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By W. F. Holbrook, T. L. Joseph

The problem of sulfur control is important in all blast-furnace operations but particularly for certain grades of steel because of the rigorous specifications. During the past decade the tendency has been to produce pig iron and hot metal of lower sulfur content. As most of the sulfur in the iron originates in the fuel, the use of high-sulfur coke demands modifications in practice to meet sulfur specifications. While it is generally known that high temperatures and more basic slags favor desulfurization, operators as a rule produce the most acid slag that will permit proper desulfurization. The viscosity of blastfurnace slag has been thought to be related to desulfurization and has been investigated by Feild and Roysterl, also McCaffery and his associates*, but apparently no systematic effort has been made to determine the relative desulfurizing powers of slags over the range of composition encountered in practice. Development of Testing Method Because slag and metallic iron are immiscible, the desulfurization reaction is heterogeneous, occurring only at the interfaces between slag and metal bodies. In the blast furnace the newly reduced iron, containing FeS, descends through the slag bath probably either as small drops or streams, at the interfaces of which desulfurization occurs. After its descent through the slag the iron collects to form a bath of metal, at the top surface of which desulfurization reaction probably occurs. These two steps have been thought to account for all desulfurization in the blast furnace. Kinney3 found that metal removed from the blast furnace just above the tuyeres contained three to four times as much sulfur as the iron at cast. Herty and Gaines4 conclude that most of the sulfur is removed while the metal is passing through the slag. Mund, Stoecker

Jan 1, 1936

By W. F. Holbrook, T. L. Joseph

The problem of sulfur control is important in all blast-furnace operations but particularly for certain grades of steel because of the rigorous specifications. During the past decade the tendency has been to produce pig iron and hot metal of lower sulfur content. As most of the sulfur in the iron originates in the fuel, the use of high-sulfur coke demands modifications in practice to meet sulfur specifications. While it is generally known that high temperatures and more basic slags favor desulfurization, operators as a rule produce the most acid slag that will permit proper desulfurization. The viscosity of blastfurnace slag has been thought to be related to desulfurization and has been investigated by Feild and Roysterl, also McCaffery and his associates*, but apparently no systematic effort has been made to determine the relative desulfurizing powers of slags over the range of composition encountered in practice. Development of Testing Method Because slag and metallic iron are immiscible, the desulfurization reaction is heterogeneous, occurring only at the interfaces between slag and metal bodies. In the blast furnace the newly reduced iron, containing FeS, descends through the slag bath probably either as small drops or streams, at the interfaces of which desulfurization occurs. After its descent through the slag the iron collects to form a bath of metal, at the top surface of which desulfurization reaction probably occurs. These two steps have been thought to account for all desulfurization in the blast furnace. Kinney3 found that metal removed from the blast furnace just above the tuyeres contained three to four times as much sulfur as the iron at cast. Herty and Gaines4 conclude that most of the sulfur is removed while the metal is passing through the slag. Mund, Stoecker

Jan 1, 1936

By Richard E. Goodman

Rock slope stability assumes different roles in decision making as a mining venture develops and, accordingly, geological investigations vary in thought and in deed according to the project stage. During exploration, stability questions bear on the decision whether or not to mine, because their resolution affects the cost of mining. During the mine lay-out stage, stability is a key factor in deciding how to mine. Finally, during mine operation, stability affects safety and operating costs, and therefore, pit productivity. Among the many considerations influencing stability, rock mass classification and geological environmental factors count heavily. The rock mass behavioral class depends upon the rock character and its degree of fracturation. The loads on the rock are derived from the environmental factors of initial state of stress and the water regime. Of course, stability is also tied to a host of non-geological factors such as the pit geometry, mining style, relative wealth, philosophy of mine operation, and accident history. METHODS OF STABILITY EVALUATION AND CORRESPONDING INVESTIGATORY TOOLS The purposes of the geological investigation are to identify and evaluate the geological factors influencing stability so that the vital questions can be analyzed. The tools selected for stability analysis reflect the stage of development of the mining scheme and therefore the geological investigations, which must furnish data for the analyses and must vary with the project stage. For exploration and mine planning, the methods of resolving stability questions must meet the need to evaluate numerous parameters in limited time; examples of such methods are stereographic projections (1, 2, 3) charted solutions(4) and simple physical models (5, 6). An investigation program that employs geological mapping tools and aids to mapping, such as geophysics, with relatively few rock tests, is in balance with these analytical methods. First data about the water regime are obtained from observed piezometric levels, and the initial state of stress is queried simply through geological inference and experience. In the layout stage more elaborate physical scale models (7) and perhaps finite element models (8, 9, 10, 11, 12), will be relevant to the design process and therefore much more detailed geological investigation is required. The scale of observation of both rock material and joints will be significantly enlarged and mechanical tests of rock and joint strength

Jan 1, 1972

By George F. Comstock

It has been reported recently1 that the hardenability and toughness of forging steels may be improved appreciably by the use of complex deoxidizers containing titanium, aluminum, and vanadium. In the light of these results, an investigation of similar deoxidizers containing various hardening and grain-refining elements was proposed. These deoxidizers included three manganese-silicon-aluminum-titanium al- loys with and without boron and calcium, three titanium-aluminum ferroalloys with vanadium, molybdenum and zirconium, respectively, an aluminum-zirconium ferroalloy, and plain ferroboron. Their approximate compositions are given in Table I, and the results obtained so far in the investigation of their effects on steel are presented in the later parts of this paper. Method of Using Deoxidizers The deoxidizers listed in Table I were used in three series of 0.40 per cent carbon steels, which were melted in a small basic-lined induction furnace and cast in the form of I7-lb. ingots about 2.75 in. square. The basic charge for each heat was 10 lb. Armco iron and 6 lb. of clean low-carbon sheet-steel scrap. When melted, this charge was deoxidized with 15 grams of ferrosilicon, then about 500 grams of pig iron was added to give the desired carbon content. After the pig iron was dissolved, nickel shot and high-carbon ferrochromium were added if required, and additions of 80 per cent high-carbon ferromanganese and 50 per cent ferrosilicon were made to provide the proper amounts of those elements. Aluminum was added to every heat 2 min. after the silicon, for grain-size control and preliminary deoxidation. One minute later the special deoxidizer was added to the clean bare surface of the molten steel, and 2 min. afterward the steel was poured into the ingot mold, using a hot-top and pipe-preventing compound after pouring to decrease the depth of the shrinkage cavity. The special deoxidizers were used mostly in the form of about ½ to¾-in. lumps,

Jan 1, 1942

By George F. Comstock

It has been reported recently1 that the hardenability and toughness of forging steels may be improved appreciably by the use of complex deoxidizers containing titanium, aluminum, and vanadium. In the light of these results, an investigation of similar deoxidizers containing various hardening and grain-refining elements was proposed. These deoxidizers included three manganese-silicon-aluminum-titanium al- loys with and without boron and calcium, three titanium-aluminum ferroalloys with vanadium, molybdenum and zirconium, respectively, an aluminum-zirconium ferroalloy, and plain ferroboron. Their approximate compositions are given in Table I, and the results obtained so far in the investigation of their effects on steel are presented in the later parts of this paper. Method of Using Deoxidizers The deoxidizers listed in Table I were used in three series of 0.40 per cent carbon steels, which were melted in a small basic-lined induction furnace and cast in the form of I7-lb. ingots about 2.75 in. square. The basic charge for each heat was 10 lb. Armco iron and 6 lb. of clean low-carbon sheet-steel scrap. When melted, this charge was deoxidized with 15 grams of ferrosilicon, then about 500 grams of pig iron was added to give the desired carbon content. After the pig iron was dissolved, nickel shot and high-carbon ferrochromium were added if required, and additions of 80 per cent high-carbon ferromanganese and 50 per cent ferrosilicon were made to provide the proper amounts of those elements. Aluminum was added to every heat 2 min. after the silicon, for grain-size control and preliminary deoxidation. One minute later the special deoxidizer was added to the clean bare surface of the molten steel, and 2 min. afterward the steel was poured into the ingot mold, using a hot-top and pipe-preventing compound after pouring to decrease the depth of the shrinkage cavity. The special deoxidizers were used mostly in the form of about ½ to¾-in. lumps,

Jan 1, 1942

By David Cole

The appearance of the Wilfley table in 1897 marked an epoch in the art of concentration of ores. The table has merited and received an almost unprecedented measure of public approval, lasting through its whole patent life. It has been very little improved in itself, or improved upon by competitors. The new machines bidding for popular favor have been of the Wilfley family, but without ability to do markedly better work than the parent machine, or to do more work except by the use of superimposed decks, the latter having obvious disadvantages. During the past three years there has been developed at Morenci a new type or arrangement of riffles applicable to the Wilfley type of concentrating table which corrects many of the objectionable features or limitations of the older system and obviates most of the difficulties encountered with it, particularly when handling ores having a low ratio of concentration. The new type is known as the Butchart1 riffle system, and its general arrangement is shown in Figs. 1 and 2. The old system accomplished stratification satisfactorily but has

Jan 1, 1915

By E. J. Duggan

The mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. Description of Ore The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chal-copyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. History Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From r926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for inolybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12.500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. TO increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. Crushing There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. a).

Jan 1, 1943

By E. J. Duggan

The mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. Description of Ore The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chal-copyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. History Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From r926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for inolybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12.500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. TO increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. Crushing There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. a).

Jan 1, 1943

By Gail Moulton

THE oilfields in Illinois are located principally along the LaSalle anti-cline in the southeastern part of the state. The oil produced comes most largely from beds of Pennsylvanian and Mississippian age, but there is an increasing important production from deeper horizons, as will be noted under new development. The oil is largely of paraffine base, and has an average specific gravity of about 35° Baume. The rank of Illinois as an oil producer has dropped from third since 1913, when its production was exceeded only by that of California and Oklahoma, to eighth in 1923. This has been the result of failure to find new pools of considerable size to offset the declining production of the old wells and of the greatly increased production of some of the other states. The noteworthy features of petroleum production in Illinois during 1924 are: the low rate of decline for the state as a whole; the smaller percentage of dry holes; and the larger average initial yield of the new wells.

Jan 3, 1925

By Robert E. Ennis

This somewhat misnomered "continuous" face section is, in most mines today, still far from perfect continuity. Achieving uninterrupted flow of coal from the mining machine to the outside has been a long sought for goal. Clearly, the largest problem to be solved in continuous mining today is one of increasing the cutting time of the miner. Full machine utilization can never be achieved with the new high capacity miners as long as their performance is tied to that of the shuttle car. A recent development has proven, however, that truly continuous mining and haulage is possible. The Bell & Zoller Coal Co. working with the Research and Development Department of Hewitt-Robins Inc. developed and perfected a new mining system which offers advantages in terms of increased productivity at reduced mining costs and enhances the safety of personnel at the face. This new system involves the use of two pieces of equipment that form an unbroken link between the mining machine and the mine belt haulage system. These pieces are a Mobile Bridge Conveyor and an Extensible Panel Conveyor.

Jan 4, 1965

By C. W. Arnold, H. L. Stone, D. L. Luffel

The purpose of this research is to determine (I) the efficiency of small banks of enriched gar driven by methane in displacing oil from a porous medium and (2) the effects of variation in bank size and composition of that efficiency. Most of the experiments were conducted in a sand-packed tube 20-ft long and 1/2-in. in diameter. The hydrocarbon system generally used was methane, butane and decane at 2,500 psia and 160°F. The results of these experiments indicate that, in the regions contacted by the gas, a small bank of an oil-miscible gas driven by methane can displace all of the oil in a piston-like manner. If the enriched gas is of such composition as to remain immiscible with the oil, displacement of oil is less efficient than for the miscible case, and the gas bank travels through the sand with a velocity less than that of the driving gas. These data along with theories discussed imply that smaller banks and less total gas are required when the enriched gas and oil are miscible. INTRODUCTION Widespread application of enriched-gas drive to the recovery of oil rests upon a key factor — the use of limited quantities, or "banks", of enriched gas. At the present time, the value of liquefied petroleum gas or other enriching agents discourages their use in a continuous injection technique, or even in a large bank, except in a few isolated reservoirs. If small banks of enriched gas driven by methane were as effective in displacing oil as is continuous injection, the enriched-gas drive process might be applied to a larger number of reservoirs. Previous research on the mechanics of the enriched-gas drive process reported by Stone and Crurnpl and by Kehn, Pyndus and Gaskell has utilized continuous injection of enriched gas. This work has shown that two types of displacements occur. With gases containing sufficient intermediates. the oil is displaced misciblv and complete recovery is obtained from the regions swept. When gases are used which contain insufficient intermediate hydrocarbon for miscible displacement, oil is displaced immiscibly. In the latter type, selective solution of the intermediate hydrocarbons causes a swelling and reduction in viscosity of the oil and leads to an increased recovery over that obtained by dry-gas (methane) drive. The size of the enriched-gas bank necessary for efficient displacement of oil is determined by those factors which cause deterioration of the bank. A differentiation may be made between those factors which operate on a microscopic scale and those which act on a macroscopic scale. On the smaller scale, the enriched gas mixes in the direction of flow by diffusion and convection with the fluids immediately preceding and following it. On the larger scale, the gas may by-pass the oil by flowing through permeable streaks, by overriding the oil because of density difference, or by fingering because of unfavorable viscosity ratios. In such cases, the enriching material tends to mix with the oil both laterally and in the direction of flow. The increase in effective area available for diffusion and dispersion of the enriching components leads to a faster degradation of the bank and a need for a larger bank than is necessary for those cases in which no by-passing occurs. The effects of such macroscopic factors in the deterioration of enriched-gas banks have been reported in a separate paper by Blackwell, Terry and Rayne. The present study was confined to the factors which operate on the smaller scale, in particular to the behavior of banks of enriched gas in sands uniformly swept by the gas. Experiments were designed to answer the following questions. 1. Can small banks of enriched gas driven by methane be used to secure oil recoveries comparable to those obtained by continuous injection of enriched gas? 2. What is the optimum bank size (the minimum bank size necessary to obtain a recovery comparable to that obtained by continuous injection of the same enriched gas)? 3. How many total pore volumes of gas must be injected to obtain the maximum recovery when the optimum bank size is used? 4. What is the effect of varying the number of enriching components in a gas bank? This report describes the experimental investigations and discusses the results in terms of their significance to reservoir behavior.

By Pradeep K. Rohatgi, Clyde M. Adams

An analysis of heal conduction in the freezing of a solution has been performed. The case considered is the semi-infinite mass of solution, bounded by a con-stant-temperature plane with one-dimensional heat flow, Numerical cooling rate and freezing rate distributions hare been calcutated for the particular case of unidirectional freezing of aqueous solutions against. a cold copper chill. The solution at each level parallel to the chill surface is subjected to a spectra of frcezing rates; the freezing rate is inilially lour., passes through a maximum, and then decreases until solidification is camplete. Arange and maximum freeeing rates are inversely proportional to the square of distance from the chill surface. The temperature of the chill surface in contact with the solution remains constant during solidification. Experimentally, it is observed that deudrite sparing increases linearly with distance from the chill surface; in the light of heat -flow analysis this means that the dendrile spacing is imersely proportional to the square root of the maximum freezing rate. SOLIDIFICATION of liquids is accompanied by rejection of latent heat at the solid-liquid interface. The over-all rate of liquid to solid transformation is governed by the rate of heat extraction. Analysis of the heat transfer in the liquid during freezing is important since it determines the freezing rate and resulting structure. A considerable amount of work has been done on metallic systems since solidification is frequently a major step in processing metals. Unidirectional freezing from heat sinks has been treated by several investigators.1-4 Much of the work has been limited to the simpler case of substances such as pure metals which freeze isothermally at a unique transformation temperature. In this case the thickness of the solid zone y, next to a chill surface. is a parabolic function of the time. 8. where a and b are constants. Eq. |1| is a close approximation giving the progress of solidification for the case in which there is a small but finite thermal contact resistance at the interface between the freezing liquid and the chill.' It has been used with some success in connection with large steel ingots4 when the therma1 contact resistance is negligibly small (as is the case when water is brought into contact with the metal chill). The constant, b. becomes zero and Eq. [l] is then exact. even when the thermal properties of the chill and the freezing liquid depend on temperature. Analysis of unidirectional solidification of solutions is more complicated since the latent heat is evolved over a range of temperatures. Next to the all-solid zone in contact with the chill surface there is a solid plus liquid zone in which solidification takes place. Although the equations presented are considered quite general the numerical and experimental results are for freezing of aqueous solutions from a copper chill. I) EXPERIMENTAL PROCEDURE Solutions of potassium chloride, sodium chloride. and lithium chloride have been used. These solutions form a eutectic: a typical phase diagram is shown in Fig. 1. A solution at its liquid us temperature is poured into a tygon tube mounted on a relatively large copper chill. The chill is immersed in a freezing mixture of dry ice and acetone at -70°C. A thermocouple connected to a high-speed recorder was mounted near the top of the chill to measure the fluctuation in the temperature of chill surface in contact with the solution. In this system the progress of solidification can be visually observed through the tygon tube. After the specimens freeze to a desired height they are removed from the copper chill and stripped from the tygon tube. Longitudinal and transverse sections are cut from the specimen for micrographic examination. Specimens in the form of a thin slice are mounted on a transparent glass slide by freezing small drops of water around the periphery. They are then polished on a series of emery papers; the final finishing is done on a soft tissue paper. A Leitz Biological microscope was used for structural observations; photography was

Jan 1, 1968

By Gail Moulton

THE oilfields in Illinois are located principally along the LaSalle anticline in the southeastern part of the state. The oil produced comes most largely from beds of Pennsylvanian and Mississippian age, but there is an increasing important production from deeper horizons, as will be noted under new development. The oil is largely of paraffin base, and has an average specific gravity of about 35° Baume. The rank of Illinois as an oil producer has dropped from third since 1913, when its production was exceeded only by that of California and Oklahoma, to eighth in 1923. This has been the result of failure to find new pools of considerable size to offset the declining production of the old wells and of the greatly increased production of some of the other states. The noteworthy features of petroleum production in Illinois during 1924 are: the low rate of decline for the state as a whole; the smaller percentage of dry holes; and the larger average initial yield of the new wells.

Jan 3, 1925

By Gill Montgomery

The strong upward curve of fluorspar consumption continued through 1968, with domestic producers unable to furnish more than 30% of U.S. requirements. Stocks of all grades were quite short at all points during most of 1968. This resulted in price increases as both domestic and foreign producers sought to cover spiralling labor costs. Further increases are probable for 1969. All Illinois mines were closed by strikes in January and February. The strong trend in metallurgical imports reflects the increased use of fluorspar in steelmaking, as new oxygen process steel plants come into production. The growth of use of all fluorocarbon chemicals continues to require additional acid grade fluorspar, with military uses a major factor. Several new aluminum smelting facilities coming into production are adding to the aluminum fluoride needs, bringing further pressure upon the world supply of acid grade fluorspar. The world-wide consumption pressure on all grades of fluorspar is pushing prices up, and no end to the tight supply is in sight, even assuming that all the new projects cited here are finished on schedule.

Jan 3, 1969

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By W. R. Bull, J. R. Roos

A method is described whereby the true financial value of each constituent of each concentrate can be calculated. Each value is then compared with what this value would be in a perfect separation. The difference in each case represents the loss incurred by each aspect of an inefficient separation. The relative magnitudes of these losses indicate those avenues of plant investigation and research that offer the highest potential reward. The metallurgical efficiency of a concentration operation is difficult to express because there are two criteria of performance, namely, grade and recovery. Neither of these alone gives any indication of the effectiveness of an operation and, if they are combined in some formula, it must be empirical and unreliable because of many external factors. Since milling is an operation performed as one step in the most profitable overall method of producing a metal or other end product, the main objective of mill operation should be to make a maximum contribution to this maximum profit. When concentrates are sold to custom smelters, the income from the operation, and the costs, may be calculated. When concentrates are further processed by another division of the same company, and there is no formal financial relationship between, say, the milling and smelting divisions of the company, the mill superintendent's job is less well defined. But even so, if the mill superintendent works as though he were selling his concentrates, and uses reasonable, if fictitious, conditions of sale, then he can get a good indication of the effectiveness of his operation by calculating his "Economic Recovery". ECONOMIC RECOVERY This is a method used as a basis of operation by several mills and has been described in the literature. Very briefly, the "standard" of operation is taken as that of a mill separating the ore into perfectly pure concentrates with 100% recovery of each mineral in its respective concentrate. For example, a 4%Pb, 4% Zn ore (galena and sphalerite), would provide a lead concentrate weighing 4.62% of the feed and assaying 86.6% Pb and a zinc concentrate weighing 5.97% of the feed and assaying 67.1 % zinc. These perfect concentrates are (hypothetically) shipped to the appropriate smelters, freight and treatment charges and payments calculated, and the net return obtained. This is the optimum financial return for the operation of the mill and is used as the standard. Now the return for the actual, real-life, mill operation is calculated in the same way, and the ratio of the two is termed the Economic Recovery. Many additions and modifications may be made in any particular case. For example, if the zinc mineral is marmatite, the optimum grade of the zinc concentrate may be set lower than 67.1%. If silver is present, it will be recovered 100% in the lead concentrate in the perfect operation. Similarly, cadmium will be recovered 100% with the sphalerite. Mr. Weiss' paper' mentions a case where, because of extremely fine intergrowth between the valuable mineral, chalcocite, and pyrite, the optimum grade of copper concentrate was taken as 45% Cu instead of the theoretical 80%. Each operation will present its own possibilities in this way.

Jan 1, 1970

By H. A. Wheeler

It may surprise some of our members to learn, that, among the industries based on the mineral resources of the United . States, clay now ranks third, being exceeded in value of product only by pig-iron and coal. The present annual value of clay-products exceeds the value of the outputs of gold or silver, and also exceeds the combined values of copper, leacl, zinc, aluminum, mercury, antimony and nickel. These facts are shown in Table I. which gives the statistics for the year 1902 :

Jan 1, 1905

By Roshan B. Bhappu

Selenium is found most frequently as an accessory mineral in the ores of lead, copper, and nickel, and is recoverable as a byproduct in the treatment of many of the ores of these metals. The geographic distribution of the selenium industry closely follows that of electrolytic copper, with the United States and Canada producing approximately 75% of the total output of the non-communist world. This situation is problem number one for selenium-its dependence upon the electrolytic refining of copper- for this byproduct source is not sufficiently flexible to permit a normal supply-demand balance. Furthermore, selenium has become a strategic material in short supply. Thus, new sources are being sought by both government and private interests. Although considerable attention has been given to this effort during the last few years, hardly any research has been published concerning the recovery of small amounts of selenium present in low-grade deposits. In such ores the metal occurs as native selenium or as selenides of iron, copper, silver, lead, and bismuth. In recent years, some native selenium (in both the gray and red form) and selenides, especially iron selenide (ferroselite), reportedly have been found in the sandstones of the Morrison formation near Grants. New Mexico. Similar occurrences have been reported in the Hagan coal basin, in Sandoval County, New Mexico. Samples from New Mexico ores have contained from a trace amount to as high as 0.425% Se. Some of the samples from uranium mines in the Grants area represent mixed ore and mill feed fur uranium reduction plants. In such cases, the selenium usually is reported as an impurity and is lost in the tailings. At present, no attempt is made to recover this waste metal.

Jan 8, 1962

PRESIDENT SAMUEL A TAYLOR, District 3 PITTSBURGH, PA PAST PRESIDENTS WILLIAM KELLY, District 5 IRON MOUNTAIN, MICH J V W. REYNDERS, District 0 NEW YORK, N Y FIRST VICE-PRESIDENT E DEGOLYER, NEW YORK, N Y TREASURER CHARLES F RAND. . . NEW YORK, N Y

Jan 1, 1923

By Paul T. Allsman

Recently the U.S. Bureau of Mines completed organizational changes that, for the first time since the beginning of World War 11,. re-emphasize mining research as an important part of its responsibilities. A progress report follows on the Bureau's improved program for mining research. During World War II the Bureau's mining research lapsed in favor of short-range, more urgent activities. Since the end of the war mining has grad- ually returned to an economy where reduced costs and increased efficiency are essential to a vigorous industry, and demand for improvement through expanded research is again pre-eminent. In recognizing this situation, the Bureau is again emphasizing mining technology and has reinstituted some of its earlier mining research activities. Recent research, however, follows a more scientific approach in solving mining problems because new technology, particularly new instrumentation, has developed rapidly during the past decade.

Jan 7, 1960