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Secondary minerals are the result of a process of concentration and enrichment and are commonly richer than the primary minerals of the same deposit. Secondary ores that contain abundant sulphides are commonly distinguishable in the field from primary ores; this distinction is more difficult with ores that carry a small quantity only of sulphide minerals, and microscopical examination of the ore in thin section is advisable in all doubtful cases. It is frequently difficult, and in some cases impossible, to determine the primary or secondary character of quartzose gold ores. The safest guide is the presence or absence of gaseous or fluid inclusions in association with the ore minerals; such primary association coupled with a lack of traces of oxidation is indicative of primary origin. Any ore that carries traces of oxidation, such as iron stains, endritic oxides of manganese, kaolin and so forth must be considered to have been acted upon by surface agencies, and secondary enrichment must be suspected. Further examination, however, may indicate that the contained value is primary, and, therefore, persistent in its vertical distribution. The manner of occurrence of secondary minerals is frequently characteristic; secondary minerals, being due to the action of surface waters, are commonly found in cracks through earlier minerals, or as crusts lining vugs or surrounding nucleal masses, or in general, in some distribution later than and not conforming to the arrangement of the primary minerals. Secondary sulphides are often found as soft black pulverulent masses or coatings, although they frequently occur well crystallized or as

Jan 1, 1932

CONTENTS PAGE WARNER, ROBERT K.-Efficiency of Screening. Discussed by Robert C. Canby,Robert K. Warner, W. 0. Borcherdt, V. E. Flanagan 1 FAHRENWALD, A. W.-Surface Reactions in Flotation. Discussed by Arthur F. Taggart, A. M. Gaudin 7 Efficiency of Screening Discussion of the paper of ROBERT K. WARNER, presented at the New York Meeting, February, 1924, and issued, as Paper No. 1293-M, with MINING AND METALLURGY, January, 1924. ROBERT C. CANBY, Wallingford, Conn.-On page 2, the author speaks of the importance of using a testing sieve clothed with the same material as the screen being studied; does he regard this as essential? In practice, how does the screen efficiency calculated by the difficult-grain method compare with those computed by the usual undersize method? Is it possible by the author's method, to obtain a measure of efficiency that can safely be compared with that of another screen sizing a different ore? ROBERT K. WARNER.-The use of a testing sieve clothed like the mill screen being studied is desirable but not essential for routine work about a mill where the character of the feed varies little, as regards the shape of the grains and the sizing test. Close agreement in size of aperture between the sieve and the screen is, however, important if the data obtained are to be valuable to others working on a different feed. If the testing sieve is coarser than the mill screen, the efficiency will appear unfairly low; whereas if it is too fine one might get efficiencies greater than 100 per cent. because of the apparent oversize in the undersize. In either of these cases the usefulness of the difficult-grain method of calculation would be reduced. As screens wear, the undersize grows appreciably coarser so that it is desirable to test screens only when the cloth is new or else to provide testing sieves that are worn about the same amount. The futility of making efficiency observations on a. distorted or broken screen surface is obvious.

Jan 6, 1924

By Roy D. McLellan

Electrolytic production of cadmium at the Great Falls plant started in the first part of the year 1925. Prior to that time, an experimental unit had been in operation for a few months during the year 1922. At present the plant is the largest producer of cadmium, not only in the United States but in the world. Electrolytic cadmium is characterized by its very low content of impurities. The main impurities, which are lead, copper, zinc and iron, do not aggregate more than 0.05 per cent., leaving a cadmium content better than 99.95 per cent. The consumption of cadmium has increased rapidly since the war, with the growing use in the plating industry. Cadmium plating has come into prominence because of its connection with the automobile and radio industries. As a protective coating, a deposit of cadmium is superior to zinc. Cadmium is also used extensively in alloys. Important among these is the alloy with copper used in telephone and trolley wire. In proportion of 0.5 to 1.2 per cent., cadmium adds materially to the strength and wearing qualities. Practically all commercial zinc concentrates contain a small amount of cadmium. In the electrolytic zinc process, the cadmium that is dissolved in the regular leaching operation is removed from solution with metallic zinc, generally added in the form of zinc dust. This precipitate also contains the copper taken into solution and the excess of zinc dust, which is required in order to insure a complete removal of all impurities. The precipitate is treated by a separate process to recover the copper, cadmium and excess zinc. In order to make the metals more soluble and to render insoluble such impurities as iron, arsenic and antimony, the precipitate is roasted. The roasting is carried out in gas-fired McDougall furnaces, with a roasting temperature of about 700' C. The roasted residue is leached with dilute sulfuric acid, spent electrolyte from the zinc electrolyzing, containing 10 to 12 per cent. free H2S04. The leaching is done in cylindrical tanks of 30 tons capacity, agitated

Jan 1, 1930

By M. Kaitz

A continuing discussion in both the petroleum engineering and economic literature is directed to the difficulties encountered in the use of discounted cash flow rate of return (DCF) as a measure of investment worth. Although useful in most Instances, DCF has been criticized because it is time-consuming in its trial-and-error solution, theoretically invalid, not adaptable to cash flow streams yielding multiple return solutions and not entirely reliable in selecting between mutually exclusive investments. The theoretical invalidity of DCF stems from the reinvestment assumption implicit in the calculation that earnings are reinvested at the DCF rate. The emerging consensus of the economic literature is that the net present value or the present worth of the net cash flow stream (discounting at the average opportunity or cost of capital rate) is more correct and reliable. Other criteria proposed have been ratios of net present value divided by initial investment or by present value of all investments in a project. All of these criteria are simple to determine and explicitly assume a reinvestment rate for [he income generated by a project. This paper develops and discusses a theoretically valid profitability criterion which is simple to compute and retains the appeal of a percent return on investment. It is called "percentage gain on investment" or PGI. It measures the gain an investment is expected to realize over like capital invested in the average opportunity and explicitly considers reinvestment potential. Why add another Concept to the large array of investment criteria now available, any one of which, or perhaps a combination of several, appears to embrace the form's (or individuals) objectives? The answer is that not one of the existing criteria provides both a readily comprehensible and theoretically valid measure of risk coverage that has general application. The proposed PGI does fulfill these requirements. INTRODUCTION An ancient expression warns that "one must yield to the times" — there are better ways of doing things. A review of the petroleum engineering and economic literature on one topic alone, measurement of investment worth, certainly is witness to this truth. In use for a number of years, the DCF has recently received attention, directed mainly to its theoretical invalidity. Several alternatives to DCF have been proposed to provide a valid, simply determined criterion to describe investment worth and to overcome the criticisms previously mentioned. This paper introduces another method called percentage gain on investment (PGI) and is proposed as but one of several yardsticks that should be used in making investment decisions. MEASURES OF INVESTMENT WORTH This paper will consider only those criteria which give weight to the time pattern of future earnings. These criteria are usually compared with an average opportunity rate or cost of capital of the firm to judge the relative worth of the investment. For purposes of demonstration, a 9 percent average opportunity rate will be used throughout this paper. Implicit in the DCF calculation is the assumption that earnings are reinvested at the DCF rate. Some argue, though, that there is no reinvestment assumption,' that the DCF rate is simply that maximum rate of interest one can pay on the investment over the life of the project and break even. The determination of DCF is accomplished by discounting the net cash flow stream at that rate (DCF) which will yield zero. The question is: why should reinvestment potential be explicitly considered in calculating return on investment or other criteria measuring economic worth? Perhaps the answer lies in consistent or equal treatment of future cash flow. It appears entirely illogical to give different present worth value to $1 received, say, 10 years from now, which is the circumstance resulting in comparing projects with different DCF returns. In fact, $1 received in 10 years has the same value regardless of which project generated the income. DCF return thus favors investment projects which are expected to provide early income as compared to those providing long-term income. Not surprisingly, the controversy on reinvestment assumption is an old one. Hoskold' discussed this same problem in 1877. He considered the future income from mineral properties as an annuity or a series of fixed future payments. (As will be demonstrated, his equation can be modified for variable income.) Prior to Hoskold, the value or what one could pay for the mine, was determined with the use of standard, single-interest tables. Here is what Hoskold said with regard to these tables. "This table, and others of its kind, to be found in most works on annuities, is constructed correctly according to

By R. S. Kinasevich, D. D. Crabtree, D. W. Fuerstenau, T. P. Meloy, A. L. Mular

This paper presents details of the concept that size reduction in comminution machines takes place by three mechanisms; namely impact, abrasion, and chipping grinding. Experimental evidence is presented to show that each of these comminution mechanisms produces particles of a characteristic size distribution. The actual product from a comminution device has been postulated to be the result of a composite of these three types of mechanisms. Composite curves are demonstrated for a material following the Gaudin-Schuhmann, Gaudin-Meloy, and Rosin-Rammler size distributions. About forty years ago, Gaudin1 showed that grinding may be considered a combination of impact and attrition. Since that time, major efforts have been devoted to studies, both experimental and theoretical, of impact and compression fracture, while attrition has been almost wholly neglected, at least quantitatively. A significant attempt to distinguish between attrition and impact grinding is the work of Tanaka2 in Japan. By means of an instrumented abrasion grinder and a drop weight device, he determined coefficients relating the increase in surface area of various materials to the energy expended by the two methods of comminution. He showed that physical properties of a material determine its response to attrition and impact grinding. By extending his ideas to ball milling, he concluded that to a large extent the basic comminution process in a ball mill is due to the impact of falling media. Because Tanaka used a mass of particles in his drop weight device, however, he did not obtain pure impact size reduction. The purpose of this paper is to define basic mechanisms of size reduction, to verify that each has a characteristic and meaningful distribution curve, and to show how a combination of these mechanisms results in a size distribution very like that produced by grinding mills. DEFINITION OF COMMINUTION MECHANISMS The product from a grinding device will be considered to be the sum of the contributions resulting from the action of three basic mechanisms of size reduction which operate in a ball and rod mill. These are impact grinding and two forms of attrition, namely abrasion grinding and chipping grinding. Each is defined below. Impact Grinding: If a straight line is obtained when the cumulative weight fraction finer than a given size is plotted on log-log paper as a function of that size, the slope of the line is defined as the distribution modulus. A number of investigators 3,4,5 have shown that impact fracture of single specimens of different composition and shape results in a distribution of daughter particles that exhibit a distribution modulus of unity. This has also been true for materials broken in slow compression. Moreover, under slow compression or impact fracture conditions, theoretical distribution equations6*' which approximate the Gaudin-Schuhmann distribution when the distribution modulus is close to unity agree reasonably well with the experimentally obtained products. If one now considers a comminution process where many impact or compression fracture events are taking place, then any rational summation of the individual events yields, in the finer size range, a Gaudin-Schuhmann distribution whose distribution modulus is also unity. If the size distribution of the product from a mill has a distribution modulus of unity, it is implied that particles are fractured in a manner which results in a comminution event having a distribution like that obtained for the breakage of single specimens. These impact events would occur when a particle is smashed between balls or between a ball and the mill liner. Attrition, on the other hand, is that form of grinding which cannot be called impact grinding, and comprises both abrasion and chipping grinding. Abrasion Grinding: Abrasion is a form of surface wear resulting when particles rub against each other or against balls or the mill liner of a tumbling mill, somewhat like a piece of wood being rubbed over a piece of sandpaper. The surface wears at a more or

Jan 1, 1964

By David B. Brooks, William A. Wallace, James R. Dunn

As the conflict between the mineral industry and preservationists steadily increases, it becomes urgent to determine as precisely as possible the costs of developing vs. not developing our domestic mineral resources so that enlightened mineral policy decisions can be made. A simple discounted cash flow equation is suggested as an approach to assigning a public-interest dollar value to mineral resources. Present value of a piece of land underlain by a mineral resource is taken as equal to the discounted cash flow derived from the resource corrected for present and future surface land values. The equation is effective as a base for mineral aggregate resource planning for metropolitan areas if transportation cost differences are substituted into the cash flow component. Suitably modified, the equation can be fitted readily into an existing computer simulation program for aggregate mineral resource planning in metropolitan areas. However, because of the lack of accurate input, the equation is less effective for other mineral commodities such as base metals and coal. Nevertheless, within a somewhat philosophical context, the equation still has utility and is effective in defining the problem of public interest values for minerals other than mineral aggregates.

Jan 1, 1972

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 Dae S. Young

A new probabilistic approach is introduced for slope stability analysis including adverse impact of blasting vibrations. The material strength within a slope area is randomized in this bi-variate joint probability analysis. To be a completely general engineering method, the new probabilistic approach employs the random variable transformation technique: the Hermite model of the Gaussian transformation function, which transforms the experimental histogram of strength values to the standard Gaussian variables. Because bi- normal joint probability is analyzed on the true probability region projected on the Gaussian transformed variable plane, it is an exact solution of slope stability based on the available sample data. Details of the probabilistic approach are given through a case study on an iron ore mine. The case study demonstrated that the probabilistic approach is effective and simple enough to be a routine engineering practice for slope analysis.

Jan 1, 1984

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By Millard K. Shaler, Sydney H. Ball

During the past 50 years the attention of mining-men has been turned to Africa, and within the past decade prospecting-expeditions sent into Central Africa have resulted in the open-ing-up of several mines. Having recently spent two years in the Belgian Congo for the Societe Internationale Forestiere et Miniere du Congo, it is our purpose to outline briefly the mining-conditions in that part of Central Africa. Before it became, in 1908, a Belgian colony, this region was known as the Congo Free State. 11. Location and Area. The Belgian Congo lies in southwestern Central Africa, west of the continental divide, as illustrated in Fig. 1. It has a coastline of but 20 miles on the Atlantic, but widens rapidly eastward to points 5 north and 14 south of the equator. Its area is 908,000 sq. miles, which is almost one-quarter that of the United States and Alaska, or more than one and one-half times that of Alaska. Stated in other terms, its area is nearly one-third that of the United States, exclusive of Alaska and the insular possessions, or more than that of all the United States east of the Mississippi river. III. Topography and Drainage. From north to south, Africa is made up of three topographic elements: 1, the Atlas mountains, a highly-accentuated region

Jan 1, 1911

By V. A. Josendal, J. W. Wilson, B. B. Sandiford

Two radioactive tracers have been tested as a means of determining core saturation in multiphase flow studies. Cesium chloride was tried as a water-phase tracer, but complications in its use in low permeability cores resulted from sorption of cesium by the core or water-wet pads. IodoI3'ben-zene proved very satisfactory as an oil-phase tracer. The synthesis of iodobenzene from the sodium iodide as received from Oak Ridge is simple and direct. The tracer is insoluble in water and there was no evidence of sorption by any of the core materials used. use of the method to determine saturation profiles during capillary and dynamic desaturations and relative permeability measurements on oil-water and oil-gas systems is described. Comparisons of the dynamic and capillary methods of relative permeability determination were made using the tracer to check core saturation and saturation distribution. Other experiments are also described in which mobility of the oil phase at various saturations was measured by displacing labeled oil by flowing inactive oil. Similar experiments were made using water labeled with cesium 134. INTRODUCTION To be fundamentally sound. any laboratory method of determining relative permeability must meet, among others, the following requirements: (1) the core saturation between the pressure taps must be uniform, and (2) there must be uniform pressure difference between the phases in the region between the pressure taps. In practice it may eventually be shown that appreciable departure from either or both of these conditions may be permitted without introducing sensible error. but the magnitude of such error remains to be established. There is need then for methods of saturation determina- tion which will determine the saturation profile in a core as well as the overall average saturation given by the gravimetric or material balance methods usually employed. Methods which have received consideration include resistivity, X-ray absorption, gamma ray absorption, neutron diffraction, and radioactive tracers. The resistivity method has use in the determination of profile, but needs checking by independent means.' The X-ray method has been used successfully, but requires elaborate equipment and calibration.23 4 At present gamma ray absorption is practical only for very large cores. The neutron methodo does not appear adaptable to measurement of saturation variations in short cores since it is a scattering method. Russell, Morgan and Muskat' employed radiovanadium in a study of the mobility of interstitial water. Coomber and Tiratsook used radioiodine as an oil-phase tracer and measured profiles in unconsolidated sand packs. ADVANTAGES OF RADIOACTIVITY METHOD A thorough discussion of the theory of radioactive tracers and the technique of using them to study the movement of fluids in sands is given by Coomber and Tiratsoos and need not be detailed here. The method requires less equipment than the X-ray method; the counting problem is similar, but the generation and regulation of the radiation, which is a major problem in the X-ray method, is not required in the radioactivity method. In the X-ray method it is customary to add up to 20 per cent of absorber, usually an iodine compound, to the phase to be labeled. while in the radioactivity method a mere trace is sufficient. A possible disadvantage of the radioactivity method is that a disproportionate amount of the activity comes from the portion of the core nearest the counter, whereas the X-ray gives a true average saturation in the region traversed by the beam. 011 the other hand, this characteristic of the radioactivity method renders it peculiarly adaptable to study of certain types of radial variations in saturation. As with the X-ray method, a tracer may be added to either the oil or water-phase. An oil-phase tracer is more generally useful. however. since it permits work on oil and gas in the

Jan 1, 1952

By A. Aly Selim, Bruce D. Hanson

A computer model capable of simulating the basic production system in a variety of underground mining methods has been developed at the U.S. Bureau of Mines Twin Cities Mining Research Center. The simulation of the basic production system at the working face, which includes drilling, blasting, mucking, and roof bolting, is described in this paper. The model is event-oriented; that is, the status of the system is updated at the occurrence of an event. All event or activity times, including failure and repair of equipment, are treated as stochastic variables. A location dimension, in addition to the time dimension, is used, allowing the simulation of systems with a number of stopes or faces. To aid the user in determining the efficiency of his system, the model provides several reports. An event or activity report gives the frequency of occurrence, total active time, and cumulative delay statistics for each event. An equipment report delineates the availability and utilization of all equipment. A production report provides information on various performance measures such as tons produced and holes drilled. The cost report gives the cost of various elements in the system. Operating data from an actual mining situation were used in validating the model. In addition to the description of the model, examples of simulation experiments which were performed on a room-and-pillar mine system are given. Use of the model to determine methods for improving the efficiency of the system is discussed.

Jan 1, 1976

By S. J. Green, R. D. Perkins

Little data exists on the high strain rate behavior of geologic materials. Uniaxial stress tests by Kumar1 and by Serdengecti and Boozer2 present some results to strain rates in the range 10 to 103 per sec. The two references cited are the only ones the authors found for high strain rates. Techniques for experiments at high strain rates are not well developed and, in general, leave questions with respect to interpretation of data when the material behaves in a brittle manner. The most reliable data at very high strain rates (104 to 106 per sec) are found in some recent flat-plate work3-5 where experimenters have impacted one plate into another, thereby creating a uniaxial strain loading. The standard uniaxial stress-strain curves cannot be obtained directly from these tests; however, yield (or fracture) may be studied by measuring the "elastic" wave magnitude. These data then may be compared to uniaxial stress data by resorting to some preassumed yield/fracture criteria. Even with the most liberal interpretation of existing data, little can be said of the effects of high strain rate on stiffness, yield/fracture, and post-fracture behavior of geologic materials. It is the purpose of this work to begin to add to the understanding. The data presented here are intended to provide material information for use in calculating the response of geologic materials to high rate deformations. These data can be applied directly to determine parameters such as the elastic modulus and Poisson ratio as a function of stress. Furthermore, these data can be used to derive an understanding of yield/fracture and the effect of rate on the yield/fracture surface. The work presented is a first step, and is

Jan 1, 1972

By C. W. Robinson

Transportation is the minerals business. Once upon a time the geologist, the engineer and later the metallurgist reigned supreme, but the leading role in mineral development today is the economist-especially the transportation economist. Obviously this trend is more important for large volume, low value mine-to-market movements of such commodities as iron ore, phosphate, potash, coal and petroleum. But, transportation is of sufficient importance to all mining activities to be included in forward planning. Civilization has been influenced by transportation developments to an extent little understood or appreciated. Although industrialists are capable of intelligent historical analysis of this phenomenon, too frequently they have failed to project, or anticipate those changes in transportation techniques which will influence the pattern of industrial growth.

Jan 1, 1971

By R. E. H. Pomeroy

Early in 1917, it became evident, owing to existing and pending market conditions, that a substitute for crude petroleum must be found for firing the smelter furnaces. After a review of the plants then existing, it was deemed advisable to depart from their practice and to adopt the following described system of distributing pulverized coal to the furnaces. The principal advantages of the proposed system, which influenced this decision, were: (I) Equal safety; no pulverized coal is stored at the furnaces; (2) greater ease of operation, furnace fires being controlled by regulating valves in burner branches, as in gas firing; (3) better organization; all machinery, including pulverized coal feed, is under one roof, and the organization is separate from the furnace department; (4) greater cleanliness; all machinery is under vacuum; (5) greater flexibility of application, the coal and air mixture can be piped where needcd. The design of the plant is the result of the combined efforts of the local engineering staff and of the machinery manufacturers. Many new features were embodied in this design to insure greater safety, cleanliness, and efficiency of handling. After 14 months of continuous operation, the plant has proved entirely satisfactory. The building is of structural steel, covered with corrugated steel, with many windows and is painted light gray inside and out. The building floor is of concrete with ample drains to the sewer, and the building is equipped with fire and service water piping. Fig. 1 shows the east side of the building and the incline conveyor from the raw-coal bins. The plant is operated entirely by electric power. All alternating current is 550 volts, 3-phase, 60 cycles. The pulverized-coal feeding niotors and control mechanism are operated by 220-volt direct current supplied by motor-generator sets in the building. Autamatic push-

Jan 1, 1921

By B. D. Cullity, A. Freda

It is well known that deformation by cold drawing or swaging produces a kind of preferred orientation called fiber texture in metal wires. Such textures have been extensively studied by means of X-ray diffraction, chiefly by techniques in which the diffracted X-rays are recorded photographically. This paper describes some results obtained with a new method, involving the use of a diffractometer. X-RAY METHOD Since the pole figure of a material having a fiber texture is symmetrical about the fiber axis, there is no need to present the entire pole figure. Instead, texture data can be summarized by means of a curve showing the variation of pole density along any meridian of the pole figure, if the north-south axis of the latter is made parallel to the fiber axis. The position of a pole on the meridian is specified by the angle 9 between the pole and the fiber axis (wire axis). The X-ray method1 used here to measure pole density as a function of 9 actually consists of two methods, referred to as A and B. In method a, X-rays are reflected from the wire surface. Pole densities can be measured by this method from O = 90 deg down to a limiting value which is dependent on the Bragg angle 8 and is typically of the order of 30 to 40 deg. Method B involves diffraction from the wire cross section and covers a range of 0 values from 0 up to an angle sufficient to overlap the measurements made by method A. By a combination of the two methods, a complete pole-density curve can be obtained. However, method A alone, which requires no specimen preparation other than etching, can often reveal enough of the pole-density curve to provide complete information about texture components. Both methods require that the measured diffracted intensities be corrected for changes in absorption and diffracting volume with changes in 9. A11 intensities I reported in this paper are corrected diffracted intensities; they are propor-

Jan 1, 1960

By John A. Church

IT is a well-known fact that under similar conditions a ton of pig iron can be made from any ore with less fuel when charcoal is used than when coke or anthracite is employed for heating. The cause of this superiority is entirely undiscovered, and no explanation has yet been offered which has won even temporary acceptance. Such as they are the hypotheses heretofore presented are of two general kinds, and they come from those two divisions of the working force which are so often antagonistic in all professions, the scientific and the practical men. The theory of the scientific men is drawn from the chemical reactions which are known to go on in the furnace, and it is well expressed in a recent paper by Prof. Akerman, of the Stockholm (Sweden) School of Mines. It makes the superiority of charcoal as a fuel in the blast furnace to depend upon its quick and thorough reduction of carbonic acid. This property belongs to charcoal in virtue of its porous structure. Its cells have a diameter of only one twenty-four hundredth of an inch, and one grain of charcoal has been computed to have a surface of two and four-tenths square feet. According to the theory in question, carbonic acid is formed by the first impact of the air upon the fuel; but in charcoal furnaces the acid is immediately reduced to oxide by taking the necessary amount of carbon into combination, and the gas reaches its highest reducing power even in the near vicinity of the tuyeres. Coke, being harder, does not produce so strong a reducing gas in that part of. the furnace. Carbonic acid is formed as in the case of charcoal, and remains longer before its reduction to oxide is completed. Anthracite is the

Jan 1, 1879

By M. C. Dow

INTRODUCTION BELT conveyors generally are conceded to be the most economical method yet devised for the transportation of large quantities of bulk materials within plants. Belts are coming into greater use for transportation over long distances and are replacing other forms of long-haul conveyance. It is not the intent of the writer to comment on the design of belt conveyors, but to offer some suggestions for improvement in the operation of conveying belts, based on many years experience with belts in several stonecrushing plants. It is the contention of this writer that opportunities for improvement not fully realized by most operators exist in many plants and that a study of the individual conveyors can result in lower unit belt costs. GENERAL RELIABILITY It is the fool-proof nature and general reliability of the belt conveyor that is responsible for its neglect. Even when abused, a belt will carry large tonnages and last a long time, often outlasting the personnel responsible for its care. Some operators look upon the purchase of a new belt as an expensive outlay every so many years, rather than a constant and steady cost amounting in belt wear alone to several dollars every operating day. This belt wear for all conveyors in one stone-crushing plant known by the writer costs 1 1/2 cents per ton or $12 to $15 per hour, based on the present cost of belt replacements. The fact that these replacement costs are double those of prewar quality emphasizes the necessity of improving the conditions that make belts wear out. Some of those conditions are the responsibility of the conveyor designer, and if a serious original error has been made, this handicap cannot be overcome by the operator. If the belt is too narrow, too steep, or too fast for good loading, the operator finds that he has an inherently expensive belt. The other factors that decrease the life of a belt are more or less under his control. The two factors of most importance are the loading conditions and the alignment. Most of the cover wear and carcass cutting occurs at the loading point and if the alignment is not good under all conditions, edge wear or damage may result, Fig I. LOADING THE BELT It is difficult to load a belt in an ideal manner; that is, to flow the material on the moving belt at a steady rate, without impact, without turbulence, and with a velocity equal to and in the same direction as the belt. An inclined belt is especially difficult to load and particularly when its speed is over 300 or 350 fpm. Inclined conveyors require longer skirt plates to prevent spillage until the load settles. It does not seem reasonable to expect an inclined conveyor to carry the same load as a horizontal belt of the same width and speed, because of the difficulty in loading

Jan 1, 1947

By James Norman, Benjamin S. Lindsey

Barite (BaSO4) is the most important industrial barium mineral from the standpoint of quantity consumed. In 1938 the amount was 365,000 tons. Its uses are numerous, some of the more important being in the production of pigments, as an inert filler in a wide variety of products, as a weighting material in oil-well drilling muds, in the manufacture of some binds of glass, and in the production of barium chemicals. The barite studied in this investigation occurs as the massive variety in a shale replacement formation near Magnet Cove, Ark. It is intimately mixed with quartz and small amounts of iron oxide and residual shale. Complete liberation of barite and quartz is not obtained when the rock is ground to I00 per cent through 325 mesh. The crude rock with which all test work was done assayed 85.1 per cent BaS04, II.II. per cent SiO2, and 2.85 per cent R2O3. Its specific gravity was 4.03. It was found in the course of the investigation that the barite contained a small amount of uniformly distributed carbonaceous material, so intimately dispersed that it could not be seen with the microscope, but gave a definite gray cast to the barite concentrates. The color of the barite in polished sections of the crude rock ranged from gray to almost black, while finely ground powder was a reddish gray. The reddish hue was due to the iron oxide present. In 1939, under the supervision of Oliver C. Ralston,* the Bureau of Mines, in cooperation with the Milwhite Company, of Houston, Texas, undertook to study the flotation of barite from Magnet Cove, Ark. Flotation of several southern barite ores had already been investigated by the Bureau of Mines.' The Magnet Cove barite was unique, however, in that it had to be ground to such an extremely small particle size for liberation. The flotation of non-metallic minerals in the presence of a large amount of slime is usually very difficult, and when done is an innovation calling for conditions different from those under which granular material is separated. Laboratory Flotation Work Preliminary investigation showed that the flotation reagents recommended by O'Meara and Coel for separating barite and siliceous material were suitable, but much larger quantities of reagents were required to separate such finely divided material. Moreover, it was found that careful control of the frothing was necessary. When oleic acid and sodium silicate were used a voluminous froth was observed at the beginning of a test, and toward the end there was not enough froth to effect complete recovery of barite. The addition of pine oil as a frother gave proper frothing for the duration of the test, and the substitution of coconut fatty acid (crude lauric acid) for oleic acid produced a more heavily burdened froth. Two different paper-mill fatty-acid products were tested

Jan 1, 1943

By Benjamin S. Lindsey, James Norman

Barite (BaSO4) is the most important industrial barium mineral from the standpoint of quantity consumed. In 1938 the amount was 365,000 tons. Its uses are numerous, some of the more important being in the production of pigments, as an inert filler in a wide variety of products, as a weighting material in oil-well drilling muds, in the manufacture of some binds of glass, and in the production of barium chemicals. The barite studied in this investigation occurs as the massive variety in a shale replacement formation near Magnet Cove, Ark. It is intimately mixed with quartz and small amounts of iron oxide and residual shale. Complete liberation of barite and quartz is not obtained when the rock is ground to I00 per cent through 325 mesh. The crude rock with which all test work was done assayed 85.1 per cent BaS04, II.II. per cent SiO2, and 2.85 per cent R2O3. Its specific gravity was 4.03. It was found in the course of the investigation that the barite contained a small amount of uniformly distributed carbonaceous material, so intimately dispersed that it could not be seen with the microscope, but gave a definite gray cast to the barite concentrates. The color of the barite in polished sections of the crude rock ranged from gray to almost black, while finely ground powder was a reddish gray. The reddish hue was due to the iron oxide present. In 1939, under the supervision of Oliver C. Ralston,* the Bureau of Mines, in cooperation with the Milwhite Company, of Houston, Texas, undertook to study the flotation of barite from Magnet Cove, Ark. Flotation of several southern barite ores had already been investigated by the Bureau of Mines.' The Magnet Cove barite was unique, however, in that it had to be ground to such an extremely small particle size for liberation. The flotation of non-metallic minerals in the presence of a large amount of slime is usually very difficult, and when done is an innovation calling for conditions different from those under which granular material is separated. Laboratory Flotation Work Preliminary investigation showed that the flotation reagents recommended by O'Meara and Coel for separating barite and siliceous material were suitable, but much larger quantities of reagents were required to separate such finely divided material. Moreover, it was found that careful control of the frothing was necessary. When oleic acid and sodium silicate were used a voluminous froth was observed at the beginning of a test, and toward the end there was not enough froth to effect complete recovery of barite. The addition of pine oil as a frother gave proper frothing for the duration of the test, and the substitution of coconut fatty acid (crude lauric acid) for oleic acid produced a more heavily burdened froth. Two different paper-mill fatty-acid products were tested

Jan 1, 1943