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By Henry Schwellenbach

LABOR shortages, rising wages, and changes in moisture and fines content of the ore necessitated a review of the flowsheet at Warren Foundry & Pipe Co's. Mt. Hope, N. J. Mines Div. This plant had been designed prior to World War II, and started operations in January 1945. When this operation was designed there was an oversupply of labor in the area, wages were at a reasonable level, the run-of-mine ore was low in moisture, and there was a limited amount of fines. These factors all favored use of hand picking for cleaning rougher lump concentrate. Since that time, many changes have occurred. The proximity of defense industries and large military installations, has created one of the most acute labor shortages in the country. This labor shortage together with present wage levels, and the increase in moisture and fines in the ore created several problems. The increase in moisture and fines may be attributed to changed mining methods, the extensive use of sprays, and the opening of new mining areas bringing an increase in ground water. [ ] Two problems were presented by the increase in moisture and fines. The first was that lump ore arrived at the open hearth in an unacceptable condition, due to the segregation enroute of the large amount of fines that had adhered to the lumps. The second problem was that high magnetite content fines coated the large lumps, making it difficult for pickers to distinguish middling by other than the relative weight. [ ]

Jan 1, 1952

By John D. Morgan, Sylvain J. Pirson

An instantaneous and continuous analysis of a mineral suspension should be of great value in controlling various mineral preparation processes. Described herein is a method of analysis based on the use of high-frequency alternating current, which, under certain restrictive conditions, could be of use. Theoretical Considerations In a suspension of conductive minerals, the electrical effect primarily responsible for the variations in electric conductivity is the capacitance. In Fig Ia, let an alternating voltage E be applied across two electrodes F and F', between which are two rows (S and T) of mineral particles, all of approximately the same size, immersed in a fluid. Let the number of rows of particles between the electrodes be R and the number of particles in each row be N. If A is the area of a particle, D is the distance separating two particles, and k is a constant depending on the fluid, the capacity, CO, of the particle system will be given by: KRA C°= ND Referring now to Fig Ib, let R, the number of rows, be held constant, but let N, the number of particles in each row (S and T), be increased X-fold. The capacity, C1, of

Jan 1, 1949

Nominating Committee.-On the recommendation of the President, the Board of Directors, at their meeting on Apr. 25,1913, appointed the following Nominating, Committee to nominate officers and directors for next year: John Hays Hammond (Ex-President), Chairman. Robert H. Richards Chairman of Boston Local Section. Richard S. McCaffery, . . . . Chairman of Spokane Local Section. Chester F. Lee Chairman of Puget Sound Local Section. Frederick G. Cottrell, . . . . San Francisco, Cal. Charles W. Goodale Butte, Mont. James W. Maleolmson, . . . . Kansas City, Mo.

Jan 5, 1913

By Rudolph G. Wuerker

THE Krupp-Renn Process,[1] has been successfully used to treat low-grade iron ores, laterites, titaniferous sands, and other minerals, and before World War I1 25 units were built by the Krupp-Grusonwerk, Magdeburg, Germany. J. J. Howard[2] has shown theoretical advantages of the pyro-metallurgical treatment over conventional approaches; and the process was described by R. Durrer[3].

Jan 1, 1951

By A. R. Denison

The area discussed in this report comprises a group of 38 counties lying in the northeast corner of Texas. It covers all, and extends beyond, the borders of what is commonly described as the East Texas Basin. It includes two major oil fields, Van and East Texas; two major gas fields, Bethany and Waskom; and a number of fields, both oil and gas, of lesser importance. All commercial oil production comes from beds of Upper Cretaceous age and almost exclusively from the basal member, the Woodbine sand. Gas is produced from both Upper and Lower Cretaceous beds with the newer production from the older beds. East Texas Field This field continued as in 1932 and 1933 to dominate the drilling and production situation throughout the oil business. A total of 3681 producing wells was completed in this field during 1934. This is an increase of 1215 over the 2466 producers completed during 1933. Total wells

Jan 1, 1935

By J. T. Lapsley, I. I. Cornet, A. E. Flanigan, R. Hultgren, J. E. Dorn

With the greatly expanding use of magnesium during the war, it appeared necessary to the War Metallurgy Committee that procedures of heat treating common magnesium casting alloys be investigated systematically. Accordingly, the National Defense Research Committee of the Office of Scientific Research and Development placed a research contract with the University of California, supervised by the War Metallurgy Committee. Research on the solution heat treatment and aging of magnesium alloy AZg2, (formerly ASTM No. 17; also known as Dowmetal C or AM26o) which was done under "Restricted" Project NRC-2I, in 1942-1943, and reported in detail in a final report to the OSRD on September 3, 1943, is the basis of this paper, which has been released for publication by the OSRD. American experience with magnesium alloys prior to World War II was limited, although some of these alloys had been in commercial use since 1909. Requirements of the aircraft industry for light alloys made it imperative to understand and utilize the advantageous strength-weight properties of magnesium castings. Magnesium casting alloys are commonly solution heat- treated to increase their ultimate tensile strength and ductility; this treatment may be followed by an age hardening at elevated temperatures, which raises the tensile yield strength but diminishes the ductility. In the present investigation, AZp was studied to determine the effects on properties of various solution heat treatment and aging schedules. The principles of heat treatment of magnesium alloys are the same as for other non-ferrous alloys. The alloy studied, AZp, has a nominal composition of 9 pct aluminum, 2 pct zinc, and minor constituents. Fig I shows, by the projection of isothermal sections on the concentration plane, the solubility surface1 of the ternary magnesium-rich solid solutions as determined by X ray 'analysis. From this figure it may be seen that above approximately 375°C (707OF), AZp alloy is a single phase alloy at equilibrium; below this temperature the equilibrium condition is a two phase alloy. The beta constituent which precipitates from the solid solution exerts a hardening effect. Because the zinc concentration is only 2 pct, ancl because of its solubility relations, AZqz resembles a binary alloy of magnesium with aluminum. The solubility curve1 of Fig 2 shows how rapidly the solubility of the beta phase decreases with decreasing temperatures. Unlike many aluminum alloys, magnesium alloys do not age harden at room temperatures; precipitation hardening of AZ² alloy is performed at about 177°C (350°F). At this temperature about 3 pct aluminum

Jan 1, 1949

By Harrison Everett Ashley

Slimes are usually defined as all material passing a certain sized sieve, which is invariably the finest sieve employed by each metallurgist in his tests; 100-mesh and 200-mesh have been taken as the limits by different writers in recent volumes of the Transactions; 150-mesh is common practice. The reader of the first two volumes of Richards' is impressed by the inflexibility with which all grinding-arrangements are praised or condemned according as they make little or much slime. Until recently, one of the principal aims of the mill-man was to keep down the slimes. He was obliged to have separate apparatus for sand- and slime-treatment. The latter was sometimes puzzling and erratic. Often it was found most economical to omit treatment of the slimes entirely, or to store them in immense basins or heaps, waiting for possible future improvements in methods of treatment. Recently, improved processes of agitation and filtration have made possible the successful extraction of values from many slimes. The tendency, when the metal is exceedingly finely disseminated, is to cut out all sand-treatment and to slime everything. These processes have been almost entirely mechanical in their conception, and it is the object of the present paper to show that certain chemical considerations may greatly hinder or assist their successful operation. It has long been a source of dissatisfaction to chemists that their analyses fail to be of much help to the clay-worker; that it is a matter of common occurrence for clays of closely identical ultimate chemjcal constitution to show the widest possible divergence in physical properties.

Jan 1, 1911

By W. E. Stiles

A method is presented for predicting the performance of water flooding operations in depleted, or nearly depleted, petroleum reservoirs. The method makes use of permeability variations and the vertical distribution of productive capacity. From these two parameters can be calculated the produced water cut versus the oil recovery. Derivations of the mathematical analogy is shown and sample calculations and curves of prediction are presented. Comparison is made of the predicted and actual performance of a typical 5-spot in an Illinois water flood. INTRODUCTION The use of water as a flooding medium in both depleted and "flush" oil reservoirs is gaining greater recognition and acceptance. Many of the shallower fields, depleted by primary production, have been and are being subjected to water injection in order to obtain some part of the large volume of oil remaining after primary production. Some of the earlier water flood installation proved highly discouraging and the value of water flooding was often questioned. Many of these earlier floods were haphazardly selected and developed as little was known of the physical characteristics and contents of the producing formations. The prior evaluation of the flood performance was impossible. During the past decade the development of the required reservoir engineering tools-—core analysis, reservoir fluid analysis, electric logs, fluid flow formulae, etc.—has allowed the engineer to construct and apply the methods which are presently being used to evaluate the economic and mechanical susceptibility of a reservoir to flooding. This discussion will present a method for taking into account the effect of permeability variations in predicting the performance of water floods in depleted reservoirs. PERMEABILITY AND CAPACITY DISTRIBUTION It is generally agreed by most investigators that in a single phase system fluid will flow in a porous and permeable medium in proportion to the permeability of the medium. Producing formations are usually highly irregular in permeability, both vertically and horizontally. However, zones of higher or lower permeability are often found to exhibit lateral continuity. Thus, while structurally comparable stringers in adjacent wells may differ several fold in permeability values, they usually bear resemblance as being part of a general continuous higher or lower permeability section. It is generally agreed that where such stratification of permeability exists, injected water sweeps first the zones of higher permeability, and it is in these zones that "break-through" first occurs in the producing well. It is a basic assurnption of the presently described method that penetration of a water front follows the individual permeability variations as if such variations were continuous from input to producing well. This is admittedly not rigorously true, but can be justified as making possible a simplifying mathematical approach to an otherwise extremely complicated three dimensional flow problem. As a basis for study of the lateral flow of fluids in formations of irregular permeability, the irregularities may be conveniently represented by a permeability distribution curve and a capacity distribution curve. In obtaining these curves, the permeability values, regardless of their structural position in the formation, are rearranged in order of decreasing permeability. If these permeability values so arranged are plotted against the cumulative thickness, a permeability distribu- tion curve is obtained. This curve may then be likened to a "smoothed" permeability profile of the formation. In making comparison between different distribution curves it is convenient to state the permeabilities in terms of the ratio of the actual permeability values to the average permeability of the formation. These ratios termed "dimensionless permeabilities", are used in this paper rather than the permeabilities in terms of millidarcys. The capacity distribution curve is a lot of the cumulative capacity (starting with the highest permeabilities) versus the cumulative thickness. The capacity and thickness are given as fractions of the total capacity and thickness. Mathematically, the capacity distribution is the intergration of the permeability distribution curve. In practice it is convenient to first obtain the capacity distribution curve and derive from it a smoothed dimen-sionaless permeability curve. The method of obtaining the capacity distribution curve is illustrated in the successive column of Table 1, in which capacity and thickness are derived as fractions of their respective totals. If only a small number of permeability values are available, it is generally desirable to smooth the resultant curve. This has been done to give the capacity distribution curve shown in Figure 1. The differentiation of the capacity distribution curve to obtain the permeability distribution curve is shown in Table 2. Here, the capacity values are read from the smoothed curve at intervals of cumulative thickness, and the increments of capacity are divided by the increments of thickness to obtain the dimensionless permeability, K. Due to this stepwise procedure of calculation these premeability values must be plotted at the midpoints of the successive increments of thickness. The curve so obtained from these data is shown in Figure 1. The total area under the K' curve is equal to unity.

Jan 1, 1949

By W. E. Stiles

A method is presented for predicting the performance of water flooding operations in depleted, or nearly depleted, petroleum reservoirs. The method makes use of permeability variations and the vertical distribution of productive capacity. From these two parameters can be calculated the produced water cut versus the oil recovery. Derivations of the mathematical analogy is shown and sample calculations and curves of prediction are presented. Comparison is made of the predicted and actual performance of a typical 5-spot in an Illinois water flood. INTRODUCTION The use of water as a flooding medium in both depleted and "flush" oil reservoirs is gaining greater recognition and acceptance. Many of the shallower fields, depleted by primary production, have been and are being subjected to water injection in order to obtain some part of the large volume of oil remaining after primary production. Some of the earlier water flood installation proved highly discouraging and the value of water flooding was often questioned. Many of these earlier floods were haphazardly selected and developed as little was known of the physical characteristics and contents of the producing formations. The prior evaluation of the flood performance was impossible. During the past decade the development of the required reservoir engineering tools-—core analysis, reservoir fluid analysis, electric logs, fluid flow formulae, etc.—has allowed the engineer to construct and apply the methods which are presently being used to evaluate the economic and mechanical susceptibility of a reservoir to flooding. This discussion will present a method for taking into account the effect of permeability variations in predicting the performance of water floods in depleted reservoirs. PERMEABILITY AND CAPACITY DISTRIBUTION It is generally agreed by most investigators that in a single phase system fluid will flow in a porous and permeable medium in proportion to the permeability of the medium. Producing formations are usually highly irregular in permeability, both vertically and horizontally. However, zones of higher or lower permeability are often found to exhibit lateral continuity. Thus, while structurally comparable stringers in adjacent wells may differ several fold in permeability values, they usually bear resemblance as being part of a general continuous higher or lower permeability section. It is generally agreed that where such stratification of permeability exists, injected water sweeps first the zones of higher permeability, and it is in these zones that "break-through" first occurs in the producing well. It is a basic assurnption of the presently described method that penetration of a water front follows the individual permeability variations as if such variations were continuous from input to producing well. This is admittedly not rigorously true, but can be justified as making possible a simplifying mathematical approach to an otherwise extremely complicated three dimensional flow problem. As a basis for study of the lateral flow of fluids in formations of irregular permeability, the irregularities may be conveniently represented by a permeability distribution curve and a capacity distribution curve. In obtaining these curves, the permeability values, regardless of their structural position in the formation, are rearranged in order of decreasing permeability. If these permeability values so arranged are plotted against the cumulative thickness, a permeability distribu- tion curve is obtained. This curve may then be likened to a "smoothed" permeability profile of the formation. In making comparison between different distribution curves it is convenient to state the permeabilities in terms of the ratio of the actual permeability values to the average permeability of the formation. These ratios termed "dimensionless permeabilities", are used in this paper rather than the permeabilities in terms of millidarcys. The capacity distribution curve is a lot of the cumulative capacity (starting with the highest permeabilities) versus the cumulative thickness. The capacity and thickness are given as fractions of the total capacity and thickness. Mathematically, the capacity distribution is the intergration of the permeability distribution curve. In practice it is convenient to first obtain the capacity distribution curve and derive from it a smoothed dimen-sionaless permeability curve. The method of obtaining the capacity distribution curve is illustrated in the successive column of Table 1, in which capacity and thickness are derived as fractions of their respective totals. If only a small number of permeability values are available, it is generally desirable to smooth the resultant curve. This has been done to give the capacity distribution curve shown in Figure 1. The differentiation of the capacity distribution curve to obtain the permeability distribution curve is shown in Table 2. Here, the capacity values are read from the smoothed curve at intervals of cumulative thickness, and the increments of capacity are divided by the increments of thickness to obtain the dimensionless permeability, K. Due to this stepwise procedure of calculation these premeability values must be plotted at the midpoints of the successive increments of thickness. The curve so obtained from these data is shown in Figure 1. The total area under the K' curve is equal to unity.

Jan 1, 1949

By P. S. Bagga, P. T. Luckie

Liberation (the process of destroying the interlock between unwanted materials, such as mineral matter and pyrite, and coal) is one of the most important precursors to the benefication of raw coal in order to produce high grade marketable coals at economical recovery levels. Liberation is achieved usually by comminuting the raw coal. However, because of the significant interaction between the production of fine particles to achieve liberation and the cost of separating fine particles, the coal industry needs a method projecting the change in washability with size reduction. Such a projection can be combined with size reduction and separation simulators in order to examine completely beneficiation circuitry to produce coals of stringent specification. This paper presents the development of an appropriate liberation model for coal and its solution by Monte Carlo simulation to predict the expected washability of the raw coal as a function of comminution. It includes model validation with a known binary mixture as well as with a known teritary mixture.

Jan 1, 1983

By Eugene McAuliffe

THE last obtainable figures of the value of the coal mining investment are those contained in the U. S. Census Reports, from data gathered in 1919. The values shown therein and set forth below cover lands and equipment and do not include distribution and marketing facilities. The working capital necessary to meet pay rolls to purchase material and supplies and workmen's compensation insurance are likewise excluded.

Jan 1, 1926

University of Alabama University, Alabama Mining and Metallurgical Society WILLIAM R. HIGGS, President ROBERT E. MEAD, Secretary J. R. CUDWORTH, Faculty Sponsor MILTON H. FIBS, Counselor University of Arizona Tucson, Arizona Miners' Society LEE BLESSING, President R. W. CORY, Secretary E. P. MATHEWSON, Faculty Sponsor B. R. HATCHER, Counselor University of California at Los Angeles Los Angeles, California Student Chapter, A.I.M.E. ARCHER WARNE, President ROBERT JOHNSTON, Secretary E. K. SOPER, Faculty Sponsor EMILE HUGUENIN, Counselor Colorado School of Mines Golden, Colorado Student Chapter, A.I.M.E. J. Ross REED, President ANDREW PERRY, JR., Secretary J. BURNS READ, Faculty Sponsor A. W. NEWBERRY, Counselor

Jan 1, 1936

Jan 1, 1968

By H. S. Kalish, F. J. Dunkerley

INTRODUCTION AND PREVIOUS WORK THE determination of the low temperature tensile properties of tin and tin-lead alloys was initiated as part of an extensive research program on the phasial equilibria of tin binary systems below 13.2°C and the property changes associated therewith. A survey of the literature revealed that, although some room temperature tensile data were available for tin and a few tin-lead alloys, there existed no survey report of the low temperature tensile properties of these alloys. Accordingly, the tensile data for tin and a series of pure tin alloys containing from 0.01 to 50 pct lead were measured at eleven temperatures from -196 to +20°C and are presented here in survey form. In addition to the importance of these tensile data as a preliminary to certain fundamental research, this information promises also to be of commercial importance because of the 'extensive use now being made of tin-lead solders in refrigeration and gas liquefaction equipment. Occasional failures of the solders in low temperature service have been attributed by some metallurgists to low temperature mbrittlement, while others hold that failure results from the transformation of the tin-rich phase of the solder from the solid white (beta) to the powdered gray (alpha) form. From the data previously available in the literature, neither of these explanations can be completely corroborated. Cohen and van Lieshout1 have established the beta to alpha tin transformation at 13.2°C and have shown that strain accelerates the transformation. Further, they showed that although I pct lead retarded the transformation, it was not eliminated. It will be pertinent, then, to note whether the beta to alpha transformation of the tin-rich phase will occur in a short-time tensile test. Polanyi and Schmid2 studied the tensile . properties of tin single crystals at -185°C and report that tin is brittle at-this temperature. In 1946, Kostenets3 reported that pure tin was brittle at -196 and -253°C, while lead remained ductile at both these temperatures. He also studied the tensile properties of tin-lead alloys containing from 10 to 75 pct lead at -253, -196 and +17°C. The ultimate tensile strength increased more than two times in going from room temperature to -196°C and about three times that value on further cooling to -253°C. All the alloys were brittle at -6°C except those containing more than 75 pct lead. Even at -253°C, the 75 pct lead alloy retained 36 pct elonga¬tion in a 30 mm-ga. length. Creep is also an important phenomenon

Jan 1, 1948

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By A. T. Yu

The energy crisis and the resurgence of mining activities highlighted 1972-73. Added to the impetus for more effective materials handling systems has been the continued upward trend of inflation and labor costs. The need for addressing the fundamentals has led to the trend toward an overall systems approach to materials handling problems.1 The decision to move forward with large mining ventures such as the $150 million Hibbing taconite comple2 and the immense Carajas iron project must base themselves on indepth studies of a host of factors viewed in the context of an integral mine-transport-market system. What's important is that each ton of pellet or ore must be an economical commodity at the end users' furnace. The project can still be viaable even if certain components in the entire network appear to be individually costly.

Jan 2, 1974

By Oliver S. North

Reserves of perlite rock in the western section of the United States are immense. A geological report prepared for the Union Pacific RR showed proved tonnage of over 400 million tons in southern Nevada alone. Yet the perlite industry is not large, and most of the mines are relatively small. There are indications that perlite processing, thus far largely in the hands of relatively small com¬panies, is becoming concentrated in a few large companies. For economic reasons the bulk of perlite sold in crude form has come from three or four pro¬ducers, although many individual operators have sold small quantities to limited lists of purchasers. Perlite is one of ultra-lightweight aggregates. Owing to its outstanding qualities as an aggregate, variety of uses, and ultimate economy of use, it has enjoyed considerable popularity in this country in recent years. Marked foreign interest also has been evinced in this material.

Jan 2, 1955

By Harry C. Wiess

PETROLEUM has come to be one of the most important and essential of the mineral re- sources of the nation. It is the most advantageous source of mineral fuels and of lubricants, and as such it has provided in large measure for the phenomenal development in our modern means of travel and in the facilities for transportation by land, sea, and air. Without it the remark- ably rapid industrial and social progress in our country during the past thirty years could not have occurred. In view of its important place and the fact that petroleum is irreplaceable and existent in limited quantities in nature, its conservation is essential.

Jan 1, 1939

By F. M. Morris

A large reserve of thick coal in southwest Virginia was developed by Clinch-field Coal Co. in 1957-1958 to produce a nominal rate of 1500 tph raw coal. Operation features coal cleaning in transit. Refuse removed averages 35 pct. Evolution of plant from initial conception to completion is discussed, selection of means applied is explained, and performance to date us design expectation is described. The long-range plan calls for ultimate handling capacity of 40,000 tpd raw coal with anticipated clean coal capacity of 26,000 tpd on a three-shift basis. From a materials handling viewpoint, the Moss No. 3 operation is principally of interest as an ensemble. Generally speaking, it uses time-tested equipment and ideas but some of these are employed on a scale that may be new in the industry. At present about 20,000 tpd of raw coal are being handled. This is expected to increase to 40,000 tpd as soon as business conditions justify it. The purpose of this paper is to discuss the evolution of the plant as to materials handling practices, to describe briefly the equipment and methods used, and to comment on performance in certain areas. The subject divides itself naturally into four phases: 1) operations at the mines, 2) transportation from mines to plant, 3) raw coal handling into the plant, and 4) transportation of refuse away from the plant. OPERATIONS AT THE MINES The coal reserves for Moss 3 are in southwest Virginia where Dickenson, Russell, and Buchanan Counties come together. This area contains about 15 square miles and over 100,000,000 tons of coal. Here the No. 4 (Tiller) and No. 5 (Jawbone) seams of the Norton Formation lie so closely together that for practical purposes, they constitute one seam of coal. This seam, which is called the Thick Tiller, varies from 10 to 18 ft in thickness and underlies Sandy Ridge, a mountain cresting between 2400 and 3300 ft above sea-level. At 18-ft seam height (which is considered to be the maximum practical mining height) the parting between seams will be about 3 1/2 ft thick and each bench of coal will be about 7 ft thick. Depending on the amount of impurities in the seams Drover. total reject in this height coal may approximate 50 pet by wt. It will average about 35 pet for the property as a whole. The first move in developing this resource was made in 1953 when contour maps of several square miles around Duty, Va., including all the known outcrop, were made by photogrammetry. At this time, it was felt the prospective operation would be served by the Clinchfield Railroad and a photogrammetric route survey was made by this railroad from Haysi to Duty. Study of the resulting maps indicated only one site—adversely owned—which might accommodate the size washing plant to be erected. Water resources of a dependable nature seemed nonexistent. In 1954 bulk washability tests were made on the Tiller Bench at the Moss No. 1 preparation plant. The tests indicated this portion of the seam, mined separately, would wash to 4 pet ash with good recovery. Also in 1954, development of Moss No. 2, south of Sandy Ridge, was begun. This mine is in the Tiller Seam where it is about 100 ft below the Jawbone Seam. The reasons for developing a mine in the normal Tiller Seam before tackling the Thick Tiller seemed compelling: Railroad service could be established quickly, communications were better (though not good!), more was known of the seam (it had been mined in the years 1911 to 19241, and there was no essential property to be acquired. After some legal skirmishing, the Norfolk & Western Railroad was granted the right to serve the new mine. Three decades earlier the old mine had been served by the Clinchfield Railroad. The event which triggered active development of Moss 3 was the Appalachian Power Co.'s decision in 1956 to build a 450,000-kva power plant on Clinch River at Carbo. This solved the problem of marketing the steam coal which inevitably must be a product of a mine in the Thick Tiller. Management promptly decided to build the preparation plant at Carbo where an excellent site was owned; where railroad service existed; where telephone service could be obtained; and where roads, bridges, and water supply were tolerable.

Jan 1, 1961