Mining Methods and Costs

The Globe Mining District is in the southeast central part of Arizona, in Gila County. Globe, with a population of about 7000, is the terminus of the Arizona Eastern R.R., a branch line 130 miles long that connects with the Southern Pacific R.R. at Bowie.

In 1874, prospectors crossing the Pinal Mountains from the west located what is generally known as the Old Dominion mine. For several years, it attracted little attention, because of the greater interest aroused by the discovery of high-grade silver ores in some of the foothills northeast of Globe. About six years later, the prospector turned his attention to the abundant copper ore revealed by surface workings along the Old Dominion vein, and, in 1884, the Old Dominion company erected two 30-ton furnaces. From 1888 to 1893, the Old Dominion company is said to have maintained an average annual production of about 8,000,000 lb. of copper. Until Dec. 1, 1898, all supplies had to be freighted into Globe by wagons and the mines of the district operated intermittently because of high expenses, but with the advent of the railroad the Old Dominion company continued to be a large and steady producer.

In the Globe district, the production of copper far exceeds in importance that of any other metal. There are four operating companies along the Old Dominion vein and the total annual product of these properties for 1923 was about 45,000,000 lb. of copper.

The unit size of the mineral tracts in the district is the regulation mining claim, 600 ft. wide by 1500 ft. long, and all ownerships are held in fee.

Globe is 3600 ft. above sea level, and lies between the Apache Mountains to the east and the Pinal Mountains to the west. The principal drainage of the district is northward through Pinal creek into the Salt River. The general slope from the high point along the vein, where the Superior & Boston mine is, to Pinal creek, where the Old Dominion mines are, is about 250 ft. to the mile.


The oldest rocks in the district are pre-Cambrian crystalline schists known as the Pinal schist, which are the basement upon which all the later rocks were deposited. These latter rocks comprise shale, conglomerates and quartzites with a total thickness varying from 500 to 800 ft. and are thought to be Cambrian in age. Overlying these rocks is a series of limestones, known as Globe limestones, that vary in thickness from 300 to 500 ft. and range in age from Devonian to Pennsylvanian.

These rocks have been cut by numerous faults, and following or accompanying the faulting large sills and masses of diabase were intruded between the sedimentary beds. A long period of erosion followed, during which the region was deformed by further faulting to its present topography and during which the original ores were deposited.

The main fault in this district and the one along which most of the mining is carried on is known as the Old Dominion fault; it varies from 3 to 50 ft. in width and is developed for a length of approximately 3 miles. The fissure has a variety of strike and dip but is roughly north-east and southwest, with a dip of about 80° to the south.

The fault is fairly conspicuous and is easily followed, except where it is wholly in diabase, when its course is marked by a zone of brecciation stained with hematite and salts of copper.

The vein is commonly made up of brecciated shale or quartzite and mineralized with oxide of iron and the ores of copper, the overburden varying from 200 to 600 ft. The mineralogical character of the ores along the vein is simple. The oxidation of the sulfides has resulted in simple products. The pyrite and chalcopyrite have their sulfur replaced by oxygen, carbon dioxide, or silica and become hematite, limonite, cuprite, malachite, or chrysocolla. The secondary sulfides recognized in the district are chalcocite and bornite. Native gold, silver, and copper have been observed in small amounts within the zone of oxidation.


Most of the exploration along the Old Dominion vein has been done by test pitting, tunneling, trenching, shaft-sinking, drifting, and cross-cutting. All development is carefully sampled, the outline and extent of ore body carefully determined, and ore estimated on a basis of 11 cu. ft. per ton. The production, as indicated by exploitation, has proved the method sufficiently accurate.

Change in Mning Method

The principal mining method formerly in use was the square-set method but the decreasing copper content and the increasing cost of timber, together with the increasing cost of labor and supplies, made it imperative that a cheaper method be substituted. In some places along the vein where the ground is heavy and where it is imperative to keep timber close to the working face, square-setting is used, but in general that method has been superseded by newer and more economical methods; the selection of method depends on the size and shape of the orebody and the character of vein filling and walls.

Sampling and Estimating Reserves

The method of sampling is far from elaborate; grab samples only are taken from each round blasted daily. A record is kept of all samples taken in the block of ore lying between any two raises and the numerical average for the month is assumed to represent the value of the ore mined that month.

When estimating the reserve, which is done the first of every year, the area of each block of ground remaining between any two raises is measured on the profile tracings with a planimeter. This area multiplied by its average width gives the contents in cubic feet. This figure divided by 11 (11 cu. ft. ore in place is equivalent to 1 ton) is reported as the tonnage for that block. The numerical average of the year’s samples of ore broken, together with the assay values of the drift over this section, is reported as the assay value of the block remaining.

The total ore reserve is computed by multiplying the number of tons by the per cent, for each block. This product divided by the total reserve tonnage gives the average per cent., which practically checks with the heads reported by the mill. Blocks of ground that have been worked out have been found to check within 0.5 per cent, on both tonnage and value reported.


The section of the Old Dominion vein along which the Iron Cap Copper Co. is mining is about 3500 ft. long, and the width of the vein varies from 3 to 40 ft. with an average dip of about 80°. The vein material is a hard brecciated quartzite or shale between fairly good walls. The distribution of values through the deposit is irregular and some sorting is resorted to. Very few waste bodies occur in the ore zone, however, and when found are usually left in place until stopes are ready for waste filling; they are then blasted down and become part of the gob.

The inclined cut-and-fill system is used throughout the mine. This method requires very little timber; stope floors are carried on an incline of about 34°, which eliminates most of the labor of shoveling but which is not so steep as to constitute a hazard from rolling boulders.

The average stope temperature is about 78°, average relative humidity 88 per cent. Production is approximately 150 tons per 8-hr. shift of 50 men. This includes all underground labor, but only about 50 per cent, of the shift are on actual stoping operations.

Fig. 1 shows the successive steps from the starting of a stope to its finish. This plan calls for a main shaft for the handling of all men and materials and the opening up of the mine by a series of levels placed approximately 100 ft. apart.

The Iron Cap Copper Co. has two three-compartment shafts, each compartment 4½ by 5 ft. in the clear, timbered with 10 X 10-in. timber sets on 5-ft. centers and lagged with 2 by 12-in. lagging. The Iron Cap shaft, the only one at present operating is 1540 ft. deep, and is in the hanging wall. Commencing at the 800-ft. level, stations 15 ft. high by 40 to 60 ft. long are cut every 100. ft. Station sets are 10 by 10-in.

timber on 5-ft. centers with a drop of 6-in. on each set, leaving the back end of the station 11 ft. high.

Loading pockets were cut above the 1100-ft. level and under the 1300-ft. level; raises driven from each pocket accommodate the ore mined on three levels.

The shaft is so conveniently located with respect to the vein that all ground between shaft and vein constitutes the length of the station. Drifts 5 by 7 ft. are then driven along the footwall, no timber being used until stopes are started. Fig. 2 shows the arrangement and dimensions of shaft, station crosscut, and drift.

Mining Methods

As soon as the drifts have advanced far enough, raises with a minimum cross-section of 5 by 10 ft. are driven on 125-ft. centers. The sill floor sets of these raises are timbered with 10 by 10-in. timber. Posts are 8½ ft. high, sets placed on 5-ft. centers. Above the sill floor set, however, the timber is 8 by 8 in. Posts are 5 ft. 4 in. long.

The manway only is timbered with what is termed a “clap-me-down” set. The posts are set as nearly over each other as possible,

the cap is cut to fit the ground and is well blocked; 3-in. by 12-in. by 6-ft. lining boards keep the manway clean. A 6 by 6-in. sprag in the chute end flush with the last cap serves as staging for the machine men.

All headings are given definite numbers, which indicate to a certain extent their location with respect to the shaft. Headings to the east of the shaft have an even number and headings to the west, an odd number. Raises are numbered consecutively from the shaft in each direction. For instance, 902 raise No. 4 would be the fourth raise east of the shaft on the 900-ft. level. Stopes are designated as 902 stope east or west of raise No. 4.

Stopes may be started as soon as the raise has been holed through to the level above. The back of the original drift is first broken down to a height of 15 ft. and the ore mined from footwall to hanging wall. This sill-floor stope is then timbered with 10 by 10-in. sets with square framing. Sets are placed on 5-ft. centers; posts are 8½ ft. long. If the vein is not over 8 ft. wide, the cap is cut to fit the ground. In some cases where the vein is 6 ft. wide or less, and walls are exceptionally hard, hitches are cut to receive the cap and no posts are used.

A temporary chute is placed 25 ft. on each side of the original raise; permanent chutes are placed 50 ft. on each side of the original raise with a manway between.

Filling Stopes

As soon as all sill-floor timber has been placed, the sets are lagged over with a double floor of 3 by 12-in. planks and stoping starts on the first floor at the original raise. The ground is blasted out around the raise as high as safety will permit, the ore is then removed and filling poured in from the level above, the waste taking its own angle of repose. When the waste filling is about 3 ft. from the back of the stope it is roughly leveled off and floored with 3-in. by 12-in. by 5-ft. planks. Cleats of scrap timber are nailed to the floor to enable machinemen to move around easily and a cut about 6 ft. high is taken each side of the raise. When this cut is completed, the floor is taken up and piled out of the way and the opening is filled as before; the flooring is again laid and another cut taken. The flooring is used until it is worn out.

As the stope passes the temporary chute, the timber from this chute is salvaged for use elsewhere. As the toe of the incline reaches the permanent chute, the original raise timbers are salvaged as the stope progresses and are used to build up the permanent chutes and manway.

In all main drifts 2-in. air lines and ¾-in. water lines are carried, 1-in. air lines and ½-in. water lines are run down each original raise and up each center manway, so that drilling connections may be made at either the top or the bottom of the incline.

As stopes hole through to the level above, drift timbers are caught up and held in place until they can be supported by 8 by 8-in. sets placed upon the filling; these are eventually filled in and the original drift left intact.

Waste fill for stopes is obtained from development work, from shrinkage stopes above the ore zone, and from old filled stopes where no damage can be done by allowing them to cave.

Drilling and Blasting

All stoping is done with wet hand-rotated stopers using 7/8-in. quarter-octagon hollow steel. The starter bit is 1¾ in. and decreases 1/8 in. on each length of steel. Thirty-five per cent, gelatin powder is used in blasting all holes in stopes.

A grab sample is taken from all rounds blasted and a copy of the assays is furnished to bosses daily. All broken material is hand trammed in 16-cu. ft. end-dump, roller-bearing cars run on 12-Ib. rails.

The method adopted for drilling drifts, raises, and shaft sinking do not call for any special mention. This work is usually done on contract, the company providing all tools, equipment, and supplies; and the contractor providing all labor. The average price paid for drifting is $4.40 per linear foot with a minimum cross-section of 5 by 7 ft. The price for raises with a minimum cross-section of 5 by 10 ft. is $4.50 for the first 50 ft. and $5 per foot for the remainder. Shaft sinking averages about $50 per foot, depending largely on the character of rock being drilled and the amount of water likely to be encountered. All drifting and shaft sinking are done with water Leyners using 1¼-in. hollow round steel with double-taper cross bits. The starter bit is 2 in. and decreases 1/8-in. on each length of steel.

Blasting in all development work is done with 40 per cent, 1 1/8-in. gelatin powder. All development is carefully sampled and accurate assay maps brought up to date every 30 days. About 1 ft. of development is done for every 12 tons of ore mined.


Tables 1 and 2 show average detailed costs of stoping and development work per ton of ore for 85,211 tons mined in 1923. These costs constitute approximately 50 per cent, of the total cost of mining. At

present 30 miners break all ground in development work and stopes, including waste to fill stopes, and supply 300 tons of ore daily. There is an average of 100 men employed underground and a total of 135 at the mine. All labor in stopes is on a “day’s pay” basis.

Machinery and Surface Plant

The surface equipment consists of an Allis Chalmers double-drum hoist driven by a 250-hp., 440-volt, 25-cycle electric motor, each drum holding 1800 ft, of 1 1/8-in. 6 X 19 Lang lay cable. All hoisting is done in counterbalance at a speed of 700 ft. per min. Ore is hoisted from the pocket through two compartments in 4-ton skips and dumped direct into a choke feed Austin No. 7½ gyratory crusher, which breaks to about 2 in. This material then passes through a trommel and the oversize is fed to a Symons 48-in. vertical disk crusher, the final product being ½ in. A belt conveyor carries it to the ore bins, and from there it is taken to the mill, ½ mile away, by a Westinghouse 6-ton locomotive operating over a 24-in. gage electric railroad on 550-volt d.c. current.

A cage for hoisting men is hung under the skip, and provision is made for connecting a second cage underneath if necessary. An unbalanced dinkey cage is operated in the third compartment by a 10 by 18-in. duplex, direct-acting sing!e-reel Ottumwa hoist; this cage is used only to lower supplies. Skips and cages are both equipped with safety catches and are inspected daily.


At present, the water is handled by electrically driven pumps in two lifts, from the 1500 to the 1300-ft. level, and from this level to the mill on the surface. Both pumps are on the 1300-ft. level. A Lane & Bowler 500-gal. deep-well pump, six-stage, driven by a 40-hp. motor lifts the water from the 1500-ft. level through an 8-in. pipe and discharges into two 60,000-gal. concrete sumps on the 1300-ft. level. The water gravitates into a 300-gal. Aldrich quintuplex plunger pump driven by a 150-hp. motor and pumps through a 6-in. pipe direct to the mill on the surface, about ½ mile from the collar of the shaft.

Air Compression

Air for drilling is furnished by a steam-driven 3000-cu. ft. O. R. C. Ingersoll-Rand compressor. A Sullivan 1500-cu. ft. tandem, compound, direct-connected, steam-driven compressor is idle at present but can be used in an emergency.


Ventilation is provided by one Sturtevant multivane fan pulling 65,000 cu. ft. at 3½-in. water-gage pressure, driven by a 75-hp. motor, belt-connected. This exhaust fan is at the collar of the Williams shaft and is operated 14 hr. per day. Air is drawn down through the Iron Cap shaft to the lowest mine level and allowed to work upward through the stopes, finally finding its way out through the Williams workings and up the Williams shaft.

Lighting and Signaling

Stations are lighted by 32-c.p., 100-watt, 110-volt electric bulbs; main tramming drifts are lighted by 16-c.p. 40-watt, 110-volt electric bulbs. Lights in stopes are provided by carbide lamps carried by each miner.

Western Electric, local-battery system telephones are located on each shaft station and at the collar of the shaft. All signals between cagers and hoisting engineer are over an electric signal system, supplemented by rope bell.

All electric power used is bought from the Inspiration Copper Co. at Miami, Ariz., and brought over a private line a distance of 7 mi. All steam power is furnished by three 250-hp. Babcock & Wilcox water- tube boilers.


A Member.—What do they put on the filling?

A. L. Walker.—They put lagging and use that lagging until it is worn out, gradually moving it from one place to another.

A. Neustaedter, Roselle Park, N. J.—Do they put up square-set raises?

A. L. Walker.—They put up square-set raises, but use square sets only when it is imperative. If the ground is at all soft or dangerous they use stulling.

A. Neustaedter.—Cribbing would not do.

A. L. Walker.—It might in certain cases. Of course, in the old days, the square-set system was used altogether. In the Old Dominion mine, where the orebodies were 40 or 60 ft. wide and sometimes 60 ft. high, square sets were used altogether.

A. Neustaedter.—Do they work the square-set stopes on the rail too?

A. L. Walker.—Yes; all the orebodies above the eighth level in the Old Dominion were worked with the square-set system. About 1893, however, that system became so expensive that we developed a system of heavy stulling to support the roof whenever possible.

Mining Methods of Michigan

The Marquette range, on which are situated the iron mines of Marquette County, together with a few in Baraga County, Mich., extends from a point 10 miles southwest of Marquette westward for 30 miles. The tracts are usually a multiple of the standard 40-acre parcel, which is the smallest government subdivision of a square mile or section.

About half of the mines are held in fee, these being owned by the older mining companies. Some of them date back to about 1880, and a few are as early as 1850. During the past 40 years most of the mines opened have been on leased lands, the royalty being either for a stated amount per ton or a percentage of the selling price of the ore at the lower lake ports, or the selling price on board of cars at the mine.

The first merchantable body of ore discovered in the Lake Superior district was found, in 1845, at what is known as the Jackson mine, near Negaunee. This ore was hard hematite. After unprofitable attempts had been made, in 1848, to smelt it in forges, shipments were begun to lower lake ports, which increased rapidly upon the completion of the locks at Sault Ste. Marie, in 1855, and the railroad from the mines to Marquette, in 1857.

Beginning with the Pioneer, in 1858, a number of small charcoal furnaces were built to smelt a part of the product. At various points in the upper and lower peninsulas, large charcoal furnaces are still making iron from the ores of the Marquette range. In connection with these furnaces are byproduct plants.

As the deposits were opened up, soft ore was encountered, but for a few years this was disregarded, as only the hard ore was used. Underground mining was begun about 1880 and during the next few years the open-pit mines producing high-grade ore were exhausted.

The first mines in the district were owned by the Jackson, Cleveland, Lake Superior, Lake Angeline, Champion, Iron Cliffs, Humboldt, Republic, and Michigamme companies. Many of these companies have been merged in the holdings of The Cleveland-Cliffs Iron Co., which controls most of the tonnage of the district.

Ore shipments are made from Marquette through the docks of the Duluth, South Shore & Atlantic railway, constructed in 1857, and from the Lake Superior & Ishpeming railway, constructed in 1896. A portion of the tonnage is also shipped over Chicago & North Western to Escanaba. Since the above dates, larger and more modern docks have been built.

The average number of men employed in the district for 25 years is 4215. The average annual shipments from 1902 to 1920, inclusive, have been 3,827,659 tons; the total shipment to the end of 1921 is 137,237,513 tons.

Geology of District

The geology of the Marquette range is described, in Monograph 52 of the U. S. Geological Survey, by Van Hise and Leith. The iron formations occur in the Huronian series of the Algonquin group of pre- Cambrian rocks. The sedimentaries, in which occur the principal mines, stretch from Marquette through Negaunee, Ishpeming, and Champion to Michigamme and Republic, with a separate area at Gwinn. The series consists mostly of quartzites and slates, interbedded with them being the jaspers of the iron formations. All of these rocks are faulted and folded and are crossed by dikes of greenstone or diorite. The Negaunee iron formation, or jasper, in which most of the mines occur, is a combination of iron oxide and silica containing, according to the U. S. Geological Survey, about 29 per cent. iron. No commercial use has been made of this material. Its greatest thickness, where proved by drilling at Negaunee, is over 2000 ft. It rests upon slates belonging to the formation, which is locally known as the Siamo, and is overlaid by quartzites or slates of the Goodrich formation. The soft ores occur as secondary concentrates either near the base or near the top of the jasper, the latter resting upon interbedded diorite intrusions. These orebodies are often limited in depth by dikes or faulted portions of diorite or slate, which serve as impervious bases on which concentration has taken place.

The hard ores occur only at the top of the Negaunee formation, being underlaid by a few hundred feet of hard-ore jasper, which again lies on the jasper of the soft-ore formation. The hanging wall of the hard ores is quartzite or slate.

The Michigamme slate formation, which overlies the upper quartzite and the Negaunee formation, contains interbedded iron formations, which in places produce limonite ores.

The Gwinn, or Swanzy, subdistrict of Marquette range is placed, by the U. S. Geological Survey, in the Michigamme formation. Here the ore is a soft hematite, found at the base of a jasper 100 ft. or more in thickness, and resting on comparatively thin beds of black slate and quartzite or arkos overlying the granite. The iron formation is overlaid by the slates of the Michigamme formation.

The physical structure of the ores on the Marquette range is excellent, none of them having a fine enough structure to be objectionable to furnace men. The hard ores are used in lump form in the open-hearth processes; they are crushed for use in the ordinary blast furnace.

Description and Topography

The district is about 800 ft. above Lake Superior, or 1400 ft. above the sea level. The surface is hilly and rocky. Lakes and swamps are bordered with terraces of glacial origin, above them rising rocky hills of iron formation, quartzite or diorite, the tops of which are from 50 to 200 ft. above the glacial terraces. Because of the proximity of Lake Superior the summer climate is cool. The prevailing northwest winds bring heavy snow falls from the beginning of November until the middle of April. The winter temperature is modified by the lake, which never freezes over entirely. The average yearly rain fall, which includes the equivalent in snow, is about 32 inches.

Mining timber is brought in by rail. Practically all the white pine of the Upper Peninsula has been removed, and in the neighborhood of the mines, most of the hardwood also has been cut. There still remain districts where there are large stands of hardwood, together with tamarack, hemlock, and cedar.

Labor conditions have been excellent. The mining population, derived chiefly from northwestern Europe, has been industrious and thrifty. Most of the men own their homes, the lands upon which they stand being either purchased or leased from the companies on easy terms, which induce building. However, since the war, many workmen have been attracted by high wages to the large cities, this movement being accelerated when the mines were shut down during the recent depression.

The ore is transported by rail from the mines to the docks in hopper- bottomed cars of 50-ton capacity. It is there dumped into pockets of 200-ton capacity. In loading vessels, the ore is delivered through spouts, which are lowered to the hatches. These spouts are 12 ft. from center to center, and the hatches on the boats are 12 ft., 24 ft., or a multiple of 12 ft. apart.

Pumping at the mines varies from a few hundred to about 3000 gal. per min. The overlying sand and gravel are often saturated with water, but owing to embedded clays and other courses, this is seldom drained until broken by the extraction of the ore, under the caving system.


The earliest explorations were for the purpose of finding the ledge or deposit from which had come the many broken masses of hard ore found lying upon the surface or in the glacial material. Succeeding explorations were conducted by test pitting through the overburden, drilling the same way from outcrops, or tunneling into the rocks themselves. In many cases, shallow shafts were sunk, from which drifts were driven.

The diamond drill was used, as early as 1869, for deep holes in hard orebodies, but its use was not customary until about 1878; since then it has become the usual method of exploring for both surface and underground work. The churn drill has not been used much for exploring, owing to the hardness of the rock capping to be penetrated.


As all explorations of orebodies at present are by diamond drilling, the only sampling is that of the core and sludge, from the drill holes. These are collected after each 5-ft, run and later analyzed; the weighted average of the two, figured on the proportion of each covered by the length of the run, is the analysis for that run. These analyses, combined for the entire hole, give the depth and grade of known ore encountered. Duplicate samples of core and sludge from each run are preserved for reference. If the presence of soluble sulfur is suspected, the amount of water pumped down the hole and the amount coming out are measured, and samples of this water taken at definite intervals, these samples being analyzed for sulfur. This is then combined with the analysis of insoluble sulfur in the core and sludge. When drilling through hard ore, almost complete recovery of core can be made; while in soft ore practically no core is obtained, and the only analysis is that of the sludge. The sludge is collected by causing the water from the drill holes to flow into boxes 4½ by 1½ ft., with two baffle boards, in which the particles of ore held in suspension settle out.


The methods of estimating are those customarily used in the Lake Superior district, both for ore found by diamond drilling and for developed ore underground. This is a comprehensive subject and should be treated in a separate paper. Plans and cross-sections are made both of explorations and mine workings to show the area and depth of the deposits as soon as ascertained. The limits of formation are shown on these drawings as a result of careful geologic examination. Considerable latitude for judgment must be permitted in the case of partly developed orebodies, which for the purpose of estimating the cost of development must be divided into ore in sight and prospective ore. The number of cubic feet per ton varies from 8 to 9 in hard ores and is about 12 in soft hematites, while for limonites it is as high as 13 or 14. The U. S. Geological Survey states that the soft ores of the Marquette range average 12 cu. ft. to the ton. This is borne out by a series of careful tests, which have recently been made in several mines.

Accuracy of Method

Due to the irregular shape of the various orebodies, large discrepancies have often been found between careful estimates based on exploration and the tonnage eventually recovered. Drill holes in some cases follow chimneys, the formation proving barren except for a small cross-section in the vicinity of the hole. On the other hand, some deposits have proved to be much greater than estimated by drill holes. Reasonably correct estimates can be made in shallow deposits from a number of short holes at close intervals, but the difficulty increases for depths over a few hundred feet, due to the deviation of the diamond-drill holes and the lack of knowledge of the geology of the formation.

In the case of shallow regular orebodies, as on the Mesaba range, the tonnage can be accurately estimated and the percentage of extraction determined. This is not the case on the Marquette range, where the ore deposits, as a rule, are deep and extremely irregular.

History of Principal Mining Methods

The early shipments of iron ore previous to 1860 were made largely from loose masses found scattered on the surface. After this source was exhausted, open pits were developed, some of which continued to operate until about 1880. Drilling by hand, blasting with black powder, loading into carts drawn by horses or mules, and again loading into 7-ton cars on the railroad constituted the usual method. Shafts were started for collecting the water in the pits, so that it could be pumped. As the pits grew deeper and available ore seams were followed, winzes or stopes were sunk, the ore being raised by horse whims. Most of the early appliances were introduced from Cornwall, whence the first miners came. Details of the mining methods, including cost, are given in Volume 1, Geological Survey of Michigan, published in 1873.

As the open pits were continued, they increased in depth until a point was reached, on account of the ore dipping under the rocks, where it was not profitable to remove the overburden. It was then necessary to sink inclines and provide mechanical means of hoisting. The hard ore was then mined in open stopes, pillars being left to support the capping.

From 1875 to 1880, much of the mechanical equipment was introduced, including rock drills, electric lights, and electric signals. At this time dynamite replaced black powder as the chief explosive. Hard-ore mining continued in about the same manner at depth. The breast stoping method of room and pillar was used, in which only enough ore was left in pillars to support the hanging.

The first soft ore was found near hard ore deposits. The demand for this class of ore was limited, previous to 1880. It was mined in open pits, but when it became unprofitable (on account of the increased depth and the dipping of the ore under the rock, to remove the overburden) underground mining was started. This soft ore could not be mined by the open stope and pillar method; timber was necessary to keep the places open. The principal method of mining soft ore was, in the middle eighties, almost entirely by the square-set system of rooms and pillars. The rooms were usually three sets, or 21 ft., wide and as long as the orebody. As a rule, the pillars were of the same width as the rooms. In many instances the rooms were carried to a height of ten or twelve sets, or 70 or 84 ft. After the ore had been mined in the rooms, an attempt was made to get what was left in the pillars by raising and running them. This system was extremely wasteful, as the ore soon became mixed with rock and the grade lowered so that work had to be stopped.

The caving system was introduced by miners from the north of England, where it originated. The method there was to mine from a sub-level immediately below a mat of timber, which was kept propped up until the retreat of mining began. Every effort was made to maintain this timber mat, for when it was destroyed a new mat had to be made at great cost. Modifications of this caving system were introduced during the early eighties, until it became the accepted method about 1887 to 1890 (see J. P. Channing in the Lake Superior Mining Institute, Volume 19); the introduction of the caving system was hastened by the decreasing local supply of large timber.

Permanent shafts in the foot wall were rare before 1895. Timber was used exclusively for shafts, shaft houses, and trestles. About 1900, steel replaced wood in shaft houses and, about 1910, concrete and steel were used for shaft lining. The most recent practice, inaugurated in 1919, is to build enclosed shaft houses of reinforced concrete. Electric haulage was introduced in 1892, since which time its use has become general.

Certain changes in mining conditions were brought about by legal restrictions, though in justice to the mining companies it should be said that much of the beneficial legislation enacted was prompted by the managers. The election of the mine inspector in each county, provided for in 1887, assured a greater degree of care for the safety of the workmen, which was increased after the passage of Michigan’s workmen’s compensation law in 1912. All large companies now have a department for safety inspection and first-aid training, although these are not required by law.

Early operations were wasteful, because of the lack of system and the necessity of marketing only the higher grades of ore. As the number of leased properties increased, it became necessary that all ores should be taken out in a workmanlike manner and unnecessary waste prevented. The principal causes of loss in the early years of the district were: (1) the lack of preliminary exploration and the beginning of caving before the limits of the orebodies had been determined; (2) the fact that softer material could be mined more cheaply than hard, which consequently was left in place; (3) in attempting to reduce the cost too great a vertical distance was taken between sublevels; (4) the lack of proper maps and systematic method of laying out the work.

The larger and more progressive mining companies, realizing these mistakes, inaugurated geological investigation, systematic development and close supervision, which resulted in the present methods by which losses have been reduced to a minimum.


The demand for crushed soft ores for charcoal furnaces necessitated the introduction of crushers in shaft houses.

Attempts have been made to concentrate some of the lean ores but, principally because of the intimate association of silica with the iron oxide, these have failed. The first concentrating plant in the district was built at the Jackson mine in about 1880. It failed because jaw crushers and rolls could not be made hard enough to withstand the wear and tear (manganese and other alloys of steel were not then in general use) and because, regardless of how fine the crushing might be, the particles of iron oxide and silica were too closely associated to be separated. Magnetic concentration was attempted, about 1890, by Thomas A. Edison at Humboldt and Michigamme by means of machinery introduced from Sweden for the magnetic treatment of magnetites. These attempts failed, owing to the small amount of ore available for concentration. At the American-Boston, a concentrator for soft ores was used until the mine was shut down. This method depended on a special structure of low-grade ore. Crowell & Murray’s “Iron Ores of Lake Superior” gives valuable information on the various ores of this range.

Mining Methods in Use

There are no rich ores close enough to the surface to permit open pit mining. Doubtless many deposits, now exhausted, that were-opened underground could have been stripped with modern equipment and the ore mined at a profit. Only the lean ores are now mined in open pits. The factor deciding the question of open pit or underground mining is the cost of stripping the overburden as compared with underground cost. Climatic conditions do not interfere, as shipments are made only in the summer. In open-pit mining, the systems used are: steam shoveling directly into standard railway or narrow-gage cars; milling into raises to underground drifts, then tramming the ore to a shaft, where it is dumped, hoisted, and run into ore cars.

Varying conditions necessitate a number of minor differences in mining methods, because hardly any two orebodies in the district are of the same size, shape, or physical structure. Underground mining may be separated into hard- and soft-ore mining. The hard ores are comparatively unimportant, as only a few mines on this range contain this grade. The systems used are either breast stoping into rooms and pillars, or shrinkage stoping if the vein is narrow, steeply dipping and has firm hanging and foot walls.

Underground Mines

Most of the soft ore mined on the Marquette range is won by the top-slicing method. The typical orebody is a large mass with width exceeding thickness, lying in a basin of slate or diorite, or both, with a flat pitch and with the overlying jasper for a hanging wall or capping. The width may be as great as 1200 ft., the thickness 200 ft., and the length indefinite. In orebodies of such shape and dimension, top slicing is the only system by which great loss of ore can be avoided and the ore obtained without the grade being seriously affected by its being mixed with rock. The top-slicing system is flexible, in that any horses of jasper or large dikes can be left. By daily sampling from the breast of the working places, the ore can be hoisted and shipped, or stocked, according to grade. In large orebodies, slicing is carried on at different elevations, as extremely large sublevels are practically impossible to keep open. A large product can soon be obtained by starting work in a number of different places, each of which must be immediately below the hanging jasper. The disadvantage of this system is the amount of timber required and the heat generated by the decay of the timber. However, as practically all of the mines have two openings to the surface, good ventilation can be obtained by natural or mechanical means and the temperature kept within reasonable limits.

Mine Opening

Mine openings are entirely by shafts, all modern ones being vertical; the deepest in the district is about 2500 ft. The size regarded as standard is 10 ft. 10 in. by 14 ft. 10 in. inside. This is divided into four compartments, the arrangement of which is shown on Fig. 1. Modern shafts are constructed with concrete walls and steel sets; some are circular, others rectangular. Great care is taken to locate them in the solid foot wall at a point where they will not be disturbed by caving.

The arrangement of the loading and discharging pockets is shown in Fig. 2. There are usually three 60-ton storage pockets, from which the ore is drawn into two measuring pockets, each holding enough for one skip. In some cases, additional storage is gained by raises from one level

to another, thus avoiding the expense of installing pockets on each level. In some mines storage raises at the shafts, 200 ft. high, are satisfactory.

Underground Development Plans

In Fig. 3 are shown the main levels, sublevels, raises, and chutes, with their relative dimensions and intervals.


Drilling and Blasting.—In drifting through hard rock, water-feed hammer drills on cradles are used, while for drifting and raising in ore hand machines of the jackhamer type, fitted with auger bits, are employed. For rock raising, the stoper is in common use. For shaft-sinking, air or water-feed hand sinking machines are used. The size and shape of the steel and bits are shown in Fig. 4; Fig. 5 shows the arrangement and depth of holes for cuts in ore and rock drifts. For tamping,

paper bags filled with fine material are supplied. In most soft ores, the explosive is 40 and 50 per cent, low-freezing ammonia; while in some of the harder ores 50 and 60 per cent, gelatine is used, occasionally 80 per cent.

gelatine is necessary. The air pressure at the drill is usually between 70 and 80 pounds.

Drifting and Sloping.—In laying out the main haulage levels, parallel crosscuts at intervals of 150 ft. are driven. From these crosscuts, raises are put up at intervals of 40 to 60 ft.; these raises are carried through to the capping. At the top of the raises, immediately below the jasper, sublevels are started. Crosscuts are driven to the proper limit and slicing is commenced. As the ore is removed, the floors are well covered with either lagging or 5/8-in. covering down boards. Succeeding sub-levels are driven at intervals of from 10 to 12 ft. On each sublevel, during the process of slicing, the floors are well covered. Fig. 3 gives details for position of drifts and raises.

Timbering.—In the main and sublevel drifts, round hardwood, hemlock, and tamarack timber are used. On this range, timber has not been chemically treated, although preparations are being made to do this. It is thought that before long all of the timber for main levels will be chemically treated.

Timber that has been framed on the surface is delivered at the bottom of raises on timber cars. From these points it is hoisted to the working places by means of small air hoists. It is not the practice, on this range, to remove any timber during the process of working the caving system. The details of the timber framing are given in Fig. 6.

The legs of the standard level set are usually 8 ft. long, though 9-ft. legs are used at the bottom of raises and in some haulage drifts. Logs are delivered in 8-ft. and 16-ft. lengths, the 7-ft. stub from a 9-ft. leg being used as a short leg or cap in sublevels. For main-level sets, legs are 12 to 14 in. in diameter; for sublevels, 8 in. to 10 in. and 10 in. to 12 in. Sets are usually 5 ft. apart and are braced as shown. A longer brace, behind the lower one shown, is spiked to both legs. Lagging and, where necessary, blocking is placed above the caps.

Specifications for Timber

Stull Timber (Legs and Caps)

Hemlock, hard maple, soft maple, yellow birch, tamarack, and Norway pine. Must be sound, straight, and green. Ash, white birch, poplar and balm of gilead not accepted.

Tamarack in top diameters of 8 and 9 in. containing not more than ¼ in. sap rot in depth, accepted; in diameters of 10, 11, 12, and 13 in., containing not more than ½ in. sap rot in depth. Lengths, 8 and 16 ft.


Tamarack, spruce, pine, hemlock, maple, and yellow birch. Must be sound, straight and green. Top diameter, 6 to 8 in. Lengths, 5 ft. 4 in., 10 ft. 8 in., and 16 ft.


Straight sound cedar, tamarack, and spruce (10 per cent, jack pine permitted).
Round, 3- to 4½-in. top, larger than 4½ in. to be split.
Split, not less than 2½ by 4 nor greater than 3 by 6.
In 5-ft. lengths, 160 cu. ft. per cord.
In 7- and 8-ft. lengths, 128 cu. ft. per cord; not less than 125 pieces.

Covering-down Boards

No. 3 maple and birch. To be resawed from 2-in. maple and birch hearts to 5/8 in. To be sound, green lumber.
Widths, 6 in. and wider.
Lengths, not under 6 ft. and not over 9 ft.

Underground Track Ties

4 ft. 6 in. long, in- thick, 4½-in. face.


10 ft. long, cut from 20- and 30-ft. lengths, 3 to 4½ in.

Underground Sampling.—All working places are sampled at intervals of about 5 ft.; these samples make it possible to grade the ore. In addition, the ore from each chute is sampled as the motor cars are filled; this gives a check sample on the grade. All cars, before they are dumped into pockets at the shaft, are sampled, the samples being kept separate, according to mine chutes. On the surface, during the shipping season, a sample is taken from each skip as the ore runs into the railroad car; during the stocking season, it is taken from each car before it is dumped on the stock pile. The samples taken on the surface are the ones reported for the grade, those taken underground are simply used as a check.

Loading Machines and Scrapers.—Two heavy types of loading machines (the Shuveloder and the Hoar) have been successfully used in main level drifts. These machines are not practical, except on main levels. For sublevels, the John Mayne sublevel loader, Fig. 7, has been successfully used for over two years; this machine will be on the market in a short time. It is simple in construction and has few moving parts. It stands up under continuous work and can be operated by any miner. Records for a year for a gang using this loader show an increase over hand shoveling of 93.6 per cent, in tons per man per day; a decrease in

price to contractors of 32.6 per cent., and an increase in monthly earnings of miners of 19.5 per cent. Scrapers operated by double-drum air hoists have been used successfully in a limited way in both hard and soft ores. They load rapidly in a straight drift up to 75 ft. from a raise.

Tramming and Haulage.—Electric haulage is used almost exclusively in the district. Direct current is generated from alternating by rotary converters, situated on the surface or underground. Locomotives are 6 tons in weight and the current is received from an overhead trolley. The ore cars are of steel, side dumping with saddle backs, of 64 cu. ft. capacity, or approximately 4 tons of ore. The gage of tracks is 30 in. and the weight of rail 30 to 40 lb. The grade is 0.5 per cent, with the load. The cars are usually equipped with roller-bearing wheels and are dumped by hand at the shaft though the most recent installations have been rotary dumps, using round bottom cars.

Hoisting.—At most of the mines electricity is used for hoisting, although some still use steam. These hoists vary in horsepower from 400 to 900, depending on the depth of the shaft and the size of the skips. As alternating-current hoists of this size throw a heavy variable load on the power line, the newer hoists are equipped with flywheel motor-generator sets following the Ilgner system. With the skip hoist operated by direct current, an approximately constant power load is

maintained. The drums are from 8 to 10 ft. in diameter and from 8- to 14-ft. face. Hoisting is usually done in balance, at a speed from 1000 to 2000 ft. a min. For a 4-ton skip, which is the average size, a 1¼-in. plow-steel rope is used, except in the deeper shafts, which use 1 3/8-in. ropes. At almost all mines, a separate hoist is provided for the men. The cages have a capacity of from 24 to 30 men, and the speed, when men are being handled, is about 800 ft. per min. The hoists are provided with Lillie overwind device and the cages are equipped with safety catches and are balanced by counterweights. The counterweights are cast-iron cylinders, operating in 12-in. iron pipe. Most of the head frames are of steel, though one mine has an enclosed concrete structure.

Pumping.—Both plunger and centrifugal pumps are used, the motive power usually being electricity. These have capacities from 500 to 1600 gal. per min. against heads from 500 to 2400 ft. The amount of water pumped varies, in the different mines, from 150,000 to 3,500,000 gal. per day.

Air Compression.—Most mines are equipped with two-stage inter-cooled compressors of a capacity of 2000 cu. ft. per min.

Ventilation.—Nearly all mines have two openings and sufficient ventilation is therefore provided by natural means. In a few cases, where there is only one shaft and the connecting drift with another mine does not furnish sufficient ventilation, fans are installed underground. These are multivane blowers, with a capacity of 40,000 cu. ft. per min. against 3 in. water-gage pressure, operated by 50-hp. 2200-voIt a.c. motors. Where such installation has been made, the fan is placed on the bottom level, the cage compartment being used as the intake and the skip compartments as the outlet on the upper level, the lathing between these compartments having been made tight. By means of doors on the various levels, the air is forced through all the working places and discharged into the skip compartment on the top level. In order not to interfere with haulage, these doors are opened and closed by pneumatic cylinders controlled, from a distance, by ropes, red and green lights indicating the position of the doors.

Lighting.—In the shaft houses and on the underground plats, 55-volt, a.c. lamps are used; while in the main haulage levels 250-volt d.c. lamps are installed, which obtain electricity from the trolley wire. These are placed at all switches and at intervals from 100 to 200 ft. along the drifts. All men in the mines use carbide lamps.

Telephones.—Telephones are installed at all main level plats, in pump rooms, at underground hoists, in the lander’s station in the head frame, in hoisting houses, and in the various surface buildings, such as office and shops. Signaling in the shaft is by means of a.c. electric bells, all mine plats being connected to the engine house on two lines. The repeating bell system is used for operating cages. The cage rider is not permitted to open the door of the cage until the stop bell has been received from the hoisting engineer. A wire-pull bell is always installed in the cage com¬partment for emergency signals. No method of signaling in the haulage ways is used except that, on large levels operating more than one electric motor, there are colored lights at the entrance to each crosscut, which are automatically lighted when the motor is in that crosscut.

Disposition of Ore after Reaching Surface

During the shipping season, which is from May 1 to Dec. 1, ore from the skips goes directly into standard railway cars and is hauled to the ore docks at Marquette or Escanaba, for shipment by boat to lower lake ports. During the winter months, ore is stockpiled from trestles about 40 ft. high, usually built with wooden bents. As wooden trestles must be torn down each shipping season and re-erected in the fall, permanent steel stocking trestles are used at mines of a long life where there is sufficient ore to warrant the larger initial expenditure.

Safety and Welfare Work

Most of the large mining companies employ safety inspectors, who make regular or periodical trips through the mines. A book of safety rules is given to each employee, who signs a receipt for it. Examinations are held on these rules by a special committee. Failure to pass is sufficient cause for dismissal. Once a year, a special committee of workmen visits other mines to compare the safety appliances with rules as enforced in the different properties. This committe makes a number of safety suggestions, which are considered by the officials and almost invariably are accepted. At regular intervals, a certain number of men in each mine are trained in first-aid and apparatus work. Once a year, teams are selec¬ted from the various mines and field meets are held to compete for prizes in first-aid work.

The amount of welfare work varies with the different conditions. Some of the companies have hospitals, attached to which is a staff of doctors. In these hospitals, not only the cases resulting from mine accidents are treated, but also cases of sickness in the community. Nurses are employed, who visit the houses of the workmen, give help, and advise in case of sickness. A small uniform charge is paid monthly by all employees to cover doctors’ fees for ordinary consultation and visits, while the amount paid by the company for compensation takes care of cases of accidents. Some companies also have a system of pensions. Most of the companies take an active interest in the community by providing in the mine locations, where there are no theaters and other means of recreation, club houses with reading rooms, gymnasiums, bowling alleys, moving-picture machines, etc.

Precipitation Efficiency of Zinc Dust

It is generally realized that in cyaniding the precipitation efficiency of zinc dust is due to the fine division or extended surface of its metallic particles; but frequently it is thought that the presence of other metals, say2 to 3 percent, lead, is advantageous, causing more complete precipitation. The results of testing about fifty brands of commercial zinc dust have led to the conclusion that there is a distinct relation of precipitation efficiency to fineness and that the effect generally can be estimated by examining the size of metallic particles. The presence of lead was not found to be of any importance.


Generally the term “97 per cent, to pass a 350-mesh screen, 95 to 97 per cent, uncombined metallic zinc” is used by the leading European exporters. Among the many methods of determination of metallic zinc, I have found the iodine test (iodine in potassium iodide) very satisfactory and rapid. It has been controlled by the other methods, samples of the same product having been sent to three different analysts:

Iodine method………………………………………98.11 per cent, metallic zinc,
Volumetric method………….07.63 per cent, metallic zinc, Ledoux Co., New York
Bichromate method…………..98.16 per cent, metallic zinc, Wataon Gray, Liverpool, England
Bichromate method…………..98.20 per cent, metallic zinc, Norway Inst, of Tech., Trondhjem

Determination of Precipitation Efficiency

The method devised by W. J. Sharwood was used. The method proved to be satisfactory, the tests being merely for the comparison of different samples, hence the personal factor in manipulation was eliminated. As nearly all tests showed more zinc in solution than was accounted for by the silver precipitated, the term “dissolved zinc” was introduced—it means zinc dissolved by the action of cyanide and oxygen:

Zn + 4KCN + H20 + O = K2Zn(CN)4 + 2KOH;

or more probably, resolution of its equivalent precipitated silver, as tests stirred two hours showed more “dissolved zinc” than those stirred

one hour. To determine “dissolved zinc” the silver precipitate was dissolved in nitric acid, silver titrated with thiocyanate, and solution titrated with ferrocyanide (after removing silver precipitate) giving the amount of intact metallic zinc left in the zinc dust silver precipitate. The difference between active plus intact zinc and total metallic zinc is “dissolved zinc.”

Sources of Zinc Dust Tested

Samples 1 to 4 are American zinc dust; 6 to 8 are Norwegian, electrothermic fumed dust; 9 to 14 are from an electrothermic experimental plant; 15, origin is unknown, sample was furnished by Ste. Generale de Commerce & Exterieur, Paris; 16 is Belgian dust; 17, German; and 18, electrothermic blue powder (byproduct from electrothermic zinc smelting). Samples 4 and 5 were atomized, the other distilled.

Precipitation Efficiency as a Factor of Fineness

The microscopic examination showed that the distilled zinc dust consists of almost perfect spherules. The appearance is almost clean and metallic except in samples 16 and 17, where numerous particles of oxide are shown. The atomized dust (samples 4 and 5) has a coke-like surface and is very coarse; especially sample 5, which was made by a 100 lb. air pressure.

The number of spheres in a pound zinc dust, assuming the specific gravity as 7. is 0.1235/d³, hence

if the diameter is 1 mm., 1 lb. will contain 123,500 particles
if the diameter is 0.1 mm., 1 lb. will contain 1235 million particles
if the diameter is 0.01 mm., 1 lb. will contain 123,500 million particles
if the diameter is 0.003 mm., 1 lb. will contain 4600 billion particles

In sample 15, spherules of 0.003 mm. diameter were found to be pre-dominant, hence the number of particles in 1 lb. (89 per cent, metallic zinc) is 4600 billion X 0.89 = 4100 billion. In the precipitate was left 28 per cent, of the dust’s metallic zinc content, the diameter of remaining intact zinc spherules is then: 4100 billion = 0.1235/d³X 0.28 and d = 0.002 mm. The original spherule was 0.003 mm., hence the thickness of active surface is 0.0005 mm.

In sample 8, the major particles were of 0.004 mm. diameter. The number of particles in 1 lb., is 2000 billion. In the precipitate was left 43 per cent, of the metallic zinc, then the diameter of the intact particle is 2000 billion = 0.1235/d³X 0.43 and d = 0.003 mm. The thickness of active surface is 0.0005 mm. and so forth.

In the same manner, the efficiency of a zinc dust may be estimated on the basis of fineness as:

The diameter examined being 0.002 mm., the efficiency is 88 per cent.
The diameter examined being 0.003 mm., the efficiency is 70.3 per cent.
The diameter examined being 0.004 mm., efficiency is 58.0 per cent.
The diameter examined being 0.008 mm., efficiency is 33.2 per cent.
The diameter examined being 0.01 mm., efficiency is 27.0 per cent.
The diameter examined being 0.03 mm., efficiency is 9.7 per cent.
The diameter examined being 0.07 mm., efficiency is 3.5 per cent.
To get a fair comparison between the found efficiencies and those from fineness estimated values, it is necessary to eliminate what is called “dissolved” zinc. This is possible by figuring the precipitation efficiency from the difference of metallic zinc left in precipitate; which is here called “true efficiency.”

From the foregoing data there is little doubt as to what role fineness is playing. The consumer frequently calls for high content of metallic zinc, but mostly he buys in accord with the efficiency obtained in practical running. The producer should, therefore, direct his attention to improving the fineness—under maintenance of the highest content of metallic zinc—until it becomes really fume.


G. M. Brown, New York, N. Y.—Andre Dorfmann of the Mclntyre Porcupine Mine made a similar test, some years ago, relative to consumption of zinc and the results he obtained confirm the statements in this paper. In addition to the amount of zinc left in the precipitation presses, he determined the amount of zinc in the barren solution. As this solution was recirculated through the system, he also determined the amount of zinc precipitated from solution, in the ball-mill, tube-mill and agitators, before the solution was again returned to the precipitation presses.

Charles E. Locke, Cambridge, Mass.—The thing which struck me in looking through the table is that the maximum figure is about 60 per cent, efficiency, when based on the metallic zinc content, and the ordinary efficiency, if I interpret it correctly, ranges from a maximum of 57 per cent, with the finest dust down to 6 per cent, with some rather coarse samples of zinc dust.

Ecuador Mining Methods

The mines operated by the South American Development Co. are located in the Zaruma mining district of southwestern Ecuador. They are near the old mining town of Zaruma, which is the only important city in the canton of the same name. The district is situated in the upper end of a valley lying between two spurs of the Cordillera, one of which may be considered the Coast Range and the other an intermediate range. The mining camp proper, known as Portovelo, and the plant are 2.4 km. south of Zaruma, on the north bank of the Amarillo River, a tributary of the Tumbes River, which flows southwestward through Peru to the Pacific.

Portovelo is difficult of access; it is reached from Guayaquil by means of river steamers and muleback. Embarking at Guayaquil, the route is down the Guayas River, across the Jambeli Canal, and up the Santa Rosa River to Santa Rosa, a distance of 177 km. From Santa Rosa, the road or trail follows the Santa Rosa River, then across the summit of the Coast Range into Portovelo, a distance of 74 km.—a two-day trip on muleback.

The mining property of the South American Development Co. comprises 297 lode claims and 142 placer claims. The combined lode claims cover an area, roughly rectangular in shape, 9000 m. long from north to south by 4000 m. wide; the northern edge of the rectangle is about 3000 m. north of the central plaza of Zaruma. Mineral lands in Ecuador are held by virtue of an annual tax per claim paid to the central government, no surface rights being included.


The Zaruma mine has been worked by white men since 1549, probably before that it was worked by Indians. The followers of Pizarro, led by traces of gold in the sands of the Tumbes River, followed that stream to its source, worked the gold veins that they found and established the city of Zaruma. Among the mines worked by them were the Sesmo, Leonora, Viscaya, and Mina Grande. Only the soft ores at the outcrop were mined, making many shallow openings, none being of any considerable vertical extent. Traces of these old superficial workings have been important in outlining recent development.

Foreign capital became interested in the district in 1875, when a Chilean company was formed to work several of the old mines. Its efforts met with little success, as did the attempts of a few small, contemporary, native companies; in 1880, an English company, the Great Zaruma Mining Co., was organized to take over the properties. Later this was reorganized as the Zaruma Gold Mining Co. and carried on operations for several years; it spent considerable money on road construction and the building of a 20-stamp mill. Sporadic efforts to open up the Fortovelo mine were made, some stoping was done, and some bullion shipped. Later, the South American Development Co. acquired the rights and property of the English company, together with those of other smaller companies; it has since increased, its holdings through purchases and denouncements. Since 1900, there has been continuous exploitation of the mines of the Zaruma mining district by practical modern methods.


At present, the mines are operated in three units, known as the Portovelo or shaft mine, the Soroche, and the Jorupe. In addition, considerable work has been done in opening and prospecting other old mines of the district.

The mines are located 610-914 m. above sea level. The district is rugged, being cut by many small ridges and streams. The climate is tropical to subtropical, the temperature ranging between 82° F., at mid-day, to 62° F., at night. The dry season extends from about the first of June to the last of December, with occasional showers, and the wet season from December to June, with late afternoon showers; the average annual precipitation is about 70 inches.

Power is obtained from the Amarillo River through two canals, the water being used on turbines or Pelton wheels and the power used directly or converted into electricity to be transmitted and used in various operations.

For timber, the mine is dependent on native supplies; that within a radius of 5 miles has already been used. Native contractors supply all timber, dragging it in with mules during the dry season. The maximum size of round timber is 12 in.; and of squared 9 in. in lengths up to 10 ft. Timber is not seasoned, for it appears to last better when placed underground in a green condition. The hard, heavy woods, sanon and amarillo, are used extensively for timbering and tarapo in 4-in. poles for lagging. While the woods are very hard, they are short- fibered and do not withstand blasting well.

The mine freight from the coast, averaging about 750 tons yearly, is brought in on muleback during the dry season, by contract. The cargoes are arranged in loads of about 200 lb. per mule. Some heavy pieces are slung between two mules while a string of mules is used to carry cables, each carrying sufficient coils to make a load of 200 lb.; exceptionally heavy pieces are lashed to bamboo poles and carried by relays of men. The company has improved the trail in many places, cutting out fords and paving muddy places. The movement of freight from the coast costs about $35 per ton, hence high-grade material is used throughout the operations— high-strength dynamite, high-strength cyanide, special alloy steels, etc. Supplies are ordered through the New York office and purchases are made to the best advantage either in the United States or Europe. Delivery is made at Puerto Bolivar, where it is placed in river steamers that carry it to Santa Rosa, the lower terminal of the mule trail. The long time between ordering and delivery at mine necessitates placing the orders nearly a year in advance.

The mine labor is nearly all native Ecuadorian (a mixture of Spanish, negro and Indian); a few Columbians, Peruvians, and negroes also are on the pay roll. Bonuses are paid to men working 20 shifts or more per month, but the many religious holidays interfere materially with steady work. Nearly all work is done by contract. There is one pay day per month, but the men may draw from day to day the greater part of what is due them.

The Ecuadorian unit of currency is the sucre, with a par value of $0.4878. Recently it has fluctuated through wide ranges, therefore it is practically impossible to give the operating costs in dollars.


Granites and syenites, connected with gneisses and crystalline schists of Archean age, are the dominant rocks of the eastern range or main Cordillera, while in the Coast Range and inter-Andean country greenstones and porphyries are found in connection with Cretaceous formations.

The gold-quartz veins worked by the South American Development Co. occur in a belt of greenstone. The dominating structural feature (shown in Fig. 1 by the heavy line) is a major fault zone, known locally as the Abundancia fault, that has been traced several kilometers on the surface. Its general strike is N. 3° W., with an average dip of 65 to 68° to the east. Underground workings have opened up this fault for about 2000 m. along the strike. It is always strong and well defined, with a heavy gouge indicating a great deal of movement, and contains more orebodies than any of the other known fissures. Several lesser fractures making away from this fault at small angles are ore carriers; these are known as Cantabria, Portovelo, Soroche, Twenty-six, Nudo, and Quebrada veins. Tamayo and Jorupe veins are outlying fractures; their relation to the Abundancia fault has not been proved but they are commercially valuable. The San Guillermo vein, worked in the Soroche mine, is without doubt a development on the Abundancia fault proper and should not have a separate name. The Agua Dulce vein, which is in the development stage and on which some ore has already been found, may . be an extension of the Cantabria system of mineralization. What is known as the Portovelo vein, in the southern part of the property, is really an Abundancia fault, and should be known as such.

The rock in either wall of the fault is dacite. Near the veins it is considerably altered. The width of the altered dacite at the southern end is about 15 m., while at the northern end where there are more branch veins, the zone is 200 m. wide.

About 2.7 km, from the Soroche mine is a sharp conical peak of rhyolite; it is quite possible that the veins of the Zaruma district are genetically related to this intrusion of rhyolite.

The veins as developed along the Abundancia and Portovelo faults are composed of an intergrowth of quartz and massive calcite, with subordinate amounts of iron and copper sulfides, sphalerite, and galena. Considerable gouge is present on the foot wall, indicating much movement; and in some of the stopes, pronounced brecciation and recementing of the breccia is distinctly evident, showing at least two stages of mineralization. Some of the branch veins have different characteristics, indicating different periods of mineralization. The Twenty-six vein, in the upper levels, is practically pure quartz; but, with depth, it approaches the Abundancia type. Soroche and Tamayo veins consist of ribs of hard quartz, alternating with soft sugary quartz and without calcite. Cantabria, Nudo, and Agua Dulce veins, while having quartz and calcite in normal quantity, carry an excess of sulfides. Jorupe vein carries little calcite but large amounts of sulfides. It appears as though the gold-bearing solutions were introduced into the various fissures through the Abundancia fault, but penetrated only a certain distance from it. In all veins where high-grade ore is found, tetrahedrite is present.

The major fault, the Abundancia, represents the oldest of a series of faults; all others are of minor importance. The other veins never break through the Abundancia fault, but have a tendency to flatten against it and parallel it before they pinch out.

Vein Descriptions

The veins as opened by underground workings on A level, or projected on to that level from others where they have been opened up, are shown in Fig, 1. Abundancia fault, while easily traceable throughout the lateral extent of the underground workings, is not always marked by the pres-

ence of vein material. The sheets of quartz and calcite, some minable and others almost barren of precious metals, occur in large lenticular masses. They have the strike of the fault plane, in general N. 3° W., and its dip of 65°-68° to the east. These lenses may extend vertically 200 m., while the greatest length opened up on any one level is about 80 m. Stoping widths of 1 to 6 m. are found. The lenses have a rake to the north with dip of 65°-70°, and may be found anywhere along the fault; seemingly, there is no rule as to the position where one may be expected, although large orebodies have been found near the junctions of Abundancia vein with the Portovelo and Cantabria veins. As a rule, the extent of the orebody is less than that of the quartz lens; or in other words, the lens is not all ore. Some lenses have more than one oreshoot, and others are entirely barren of ore. They apex at different altitudes, not all of them cropping at the surface, and pinch out at variable depths. The valuable portion of the vein may lie on the foot wall at one elevation, and on the hanging wall at another, the oreshoot proper consisting of overlapping lenses within the main lens. The difference between ore and waste is not discernible to the eye, hence the necessity for close sampling. True walls are present, with considerable gouge, especially on the foot wall. A false foot wall, characterized by a strong slip, is often found. Between it and the true foot wall, there is commonly from 0.30 to 0.75 m. of low-grade quartz. The walls will not stand after mining the vein, but break well back on both sides to parallel slips within the fault zone. Many tongues of quartz follow minor fractures out into the hanging wall away from the main body, but never beyond the hanging wall of the fault zone proper.

Twenty-six vein is short, only 100 m. long on the longest level. It has only one oreshoot with a depth of 150 m. In the upper part, it is a hard white quartz averaging 1 m. wide. Calcite appears in depth, where it grows wider and is of low grade.

Cantabria vein is a fracture making away from Abundancia fault to the northeast. It holds many orebodies, generally lenticular in form. The ore extends from 150 to 180 m. below the surface. The most intense mineralization is near the fault; here the company worked to a depth of 200 m., the length of the oreshoot being 85 m. and its width 4.5 m. Cantabria is the most base of all veins worked, galena, sphalerite, chalcopyrite and pyrite being present. Its walls are not well defined except at the junction with Abundancia fault.

Soroche vein makes into the Abundancia fault from the foot wall side; it has made several good sized orebodies. It has a banded structure of alternate ribs of hard and soft quartz with little calcite or sulfides. There are no walls. Stringers of quartz lead out from the main body into the foot and hanging walls and many overlapping kidneys of ore are found. The vein itself is badly broken and has numerous pockets of high-grade ore. The average stoping width has been 2.74 meters.

Tamayo vein has not been connected with Abundancia fault so far. Its ore lies in overlapping, slightly offset lenses of quartz that strongly resemble those of the Soroche vein. Some sulfides are present and it is strongly oxidized in the upper levels.

The Jorupe vein is composed of very hard massive quartz and the base sulfides. Several oreshoots have been developed, on one of which stoping has been started.


The ore deposits of fault planes and fissures of the Zaruma district have a dip of 50° from the vertical, usually to the east. The oreshoots vary in length and width and may apex or bottom at any elevation. They rake usually toward the north. There is no secondary enrichment and values bear no relation to topography. On account of thick soil and undergrowth, outcrops of ore are found with difficulty, except upon the sharp ridges and in the deep ravines, where the surface is worn to bed rock.

Preliminary exploration is generally carried on through adits. Depressions resulting from the cavings of old workings, dumps, etc., have often guided exploration. The abandoned workings are seldom found to contain ore; they were quite thoroughly mined to the water level, but they indicate where ore may be found at greater depths.

Prospect tunnels are driven on the vein or in the walls, depending on the character of the formation. Where the vein is wide, frequent crosscuts are made to the wall; and where it pinches, crosscuts are extended into foot and hanging walls in search of overlapping veins. Following the development of ore on one level, additional levels are opened 30 m. above or below the first, depending on the surface contour, and raises are put up at intervals of 30 meters.

Close sampling is the practice wherever a vein is exposed. Channel samples are moiled from the face and back 1 m. from each other, about 15 lb. being cut per meter. The samples may represent the whole width of the vein, or may be split as the character of the vein changes. The usual records and sample maps are preserved.

In the stopes, the faces are frequently sampled under the direction of the mine superintendent for the purpose of directing the daily work. Once every 6 months, the entire stope surface is sampled and mapped just as the drifts are sampled, and tonnage estimates are made, using 2.9 tons per cubic meter. Because of the indefinite walls at many places, the tonnage recovered is usually larger and the values lower than estimated from moil sampling.


In the early Spanish days, mining was conducted in the customary crude manner. The ore was followed in depth to water level, usually about 100 ft., leaving low-grade ore for pillars. The ore was carried out, the men using chicken ladders, steps cut in the walls, or built of masonry. The Spaniards left no ore that can now be mined at a profit.

The first work done by the South American Development Co. was through adits, the principal working adit being known as A level. Many large oreshoots were found above this level and worked out with filled stopes on the rilling system. The usual procedure was as follows: After the development of an orebody by drifting, a lateral was driven parallel to it about 6 m. in the foot wall, and crosscuts were broken through from the lateral to the drift at 20-m. intervals. Raises were then driven through to the surface from alternate crosscuts. The orebody was next silled off from wall to wall, as high as the ground would stand without timbering. Rills were then started from the raises, and the ore stoped above the fill which was introduced from mill holes on the surface. The broken ore was shoveled from the toe of each rill and later was handled through chutes built in the crosscuts. Costs were low by this method.

Some time later, the shrinkage system was adopted. This change was hastened by the fact that the stopes were located farther and farther away from the surface, making it increasingly difficult to get the cheap surface fill. Moreover, this fill was not altogether satisfactory. It was composed of surface wash, badly weathered, mostly very fine and not possessed of sufficient body for good fill. As it was worked into the stopes in quantities, a mill hole was formed around the top of each filling raise or chimney. During the rainy seasons, these mill holes had to be bulkheaded off to hold back the accumulated mixture of surface wash and water.

In order to work the deposits on the Abundancia and Portovelo veins lying below A level, the American shaft was sunk; it is collared in the hanging wall a short distance south of the intersection. This location was advantageous because it was close to many large, high-grade stopes on both leads, but disadvantageous because of the temporary loss of good ore left in the shaft pillars. At this time, stoping by shrinkage reached its maximum development. Shrinkage stoping of the narrower oreshoots on the northern end of Cantabria vein was a success. The walls were firm, standing very well, and the ore suffered but very little dilution through the admixture of waste. On the wide orebodies of the Cantabria vein near its intersection with the Abundancia fault, and on the ore-bodies of the Abundancia and Portovelo veins, shrinkage stoping was not so successful. After drifting and opening the veins to their full width, the stopes were either cut out for a height of about 4.57 m. above the rail, timbered, and breaking carried up on the timber; or box-hole raises were driven for chutes, the raises being lengthened both ways on the strike of the vein, connected a short distance above the level, and the stope carried up from them. This system was efficient as far as low breaking costs were concerned, but a large percentage of the ore in any orebody was lost. In the first place, it was impossible to delineate clearly the outlines of an orebody as irregular as these; and, in the second place, much ore was lost through caving of the walls in great slabs so that the ore was hung up and left behind in the stopes. Even when the walls did not cave, there was a loss in the ore that clung to the foot wall when drawing the stopes. There was also dilution of the ore through admixture of waste from the walls.

In late years the shrinkage system of stoping has been abandoned almost entirely, there being a reversion to filled stopes of the horizontal cut with back-fill type, or the modified rill with waste obtained from development on the upper levels or broken in waste raises at the head of the rill. Costs are higher, because of the extra cost of breaking and handling fill and the extra handling of ore in the stopes, but a close examination of the records shows that up to 40 per cent, more ore is obtained from a given orebody. The ore is also kept much cleaner, as considerable waste is sorted out and left in the stopes for fill. In high-grade stopes, however, there is a chance for loss through the mixture of fines with the fill. After blasting rich ore, it has paid, in some instances, to remove the upper layer of fill, to a depth of 0.61 m. (2 ft.) and send it as ore to the mill. The development of the Soroche mine, or that part of the property lying adjacent to the northern end of A level and extending about 152 m. to the surface above, which has taken place during the last four years, is chiefly responsible for the change in practice. The ore-bodies in this mine are characterized by the lack of definite walls, the overlapping lenticular structure being decidedly pronounced, and the walls extremely shattered and heavy. Many of the stopes must be closely timbered and then filled tightly as soon as possible to prevent caving. The finding of so much ore in the walls led to the adoption of this method on the other veins with similar results.

Previous to the installation of the modern cyanide plant, which began operations in April, 1919, the milling practice was straight amalgamation, followed by a rough classification and cyanide leaching of the sands. The extraction was low, as well as the tonnage. For some years, selective mining was carried on and the mill heads kept at $18 or better, so that operations might be carried on at a profit. This resulted in taking the cream of an orebody, with much lower grade material left unbroken as pillars and on the ends of the old stopes. A considerable part of the present tonnage is derived from working the edges of these old stopes; the cut-and-fill system is admirably adapted to the extraction of this ore. The operation, on the whole, is similar to that of any cut-and-fill stope, except that the old adjacent stope must be filled as well as the actual working stope; see Fig. 6.

Mining Methods

Development Plans

The combined Portovelo and Soroche mines are now worked through a shaft and two adits. The collar of the American shaft and A level are at the same elevation. The levels above A level, known as B, C, D, etc., are driven 30 m. apart. Mining is being done on D and E levels, while A, B, and C are on development work. Ore from above A level is passed through rock chutes to that level and transferred to the mill by mule train. Waste is used in filling stopes either above or below A level.

Eight levels are turned off from the American shaft at intervals of 30 m. each. The shaft has two compartments down to the seventh level, then three to the ninth, and two below to the depth of 320 m. Sinking will continue to a depth 343 meters.

The greatest lateral development is on the third level, which has been opened up for 420 m. to the south of the American shaft and for 920 m. to the north. The levels of the Soroche mine overlap those of the Portovelo mine (American shaft) on the south and extend 360 m. beyond the north face of the third level. Increased depth is gained by following the Abundancia fault to the north, the country rising in that direction. The combined length of the lateral openings of the Portovelo, Soroche, and Jorupe mines, including the outlying prospects, is over 35 kilometers.

The usual method of development is to drive a crosscut, to the vein to be developed, from the American shaft, in the case of the Portovelo mine, or from the working raise in the Soroche, perpendicular to the Abundancia fault, then turn off drifts both north and south. Drifts are run on the foot wall with crosscuts to the hanging wall at frequent intervals, in high-grade ore; and at greater intervals in low-grade ore or waste. When the faulted area is barren, laterals are often run paralleling it, with occasional crosscuts to the fault for exploration. This avoids the necessity of timbering, which must always be done when following the faulted area.

When an orebody is outlined, a raise is generally driven to the level above, for the purpose of further developing the block, for ventilation, and to serve as a working raise for stoping. The method of stoping to be used is then decided. If the orebody is to be stoped through a lateral, this is driven in the foot or hanging wall, and chute raises are broken through; otherwise the stope is cut out or box-holed over the drift for stoping by shrinkage or by the cut-and-fill method. During the last ten years, about 25 tons of ore have been developed for each meter of development.

Nearly all development is done by contract. The contractor pays for all supplies, including timber, but excepting tools, drill steel, oil and air. Track and pipes are laid by the company; it also pays for the timbering in headings.

Sinking, Stations, Pockets, Etc.

Native labor is extremely inefficient on sinking, 10 m. per month being about the maximum attainable. For drilling No. 55 Clipper, No. 95 Waugh drills, and B.C.R. Jackhamers are used. Blasting is done with 1 in., 60 per cent, gelatine, using ordinary fuse and caps. Local timber will not withstand the shock of electric blasting and delay-action primers are too complicated for the native workmen.

Ordinary shaft sets are used; 7-in. timbers being placed on 6-ft. centers. Formerly bearers were of squared timbers hitched into the walls. At present, bearers are made of concrete, poured in place, and reenforced with old rails and twisted rods. The hoisting compartment is lagged throughout on all sides with 3-in. plank; the other compartments are lagged only where the ground requires it. The guides are of 4 by 6 in. timbers, with notched lap joints, and are lag-screwed to the end plates and centers. Two small auxiliary hoists have been used for sinking, the dirt being raised in buckets to a pocket at a level above. A No. 6 Cameron sinking pump handles the water during sinking operations.

The stations at the levels are small, no car capacity being required. Ore pockets up to 175 tons capacity and waste pockets of 10 to 15 tons capacity are provided at each level. Grizzlies, made of 30-lb. rails and having 1-ft. square openings, cover each pocket.


The drilling machines used in the headings are the No. 21 Turbro with 1¼-in. round hollow steel and the No. 50 Clipper and No. 93 Waugh with 7/8-in. hollow hexagonal steel. The Turbro is too heavy for the average native and is only used in very hard ground. Average advance is about 20 m. per month. There is nothing peculiar in the method of advancing the faces, except that in the hard veins, consisting of quartz and massive calcite, a larger number of holes is necessary; the calcite makes the ground tough and hard to break. Headings are driven 5 by 7 ft. in the clear and the average round pulls 1.1 to 1.2 meters.

In addition to the machine work, over one-half of the total advance is made by hand, chiefly because a great saving on compressed air is made, the compressor capacity not being sufficient to do all of the necessary development by machine. Hand work is also cheaper from all other angles. Single-jackers are paid lower wages than machine-men, explosives are used more efficiently, and there are savings on oil, piping, drill steel, drill repairs, etc.

Raise Practice

Most of the raises driven are stoping raises for blocking out the ore and for ventilation; later these serve as stope manways. Other raises are driven solely for ventilation and for permanent manways or safety exits, or for transfer raises for ore or waste. All raises in waste are driven as two-compartment raises; those in ore may have two or three compartments.

Drilling is done with No. 16V Waugh stoperor No. 71 Waugh stoper; 50 to 35 per cent, gelatine is used for blasting. An advance of 50 ft. per month is considered good work. A Little Tugger hoist is part of the equipment for all raises. Two-compartment raises are carried in the country rock, 11 by 5½ ft. over all. These are timbered as raised and lagged with polls or planks. Manway platforms are placed at intervals of 30 ft. The ends of the timber, as a rule, are not notched, cleats being used instead of notches. Centers are placed every 5 ft., and the rearing, or chute lining, is of 3-in. plank as before. Bulkheads are carried over the manway in all cases.


The shrinkage method of stoping was formerly used extensively, but as the walls are unsuited to its use, it has now been abandoned almost

entirely and will not be considered here, except as to the first steps which are practically the same for the cut-and-fill method that is now commonly used. The stopes are either cut out for timber, or box-holed for chute raises and manways; see Fig. 2. With the first method, two cuts are taken out of the back of the drift from wall to wall, leaving the back of the cutting-out stope about 16.4 ft. above the rail. The muck from this first operation is then cleaned up and timbering started. When the veins are narrow, sets of caps and posts are placed, the caps reaching from wall to wall. On the wider veins, timber cannot be obtained long enough for this purpose. The sets are placed and blocked temporarily, then pole lagging is placed back of the posts of all sets, and the space between the lagging and the walls filled with waste to the level of the caps. Sets are placed 4½ ft. center to center.

Chute mouths of double 3-in. plank are built at every third set. Two manways are generally carried up with the stopes, either timbered in the usual manner or cribbed. When a stope is started by box-holing, the raises are spaced at intervals of 16.4 ft. They are opened 5 ft. long on the level, this length being increased in both directions with each succeeding round, until they are connected about 16.4 ft. above the back of the level, leaving triangular shaped pillars between the chutes.

Practically all of the present stopes are worked by the cut-and-fill method. They are either carried on timber or worked through stoping laterals. In the first case (see Fig. 3) the stope is cut out and timbered in the same manner as a shrinkage stope, except that chutes are built in every fifth set. Cribbed chutes and manways are carried up through

the stopes. With soft ores, cribs of round timbers, unlined, are sufficient; with hard ores, the unlined cribs wear rapidly and are a constant source of trouble. Cribs of squared timbers lined with 3-in. plank are better for this class of muck. The round, unlined cribs are carried with the dip of the vein, no offsets being necessary. Cribs of square timber, plank lined, are carried up vertically with offsets, to facilitate the nailing of the lining in a secure manner. When stoping from a lateral, the lateral is driven parallel to the orebody at a distance of 5 m. in the wall; see Fig. 5. The foot wall is the preferable location, but in the case of parallel veins separated by a small width of country rock, one of the laterals must be driven in the hanging wall. The chutes are raised from the laterals to the stopes each 8 m., and are carried up with cribbed timbers. When working stopes in heavy ground, timber as well as filling must be used; see Fig. 3. Stope timbering is done after the manner of drift sets, with caps from wall to wall. The posts are stood on footboards placed on the fill, and the sets are 8 ft. high; they are spaced from 5 to 8 ft. center to center. Stringers of 6-in. round sticks are laid from cap to cap, about 1 ft. apart, and blocked down from the back. Filling closely follows breaking in these timbered stopes, the fill being kept close to the back. The posts are lost in the fill, but the caps and

stringers are recovered to be used on the next cut. Breaking of ore and waste in the cut-and-fill stopes is done with stoping machines of the No. 16V or No. 71 Waugh type, or with jackhammers of the Waugh No. 95 or Ingersoll-Rand BCR-430 types. As a rule, one or more horizontal banks are carried across the stope followed by filling, which is obtained by driving raises and crosscuts into the foot and hanging walls. A small

amount of fill is also obtained by sorting the ore in the stopes. In some of the stopes, part of the fill is obtained from development waste on the level above, this waste being dumped through the stoping raise. When waste is obtained in the latter manner, the stope may be rilled to this raise; see Fig. 4. Other rills are made by breaking waste in the end of the stope by raising on the barren vein, and then rilling to the waste raise; see Fig. 4. These raises serve the double purpose of providing waste and developing the vein. It is important that a part, at least, of the fill should be broken in crosscuts or raises in the walls, as these workings have the added value of being good prospects. They need not be long, as the overlapping lenses are always found within a few meters; often they are separated by only a shell of waste. The advantages of the cut-and-fill system are 100 per cent, recovery of the ore in any oreshoot; the ore is kept cleaner, as sorting can be carried on in the stope; and there is no admixture of wall rock, as when pulling a shrinkage stope.

The drilling in all stopes is done on a contract basis, the machine men receiving a fixed rate per meter of hole drilled. All holes are spotted by native stope bosses, and measured by them before blasting. Bonuses are paid to those machine men that work steadily and drill more than a

given meterage per month; the bonus takes the form of a higher rate for all meters drilled. Loading and blasting is done by the drillers under the supervision of the stope bosses.

An analysis of the cost sheets will show that, for the period from 1902 to 1908-9, the direct mining costs, including development, amounted to S/9.40 — S/9.50 per ton milled. During this time, the average yearly development amounted to 925 m. at a cost of S/57 per m. This period represents that in which the large surface stopes were mined by the rilled cut-and-fill method. The ores near the surface were softer, much more easily and cheaply mined, and the surface fill was cheap. Moreover, many of the stopes were above A level, so that pumping and hoisting costs were low.

During the period 1909-19, the direct mine costs were S/8.00-S/8.10 per ton milled; this includes an average yearly development of 1530 m. at an average cost of S/49 per meter. During this time, when shrinkage stoping was at its height, therefore, the mine costs were much lower in spite of increased development. A great part of the tonnage came from the American shaft, so that hoisting and pumping charges were inevitably higher. The cost of supplies during the latter years of the period, the years of the war, also was high. The bulk of the tonnage, however, came from the enormously large bodies of ore lying within a short distance of the shaft. A few stopes were sufficient to give the desired tonnage, and close supervision was possible; this fact accounts for cheap breaking, tramming, and total mining costs. The great disadvantage of this cheap mining lies in the high loss of broken and unbroken ore left behind in the caves that resulted when the stopes were pulled, or overlooked in the walls during actual mining.

From 1920-22, inclusive, the average direct mining costs were S/12.40-S/12.50 per ton milled. This seems to be a large increase, but can be accounted for in several ways. Shrinkage stoping had been almost entirely discontinued, the cut-and-fill method, by back filling or with the modified rill, having been adopted once more. This method undoubtedly requires more labor, resulting in increased costs. A minimum increase of 20 per cent, in the wages of all mine labor also went into effect at the beginning of this period. The costs of all supplies have continued to be high. Moreover, there has been a big increase in the quantity of development work done. In this period, an average yearly development of 3796 m. at a cost of S/51 per m. was done. While the cost per meter compares favorably with the costs of the preceding periods, the greatly increased meterage shows its effects in the total mine costs. Most of the big orebodies near the shaft had been worked out before this time. The tonnage came from smaller bodies at long tramming distances, or from the edges of worked-out and caved stopes. A great deal of filling had to be put into these old stopes before the remaining ore could be mined. Mining of the orebodies in the heavy ground of the Soroche mine has necessitated the use of a great quantity of timber, the first cost of which, together with its maintenance, constitutes a big percentage of the increase in total costs. More bosses were needed to cover the scattered workings and even then the supervision was not as close. The big advantage of this method of mining lies in the greatly increased tonnage gained from an orebody. All things considered, the cut-and-fill method of mining, with the stopes worked through stoping laterals, seems best adapted to the orebodies of the district. A tabulation of costs is as follows:

Records of Unit Production

The following figures have been obtained from one year of normal operation (1922).

Tons per man per hour for all underground and surface labor exclusive of office………………………………………………………………………0.0534
Man-hours per ton for all underground and surface labor exclusive of office …………………………………………………………………………………18.73
Tons per man per hour for all underground labor………………………………..0.0614
Man-hours per ton for all underground labor………………………………………..16.28
Tons per man per hour for all surface labor…………………………………………..0.4087
Man-hours per ton for all surface labor…………………………………………………..2.45

Classification of Labor, Expressed in Percentage of Total

The percentage labor turnover is unknown, but it is very high.

Records of Unit Supplies

Pounds of dynamite per ton of ore mined………………………………………..2.24
Pounds of dynamite per meter of development………………………….15.45
Pounds of dynamite per foot of development………………………………..4.71

Safety and Welfare Work

No safety organization of any kind is maintained, this work being left in the hands of the mine bosses. Due to the inability of the natives to take care of themselves, close attention must be paid to all working places. Obviously, the accident rate with this class of labor is high; the only surprising feature of it being that serious accidents are not more common. All serious accidents are reported to the local authorities in compliance with a compensation law passed in 1922. The company, however, has had an agreement with its employees for several years which is much more favorable to them than the compensation law.

The company hospital is the most modern and best equipped in the Republic. An American surgeon and trained nurse, together with an Ecuadorian doctor, are in attendance. All employees and dependent members of their families are given free treatment and medicines for all classes of cases without restriction. Many outside cases are also treated at the company hospital, this work being done for a nominal fee or free of cost. All prospective employees are subjected to a hospital examina¬tion, as a result of which they are passed for work or rejected as unfit. Those qualifying for employment are treated for hookworm, and classified as to their fitness for surface or underground employment. Many sanitary measures have been instituted, and strict enforcement of sanitary regulations is maintained.

A large boarding house is operated by the company for the benefit of the Ecuadorian employees, they having the option of boarding with the company or receiving an allowance of 50 centavos per day above wages. This boarding house is operated at a loss, using the amount of 50 centavos per day per man as a basis to figure from. Dwelling quarters at a very low rent are also provided for all employees who wish them.

The members of the staff are quartered in frame and concrete houses; married men being furnished with houses, and bachelors with single rooms. A boarding house is also operated for the benefit of the single men on the staff. House rent, lights, fuel, water, etc., are furnished free, as well as ice and distilled drinking water. Recreation is provided for in the shape of a clubhouse, tennis courts, baseball field, and swimming tanks. There is also a company stable, where staff employees horses and mules are cared for at a very small charge.


Rudolph Emmel, Quayaquil, Ecuador.—All the labor is very inefficient and the bonus system is the only way we can get anything at all. Drilling is done by paying so much per meter of holes. Drifting is done by contract on the basis of so much per meter.

George A. Packard, Boston, Mass.—That accounts for the apparently very low efficiency in man-hours per ton. At Cornucopia and at Jarbidge, the drillers in stopes practically average 2 tons per man per hour, or 1½ hour per man per ton; whereas at Zaruma the average is 0.05 ton per man-hour or practically 19 hours labor. Later the author shows that machine men and helpers on developing and stoping make about 32 per cent, of the labor, which would mean 60 man-hours per ton of ore or twelve times the requirement at Cornucopia and Jarbidge. This figure includes both the tons per man per hour in the shrinkage stoping and in the other; what is the production on shrinkage stoping alone?

Rudolph Emmel.—I have no figures on shrinkage stoping here; the figures given include all of the developments and we are doing an abnormal amount of development work. For a mine producing 225 tons, 4 km. a year is a large amount and tends to bring down that figure. However, the labor is very inefficient; it is incredible the small amount of work that one of those men can do in a day.

Fred Hellmann, New York, N. Y.—My experience in the mines of South America has been entirely different from that. The Chileans would compare favorably with drill men in any part of the world. I consider the Chilean one of the best drill men in the world.

R. M. Raymond, New York, N. Y.—It is largely a matter of the make-up of the men. Mexicans are rather fond of mechanical work. They make excellent men for running drills, do not mind dust, and work fairly steadily. They are good at any kind of machinery, even running hoisting machines. They are showing themselves adapted to such work in a surprising manner. How do the Bolivians compare with the Chileans?

Fred Hellmann.—The Bolivians cannot compare with the Chileans. They are not nearly as intelligent. The Inca race, as you know, was conquered by the Spaniards, and their present conditions testify to the inaptitude of the race and its weakness.

R. M. Raymond—How do the Chileans compare with the Kaffirs as to man power?

Fred Hellmann.—The Kaffir must be taught everything when he enters the mine, while the Chilean is much more intelligent. The Chileans are essentially a mining people. You can induce the Kaffir to do things, and under certain conditions he is a wonderful worker. The Kaffir can be taught to run a machine but it will take him much longer than the Chilean. I do not know how you could state it in percentage, but you would get possibly twice as much work out of the Chilean miner as out of the Kaffir.

In Bolivia, there are large mines but the results obtained could never be had by our methods of mining. If an American should go there with the intention of lowering the cost of getting out the ore by using more modern methods, he would soon find his mistake. They mine about as cheaply as is possible, though their methods are somewhat primitive. Access to the mine is usually given by a spiral stairway—the so-called Boca Mina—up which the ore is carried on the backs of the workers. It is hard to believe that you cannot improve on that, but the supply of labor is so great and the comparative cost of it so small that when you figure the cost of installing and operating machinery in a country that is not mechanically inclined, you find that the present is the better method. Of course mining 50,000 tons of ore a day is another kind of mining; that is done with steam shovels.

Arthur Notman, New York, N. Y.—The report of the Union Miniere du Haut Katanga for 1923 indicates roughly the number of men employed by the company for that year. There was no classification of labor or segregation as to operation and construction given in the report, but taking the whole number of employees as given, an output of 35 lb. of copper per man per day is indicated. Somewhat similar figures for the Chile Copper Co. indicate, for 1923, an output of 115 lb. per man per day. The output of the porphyry copper mines in this country runs from 150 to 190 lb. per man per day. Even after making due allowance for perhaps an extraordinary proportion of the labor on construction work in the Congo, it does seem as though the much higher grade ore and lower wages were pretty well offset by the low labor efficiency.

Fred Hellmann.—You have an entirely different type of labor in the Congo. The northern tribes and the tribes of central Africa are not nearly as husky and strong physically as the natives of the more southern parts. They are much weaker and much more subject to disease and death when they do work, especially under changing climatic conditions.

C. F. Jackson, Skouriotissa, Cyprus.—Labor is our chief problem; the technical problems would not be so bad if our labor did not keep undoing our work. For example, the other day coming out the main intake airway of the mine, we found a canvas had been stretched across the portal by some workmen employed nearby who wanted to shut off the chill air; sometime ago we sealed off a small fire but not long after some native opened a 9-in. hole in the wall with a pick. These people tear out ventilating doors, remove brattices, close doors that are supposed to be kept open and leave open doors that should be kept closed, and pick into pillars that are necessary to the maintenance of important airways and roadways. With good labor the work would be comparatively easy, but the Cypriot is the worst workman in the world to deal with; there is no punishment that he cares anything about so that it is next to impossible to discipline him.

Rudolph Emmel (Author’s reply to discussion).—The figures of 2 tons per man per hour at Jarbidge compared with 0.05 ton per man per hour at Zaruma is a comparison of the output of men directly engaged in stope drilling at Jarbidge with the output of all men connected with the work of the mine, underground and surface, at Zaruma. The figures used for comparison should be 0.4 ton per man per hour at Jarbidge (mining by shrinkage) against 0.05 ton per man per hour at Zaruma (mining by cut-and-fill methods). The mines at Zaruma also are undoubtedly carrying a much heavier burden in the nature of development work necessary than those at Jarbidge.

As to the efficiency of the Chilean miner, one must bear in mind that the Chilean is working under entirely different conditions from those that prevail in Ecuador. The Chilean is the product of the temperate zone, is physically and intellectually much superior to the Ecuadorean miner, and has had the benefit of a much closer contact with American and European labor and methods. We all know that the Mexican, when in competition with American labor, has developed to the point where he works side by side with miners of all nationalities in the United States, and has supplanted the higher priced labor to a great extent throughout the West. Chile is a very cosmopolitan country, is one of the most advanced of the Latin-American nations, and the Chilean workman has undoubtedly benefited therefrom. The Ecuadorean workman physically is a poor specimen, he has generations of hookworm and malaria and other diseases back of him, is working in hot, poorly ventilated mines in a tropical country, and has come into no competition or association with the outside world.


Geology and Mining Methods

Latouche, or the Beatson plant of the Kennecott Copper Corpn., is located in the Prince William Sound district of Alaska about 80 miles west of Cordova and 60 miles from Seward.

Ore was discovered and claims located on Latouche Island in July, 1897. The mines were worked in a desultory manner until 1910, when the property was acquired by the Kennecott Copper interests. The first shipment of ore was made in 1904; only the higher grade ore was mined and no attempt was made to treat the ore until it was taken over by the Kennecott company. Since that time the mine has been developed to produce 1500 tons a day and a mill has been erected for treating this tonnage, using flotation entirely. The orebody is more or less lenticular in shape with a maximum width of about 280 ft. and a length of about 800 ft. The southern end of the lens is split by a horse of waste for about 400 ft. The hanging-wall limit is a well-defined fault associated with a band of pyrrhotite, having an average dip of 60°. There is no defined foot wall, the value of the ore governing the limits of the mining; in one part of the mine, however, it is defined by a minorfault. The orebody is in a shear zone of the country rock of graywacke and slate. The principal mineral is chalcopyrite associated with pyrite and quartz.

The mine is situated a few hundred feet from tidewater; the mill is on the beach and the ore is hoisted direct from the mine through a vertical shaft into the mill bins. Until the past year practically all the mining had been conducted in open pits. During the past two years, the ore above the 200-ft. level has been developed and a system of stoping devised to recover this ore; however, this system has not been in use long enough to give any definite results as to costs or efficiency. It consists, principally, of dividing the orebody into stopes, across the full width of the orebody, 70 ft. wide with a pillar 30 ft. wide between the stopes. Raises are driven through the stopes at intervals of about 60 ft. and all the ore is broken by drilling from these raises; in other words, similar to a shrinkage stope, doing the drilling from the raises instead of setting up on the broken ore as is customary. The ground is very much broken up by clay slips and seams running in every direction. Shrinkage stoping had been

unsuccessful because of the time necessary to bar down and the impossibility of making the back safe, because of these slips. On the upper level, in the higher grade ore, some square setting was done, but the average grade of the ore precludes the use of this system throughout the mine.

Some exploration work has been done with the diamond drill, but for the greater part drifts and crosscuts have been driven to explore the ground. In sampling, all crosscuts and drifts, as well as raises, are sampled in 5-ft. intervals and a double groove 1½ in. deep by 6 in. wide is cut in each working place. It has been found that results, even with such large samples and double sampling, are approximately 20 per cent, too high. It is estimated that 12 cu. ft. in place and 20 cu. ft. broken produce a ton of ore.

Underground Mines

Mine Openings, Shafts, or Tunnels

The mine has two entrances, a vertical shaft and a main-level tunnel. The main-level tunnel is 970 ft. long from the portal to the central point of the orebody. It is about 7 by 7 ft. in size. Where timber is used, the measurements are 6 by 6½ ft. inside the timber. The shaft is 360 ft. deep with three compartments, each compartment being 5 by 5 ft. inside the timbers. Loading pockets are located below the main level and below the 200-ft. level.

Underground Development Plans

The 200-ft., 150-ft., and main levels are shown in Figs. 1-3. The curves on the 200-ft. level have a radius of 35 ft.; on the 150-ft. level a radius of 30 ft. is used. All the main haulage tracks have 35-lb. rails; on the 150-ft. level, 16-lb. rails are laid, this track being used only for hauling supplies. Track grade is carried uniformly at 0.5 per cent.

The main compressed-air line is 6 in. No. 20 gage galvanized ventilating pipe, 10 in. in diameter, is used.


Drilling and Blasting.—The following is a list of drills used, together with the size of the steel:

The single-taper cross bit is used on all steel. Holes are tamped with mud enclosed in parafine paper cartridges, and firing is done with fuse and caps or with electric blasting caps and magneto.

The explosives used are 40 per cent, gelatine and 40 per cent. Red Cross manufactured by the du Pont company.

Tramming and Haulage.—Granby 3 and 4-ton side-dump cars are used, dumping direct into the shaft ore pockets.

Baldwin-Westinghouse storage-battery 4½-ton locomotives furnish the motive power. The rails are 35 lb., the gage 24½ and the grade 0.5 per cent.

Underground Storage and Dumping.—The underground storage consists of the two shaft ore pockets, each with a capacity of about 250 tons. The cars dump directly into these pockets, the ore then being drawn to the measuring pocket in the shaft and then loaded into the skips.

Hoisting.—The man hoist is a Coeur d’Alene Hardware & Foundry Co., electric hoist having a 36 by 36-in. drum driven by a 100-hp., 440-volt

motor. A single-deck cage with spring safety dogs and ¾-in. plow-steel rope is used.

The ore hoist is a Wellman-Seaver-Morgan double-drum geared electric hoist having 60 by 42-in. drums driven by a 150-hp., 440-volt Westing- house motor.

The cable used is 1 1/8-in- plow steel; the skips hold 4 tons and are operated in balance; the hoisting speed is 600 ft. per min.

Pumping.—The pumping is not a serious problem; the following electrically driven pumps take care of the entire mine: One Gould triplex pump, 70 gal. capacity; one Gould triplex pump, 275 gal. capacity; one Gould triplex pump, 80 gal. capacity.

Air Compression.—One Chicago Pneumatic Tool Co. compound compressor of 1730 cu. ft. capacity driven by a direct-connected, 2200- volt, 295-hp., synchronous motor supplies the mine with compressed air.

Ventilation.—Natural ventilation has been relied on but with increase in the amount of bulldozing on the 150-ft. level, attendant with increased mining underground, an electrically driven fan of 30,000 cu. ft. capacity is being installed. The ventilation system is shown by the dotted lines and arrows on the accompanying figures.

Lighting.—The mine is lighted by electricity, the circuit being 110 volts. Several types of globes have been used—carbon, mazda type B, and mazda mill type. The mazda mill-type 50-watt globe has been found the most satisfactory. Small carbide cap lamps are used by the miners.

Telephone.—The telephone is used for general underground communication. Electric signals are used in the shaft.

Records of Unit Production

For the year 1922, the total production was 274,863 dry tons (2000 lb.) ore milled. Tonnage broken amounted to 349,071 tons. The following data are based on the tons broken and not on the tons milled.

Work in the mine is done either on day’s pay or on the bonus system. Bonus is paid on footage driven, in development work, or on footage drilled, in stoping.

Bluff mining, 92,365 tons broken.
Tons per man per hour……………………………………………9.68
Man-hours per ton…………………………………………………..0.103

Stopes, 226,199 tons broken.
Tons per man per hour…………………………………………….9.30
Man-hours per ton…………………………………………………….0.107

Stope preparation and development in ore, 30,507 tons broken.
Tons per man per hour…………………………………………….0.661
Man-hours per ton…………………………………………………….1.51

No work done in rock or waste.

All underground labor:

Tons per man per hour………………………………………….1.76
Man-hours per ton…………………………………………………0.57

Surface force:


All mine labor;

Tons per man per hour……………………………………………1.67
Man-hours per ton…………………………………………………..0.60

Total labor turnover for the year 1922 for the entire plant was 248 per cent.; for the mine alone, 235 per cent.

Total labor cost, expressed in percentage of total cost of mining, was 59.1 per cent.

Records of Units of Supplies Used

Explosives used on ore-breaking on bluff ore was 0,3 lb. per ton; on stoping, it amounted to 0.3 lb. per ton broken.

All timber, lagging, and poles used amounted to 0.020 ft. B.M. per ton ore broken.

Power required was:


Coal Mining

The coal-mining districts of Washington are mainly west of the Cascade Mountains; Fig. 1. The mines are on the foothills of the slope, the lignite fields of Lewis and Thurston’s counties extending into the valleys west of the mountains. An exception is the Roslyn field, a small but important area on the eastern slope.

If western Washington is divided into three areas from north to south, the bituminous coal fields and most of the active operations will be found largely in the middle area. The southern portion contains lignites over a large area. In the northern portion, mining is practically restricted to one mine, the Bellingham, which is working a sub-bituminous coal seam near the city of Bellingham.

The degree of alteration that the coal of a particular field has undergone may be gaged roughly by the position of the field with reference to the Cascade Mountains. The lignites of Tenino, Tono, Mendota, and Castle Rock, on the railway line connecting Seattle and Portland, occur in a region of low relief, in which the Eocene coal measures have suffered only minor disturbances. Subbituminous coals occur in the foot hills, as at Renton, Newcastle, and Bellingham. Coals high in fixed carbon are found nearer the mountains, where the measures have been sharply tilted and folded, as at Black Diamond and Carbonado. The two typical anthracite fields are located still higher, in the rugged mountains where outcrops appear both in deep gorges and on ridges at elevations approaching 5000 ft. At Carbonado is found the only semianthracite seam worked on a commercial basis. The occurrence of this seam in a region of bituminous coal is due to the nearness of an igneous sill, which has changed its character from bituminous to semianthracite.

Purpose of Paper

On account of the various faults and dips and the varying nature of the wall rock, many difficulties are encountered in mining and winning the coal. The paper gives some of the methods used in mining and, in some detail, the conditions that affect the efficiency of these methods.

The paper does not discuss the geology of the district, except incidentally, as numerous reports have been issued on the geology; nor does it attempt to discuss the methods of preparation of the coal for market, although

this is an important part of the operations. Details of underground methods that have been tried and found successful are given as well as those that have not proved successful.

General Geology

The coal-bearing rocks of Washington are of Eocene age and correspond, in time of deposition, with the lignites of the Dakotas, Montana, and Texas. In Washington, however, mountain-building forces were active in late Tertiary time, faulting took place on a large scale, intrusive agencies were active in some fields, and much of the coal was changed in rank and structure. In general, the surface has a rugged contour in the different districts. The covering of glacial material masks the continuity and only a few of the fields are well marked. Because of this, the fields have generally been subdivided arbitrarily, according to the county in which they occur; and reports on the coal areas of the state have been differentiated in this manner.

All coals contain foreign material in the form of partings or binders, but the coal of Washington includes more of such partings than the coal of many other parts of the United States. The old swamps in which the coal was deposited were apparently subject to many floods which washed in mud and sand from the surrounding higher lands. These materials formed a shale, sandstone layer or parting in the coal bed, which detracts greatly from its value, for the partings must be removed before the coal can be marketed. Removal is generally difficult and expensive. This foreign material cannot be gobbed in the mine, on account of the heavy dips, but must be handled with the coal and eliminated at the cleaning plant.

Mining Methods

A study of the various factors involved in mining coal in the different counties shows that the state can be divided into two divisions based on the methods of mining. These two divisions are Kittitas, Thurston, Lewis, and Whatcom counties, which will be designated division 1, and King, Pierce, and Skagit counties, with a future development in Whatcom, called division 2. The production of the two divisions is approximately equal under normal conditions. The mining methods in division 1 do not present any problems that might be classed as unusual. They do not vary from those in many parts of the country and are not discussed here. As a whole they are flat workings, the coal seams are clean, the rock can be separated at the face and gobbed, and the roof is good. The workings of division 2 are on the high-dip seams of the state, where everything between walls must be removed in mining and the rock and accompanying refuse removed in cleaning plants on the surface. The walls, as a rule, are bad and the methods of working are diversified. The daily and yearly output per man in division 1 is greater than that of division 2; the payment of miners in division 1 is on the tonnage basis, while in division 2 it is mostly on the contract yardage basis. Practically all the coal in division 1 goes directly from the mine cars into the railroad cars, while all that of division 2 must be washed or prepared for the market. The coal of division 1 is used mainly by the railroads, while that of division 2 finds its market as a domestic fuel, for coke, gas plants, steamship, and miscellaneous trade. Although the production is smaller in division 2, more men are employed, the ratio being roughly 3 to 2. To show the methods of mining that depart somewhat from the ordinary methods, mines have been selected that, because of physical conditions, present problems that are typical of the district.

Mining Methods in Pierce County

Nature of Coal Seams Worked

The method of working steep seams described is practiced, with slight modifications, in the different beds of the Pierce County coal measures, especially where explosive gas is likely to be met. In general, the method is always employed where gas is found and where the roof and bottom are troublesome. It is used on all dips where the coal will run, that is, on dips to 65°; whenever a dip exceeds 65° to 70°, the working place becomes unsafe and the bed is hard to work. In such cases, chutes are driven across the pitch to give 45° to 50° pitch. The coal is mined by the chute-and-pillar system, for the character of the walls is such that mining by wide breasts is usually impracticable. The coals do not fire spontaneously. The seams worked by this method vary in thickness from 3 to 25 ft., although in Pierce County they do not reach the maximum thickness.

Method of Development

The gangway, Fig. 2, which is the intake airway, and the counter gangway, which is the return airway, are driven parallel. The counter gangway is up the pitch from the gangway and the stumps are from 20 to 50 ft. thick. Chutes are driven 6 ft. wide from the gangway to the counter on 50-ft. centers, from which point they are driven 8 to 10 ft. wide up the pitch to the top of the block. Lifts are arranged so that the chutes will be about 400 ft. long. It has been found impracticable to make blocks over 50 ft. in length, although the law permits a distance of 60 ft. Crosscuts are driven, varying from a hole just large enough to crawl through to 6 ft. wide. Between every other chute, a “half chute” 6 ft. wide is driven from the main gangway to the counter gangway. Where much gas is encountered or conditions warrant such action, half chutes are driven between all the chutes. Each half chute is a traveling way between the gangway and the counter gangway, and from it access can be made to two chutes at all times.

As shown in Figs. 2 and 3, a board brattice is carried from the counter gangway to the face of the chute. The inby compartment is used for a manway. The boards are nailed on the coal side to a line of props up the center of the chute; these props are set from 2 to 4 ft. apart, depending on the character of the walls. Hand rails are fastened to the props on the manway side and steps on the bottom are made by placing a prop on the upper side of the brattice props, setting one end in a hitch in the rib. Often a ladder is laid on the floor of the manway and nailed to props, which are placed on the floor at regular intervals to be used for steps.


The details of ventilation are shown in Fig. 2. The main ventilating fans are usually of the exhaust type and the main haulage roads are used as the intake airways. In all mines where explosive gas is generated, approved electric lamps are used by the workmen.

The chute between the gangway and the counter gangway is always kept full of coal, thus preventing leakage of air from this source. At every crosscut and on the counter gangway, canvas curtains 2 by 2 ft. are hung in the brattice: access across the chutes can be made through these openings at all times. At each crosscut, wings are built of boards from the bottom to the roof on the manway side of the chute, starting at the chute brattice below the canvas curtain and extending to a wing post set in the center and slightly back in the crosscut. These wings not only deflect the air into the crosscut, but prevent any coal from passing below the crosscut as it runs into the coal side through the canvas curtain, and eliminates danger to men coming up the chute.

Batteries or bulkheads are built at every other crosscut (if necessary, at every crosscut) and the chutes are kept filled with coal. This not only arrests the coal and reduces breakage, but aids ventilation and makes it safe to pass from one chute to another through the crosscuts. A chute starter runs coal from the batteries when necessary.


Much of the shooting is done on the solid, which requires that the opening shot be fired first and other requirements of the law complied with. Holes are drilled with ordinary hand machines, which are either post machines or breast augers, although drills of the auger type, such as the Waugh “Ninety” are finding favor where compressed air is available. Power drills are a distinct advantage, especially in a thick seam or a high chute where it is difficult to set up a post machine and the coal is too hard for a breast auger. In some instances, half of the miner’s time is used in drilling and this time can be reduced considerably. Compressed air in the place also aids ventilation. In most cases only one miner works in each chute. In one seam averaging 4 ft. and a coal of average hardness, the rate of advance in 8-ft. chutes is 8 ft. per shift of 8 hours.

Drawing the Pillars

In Fig. 3 is shown the method of pillar drawing. Four chutes are usually driven to the level above in advance of the pillars, three of which are usually working at one time. The pillars are drawn on the inby side of the pillar as follows:

Pillar No. 30 is finished. When a pillar is to be drawn, the crosscuts of the chute are well timbered and a cog, 1 in pillar 33, is built on the under side of the upper or crosscut as the case may be; often right up against the face or top of the workings. This cog is placed as close as possible to the inside rib of the chute. The corner A of the fourth block is then worked off and, as soon as space permits, a second cog 2 is placed 4 to 5 ft. from the first. A temporary battery of props is built above this cog to aid travel across the face and to prevent anything falling from above and injuring the men below or knocking out the props. The cogs and batteries are built of 6- or 7-ft. props, depending on the thickness of the seam.

Attacking a block in this manner is called “taking off the angles.” The process is continued until half of the block is worked off, as in block 2 of pillar 31, leaving what is called the “tail.” The same course is pursued in the next block below, as in 1 of pillar 31, after which the tail or half of the block above is worked and run out. When a tail is to be run out, as shown in block 3 of pillar 32, part of which has been run out, the temporary battery above the cogs is replaced by spouts or wooden chutes. The coal is then run out between the cogs through the spouts, after which a permanent battery is placed above the cogs. In this way the pillars are drawn to the first block, from which only an angle is taken off as shown in pillar 30, and a cog and permanent stopping are placed in the chute neck. By working as shown, the roof breaks above the cog lines and is held up by the cogs. With good roof, it is not unusual to recover 90 per cent, of the coal.


The timber is cut on the surface, at the mines; the lagging is split or sawn timber. Split lagging is the best in gangways, heavy ground, and in chute batteries or bulkheads. Ordinarily, the gangway sets are made of heavy poles, set 6 ft. apart, and consist of two legs and a collar, above which split lagging must be placed to safeguard against the coal or shale roof. Ties are usually made of sawn lumber, the size and spacing varying with the gage of the track.

Timber, where possible, is brought down from the surface to a gangway or top counter gangway above the active mine workings. Chutes are driven from the working level to this gangway, or top counter, and one of these is used as a timberway from which props are taken to all parts of the pitch.

On the pitch, props are set 3 ft. or more apart, depending on the walls, above which cap pieces are placed if the roof is bad, otherwise they are set in a hitch. They are slightly underset on the top or bottom, depending on which wall is apt to move ahead of the other; if the bottom is bad, sills are used.


Where the mines are opened as slopes, the hoisting is done by means of steam or electric hoists. In small mines, unbalanced or single-rope hoisting is used; but in the large operations, the partly balanced or two-rope hoisting system is common. The loads hauled vary according to the dip and size of the mine.

Mine cars of from 1 to 2 ton capacity are used and, except for some short hauls when mules are used, haulage on the gangways is performed with electric locomotives of both the trolley and storage-battery types.

As the cars are loaded from the chutes on the gangway and the coal spills more or less over the sides of the car, it has been found economical to place the drainage ditch on the hanging-wall side of the gangway.

Crosscuts on Extra Heavy Dips

When the dip exceeds 65°, it is advisable in most instances to drive angle chutes where the chute-and-pillar system is used. The question arises whether the crosscuts should be driven at right angles to the chutes, that is, angle crosscuts, or driven on the strike of the seam. The general practice favors the level crosscut, regardless of whether the mine is gaseous or not. Probably the determining factor in working a steeply dipping seam is the distribution of timber and material on the pitch, and this is greatly facilitated when the crosscuts are driven on the strike of the seam. Other advantages of the level crosscut are better escape-ways for men in pillar drawing, better opportunity to keep the places clear of gas, and greater ease in traveling. A disadvantage is the shoveling necessary when driving the crosscuts. The advantages of the angle crosscuts are that no coal must be shoveled and if desired they can be used as chutes at any time, especially when drawing the pillars. However, the crosscuts must be well covered so that no coal runs into them. There are times when it is a distinct advantage to drive angle crosscuts even when the chutes or breasts are driven on the full pitch when the dip is less than 65°; this is described under the Newcastle operation.

Unusual Methods Employed by Carbon Hill Coal Co.

The mines of the Carbon Hill Coal Co. are located at Carbonado, Pierce County, Washington, on the Northern Pacific Ry., 53 miles from Seattle. They are typical of the district. The mines, because of folding and faulting and the topographic and physical conditions of the seams and walls, present many interesting and complex features of operations; all angles of dip are found and the methods of working the seams vary accordingly. On this property will be found all seams of major importance that have been developed in the district.

The Carbonado mines have been the largest producers in the district; their total output has been over 7,500,000 short tons. Normally, about 1000 tons per day are produced. The product of the mines is a high-grade bituminous coal ranking with the best in the northwest.

Method of Entry

All the Carbonado mines have been opened as water levels starting above high-water level in the Carbon River canyon. This canyon, in places, is 50 to 80 ft. wide and is about 400 ft. deep near the openings. The river has cut through the sedimentary rocks and has exposed the seams.

Twelve workable coal seams appear in this series, first identified by Doctor Willis as a continuation of the series at Wilkeson, Burnett, and Spiketon. The tops of the seams have been eroded and the outcrops are usually covered with gravel. Because of the available coal above water level, only two seams, No. 12 (Miller) and No. 11 (Wingate), have been worked to any extent below water level. All the Carbonado seams that have been worked are shown in Fig. 5.

Structural Geology

The structure of the Carbon Hill Coal Co.’s property is divided by the Willis fault into very distinct parts. North of the fault, the structure

is one of intense folding and thrust faulting; south of the fault there is only a monoclinal dip to the west. In this section, the whole series appears and is disturbed only by the Miller fault and a small fold in the southwest.

Dikes and sills of igneous rock have been intruded into the measures in sections 9 and 16. Bed No. 8 has been entirely burnt out by small sills and dikes. The presence of anthracite coal in south No. 3 seam, the equivalent of Big Ben, is due to the intrusion of a large sill into the measures below. Fig. 6 gives sections of some of the seams.

Working a Steep Coal Seam by the Longwall Method

The methods of working seams No. 6 and No. 12 (Miller) at Carbonado were changed to a modified longwall method in preference to the breast-and-pillar and chute-and-pillar methods adopted by the other mines.

The writer had occasion, in 1916, to note the successful operation of the longwall method as applied to a short lift in seam No. 6, and to seam No. 12 during a period of two years, when the level on which the system was being worked was worked out. Several attempts were made to work seam No. 12 by the breast-and-pillar and chute-and-pillar methods, but they failed because the roof could not be kept up. Under J. F. Menzies, a successful longwall method was developed.

The coal of the Miller bed, shown in Fig. 7, is from 4 to 4½ ft. thick and is used as a domestic and steam fuel. The seam is reached by a rock

tunnel 600 ft. long from the Wingate slope on the second level. Because of the impossibility of profitably working the seam by breasts or chutes, little has been done on the seam compared to that done on the other beds, and the second level is the lowest level worked. The dip here decreases south from the rock tunnel to the slope, the coal having an average dip of about 38°, which is a little too flat for the best results.


The level worked at the time was originally opened for chutes and pillars, consequently the practice adopted was not the same as will be used in future development. The gangway, which is the intake airway, was driven in the coal and timbered by three-piece sets consisting of 9-ft. legs, 8-ft. collars, and with lagging between sets.

The air travels from the gangway to the longwall face, through a counter gangway, 4 by 4 ft., driven parallel to the main gangway and about 25 ft. up the pitch from it. The air then circulates up the longwall face to the old gangway above. It was practically impossible to keep this gangway open as a return airway, so a rock tunnel was driven in the foot wall parallel to the top gangway and about 20 ft. from it, as shown in Fig. 9. This tunnel is 7 by 6 ft. and was driven for $19 per yard, which would be about the same rate paid at this time for such a tunnel. Crosscuts, 4 by 4 ft., are driven from the tunnel to the top gangway at intervals no greater than 50 ft. to tap the longwall face; their distance apart depends on the condition of the face.

This tunnel is the return airway for the longwall face, and timber, as required, is brought through it and taken down the longwall face by timber packers. The first cost of driving the top tunnel is slightly higher than driving in the coal, but when the cost of retimbering, general upkeep, and value of a reliable and permanent airway and escapeway are considered, the tunnel is the cheaper. Chutes connecting the gangway and counter are driven up the pitch 25 ft. apart, as shown in Fig. 9.

Opening a Face

The method of opening the longwall face, shown in Fig. 8, was as follows: The first two chutes were driven narrow up the pitch from the gangway with a 12-ft. pillar intervening. The crosscuts above the

counter gangway were driven about 4 by 4 ft. with a 50-ft. block between them. As soon as the chutes were up one block, the longwall face was begun, there being room for one miner, who started in chute 2 and drove angle a to meet angle b, which was simultaneously driven by the counter miner before he proceeded with the counter gangway. Skip, or slice, a, about 6 ft. wide, was then continued up the pitch, its faces (a², a³, a4, a5) being at all times about 18 ft. behind the face in chute 2. When skip a was up about 18 ft. from the point of the angle, skip b was started and continued up the pitch. This sequence was repeated on the skips following. When skip a reached point a5, the face was as shown by the heavy line a5, b4, c3, d2, e1. A miner then proceeded to drive an angle from chute 3, to tap angle 4, and the longwall face proceeded.

Before chutes 1 and 2 tapped the gangway above and before the rock crosscut was driven from the rock tunnel at top of chute 1, the air traveled up the last crosscut from the gangway to the counter gangway, back along the counter gangway to the longwall face, up the longwall face and then down chute 1 to the counter gangway, through which it returned to the return airway and to the fan. The usual practice in this field, when the gangways are driven as water levels or have a gangway above to be used as an aircourse, is to drive but one rock tunnel connecting the seams and then drive a chute, or pair of chutes, through to the surface or to the gangway above for a permanent airway. Until this is done, a brattice, flexoid tubing, sollar, or air box is carried in the rock tunnel, making two compartments for the purpose of ventilation. Very often this is connected to the main counter gangway, or airway, but more often a small booster fan is used which is very efficient even for long distances when flexoid tubing is used.

Outside the longwall face in chute 2, no attempt is made to keep this chute open after angle chute 3 is tapped by the longwall face. The chutes 1 and 2 are driven with a small pillar between them in order that they may be rushed, and there is little possibility of profitably recovering the pillar separating them.

Advancing the Longwall Face

The longwall face developed is shown in Fig. 9. The miners took a 6-ft. skip each, keeping about 18 ft. apart and driving through to the top counter gangway or level. As soon as the skip was finished, the miner dropped back to the bottom of the longwall face and started another 6-ft. skip. The longwall face, in Fig. 3, was about 500 ft. long, 30 miners were working upon it, and the output was about 250 short tons per shift of 8 hr. The longwall face, counter gangway and chutes were worked but one shift, while the gangway was operated double shift.

A sheet-iron chute was used to carry the coal from the men to the gangway. This was kept full and batteries were placed at intervals, where necessary to keep the coal from rushing. Four buckers were employed to keep the chute in shape and run the coal. This chute was moved to the longwall face once each week, when the longwall face was not working.

The longwall face advanced inby from 18 to 24 ft. each week, so that the chute ordinarily was never more than 24 ft. from the face. When the chute was moved to the longwall face, a center post was set under each stringer of the sets outby and next to the chute; these strengthened these sets and protected the chute. The sets outby, and next to the strengthened sets, were knocked out, so that the weight was taken off the faces and the roof was allowed to sag and cave behind, as shown in Fig. 11.

A wing was kept below each miner to prevent things from falling upon the man below, and also to carry the coal into the chute. Owing to the broken ground near the old gangway at the top of the longwall face, cogs were built to keep the top open so that the timber could be brought

down from above and insure an open place until the next crosscut was open from the rock tunnel.


In Fig. 11 are shown the typical timber sets used on the pitch and the lowering of the roof behind the longwall face. These sets are composed of two 4- to 5-ft. legs, usually set in hitches in the foot wall, and a 6-ft. stringer, all of about 8-in. timber. As a rule, lagging is used above the stringers. Each miner, on an average, put in two sets during each 8-hr. shift for which he was paid at the rate of $2.30 per set. The sets were placed on 4½-ft. centers on the pitch and the stringers were set end to end; in the event of a squeeze, a cap piece may be put under the joint and a post set under this, thereby strengthening the joint materially.

As the face advanced before a set, the roof was kept up by lagging and temporary posts. As soon as a sufficient distance was made for a set, a hitch was cut into the rib, about 6 in. in depth, and the end of the stringer was placed in it. A post was then set about 1 ft. from the other end, the other post was set about 6 in. from the rib, and the temporary posts were knocked out. The lagging above the sets can be reenforced if necessary. The timber on the pitch was taken by timber packers on the opposite shift to that in which the miners work, to each skip face of the longwall by way of the rock tunnel above.


A good current of air was traveling at the face at all times. Open lights were permitted and the miners blasted the coal whenever they thought it was necessary. As there were always two free faces, only

small shots were fired. At the time of working this mine, whenever powder was used, it was of the permissible class and was detonated by a No. 7 cap using fuse.

Wages and Hours of Labor

The labor employed at the mines at the time of working this coal seam was all union. The scale of wages was regulated by an agreement between the company and the United Mine Workers of America. In most cases, the miners were working by contract, and the average wages were above the scale rate for day work. The minimum wage was $3.15 per shift of 8 hr. (1916), as compared with $5.25 today (1923). The miners on day work received $3.80 per shift of 8 hr. and were furnished with all tools and blasting supplies, as compared with $6 under the same conditions today, with the exception that the mines in this district are non-union.

Advantages Gained by Longwall Method

It has been proved that, by longwall methods, a larger tonnage per man can be maintained and a larger percentage of lump can be produced. It is stated, however, that the cost per ton is slightly higher than by breast-and-pillar system or chute-and-pillar, This was not true at this mine. Using the other system, the coal was not mined at a profit; but with longwall, the cost was reduced to a point where a profit was made.

This was the result of several causes. What was formerly a safety-lamp mine, because of trouble from gas arising on account of faulty ventilation, became an open-light mine as there were no places for gas to lodge on the longwall face. When breasts were used, the air had to circulate up and down between the crosscuts, which were kept open with difficulty for but a short time; with the longwall method there were no crosscuts, the ventilation was ascensional, and had only one face to sweep. There was no upkeep on the return airway, as it is driven in the foot-wall rock. Although a large amount of timber was required, it was no greater than was formerly necessary, because under the breast-and-pillar system the breasts had to be well timbered until the pillars were drawn. In the longwall method, the timber can be taken to the face more quickly for it must be moved down one face only while in the breasts it had to be distributed through the crosscuts and packed up to the faces. Less powder was used and a larger percentage of lump coal was obtained because there were a greater number of free faces. The work was concentrated and the longwall face permitted a more frequent and closer inspection of the working places by the mine officials. In the longwall system, practically all the coal was recovered, but the breasts could not be kept long enough for even their limit to be reached, and the pillars had to be worked by small skips or lost.

Advantages Gained by Driving Gangways in Foot Wall

When driving gangways in the foot-wall rock, the first cost of the gangway will be higher; but this will be overcome by the increased recovery of coal above this passage, the gangway will serve as a permanent airway for the level that may be driven below, and will require but a small upkeep, little retimbering being necessary. Ordinarily, no timber is required when driving is done in the foot-wall rock and the tracks will always be in good shape.

Discussion of Longwall Methods in Pitching Seams

If larger coal were the only factor to offset the higher longwall cost under ordinary conditions and the profits were increased as the result of higher sales realization, it would naturally leave nothing to be desired. However, in a seam of this kind, anything that increases tonnage per man per day will lower cost correspondingly. The illustrations in the foregoing description show that there are no pack walls, as in a regular longwall system, no roadways to be maintained to the face, no brushing is necessary. If the method can be pursued, unless dangerous and uncontrollable caves prevent its operation, less timber is required. The face must be kept advancing and open, as developing a longwall face for ventilation and working is slow and expensive. If squeezing is troublesome, due to the roof sagging or the bottom heaving or sliding, or if there is a combination of these characteristics, such a pitching seam can be profitably worked by a longwall method if it can be worked without loss by any other method. In the writer’s opinion, if such a seam is worked with a face not exceeding 400 or 500 ft., lower costs will result if the mine is worked steadily enough to keep the face open and the timber can be easily distributed from the top and through a counter or old gangway or similar opening.

The question is raised why the system is not used more and why it has been discarded in several instances where it has been tried. The writer has studied some of the cases and offers the following suggestions: Probably the two major causes for such failures as have been observed are: first, inexperience and lack of interest on the part of the immediate officials in charge; second, the failure to substitute the contract system in place of day work, a difficulty that may be due to the attitude of the labor union. It is necessary to keep the face advancing and to have regular shooting times. Timber is most efficiently distributed on the opposite shift from that on which the miners are employed. Congestion of coal in the delivery chutes is a drawback and can be avoided only by keeping chutes cleared as coal is made. Chute starters are required to look after coal in the chutes, as congestion will surely result if. the face is too long. One chute with two outlets will handle 200 to 300 short tons per 8-hr. shift on a longwall face 400 ft. long.

The writer has never seen a longwall face worked successfully on steep dipping seams when the face is carried up the pitch, as in the method of overhead stoping in metal-mining practice. Shales and sandstones with more or less carbonaceous material separating them do not permit this method, at least in this district. As the coal is withdrawn, the roof breaks; and if it be strong, there is an area open that is entirely too large to be safe. A cave is almost sure to follow and it breaks along the face, which is lost, causing a wild and dangerous place. There is also greater danger in facing large niggerheads that the coal might contain. It appears that the face must proceed inby or outby along the level haulage road, and not be worked in sections up the pitch, which approaches a wide breast. An area once removed should be of no further use and the quicker it can be allowed to fill with waste the better.

Working a Thick Seam of Two Benches on a Heavy Pitch

Location and Description of Seam

On the north and east side of the Carbon River, seam No. 3 (Fig. 10) is known as the Big Ben and is a high-grade, coking, bituminous coal. Operations were started on the seam in the West Douty measures in the southwest quarter of section 4 and extend south into the northeast quarter of section 9.

On this side of the river, an average section of the seam has 5 ft. 6 in. of coal in the top bench, 5 ft, 4 in. in the lower bench and 7 ft. 7 in. of parting, which is mostly shale (see Fig. 6), although it contains some bony coal and carbonaceous shale. This parting must be left in the mine as it constitutes waste that would be prohibitively expensive to transport and remove in the cleaning plant.

The hanging and foot walls are both good; and if it were not for the extreme thickness of the seam, the inability to gob the waste, and the pitch of the seam, the ordinary prop and cap, or single stick, method of timbering and working would fill all requirements so far as the main walls are concerned. As the pitch varies from 45° to 70° and the lift is approximately 900 ft. long, an entirely different method of working than has been practiced in the field heretofore is necessary.

Factors Deciding Method of Working Adopted

To mine each of the benches of coal in this seam separately naturally raises the question as to which, the upper or lower bench, should be taken out first, and to what extent the workings of one bench can be kept in advance of those of the other bench. When solving a problem of this nature, several factors must be considered, such as the pitch of seam, thickness of seam and its benches, thickness of intervening strata or parting, the hardness and tendency of this parting to swell or to slide, caving habits in general of main walls, any peculiar features of coal to be worked, presence and extent of faulting of strata, and, a most important factor, the length of the lift.

In this particular instance, the main roof and bottom are fairly good. The roof stands well and the bottom, under normal conditions, does not swell or slide to any great extent. There is little difference in the behavior of the coal benches, although the bottom bench does not work quite as freely as the top. As in any seam, the top bench requires less timber as it has the better roof, and the bottom bench the better foot wall. There is, therefore, little to determine from the individual characteristics of the coal benches which should be worked first. However, small faults cut the seam at various points, horses appear, and the intervening shale parting varies considerably. This materially affects any method and is further aggravated on account of the scarcity of experienced pitch timbermen to repair chutes on this long lift.

The problem therefore resolved itself into how the intervening bench of impurities would act, whether the chutes could be, as a matter of safety, economically kept open long enough to recover the coal, and how the main roof would act and affect the lower seam workings. Experience of the management over a period of two years has proved that with an ample supply of timber such a seam can be profitably mined under normal market conditions.

Chute-and-pillar Method Used

As a chute-and-pillar method played an important part in the various methods of working, this method is described.

The usual method of opening up the pitch and the general practices are the same as have been described, with the exception of differences in the method of driving chutes, conducting the air current, and arrangement of manways. The method of drawing pillars is exactly the same. The method of chute driving is common practice at the Carbonado and Wilkeson mines.

As shown in Fig. 12, the chutes are driven up the pitch 3½ ft. high and 4 ft. wide with no permanent brattice for ventilation. As a rule, the chute is driven on a bottom or top bench of the seam, depending on the nature of the walls, which may both be of coal or bone. However, but one wall is usually of coal and preferably the hanging wall, for the loose material must run over the bottom.


Timbering of the chute will necessarily vary with the ground, but in the Big Ben seam posts with a cap piece are set every 5 ft. on the pitch in the chute which is on the bottom bench. These last only for the time required to drive the chute but a few blocks and are principally used to carry the canvas brattice used in ventilating the chute while driving between the crosscuts. They also enable the miner to travel to and from the working face for a distance of one block. A step is hitched in the ribs every 5 ft., as shown, and a step made at the posts as shown. These are destroyed later by the loose material running down the chute.


In Fig. 12, the arrows show the direction of the air current used for ventilation when the gangway is used as the intake. The mines in the Carbonado district have been principally opened as water levels, for which reason chutes are driven to the surface at certain intervals to be used as timber and air passages. Whether these are used as an intake and the gangway the return, or just the reverse, depends on whether or not the mine is gaseous. If electric haulage is used on the gangways, the gang-

ways must be the intake airway. No attempt is made to keep the first crosscut or counter gangway open as it is not necessary for an airway other than at the time the area is being worked.

If the gangways are used as intakes, a blower fan and doors are necessary at the main opening (which is not the case at any mine in the Carbonado district), or small exhaust fans are placed on the air chutes of the various seams eliminating the doors on the main haulage road.

This is the general practice here if the gangways are the intakes. The most common practice in the Carbonado district is to use the gangways as the return airways and to place one large exhaust fan near the mouth of the main opening, using doors in this passage and utilizing the air chutes on the different seams as intakes if several seams are worked from one main cross-cut tunnel or opening.

Method of Driving Narrow Chutes

Such a chute can be rapidly and more cheaply driven under the Big Ben method, even if less coal is loosened, provided the chute will stay open and the gas is not troublesome. If the chute will not stay open, it becomes a question of maintenance, and probably a different method of timbering and chute driving would be used. It is impossible to use the chutes for traveling ways, for which reason about every fourth chute is made into a manway and a permanent ladderway placed in it, over which no coal is run during its life as a manway. A timber box, such as is shown in Fig. 2, is placed in the chute, outby, adjacent to the manway chute. Batteries are placed every two blocks to enable persons to cross the chutes safely at these points for any purpose, such as distributing timber.

Pillar coal is the cheapest coal and is usually the source of profit, and the advantage of the system is the speed with which the chutes can be driven. If the chutes stay open, there is a large saving in the timber and timber distribution on the pitch. Because of the nature of the seams, there are cases where a larger chute cannot be kept open and for long lifts it is very advantageous to maintain only a small chute.

Disadvantages of Narrow Chutes

The system has some distinct disadvantages. The work must be well balanced between chutes and pillars as very little coal is obtained from the narrow chutes; if the seam is at all gaseous the faces cannot be kept clear because of the canvas brattice and other ventilation difficulties on the heavy pitch. There is a great deal of trouble from the blocking of the chutes by timber and large pieces of rock, or niggerheads; such chutes must be freed and chute starting becomes a hazardous occupation.

The author does not recall an instance where a chute starter was caught in the wider chutes where manways are kept separate by a brattice in the chute, but in the narrow chutes workers have been suffocated while removing a block. The wide chute offers one remedy for the prevention of accidents from this cause. It is obvious that with a manway in the chute it is easy and safer to take off a board or jar the chute to start the coal, whereas it is necessary to go up a narrow chute, place a charge of powder, and blast away the obstruction—an extremely dangerous practice if the mine is dusty.

Precaution Used in Chute Starting

The following safe and practical precautions should be taken when starting chutes; in this district, no starter has lost his life when these precautions were taken. As a rule, a chute blocks just below a crosscut, which means that it must be faced for at least about a block. A starter should never go alone to start such a chute; he should have a companion, who remains at the open crosscut below. Before the starter goes up the chute, a grizzly should be constructed over the chute at the lower crosscut, by placing timbers across with openings large enough for the loose fine coal to go through. If the coal rushes or breaks away, as it sometimes does, and catches the starter while in the chute, he goes down ahead or with it and is caught at the grizzly. The fine coal passes through the grizzly and his partner can easily and safely rescue the starter.

Timber Packing

The foregoing method of carrying narrow chutes is practiced in lifts that have reached over 1200 ft. For many years, the workings have been confined to water levels. The timber is usually taken into the mine through combination air chutes and timber chutes driven to the surface. As in other occupations in coal mining, it has been found that when timber packing is done by contract the results are more efficient, The cost of handling timber is an important factor and unless it is carefully supervised the costs are soon on the red side of the ledger.

In the Carbonado district, a counter gangway is defined as a cross-cut made large enough to handle timber and material by tramming and which can also be used as a main ventilating passage. These openings are usually made about four crosscuts apart, or about 200 to 240 ft. apart on the pitch.

In the Big Ben mine at Carbonado, the timber packers are not working on contract but are paid a day’s wage. Under these conditions, on a pitch of about 60° and where the timber is brought in at the top, or 13th crosscut, which is called the 13th counter, five men can pass 80 props per hour from the 13th to the 8th crosscut or counter. These props average 6 ft. in length but run from 5 to 9 ft. in length. During the 8-hr. day, five men will pass 320 props, 160 from the 13th to the 8th counter, and 160 from the 13th to the 4th counter. It takes nine men 1½ hr. to pass 40 props four blocks through the crosscuts and land them one-half block above or below the crosscut along which they are being passed. In 1 hr., using a timber truck on a counter, nine men can move 40 props along the counter for four blocks and then down the pitch one and one-half blocks.

Methods of Working Tried in Big Ben Seam

Rock Chutes Driven between Coal Benches

Various methods of working the Big Ben seam have been tried, the first of which was to open up the bottom bench as has been described. At first, fourth, and eighth counters, rock chutes were driven to the

top bench, as indicated in Fig. 13, which shows actual pillar measurements in this section plotted on the plane of the seam.

A Modified Longwall Method Used in Top Bench

After the rock chutes were driven to the top bench, the attempt was made to remove the top bench ahead of the bottom bench, using a longwall method on the top bench. The face advanced up the pitch, and failed for the reasons given under the discussion of longwall methods, but the worst result was the loss of the lower seam. The causes of this were due to the large area worked ahead on the top bench. The coal running down the chutes in the lower bench wore the chute to a width of 15 ft., or more; and because of the heavy pitch, the coal ran against the roof and ultimately wore through to the top seam. The whole area then became wild and uncontrollable, poor pillars resulted and a squeeze started, which overran these workings and extended to the gangway, resulting in a loss of considerable coal and causing heavy maintenance. Much of the difficulty on the lower bench probably could have been avoided in so far as the widening of the chutes was concerned if a large maintenance force of experienced timbermen had been available.

It will be noticed (see Fig. 3) that the top bench was worked in advance of the lower bench for the greater part of the area of the workings, thus leaving large sections of the main roof to act as it would above the parting between the two benches. Here lies the worst danger of this method of working the top bench first, because the main roof does not break immediately and there is no way of telling when it will cave. When the main roof does cave, it breaks through the bench above the bottom-seam workings, caving them in. This happened on two occasions and it was fortunate that no one was in the workings at the time. The method was unsafe and so was abandoned.

Angle-chute Method in Top Bench Worked in Advance

To further test the method of working the top bench ahead of the lower bench, a system of angle chutes and crosscuts was tried on the upper bench. The bottom seam was opened as before and at each block rock chutes were driven through the parting to the top bench or seam.

It was found that whenever a considerable portion of a pillar on the top bench was extracted in advance of a similar operation on the lower bench, the tendency of the foot wall of the top bench or the parting was to bulge and crumble, and thus be liable to slide. In this condition, it afforded poor material for roof protection while the lower bench was being extracted. For this reason, it was found advisable to keep the lower bench workings about one block in advance of similar pillar workings on the top bench.

The chutes and crosscuts on the top bench were driven at an angle of about 45° across the pitch and starting at the top of the chutes of the lower seam, the top seam pillar workings were worked in advance of similar workings on the lower bench for a distance of from one to four blocks. The same results were obtained and a cave from the main roof broke through the parting between the seams and nearly resulted seriously. Experience has demonstrated that this practice is unsafe as there is no way of telling the condition of the main roof once the top seam is removed.

It was found, over a period of one year, that when the top seam workings were worked ahead of the lower seam workings by the longwall and angle-chute methods described, the recovery was 44 per cent.

Bottom, Bench Worked in Advance and Angle Chutes in Top Bench

Because of the experience just described, it was decided to work the lower bench in advance of similar workings on the top bench by starting at the top crosscut and working the lower bench pillars one block ahead of the top bench pillars and, further, to open up the top-seam chutes only as required and drive these chutes on an angle across the pitch.

Accordingly the chutes were opened on the lower bench and rock chutes driven to the top seam as required, starting at the top of the pitch workings. The pillars on both benches are recovered by the angle-and-tail method, or regular pitching seam practice, which has been already described.

The mining method is shown in Fig. 14. A rock chute 1 is driven from the next to the top crosscut of the bottom-bench workings between chutes c and d, and slightly to one side of chute c, so as not to weaken the chute and to facilitate traveling to and from the top-bench workings. These rock chutes serve for ventilation and passages through which the coal from the top-bench workings can reach the lower bench chutes through which the coal is delivered to the gangway.

Angle crosscut 2L, on the top bench, is then driven with about 45° pitch toward 2R, previously driven from rock chute 1 of chute b; this makes a connection for ventilation. Angle chute 2R on the top bench is then driven. The half-block 3 on the bottom bench is then removed; this is called taking off the angle. A similar angle 4 is then removed on the bottom bench. The half-block 5 between chutes a and b is then removed; this portion of a pillar is called the “tail.” Block 6 on the top bench is then removed through chute 2B from rock chute 1 in chute a of the bottom bench. As soon as the tail 5 between chutes b and c is removed, the V-shaped piece of coal on the top bench vertically above this tail, and between angle chute 2R and angle crosscut 2L of the top bench will be removed. As soon as the tail 8 of chute a on the bottom bench is removed, the coal of the top bench vertically above angle 4 will be removed from angle chute 2R2 and rock chute x in chute a of the lower bench. The workings are advanced inby in the same manner, the pillar workings on each bench retreating toward the gangway.

By using the angle system, it was expected that slides would be averted in the top-bench chutes. Although the bottom heaved somewhat in the top-seam chutes, this was not serious in a distance of one block; but when driving the angle chutes, the high ribs sloughed off and the chute ribs, although lagged, would run on the high side. The high rib sloughs in a level crosscut and this coal can be allowed to stay there, but this is not so in an angle crosscut or chute on this pitch. An excessive amount of extra timbering is required.

Angle Chutes and Angle Crosscuts Abandoned

The method of driving angle chutes and crosscuts on the top bench was abandoned and the present method of driving the top-seam chutes straight up the pitch was adopted. After several months’ trial, the method has demonstrated its superiority over the former methods. The most successful method developed is to open the seam by narrow chutes and work the pillars by the usual method, but to work the bottom-bench pillars first, one block in advance of similar workings on the top seam, starting at the top and running the coal down through the chutes on the lower seam. The top-seam chutes are offset about 6 ft. from the lower seam chutes, are driven straight up the pitch, and are opened only as required to keep up with the lower seam pillar workings.

Precautions Taken in Drawing Pillars

Before a pillar is started on the top bench, after the top block in each bench has been removed, the battery holding back the caved material in the lower bench workings is blasted out and the caved material is run into the lower bench area excavated under the area of the top bench to be worked. This caved material is caught by a battery in the lower bench workings above the area being worked here and which is one block in advance of the top-bench pillar workings. This gives a better foot-wall support for the top-bench workings and has been found indispensable for this purpose after one block has been removed and a cave has occurred in the lower level^ bench workings. However, the workings generally stand open until one or two blocks are removed on the bottom bench. Caving is then prevented with difficulty and the pillar workings must proceed rapidly to avoid losing a pillar before all the coal is removed, and it is only by filling with the caved material that the intervening rock bench can be kept in place long enough to remove the top bench of coal.

The method of filling the lower bench workings with caved material and drawing the pillars is shown in Fig. 15. In (a), the development work for the chutes and crosscuts is complete on the bottom bench, in which a is the face of the bottom-bench chute. In (b), the top block of the bottom bench has been removed and a rock chute and coal of the top bench has been removed; the same procedure has taken place in the sixth block (c). In (d), the procedure shown in (b) is repeated in the fifth block. The battery at the cog line in the seventh crosscut is then blasted out and the caved material run down against the battery in the sixth crosscut filling the space in the lower bench workings below the coal in the top bench of the sixth block, which is then removed. This procedure

is repeated until the entire section of pillars on both benches is removed, as shown in (h).

Over a period of 9 months, it has been found that by this method of working the bottom bench in advance the recovery so far has been 70 per cent. This will be increased for most of the loss to date has been due to the losses in the angle chutes where the ribs ran away. The greater portion of the coal still in is represented by the present live workings and is being recovered.

Discussion of Methods

It is apparent that the regular rock-chute mining method is not strictly followed at the Big Ben mine, but, as in the regular method, all gangways, airways, and counters are developed first in the lower seam.

The question might arise why the top seam could not be worked out entirely from rock chutes on the gangway on the lower seam and then allowed to cave, and the lower bed then worked in the regular manner by narrow chutes and the pillars drawn. Objections to this method are that the parting between the coal benches heaves and slides before the main roof caves, on account of the character of the parting. The presence of small faults and the heavy pitch destroy what is later to be the hanging wall of the bottom bench. The lift is so long that a squeeze comes on the lower workings and the chutes on the lower seam rapidly become uncontrollable; this means the making of man ways and timber- ways on both seams, while the method last described requires this to be done but once, and that in the lower seam.

It is not an uncommon method to work hard firm seams, such as anthracite, on a pitch and have the operations carried on simultaneously in both seams. However, the nature of this seam is entirely different and it is not a common practice to carry on pillar workings in the bottom bench in advance of similar workings in the top bench, as has been described in the foregoing paragraphs.


It is the author’s opinion that when the parting between the two benches of coal is firm and thick enough and one or both of the walls at the top bench are inclined to heave or sag to such a degree as to allow the main roof to break over the waste workings in the top seam or bench, an excellent method is offered of avoiding bumps in the workings of either seam or bench in deep mines, when the walls of the lower bench are firm. If the lift is made such that a longwall method, as described in the Miller mine, is used, the main roof will break but owing to the filling of the top bench workings with waste no serious consequences result. The bottom bench can then be worked and caves obtained as desired. However, if this is not done the top bench is usually lost, the lower bench workings remain open for a considerable area, and when the main roof does let go such a tremendous pressure is instantly thrown on the adjacent workings that a serious bump results, crushing the pillars, breaking the timber, and at times caving in the section of the mine. This is especially true in a region of faulting. If the coal makes much gas, a large quantity of gas is often liberated from the crushed pillars at the same time. If the caving of the roof can be regulated, the bottom will not give any serious trouble from bumps.

In a pitching seam, as in any other, when the lift is over 450 ft. long, trouble is experienced in one way or another. From the start, due to the shortage of inexperienced labor, it has been difficult to maintain the chutes and, even with plenty of such labor, the cost of chute maintenance in a seam of this kind is very high. The lift is entirely too long and the high maintenance cost is largely due to this cause. In the present operation of the Big Ben seam, the gangway has reached its limit and the workings have not been extensive enough to more than demonstrate by experiment the best method of working this seam to follow in future operations. If the present lift were made into two, the present mining system in a new development would make a more profitable mine. Further, the output would be more flexible. Less territory would have to be kept open for the same output or the same territory opened could be made to yield a larger output and what is now a struggle to yield 200 tons per 8-hr. shift could be made to yield 300 tons in the same time at a lower cost per ton.

Machine Mining at Newcastle Mine, Pacific Coast Coal Co.

Location of Mine

The Newcastle mine is located in King County, Wash., 22 miles by rail from Seattle on the Pacific Coast R. R. At present, the output is

about 900 tons of coal per 8-hr. shift, 300 men being employed. Five workable seams are present on the company’s property, which is distributed over three sections; viz., 27, 26 and 25, T 24 N, R 5 E, W.M. The seams have an average dip of 40°, and are separated by the usual shales and sandstones. The coal beds worked are shown in Figs. 16 and 17.


The geology of this area has been fully described. The coal mined is subbituminous and finds its principal market in the Puget Sound cities and contiguous territory.

Method of Development

As the principal operations in the Newcastle mine are confined to the Muldoon seam and, to a lesser extent, the No. 4 seam, and the methods of mining are identical in both, the description of the Muldoon will apply to No. 4. Sections of these two coal beds are shown in Fig. 21.

The character of the walls and the coal of the Muldoon and No. 4 seams are such that breasts varying in width from 15 to 70 ft., can be mined to advantage with a lift from 400 to 800 ft. on the water level.

Experience has demonstrated that this mine cannot, under present conditions, be economically worked mining the coal on the pitch on the advancing method, for the walls, although excellent so long as only the gangways and counter gangways are driven, become very bad and the gangways are squeezed and expensive to maintain if kept open for a long time. The coal in the gangway and counter stumps became crushed to such an extent as to be too fine to be profitably mined; the worked out areas were apt to fire, due to spontaneous combustion; and the amount of black damp evolved and difficulties in keeping the airways open made the ventilation expensive.

Retreating System Used

Because of the foregoing, this mine has been worked on the retreating system. The lateral extent of the mine workings on the 4th level was 8000 ft. from the slope on the east side, and 4500 feet on the west side. The mine has been worked to the 4th level from the present slope, which is about 1760 ft. long and dips about 40° (see Figs. 16 and 17.) A water level was worked and called the first level. It takes about five years to sink the slope one level, a distance of 500 ft, complete the main airways, and drive the gangways and counter gangways to the boundary and prepare the new level for retreating operations. Taking all things into consideration, the gangways are advanced at the rate of about 2000 ft. per year. There is no deviation from the ordinary method of gangway and counter gangway driving already discussed, with the exception that chutes are not opened except at 300-ft. intervals, these being driven to the counter gangway only and used for ventilation and dump chutes for the disposal of the coal from the counter gangway. By this procedure, but few stoppings are required and all of the air possible is conveyed to the inside end of the workings.

Ventilation, when developing ahead of the last dump chute, is accomplished by booster fans, electrically driven, and air boxes. The dump chutes are so spaced that they can be later used for regular chutes.

On this dip it has been found that a 70-ft. pillar can be worked, 35 ft. on each side, from 20-ft. breasts. The breasts are therefore opened on 90-ft. centers, except where an old dump chute is utilized, when the pillar is proportioned to suit the conditions. The system of working is shown in Fig. 18.

Method of Driving Breasts and Drawing the Pillars

As shown in Fig. 18, the first breast is driven at or near the face of the gangway, although it is preferable to have the gangway extended far enough beyond the last chute to provide room for several cars to be loaded at the last chute. Breast 2 is also started and the relative positions of

the breasts advancing up the pitch are shown by breasts 4 to 10, breast 4 having reached its limit, which is from 50 to 75 ft. from the low rib of the old gangway above. The chain pillar is varied in pitch length according to the dip of the seam, tendency for the coal to run, nearness to the face of the old gangway, and amount of water running out of the old gangway that would have to be pumped. The two principal reasons, in their order of importance, for leaving this chain pillar in are to confine as much water as possible on the level above and to serve as a barrier for the blackdamp in the old workings.

A level is always sealed off near the main airways and slope and the blackdamp is confined to the sealed-off areas. The advantage is taken of this condition and pipes, with valves and hose attachment, are placed at these stoppings so that, in case of a fire, the blackdamp, which in most instances is under greater pressure than that of the mine atmosphere, can be made to flow effectively through pipes to any desired part of the mine, and the workings near the stoppings can readily be flooded with this inert gas.

Details of Breasts

When driving a breast, it is mined 10 ft. wide to a point about 10 ft. from the second crosscut, where it is widened to 20 ft. The manway, with a ladderway, also serves as a passageway for air, timber, compressed air and water pipes. This manway is driven about 5 ft. wide and is separated from the coal side of the breast by a board brattice. The coal side is kept full of loose coal (worked full) up to the top crosscut outby, to prevent breakage and aid ventilation. As no traveling is done in this system, except through the open crosscut at the top and the bottom counter gangway, but one battery or bulkhead, is ordinarily used; this is placed at the counter gangway, as shown in breast 3. However, unless the chute can be kept full, it is advisable to have bulkheads above the counter gangway, at every crosscut, to eliminate the breakage of the coal. It is an easy matter to remove the bulkhead boards holding back the coal if there is enough to keep the entire coal-side full, and to replace them if necessary. The main objection to too many bulkheads is the time lost as the result of large chunks, timber, etc., becoming lodged at the bulkheads when running the coal. Wings placed at each crosscut deflect any loose coal from the manway side through a small opening in the brattice to the coal side of the breast.

Angle Crosscuts Used

Above the second crosscut, and sometimes above the counter gangway, all crosscuts are driven on an angle across the pitch, half way from each side of the block. These are driven as angle chutes 10 ft. wide on a pitch sufficient for the coal to run, generally using sheet-iron. They serve not only the usual functions of a crosscut but as wings or coal chutes when drawing the pillars, for which purpose an ordinary crosscut is worthless. If such an opening, called a wing, is driven at the time a pillar is being removed, the drawing of the pillars is slowed down materially. They are longer and cost more to drive than a level crosscut, but the cost of a wing and a level crosscut more than offsets this additional first cost. The only reason for driving the second crosscut level is to get a hole through for ventilating as rapidly as possible, so as to afford storage room for coal, otherwise not possible.

Drawing the Pillar

As soon as breasts 1 and 2 are finished, the pillar, which is called the top block, is removed as is shown in the top, or sixth, block between breasts 4 and 5, half of the block being removed from each side, the coal running down the angle crosscut into the breast.

The first thing to do before starting a pillar is to build a battery as is shown in breast 4. On this dip, the batteries are made of heavy posts set from 3 to 5 ft. apart across the breast and a lagging of props is placed above them. Some of the lagging is left off as long as coal is being run from above the battery, which would be the case with the lower battery at the 5th crosscut in breast 4. If a cave is likely to occur, all the lagging can be put in place. Before a battery is left, it is banked with coal on the high side, which serves as a cushion for the caved material to land against. The battery protects the pillarmen from the danger of being shut off by a cave; in addition, the small block x is used. This piece of coal is left in, and when the squeeze comes upon it, crushes and runs against the battery serving as a cushion; or it is blasted out for the same purpose before the battery is left. The pillar is then breasted as shown in the third block between breasts 2 and 3, and the top block between breasts 4 and 5, but before it is advanced to any extent, an opening is made to the breast for ventilation.

Whenever possible, the operations are carried on simultaneously from both sides of the block; if not, the procedure shown in the third block between breasts 2 and 3 is followed. A small pillar of coal, y in breast 2, is left to hold back any cave that may occur in the breast and as a support for the roof; it may or may not be recovered, depending on the state of affairs existing in the old breast. The pillar face advances as a breast, shown in the top block between breasts 4 and 5 and the fourth block between breasts 2 and 3.

The custom here, as through all the district, is to call a working place a chute, breast, or room as long as the face is advancing, is to be advanced, or is standing idle with the pillars still in. As soon as the removing of a pillar is started the working place is called a pillar; and breasts 1 to 4 would be known as pillar 1, pillar 2, etc. The pillar draw¬ings retreat toward the gangway and the workings proceed outby.

The gangway stumps are removed as soon as block 1 is removed. The method used is to breast up the entire stump. This is done by carrying a face up the pitch from the gangway to the counter gangway the entire width of the stump, or 70 ft., the coal being loaded into mine cars through chute spouts placed apart a distance equal to that between the centers of two cars placed bumper to bumper. In this way storage for coal is provided in the stump area, several cars can be quickly loaded, and the stump removed with the maximum speed.

In Fig. 19 is shown a section of the mine and indicates the progress made in an area selected for the reason that it shows a condition contrary to that desired. In this section, the mine was worked inby and not on the retreat, because there was an unusual demand for coal and the gangway had not reached its limit. In the workings shown, the mine was closed down for four months, during which interval the pitch workings caved tight although timbered with four-piece sets of 8 by 8-in. timbers, the collar being 10 ft. long and supported by three posts, placed about 6 ft. apart on the pitch, lagged between, and collars end to end across the breasts. The counter gangway was reopened and kept open with great difficulty, and the gangway has had to be retimbered several times. These conditions on the gangway prove that this seam of coal

cannot be worked advancing along a gangway of normal length even if large pillars are left in and only the breasts worked. The pillars left in become so badly crushed that the coal could not be profitably mined.


Unless the roof is exceptionally good and likely to remain so, the gangways are timbered with three-piece sets spaced 7½ ft. between notches at the collar, 7 ft. in the clear above the rail to the underside of the collar, and 11 ft. in the clear between the legs at the rail.

Where the roof is good, single-stick timbering is used on the pitch. The general practice is to use sawed timber of not less than 7 by 7 in. in cross-section, and cap pieces with the props. If the roof is bad, four-piece sets are used, as has already been described.

All timber used on the pitch is carried by the miners. There are distinct advantages in having this work done by the miners when the latter are working on the contract system, which is the case at this mine. Then each working place gets its proper supply of timber as needed; there are no timber packers; there are no delays on the pitch due to the passing of timbers; there is no obstruction of crosscuts due to timber being stored in them; and a more accurate record of the timber used and its distribution is possible.

All timber is placed on the gangway in front of each breast on the shift when no coal is being transported and the miners remove most of it above the chute trapdoors before going up the pitch. An order for timber needed in each working place is given by the miners to the district fireboss, who in turn delivers it to the timber distribution supervisor who sees that the order is filled.

To be successful, this system is carried as part of the contract with the miner, who is paid for packing the timber and putting it in place, but receives no payment for timber not in place. Contract rates for timbering and timber packing are discussed under wages and hours of labor.

Mining the Coal

All mining is done on the contract system. Both mining machines and coal picks are used. As the coal works freely when the roof pressure is brought into play, it is not necessary to undercut or shear the coal in the pillar workings with machines, except when driving the pillar wings and removing the top block, or starting the pillar face in the block. The machines are not used in driving ordinary crosscuts and counter gang¬ways, where the shooting is done on the solid and where there is not room to use a machine to advantage. They are used at times in the gangway and always in the breasts, chutes, and angle crosscuts.

Factors Determining Selection of Mining Machines

Since the introduction of fuel oil and the development of hydroelectric power, markets for steam sizes of coals of the subbituminous rank have been closed, necessitating the marketing of this coal as a domestic fuel. This requires a greater production of the larger sizes. As the mines increase in depth, this becomes more difficult and the cost of production increases with the depth unless something is done to offset the added expense.

With the substitution of high explosives for black powder and with a type of miner not skilled in the use of a pick, production increased through the practice of solid shooting, which does not make large sizes of coal. Something had to be done to increase production and at the same time give more lump coal. These circumstances have necessitated the extensive use of machines for doing the work formerly done by skilled miners, namely mining the coal at the face instead of shooting off the solid, also doing it at a greater rate than is possible for even a skilled pick miner. It is a fact that in production alone, with a very keen market, even for fine sizes, the use of machines has decided the difference between the black and red side of the cost sheet of this mine, on account of the greater progress made.

Method of Operating Machines

The machine used is the Sullivan “Post Puncher.” The method of setting up the machine is essentially the same regardless of where it is to operate, but its position is a matter of much importance. Ingersoll machines of the same type were also used. It is advisable to use breasts 20 ft. wide in the Muldoon seam although 40-ft. breasts have been used on the upper levels and on the present level in No. 4 seam. In a 20-ft. breast, there is plenty of room for two miners and not enough for three

to work to the best advantage. They divide their work in such a way that one runs and takes care of the machine, and the other packs and sets the timber and assists the machine miner to move the heavy part of the machine. They work jointly in all operations to their best advantage and divide the earnings of the place equally.

As shown in Fig. 20, the first cut is made from a post set from 18 in. to 2½ ft. from the face and about 7 ft. from the rib of the coal side of the breast, although if the dip of the seam and character of the coal are such that the coal is not apt to break out without warning, the cuts are started alternately at each rib to avoid moving the machine as much as possible. It is good practice to work toward the manway side of the breast and the machine is placed to suit the conditions in the place.

A cut 8 ft. in depth is put in, using different lengths of extension bars. One miner operates the machine, swinging it by a worm crank, with one hand, and feeding the cylinder forward with the other hand. The cuttings fall out of the cut because of the dip of the seam. Time is saved if two posts and sets of blocking are available, and while the machine runner is making the cut the other miner sets up the second post. When the rib cut is completed the transfer of the machine, which weighs about 225 lb., to the other post is quickly made and the second cut put in.

The machine and posts remain in the breast near the face until the breast is completed, as there is no shooting of coal to injure the machine

nor loading of coal to interfere with it. The main precaution to be observed is the placing of the post so that the coal will not be apt to break out and catch the machine runner. Sufficient coal is kept near the face, by means of lagging above the props, to furnish solid footing for the machine runner. If there is likely to be any danger of the coal breaking out, the pieces of coal x and y will not be mined; probably they will have to be blasted out.

In mining the Muldoon seam, the cut is made in the band of bone about 5½ in. thick and 2 ft. from the bottom (see Fig. 21). In No. 4 seam, the cut is made in the 6-in. piece of coal near the roof. The position of the swinging gear on the post depends on whether the coal is mined near the top or bottom. Very little blasting is required, for the coal once mined soon works free and can be easily taken down with a pick; if shooting is required, the shots are very light.

Progress Made with Machines

The average number of square feet that can be cut in an 8-hr. shift using mining machines in the Muldoon and No. 4 seams follows:

In a breast 40 ft. wide; 120 sq. ft. for three men, or 40 sq. ft. per man-day.
In a breast 20 ft. wide; 100 sq. ft. for two men, or 50 sq. ft. per man-day.
When using machines, a breast 15 ft. wide is not advisable as it is too wide for a bad roof and too narrow for a good roof.
In chutes and angle crosscuts 10 ft. wide; 60 sq. ft. for two men or 30 sq. ft. per man-day.

These averages include time taken up in the working place for timber packing, setting timber, chute building, and all other work necessary in the place. The results obtained when timber packers were used were much less.

The average number of square feet that can be cut with mining machines in the Muldoon seam when not timbered with sets is:

In a breast 40 ft. wide; 240 sq. ft. for three men, or 80 sq. ft. per man-day.
In a breast 20 ft. wide; 160 sq. ft. for two men, or 80 sq. ft. per man-day.
In a breast 15 ft. wide; 130 sq. ft. for two men, or 65 sq. ft. per man-day.
In a chute 10 ft. wide; 100 sq. ft. for two men, or 50 sq. ft. per man-day.

These averages include time taken up for chute building, timbering, timber packing, and all other necessary work.


Eli T. Conner, Scranton, Pa.—In 1910, I made a professional visit to the Carbonado mine, in which, as I understand the paper, most of the work described has been done. At that time the manager was a Mr. Davis, who had spent some years mining in the southern anthracite regions of Pennsylvania, where the coal beds stand on rather steep pitches.

He said that he had introduced in Washington many of the methods generally practiced in the southern anthracite region. While the pitches of the beds at Carbonado were about the same as those of the southern anthracite region, they were thinner than the Great Mammoth bed and some others.

The practices described appear to be accomplishing somewhat better recovery than the average experience in the thicker beds of Pennsylvania. In the steep pitching measures of the Pennsylvania district, it is a general practice to open from the gangway with chutes square up the pitch; then after passing the airway, which usually is 40 ft. above the gangway, and parallel thereto, chambers or rooms are extended to the rise, gradually widening to 24 to 30 ft., which chambers are advanced to the limit, usually from 250 to 300 ft., keeping the chamber filled with coal and carrying a manway on each side. The plans described differ from the Pennsylvania practice, in that chambers to the rise are driven narrow and not square up the pitch, which reduces the pitch of the chutes materially. This practice permits of using an open chute, down which the coal is carried to the gangway. By this practice the yield in large coal is substantially increased, or, putting it another way, the breakage of coal incident to working steep pitching measures with full chambers is substantially reduced by the method described. I have seen this same method successfully conducted at Bankhead, Alberta, where the necessity for care in the handling of coal, by reason of its friability, is imperative.

The practices described are an advance upon the ordinary methods that for many years have been practiced in the anthracite region of Pennsylvania.

Simon H. Ash (author’s reply to discussion).—At the present time I do not know of any mine in the heavy dips of western Washington that is working with chutes such a seam that the dip would be from 20 to 25 per cent., as they would be prohibitively long from the standpoint of maintenance. To run cars inclines have been driven on a dip of 20° to 30°, Where the coal is run down a chute, the chute has a dip of 30° to 45°.

In the Pierce County district, the coals are very friable, breaking easily. Although lump coal is desirable, it is not available and the low ash content becomes the desirable factor rather than the larger sizes. The cleaner coal is found in the smaller sizes for where the coal does hold together it is due to a binder of bone or rock. A growing practice is to crush all coal over 2½ in. round opening, to separate the coal from the bone and rock, and then wash the resultant product with fine coal jigs and tables, the latter giving the lowest ash product. This eliminates the expensive rock picking on picking tables, only enough men being used to remove niggerheads and timbers that are in the mine run. A Bradford breaker at one mine is reducing this cost still further. The following shows the average percentage of sizes as compared with the ash content :

The chutes are worked full more to protect the ribs and brattice of the chutes from rock, niggerheads, and running coal. In some instances, it is necessary to place lagging in the chutes to permit the coal to build up above them so that the running coal will not wear the bottom, causing the bottom to wear and become lost. Further, the chutes 4 to 10 ft. wide are driven at considerable expense and the profit is made in pillar coal. When possible breasts (15 to 40 ft. wide) are worked, as they yield more and cheaper coal; but as a rule the walls will not stand and the maintenance is high.

If the coal is firm enough to yield lump, there is little danger of the ribs sloughing off and angle chutes or breasts can be driven on any grade. However, the system usually followed is to drive a chute or incline on such a grade that cars can be handled; they are then called planes. The rooms are then driven on the strike of the seam and the coal won from these planes and rooms. As in other localities, a set is called a panel or battery. Planes are placed about 600 ft. apart and electric hoists are used for handling the cars on the plane.

Chloridizing Plant

The chloridizing mill of the Standard Reduction Co. is located about 75 miles south of Salt Lake City on the Tintic branch of the Denver & Rio Grande Western R. R. and 12 miles from the Tintic Standard mine. The daily capacity is 200 tons of a siliceous, low-grade, silver-lead ore from this property. It has operated continuously since it was started in January, 1921.

The process consists essentially of a chloridizing roast followed by a percolating leach with a nearly saturated solution of common salt, acidified with sulfuric acid, the precipitation of silver on sponge copper and of copper and lead on tin-plate cuttings. The precipitates are shipped to a smelter. Some of the general ideas involved are said to have been used by Augustin in England, in 1840. A number of textbooks treat of the subject, especially the chloridizing roast followed by a leach with sodium hyposulfite or amalgamation. The process was revived in this district by Theo. P. Holt, N. C. Christensen, the Bureau of Mines, and others.

Nature of Ore Treated

The average assay of the ore treated during 1924 is as follows:

Gold, ounces per ton……………………………………………………………0.025
Silver, ounces per ton…………………………………………………………..18.26
Copper, per cent……………………………………………………………………..0.30
Lead, per cent………………………………………………………………………….5.00
Silica, per cent………………………………………………………………………65.00
Iron, per cent…………………………………………………………………………10.00
Lime, per cent…………………………………………………………………………0.70
Sulfur, per cent……………………………………………………………………….3.00
Arsenic, per cent…………………………………………………………………….0.70

The silver is finely disseminated and occurs as native, combined as a sulfide and, to a very small extent, as the chloride. The lead may be present as carbonate, sulfide, or sulfate.

Preparation of Ore for Roasting

The ore is received in standard, bottom-dump, railroad cars, crushed to 3 in. by a Kennedy 6F gyratory, then to ¾ in. by a 36-in. horizontal Symons disk. Finally the ore, with 8 per cent, salt, is run through two sets of Allis-Chalmers rolls, 16 by 48-in., working in series and in closed circuit; the final product passes through an 8-mesh screen with a clear opening of 0.071-in. Three Mitchell and two Colorado impacts are used in the roll circuit. The results of a screen test on the roll product and the distribution of the metals are as follows:

After grinding, the ore-salt mix is sampled by a mechanical sampling device in batches of 70 tons, each batch being run to a separate bin. For the purpose of furnace control, the sample is tested for its reducing power on litharge, which test indicates its fuel value. The latter is then adjusted to suit the requirements of the subsequent roasting operation by the addition of coal dust; this usually amounts to between 1 and 2 per cent. Before passing to the bins over the roasters, the mix is moistened with just enough water so that it will stick together as a ball when pressed in the hand. The actual amount of water needed will vary according to the fineness of the ore, but is approximately 7 per cent.; the ore, as received at the mill, has a moisture content of 2 to 3 per cent. The mix is now ready for the roasters.

Ore Roasted in Holt-Dern Furnaces

There are nine Holt-Dern blast roasters. These consist essentially of a row of reinforced-concrete boxes 7 by 9 by 5 ft. deep inside, set end to end; on the bottom are mechanically operated grates with hoppers underneath. On the long side, and 30 in. above the grates, are two double work doors which run the full length of each furnace. Above the furnace are the charge bins, with four segmental gates for each furnace. Leading into the hopper under the grates is a pipe through which an air blast is supplied at 8 oz. pressure by a direct-connected Sturtevant fan. A common flue, through which the gases are drawn, runs the full length of the furnaces.

These furnaces are operated as follows. Starting with a bed of hot calcines, about 10 in. deep, on the grates, sufficient ore mix is let down from the bins to fill the furnace even with the bottom of the work doors. After leveling, by hand, the air gate is opened and the sulfur and coal in the charge, ignited by the hot calcines on the bottom, gradually burn upwards, and as a rule, quite evenly over the whole area of the furnace. This takes a little over 3 hr.; at the end of this time, the whole mass is at a dull red heat, or about 700° C. The air gate is then closed and the grates put into, motion so that the charge is shaken into the hopper below, leaving enough hot calcines on the grates to ignite the next charge. The operation is then repeated. With each “drop,” 4¼ tons of calcines are obtained or 25 tons per furnace per 24 hr.

Leached by Percolation

As soon as convenient after shaking the calcines from the roasters, the gates at the bottom of the hoppers are opened and the calcines run into a concrete launder through which a stream of brine is flowing. This flushes the calcines into one of six concrete leaching tanks. These tanks are 28 ft. in diameter by 11 ft. deep, inside, and have a filter bottom, made up of crushed quartzite and two 3-in. earthenware cocks for discharge. A tank will hold about 225 tons of calcines when filled to within 8 or 10 in. of the top. After leveling, leaching is commenced. The effluent liquor is received in two concrete sump tanks of the same size as the leaching tanks. The first, and richer, part of the solution is received in one of these and is designated “pregnant solution.” It ordinarily carries about 3 oz. per ton of silver and 14 lb. of lead. The subsequent solution is received in the second tank and is called “weak.” This weak solution is used for sluicing the calcines from the roasters and for the first 24 to 48 hr. of the leaching period. After precipitating the metals from the pregnant solution, a barren liquor is obtained; this is used as the second leach solution over the next 48 hr. period, being received in the weak sump after passing through the leaching tank. Finally, each tank is washed for 8 hr. with water, to replace the last solution, then drained and sluiced through two bottom gates to the sump.

Summary of Leaching Cycle

The amount of solution that will run through a tank of calcines in 24 hr. varies from 200 to 300 tons.

Precipitation on Sponge Copper and Scrap Iron

The pregnant solution is pumped by air lifts from the sump tank to the silver precipitator. This really amounts to a four-compartment Pachuca tank with an air lift in each compartment for agitation. Each compartment is 11 ft. 4 in. by 11 ft. 4 in. in cross section and 10 ft. deep, to a pyramidal bottom, which adds 8 ft. to the over-all depth. It is built of reenforced concrete. In this, the solution is agitated with sponge copper to precipitate the silver, and flows through the four compartments in series; the fine copper is added intermittently as needed. When a sufficient amount of silver has accumulated in the first compartment, the solution is bypassed, that remaining in the compartment is decanted and the precipitated silver run to a filter box. Before shipping, this material is treated as noted later.

The effluent from the silver precipitator runs to eight concrete boxes, varying in depth from 18 in. to 3 ft., by 5 ft. wide and 30 ft. long, and filled with tin-plate cuttings. The boxes are provided with wooden grids, on which the cuttings rest, and with four baffles to interrupt the flow. Additional scrap is put in every day and one box is thoroughly cleaned each day, the precipitated copper being washed to a settling box. Part of this is used as the precipitant of the silver; the remainder is shipped to the smelter. It will contain about 100 oz. of silver per ton and 50 per cent, copper.

The solution flowing from the eight copper boxes is pumped with an “Olivite” centrifugal to a rectangular concrete tank containing about 1260 ft. of 1¼-in. copper pipe through which low-pressure steam is passed. Thus the solution is brought to a temperature of 75° C. and is then passed through fifteen additional boxes, similar to the copper boxes and likewise filled with tin-plate cuttings. In these, the lead is precipitated. It is necessary to add live steam also to these boxes, to maintain the temperature. The boxes are kept as full as possible with the cuttings and two of them are cleaned each day, the cuttings being removed and washed and the lead sluiced to a drain box; this is shipped to the smelter without drying. A partial analysis is given below:

Gold, oz……………………………………0.005
Alumina, per cent……………………3.00
Silver, oz……………………………………..6.58
Zinc, per cent……………………………0.50
Lead, per cent………………………..70.27
Arsenic, per cent…………………….0.10
Copper, per cent……………………..5.12
Antimony, per cent…………………0.10
Insoluble, per cent…………………..1.70
Ron, percent…………………………….5.52
Moisture content as shipped 21.38 per cent.

When using tin-plate cuttings, it has been found advantageous to remove first the tin coating; this is accomplished by treating them with a solution of caustic soda containing a small amount of litharge. The following reaction is involved.

2PbO + 2NaOH + Sn = Na2O.SnO2 + 2Pb + H20

The lead oxide is obtained by simply heating the lead precipitate in contact with air. The bales of scrap are loosened, placed in shallow iron boxes, and the caustic solution circulated through it, having in the circuit a small steam coil for heating. The tin at present is not recovered. The method is outlined in Schnabel, Vol. 2, page 541.

Additional Treatment given Silver Precipitates Before Shipment

The precipitate, as taken from the silver precipitator, will run 30 per cent, silver (8750 oz.), 15 per cent, copper, 2 per cent, lead, 25 per cent, arsenic, and 1 to 2 per cent, antimony, the remainder being largely insoluble, iron, and alumina. After washing and draining, this is placed on a small reverberatory hearth and heated slowly with an oil flame to dry. The temperature is then increased somewhat, when about 60 per cent, of the arsenic will be volatilized, the fume being caught in small bags. When the fumes are no longer emitted, the material is brought to a dull red heat and the copper oxidized. The product is then removed from the furnace, the lumps broken, and leached with a hot 25 per cent, sulfuric acid solution; this reduces the copper to about 1 per cent, and the arsenic to less than 0.75 per cent. Finally it is dried, sacked, and shipped, by express, to the smelter. It will run from 10,000 to 14,000 oz. silver per ton.

Recoveries and Costs

The recoveries of both silver and lead have gradually improved and, at present, the following can be consistently obtained. Gold none; silver, 89.8 per cent.; lead, 65.7 per cent.; copper, 52.2 per cent.

The following cost data represent the average for the year, 1924.

Department 1, unloading, crushing and grinding; 2, roasting; 3, leaching and precipitating silver and copper; 4, silver-product treatment; 5, lead precipitation; 6, chemical laboratory; 7, undistributed; 8, office and supervision.

Average labor wage per 8 hr. day…………………………$ 5.00
Salt costs, f.o.b. mill, per ton……………………………………..4.00
Slack coal, f.o.b. mill, per ton…………………………………….3.05
Tin-plate cuttings, f.o.b. mill, per ton……………………18.00

The cost of all experimental work to improve recoveries or operation is included in the above.

Miscellaneous Observations


Probably the roasting operation is the most satisfactory step in the process, whereas formerly roasting apparently caused much trouble on account of the volatilization of the silver and the skill required to obtain good chloridization. Using this method, with a reasonable amount of attention, there is only a negligible silver volatilization loss and a good conversion usually results. While it is an intermittent operation, two men per 8-hr. shift will roast 75 tons of ore and have time to spare.

As in most furnace operations, some points must be carefully watched, especially those regarding the preparation of the mix. While 8 to 10 per cent, of salt is used, a satisfactory chloridization may be obtained in the furnace with 5 to 6 per cent.; the balance is used to maintain a high chlorine concentration in the leaching solution. No detrimental effect has been observed when using this large excess, unless it is when the furnaces are running a little hot; then the salt may fuse, making the calcines slightly more difficult to shake through the grates. A thorough mixing of the salt is essential. The salt used is commonly known as “smelter salt” and is obtained from the Morton Salt Co. at Burmeister, Utah. It is shipped in bulk in bottom-dump cars and is handled the same as ore. In size, the crystals vary up to possibly ½ in. Salt containing a large proportion of fines, or “dairy salt,” is more difficult to pass through the plant, as it hangs up in all the bins. The salt, as received, is quite pure, samples usually showing a chlorine content equivalent to 97 per cent: NaCl.

It has been found in blast-roasting Tintic Standard ore, that for the most favorable operating conditions the sulfur permissible in the mix must be between 2 and 4 per cent. Good chemical results have been obtained with 1 per cent, sulfur, the balance of the fuel necessary being made up with fine coal; but the tendency on low-sulfur charges is toward uneven burning and a rapid loss of heat during the recharging period; Above 4 per cent, sulfur, Standard ore fuses too easily so that the resulting calcines are caked or sintered in hard lumps, which require a long time to shake through the grates. It is not the additional sulfur in itself that causes the fusion but the fusion point of the whole charge is lowered by the addition of the sulfide ore. With a high-sulfur charge, when the rate of burning is decreased by decreasing the air supply, the fusion still takes place.

The sulfur content is determined by fusion with litharge; this is supposed to be the sulfide sulfur but, of course, any other material that will reduce litharge will be reported as sulfur. It is the heating value of the charge that is sought. Some idea of the requirements in the charge may be obtained by noting that with a 3 per cent, sulfur 1½ per cent, coal is used; and as the sulfur changes, the coal is varied using the ratio sulfur: coal = 1:0.65. This method of adjusting the fuel value is purely empirical but commercially uniform results are obtained. It would be desirable to have a calorimeter determination on the roaster charge, but this is difficult for the total heat value is so low that an undue portion of some substance with a high heat value must be added in order to get the calorimeter charge to burn. However, some good work along this line has been done by the Salt Lake station of the Bureau of Mines and it is quite possible that a more satisfactory method for furnace control will be worked out.

The quantity of water iii the mix is important. Water is added primarily for the purpose of agglomeration and so forming a more porous orebed. It is supposed to assist in the chloridization, also, by the formation of hydrochloric acid. Too much water makes a hard calcine; too little makes slow roasters with a tendency to be “spotty.” If the ore just sticks together when pressed in the hand, it is about right. This is another unscientific procedure but moisture determinations are useless; 7 per cent, water is close to the amount usually needed with the present size of ore feed.

It would probably be difficult, surely slower, to blast-roast the ore if ground finer than it now is although no work has been done with increased blast pressure.

It does not require a great deal of skill to operate the roasters properly but the results obtained are largely dependent on the conscientiousness of the firemen. The roaster should be “dropped” immediately the operation is completed, thus the heat in the charge is conserved and a good ignition obtained on the next round, also; the calcines remaining on the grates should be leveled and care taken that there are hot calcines over the whole earth area. Spots that are a little cold should be covered with hot calcines from one of the other furnaces, or the next charge will develop dead spots, which must then be shoveled out or poor results obtained.


Presumably, the silver has been converted to the chloride or sulfate “in the roasters. (About half of it is soluble in strong ammonium hydroxid.) A strong brine is used for leaching but after having made the round trip through the plant a few times, it contains small amount of many substances. A number of determinations are made every day on the pregnant solution, the following being an example.

Specific gravity………………………………………………………………1.24 to 1.30
Acid, lb. per ton expressed as H2S04………………………….2 to 5
Silver, oz., per ton………………………………………………………………..3 to 5
Lead, per cent. (10 to 30 lb. per ton)…………………………0.5 to 1.5
Copper, per cent……………………………………………………………………0.1
Chlorine, per cent…………………………………………………………….12 to 15
Sulfur, per cent……………………………………………………………..0.75 to 1.2
Iron, per cent…………………………………………………………………………1 to 2

The amount of “ic” salts in solution is so small as to be almost indeterminable, the solution oxidizing very slowly in contact with air. This is unfortunate as the higher oxidized forms of both iron and copper, when dissolved in the brine solution, are good solvents for metallic silver and the sulfide, should these chance to escape the action of the roasters.

It is interesting to note the effect of the addition of very small amounts of copper sulfate to a fresh brine solution in its action on Tintic Standard ore without roasting, as shown by Fig. 1. The sample was ground to pass 120 mesh and leached by agitation, first with brine to which different amounts of copper sulfate were added and then, for the sake of comparison, a second portion with a brine carrying different amounts of sulfuric acid.

A small amount of free acid is necessary for consistent results in the solution of the silver. At times, especially when a rapid leach is made, a neutral brine dissolves the silver, but in the routine of plant leaching it is decidedly unsafe to allow the solution to approach neutrality more closely than is shown in the analysis. The silver is dissolved when using a solution short of acid and is then precipitated in the leaching tank, for while the top portion of the tailings will have a normal silver content the lower portions will steadily grow richer until they contain more than the original heads. But a small part of this silver can be dissolved when the tails are subjected to further leaching with solutions highly acidified.

Just what substances in the calcines cause this precipitation have not been determined. Lime, which exists in the ore up to 1.5 per cent, zinc, which seldom runs as high as 0.3 per cent., and metallic iron, introduced in grinding, have been investigated as possible interfering elements but no definite data obtained show that any of these could be the cause of the trouble.

The acid content of the solution is maintained for the most part by the direct addition of 66° sulfuric acid, although a small amount is absorbed by passing the solution through a spray chamber in the roaster flue system.

Iron even in the “-ous” state probably aids in the solution of the silver. Total iron seldom builds up as high as 2 per cent, in the solution in spite of the fact that all the precipitating of the metals is, in reality, done with scrap iron and no effort is made to remove it.

Trouble has been experienced when working with a solution saturated with respect to salt; i. e., one from which salt will separate on standing a short time. Being a denser, more syruplike liquid, it percolates more slowly and it does not seem to have the dissolving power for silver that a slightly weaker solution has. This is contrary to accepted solubility data as to silver chloride in brine. More acid will not correct the trouble; in fact this difficulty is not at all times apparent. No good reason has been found as to why this is so.

The mill solutions will carry between 25 and 30 oz. of silver per ton; as this concentration is not approached in practice, the solution has ample carrying capacity for silver. Dissolving the lead, however, is quite a different problem. Nearly all of the lead in the calcines is considered as being present as the sulfate and not as the chloride. It is well known that the amount of this substance that a brine will carry is dependent on the solution temperature, chlorine concentration, and sulfate content. The most difficult of these to control is the sulfate content and, while a number of schemes for removing this have been suggested, including freezing, evaporation, and the addition of various reagents, few have much merit commerically. By keeping the sulfate content of the leaching solutions down to 2 per cent, or under, expressed as Na2S04, it would be possible to obtain about 1¼ tons additional lead per day.

The effect of sulfates is shown in Table 1 (and graphically in Fig. 2), to obtain which, an excess of lead sulfate was left in contact with a brine solution containing 26 per cent, salt and the different amounts of sodium sulfate given, until it would dissolve no more. Also, the improvement by increased temperature is shown. It will be noted that the solutions with the large amounts of sulfate are not benefited as greatly by raising the temperature as those low in sulfates.

Calcium chloride was long ago recommended as a precipitant for these objectionable sulfates, but 70 to 75 per cent. CaCl2 costs $41.90 per ton delivered to the mill and it would require at least 5 tons a day. That this reagent improves the solution as a lead solvent is shown by the following experimental data. In this case, the plant solution was treated with different amounts of the calcium chloride to obtain the varying sulfate contents indicated. Lead sulfate was left in contact with frequent stirring until the solution would no longer dissolve it.

Slacked lime may be used in place of part of the calcium chloride, but alone it apparently acts as a precipitant for the sulfates only when there is iron in the solution, the resultant precipitate being a basic sulfate of iron. This is a very disagreeable material to handle as it is bulky and gelatinous. Also, lime acts very slowly and requires long agitation with the liquor to obtain efficient results. Finally, it is preferable to carry iron in solution.

To discard enough solution each day to control the sulfate content has been suggested, but this had no attraction commerically as it would require some 50 tons of salt.

About the most feasible plan, probably, is to increase the number of leaching tanks, thus allowing sufficient time to pass the desired amount of solution through the calcines; then the only added operating expense would be the cost of circulating the solution. In line with this, one of the tanks was held in the mill circuit for nine days as an experiment and an extraction of 92.5 per cent, of the lead was obtained.


Each day 1000 tons of pregnant solution are delivered to the precipitating department and the silver precipitated first by means of the copper afterwards obtained in the iron boxes. Working in this manner, the copper is never completely replaced by the silver. When the material reaches a copper content of about 15 per cent., the remaining copper behaves as though it were coated with some protecting substance and the silver begins to dissolve. The difficulty has been attributed to arsenic which is thrown down in the metallic state in both the silver precipitator and iron boxes. The remaining copper is not soluble in weak acids.

Of the precipitation of copper on iron, little need be said as the operation is common practice. As the copper product is used in an agitating apparatus with a continuous overflow, it is desirable to have it coarse or granular so that it will not float out of the silver precipitator. Large pieces of cast scrap give a more granular precipitate than light tin plate but the latter has the advantage of increased surface and makes a reagent free from adhering foreign matter that usually accompanies ordinary scrap iron. On the other hand, where it is necessary to keep the precipitator boxes filled at all times, the cuttings are more difficult to wash.

In the precipitation of lead on iron, the solution must be maintained at a relatively high temperature in order to get a sufficiently rapid action; 75° C. secures satisfactory results with the present precipitating capacity, but the solution must come in contact with the iron and not allowed a chance to bypass. As now conducted, the cleaning of the boxes calls for a high labor charge, but without doubt this can be greatly improved should it be decided to continue the use of tin-plate cuttings as the precipitant.

Structural and Mechanical Features

The process described was adopted after numerous tests made on the ore with various processes, such as concentration, flotation, volatilization, and cyanide, as it gave a higher and more consistent recovery at a reasonable cost than any of these. Owing to its nature, however, materials that could be used in the construction of the plant were practically limited to wood, siliceous concrete, and rubber. The structural and mechanical features may be of some interest. The general flow plan, Fig. 3, approximately indicates the arrangement and flow of ore and solu-

tions through the mill. The general ground arrangement of the plant, which is situated on a hillside, having a slope of 29° is shown in Fig. 4.

The railroad, entering the plant below the main mill building, delivers ore, salt and coal to bins, and the preliminary crushing plant, from which they are hoisted, in 45-cu. ft. skips, up a double-track incline, to a conveyor distributing to the storage bins at the top of the mill. A service tramway, with skip operated by a 50-hp. hoist, runs from the bottom of the hill alongside the mill building to the top ground floor, serving all floor levels, together with the machine shop, laboratory, warehouse, crushing plant, and carpenter shop, which are situated along this tram. The warehouse is also on the railroad; and all materials received can be delivered to any department of the plant with this skip.


The buildings are of wood, the sides being covered with a double thickness of 1-in. boards, with 40-lb. building paper between, and the- roofs with extra heavy Rubberoid laid on 1-in. boards.

The main mill building is approximately 282 ft. long and 182. ft. wide, with an extreme height of 40 ft. and an average height of 24 ft. There are no special features in the design, but all floors or sections spanned by trusses are of the same width, 33 ft., so that all trusses are exactly alike, which results in economy in construction. The slope of the roofs was

so made that, with this span, a minimum of material; consistent with required strength was realized. All bents are 13 ft. wide. All foundations and retaining walls are of reinforced siliceous concrete, and all the ground in the wet part of the mill is covered with a concrete coating, terminating in a general drainage sump in the lower end of the mill. Figs. 5 and 6 show the general plan and general cross-section, or sectional elevation of the mill.


The tests indicated that the finer the crushing, the better was the recovery; but at the same time a granular product was necessary for the roasting and leaching operations. So in the design,. attention was first directed to this step in the process and a crushing scheme was adopted and equipment selected that would fulfill this condition to the fullest extent possible. These are indicated on the general flow plan; the screen analysis previously given shows the product realized.

A gyratory followed by a Symons disk for the preliminary or coarse crushing, and large rolls and screens, in series, the screens preceding the rolls, so that fines are eliminated as fast as produced without further grinding, were adopted as being the most suitable for producing the desired result. The centrifugal action of the Symons in immediately discharging everything below the size to which the disks are set, produces a minimum of fines; and as the ore is dry, the machine gives no particular trouble. One set of manganese-steel disks crushes between 40,000 and 50,000 tons. An electromagnet is suspended over the short conveyor belt between the gyratory and Symons disk to remove tramp iron. Expressed in terms of original ore, the life of the coarse roll shells is about 25,000 tons; and of the fine roll shells about 30,000 tons. Each shell weighs 3000 lb.

The gyratory and Symons disk are driven from a line shaft by a 50-hp. motor, and run about 5 hr. out of the 24. A 125-hp. motor drives the rolls and elevators through a line shaft; they run about 20 hr. daily.

The fine ore and coal are withdrawn from the fine storage bins into a hoppered scale car, carrying 3000 lb., which is propelled by trolley and discharges into an ordinary tilting concrete mixer, which, in turn, delivers through hopper and belt feeder to an elevator. A belt conveyor receives the discharge of the elevator, and delivers it to a paddle mixer, where it is moistened. A shuttle conveyor directly beneath the paddle mixer distributes the product to the roaster bins.

Holt-Dern Roasters

The inception and early development of the roaster is described by Theo. P. Holt; as used in this plant, it is shown in Fig. 7. There are nine of these—seven in one bank and two in another. The general structure is of reinforced, siliceous concrete. The roasting chamber sides are ¼-in. steel plate, lined with 6 in. of concrete; the ends, which are formed by the 10-in. partition walls, have also an additional 6-in. concrete lining. This lining gradually disintegrates and must be renewed about once a year.

At the bottom of the roasting chamber, there are fifteen rocking grates, 7 ft. long spaced 7¾ in. apart. Each consists of a 2¼-in. square steel shaft, passing through the cast-iron grate bars, which are made in sections 21 in. long, and are shown in cross section. They have four longitudinal ribs, 1½ to 2 in. high, 90° apart, the vertical ribs being solid, while the horizontal ribs are notched for free passage of the air through the grates. When the top rib is worn down, the shaft is turned over and the opposite rib used; when both are worn they are replaced with new bars. These bars wear about 1½ year. The grates rock through an angle of 60°, and are actuated in pairs by segmental gears on each shaft, which are given a reciprocating motion through connecting rods by two main cranks. These are revolved through bevel gears and pinions from a line shaft, each pinion being attached to a friction clutch, which is keyed to the shaft. Thus half of the grates in a roaster can be operated at a time, and the starting load is only one-half as large. The grate shafts pass through stuffingboxes, with glands in cast-iron plates, in each side of the roaster. On the gear, or driving, side, they turn in rigid bearings that are supported on a cast-iron filling piece resting on a 12-in. I-beam; on the opposite side, the glands of the stuffingboxes serve as bearings. The bearings are specially designed, with caps fitting

accurately in deeply machined grooves and a dowel between the bases and the cast-iron support, so that the shafts are held firmly in place. The gears are thus held in mesh always on their pitch lines and there is no sliding contact, so that the wear of the teeth is reduced to a minimum and any lost motion is prevented in the movement of the grates—an important feature. The whole mechanism is so designed that any bearing, or any other part, can be quickly and easily repaired or replaced, so that there may be no delays in the operation of the roaster and the roast can always be quickly discharged. The line shaft is driven by a 15-hp. motor on each end, one of which is a spare; from 2 to 5 hp. is required for each motor after starting. If the charge is “hard,” or partly sintered, double this power is sometimes required to start the grates. From 5 to 15 min. are required to “drop” or discharge the roasted charge.

The Sturtevant gas blower, supplying air to the nine roasters at 8 oz. pressure, has a capacity of 15,000 cu. ft. per min., and is direct connected to a 75-hp., 1800-r.p.m. motor. There are two of these, one being kept in reserve.

The gases issuing from the roasters have a temperature of only 35° to 55° C., so that an exhaust fan is necessary to remove and discharge them through an absorbing chamber and short stack. This fan is

72 in. in diameter, 35 in. wide, and is housed in a concrete casing. It is driven by a 15-hp. motor at 250 r.p.m. Originally, the fan runner was completely rubber covered; now only the spider is rubber covered and the blades, which are of 3/16-in. steel, are painted with six coats of elaterite paint. This coating lasts three to four months. The stack is 5 ft. in diameter and 40 ft. high and rests on top of the fan casing. It is made of 3 by 6 in. plank with round iron bands, similar to wood-stave pipe.

Leaching Tanks, Precipitating Boxes, and Tanks

These are all of reenforced, siliceous, concrete, and great care was exercised in the designs, preparation of the materials, and placing of the concrete. The aggregate was composed of crushed quartzite, taken from one of the mine dumps, and siliceous sand, part of which consisted of fines from crushing the quartzite, the SiO2 content being 96 per cent. The walls and bottom of tanks are 8 in. thick, while those of the precipitation boxes are from 5 to 7½ in. Test blocks were made in all cases, and a mixture made up to stand 2500 to 3000 lb. per sq. in. in compression, while the reenforcing steel was calculated on a basis of 10,000 lb per sq. in. safe tensile strength. The true mix varied from 1:3 to 1:4, and the maximum size of the aggregate was 1¼ to 1½ in. The proportions of coarse and fine aggregate were about 65 and 35 per cent., respectively, and none of the concrete has shown any penetration or leakage of solution. There are two circular discharge holes 13 in. in diam. in the bottom

of each tank; and on the underside a circular dovetail groove, 5/8 in. deep, 1¼ in. wide on the bottom, and 1 in. on top was cast around each hole. A soft-rubber packing ring to fit this groove, and thick enough to project 5/8 in. from the concrete is inserted. The gate, made of two thicknesses of 3-in. plank, doweled together with wooden pins, is drawn up against the rubber packing ring by means of a rubber-covered 1¼-in. rod extending through the tank and a beam 2 ft. above the top, which is supported, by two posts resting on the tank bottom. The two 3-in. stoneware cocks for discharging solution are screwed into wooden nipples set in the sides of the tank just at the bottom. The quartzite gravel, forming the filter bottom, rests on triangular strips that are supported by 2 by 4 in. pieces lying on the tank bottom. These strips were made by ripping 6 by 6 in. sticks diagonally, and are held apart by small pieces.

The silver precipitators were made of about the same mix as the leaching tanks, with the same proportion of steel; but a 1:5 mix was used in the construction of the copper and lead precipitation boxes. There is no deterioration or penetration of the concrete, even by the hot solutions in the lead precipitation boxes; but as these are decanted and washed out at frequent intervals with cold water, cracks occasionally develop, which are simply chipped out and filled in with a 1:1 cement mortar. On the other hand, the walls of the absorber chamber, where the roaster gases are drawn horizontally through falling sprays of solution, do not endure. The cement is gradually dissolved and the walls disintegrate. They are now protected with elaterite painted plank, which lasts over a year. The sides of the fan casing have also softened to a depth of ¾ in., but they remain this way without disintegrating and it has been unnecessary to repair them.

Air Lifts

Solutions are pumped from the sump tanks at the lower end of the plant to the sluice launders under the roaster hopper discharge gates, the absorber, precipitators, and the leaching tanks with four air lifts, made of bored redwood pipe, two of which are 5 in. and two 4 in. inside diameter. These lifts are supported by the tower shown near the lower end of the mill in Fig. 6. The net lift is 60 ft. above the top of the tanks, and submergence is obtained by a two-compartment concrete-lined shaft 60 ft. deep, connected with the sumps. A 1-in. rubber-covered air hose enters each of the pumps a few feet above the top of the tanks, and extends down to about 18 in. from the bottom. The pipe lengths are from 8 to 14 ft. and the joints of the submerged portion are held together with four 2 by 10-in. planks 5 ft. long, doweled to the pipe with wooden pins. Iron clamps, 2 to 3 ft. apart, are put on the unsubmerged part to prevent splitting. Each pump terminates, on top, in a discharge barrel 32 in. in diameter and 42 in. deep, which discharges into a launder. Air is supplied at 40 lb. pressure for the air lifts and silver precipitators by three motor-driven compressors, having a total capacity of 900 cu. ft., which is more than ample, so that when repairs are necessary on any one, pumping is not interfered with. The 5-in. pumps will each handle 130 gal. permin., and the 4-in., 110 gal. These air lifts are quite satisfactory, except that occasionally a crack develops in the submerged portion, requiring the pulling of the pump and replacing of the broken length; which is slow, arduous, and somewhat expensive. Two 2-in. centrifugal “Olivite” pumps handle barren solution with lifts of about 20 ft. In these, the casing is lined and the runner covered, with a composition that is not affected by the solutions. They discharge through rubber-lined pipe, which so far has shown no deterioration. These pumps may, in time, replace the air lifts.

The solutions are carried about the plant in wooden launders made of 2-in. plank, held together with clamps, but the wood shrinks and softens, especially from the hot solutions, making it very difficult to prevent leaks. To overcome this, the solution launders are lined with a cement mortar, made of one part siliceous sand, one part quartzite gravel of a maximum size of 3/8 in., and a two-thirds part of cement. Wire netting, 1-in. mesh, is placed in the launder, conforming to the sides and bottom, together with some ¼-in. rods, the launder partly filled with the mortar, and the inside forms, made in 6- or 8-ft. lengths, are then set in and pressed into the mortar, forcing it up on the sides, when it is rammed and levelled up to the top. The bottoms of the sluicing launders carrying the hot roasted ore in solution are lined with concrete slabs of the above mixture, 30 by 16 by 1½ in., cast in separate molds and cured for several days. Slabs 7 in. wide are used on the sides. A concrete air agitation tank, 8 ft. deep and 4 ft. in diameter, in which the oxidized copper is leached out of the silver precipitates with a hot 25 per cent, sulfuric-acid solution, slowly disintegrated, but a sulfur-and-sand lining, made of equal parts of melted sulfur and fine sand, thoroughly mixed and poured in a form, making it 1½ in. thick, has stood very well. Care must be used to keep the temperature of the sulfur just above the melting point, and to pour the mixture quickly before the sand settles out.

Heating Solution

Steam for heating the solutions is supplied by three boilers, two of which, aggregating 110 hp., supply steam to the copper coils, the condensation returning to the boilers through a trap. The coils become coated with a hard scale, precipitated out of the solution, which gradually lowers their efficiency until it becomes necessary to knock off the scale. The other boiler, a 125-hp. return tubular, supplies live steam direct to the solutions in the precipitation boxes, after it has passed through the coil box, and to any other places where steam is required. While this is the normal method of running, a double system of steam piping permits the steam from any one, two, or the three boilers to be run to either the coil or live steam systems. The three boilers consume about 11 tons of slack coal per day, and 1000 tons of solution |are heated from 48° to 75° C.


Water supply for the plant is pumped from a spring at the base of the hill to a 140,000-gal. wooden storage tank above the mill, by a 250- gal. per min., triplex pump, driven by a 30-hp. motor, running 18 hr. daily; the net lift is 225 ft.

Power is delivered to the plant at 44,000 volts, by the power company, transformed to 2200 volts for all motors over 30 hp., and to 220 volts for all under that size. The motors are all wired with three-conductor lead-covered cable, and all wiring for lights is lead encased. The total con¬nected load amounts to 700 hp., but the average maximum demand reading is about 500 hp.

Well-equipped machine and carpenter shops enable practically all repairs to be quickly made at the plant. As the process as a whole is rather destructive, the repairs are considerable, but the plant has now been in operation 4½ years without a shutdown.