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.
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
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.
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.
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.
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.
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.
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.
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:
Alumina, per cent……………………3.00
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
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.
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.
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.
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.
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.
To the average person, the purport of the items and figures on the balance sheet of a mining company are hazy and the real financial condition of the company is cloaked in obscurity. It is also likely that the satisfaction of the stockholder in his investment, as well as the determination of the prospective investor, rests more on the reputation of the company, his confidence in the management, the stated earnings per share and the dividends paid, than on a comprehension of the information contained in the balance sheet.
The balance sheet of any company at best can only closely approximate the true financial condition as of a given date. Any item of asset, except cash, can probably be altered if another’s viewpoint is applied. It is difficult and expensive to drill and explore an orebody to determine the entire tonnage of ore therein. Those inherent difficulties render it practically impossible to prepare an absolutely correct balance sheet of a mining company.
It is possible, however, to present the facts of a mining company more intelligently and comprehensively than is usually done. With this in mind, a balance sheet of an imaginary mining company is submitted, with supporting schedules.
The same fundamental principles hold true whether we consider a large or a small company. Less complexity is involved in dealing with the affairs of a smaller company. Therefore, for this purpose, the assumed company operates but two mines and a smelting plant. It has been deemed advisable to introduce the features of the operation of a mine acquired prior to March 1, 1913 and one after that date. The presentation therefore is concerned in one mine with the appraisement as of March 1, 1913 and in the other with the appraisement by reason of discovery. A consequential factor pertains to dividends paid from depletion reserves and therefore free of surtax to the stockholders.
While supporting schedules, when necessary, give details of certain items on the balance sheet, it seems desirable to give further explanations with respect to some of them. This seems to be important because some mining companies do not treat the items as indicated on the statements submitted.
Mining Properties.—These properties are shown in the Assets at their appraised values, less depletion sustained.
Concentration and Smelting Plants.—These plants are shown in the Assets at their cost less depreciation.
Unrealized Appreciation.—This item is frequently included by mining companies in the Surplus Account. In the writer’s opinion this treatment anticipates profits and is misleading.
While the scope of this paper does not contemplate the consideration of the mechanics of accounting, it may be of interest to state that the mine product is charged with the depletion based on the original cost and like credit given to the Depletion Reserve Account pertaining to the original cost. This reserve account, while carried separately on the books, is deducted from the original cost on the balance sheet. The depletion based on the appreciated value is charged to Profit and Loss Account and credited to Depletion Reserve Account pertaining to appreciated value. While a separate account for such depletion is kept on the books, the depletion reserve is deducted from the original appreciated value in preparing the balance sheet, as is indicated on Schedule Nos. 1 and 2.
The yearly amount of depletion on account of appreciated value is transferred from Unrealized Appreciation to Realized Appreciation.
It will be observed that the current earnings since March 1, 1913, have been distributed. Such earnings must first be distributed before dividends not subject to surtax can be declared out of Depletion Reserves or Surplus as of March 1, 1913.
We are dealing with a company that is not in the process of liquidation but one that intends to reinvest in new properties, its depletion recovered from the original cost of the properties now in operation. For this reason the depletion sustained on the original cost is carried in a Depletion Reserve Account, from which dividends are not intended to be paid.