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Jan 1, 1941

Jan 1, 1940

By M. A. Mahtab, J. D. Dixon

The results of a parametric study of the influence of natural features (stress field, rock strength, and strength and orientation of fractures) as well as the influence of induced features (undercut span, boundary slot, and other boundary-weakening measures) on the cavability of an undercut block of ore in mining by the block- caving method are presented. The two-dimensional finite-element analysis technique was used for this study. A coarse "mesh," along with linear homogeneous assumptions for rock properties, was used to model the problem of initiation of caving in a block of ore. Interpretation of the test results suggests that weakening the "abutments" of the block as well as certain orientations of natural fractures exercise significant influence on cavability. These findings hold promise for immediate application to known situations in block-caving mines.

Jan 1, 1977

Jan 1, 1941

By M. E. Tyrrell, Alan H. Goode

The application of a simple, two-dimensional computer technique for evaluating rock burst potential in cut-and-fill stopes was investigated. The principal purpose of the study was to compare the effectiveness of high-modulus fill and destressing as aids in the control of rock bursts. Two conclusions that are probably of general applicability are drawn. First, high-modulus fill is ineffective as a means of increasing the mine stiffness and acts to control rock bursts only if it supports enough load to reduce substantially the stresses on the unmined vein material. Second, destressing is most effective if it is performed when the vein material is near the point of failure.

Jan 1, 1975

By E. Tulcanaza

The practical significance of the in situ grade variability is usually neglected. Although recognized, often times it is either ignored or just considered as a vague guideline in many of the decisions dealing with exploration, reserve estimation, and mine planning. This paper stresses the significance of grade variability for exploration and mining, and illustrates its liaison to the structural features of the ore deposits, as well as to the mining parameters used for mine extraction. Some geostatistical results obtained from a silver mine, tungsten mine, and copper mine in South America are discussed.

Jan 1, 1985

By W. E. Anderson, T. M. Brady

As part of a larger study to identify factors influencing the practical operating life of wire rope used on large draglines in surface coal mining, field trips to operating surface coal mines were made during late 1975. Wire rope performance information was obtained from personnel at 13 mine sites west of the Appalachians. Seventeen pieces of equipment, utilizing bucket sizes ranging from 30 to 220 cu yd (23 to 168 m), were documented. Draglines at large US surface coal mines can be represented as 120 machines having 60-cu yd (46-m3) buckets, moving 1.8 X 109 cu yd (1.4 x 109 m3) of overburden during 6650 hours of productive availability per year. Based on the limited field study, the calculated downtime resulting from rope-related causes averages 105.6 hours per year for each typical 60-cu yd (43-m3) dragline. The total loss in production from this source alone is thereby estimated at about 25 x 106 cu yd (19 x 106 m3) annually. Although study findings did not reveal anything about dragline rope life that wasn't known or practiced by one operator or another, most operators can expect improved rope life by periodic review of their practices and by adopting one or more life-improvement measures employed at other sites. Since rope productivity was found to vary by a factor of five among the sites studied, considerable improved average performance is possible. As little as 10% overall improvement would increase national productivity by 2 million cu yd (1.5 X 106 m3) of solid overburden removal. Associated savings in rope costs would be some $2 million annually, plus a large reduction of labor-related expenses. This report outlines preliminary findings which may prove useful to many operators in extending wire rope life. Some of these modifications or adjustments in maintenance/operating practices would not require capital investment. During the field survey, no company was found which had tried cost reductions by using presocketed dump ropes, or dump rope sheaves that opened like a snatch block, so these are suggested as possible improvements requiring modest capital input. Data Acquisition During the field visits, mine operators provided rope life information about their machines. Some of the rope life data consisted of gross averages; other data was in the form of detailed records, including rope construction and manufacturer, cutoff and resocketing periods, end-for-end switching, and downtime for rope change-out. Trends of average hoist rope life and average drag rope life determined from the field study are shown in [Fig. 1] as a function of bucket size. The three lines on the graph have slopes of 1, M, and 1/4. A comparison of the data shows that hoist ropes moved about twice as much material as drag ropes prior to retirement, although more than a two-to-one scatter within each category is evident. As expected, the larger machines with the larger buckets moved more total volume of material. The four pairs of solid symbols represent those data for which detailed rope life histories are presented in subsequent figures.

Jan 1, 1980

By B. H. Sage, R. T. Carter, W. N. Lacey

The phase behavior of a number of binary systems consisting of parafin hydro-carbons has been determined in recent years. In addition, the composition of the coexisting phases of mixtures of crude oil and natural gas has been investigated.~ However, the latter study did not include information concerning the influence of the concentration of a component upon its distribution between two coexisting phases. Owing to the paucity of such information it was considered desirable to investigate in some detail the composition of the coexisting phases in the methane-propane-n-pentane system throughout the greater part of the two-phase region at a tempera-ture of 100°F. Information of this nature also permits the establishment of the critical behavior of this ternary system. Method The compositions of the coexisting phases in this ternary system at 100°F. were deter-mined by withdrawing samples from a pres-sure vessel containing a heterogeneous mixture of the three components that had been brought 10 equilibrium by mechanical agitation at the desired pressure and tem-perature. The samples were withdrawn under such conditions that the equilibrium was not disturbed and the composition of each of the phases was ascertained by means of the conventional low-temperature fractionation procedure. The apparatus employed in this investi-gation was nearly the same as that utilized in an earlier study of the methane-pentane system.~ Several minor modifica-tions in the arrangement of the apparatus were made in the interest of simplifying the procedure and avoiding the possibility of condensation in the tubing connecting the equilibrium apparatus with the analyti-cal equipment. The pressures were meas-ured by means of a pressure balance with a precision of 0.5 lb. per sq. in. It is believed that the equilibrium pressure was ascer-tained with an uncertainty of not more than 2 lb. per sq. in., even though small varia-tions in pressure obtained during the withdrawal of the samples. The tempera-ture of the equilibrium equipment was maintained by use of an automatically controlled oil bath. The bath temperature was determined by means of a copper-constantan thermocouple, the electromo-tive force being determined by means of a potentiometric circuit of such a nature as to reduce the uncertainty of the tempera ture to less than O.I f. The fractionating column employed in the determination of the composition of the gas phase had a diameter of 0.12 in. and was approximately 4 ft. long, while the column employed in the analysis of the liquid phase was 0.18 in. in diameter and about 5 ft. long. Each column was equipped with the conventional constant-volume receivers, which were maintained at a constant temperature by means of an automatically controlled oil bath. In nearly every case the distillation in thr neighborhood of the transition between

Jan 1, 1941

By W. N. Lacey, R. T. Carter, B. H. Sage

The phase behavior of a number of binary systems consisting of parafin hydro-carbons has been determined in recent years. In addition, the composition of the coexisting phases of mixtures of crude oil and natural gas has been investigated.~ However, the latter study did not include information concerning the influence of the concentration of a component upon its distribution between two coexisting phases. Owing to the paucity of such information it was considered desirable to investigate in some detail the composition of the coexisting phases in the methane-propane-n-pentane system throughout the greater part of the two-phase region at a tempera-ture of 100°F. Information of this nature also permits the establishment of the critical behavior of this ternary system. Method The compositions of the coexisting phases in this ternary system at 100°F. were deter-mined by withdrawing samples from a pres-sure vessel containing a heterogeneous mixture of the three components that had been brought 10 equilibrium by mechanical agitation at the desired pressure and tem-perature. The samples were withdrawn under such conditions that the equilibrium was not disturbed and the composition of each of the phases was ascertained by means of the conventional low-temperature fractionation procedure. The apparatus employed in this investi-gation was nearly the same as that utilized in an earlier study of the methane-pentane system.~ Several minor modifica-tions in the arrangement of the apparatus were made in the interest of simplifying the procedure and avoiding the possibility of condensation in the tubing connecting the equilibrium apparatus with the analyti-cal equipment. The pressures were meas-ured by means of a pressure balance with a precision of 0.5 lb. per sq. in. It is believed that the equilibrium pressure was ascer-tained with an uncertainty of not more than 2 lb. per sq. in., even though small varia-tions in pressure obtained during the withdrawal of the samples. The tempera-ture of the equilibrium equipment was maintained by use of an automatically controlled oil bath. The bath temperature was determined by means of a copper-constantan thermocouple, the electromo-tive force being determined by means of a potentiometric circuit of such a nature as to reduce the uncertainty of the tempera ture to less than O.I f. The fractionating column employed in the determination of the composition of the gas phase had a diameter of 0.12 in. and was approximately 4 ft. long, while the column employed in the analysis of the liquid phase was 0.18 in. in diameter and about 5 ft. long. Each column was equipped with the conventional constant-volume receivers, which were maintained at a constant temperature by means of an automatically controlled oil bath. In nearly every case the distillation in thr neighborhood of the transition between

Jan 1, 1941

By J. Wright, R. J. Tyler

Changes in the pore structure of hematite pellets during reduction to iron were investigated under a variety of experimental conditions. Pore structures were determined by a mercury penetration technique. The reduction temperature was found to be the major influence on structure, the pore system being bimodal below 7.50°C and unimodal above 900°C. Pore sizes ranging front 0.04 to 20 m were observed in complete metabolized pellets produced over the temperature range 650 - 1050°C. Gas composition and reduction rates also influenced the pore structure of the metabolized pellets, although to a lesser extent than the reduction temperature. An apparent activation energy of 24 kcal/mole was derived for the pore formation process.

Jan 1, 1979

By Roy J. Greenfield, Charles H. Stoyer

A geophysical survey was made in Kylertown, Pa., in an area where the ground water is polluted with acid mine drainage. Since acid mine water is a good electrical conductor, both direct-current electrical and electromagnetic loop methods were examined for their ability to determine the subsurface electrical conductivity structure and from this to imply the location of the acid water. Consistent results were obtained with the two methods and the interpreted subsurface resistivities were consistent with the electrical resistivity measured for water samples. It was demonstrated that fracture traces which act as conduits for ground water flow can be located with electromagnetic surveying methods. Thus geophysical methods for measuring subsurface electtrical resistivity can be used to advantage for monitoring the environmental impact on the ground-water system of acid mine water.

Jan 1, 1977

By John V. Beall

With some difficulty we framed the picture of the village on the mountain getting in the railroad bridge and tunnel at the base. It was a charming scene of red-tile roofs on a green mountain in the golden autumn sunshine with the black hole of the railroad tunnel looking like something out of the movie, "The Train." We clicked the shutter, advanced the film and started back to our rented Audi. At which point there was a hail from the other side of the railroad bridge. A man in uniform was brandishing an object in the air which looked suspiciously like a rifle. We peered around. We were the only other persons in view. He must be yelling at us. He was now closing the distance at a brisk, determined gait but was still 100 yards away. The car was handy but the road was straight, paralleling the railroad tracks for some distance. We dismissed the idea of running for it. Who knows, he might shoot. When he came up we didn't have much of a conversation because it had taken us a week to learn three words of Yugoslavian-zhiveli, long life, huhvalla, thank you, and schlivavitza, the national drink (plum brandy). None of these seemed appropriate. We got the Idea that photographs of railroad tunnels and bridges were prohibited. The guard, for such he seemed to be, took the Audi plate number, administered a last scowling reprimand and stalked off. This incident was the only one in a 15-day visit to Yugoslavia which we would have associated with "Iron Curtain restrictions."

Jan 1, 1970

By S. R. Mitchell, M. Weissman, J. H. Brown

It has been demonstrated that for fine grinding operations, the energy input (E) and the product size modulus (k) are related by an equation of the form E = Ak1-n (1) where "A " and "n" are constants.(1) Also, it was suggested that "A" is determined by the comminuting device and that "n" depends primarily on the rock. It was the purpose of this investigation to study these constants by determining the form of an energy-size reduction relationship over an extreme range of energy inputs, and at the same time, to evaluate the minimum particle size produced during grinding. The comminution system consisted of a stainless steel laboratory ball mill in which quartz was ground at 60% solids for times ranging from 1 minute to 700 hours. The mill which measured 10" diameter by 12" long was loaded with about 60 pounds of balls ranging in diameter from 3/8" to 1-1/8" and was driven at 55 r.p.m. Quartz was selected for the tests because it is a brittle material for which size distribution data were expected to show a constant distribution coefficient regardless of product size.(1) This permitted specification of product size by a size modulus alone. The quantities measured were the time of grind (measure of energy input) and the specific surfaces of the products. Since no accurate method is available yet for determining particle size directly in the very fine size ranges specific surface data were used for this purpose. The relationship between these two quantities may be expressed as where Cis a constant which includes a shape factor, p is the specific gravity of the solid,_S is the specific surface and D is an equivalent particle diameter. For a material in which the distribution coefficient does not change with size the term (l/S) may be substituted for k in equation L if the proportionality constant "A" is changed correspondingly. This substitution was made in this work but in addition, screen analyses were made on several products to verify the equivalence of the two methods of size specification. EXPERIMENTAL DETAIL QUARTZ Quartz from two sources was tested. One series of tests was made on a clear, high purity crystalline quartz from Brazil and Arkansas, and another series was run on massive quartz fromNewHampshire. Grinding charges of the high purity material consisted of 2000 gram samples of -3-8 mesh particles, while the charges of the New Hampshire quartz were 500 gram samples of a -35+48 mesh screen fraction. The difference in the sizes of these two series of charges was not significant in view of the extreme size reduction accomplished, and the New Hampshire samples were smaller to permit ginding to finer sizes in shorter times than required for 2000 gram lots. SAMPLING AND ANALYSIS Separate charges were ground for all tests in which the grinding times were less than one hour and for all tests on New Hampshire quartz since these products were difficult to sample accurately without affecting the entire charge. During the long time grinding tests the mill was

Jan 1, 1961

By W. R. Bendy

GRINDING is a large and costly part of Portland cement manufacture. Prior to clinkering in the rotary kiln, raw materials are ground to a fineness of 80 to 90 pct passing 200 mesh. Then, after burning and cooling, the resulting clinker is ground to about 92 pct passing 325 mesh. In the cement industry the most favored method of grinding has always been by impact and attrition of a ball charge in a rotating mill. Other types of mills, in which materials are ground between die rings and rollers, or between die rings and large balls, are sometimes used for single-stage grinding and often for preliminary grinding followed by grinding in ball-type mills. Their efficiency is usually high, but maintenance and repair costs are high also, and the ball-type mill continues to be the most widely used. Together with 4 pct gypsum to retard setting time, the kiln run clinker—consisting of hard, semi-fused lumps that may have been crushed to — 1/2 in.—is ground to a Blaine surface area of 3000 sq cm per gram and about 92 pct passing 325 mesh. Approximately 32 kw-hr per ton, or 6 kw-hr per 376-1b barrel, are consumed in the grinding mills alone, not counting auxiliaries. Open Circuit Clinker Grinding: Many early clinker grinding plants employed two-stage, open circuit grinding. Comparatively short mills of large diameter, loaded with balls 4 in. and smaller, first reduced the clinker to 95 pct passing 14 mesh. Tube mills of smaller diameter and greater length, loaded with balls 11/4 in. and smaller, carried out the fine grinding operation. An early development was to assemble two or more stages of grinding in a single mill having two, three, or four compartments separated by division heads. There were difficulties in balancing the compartments, but compartment mills were popular in the 1920's because the layout was simple and there were no elevators and conveyors. Oversize Particles: Open circuit mills, whether separate ball and tube mills or compartment mills, always encountered the problem of tramp oversize in the product. A small percentage of clinker survived passage through the first grinding stage as particles—perhaps 1/16 in. diam—that were too large for the smaller balls in the following stages to reduce to —200 mesh. When the feed rate was lowered to reduce tramp oversize, there were serious losses in grinding efficiency. Many ingenious designs were perfected to incorporate perforated screen plates or wire mesh screens in the ball mills or in the first compartments of compartment mills. Their purpose was to prevent mate: rial passing out of the first stage until it was small

Jan 1, 1958

By R. S. Silo

A new process for leaching oxidized copper ore (CuO) is presented. Its principle advantage is that it requires low cost raw materials. For large operations the only materials used are sodium chloride (NaCl) and calcium carbonate (CaCO3). In those operations of medium size, sulfur, which is abundant in the North of Chile, is used. However the process may also be applied to small copper plants provided low cost sulfuric acid is available as a substitute for sulfur. That is, it may be applied to those small capacity plants where large additional installations for the regeneration of the leaching solution are not needed. The method is comprised of six stages, namely: Crushing, Leaching, Washing of the Solution, Precipitation, Regeneration of the Leaching Solution, and Fusion. A flow diagram for this process in a large or medium-sized plant is given. LEACHING This process requires hydrochloric acid as the leaching solution for treating copper ores. This is easily produced from sulfuric acid and sodium chloride as follows: A small excess of NaCl is advisable to avoid the presence of free sulfuric acid. The experiments reported here were performed with 36 grams of HC1 per liter of solution to obtain a free acidity equivalent to 50 gpl of H2SO4 (the usual leaching solution concentration). Hydrochloric acid leaching may be done by either of the usual methods; i.e. percolation or agitation. The method selected depends on the characteristics of the ore. No matter which method is chosen, low pressure air can be used either as an air lift for circulation of the solution, or for agitation of the pulp with "Pachuca" tanks. In either case, there are no problems with respect to oxidation. Leaching by agitation produces good recovery; but at times, the problem of high consumption of leaching solution or the overproduction of grinding slimes which interfere with contact between phases negates this advantage. The percolation method was used in this study. Percolation took place in four 15-liter containers, 11 1/2 in. diam. by 12 in. height, having hemispheric bottoms. Each container was fitted with a wooden bottom having 1/4 in- holes. The air lift was a 3/4 in. plastic tube 10 in. high. Air at 6 psi was supplied from a rotating laboratory compressor. 10 kg of ore were charged to the container and counter-current leaching was used. The leaching solutions were fed through the center pipe in order to use the ascending solutions to reject the air in the charge. The solution circulated through the air lift in order to facilitate the movement according to the increase in density produced by the reaction between the phases present. Specific observations made in this report cannot be generalized because the leaching characteristics are dependent on the nature of the ore used, the method of leaching, and the production of primary or secondary slimes while crushing or leaching.

Jan 1, 1967

By F. N. Pease, C. B. Dudley

Those who have worked a good deal on the chemistry of pig-irons of which the physical properties were known, cannot fail to have been struck by the remarkable want of relation between the physical properties of different samples, and the ultimate chemical composition. It is very well known among foundrymen that irons which will give say 4 inch of chill, and are, therefore, approximately the same in chemical composition, are very far apart, many times, in their physical properties. The foundryman's tests for strength are, of

Jan 1, 1886

By Michael A. Zappia

INTRODUCTION The San Manuel mine of Magmacopper Company is located approximately 72 km northeast of Tucson, Arizona. The mine, concentrator, smelter, electrolytic refinery and continuous rod casting are located on the west side of the San Pedro River valley. Ore Production: 58 1W mtpd Production To Date: Over 340 000 000 mt since 1956. Ore Grade: 0.70% net sulfide cu. As conceived at the present time, the San Manuel ore body is the lower half of what was originally an elliptical (In cross sect ion) cylinder same 2400 m long with major and minor cross sectional axes of 1500 m and 760 m. After emplacement, the cylinder was diagonally sliced along its long dimension by the San Manuel fault, which separated it into halves, moving the upper half same 2500 m down dip and bringing a wedge-shaped blanket of tertiary conglomerate into position over the lower segment of the cylinder. The lower segment is the San Manuel ore body. The upper part, moved down dip, is the Kalamazoo ore body. Originally, these were one ore body. The San Manuel ore body is a structurally weak and highly fractured quartz monzonite, the principal ore minerals being chalcopyrite, with small amounts of molybdenite. The present mining method is full gravity diagonal retreat panel caving by block. This is shown on Figure 2 , (Block Sequence 56, 57, 58, 59 --on the following page.) The ore mined is hoisted to the surface for transportation to the concentrator. BLOCK CAVING MINING SEQUENCE HISTORY The San Manuel mine was planned and developed as a full gravity block caving mine. It continues as a full gravity block caving mine. Since its initial start, the major changes in mining have been in the areas of ground support, rate of draw and undercut sequence. Ground weight resulting from block caving plays a dual role. It can cause extensive damage to ex- traction openings, yet it is necessary to break the ore into easily handled size. The ground- weight problem resolves itself into finding ways to make it manageable. The initial early ground support in the production areas consisted of timber sets and various combinations of steel and timber sets. This sup- port system proved unsatisfactory and a new support system utilizing concrete was developed. Concrete support, 30 cm to 45 cm thick, proved satisfactory. Rates of draw were experimented with. A draw rate of 45 cm to 61 cm (32 to 45 mt) per draw point per day was determined as optimum in con- trolling ground weight. Undercut sequence or mining sequence was recognized as a major contributing factor to ground weight. This is the most difficult factor to change. The lead time from primary development to undercut can be as much as two years. Initial Checkerboard Pattern The initial undercut sequence was in a checker- board pattern (Figure 2). Blocks Sequence 1, 2, 3, 4, 5, 6--next page.

Jan 1, 1981

By R. E. Head

Under existing economic conditions, the treatment of gold ores occupies an outstanding position in metallurgical activity. The increased price of gold has automatically brought about a reclassification of old dumps and low-grade orebodies, and many that formerly were considered valueless are now being vigorously investigated and worked. This situation has developed problems that have taxed the ingenuity and ability of metallurgists to the utmost. Some gold ores give a high extraction of gold by standardized methods of treatment; other ores of apparently similar character yield but a relatively small percentage of their gold when treated by the same processes. Surface Contamination The experience gained by microscopic study of numerous gold ores and tailings has shown that there are pronounced differences in the physical characteristics of the minerals composing them, more especially of the gold. Repeated studies of tailings from flotation and cyanidation of gold ores has established the fact that surface contaminations on gold particles are often directly responsible for high gold losses; comparison of gold particles isolated from flotation concentrates and the resulting tailings have showed that the clean gold has been recovered and the tarnished or contaminated gold invariably lost in the tailing. Obviously, it is not possible to make such a comparison of gold ores treated by cyanidation, as the clean gold has been taken into solution; but when the gold found in cyanide tailing has a tarnished or coated surface, one may infer, with a reasonable degree of certainty, that the clean gold has been extracted. This premise is supported by experimental evidence obtained by isolating particles of tarnished gold and exposing them to cyanide solution in small parting cups. In one such experiment, tarnished gold particles picked from a cyanide tailing showed but slight evidence 01 dissolving at the end of 27 days. In this test, a cyanide solution of 1.6 11). per ton was used and the leach solution was decanted arid renewed every

Jan 1, 1939

By R. E. Head

Under existing economic conditions, the treatment of gold ores occupies an outstanding position in metallurgical activity. The increased price of gold has automatically brought about a reclassification of old dumps and low-grade orebodies, and many that formerly were considered valueless are now being vigorously investigated and worked. This situation has developed problems that have taxed the ingenuity and ability of metallurgists to the utmost. Some gold ores give a high extraction of gold by standardized methods of treatment; other ores of apparently similar character yield but a relatively small percentage of their gold when treated by the same processes. Surface Contamination The experience gained by microscopic study of numerous gold ores and tailings has shown that there are pronounced differences in the physical characteristics of the minerals composing them, more especially of the gold. Repeated studies of tailings from flotation and cyanidation of gold ores has established the fact that surface contaminations on gold particles are often directly responsible for high gold losses; comparison of gold particles isolated from flotation concentrates and the resulting tailings have showed that the clean gold has been recovered and the tarnished or contaminated gold invariably lost in the tailing. Obviously, it is not possible to make such a comparison of gold ores treated by cyanidation, as the clean gold has been taken into solution; but when the gold found in cyanide tailing has a tarnished or coated surface, one may infer, with a reasonable degree of certainty, that the clean gold has been extracted. This premise is supported by experimental evidence obtained by isolating particles of tarnished gold and exposing them to cyanide solution in small parting cups. In one such experiment, tarnished gold particles picked from a cyanide tailing showed but slight evidence 01 dissolving at the end of 27 days. In this test, a cyanide solution of 1.6 11). per ton was used and the leach solution was decanted arid renewed every

Jan 1, 1939

"The Washoe Smelter is situated about two miles east of the City of Anaconda. The smelter site includes about 240 acres and peculiarly adapts itself in topography to the efficient handling of material. By reason of its being on a hill side it is possible to make the level of delivery of the product of one building the level of the charge floor of the next succeeding one.THE ORES: The ore treated comes principally from the Butte district and is hauled by the Butte, Anaconda & Pacific railway, a distance of twenty-eight miles, in trains of fifty 50-ton bottom-dump cars to the railroad yards at East Anaconda. At this point these trains are• split up according to their destination,—that is, as to whether they go to the concentrator and sampling mill ore bins, or to the storage bins. In each case they are made up into trains of eight or sixteen cars, according to the size of engine to be used, and are carefully weighed before delivery to their respective bills.Some idea of the magnitude of the works may be obtained from the following figures:Amount of ore that can be treated in 24 hours…12,000 tonsLime rock from adjacent quarries…2,300 tonsCoke used…500 tonsCoal for reverberatory use…500 tonsCoal for power use…50 tonsWater, per minute, 35,000 gallons.Men employed in and around Anaconda, 3,000. Monthly payroll for above, $300,000.00."

Jan 1, 1913