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.