 hoister, 162 
 
 lamps, 481 
 
 locomotive, 277 
 
 motive force, 167 
 
 motor, 182, 381 
 
 pump, 375 
 
 signalling, 214 
 
 symbols, 174 
 
 transmission of power, 1 78 
 
 units, 167 
 
 wires, 164, 174 
 
 Elevation of the outer rail, 264 
 Endless cable, 289, 298 
 
 end lines, 16 
 End on, 52 
 
 Engine capacity, 154, 158, 276, 280, 
 282 
 
 foundation, 117 
 
 haulage, 290 
 
 horse-power of, 88, 115, 312, 373 
 
 plane, 283 
 
 plant, 87 
 
 second motion, 133 
 
 speed, 114 
 
 Engines, classifications, 102, 108 
 English system of tunnelling, 583 
 
 and French measures, 714 
 Entry, double, 19, 69 
 
 ventilation, 405, 406 
 Equivalent orifice, 403, 430 
 Exhaust fan, 423 
 Expansion-bits, 639 
 
 -joints, 383 
 
 of air, 334 
 
 of steam, 102 
 Exploitation, 34
 
 7 i8 
 
 INDEX. 
 
 Exploratory work, 10 
 Explosions, 499 
 
 force of, 395 
 
 barometric relations, 396 
 guidance after, 475 
 precautions against, 474 
 Explosive, compressed air, 424 
 
 currents, 477 
 
 definition of, 672 
 
 gas, 393 
 Explosives, 671 
 
 accidents with, 67, 492 
 
 flame from, 682 
 
 gases from, 492, 68 1 
 
 storage of, 680 
 Extinguishing fires, 497 
 Extra-lateral rights, 13 
 Eye, see Pick. 
 
 FABRY fan, 422 
 Face on, 52 
 Face-entry, 59 
 Fahrkunst, 253 
 Fainting in shaft, 310 
 Falls of ground, 489 
 
 of roof, 43, 489, 547 
 Fan, chimney, 100 
 
 closed- running, 425 
 
 design, 435 
 
 efficiency, 434 
 
 Guibal, 429 
 
 relation between mine, 432 
 
 theory, 429 
 
 velocity, 427 
 
 ventilation, 420 
 Fault, 28 
 
 Federal mining laws, 14 
 Feed-water, 89 
 Ferranti electric system, 97 
 Fiery mines, 447 
 Filling system. 65, 82 
 Fire, causes of, 495 
 Fire-damp detection, 391,471,477, 
 
 49.3 
 
 Fire-setting, 616 
 Firing, barrel system, 428 
 
 electric, 650, 691 
 Firing, hand, 97 
 
 -needle, 427 
 
 First-motion engine, 77, 430 
 Fissure veins, 5 
 Flameless explosive, 682 
 Flash-point, 240 
 Flat rope, 198 
 
 Fleuss diving apparatus, 501 
 Float, ii 
 Flue-gases, 95 
 Flushing. 68 
 Foot- wall, 78 
 
 Force-pump, 348 
 Forcite, 432 
 Forepoling, 533 
 Forfeiture, 15 
 Forge, 630 
 Framing-machines, 574 
 
 -tools, 572 
 Free air, 303 
 Freezing system, 543 
 French and English measures, 714 
 Frequency, 173 
 
 Friction, coefficient of, 193, 329, 
 441 
 
 -gear, 136 
 
 -hoisters, 136 
 
 of air, 329 
 
 of cars, 238 
 
 of water, 189 
 Fuel consumption, 88-99 
 
 -valve, 96 
 Fulminate, 676 
 Furnace ventilation, 418 
 Furnaces, 417 
 Fuse, electric, 676, 692 
 Fusee, 76, 676 
 
 GAD, 624 
 
 Galleries, dimensions of, 26, 560 
 
 Gangue, 4 
 
 Gangway, 26, 258, 560 
 
 centre props, 555 
 Gas effects upon life or flame, 392 
 
 occlusion, 391 
 Gas-engine, 124 
 Gaseous mines, 61, 387 
 Gases, 386 
 
 explosive, 393 
 
 from explosions, 394, 493 
 
 from powder combustion, 68 1 
 Gauge pressure, 308 
 
 of track, 262 
 Geological maps, n 
 
 theories, 5-8 
 
 German system of tunnelling, 584 
 Giant powder, 680 
 
 nozzles, 31, 377 
 
 duty of, 32 
 Glossary, 696 
 Goaf, 401 , 454 
 Gob, 40, 66 
 
 road, 44. 389, 557 
 Governors, 109 
 Grade of drift, 257 
 Grate area. 89 
 Grip drill, 618 
 Guibal fan. 423 
 Guides. 222 
 Gunboats, 224 
 Guncotton, 673
 
 INDEX. 
 
 719 
 
 HAASE'S system, 542 
 Half-end on, 52 
 Hallidie tramway, 300 
 Hammer, 627 
 Hardening steel, 631 
 Harrison coal-cutter, 659 
 Haulage, 257, 268, 276, 284 
 
 -ways, 45- 457 
 Heading, 259 
 Head-gear, 209 
 Heater, 93 
 
 Hepplewite-Gray lamp, 399, 473 
 Hoisting conveyances, 196 
 
 -engine, 127-133 
 
 -rope, 196 
 
 speed, 512 
 
 water, 338 
 Holing, 612 
 Hollenback shaft, 523 
 Hooks, 217 
 Horse-power, 88, 115, 1 68, 321-328, 
 
 382 
 
 Housing of plant, 18, 90, 210, 431 
 H -piece, 340 
 Hudson tramway, 301 
 Hydraulic engine, 372 
 
 feed, 610 
 
 mining, 30, 382 
 
 ram, 338 
 
 shield, 592 
 
 wedge, 623 
 
 Hydrogen, sulphuretted, 289 
 Hydrometer, 240 
 Hyperian logarithms, 314 
 
 ILLUMINATION, 468, 469 
 Inclined planes, 284 
 Inclines, 24, 69 
 Indicated horse-power, 115 
 Indicators, 103, 214, 320, 361 
 Induction motor, 183 
 Ingersoll coal-cutters, 66 1 
 
 compressors, 324 
 Internal gear, 136 
 Intersecting veins, 6, 16, 20 
 Inysutse wheels, 187 
 Iron in tunnels, 592 
 
 lining, 562 
 
 -mines, 36, 72, 535 
 Isothermal compression, 310312 
 
 JARS, 598 
 Jaws of rooms, 48 
 Jeffrey coal-cutter, 66 1 
 Joints of timber, 514 
 Juggler-way, 64 
 Jumper, 625 
 
 KEEPS, 230 
 
 Keystone prospecting-machine, 606 
 
 Kick back, 244 
 
 Kicking down a hole, 599 
 
 Kind-Chaudjon process, 539 
 
 Kirving, see Underholing. 
 
 Knight wheel, 188 
 
 Koepe's system of winding, 146 
 
 Kutter's formula, 192 
 
 LADDERS, 249 
 
 Lagging, 561 
 
 Lamps, 394, 398, 399, 471, 477, 479 
 
 Landings, 229, 265, 524, 573 
 
 Latches, 236 
 
 Laths, 533, 589 
 
 Laws, State mining, 14, 210 
 
 U. S. mining, 14 
 Lemiele fan, 423 
 Levels, distance between, 25, 71, 
 
 258, 560, 577 
 Lewising, 637, 676 
 Lift, height of, 25, 71 
 
 -pumps, 345 
 Limit of mining, 504 
 Line of least resistance, 686 
 Lippman's drill, 542 
 Locked-wire rope, 193 
 Locomotives, 277 
 Lode, definition of, 5, 9, 13, 18 
 Logarithms, hyperbolic, 314 
 Long-hole process, 509, 654 
 Long horn, 52, 47 
 Long tunnels, 581 
 Longwall, 664 
 Lubricating-oil, 238 
 
 MACHINE vs. hand-work, 509, 648, 
 
 659 
 
 Magic wand, 8-13 
 Magneto machine, 692 
 Man-engine, 252 
 Manganese ores, 249 
 Man-hole, see Mill-hole, 249 
 Manometric depression, 425 
 Manual haulage, 267 
 Mapping, n, 498 
 Marsault lamp, 472 
 Marsh-gas, 390 
 Masonry in mines, 529, 563 
 Mather and Platt boring system, 
 
 60 1 
 
 Mean effective pressure, 105, 118 
 Mechanical ventilators, 420 
 Mercury veins, 4 
 Meyer valve, in, 322 
 Mica, 4 
 
 Mil, circular, 169 
 Mill-hole, 36, 74, 551, 574
 
 7 20 
 
 INDEX. 
 
 Mine buildings, 18, 90, 210, 431 
 Mine, definition of, 3 
 Miner's inches, 382, 714 
 
 tools, 616 
 Mining, economy of, 46, 54 
 
 in soft ore, 75 
 
 in thick seams, 61 
 
 .n thick veins, 75 
 
 in thin seams, 38 
 
 in thin veins, 38 
 
 laws, 14 
 
 limit, 504 
 
 terms, 696 
 
 claim, 14-16 
 
 retreating, 38 
 Moil, 624 
 
 Moment of an engine, 155 
 Moss-box, 541, 604 
 Motive column, 413, 441 
 Motor, electric, 182 
 Mueseler lamp, 472 
 Multiple circuits, 457 
 
 NARROW work, 57 
 Native metal, 2 
 
 ventilation, 409 
 Nickel ores. 4 
 Nitrogen, 388 
 Nitroglycerine, 678 
 
 storage of, 427 
 Norwalk air-compressor, 325 
 Nottingham system, 47 
 
 OCCLUSION of gas, 391 
 Ohm, 1 68 
 Oil-engine, 124 
 Oil, illuminating-, 469 
 
 lamp, 468 
 
 lubricating-, 238 
 
 prospecting for, 601 
 Oil-well rig, 598 
 Oilers, self-, 238 
 Open cut, 34 
 
 -running fans, 424 
 Ore deposition, 4 
 
 bucket, 218, 308 
 definition of, 3 
 
 -shoot, 7 
 
 Outbursts of gas, 391 
 Outcrop of coal, 10 
 
 of veins, 10, 14, 18 
 Outlet, single, 507 
 Overcast, 460 
 Overhand method, 7073 
 Overloaded engine, 62 
 Overwinding, 216-492 
 Oxygen, 388 
 
 PACK-WALL, 40 
 
 Panel system, Brown's, 61, 447 
 
 Parson's steam-turbine, 119 
 
 Patenting a claim, 16 
 
 Pay streak, 6-25 
 
 Peat-mining, 31 
 
 Pelton wheel, 190 
 
 Pentice, 510 
 
 Petroleum, monograph on, 12 
 
 Phosphate rock, 4 
 
 Pick, 621 
 
 Pickets, spilling, 390 
 
 Pieler gas-detector, 473 
 
 Pike, see Pick, 
 
 Pillar and room, 38-48 
 
 Pillars, robbing of, 57, 65, 392 
 
 sizes of, 27-53, 73, 447, 531 
 
 waste in recovering, 42 
 Pinch, 6-26 
 
 Pipe sizes, 190,333,342,359 
 Pipes, iron, 194, 333, 359 
 Piston speed, 367 
 
 pump, 354 
 Planimeter, 106 
 Pneumatic system, 537 
 Poetsch sinking system, 543, 577 
 Pole-pick, 621 
 Poling, 588 
 Post drill, 618 
 Powder, 495, 671 
 
 accidents with, 681 
 
 consumption of, 677 
 Power plants, 87 
 
 drills, 641 
 
 transmission, 101, 178, 296, 335 
 
 -driven pumps, 374 
 Preparatory work, 18 
 Pressure of air, 308 
 Prop, 491, 548, 668 
 Prospecting by boring, 10, 504, 
 
 597 
 
 by wand, 813 
 
 by witchery, 13 
 
 in massive rock, 10 
 
 in stratified rock, n 
 
 surface, 10 
 Pumps, 336 
 
 centrifugal, 377 
 Pump-bob, 352 
 Pump, Bull, 347 
 
 Cook, 345 
 
 Cornish, 348 
 
 duty of, 365 
 
 electric, 375 
 
 hydraulic, 363 
 
 rod, 348 
 
 rotary, 379 
 
 sinking, 356 
 
 -valve, 343, 355 
 Pulsometer, 383
 
 INDEX. 
 
 721 
 
 Punch-drills, 597 
 
 8UARRY, 29, 35, 635 
 uarrying dimension-stone, 29, 
 623-636 
 
 RAILS, 261 
 
 Ram, 636 
 
 Ramsey caging, 244 
 
 Rand compressors, 323 
 
 drills, 639 
 Rateau fan, 425 
 Recorder, 428 
 Reels, 144-158 
 Regulator-doors, 463 
 Reheating air, 334 
 Reserves, 26 
 
 Resistance, 402, 419, 445, 448 
 Retreating, mining, 44 
 Reversible engine, 112 
 Riedler pump, 375 
 Rifled holes, 645 
 Rheostat, 184, 279 
 Rivet, pipe-, 340 
 Robbing pillars, 57, 65, 392 
 Rock-chute mining, 63 
 Rock pressure, 49 
 
 -shaft, 82 
 Room, 54 
 Root-blower, 422 
 Rope, 152 
 
 -drilling, 196 
 
 flat, 284 
 
 haulage, 284 
 
 preservation, 203 
 
 round, 201 
 
 strength of, 153, 199, 297 
 Rotary bar-cutter, 660 
 
 effort, 118 
 
 converters, 177 
 
 pumps, 397 
 Rupturing effect of explosives, 673 
 
 SAFETY appliances, 265 
 
 -boiler, 91 
 
 -cage, 226 
 
 -catch, 218 
 
 -doors, 464 
 
 explosives, 682 
 
 -lamp, 471 
 Sag of rope, 298 
 Salt-mining, 31 
 Sand-pump, 602 
 Scale in boiler, 92 
 Schiele fan, 230 
 Schmidt's rule, 28 
 Scotch system, 52 
 Schram's drill, 640 
 Second-motion engines, 33 
 
 Self-acting plane, 55, 282 
 Self-dumping cages, 231 
 Self-recorder for fans, 428 
 Selvage, 5 
 
 Sergeant drill, 641, 660 
 Shaft, 504 
 
 auxiliary, 580 
 
 -bottoms, 531 
 
 compartment, 507-515 
 
 -pillars, 27, 42, 73, 531 
 
 rectangular, 507 
 
 round, 530 
 
 shape of, 530 
 
 sinking, 22, 513, 543 
 
 site of, 20, 69, 505 
 
 size of, 23, 512 
 
 -timbering, 512 
 Shaw's gas-detector, 399 
 Shearing coal, 663 
 Sheave, 210, 290, 296 
 Shode, ii . 
 Shoot ore, 7 
 Shorthorn, 52 
 Shovels, 621 
 Shutter on fans, 427 
 Signals, 213 
 Sills, 574 
 
 Simultaneous firing, 650, 691 
 Single chute, 64 
 
 entry, 504 
 
 -handwork, 627 
 Sinking, continuous process, 509,. 
 
 . 6 54 . 
 
 in running ground, 534 
 
 Kind-Chaudron process, 537 
 
 Poetsch's system, 543 
 
 pump, 356 
 Siphon, 384 
 Skips, 223 
 Slitter, see Pick. 
 Slope, 579 
 
 -cage, 228 
 
 -carriage, 227 
 
 openings, 21 
 
 railway, 284 
 Slow powder, 673 
 Sludger, 539, 605 
 Smokeless powder, 684 
 Snap-hooks, 206 
 Soft ground, drifting in, 611 
 
 timbering in, 566 
 Spades, 406 
 Speed of drilling, 610, 626, 648- 
 
 of haulage, 268, 287, 292 
 
 of hoisting, 151 
 
 of pumping, 364-367 
 
 of ventilating current, 443, 455; 
 Spilling, see Poling. 
 Splicing rope, 108
 
 7 22 
 
 IXDEX. 
 
 Splitting air-current, 447 
 
 Spontaneous combustion, 497 
 
 Spoon, 627 
 
 Sprag, 43, 258 
 
 Spring-hole drilling, 2, 599 
 
 Spudding, 599 
 
 Square sets, 75, 518, 568 
 
 work, 67 
 Squeeze, 390 
 Squib, 617, 672, 692 
 Statutory provisions, 486 
 Stables, underground, 269 
 Stalls, 40 
 Steam as extinguisher, 102 
 
 -boiler, 88 
 
 coal; 97 
 
 condenser, 108 
 
 expansion, 105 
 
 -jet ventilation, 100 
 
 pressure, 117, 102 
 
 -pump, 361 
 
 -shovel 30, 35 
 
 -turbine, 119 
 Steel, definition, 631 
 Stent, 627 
 
 Stephenson lamp, 472 
 Stockwerke, 4-6 
 Stokers, 98 
 Stone shaft, 83 
 Stope, 39, 70, 74,459 
 
 height of, 25, 71 
 Storage of powder, 680 
 
 battery, 166-281 
 Stowing, see Filling. 
 Stripping, 36 
 Stull, 74, 548 
 Stump-pillars, 47, 57 
 Sty the, see Carbonic acid. 
 Sub-drifting, 80 
 Suction height, 367 
 Suction-pipe, 344 
 Sullivan machine, 663 
 Sulphuretted hydrogen, 389 
 Surface examination, 12 
 
 TAIL out, ii 
 
 Tail-rope haulage, 148, 285 
 
 hoisting, 147 
 Tamping, 675 
 Taper- rope, 199 
 Tappets, 639 
 Telephones, 215 
 Temper- screw, 60 1 
 Temperature of deep mines, 408, 
 
 55 
 
 Tempering steel, 632 
 Tension-wheel, 292, 296 
 Testing for gas, 398 
 Three- wire system, 178 
 
 Throttling governors, 109 
 Throughs, 465 
 Timber-joints, 516, 542 
 
 preservatives, 546 
 Timbering gangways, 555 
 
 rooms, 43 
 
 shafts, 514 
 
 slopes, 28 
 
 stopes, 75 
 Timbering, iron for, 562 
 
 methods of, 77, 514, 554 
 
 principles of, 547 
 Tin ores, 4 
 
 Tipple, 18, 88, 209, 242 
 Tools, 594-615 
 Top-slicing. 79 
 Traction, animal, 68 
 
 locomotive, 270 
 Tractive force, 271 
 Tramway, 267, 299 
 Transmission, 329 
 Traverses, 4, 31 
 Trepan, 538 
 Triger's method, 537 
 Trompe, 421 
 Tubes, 535 
 
 Tubing, 528, 535-540 
 Tubing wells, 603 
 Tunnel, 584 
 
 long, 605 
 
 openings by, 22 
 
 progress, 606 
 Tunnelling methods, 607 
 Turbines, 119, 189 
 Two- wire system, 178 
 
 ULTIMATE source of mineral, 8 
 Undercast, 461 
 Underhand stope, 70 
 Underholing, 42, 550, 620, 658 
 Upcast, 509 
 Upraise, see Mill-hole. 
 U. S. mining laws, 14 
 
 VALVES of air-compressors, 326 
 
 of drills, 638 
 
 of engines, 106 
 
 of pumps, 342-355 
 Vein, 3 
 
 definition of, 5 
 
 formation, 7 
 Ventilating current, 455 
 Ventilation, 386, 405, 406 
 
 amount of air, 400 
 
 natural, 409 
 
 of breasts, 286 
 
 of tunnels, 161, 323, 380 
 
 power required, 450
 
 INDEX 
 
 Viscosity of oils, 240 
 Voltage, 167, 18 1 
 
 WAC.OX-BOXES, 236 
 
 -breast, 49 
 Wall-plates, 514 
 Water-bailers, 335 
 
 -gauge, 401, 415, 449 
 
 heaters, 93 
 
 -level theory, 8 
 
 -pressure engines, 371 
 
 -purifiers, 94 
 
 -wheels, 186 
 Watt, 1 68 
 Weights and measures, 714 
 
 of various substances, 411, 714 
 
 Welding, 630 
 Westphalian system, 66 
 Wheels of cars, 238 
 Whim, 128 
 White damp, 389 
 Whiting system, 148 
 Winding, see Hoisting. 
 Windlass, 127 
 Wings, 230 
 Winze, 70 
 Wire rope, 197, 296 
 Wires, connecting, 164, 175, 691 
 Wolf lamp, 474 
 Working barrel, 344 
 Work of compressing air, 314, 321, 
 334
 
 UNIVERSITY OF CALIFORNIA LIBRARY 
 
 Los Angeles 
 This book is DUE on the last date stamped below. 
 
 OCT 2 9 1952 
 
 MAY 1 2 1958 
 
 a a o Ktcn 
 
 DEC 5 v 
 ncr. 5 RECO 
 
 I 2 RC*0 
 
 Form L9-25m-8,'46 ( 9852 ) 444
 
 A 000226556 
 
 TN 
 145 
 I25m 
 1905 
 
 peering 
 Library 
 
 STACK 
 
 SEP 73