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In 1974, iron pellet production in the Great Lakes region reached the 53-million-tpy level, accounting for more than 88% of the nation's pellet production. By 1978, pellet output from the Great Lakes district will climb to at least 78 million tpy. To achieve this 50% increase in output, the US iron and steel industry is investing at least $1.25 billion in new mines and pelletizing plants, mostly in Minnesota and Michigan. The biggest effort is concentrated in the Mesabi Range of Minnesota, the heart of the US steel industry's major iron ore source: here, plans for two new taconite plants-one by Hibbing Taconite Co., and the other by Inland Steel Mining Co.-and three major expansions of existing facilities are already under way to boost Minnesota's total iron pellet capacity from the current 41.2 million tpy to 65 million tpy by 1978. The facilities being expanded are those of US Steel Corp. (Minntac operations), Eveleth Taconite Co., and National Steel Pellet Co.

Jan 11, 1975

By Guy W. Crane

The geology of the Tintie mining district, fully treated, would occupy an elaborate monograph. This less comprehensive paper is devoted primarily to the occurrence and origin of the orebodies of the district, while space is given to those phases of the general geology only which seem necessary to a proper setting of such a discussion. The first geological report on the Tintic district to be issued was that by Tower and Smith, published in 1899, in the form of a paper in Part III of the Nineteenth Annual Report of the U. S. Geological Survey, and also as Folio No. 65, Geologic Atlas of the United States. About four years ago a second investigation was begun by Waldemar Lindgren and G. F. Loughlin, of the U. S. Geological Survey. With the exception of a paper on the oxidized zinc ores of the Tintic district by Mr. Loughlin, the results of their work are still unpublished. The writer's acquaintance with the district began in the fall of 1913, and has extended over a period of about two years. 1. Location The Tintic district lies about 65 miles due south of Salt Lake City, and on the west central slope of the Tintic Mountains, a short range, which forms the connecting link between the Oquirrh Mountains on the north and the Canyon Range on the south—the group constituting the first of the Basin ranges to the west of the Wasatch Mountains. The productive portion of the district embraces an area about 6 miles long and 2 miles wide, which is divided between Juab and Utah Counties. Within this area are the towns of Eureka, Mammoth, Robinson, Silver City, and Knightsville. All these points are tapped by the Denver & Rio Grande and S. P., L. A. & S. L. Railways. 2. Topography As in most of the Basin ranges, topographic relief is strong, dropping rapidly from peaks of 8,100 ft. above sea level to elevations of 5,600 ft. in Tintic Valley on the west and 4,500 ft. in Goshen Valley on the east.

Jan 1, 1916

By E. P. Bucklen, L. J. Prelaz, J. R. Lucas

Today, the future of the coal industry is extremely bright because coal can be produced at a cost which makes it competitive with other energy sources. However, the industry has been forewarned that further cost reductions must be made if it expects to remain a serious contender in the growing energy market. Surface as well as underground operations must be improved. The very same elements which affected production a few years ago still exist today and will continue to be present in the future. Improvements in production mining equipment, mining methods and auxiliary equipment have resulted in reduced costs over the past few years. Efforts must be continued toward further reducing costs in areas which have been previously improved and new areas should be investigated for possible improvements. Therefore, new techniques must be developed which will enable the industry to augment its current efforts in reducing the cost of mining. The logical place to begin evaluating the factors affecting production and cost is in the immediate face area where a substantial part of the total cost of mining is incurred. The coal industry stands to gain much through the development of new techniques for analyzing the numerous variables affecting production and by establishing methods for eliminating or minimizing the adverse effect of those variables which can be controlled. Intuition and trial-and-error methods have been the standard means used to resolve many of the factors which influence production. As a result, many elements are overlooked or ignored because their effects on production are not pronounced and cannot be readily resolved.

Jan 3, 1965

By J. F. K. Brown, E. E. Roecker

For three years The Hudson Coal Co. has used egg anthracite instead of coke in its foundry cupola. It has long passed the stage of being told it cannot be done—the metal would be cold, of poor quality, or the speed of operation would be low. From observation, from use and from test. we know that our metal is about average; the cast iron invokes no adverse comment from the machine shop and the speed suits our foundry requirements. We have there-iore resolved into practice anthracite combustion processes in an industrial furnace formerly using coke, under conditions suitable to our needs. When it comes to translating what we know into use in another cupola, however, where operating conditions may be different, or to explaining why, with seemingly the same conditions in our own cupola, we have attained sometimes a temperature of 2800° in the metal and at another time 2600°, or a speed of 5 tons per hour against 2 35 tons per hour, we cannot do it. We cannot do this because there seems to be an unbridged gap between laboratory ideas on combustion and practical ideas on com-bustiqn, which we have been unable to cross in terms understandable to us as practical consumers. This gap or bridge centers around the problem of how combustion actually takes place; and we believe it is controlled by two factors: (I) the character of the surface of the fuel and (2) the space provided and needed within which the oxygen of the air is allowed to or can act on that surface. It has long been contended on behalf of coke that porosity is an important factor in combustion, although, admittedly, there are two schools of opposing thought on this question, as revealed in technical literature. It has been understood generally that a light, porous coke is the best metallurgical fuel because there are more square inches of carbon exposed to the blast of air, therefore more useful carbon-dioxide reactions can take place in the given space between the entrance of the air and the point where useful work is done. It is now our general idea that the heavier and the harder the coal, the better the metallurgical fuel, provided the surface of that fuel is suitable for fast reaction with the oxygen. Time and Space A cupola is loaded in layers of fuel and iron. The horizontal line where the first charge of iron rests on the coal is the top of the bed coal. With us that is 22 in., usually, above the tuyeres where the air enters. This means 600 lb. of coal. Every subsequent coal layer is 100 to 125 lb. of coal, and that 100 to 125 lb. of coal melts 1000 lb. of iron at around 2700° in the ladle. This quantity of coal represents a 3-in. layer spread across the cupola. The question arises: If this 3 in. of coal will melt 1000 lb. of metal, why was it necessary to use 600 lb., or 22 in., of coal beneath and between the first charge of iron and the air entrance?

Jan 1, 1944

By E. E. Roecker, J. F. K. Brown

For three years The Hudson Coal Co. has used egg anthracite instead of coke in its foundry cupola. It has long passed the stage of being told it cannot be done—the metal would be cold, of poor quality, or the speed of operation would be low. From observation, from use and from test. we know that our metal is about average; the cast iron invokes no adverse comment from the machine shop and the speed suits our foundry requirements. We have there-iore resolved into practice anthracite combustion processes in an industrial furnace formerly using coke, under conditions suitable to our needs. When it comes to translating what we know into use in another cupola, however, where operating conditions may be different, or to explaining why, with seemingly the same conditions in our own cupola, we have attained sometimes a temperature of 2800° in the metal and at another time 2600°, or a speed of 5 tons per hour against 2 35 tons per hour, we cannot do it. We cannot do this because there seems to be an unbridged gap between laboratory ideas on combustion and practical ideas on com-bustiqn, which we have been unable to cross in terms understandable to us as practical consumers. This gap or bridge centers around the problem of how combustion actually takes place; and we believe it is controlled by two factors: (I) the character of the surface of the fuel and (2) the space provided and needed within which the oxygen of the air is allowed to or can act on that surface. It has long been contended on behalf of coke that porosity is an important factor in combustion, although, admittedly, there are two schools of opposing thought on this question, as revealed in technical literature. It has been understood generally that a light, porous coke is the best metallurgical fuel because there are more square inches of carbon exposed to the blast of air, therefore more useful carbon-dioxide reactions can take place in the given space between the entrance of the air and the point where useful work is done. It is now our general idea that the heavier and the harder the coal, the better the metallurgical fuel, provided the surface of that fuel is suitable for fast reaction with the oxygen. Time and Space A cupola is loaded in layers of fuel and iron. The horizontal line where the first charge of iron rests on the coal is the top of the bed coal. With us that is 22 in., usually, above the tuyeres where the air enters. This means 600 lb. of coal. Every subsequent coal layer is 100 to 125 lb. of coal, and that 100 to 125 lb. of coal melts 1000 lb. of iron at around 2700° in the ladle. This quantity of coal represents a 3-in. layer spread across the cupola. The question arises: If this 3 in. of coal will melt 1000 lb. of metal, why was it necessary to use 600 lb., or 22 in., of coal beneath and between the first charge of iron and the air entrance?

Jan 1, 1944

By H. Rush Spedden

When Robert H. Richards wrote Ore Dressing, the famous four-volume work published in the first decade of this century, the ore dresser was still largely concerned, with the mechanical treatment of ores. Since then, however, economics have transformed him into a mineral processing engineer who must increasingly deal with closely integrated facilities designed to produce a commodity for manufacturing. Moreover, this change will grow marked as the complex low- grade ores of the future make it ever more necessary for him not only to use the best combination of pyrometallurgical, hydrometallurgical and physical separation processes, but also to integrate these processes at the mine site. Nor are economic pressures the only ones propelling the processing engineer into a new era. The ecology drive has thrown him into a new situation in which his economic base is radically altered, and it could well precipitate as drastic a revolution in his technology as that which took place when flotation was developed.

Jan 1, 1971

By C. F. Davan, C. T. Houseman

The news pages of MINING ENGINEERING and other trade publications have for the past two years carried a great many items about new developments in phosphate and potash, reflecting the surge in world consumption of these products as fertilizer. Today the question often arises concerning how much growth can be further anticipated by the phosphate and potash industries. To answer this question, a 15-year projection of world fertilizer consumption in terms of the three basic plant food nutrients-nitrogen, phosphate and potash-has been prepared. Two of these, phosphate and potash, are basic products of the mining industry. Nitrogen, once an important product of mining, is now mainly a chemical product. Another mineral, sulfur, although not a primary nutrient, depends on the fertilizer manufacturing industry for the bulk of its market. In crop year 1962-63, there began a startling in- crease in world fertilizer consumption. In that year, consumption increased by about 10% and in the following year by 10-12%.

Jan 12, 1965

By Richard L. Gordon, Charles J. Johnson

Electric power demand doubles about every ten years, and because of the associated burgeoning fuel requirements, power stations absorb over half of the coal output in the United States. Throughout the 1960's, demand for coal utilized to generate electricity grew at an average annual rate of 6.5% but coal producers held back greater expansion because they expected nuclear plants to proliferate faster than they actually did. Even so, the growth rate is expected to continue at the 1960's rate until at least 1976. The continuation of the 6.5% growth rate beyond 1976, however, will require additional orders for coal-fired plants during the early 1970's. In fact, plants consuming a total of 130 million tons of coal must be ordered in the next few years. If the requisite plants materialize, total 1980 coal output for electrical power generation will be approximately 600 million tons.

Jan 1, 1971

By Lewis A. Smith

Chromium is one of the new metals, but considerable research has been required to determine an approximate record of its production from 1827 until the present. Its use in the form of pure metal is not extensive; of the chromite mined, about one-half is required for making ferrochro-mium, about two-fifths goes directly into refractories and the remainder is absorbed chiefly by the chemical trades. The industries that demand chromite for such uses are mainly in countries that produce little chromite, while the important chromite-producing countries are not generally consumers. The international flow of chromite is therefore of interest, as well as the political and commercial control of resources of the metal. World Production The first mining of chromite is said to have occurred as early as 1820, at the Roros deposit in Norway, but the first well-established output came from the Reed mine, in Maryland, about 1827. Thereafter the United States led the world in chromite production until 1860, when Turkey took the lead. Russia in turn became the chief producer from 1897 to 1902, inclusive. From 1903 to 1909, New Caledonia held first place, but alternated with Southern Rhodesia from 1910 to 1917, inclusive. In 1918, 1919 and 1920-1921, the United States, India and New Caledonia led, respectively. In 1922 Southern Rhodesia assumed first place and has since continued to be the world's foremost producer by a large margin. Fig. 1 illustrates the production of the leading countries from 1890 to 1929, inclusive. Fig. 2 shows the world production of chromite from 1890 to 1929, inclusive, in long tons and Table 2 gives the production by countries for the same period. The total world production for the 103 years, 1827 to 1929, ipclusive, is estimated at about 7,100,000 long tons, distributed as shown in Table 1. World Trade in 1928 In reviewing the figures for the flow of chromite in 1928, it is evident that the six countries that produce most consume little chromite and

Jan 1, 1931

By N. J. Grant, B. C. Giessen, M. Morris

The slvuclures of alloys in the system In-Bi have been investigated after (levy vapid queuching from the mell (splat cooling) to -190°C. Tuo-phase fields could be suppressed over most of the tota1 concentvalion range; five melastable phases a1, a2, ?, ?1. and ß exisl, of which four have simple elementtike structures. For example, ß has the A5 struclure of white till. Crystal-tographic data and coordination numbers for these phases aye given, as well as addilional information on the equilibrium diagram, where a new phase In5Bi3 teas found. THE phase diagrams of binary combinations of B elements from the groups B2, B3, B4, and B5 are usually of a simple type and show considerably fewer intermediate phases than combinations of these elements with A metals, transition metals, or the B1 noble metals. Aside from phases such as the typical B3-B5 compounds with the ZnS-B3 structure type, few stoichiometric phases are known; among them are In2Bi and 1nBi.l Nonstoichiometric phases include ß (In-Sn), a1, (In-Pb), and the tin-rich ? phases in the In-Sn, Cd-Sn, and Hg-Sn systems.' The first two phases have a close relationship to indium, while the latter group are known to be structurally related to a Sn. Thus, there are generally no immediate experimental data for the correlation of the crystal structures of B metals and B metal combinations with certain monotonously varying parameters, such as the valence electron concentration (VEC), or the average atomic size. Such parameters have been recognized as structure determining, e.g., for the transition 4 Sn — y phase,2,3 where nonintegral valence electron concentrations are encountered. It should be possible to extend such correlations to other binary systems, if nonequilibrium single-phase alloys could be produced over broad valence electron concentration ranges, and if their crystal structures could be regarded as being imposed by their electronic states. It has been shown in the case of the tin-rich 7 phases3 that alloy phases with typical crystal structures can be produced in B metal alloys rapidly quenched from the liquid to -190°C by the splat cooling technique due to Duwez,3-8 and reviewed in Refs. 5 and 6. The In-Bi system was selected because it extends over a valence electron concentration range in which several metastable, nonstoichiometrir phases could be expected to occur. Further, the low melting points of the known intermediate phases, In2Bi and InBi, Fig. 1(a), indicated low binding energies and thus possibly low driving forces for the formation Of the equilibrium structures. EXPERIMENTAL TECHNIQUES AND RESULTS The preparation of the quenched foils followed the practice described in Ref. 3. Master alloys were produced by melting of indium (99.97+) and bismuth (99.99+) in evacuated Vycor capsules or in an inert gas arc furnace; quantities of 20 mg were splat-cooled onto copper and silver substrates held at - 190°C; and crystal structures and lattice parameters were determined on a GE XRD-5 diffractometer using Cu Ka, radiation at -190°C and at room temperature. The duplication of each run with both substrates permitted the elimination of overlap of the substrate diffraction pattern and that of the investigated substance. The XRD patterns were usually taken from sin2 " = 0.03 to 0.35; the substrate was used as a means of internal calibration. The fractional accuracy of the lattice parameters of metastable phases is approximately 10-3 In the following, all percentages are in atomic percent. The stable and metastable phases found after rapid quenching to -190°C are listed in Table I, together with estimated concentration ranges and crystallographic data. The investigated alloys and phases present in them are given in Table 11. Except for the region between indium with 33 and 50 pct Bi, where a revision of the equilibrium phase diagram became necessary to include a new stoichiometric phase In5Bi3, the diffraction patterns taken after heating to room temperature agreed with those expected from the phase diagram, Fig. l(a). This suggests that the new nonstoichiometric phases are not stable at room temperature, thus following the observations made for the y phases based on tin.3 Results concerning the equilibrium phases In2Bi and In5Bi3 and the revision of the equilibrium phase diagram which is made necessary by the inclusion of In5Bi3 will be treated first. The Crystal Structure of In2Bi. Makarov7 had identified In2Bi as belonging to the AlB2-C32 type ; however, in a later paper the structure was revised.8 In2Bio was found to be of the Ni21n-B8ß type,21 with a = 5.496A, c = 6.57.9A, and N = 6 atoms per unit cell.9 This crystal structure was confirmed in the present work; definite evidence for the doubling of the c axis, as compared to the A1B2-C32 type, and for the proposed order structure, was found in the powder pattern. The observed lattice parameters agree well with those of Ref. 8; those measured at -190°C are given- in Table I. The Crystal Structure of In5Bi3. A new, equilibrium, intermediate phase was identified in slow-cooled, as-cast alloys; it was identified as In5Bi3. The structure has been worked out by Giessen and Grant;" lattice parameters at -190°C are given in Table I. This phase is probably identical with "In3Bi2", produced by vapor deposition techniques, but not recognized as an equilibrium phase by Palatnik et a1.11 After completion of the present work, the existence of In5Bi3 has also been demonstrated by superconductivity measurements on In-Bi alloys.20

Jan 1, 1968

By J. C. Kerkering, P. M. Daling

This paper discusses a study performed to evaluate the feasibility of transferring formal system safety assessment technology to the mining industry. A representative listing of formal techniques was evaluated to select the most useful methods for the mining industry to implement. Criteria for selecting methods included the interests, needs, and requirements of mine safety officials. This paper describes one of the selected methods, preliminary hazards analysis (PHA), and illustrates the analysis procedure for a mining example. This study also includes an assessment of the adaptability of selected methods to a user-interactive computer system for use by mine safety personnel. This paper includes a description of a conceptual computerized PHA-type safety program that facilitates performance of the safety analysis and can accommodate updates and revisions as necessary.

Jan 1, 1983

By Frederic Lahee

THE total production of crude oil in Texas during 1924 was said to be 133,613,985 bbl. as compared with 125,991,628 bbl. in 1923.1 Subtracting from these figures the yield of the Gulf Coast fields, there remains, for the other part. of Texas, a yield of 106,272,573 bbl. in 1924, as compared with 101,931,595 bbl. in 1923. This represents an increased, production of over 4,250,000 bbl. for all of Texas outside of the Gulf Coast or Salt Dome fields. Within this same region, in 1924, there were 3624 completions, of which 2352 were oil or oil and gas wells, 7 were gassers, and 1265 were dry holes. Thus, nearly 35 per cent. of the completions were dry holes. In 1923, there were 4134 completions, with 2853 oil or oil and gas wells, 2 gas wells, and 1279, or 31 per cent. of the total, dry holes. Outside of the Gulf Coast region, six major pools were discovered and developed in 1924.; two pools, previously discovered, were extended far beyond their limits at the end of 1923, and many smaller pools were opened. The major pools were the Big Lake field, in Reagan Co., the New Richland field, in Navarro Co., the Harmel, Carey, and Ragle pools, in south Archer County, and the Wortham field, in Freestone and Limestone Counties. Fields that were notably extended were the Panhandle producing area and the Luling field. Many wildcat wells have been drilled; most -of these were failures, but several opened .pools and several have had sufficiently favorable results to. lead to further drilling. The activity of the operators in searching for new production was very intense during 1924, and indications point to as feverish a campaign during the coming year. This will be stimulated by rising prices for crude oil. There are good reasons for believing that new pools will be discovered along the Mexia fault zone, in Archer, Young, Baylor, and Throckmorton Counties, in the great central region of Texas where we now have the Mitchell County and Reagan County pools, and very likely in the eastern Red River Counties including Cooke and Grayson. On the other hand, with the opening of the year, there is no new big pool in sight. An outstanding feature of the production end of the business is the increasing reliance placed on geological work and geological recommendations. Not in five or six years has there been such a demand for first-class geologists.

Jan 3, 1925

By R. T. Hukki

The purpose of this paper is to present an analysis of the unit operation of grinding and the circulating load, of the unit operation of classification and the circulating load, and of the two superimposed into one operation to clarify the overall trend of events relating the circulating load to the capacity of the closed circuit grinding system, to the specific surface area of the final fine product, and to the energy consumption in the closed circuit grinding process. The capacity of the closed circuit is frequently, if not always, determined by the classifier, not by the mill." With these words a startling idea was expressed in 1934 by Dorr and Anable.1 Today, some 30 years later, grinding in closed circuit is universally accepted. The basic curve presenting the capacity factor versus circulating load by Dorr and Anable has obviously been so convincing that not a single authoritative investigation regarding its validity seems to have been conducted. The only major subject of discussion appears to have been the relative merits of the 'high' and 'low' circulating loads. It seems to be universally agreed that much still remains to be hoped for regarding sharpness of classification. BASIC CURVES FOR CONSTANT MILL LOADING Fig. 1 presents a series of basic curves for constant mill loadings in function of new feed rate and circulating load. The mill loading indicated by each curve is the sum of hourly tonnages of new feed and circulating load. For example, at point 0 where the new feed rate is 100 t/h and the circulating load 300%, or 300 t/h, the mill loading is 400 t/h. In closed circuit grinding, the new feed rate corresponds to the rate of finished classified material produced. Although at this stage the effect of classification is excluded in construction of the basic curves, the trend indicated by these curves is strongly in favor of closed circuit operation under relatively small circulating loads. A study of the curves reveals that, e.g., a mill operating under a load of 200 t/h at point P where the new feed rate is 50 t/h and the circulating load 300% will also operate under the same total load at point Q where the new feed rate is 100 t/h and the circulating load 100%. In other words, a reduction of circulating load from 300% to 100% would lead to doubling of the new feed rate and doubling of the rate of production of finished material, without a change in the total mill loading. The obvious question is, is it also possible by a reduction of the circulating load in a closed industrial grinding circuit to increase the production of finished classified material and, if so, how and to what extent? GRINDING TESTS BY DORR AND ANABLE In a series of open circuit grinding tests reported by Dorr and Anable1 limestone was ground at various feed rates in a 3-ft ball mill. The original data are reproduced in Table I. It can be seen that with the increased feed rate production of -65-mesh material increased and the energy consumption per unit weight of the -65-mesh product decreased in the reverse ratio. The data shown in Table I have been reproduced for our purposes in Table II for a scale corresponding to grinding in a large industrial mill. This reproduction

Jan 1, 1968

By K. W. Nelson, John N. Abersold

INDUSTRIAL hygiene has been defined by Patty' as "the science and art of recognizing, evaluating, and controlling potentially harmful factors in the industrial environment." This definition implies thorough study of operations, evaluation of potentially harmful factors through air sampling, micro-analyses and other means and finally, appropriate medical and engineering control wherever indicated. The prevention of industrial health injuries is a vital part of operations of American industry today. Progress and interest in this field has increased steadily for many years, the most rapid progress having been attained, perhaps, during the last three decades. It is significant to note that there are now official agencies in 46 states actively concerned with industrial health problems and that a western field station has been established recently in Salt Lake City by the U. S. Public Health Service to augment its industrial hygiene services directed from headquarters of the National Institute of Health, Bethesda, Md. Many of the larger industries have found it advantageous to establish their own industrial hygiene departments. The American Smelting and Refining Co. is a world-wide organization engaged in the mining, smelting, and refining of lead, copper, zinc, silver, gold, by-product metals, including cadmium, arsenic, and others. In the United States there are 13 smelters and refineries, 11 secondary smelters or foundries, and a number of mines. Approximately 9000 workers are normally employed. It has long been the established company policy to seek out occupational hazards and provide safeguards for employee health. Protective equipment has been supplied to individual workers and exhaust ventilation installations have been in use in some operations for more than 40 years. All of the major units have their own medical departments which provide employees with excellent medical and hospital care. In 1937 full scale industrial hygiene studies were undertaken at the Selby Plant and were extended to most of the other smelters during the next three years. In 1945 the Department of Hygiene was organized with Professor Philip Drinker of Harvard University as Director and with Dr. S. S. Pinto as Medical Director. The department is responsible for coordinating and maintaining a program for the good health of all employees from top management down to the lowest paid day worker. It is essentially a service organization serving all of the United States plants regardless of location or size. Full and part-time physicians employed in all of the company's American plants and working in close cooperation with the Medical Director are responsible for de- termining the state of health of all the employees and giving treatment when necessary. In general, medical care is confined to accidents or illnesses occurring while the men are on the job. Among the duties of the doctors is the making of careful physical examinations of new employees and routine check-ups of old employees. In addition to medical care a primary responsibility of the department is the prevention of occupational illnesses. In this the main concern is with the working environment in relation to its effect on the worker. Environmental factors may be dusts, fumes, gases, toxic materials, heat, humidity, radiation, or noise. The objectives are: (1) Immediate control of these factors through the education of the worker, through providing the wearing of respirators or other protective devices, and through careful medical examinations and regular analysis of urine specimens; (2) a long range control program which may be accomplished by local exhaust ventilation, wetting of materials, changes in metallurgy, changes in methods of handling, or by use of special devices and special equipment. To accomplish these objectives a fine industrial hygiene laboratory was built in Salt Lake City and equipped to do routine and experimental work. Trained and experienced industrial hygienists obtain the facts by making frequent hygiene surveys. These surveys include tests of the air, studies of all processes, and careful investigation of ventilation, lighting, and general working conditions. Except in emergencies, the air contaminants and often the substances handled by the worker are sent to the laboratory for analysis by chemists and technicians specially trained in industrial hygiene methods. The findings are evaluated in terms of limits recommended by various State and Federal agencies, and in light of all available medical data. The methods used for studying the working environment involve all of the usual chemical and physical procedures employed in industrial hygiene. The Impinger, electric precipitator, thermal pre-cipitator, and filter paper sampler have been used to collect atmospheric dust and fume samples. Of special interest here is the filter paper sampler, shown in Fig. 1, which was developed by Dr. Silver-man at Harvard University. The instrument has been improved and is used very extensively in field studies. A water manometer connected behind an orifice is used to determine the rate of air flow. Calibration is effected by use of a standard gas meter or rotameter. The dust or fume is collected on a filter paper clamped between two rings, as shown in Fig. 2. The filter paper, such as Whatman No. 52, collects both dust and fume with a very high efficiency. The instrument is very convenient and easily transported. The solids collected on the filter paper are analyzed in the laboratory usually by use of a polar-ographic procedure. By this procedure it is possible to measure quantitatively in a single analysis the

Jan 1, 1952

By W. E. Brigham, P. W. Reed, J. N. Dew

By David J. Dunn

GENERAL NOTES Liner Design for Comminution Liners in crushers and mills are the mechanical link between machine and ore. The way liners transfer energy to media and ore determines to large degree the production rate, liner wear rate, maintenance cost, energy use, and availability of the comminution process. Plant capacity and mineral recovery economics, and return on investment, are thus strongly dependent upon these simple, low cost metal shapes of which so much is demanded. The complexity of dynamic relationships between liner configuration, material properties, ore characteristics, and operating parameters defy simplistic generalization. Rules-of-thumb and axioms of years past fall into disuse almost daily as new analytical methods, assisted by computers and sophisticated control systems, are applied to crushing and grinding processes. Recent literature is well supplied with references citing production benefits that improved liner designs produce. This chapter is intended to review especially promising liner designs and material developments and presents cameo case studies showing the benefits derived from their application to comminution circuits (Tables 1,2,4,5,7). Complete discussions of liner design and wear materials are beyond the scope of this chapter. Norman's comprehensive studies in references (1) and (2) and Mclvor's analyses of mill speed/configuration effects (3) are especially useful reading for designers and operators of crushing and grinding circuits. Changing trends in ore, product value, and costs of wear materials, interest, labor, and energy, make ongoing evolution of liner design equally important for existing mines and for projects still on the drafter's board. Original designs supplied as new equipment must, of necessity, be best estimates. These can usually be improved as operating experience replaces engineering forecasts. Advent of new materials or design concepts such as the Angular Spiral Liners (Korpi (4) ) or Norman's Bimetallic design (2) offer improvement potential that must be evaluated if a given circuit is to be kept competitive versus others making similar products.

Jan 1, 1982

By L. Graydon Carpenter, Donald E. Garrett

Probably the largest single tungsten deposit in the U. S. is one that has yet to produce any tungsten; it is not even listed in tables showing U. S. reserves. This deposit is at Searles Lake, Calif., where the brine contains about 70 ppm WO3. Small though the concentration is, the amount of brine in the lake is so great that it is estimated at 170 million lb of WO3 (8.5 million units), equaling the total of all other known reserves in the U. S. This dry lake is located in the basin formed by the Argus and Slate ranges in the northeastern corner of the Mojave Desert (Fig. 1). It consists of a massive salt body some 35 sq miles in extent, varying in thickness from 0 to 120 ft and containing more than 3 billion tons of salt. A central section of about L. G. CARPENTER and D. E. GARRETT are with the Research Dept., American Potash & Chemical Corp., Trona, Calif. TP 47881. Manuscript, Nov. 4, 1958. AlME Trans., Vol. 214, 1959. 8 sq miles is exposed salt; the remaining area is overlain by mud. Except for the exposed sodium chloride (10 to 15 ft thick) the body consists of a very complex and heterogenous mixture of soluble crystals, containing at least eight major phases. It is only about 53 pct solid, the voids being occupied by a dense brine which is the raw material for two chemical companies operating in the valley.

Jan 3, 1959

By J. Quigley

As the outcrop mines in the West developed into underground operations, systems of ground support were gradually evolved. In the early coal mines there was little need for support except near the dirt line in portals, where stone masonry was common. Where the top was shaley or broken, native pine props with light cross bars and legs furnished enough support even in Utah's 25-ft coal seams. As depth of workings increased. roofs and backs of the same general nature as those near the surface became more and more unstable and required more and more support. Some coal airways show this tendency very clearly. From the surface down the same type of roof shows deterioration which an experienced eye can translate into a measure of depth under surface rather than change in rock characteristics. Rock bolts, developed by various companies and by the U. S. Bureau of Mines, have become an effective substitute for timber in sections of some metal and nonmetal mines formerly requiring escessive timber support, and further use of war surplus landing mats, chain link fencing, and a new punched channel developed by one of the steel companies has enabled other mines to operate deposits where costs of timber and lack of clearance for timber support would have prohibited mining. The block caving mines have made extensive use of reinforced concrete underground to achieve similar ends under difficult conditions. Steel sets are standard in many Bureau of Reclamation projects, although these are usually covered in with concrete to make the permanent structures the Bureau's reclamation projects require. But the use of steel in mining operations is limited and has been confined principally to the iron ore mines of Michigan, Wisconsin, and Minnesota. Some mines have installed used rail as posts, caps, and crossbars, but a rail section is not suited for load carrying, and used rails are generally brittle. having a tendency to fail without warning when overloaded. European mines were the first to reach the size of worked out areas and depths of cover resulting in major roof problems. The Europeans resorted to pack walls and masonry walls, in conjunction with timber arched sets. rail arches, and combination timber and rail and steel arches. The give in these pack walls and wooden blocking was supplemented by a hinge in the center of the arch. This design is called an articulated arch Through various refinements of this principle of the support giving graduallv with the load. Toussaint-Heintzmann developed the yielding or sliding arches, in which yield is accomplished by friction in the overlapping joints of the arch. This type has gained widespread acceptance in the Ruhr and Lorraine Basin and is being manufactured by Bethlehem Steel for sale in this country. In North America the anthracite mines in Pennsylvania, followed by certain iron ore mines in upper Michigan and Canada, were the first to employ these arches to any extent. The practice was later adopted by Kennecott at Ruth, Nev., and by others. Despite high initial cost, the use of these arches is growing in many parts of the country because of their suitability in heavy ground. In its present form of manufacture the yield-able arch consists of open U-shaped rolled section with heavy beads on the edge. The open edge of the U is placed toward the wall. The section nests in another section of the same dimensions, and an arch can be built up from rolled radii and tangents of various weights and lengths. Sections are fastened together by U bolts and saddles. The lap on the joint varies from 12 to 24 in., and ordinarily the bolts are tightened with a 1-in. drive air wrench. The arches are spaced with channel struts held by J bolts and saddles. Sections can also be obtained that are composed of various combinations of radii and tangents and true circles. The joints can be placed to bear against anticipated loads and asymmetrical loads imposed by dipping strata. In the arches now being manufactured clearance widths up to 19 ft are obtainable in weights of sections from 9 to 30 lb per ft. The circular cross sections are available in the same weights ranging from 8 to 16 ft diam. At present most of the arches sold are supplied only in carload lots. It is hoped that demand will grow so that distributors can stock various weights and sections to give small operators a chance to try this new type of rock support under their own particular conditions. Several excellent papers have discussed the properties of various sections now manufactured, the dimensions of the sets obtainable, and their application under widely differing conditions. The present article will describe the methods and results of a special use of the arches at Kaiser Steel mine No. 3. Sunnyside, Utah. Problem at Mine No. 3 : In 1953 Kaiser Steel Corp. laid out Sunnyside mine No. 3 to recover coking coal left by the previous operator, Utah Fuel Co.. below workings that had been abandoned in 1928. Two seams had been worked, the upper and lower, separated by 30 to 42 ft of rock. Approximately 10 million tons of coal had been extracted from this area some 3000 ft down the itch from the outcrop to a 1500-ft depth of cover. The mine had been opened by slopes in both upper and lower seams. Sometime in the late 1920's the lower slope

Jan 1, 1960

By Hugh D. Pallister, Richard W. Smith

The Alabama flake-graphite industry has flourished only in times of war when importations of foreign graphite for crucible use have been greatly curtailed or cut off. World War I was a boom period and in 1918 some 30 plants produced 7,795,475 lb. of flake graphite. The industry languished following the war and no graphite was produced from 1929 to 1939. The Federal Bureau of Mines began investigating the Alabama graphite deposits in 1940 and worked out modern recovery methods that were applied to production during part of World War II. A prospecting program disclosed reserves of over 25 million tons of graphitic schists. In 1942 the War Production Board authorized the enlargement of the only plant then producing flake graphite and the construction of two new plants. Production of flake graphite from these three plants in 1943 is estimated at 8,100,000 lb. Increased imports from Madagascar late in 1943 shut down one of the new plants and curtailed production at the other. Alabama flake graphite has not yet been able to compete with Madagascar graphite for use in the manufacture of any but small graphite crucibles. The future of the industry must depend on meeting the higher carbon content demanded for other uses.

Jan 1, 1945

By J. Parke Channing

THE production and use of copper in the world, like that of other metals, has been of slow growth. In 1880, production in the United States, was only about 60,000,000 lb. and the world's production 325,000,000 lb. Up to 1913 there was a gradual and regular increase in production; in 1912 the production was 1,250,000,000 lb. for the United States and 2,200,-000,000 lb. for the world. At the outbreak of the war there was a slight falling off, both in the United States and elsewhere, but when it was realized that the war was not to be a short one and that copper was one of the most important things for carrying it on, there was a tremendous increase in production every-where. Roughly speaking, the average production for the years 1916, 1917 and 1918 was very close to two billion pounds in the United States and three billion pounds in the world. WAR DEMANDS STIMULATED PRODUCTION Not only did the production of copper in the United States increase, but the same may be said of .all the base metals, particularly of zinc. Much of the zinc used by the Allies came from Australia, but had been smelted in Belgium and Germany, now closed to them. We were called upon to make good this lost supply and in feverish haste built new zinc smelteries and opened new mines. The armistice found the United States with a large surplus of copper on hand, variously estimated at from 800,000,000 to 1,000,000,000 lb. In addition to this surplus in the United States, there were considerable stocks in the hands of the Allies and also immense quantities of unused brass. To. complicate the situa-tion still further, the reclamation of old brass shell cases and other used material was instituted, still further curtailing the demand for new material. This situation was- not appreciated at the time and it was assumed, that the Central Powers would immediately come into the market ? replenish their exhausted stock of copper. During the war Germany had torn down copper trolley wires and roofs, and melted up bronze bells, doors, statues and other copper-bearing material to reclaim this metal for military purposes. But Germany-was unable or at least unwilling immediately to replace her copper substitutes.

Jan 1, 1923