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By Jack W. Carlson

Considering the magnitude of today's public concern with critical mineral and fuel materials, regarding both their midterm and long term availability, the Nation is fortunate to have in place and operating the coordinated data systems of the Geological Survey and the Bureau of Mines. Under the supervision and policy direction of the Assistant Secretary--Energy and Minerals, these Department of the Interior agencies have long been responsible for important mineral activities in the Federal Government. Aside from their highly respected mineral industry publications such as the monthly Mineral Industry Surveys, annual Commodity Data Summaries and topical Professional Papers and Reports of Investigation, the Bureau of Mines and Geological Survey data bases and expertise have been readily available for use in emergency situations. Policymakers and decisionmakers in government have placed major reliance upon this ready mineral data and analytical capability in World War II, the Korean War, the recent price control era, and the Arab oil embargo.

Jan 1, 1975

By I. L. Sills

Introduction So many minerals and mineral products are used in the glass industry that it would require an unduly long paper to deal adequately with all of them. Only those that go to make up the batch for the more common types of glass will be discussed here, and even with these only, the list is a lengthy one. It is true that for the body of the glass comparatively few materials are required-as silica sand, soda-ash, lime, and feldspar. The others are used mainly to colour or decolorize the glass, or to give it some other desired characteristic, such as high lustre, greater physical strength, or opalescence; but, although they may be present in very small quantities, they are just as in-dispensable to the glass industry as are the main constituents of the batch. Sand Sand is essentially quartz, the dioxide of silicon, Si02. Most glass contains from 71J to 75 per cent silica. Sand for use in the glass industry, or 'glass sand', is obtained both from deposits of unconsolidated material, commonly referred to as 'sand', and from solid rock-sandstone, quartz, or quartzite. For the manufacture of glass, the material must possess a high degree of purity, as noted below, so that, although sand is widespread in most parts of the world, and sandstone, quartz, and quartzite are not at all uncommon, it is only in certain places that suitable deposits are found. At the present time, most of the glass sand used in Canada is imported from France, Belgium, and the United States. The French material, from Fontainebleau, near Paris, is known the world over as a very high quality glass sand. That from Belgium is nearly as good and is lower is price.

Jan 1, 1934

By Alexander Gakner

Since the end of World War II the Soviet Union has made great strides toward economic development of the country. Today it is firmly established as the world's second largest industrial power after the United States and it has committed itself before the people on both sides of the Iron Curtain to catch up with and to overtake us within a few decades. No nation could achieve such a degree of industrialization or aspire to become the world's industrial leader without a large supply of many basic mineral raw materials or at least easy access to such resources. In addition, it requires technical competence, a large skilled labor force, and a high rate of capital investment. Why should we concern ourselves with these Soviet aspirations? The Free World's yardstick of measuring a nation's well-being is the standard of living of the people in that country. Presumably even Communist governments are committed to some, however gradual, improvement in the standard of living of their people. The population of the USSR is 20 pct greater than that of the United States. Hence, if they ever achieve a standard of living equal to ours, they must by all logic reach a greater overall industrial output. However, in the U.S.S.R. the standard of living of the masses has not kept pace with the expanding industrial production and is still far behind that of the United States. Our concern, therefore, is with what the USSR intends to accomplish with its expanding industrial power: Will they produce large mineral raw-material surpluses and manufacture products to compete with us for world markets and attempt to put us out of business if not use it directly to wage war; or will they use their resources merely to produce more consumer goods for the Soviet people; or will they, along with other Communist nations in the not-too-distant future, compete with us for the world's sources of mineral raw-materials to maintain a high level of mass consumption?

Jan 6, 1960

By Joan Weinberg

Seasonally adjusted new factory orders for primary metals declined in January, February, and March, especially in the steel industry. The decline was part of an apparent slowdown for the entire manufacturing sector. Orders for all durable goods, including not only primary metals, but also clay, glass, and stone products, fabricated metal products, machinery, and transportation equipment, fell in January and February. In March, orders rose a slight 0.8%. According to the U.S. Department of Commerce, the downturns in durable goods orders in January and February were the first back-to-back declines in 3 years. They were largely attributed to a drop in orders for transportation equipment, mainly auto-mobiles and aircraft, while the small gain in March was due to a surge in aircraft orders. Higher interest rates may be partly responsible for the slowdown in durable goods orders over the past 3 months. The higher rates have also hit the construction industry--another major end-user of minerals. Construction of new homes and apartments fell by a seasonally adjusted annual rate of 5.4% in March, following a 12% decline in February.

Jan 1, 1989

By D. S. Brown

Few issues pose quite as profound consequences to the quality of life and the creation and distribution of wealth as how we protect the natural environment. No longer a peripheral but a global concern, the environment, once almost nobody's business, has now become everybody's business. It is also the single biggest issue facing mining and mineral activities. This heightened ecological awareness and sensitivity have spawned a new commitment to restore, protect, and preserve nature. But that very commitment poses difficult socio-economic trade-offs, such as: What and how many resources can be given up to prevent and reduce pollution? What are the opportunity costs associated with meeting higher environmental standards? What problems are more important or pressing than others and should be given priority? Choices have to be made that may entail unintended consequences for other societal goals.

Jan 1, 1994

By David S. Brown

What troubles me so often about debates over the environment, which I have heard since I entered government in the late 1960's, is that the issues are so often presented as a false dilemma--the either-or fallacy in logic. The either-or fallacy allows only two possible choices: Either protect the environment or exploit it. In reality, however, there may be a few more options. The diverse, often competing, values that the Earth offers require us to weigh and balance those options through a range of policy options. A quality environment and economic growth and opportunity are not mutually exclusive. First, let me acknowledge that I support a set of resource values recognized in U.S. laws and deemed important down through history for the success of free, progressive, and secure societies: the importance of access to raw materials. Increasingly, however, mineral raw materials have been viewed at best as not particularly relevant to a sophisticated, technologic society, and at worst monstrously destructive of our natural environment. Both perceptions, in my judgment, are wrong.

Jan 1, 1989

By P. A. Von Wielligh

Without minerals, man cannot possibly hope to maintain his present standard of living. Professional mining men throughout the world, particularly in mineral-exporting countries like South Africa, are faced with the increasingly difficult task of ensuring adequate supplies of minerals. The problems of gradually diminishing mineral reserves and rapidly rising mining costs can to a certain extent be overcome by the use of more-sophisticated technological techniques and by the improvement of efficiency wherever possible. However, there is growing concern in some quarters, that, despite the ingenuity of engineers, geologists, metallurgists, and scientists, the earth's mineral reserves will eventually become insufficient to meet man's needs. In order to ensure the continued well-being of the earth's population, it is therefore imperative that more attention should be paid to mineral research and the training of specialists in the field of mineral reclamation.

Jan 1, 1978

By A. Kenneth Schellinger

DURING the comminution of a brittle material in the presence of dry air, no known phase change or chemical reaction takes place. The energy changes associated with the comminution are those of the transformation of kinetic energy into heat energy and the energy absorption represented by the actual rupture of molecular and/or ionic bonds. The random nature of the actual grinding process in a tumbling-mill often precludes the effective application of a rupture force on each individual particle of brittle mineral present. The result must be a great deal of useless vibration on an atomic scale in both the brittle mineral and the grinding medium below the elastic limit, which vibration is eventually manifested as heat energy in the mill. The actual rupture of molecular, or ionic, bonds creates new surface on the brittle minerals; each square centimeter of this new surface then must possess a number of potential bonds, each of which could reform theoretically with the release of energy. The sum total of this potential energy of bonding per unit of new surface, minus bond satiation effects of the enclosing gas or liquid, must be the surface energy of the mineral. If the comminution of the brittle mineral were carried out within the chamber of a calorimeter, the energy released as heat could be determined rather accurately. Moreover, the total kinetic energy input to the comminuting mass (mineral and medium) conceivably could be derived by a suitable torquemeter and revolution counter if the whole calorimeter were revolved. The difference between these two energy figures should then represent the net energy (gross energy input—heat enerb output) used in creating new surface, gliding or slip on atomic planes, twinning, and possibly other lattice rearrangements within the mineral. If it be postulated further that with a hard, brittle crystalline material, such as quartz, the energy used to produce gliding, twinning, and other permanent lattice movements is a very small fraction of the total net energy absorbed, then this energy absorption within the grinding calorimeter can be ascribed to the creation of new surface energy alone. Such energy is, from the thermody-namic standpoint, the true effective work of the grinding device, in this case a tumbling-medium mill. The ratio of this effective work (in calories) to the total energy input (in calories) should express the thermodynamic efficiency of the comminution taking place under the conditions used in the grinding calorimeter. This approach to the measurement of the energetics of tumbling-mill grinding was suggested by some previous work by Fahrenwald and coworkers' with a laboratory ball mill and a thermometer. They measured the efficiency of grinding based on the formula: Pct Grinding Efficiency = Energy (calories) — Thermal Energy x 100 1npZ Output (calories) Energy input (calories) Efficiencies thus measured, while grinding quartz sand, ranged from 19.7 pct to 7.4 pct. Cook in 1915,' used this approach to the problem of grinding efficiencies in stamp batteries and obtained the energy input by stamp weights times drop distances and the energy output by means of thermometers and pulp flow rates. He calculated that 80 pct of the input energy to the stamp shoes was wasted as sensible heat in the ore pulp. No other studies of comminution by this method of energy differences came to the author's attention. Reasonable agreement as to the order-of-mag-nitude of comminution efficiencies in tumbling-media mills is lacking in the literature. There is a widespread impression among technical workers that the grinding efficiency of the industrial ball mill may be on the order of 1 pct, or even a fraction

Jan 1, 1952

By R. A. Deju, R. B. Bhappu

Further information on the surface characteristics of the silicate minerals was obtained by experimentally correlating the oxygen-silicon ratio of the silicates with their electrokinetic properties. Specifically, the behavior of the minerals in an electric field was investigated using zero point of charge as a parameter. A modified electrophoresis cell was used for the determination of electrophoretic mobilities. It was shown that there is an increase in the zero point of charge with increasing oxygen-silicon ratio in the silicates. Finally, a chemical interpretation was presented to describe the surface phenomena observed in the experiments. In recent years great interest in surface chemistry has produced a number of methods to study the surfaces of minerals and to observe the phenomena occurring at their interfaces. One such method is discussed in this paper. A rather complete review of the surface characteristics of the oxide minerals has been undertaken by parks,' but, a similar study has not been made for the silicate minerals. Two previous papers by the authors2'3 attempted to remedy this situation. In these papers it was demonstrated that the reaction between the silicate mineral particles and acidic water involves mainly an exchange of metal ions for hydrogen ions on the surface of the solid, leading to an increase in pH of the aqueous phase. Also, the degree of reaction was found to depend on the oxygen-silicon ratio of the silicate structure, being least for the quartz and greatest for the olivines. A hypothesis was then set forth explaining that upon the fracturing of a silicate mineral, the oxygen-metal bond (which is almost entirely ionic in character) will break more easily than the stronger oxygen-silicon bond. This results in the exposure of unsatisfied oxygen sites on one hand and active, exchangeable cations on the other. These active sites on the fractured surfaces are then available for reaction with the hydrogen ions in acidified aqueous solution. Thus, an increase in degree of adsorption of hydrogen ions occurs as the oxygen-silicon ratio increases due to the presence of a larger number of active sites on such minerals. A correlation between oxygen-silicon ratio and electrokinetic properties will be further pursued in this paper from both the theoretical and experimental points of view. Efforts will then be made to present a chemical interpretation of surface phenomena observed in the experiments. It is believed that this study may supply pertinent information on the behavior of silicate minerals in froth flotation systems. ELECTROKINETIC STUDIES Equipment: In a previous paper,2 Deju and Bhappu described a simple method for determing the zero point of charge using a potential across a cata-phoresis cell and observing particle movement by means of a Rayoscope* optical system. This setup is simple in design and operation and extremely reliable. With care, reproducible results within 1.5% accuracy may be obtained. A new setup has been developed, based partly on the design by Zeta-Meter Corp.4 and partly on previous cells used at the New Mexico Bureau of Mines. This new, modified cell is shown in Fig. 1. Although

Jan 1, 1967

By R. W. Smith, K. H. Lee

Bond type simulated closed-circuit grindability tests and batch grindability tests were run on six pure minerals plus a limestone and a taconite ore. The simulated closed-circuit tests were made at five different meshes of grind. The minerals tested varied widely in hardness and breakage pattern. For all simulated closed-circuit tests both Bond and Holmes work indices were calculated and compared. The Bond grindability testlm6 has been widely and quite successfully used for predictions of such items as ball mill power requirements and as an aid in plant comminution equipment selection. The test is interesting in that although run as a series of batch tests it is over-all a simulated closed-circuit test with a circulating load of 250%. The simulated closed-circuit feature is probably the primary reason for the notable success of the test. Perhaps the major criticism that can be brought against the test is that it is time consuming, since it normally takes a trained technician about an eight-hour shift to perform it.4 For many purposes this is an acceptable time requirement. However, if a grindability test is to be used for mill control purposes it should be performed in as short a time interval as possible, preferably under an hour. Sometimes in industry strictly batch tumbling mill tests are run and an attempt is made to correlate such batch tests with plant mill operation. Then an attempt is made to adjust operating tumbling mills based on data from the tests. The tests, however, frequently do not correlate well with operating data. The purpose of the present work was to explore some of the similarities and differences between a simulated closed-circui.t grindability test and a batch type test. Six different minerals plus a limestone and a taconite ore were tested in the experimentation. These materials varied widely in hardness, breakage characteristics, etc. Standard dry Bond grindability tests were mn on all materials at 48, 65, 100, 150 and 200 meshes of grind. Feed material was prepared by stage cmshing to -6 mesh. Initial feed volume was 700 cc. Care was taken to maintain the standard constant weight (about 20125 gm) and number of balls (283) during all the tests. Rotational speed was 70 rpm. Table I lists the materials studied along with pertinent data regarding the materials and Bond grindabilities obtained. In addition, dry batch tests were performed on all materials. These tests were made in the same mill as was used for the Bond tests. An identical ball charge and an identical feed volume were used. After rotating the mill for a predetermined number of revolutions at the same rotational speed as for the Bond test, the material was removed and separated from the ball charge. A screen analysis was performed on this product. Knowing the original size distribution of the feed and the size distribution of the product, it was possible to calculate the grindability or grams of a particular mesh produced per revolution of the mill. After the screen analysis of the product, the various fractions were recombined and placed back in the mill for further grinding. After grinding for additional revolutions of the mill the procedure was repeated. Thus, it was possible to obtain grindabilities for each material at a number of meshes of grind and for various revolutions of the mill. Grind-abilities obtained from batch tests are shown in Table 11. COMPARISON OF BOND AND BATCH GRINDABILITIES Figs. 1A through 1D show relationships between batch and Bond grindabilities. Figs. lA, lB, 1C and 1D show respectively graphs indicating the relationships for batch tests run for 100, 300, 500 and 700 revolutions of the mill. For a particular material on each graph, points are shown for 200, 150, 100, 65, and 48 meshes of grind except for several missing coarse meshes of grind for galena which did not fall conveniently on the figures. For each material and graph the point nearest the origin is for 200 mesh, the next for 150 mesh, the next for 100 mesh, etc. Considering Fig. lA, it is interesting to note that all relationships shown are essentially straight lines. There are, however, considerable differences

Jan 1, 1969

By D. C. Seidel, E. F. Fitzhugh

The Colorado School of Mines Research Foundation has developed a hydrometallurgical process for recovering nickel from oxidized ores, including both the iron-rich laterites and magnesium-rich, soft silicates. Known as the HSO-HTCP (Hydrothermal Sulfidization Oxidation-High Temperature Cementation in Pulp) Process, the system consists of feed preparation, sulfidization, oxidation, precipitation (cementation), and calcination and melting. This paper deals primarily with the sulfidization and oxidation phases of the process. Nickel sulfide ores lend themselves readily to concentration before smelting or pressure leaching, but neither of the major oxidized ore types — the iron-rich laterites and magnesium-rich, soft silicates-has been directly concentrated. Upgrading of crude ore in current practice is limited to cobbing lumps of the harder, lower grade rock (protore) from the softer, enriched silicate ore, because only a small percentage of the nickel in these ores occurs in discrete, contrasting mineral particles. There is no apparent mineralogical contrast in the superficial, iron-rich lateritic ores and, in fact, there is no firm assurance that the nickel atoms are within the lattices of the limonite minerals, or else irregularly adsorbed on the limonite. Within underlying, "rotten rock" silicates, occasional veinlets of gamierite and related nickeliferous silicates are found.' The gamierite veinlets, however, are only incidental, and the majority of the nickel atoms occur as erratic replacements of magnesium atoms in the micaceous chlorites of the ore mass. Pyrometallurgical processes account for nearly all the nickel currently being recovered from oxidized ores. Plants in New Caledonia, Oregon, Japan, Brazil and Greece smelt the ores to make ferronickels or a sulfide matte,2 and the ammoniacal leaching3 at Nicaro, Cuba, is preceded by a reducing roasL4 The only fully hydrometallurgical installation has been the politically ill-fated Freeport Nickel Co. enterprise, and this process was limited to lateritic ores.5p6 A hydrometallurgical technique which might handle both laterite and the typically richer silicate ores has been an enticing goal. The technique described here was developed at the Colorado School of Mines Research Foundation, Inc. on behalf of the Republic Steel Corp. It uses sulfur, heat, air and metallic iron to recover nickel from these ores. Inasmuch as the ores are treated as aqueous slurries, the cost of drying a plant feed that normally carried 30% or more moisture is eliminated, and there are no dust problems. Patent applications have been filed by Republic Steel Corp. on the procedures used for getting the metal into solution and on the subsequent recovery without liquids-so l ids separation of the slurry. It is referred to as the HSO-HTCP Process (Hydrothermal Sulfidization Oxidation — High Temperature Cementation in Pulp). THE PROCESS FLOWSHEET The results of various experimental studies have been combined to form the continuous flowsheet illustrated in Fig. 1. This flowsheet serves as a general pattern showing the sequence of the operations which include: (1) feed preparation; (2) sulfidization; (3) oxidation; (4) precipitation (cementation); and (5) calcination and melting. This paper deals primarily with the sulfidization and oxidation phases of the process. Feed preparation: In general, the silicate ores can be considered as earthy rather than hard, and preparation might be carried out before the addition of water in a trommel, or with water in a scrubber or log washer. This breaks up the earthy particles, allowing the hard lumps to be separated by screening. The oversize material from most ores is relatively barren and may be discarded. If there is occasion to control the iron-magnesium ratio in the plant feed, laterite, which normally breaks up readily into fine sized particles, and silicates are mixed in suitable proportions. The fine product slurry from the initial steps is mixed with elemental sulfur and fed to a conventional ball mill, although the amount of actual grinding that takes place is relatively small. The object of milling is to complete the disintegration to natural particle size and to produce an intimate mixture of the ore and elemental sulfur. This pulp, which is almost entirely -200 mesh, has a pH of about 7.0 and normally can

Jan 1, 1969

By R. B. Adair, J. S. Browning

The Bureau of Mines conducted laboratory research to determine the applicability of heavy liquid separation to the evaluation of certain mica ores and plant products. After careful standardization of the particular mica ores and associated gangue, accurate analysis of muscovite content could be obtained from results of heavy liquid separation. Where applicable, the method should be particularly suited to the analysis of ores and control of plant operations wherein savings in time would be of considerable advantage. Specifically, there are many instances where a quick plant control method should enable the production of a more uniform grade of mica concentrate. The laboratory use of various heavy liquids for separating minerals of different specific gravities has been recognized for many years.1-8 Sink-float procedures have proven their utility as a rapid and moderately accurate means for determining the mineral content of ores, test products, and plant products in certain nonmetallic concentrating operations. A standard method of analysis for mica in ore samples has not been published or generally accepted by the mica industry. Recently, the U.S. Bureau of Mines investigated methods for analyzing mica by heavy liquid separation. This report has been prepared to make the methods available to others who may need simple, relatively rapid methods of determining the mica content of ores and plant products. The heavy liquid method of mica determination has recently been adopted by one of the major mica producers for evaluating exploration and mill control samples. DESCRIPTION OF THE ORES AND HEAVY LlQLllD The weathered mica pegmatite ores used in the investigation were obtained from commercial mica op erations in Alabama, Georgia, and North Carolina. The major mineral constituents of the ores were muscovite, quartz, kaolin, and minerals of the feldspar group. The ores also contained minor quantities of biotite, limonite, and tourmaline. A comparison of the specific gravities of the major constituents of the ores (muscovite 2.76 +, quartz 2.65, kaolin 2.61, and feldspar 2.57), indicates that an effective separation of muscovite from the other minerals can be made in heavy liquids at a specific

Jan 1, 1968

By E. O. Lilge, H. Siebert

A design of a multiple-bomb leaching unit particularly suited for the study of heterogeneous reactions, i.e. sol id-liquid-gas interfaces at high temperatures and pressures is presented. The component parts and overall design are described and illustrated in detail. The system has been tested by a limited number of experiments and found to work satisfactorily. Heterogeneous reactions at solid-liquid-gas interfaces, at high temperatures and pressures, are studied in either autoclaves1'4 or bombs5'9. Autoclaves are usually fairly large vessels which have a stirring arrangement incorporated in the design. Mechanical design limits the use of these vessels to relatively low temperatures and pressures. They are particularly well suited to rate or kinetic studies. Bombs can be designed to be operated at high temperatures and pressures, and are smaller vessels; because of this, stirring arrangements are not easily incorporated. In this paper a multiple-bomb system is described, that has been designed to incorporate the desirable features of both the autoclave and the bomb. The described unit has flexibility for an over all study of leaching. Most of the component units in this assemblage are standard equipment, designed and produced by well known manufacturers. No claim is made to the design of this equipment, and although described in some detail, this is done merely for clarity in the elucidation of the system presented. The general layout of the multiple-bomb assemblage is shown in Fig. 1A and 1B. The basic units in this system are: 1) 6 stainless steel 100 ml bombs (Fig. 4)

Jan 1, 1964

By Fred C. Bond, Jen-Tung Wang

Comminution energy is principally energy of deformation before breakage, which appears as heat An empirical equation is presented which covers the entire comminution range. The new strain-energy theory considers comminution from the known principles of mechanics and the reduction ratio. Energy requirements according to the different theories the romnnred. THE present status of the theory of comminution is extremely unsatisfactory. The amount of various ores and rock products crushed and ground annually approaches the staggering sum of perhaps one half billion tons. Yet the basic concepts underlying an operation of such magnitude are unknown, and actual knowledge of what takes place in comminution is almost entirely empirical. The responsibility for this situation cannot be charged to apathy or lack of appreciation of the importance of the problem. Many attempts have been made to evolve a workable hypothesis, but none has been completely successful. General Considerations In order to break rock, it must be subjected to a stress which strains the rock beyond its critical breaking point. The stress imparts energy to the rock, most of which is released in the form of heat when the stress and resulting strain are removed. If the strain induced exceeds the critical strain, the energy is released by breaking; if the induced strain does not reach the critical point, the energy is released with removal of the stress as energy of resilience. Rock is a brittle material and it is assumed that it breaks approximately at its yield point, so that no permanent deformation results from strains below the critical. When rock is broken the total energy input is accounted for by the heat liberated, and by the surface energy of the new surface produced. A small and probably negligible amount is released as noise. Rock is commonly broken under compression and the applied stress is ordinarily compressive. However, for breakage to occur, it is only necessary that the induced strain exceed the critical value, and this strain may represent the resultant of compressive, shearing, and tensile forces. The stress-strain diagram of rocks under compression, and the modulus of elasticity as determined therefrom, should be fundamental considerations in the development of any theories of rock breakage. Crushing in jaw or gyratory crushers results primarily from a squeezing action, and reduction in a hammer mill is the result of impact. The action in rod and ball mills is a combination of impacting, squeezing, and wearing away by attrition, or rubbing. In ordinary crushing and grinding, the forces are applied at protruding points and are not distributed evenly throughout the rock. This is one of the features of comminution which theoretical considerations have not covered adequately. Another feature is the effect of impact velocities, which may greatly influence the total energy required for breakage. Impact velocities may vary from 1 or 2 fps in crushers to 10 or 20 fps for ball mills, and as high as 100 or more for hammer mills. However, according to the findings of geophysicists, the velocity of sound and of compression waves in stone is many times greater; the primary longitudinal compression waves travel at perhaps 15,000 fps, and the secondary transverse waves, which may reach the surface and cause cracks to form thereon, travel at perhaps 5000 fps. When a crack tip forms, the total surrounding stresses are concentrated in this tip, which rapidly extends throughout the rock particle.1 It would seem that the energy required to deform the rock beyond its critical strain, resulting in the formation of a crack tip, represents practically all of the energy required. Most of the primary crack tips presumably form on the surface. A simple method of designating feed and product size is necessary for the evaluation of crushing and grinding results. Taggart's suggestion2 that the 80 pet passing size be used has been found very practical. In this paper, the square opening screen size which 80 pct of the feed passes is designated as the feed size F, and the size which 80 pct of the product passes, or the product size, is designated as P. The reduction ratio at the 80 pet passing size, or F/P is designated as n. The feed and prod-

Jan 1, 1951

By Fred C. Bond, Jen-Tung Wang

Comminution energy is principally energy of deformation before breakage, which appears as heat An empirical equation is presented which covers the entire comminution range. The new strain-energy theory considers comminution from the known principles of mechanics and the reduction ratio. Energy requirements according to the different theories the romnnred. THE present status of the theory of comminution is extremely unsatisfactory. The amount of various ores and rock products crushed and ground annually approaches the staggering sum of perhaps one half billion tons. Yet the basic concepts underlying an operation of such magnitude are unknown, and actual knowledge of what takes place in comminution is almost entirely empirical. The responsibility for this situation cannot be charged to apathy or lack of appreciation of the importance of the problem. Many attempts have been made to evolve a workable hypothesis, but none has been completely successful. General Considerations In order to break rock, it must be subjected to a stress which strains the rock beyond its critical breaking point. The stress imparts energy to the rock, most of which is released in the form of heat when the stress and resulting strain are removed. If the strain induced exceeds the critical strain, the energy is released by breaking; if the induced strain does not reach the critical point, the energy is released with removal of the stress as energy of resilience. Rock is a brittle material and it is assumed that it breaks approximately at its yield point, so that no permanent deformation results from strains below the critical. When rock is broken the total energy input is accounted for by the heat liberated, and by the surface energy of the new surface produced. A small and probably negligible amount is released as noise. Rock is commonly broken under compression and the applied stress is ordinarily compressive. However, for breakage to occur, it is only necessary that the induced strain exceed the critical value, and this strain may represent the resultant of compressive, shearing, and tensile forces. The stress-strain diagram of rocks under compression, and the modulus of elasticity as determined therefrom, should be fundamental considerations in the development of any theories of rock breakage. Crushing in jaw or gyratory crushers results primarily from a squeezing action, and reduction in a hammer mill is the result of impact. The action in rod and ball mills is a combination of impacting, squeezing, and wearing away by attrition, or rubbing. In ordinary crushing and grinding, the forces are applied at protruding points and are not distributed evenly throughout the rock. This is one of the features of comminution which theoretical considerations have not covered adequately. Another feature is the effect of impact velocities, which may greatly influence the total energy required for breakage. Impact velocities may vary from 1 or 2 fps in crushers to 10 or 20 fps for ball mills, and as high as 100 or more for hammer mills. However, according to the findings of geophysicists, the velocity of sound and of compression waves in stone is many times greater; the primary longitudinal compression waves travel at perhaps 15,000 fps, and the secondary transverse waves, which may reach the surface and cause cracks to form thereon, travel at perhaps 5000 fps. When a crack tip forms, the total surrounding stresses are concentrated in this tip, which rapidly extends throughout the rock particle.1 It would seem that the energy required to deform the rock beyond its critical strain, resulting in the formation of a crack tip, represents practically all of the energy required. Most of the primary crack tips presumably form on the surface. A simple method of designating feed and product size is necessary for the evaluation of crushing and grinding results. Taggart's suggestion2 that the 80 pet passing size be used has been found very practical. In this paper, the square opening screen size which 80 pct of the feed passes is designated as the feed size F, and the size which 80 pct of the product passes, or the product size, is designated as P. The reduction ratio at the 80 pet passing size, or F/P is designated as n. The feed and prod-

Jan 1, 1951

By D. L. Harris, A. M. Gaudin

Observations were made of the distribution of mercaptan containing S35 between aqueous solution and mineral and between aqueous solution and the gaseous phase. Although equilibrium may not have been attained, adsorption of the reagent was shown to occur reasily from air or aqueous solution on sphalerite, zincite, and willem-ite and to correspond to flotation. Adsorption on quartz did not similarly occur. THE following results, presented here in condensed form,' were obtained in a preliminary study of the adsorption of n-hexane thiol, hexyl mercaptan, on sphalerite, zincite, willemite, and quartz, from aqueous solution and from a gas. Interest in this subject was aroused by a Belgian report' of effective use of hexyl mercaptan for flotation collection of oxidized zinc minerals. The relatively low boiling point, 149°C, of the mercaptan3 suggested the desirability of extending the usual measurements of partition of collector between aqueous solution and gas and between gas and mineral. It is believed that this paper presents the first measurements of this type on a flotation system. Attempts were made to carry out the measurements at equilibrium, but as the work progressed it became increasingly doubtful that this desirable condition had been achieved. To control composition and extent of the gas phase, the apparatus was a wholly-enclosed thermally-controlled glass system. Because of these constraints and the desirability of dealing with pure minerals, a scale of operations was chosen in which a few grams of deslimed mineral were used in each test. It was also necessary to choose a particularly sensitive method for mercaptan analysis, and in fact a method that would permit the experimenters to follow the approach to equilibrium. For these reasons mercaptan marked by radiosulphur 35 was used. An analysis was made for the radiosulphur by a modification of the method of Gaudin and Carr. Coarsely-crystallized sphalerite was handpicked, stage-crushed in the dry state, wet-screened on a 200-mesh sieve, and deslimed in water at about 5 microns. Further treatment consisted of a wash in dilute aqueous hydrogen peroxide, drying, removal of the dark-colored fraction in a Frantz magnetic separator, washing in very dilute hydrochloric acid, repeated washing in distilled and conductivity water, and drying. The last washings showed a conductivity equivalent to a few ppm NaC1, that is, much more than would be provided, theoretically, by a saturated ZnS solution. The material was stored dry in sealed bottles. Analyses were as follows: Zn, 62.3 pct; Fe, 0.43 pct; Cd, 0.44 pct; S, 31.2 pct; Mn, 0.001 pct. The specific surface (BET method) was 2000 cm2/g. Zincite from Franklin furnace of the New Jersey Zinc Co. was hand-picked, dry-crushed, wet-screened at 100 mesh, and deslimed at about 10 microns. After drying, the associated zinc, manganese, calcium, and silicate minerals were removed in a Frantz magnetic separator. The purified zincite was washed in distilled water and conductivity water to a conductance of less than 2 ppm equivalent NaC1, dried, and stored. Analyses were as follows: Zn, 75.1 pct; Fe, 0.9 pct; Mn, 2.78 pct. The specific surface (BET method) was 1740 cm 2/g. Willemite, also from Franklin furnace, was purified similarly. Analyses were as follows: Zn, 52.5 pct; Fe, 0.12 pct; SiO², 27.3 pct; loss on ignition, 0.13 pct. The specific surface was 1760 cm 2/g. Conductivity water (double-distilled) and demin-eralized-distilled water were used in most of the tests. The specific resistance was not less than 600, 000 ohms, and usually above 1,000,000. Radiosulphur-marked hexyl mercaptan (1-hexane thiol) was synthesized by Tracerlab, Inc., Boston. Two lots were secured several months apart. The last lot, consisting of about 0.5 g of the mercaptan, had a total activity of about 10 millicuries. Tracerlab Co. guaranteed only the activity; hence a quasi -vapor pressure determination (based upon an S analysis) of the mercaptan was made. The calculated value, 4.2 mm of mercury at 25.5' C, has been compared with that of a sample of Highest Purity 1-hexane thiol from Fisher Scientific Co. The latter had a vapor pressure of 4.5 mm of mercury at 2.5 C. Analytical Procedures The sample containing radiosulphur-marked mercaptan was oxidized to convert the mercaptan sulphur to sulphate, carrier barium sulphate being added to provide a suitable quantity of total barium sulphate in a filter cake. The precipitate was filtered and dried, and counting was carried out either in a streaming-gas (Q-gas) counter for high sensitivity or with an end-window G-M counter for convenience. The oxidized and precipitated mercaptan gave a radioactive count of 65 counts per minute per microgram in the end-window Geiger-Mueller counter and 1100 counts per minute per microgram in a Q-gas counter. For standardization of the mercaptan solution, 15 replicate analyses were made. The average deviation per measurement was about 1600 cpm in 65,000 cpm, the probable error in the mean being 275 cpm. It

Jan 1, 1955

By S. Usui, I. Iwasaki

The adsorption mechanism of dodecylammonium acetate (DAA) on mercury in potassium fluoride solutions at natural, near neutral pH was investigated. Difler-ential capacity combined with electrocapillary measurements on a dropping mercury electrode were found to provide a sensitive and reliable approach. A shift in the electro capillary maximum (or the point-of-zero-charge) towards positive polarization of the surface due to the adsorption of amine indicated a specific affinity of dodecylammonium ion for mercury surfaces. The amine was shown to adsorb competitively with potassium ion on mercury. Through correlation of the adsorption density of the amine, calculated from the thermodynamic analyses of the electrocapillary curves, and of the diflerential capacity curves, the adsorbed state and the orientation of the amine at the interface were considered. The similarities and the differences in the adsorption behavior of the amine on mercury and on such ionic solids as quartz, silver sulfide, and silver iodide were evaluated. With the promise of flotation providing a cheap and large-scale process for the upgrading of low-grade ores and magnetic concentrates, the iron ore industry has made a sustained search over the years for a flotation reagent of general applicability and high selectivity. Having shelves full of commercial reagents already on hand and anticipating more, both the users and the suppliers are in need of an experimental approach sensitive enough to differentiate the individual characteristics in each group of reagents. Therefore, the adsorption behavior of various flotation reagents on mineral surfaces from aqueous solutions has been a subject of much investigation. The current view of adsorption based on the electrical double-layer theory is formulated on the assumption that certain ions in solution act as potential-determining ions. Their concentrations fix the potential drop across the double layer and their adsorptions govern the surface charge.' Two methods used to study the adsorption

Jan 1, 1971

By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958

By W. F. Stowasser

downdraft traveling grate process to agglomerate pelletized iron ore concentrates has been successfully demonstrated in a pilot plant at Carrollville, Wis. Work there followed several years of development in the Allis-Chalmers Mfg. Co. laboratories, and the pilot plant phase was carried out in cooperation with Arthur G. McKee & Co., consultants and engineers to the iron and steel industry. End result of the process is conversion of iron ore concentrates into a form which can easily be transported and smelted in the blast furnace. Process Description The first of two process steps incorporates the art of balling and prepares the concentrates for burning. The second step consists of burning the green balls on the grate machine to the hardness required for shipping and handling purposes and for reduction in blast furnaces, see Fig. 1. Facilities are provided at the pilot plant to receive carload quantities of concentrate. The concentrates are loaded into a 50-ton bin direct from railroad cars. Because of the variable moisture content of the concentrates after shipment in an open railroad car it is necessary to repulp and refilter the concentrates to maintain a uniform and proper moisture content for the balling operation. Concentrates are conveyed to slurry tanks, and the slurry, at 50 to 60 pct solids, is pumped to a 4x4-ft drum filter. The filter provides feed of uniform moisture to the plant. Magnetite concentrates are normally filtered to produce a cake containing about 10 pct moisture, a necessary requirement for the following balling operation. The filtered concentrate is conveyed to a rotary bin table feeder which acts as a surge bin for the filter cake and delivers a steady flow of concentrates to the balling drum. It is often desirable to make additions to the concentrates as they are fed to the balling drum. These additives, such as bentonite, increase the strength of the finished green pellet and improve ballability of the concentrate. A vibrating feeder supplies additive to the feed belt, and the additive is mixed with the concentrate in the balling drum. The balling drum, shown in Fig. 3, is 8x3-ft diam. An oscillating cutting bar maintains the lining in the drum by trimming off the buildup of excess concentrate as it forms. The drum is operated in closed circuit with a lx4-ft rod-deck vibrating screen. Undersize pellets or seed pellets from the screen are returned to the balling drum until they grow to the desired size. Size of pellets is controlled by the opening in the screen deck. The formation of pellets in the balling drum is affected by many variables. Some of these are: the size distribution of the feed, the particle shape of the concentrate, the feed rate to the drum, the moisture in the concentrate, the speed of rotation of the drum, the slope of the drum, and the type of trimming obtained with the cutting bar. In this process, attempts are made to control the pellet size within the limits of % to 5/8 in. diam. The screened oversize pellets are conveyed under a coal feeder where sufficient powdered coal is added to the belt to produce desired results in the burning process. The top size of the coal successfully used has been 20 mesh, and anthracite was used in the test program. Fig. 4 illustrates the vibrating screen and the coal feeder. The pellets and free coal are conveyed together to the 5x3-ft diam- reroll drum that rolls the coal onto the surface of the pellets. This drum is also equipped with a cutting bar. The prepared pellets, containing bentonite, water, and surface coal, are elevated to the traveling grate, which consists of a continuous strand of 37 pallets. Each pallet, with a grate bar area 2 ft wide by 1 1/2 ft long, has 14-in. high side plates, Fig. 5. Feeding and distribution of the green balls to the grate is handled by a short conveyor which oscillates back and forth across the 2-ft width of the grate. An adjustable vertical plate located several inches in front of the head pulley of the oscillating conveyor controls the height of the bed and levels the moving bed of pellets. This method of feeding prevents segregation of various size pellets as well as fines and produces a uniform, permeable bed. The pallet train moves under the furnace and across four windboxes, located beneath the pallet frames, see Fig. 2. As the green pellets are deposited on the grate, partial drying of the pellets begins over a 2-ft long updraft windbox. The low temperature air reduces the moisture in the pellets in the lower level of the bed and this operation is essential to prevent sagging of the bed during later stages of the Process. The air used for this drying is recuperated from cooling the pellets on the grate, and supplemental heat, required for starting the Process, is obtained from an auxiliary burner. The pellets are then moved by the grate into the furnace and over an 8-ft windbox, designated as the downdraft waste windbox. Products of combustion are exhausted from this windbox to atmosphere. The furnace, shown in Fig. 6, is constructed with three chambers to provide downdraft drying, preheating, and ignition, respectively, to the pellet bed as it passes through. Overall length of the furnace is 5.57 ft; however, the exterior wall ends may be moved to reduce the length and also adjusted to Obtain the bed height desired, The drying, preheating, and ignition sections of the furnace are supplied with medium temperature

Jan 1, 1956

By W. F. Stowasser

Robert E. Hagen (Oliver Iron Mining Div., U. S. Steel Corp.)—Mr. Stowasser and his associates are to be congratulated on their achievements in developing the design of a pilot agglomerating plant, based on information obtained from experiments made in a laboratory sinter pot, and subsequently constructing and successfully operating this plant. In his Transactions article, May 1955 (vol. 201) Mr. Stowasser stated that the output of his heat treating strand was about 2 Lt per sq ft of grate area per day and that the fuel consumption was 0.7 to 0.8 million Btu per Lt of product. If it is assumed that the burning zone extends from the back of the ignition furnace to the beginning of the recuperation windbox, only about 1/3 of the grate area over the windboxes is utilized for burning pellets, and the actual rate at which partially dried pellets are burned is about 6 Lt per sq ft of grate area per day. It may be interesting to compare the above operating data with that obtained at the Extaca plant, where taconite concentrate is treated on a conventional sin- tering machine using limestone or hydrated lime additives with no recuperation of heat. The entire windbox area is used for burning, and a production rate of 2.5 Lt per sq ft of windbox area per day has been attained at a fuel consumption of 1.8 to 2.0 million Btu per Lt of product, the latter being considerably in excess of that reported by Mr. Stowasser, partly because a circulating load of returns amounting to 28 to 30 pct of the machine feed is used and because no attempt is made to recuperate heat from the waste gases. With the high capital and labor costs prevailing today, it seems that future development of the downdraft grate pelletizing process should be directed toward taking advantage of the high burning rate of predried pellets and increasing capacity, perhaps at the expense of fuel economy, by decreasing the proportion of grate area utilized by ancillary functions. For example, eliminating the cooling zone and using this area for burning would increase the proportion of grate area available for burning from 33 to 50 pct of the total windbox area, and a proportionate increase in capacity would result.

Jan 1, 1956