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By K. A. Brame, L. G. Austin, R. S. C. Rogers

The breakage parameters of phosphate ore were determined by batch first-ordering grinding studies in a laboratory ball mill according to the Austin, Klimpel, and Luckie procedures. The results were scaled to five larger mills of 0.56, 0.82, 8.36, 4.41, and 5 m in diameter assuming equilibrium ball charges for the full-scale mills. The results were used to predict product size distributions and capacities for batch grinding in the smaller mills and open-circuit grinding in the larger mills. The predicted capacities were in approximate agreement with experimental values except for the 5-m diam mill. The Bond method gave capacity predictions that were 4570 to 150% too high.

Jan 1, 1987

By Reed S. C. Rogers

The breakage parameters of phosphate ore were determined by batch first-order grinding studies in a laboratory ball mill according to the procedures described in the text by Austin, Klimpel and Luckie, using a slurry density of 70 weight percent. The results were scaled to five larger mills of 0.56 m diameter by 0.91 m length (2 ft. by 3 ft.), 0.82 m by 1.52 m (3 ft. by 5 ft.), 3.36 m by 7.90 m (11.5 ft. by 25 ft.), 4.41 m by 9.10 m (14.5 ft. by 30 ft.), and 5 m by 11.1 m (17 ft. by 36 ft.), assuming equilibrium ball charges for the full-scale mills. The results were used to predict product size distribustions and capacities for batch grinding in the 0.56 m and 0.82 m diameter mills and open circuit grinding in the 0.82 m diameter and larger mills using a one-large/two-small reactors in series model fitted to the residence time distribution measured in the 4.41 m diameter mill. The simulated results were compared to plant data for ranges of mill capacities. It was found that the simulated product size distributions were in good agreement with the experimental values below the 70% passing sizes, but did not agree so well in the top sizes. The predicted capacities were in approximate agreement with experimental values for the pilot-scale mills, were about 7 percent too high for the 3.36 m and 4.41 m diameter mills and about 25 percent high for the 5 m diameter mill. Use of the Bond method with the measured Work Index of the phosphate ore gave capacity predictions that were 45 to 150 percent too high.

Jan 1, 1985

By V. K. Gupta

"Effect of mill speed on grinding kinetics has been investigated with a view to using mill speed as a manipulated variable for control and optimization of the grinding circuit performance under variable ore hardness conditions. Dry and wet batch grinding tests were carried out for quartz, limestone and cement clinker in a 29 cm x 18 cm laboratory ball mill. Mill speed values selected were 55 and 70% critical. In the case of dry grinding operation, with an exception of quartz, for all materials studied the effect of mill speed on grinding rate was found to be insignificant for sizes finer than about 35 mesh. In the case of wet operation, grinding rates were found to be approximately proportional to the mill speed. The implications of these results are discussed.INTRODUCTIONIt is well known that mill speed has a considerable effect on the grinding rate of the ore particles l-3 For particles of a given size there is an optimum speed at which mill grinding efficiency is maximum (i.e. specific energy consumption is minimum) 3•4. Mill speed can, therefore, be effectively used as a manipulated variable to deal with the disturbances arising from changes in the size and hardness (grinding rate) of the ore being fed to the grinding circuit. Though, it has not been the case so far, with recent developments of ring motors2 and frequency inverters5 a continuous variation of mill speed has now become feasible, and it is expected that mill speed control will soon be a regular feature of plant operation2,6."

Jan 1, 1985

By The Mill Staff

Introduction The Hollinger cyanidation mill commenced operating in June, 1912, and 49,779,978 tons of solids had been discarded by the end of 1950. The present tailing disposal plant has been in operation since 1924 and holds 39,762,250 tons. The tailing disposal plant is on the property of Kayorum Gold Mines, Limited, about a mile and a quarter south of the mill. The area available for tailing covers 640 acres and is roughly l ¼ miles long by ¾ mile wide. Tailing was first deposited on the McKay veteran claims, l mile long by 1/2 mile wide, which were later included in the Kayorum property. This site sloped from northeast to southwest, the slope amounting to 106 feet between the northeast corner and the low land to the southwest. The low land was wet and a considerable area was covered with muskeg up to 17 feet deep. It was drained by a brook running across the property in a southwesterly direction. Later, when more ground was needed, the low-lying mining claims adjacent to the south and west boundaries were acquired to allow for the extension of the plant. This entailed some extra difficulties and expense that would not have been encountered had the whole of the present site been acquired with the original purchase. At the present time, the top of the tailing deposit stands at two elevations, but eventually the tailing deposited on the extension to the original plant will be raised to the same height as the older deposit. Previous to 1924, all of the tailing was deposited on Low land and in lakes near the mil1. Between 1924 and 1926 an increasing tonnage was sent to the McKay claims until, in 1926, all of it was being sent there. In 1936, the area of the deposit was extended and shortly afterwards all of the tailing was deposited on the new site.

Jan 1, 1951

The Millennium deposit, located in northern Saskatchewan between the McArthur River mine and the Key Lake mill, has considerable potential for becoming an operating mine. Located in the Athabasca Basin, Millennium is blessed with both the riches and the challenges that this basin offers. The potential for ground water inflows during shaft sinking, and the unconventional challenges associated with underground uranium mining, render the successful development of a new mine in this area especially rewarding. The preliminary design and engineering of the Millennium Project has tapped into the experience of other operations in the area, and has utilized new applications of geophysics technology to enhance the understanding of the local geological structure. This discussion will focus on the innovations, and the challenges specific to uranium mining, for a deposit that is poised to be one of the next mines to supplement Saskatchewan's uranium resources.

May 1, 2008

By F. D. Ford, Z. Dai

"Vale’s Ontario Operations re-introduced Cu/Ni separation at Clarabelle Mill in 2006. The mill processes ores from a number of mines in the Sudbury basin. A saleable Cu concentrate is produced at 30-60% Cu recovery from mill feed, which has to meet the quality specification. In particular, Ni content in Cu concentrate must be less than 0.6%, otherwise, penalties will apply and diminish the value of Cu concentrate, in addition to the lost Ni value. Generally, a Ni grade of less than 0.4% is obtained in Cu concentrate. More recently, while processing Cu-rich ore, Cu concentrate contained more than 0.6% Ni which is abnormally high. Detailed mineralogical investigation shows that Cu-rich ore contains more millerite in comparison to typical Sudbury Ni sulphide ore. Millerite is found to be more floatable than pentlandite and is much more difficult to depress using high-lime aeration conditioning, which is the standard practice in Cu/Ni separation to float chalcopyrite while depressing pentlandite. The current operation uses sodium cyanide at 20 g/t to depress pentlandite and pyrrhotite. However, it is found that a high dosage is required to depress millerite. It is also shown that starch and dextrin are promising in millerite depression.INTRODUCTIONPrior to bulk smelting in 1991, Cu/Ni separation was carried out at Vale’s Copper Cliff Mill, Ontario Operation. The typical Cu concentrate grade was around 30% Cu and 1% Ni. This is very high in comparison to today’s value of 0.4% Ni and would not be acceptable for external processing. However, the Cu concentrate was processed internally at the Copper Cliff smelter/refinery, so the Ni contained in Cu concentrate was not lost and no actual penalties applied. During the SO2 abatement initiative, bulk smelting was chosen because of the perceived simplicity and consequently, Cu/Ni separation was eliminated (Agar, 1991) and the Copper Cliff mill was subsequently demolished. In 2005 and after more than one decade of bulk smelting, business changes required the removal of some Cu concentrate from Clarabelle Mill bulk concentrate to provide smelter capacity for Voisey’s Bay concentrate in the Copper Cliff smelter. In essence, this strategy calls for Cu removal at Clarabelle Mill (Lawson, 2006). In this case, Cu concentrate is sold externally and it must meet the quality requirement of less than 0.6% Ni."

Jan 1, 2011

By N Robb

This paper gives an account of the laboratory testing, plant testing and final inclusion of the two-stage sulphuric acid leaching process in the Milliken mill. Operational results for September, 1962, are presented, showing a uranium extraction of 95.5 per cent, with a saving of some 16 lbs. of 100 per cent H~S04 and some 9 lbs. of lime per ton milled. Methods for effecting further improvements in the process are indicated. HISTORY P RIOR to the incorporation of two-stage leaching, the process for the treatment of uranium ores at Milliken consisted of a single-stage sulphuric acid leach, followed by a solution recovery step involving washing in thickeners and two-stage filtration. The thickener overflow solution was subsequently concentrated and purified by ion exchange, and the uranium was recovered, by two-stage precipitation, as ammonium diuranate. Figure I illustrates this flowsheet. For satisfactory extraction of uranium from the. Milliken ore, a high residual acidity, of the order of 40 grams per litre of free acid in the discharge solution, must be

Jan 1, 1963

By C S. Stevens

This paper describes the problems encountered and the steps taken to solve them in milling an ore containing relatively large amounts of pyrrhotite. A complex and greatly lengthened flotation circuit was required to recover the pyrrhotite. A very low pH had to be maintained in the cyanide circuit in order to get good recovery of the gold and maintain reasonable cyanide recovery. Introduction D ELNITE Mines Ltd., a gold mine situated about 5 miles by road south and east of the town of Timmins, Ontario, is serviced by two shafts. No. 2 shaft extends from surface to the 2875-ft. level, and No. 3 shaft, an internal winze, extends from the 2875-to the 5275-ft. level. No. 2 shaft was stopped at the 2875-ft. level because of a heavy band of chloritic schist which necessitated a drive to the north for 1000 feet, where No. 3 shaft was sunk. The latter bottomed in the schist at the 5275-ft. level. From the 2500 to the 2875-ft. level in No. 2 shaft, the ore was

Jan 1, 1965

By Charles E. Locke

SHARP CONTRAST exists in the reports so helpfully contributed by the individual members of the Milling Committee for this review. Those engaged in the milling of gold and silver ores report great activity, of which some details will be given later. Those engaged in milling ores of base metals report continued depression and little progress. Copper production in the United States is running at about fifteen per cent of capacity. Some Arizona copper mills are closed. Anaconda is milling about 45.000 tons of copper ore per month, and the Utah Copper Co. 35,000 monthly. However, in Canada the International and Falconhridge copper-nickel mills and the Noranda copper-gold mill have increased production. Also the Climax molybdenum mill in Colorado is milling 4000 tons per day. Water shortage from the drought has caused curtailment or complete shutdown of mills in some parts of the West.

Jan 1, 1935

The Hadsel Mill. By R. G HALL (Trans, vol 112 4000 words) This mill is designed to effect the entire reduction of ores by combined crushing and grinding from mine run size to material sufficiently fine for the usual operation of flotation or cyaniding The work is begun and completed in the same machine. The machine is designed on the principal of the elevator wheel, and crushing and grinding are accomplished by dropping a mixture of ore and water upon specially placed "breaker plates" in such a manner that the resultant material passes back into the buckets of the wheel for further elevation One method of removal of the finished product is to permit an overflow of wet pulp on one side of the mill through coarse screens A recent improvement provides for the removal of a pulp by a launder part way up the path of the buckets In either case the pulp passes to a classifier, and the oversize back to the mill in the manner familiar in ball-mill practice Although hardly through the development stage, the machine has given a very good account of itself in simplification of plant, by elimination of coarse and fine crushers with consequent reduction of installation cost and power consumption As compared with the ball mill alone, there is a reduction of power consumption as well as a material reduction of grinding steel There is an increase in cost of maintenance other than grinding steel, but as the machine passes out of the development stage there is a show¬ing of reduction of costs on this item It seems probable that with further develop¬ment and strengthening of the weaker parts further improvement may be made Ball Milling. By A. M Gow, M GUGGENHEIM, A. B. CAMPBELL AND W. H COGHILL. (Tech Pub 517; also Trans, vol 112. 24,000 words) This paper outlines a new philosophy in ball milling by using power as an index of mill operation and ascertaining the effects of the set and induced variables By recognizing the amount of ore in the mill as an important induced variable and recognizing that the amount of ore in the mill depends principally on the size of the discharge trunnion, ball volume, rate of feed, and speed of mill, obstacles formerly met were surmounted When the amount of ore in the mill is dealt with intelligently and the applied work is proportional to the amount of feed, series testing, which would otherwise give a hotchpotch, gives concordant results The study of ball paths and trajectories was seen to be fruitless when it was shown that the old attrition theory for slow mill speed

Jan 1, 1935

Economic Significance of High-grade Concentrates BY PAUL M TYLER AND CARLE R HAYWARD (Min & Met, Dec, 563 4000 words) Generally pyro¬mctallurgical methods yield a higher recovery but cost more per ton of ore treated Milling methods are inherently cheaper for the reason that they do not involve dissipa¬tion of heat energy in slag formation, ctc It is vaguely realized that the smelter man could do a better job if he were supplied with purer raw materials but hard¬headed mill men have been offered little inducement to do a really good job of cleaning concentrates To do this may involve some sacrifice in percentage yield and in mill capacity and unless there is a compensating saving in freight and smelter charges, it simply is not good business to go beyond a certain point in the way of making clean concentrate Opportunities exist for eliminating one or more steps in the expensive series of processes for recovering various metals, or at least reducing the volume of materials that has to be handled at the most costly steps The beneficiation of nonmetallic minerals likewise affords numerous examples of over-all economies obtainable through more refined methods of mechanical improvement All this leads to the conclusion that perfecting selective processes for isolating pure minerals (e g, galena) virtually free from other sulfides as well as gangue may be productive of industrial economies Conversely, cheapening smelting methods may invade the field of mechanical improvement of ores but, as Kiphng said, that is another story

Jan 1, 1937

By A. M. Gaudin

The study that is presented here was undertaken in order to condense information concerning comminution, and covering a great variety of conditions, into one or several rules which would be of use in the development of a systematic theory as well as in many directly practical ways. As the result of this work, the following conclusions appear justified: 1. Rocks may be classified, from the point of view of comminution, into two general classes which are (a) structurally homogeneous, and (b) structurally heterogenous. Structurally homogeneous rocks are those in which fracture takes place through grains and grain boundaries alike; heterogeneous rocks are those in which fracture passes in preference through grain boundaries. 2. If a sized product composed of structurally homogeneous rock grains is crushed, a product results in which the relation between the percentage weight of grains of various sizes and these sizes themselves follows a definite law. This is not true of structurally heterogeneous rocks. 3. Ball milling with structurally homogeneous rocks, and under certain conditions, presents exceptions to this rule. 4. In ball milling with structurally homogeneous rocks, there exists a critical ratio of size of feed to size of balls, above which the machine operates essentially by surface wear, and below which it operates essentially by breaking across the body of the particle. 5. In ball milling, with structurally homogeneous rocks, and when the size of feed grains exceeds the critical size, the particles worn off are of a substantially definite size which depends upon the ball size. 6. In the case of a typical heterogeneous rock, such as Manhattan schist, particles are crushed in such a way that there is a greater proportion of grains of the size that is the average grain-size of the rock than there would be in that range of size with a structurally homogeneous rock. This has been found to be true with all the crushers tested. 7. Different methods of crushing produce grains of different shapes. 8. The importance of sizing the — 200-mesh portion of a crushed product, in crushing-efficiency investigations, has been greatly under-

Jan 1, 1926

By Wm. C. Madge, H. P. Allen

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 toss 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.

Jan 1, 1926

By Bennett R. Bates

Nowhere in the world has cone classification in closed-circuit grinding with tube mills reached as high a state of development as on the Wit-watersrand. In the development of the Far East section the necessity of fine grinding to obtain satisfactory residues gave rise to a situation with which the cone could cope only at a great sacrifice in efficiency of tube-mill grinding. H. A. White1 recognized the limitation of the cone in his study of the theory of tube milling as early as 1915, but it was not until Ewing, Davis and Willey2 carried out their tests at Springs Mines during 1920 that the advantages of mechanical classifiers were demonstrated. During the fall of 1924 the writer supervised a series of tests with Dorr classifiers undertaken by Mr. White, Consulting Metallurgist, at the mill of the Geduld Proprietary Mines, Ltd., Dersley, Transvaal (Far East Rand). The importance of classification to obtain efficient tube-mill grinding was thoroughly demonstrated. The installation of classifiers of the above type that followed the test confirmed the results of the experimental work. Tests with Cones, Straight Type and Bowl Classifiers The test at the Geduld mill was made to determine the relative grinding efficiency of a tube mill when grinding in closed circuit with cones, straight type Dorr classifier and Dorr bowl classifier. The efficiency of a tube mill, as expressed in this article, is based on the tons of minus 200-mesh material produced by the mill. The efficiency is also expressed in terms of tons of minus 90-mesh material produced by the mills to correspond to the standard practice on the Witwatersrand. The results given for the performance of the tube mills grinding in closed circuit with cones (Fig. 1) were taken from previous tests carried

Jan 1, 1926

By Will H. Coghill

In this investigation, the low-grade red iron ores of Alabama were examined by the use of "heavy solutions."' Ores are generally ground preparatory to concentration and as long as the concentrators make suitable products the effectiveness of the grinding is likely to be unquestioned; but when suitable products are difficult to make, an investigation becomes imperative. In such a case a dissection of the ore might be made to find the cause of the trouble. Such a condition of unsuitable products exists in the low-grade red iron ores of Alabama. They are difficult to concentrate. To be sure it might be said that the concentration of iron ores is a new thing and that the ore-dressing methods are undeveloped, but a glance at the industry reveals an abundance of methods and devices for almost all imlaginable conditions. Concentrating machinery has not been tested in this investigation. When the properties of the ores have been gleaned, the machinery may be used more effectively. Alabama Low-grade Red Iron Ores The Bureau of Mines has undertaken by laboratory methods to select the best means of concentrating the low-grade iron ores of Alabama. These ores constitute the parts of the ore bed left by selective mining. If more of the ore bed could be included in the current output, the mining cost would be reduced and longer life of the district would be insured. This advantage would be reflected by added permanence of churches, homes, schools and every undertaking in the commonwealth. The red ores chosen for this study of effectiveness of grinding are too low in lime to meet the requirements of the blast furnace. If lime is

Jan 1, 1927

By R. E. Head, E. L. Tucker

In the metallurgy of precious metals, it has been standard practice for years to use cyanogen compounds, so it was but natural that early investigators in the field of flotation should consider these compounds as possible flotation reagents. In Australian Patent 9508, of 1913, alkali cyanides were described as suitable reagents for the separation of heavy sulfide minerals, the reagents being employed either in an alkaline or neutral circuit. In U. S. Patent 1236501, of 1917, to John M. Tibbet, the use of cyanide in conjunction with soda ash or lime is described. One of the claims of this patent reads, "An improvement in the cyanide process for the treatment of ores comprising, mixing the ore with cyanide solution and oil and alkali: emulsifying the oil in the presence of the ore, the alkali and the cyanide solution, floating a portion of the values in the pulp by producing gas bubbles in the mixture and removing the values floated." More recently, Sheridan and Griswold, in U. S. Patents 1421585 and 1427235, describe a somewhat similar process and have successfully introduced it in the mill of the Timber Butte Milling Co. U. S. Patent 1429544, of 1922, to Stevens, describes the use of double cyanides of copper, zinc, and lead. The success of cyanide as a reagent in the differential flotation of lead and zinc sulfides in highly complex ores is proof that these patents have a solid foundation. Although the technique of their application to particular ores has been worked out in many cases, no one, so far as is known, has attempted to describe the nature of the action of the cyanide or to determine the ideal conditions for its application. It was the purpose of the present investigation to gather information on these points. Naturally this would have been an exceedingly ambitious program had the attempt been made to study the effect of cyanide

Jan 1, 1926

By R. E. Head, E. L. Tucker

Previous investigations of E. L. Tucker and R. E. Head' related in particular to the effect of cyanogen compounds on galena, sphalerite, and pyrite, and their behavior in the presence of such compounds. The present contribution deals with the floatability of the copper minerals and of pyrite. All experiments were performed on minus 200-mesh material, the method of preparation being the same as that described in the previous article. Flotation tests were made in a small mechanically agitated machine of 50-gm. capacity designed in this laboratory by Gates and Jacobsen.2 General Naval Stores pine oil No. 5 and potassium xanthate were used in constant amount in all of the experiments; these agents were selected because of their present wide use in the flotation of copper minerals. Time additions were varied to secure four final hydrogen ion concentrations, namely, neutral, 0.0005 per cent. CaO, 0.003 per cent. CaO, and 0.02 per cent. CaO. The variation in lime addition was introduced to include the range of alkalinities employed in actual flotation practice. The minerals employed were the purest obtainable. Their analyses are shown in the following tabulation: Table 1.—Analyses of Minerals Employed in Experiments mineral Cu Fe Cao S MgO Insol. Chalcocite .............. 77 02 0.49 0.29 21.70 0.08 0.28 Chalcopyrite............. 34.50 29.20 0.23 34.10 0.07 1.42 Bornite................. 64.30 9.30 0.12 23.70 0.10 1.95 Pyrite....................... 36.90 1.12 49.70 .... 1.51

Jan 1, 1926

By Isaac H. Odell, Arthur W. Hixson, Lincoln T. Work

Although a few engineers have recognized the problem of the mechanism of filtration it has never been studied in a quantitative way. A background for a better understanding will be afforded by a survey of certain fundamental principles of filtration. In recent years, authorities have become practically unanimous in saying that the actual filtering should be done, not by the cloth itself, but by a layer of solid particles deposited upon it. Thus, a cloth may be used in which the openings are considerably larger than the particles to be filtered, provided that groups of particles form an arch or bridge over each opening. This can be accomplished if the initial pressure is so low that the arch will not be broken while it is still thin. According to Wright,' the chief advantage of this method of filtration is that the open-weave cloth does not become clogged as easily as does one of closer weave, where the fine particles often work their way into the interior of the cloth and become caught there. If the character of the material to be filtered is such that it will not form a good film, then it becomes advantageous to use "filter aids" for pre-coating the cloth with a desirable layer of solids, and the coating need be only of a very slight thickness for good results. Until the film is formed, the first runnings of the filtrate may be cloudy, often necessitating refiltering this portion. A factor which has considerable effect on the performance of filters is the concentration of solids in the pulp. According to Wright, a high concentration of solids gives a porous cake, since the flow is then slower than with a thin pulp. The porosity of the cake is also affected by the pressure: a high initial pressure tends to compact the cake, and the same effect results from fluctuations of pressure. In his studies of the relation between pore size and particle size, Hatschek2 observed that the particles in a pulp tend to arrange themselves with maximum cross-section parallel to the direction of flow and that this is more marked at the higher vclocfiies. Thus, at high pressure and high velocity, a certain pulp

Jan 1, 1926

By B. E. Mix, L. M. Barker

THe present methods of unloading ore and coarse-crushing at the Magna plant of the Utah Copper Co. are the developments of the pas five years. Hand dumping and breaking have given way to the rotary car dump and large gyratory; and the bucket elevators and stationary screens have been replaced by conveyors and shaking screens. The coarse-crushing plant may be broadly divided into two units the primary crushing and the secondary crushing, or primary roll, unit The former was erected in 1923; with the exception of the excavation the construction was carried on by forces of the company. Excavation was started in the middle of March, the first concrete was poured July 16, erection of the steel building started Oct. 30, and the unit was com pleted and operating Dec. 30. The unit has a capacity of over 24,001 tons of ore per day, and cost more than a million dollars. Remodeling of the primary-roll unit extended over a period of years preceding this some of the work having been done as early as 1918. Delivery of Ore from Mine, Yard Storage, and Handling Ori Cars to and from Primary Crushing Plant The ore is transported, in steel cars of 80 tons capacity, from the min' to the mill, where a reserve tonnage of loaded cars is maintained. Malle locomotives, handling trains of 45 cars, or approximately 3600 tons o ore per train, deliver the trains to a storage yard which contains thre tracks—a passing track and two storage tracks. All of this trackage i equipped with an overhead trolley system, so that electric locomotive can be used in yard switching. The storage tracks have a total length of 4500 ft. and provide storag space for 140 cars. The passing track, besides permitting switchin room for locomotives, is used for storage purposes between train arrivals thus giving intermittently a total storage capacity for the yard of approxi mately 200 cars. The amount of ore stored at any one time, howeve

Jan 1, 1926

By C. A. Lemke

The ores now milled at the Midvale concentrator of the United States Smelting, Refining & Mining Co. come mostly from the company's mine in the Bingham district, about 18 miles west of Midvale. Custom ores from this same district and from other western mining districts are also milled at times. The Bingham ores consist of galena, pyrite, chalcopyrite, and sphalerite in a gangue of quartz, quartzite, limestone, and porphyry. The percentage of chalcopyrite is small, usually about 0.5 per cent., while the galena, pyrite, and sphalerite combined constitute about 60 per cent, of the total ore. These three minerals vary considerably in amount and, at times, all three are somewhat oxidized. The ores also contain gold and silver, the former being mostly in the pyrite while the latter is generally with the galena. The crystallization of the economic minerals varies from fairly coarse to extremely fine. The ores are carried to Midvale by the Bingham & Garfield and the Los Angeles & Salt Lake railroads or by the Denver & Rio Grande Western railroad, in 50-ton, steel, full dump cars. The milling problem is to recover the silver-bearing galena, the gold-bearing pyrite, and the chalcopyrite in a product suitable for the lead smelter and containing the least possible amount of gangue and sphalerite and then recover the sphalerite in a product suitable for a zinc smelter. The minerals are recovered partly by water concentration and partly by water concentration combined with electrostatic treatment. Gravity Concentrator The gravity concentrator has a capacity of about 465 tons in 24 hr. The mill buildings are of wood. The mill was built in 1905 to handle approximately 90 tons in 24 hr.; since that time several additions have been built and various changes have been made in the flow sheet. The last extension, which raised the capacity of the mill to its present level, was built in 1915. The additions to the mill building have not kept pace with

Jan 1, 1926