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By Howland Bancroft

This article is not presented as a treatise on tin mines and mining in Bolivia. It deals primarily with the tin situation, and but fragmentary information is given regarding individual properties, generalizations being necessary in so short an article. Data on general subjects related to tin mining in Bolivia are necessarily abbreviated and very incomplete. It is believed that the references to be found in this article will help those interested to a more thorough under-' standing of the general conditions extant in Bolivia. Bolivia produces from one-fifth to one-fourth of the world's supply of tin. The United States consumes about one-half of the world's supply and produces practically none. Tin ore is not reduced in the United States. These facts are not generally appreciated by American mining men. Many conflicting data have been published regarding tin in Bolivia. The Transactions of the Institute contain practically no mention of this subject.1 I have made a study of mines and mining conditions in the west coast republics of South America and have been particularly impressed by the possibilities of the Bolivian field.2 In consequence the thought seems justified that the present meeting would, present a fitting occasion to bring the tin situation in Bolivia to the attention of American mining men.

Jan 9, 1913

Joseph W. RichaRds, South Bethlehem, Pa. (communication to the Secretary*): The hold experiment of Mr. James Gayley in drying the blast used in the Isabella furnace has attracted the attention of the whole metallurgical world, and has given rise to a wide discussion of the reasons for its great success. Mr. Gayley and the capitalists who backed him in his work deserve congratulations on the success of the venture, not only because it succeeded, but because they have been rewarded far beyond the maximum that could be expected from theory or probability. To state the case in its simplest terms, all that could have been predicted was: (1) that the removal of the moisture from the blaet would result in the saving of the heat required for its decomposition, representing from 3 to 5 per cent. of the heat-value of the coke used; hence about that quantity of fuel might be saved; (2) that the furnace would be somewhat more independent of variations in the weather; (3) that the campaign would be a little hotter; and (4) that the speed of the blowing-eugines would be decreased, at least in amount proportionate to the reduction in temperature of the air led to them. I do not think anyone anticipated a much larger output per day. The results actually attained were: (1) the furnace made an increase of 24.86 per cent. in the output of pig-iron per day; (2) it consumed 19.6 per cent. less coke per ton of pig-iron made; (3) the blowing-engines were run 15.8 per cent. slower; and (4) the saving in power made up for the power required to run the refrigerating-plant. In addition, the blast was hotter; the waste-gases cooler, because of their smaller volume; and there was a greater proportion of carbon dioxide to carbon

Jan 1, 1906

By H. L. Overton, J. B. Lipson

The concept of sodium single-ion equivalent activity as developed by Gondouin, Tixier and Simard,' was used to determine the filtrate resistivity-activity relationships for 150 laboratory and 49 field drilling muds. With the exception of the native and phosphate high-resistivity muds and the calcium-treated muds, the filtrate resistivity-activity relationships were found to be within 20 per cent of that for pure sodium-chloride solutions. Gondouin, et al,* previously suggested a method of treating the high-resistivity muds in quantitative electrical log analysis. A new treatment is presented which uses the P-alkalinity of the filtrate as a correlating parameter in handling the calcium-treated muds in quantitative electrical log analysis. INTRODUCTION Gondouin, Tixier, and Simard in 1957 introduced new interpretive procedures for analyzing the self-potential curve of electric logging. They re-emphasized the previous work of the Schlumberger brothers, Wyllie and others, and re-defined the activity concept in terms of a number which they termed the effective resistivity. Their procedures are particuarly effective when the connate-water resistivities are less than 0.08 ohm-m or greater than 0.3 ohm-m. For gyp muds where the calcium content is known, a procedure was presented which allows calculation of the connate-water resistivity with improved accuracy. These procedures assume that one of the fluids in the system is a pure sodium-chloride solution, but they are not limited by the assumptions. For the fresh and saline connate waters, for instance, it is assumed that the drilling mud filtrate is a pure sodium-chloride solution. For the gyp mud interpretive procedure, it is assumed that the connate water is a pure sodium-chloride solution. Brown,' on the other hand, has suggested that it would be extremely fortuitous if drilling mud filtrates could all be treated electrochemically as sodium-chloride solutions, and he cites some examples where discrepancies exist. A study has been made of the deviations of sodium single-ion equivalent activities, an of drilling mud filtrates from that of pure sodium-chloride solutions. Drilling fluids in common use on the Texas and Louisiana Gulf Coast and laboratory-controlled muds were investigated. Using a cell patterned after Gondouin, et al, (Fig. I), the equivalent activities were measured. It should be remembered here that the activity measured is merely an index which uses numbers that represent the sodium single-ion equivalent activity a,., of the mud filtrate. The membrane potential developed across the cell was determined with a null-type potentiometer. The resistivities of the solutions were measured by the four-electrode method. The resistivity system was calibrated by using saturated KC1 solution as a standard. Chemical properties of the drilling fluids studied were measured with standard API-approved equipment for the particular property in question. For the activity measurements, two shale discs were chosen to serve as cationic membranes. These two discs were both 1 1/4 in. in diameter and, respectively, 1/4 and 1/8 in. in thickness. The membrane potential, Em, was determined using the shale discs as cationic selective membranes separating two standard sodium-chloride solutions of known mean activity. Then by comparing the constant Km in the following equation, with the theoretical value of — 59 mv at 75OF, the perfection of the membrane was determined. The variation of Km for the two membranes used in this experimental work, as well as one of the many unsuitable membranes, is shown in Fig. 2. The measurement of the activity of mud filtrate involves the balancing of its combined individual activities against a standard solution of known mean activity. At the point where there exists zero potential across the membrane cell, the activity effects may be assumed equal for all practical purposes. The mechanics of making a null reading of this sort requires having a large number of different mean activity solutions. A more practical method involves

By P. van Meurs

This paper describes a technique which permits visual observation of oil displacement processes througlrout the interior of a porous structure as thick as 2 in. A model having glass walls is filled with finely powdered glass. Such a model becomes completely transparent if the powdered glass pack is saturated with an oil having the same refractive index as the glass. When water or gar, with a refractive index different from that of the oil (glass) is injected, the model becomes opaque in the space occupied by the water (gar). visual observation and photography of these displacement processes are thus rendered possible. The following processes have been studied: (I) displacement of oil by water (linear drive) in a homo-geneous formation and in a stratified formation, (2) water flooding (five-spot well pattern), (3) solution gas-drive process, and (4) oil-bank formation. INTRODUCTION Most of the scaled model experiments heretofore carried out in the Koninklijke/Shell Laboratorium, Amsterdam, have been done in steel containers, in which oil-saturated sand packs represented the oil-bearing formation. In such experiments, for instance those on the displacement of oil by water, the cumulative oil production was measured as a function of the cumulative water injection. In order to interpret the measured production data, a need has for some time been felt for a means of observing visually what happens in the course of such experiments. Some information of this kind was obtained earlier by observing the water-oil distribution in various cross sections of the sand after the conclusion of the experiment.' However, the technique used did not guarantee that the fluid distribution remained undisturbed. Moreover, the procedure was somewhat elaborate and the information obtained limited. Therefore a new technique has been developed, which is described below. DESCRIPTION OF EXPERIMENTAL TECHNIQUE AND APPARATUS Glass models simulating part of an oil reservoir are filled with ground Pyrex glass of a specific grain size, representing the porous structure of the prototype. Before an experiment is started, the porous formation has to be saturated completely with an oil having exactly the same refractive index as the Pyrex glass powder. All light-scattering surfaces of the glass grains become invisible and the formation is then transparent. If the oil is displaced by another fluid (e.g. by water in a water-drive experiment, or by gas liberated from the oil in a solution gas-drive experiment) the formation becomes opaque in the region where the oil is expelled from the rock.* The back of the apparatus is painted black and illumination is from the front; therefore, in the photographs of the experiments, the oil-bearing formation appears black, whereas the water- or gas-invaded areas are white.

Jan 1, 1958

By R. S. Sherwin

The extractive metallurgy of primary aluminum from its ores is discussed with special attention to the production of alumina from high grade ores by the Bayer process, including differences between American and European practice and a brief description of some processes for lower grade ores and the electrolytic reduction of the oxide to aluminum. METALLIC aluminum is not found in nature, but the oxides, hydroxides, and especially the silicates are plentiful. The estimated percentage of in in the crust of the earth is about 8 pct while that of iron is about 5 pct. By far the larger portion of this is combined with silica in the form of various clay minerals and igneous silicate rocks. From the point of view of extractive metallurgy of aluminum, these are low grade ores while the better qualities of bauxite are the high grade ores. There have been various definitions of bauxite but perhaps the best definition at the present time is that bauxite is a rock or earth commonly used as an ore of aluminum or its salts in which the aluminum is present predominantly as a hydrate or a mixture of hydrates and hydrous oxides. It contains varying amounts of oxides of silicon, iron, and titanium and traces of compounds of some of the less common elements. The silica is mainly combined with alumina as clay or clay minerals which are hydrous aluminum silicates, although a part of it may be present as quartz sand. On the American continents, the alumina is mainly present as gibbsite, Al2O3 . 3H2O, and the same may be said of the best known deposits of the Dutch East Indies and some of the deposits in India. In France and other countries in Europe as well as in Africa, the alumina is present mainly as boeh-mite, A12O3 . H2O, but in some of these deposits it is mixed with minor amounts of gibbsite. Some other deposits, such as those in the islands of Haiti and Jamaica, evidently contain two or more hydrates or hydrous oxides of alumina in varying proportions. Perhaps the main portion of the alumina may be present as gibbsite and boehmite with the proportion between the two varying rather widely. In the silicate minerals, including clay, the alumina is chemically combined with silica and has not been separated satisfactorily by mechanical or physical ore-dressing methods. Low grade bauxites are mixtures of hydrates, usually gibbsite or boehmite, with clay, iron oxides, etc. In some low grade bauxites, it is possible to separate a portion of the gibbsite or boehmite, which may be present as relatively hard and coarse particles, from soft or finely divided clay minerals by log washing or similar methods. This has been applied to the product of some mines or parts of them, but on other ores it is not applicable. In some cases the gibbsite or boehmite is almost as fine and soft and of nearly the same specific gravity as the clay minerals so that washing and gravity separations are not successful. The iron oxide, the clay minerals, and a part of the titanium minerals are often so finely dispersed in the ore that any of the physical mineral separation methods, including separations by gravity, particle size, flotation, and electrostatic or magnetic separation, have not been commercially SUCCESSFUL except on relatively small lots of ore. For these reasons, the only available methods of separation on the general run of ores have been methods which would be classed as chemical rather than physical or mechanical separations. Aluminum oxide can be reduced by carbon at temperatures above 1800°C to form metallic aluminum and aluminum carbide or nitride. The temperature for rapid reduction of aluminum oxide to metallic aluminum is about the boiling point of aluminum and above the temperatures necessary to reduce iron, silicon, and titanium so that the direct reduction of an aluminum ore with carbon will produce an alloy of aluminum, iron, titanium, silicon, etc., which may be mixed with carbides and nitrides. Also a large amount of the reduced aluminum may be lost as a vapor except in the presence of some alloying agent such as copper or other metals. While it is possible to refine such alloys or mixtures so as to produce commercially pure aluminum, the methods which have been found are too expensive for the present market. One direct reduction method which found limited commercial use in Germany during World War II was the direct reduction of a mixture of clay containing very little

Jan 1, 1951

By J. K. Edgar

FOR a number of years the production and use of super-purity aluminum (better than 99.99 pct) has been steadily increasing. High-grade lots of such aluminum show certain outstanding characteristics not found in lower grades containing principally silicon, iron and copper as impurities. One such characteristic is the so-called "self annealing," that is, the recrystallization of the heavily cold worked metal at room temperature.' Interest therefore is aroused as to the effects of the impurities. It is well known that the amounts of silicon and copper which occur in aluminum of better than 99.95 pct are completely soluble, at least at elevated temperatures (above 200°C). Iron, on the other hand, is known to be fairly insoluble in solid aluminum. Thus, a knowledge of the limits of this solubility would be of considerable use in studying the characteristics of the high purity metal. All investigators have agreed that iron has very low solubility in solid aluminum. Rosenhain, Archbutt and Hanson2 reported traces of an aluminum-iron phase in the "purest samples of aluminum obtainable." Gwyer and Phillips3 reported that an iron constituent was visible under the microscope in metal containing ".0088 per cent iron." Roth1 found that in the solid state the solubility of iron in high purity aluminum was 0.01, 0.01, 0.017 and 0.022 pct at 225, 320, 500 and 600°C respectively. Dix5 working with alloys containing less than 0.03 pct impurities, concluded from microscopic and chemical analyses of eutectic areas that the eutectic composition was close to 1.7 pct iron. Dix placed the eutectic temperature at 655°C'. Archer and Fink,6 using high purity aluminum-iron alloys, found that the hypoeutectic liquidus intersects the eutectic horizontal at 1.7 pct. This is in fair agreement with the investigation of the aluminum-iron system by Gwyer and Phillips who placed the eutectic at 1.89 pct iron with a freezing point of 653° PREPARATION AND CHEMICAL ANALYSIS OF ALLOYS Electrolytically refined aluminum (99.991 pct aluminum) and a high purity aluminum-iron alloy (11.45 pct iron) were used in the preparation of the alloys for this investigation. To aid in controlling the iron content of the experimental alloys, an alloy of lower iron content was prepared first from these materials. The high purity aluminum was melted and heated to about 700°C in a plumbago crucible coated with alundum cement and the aluminum-iron alloy was then added. The alloy was thoroughly stirred, fluxed and skimmed and then chill-cast into notch bars. The analyses of the three materials are given in Table I. The alloys for the determination of the solid solubility of iron in aluminum were prepared by melting 600 g total of the aluminum plus the required quantity of the aluminum-iron alloy (No. 92531). These

Jan 1, 1948

By Norman Hedley, George Barsky, S. J. Sawinson

The cyanidation of precious-metal ores is a complex chemical process. Numerous reactions occur, some of which cause an undesired consumption of alkali and of cyanide. A knowledge of these reactions, their products, and the factors that influence the rates at which they proceed should be of value in working out the most economical conditions for cyanidation; but at present, much of this knowledge is lacking. This paper is confined to a study of some of the fundamentals of cyanidation. At the start of such an investigation it is necessary to simplify the problem under consideration by eliminating as many variables as possible. The results are then, of course, not immediately applicable to practice, but it is hoped in further work to round out the information and thereby make it of more practical value. For convenience in presentation, the paper has been divided into two parts, one devoted to a study of the reaction between cyanide and gold or silver, the other to a consideration of the factors that determine the rate of dissolution. I. REACTIONS INVOLVED IN DISSOLUTION OF GOLD AND SILVER IN DILUTE CYANIDE SOLUTIONS Ever since the cyanide process for gold and silver extraction was introduced there has been discussion of the equation representing the reactions involved. In undertaking a detailed study of these factors, it was felt that the most logical first step was to determine, if possible, what reactions could theoretically take place when gold and sill-er are dissolved by dilute cyanide solutions. It is possible to determine whether or not a given reaction will theoretically take place by calculating the free energy of the reaction. This free-energy value for any given reaction is the algebraic sum of the free energies of formation of each of the reactants and products, and, if these values are known, the calculation is simple. (A chemical change always involves an energy change and the amount of energy change involved is the factor determining whether a given reaction will take place or how far it will go. In chemistry we consider what is known as the "free energy"; that is, the maximum amount of energy that can be utilized as

Jan 1, 1935

By Joseph B. Umpleby

PROGRESS in the petroleum industry in 1929 has been characterized by outstanding accomplishments in the fields of new discovery of supply, economic control of production, increased efficiency and reduced costs of recovery, and lowered costs and greater flexibility in manufacture. Resources have been developed in an outstanding degree in the Kettleman Hills. California, the Oklahoma City field (which adds a large potential area in central Oklahoma), the discovery of Woodbine sand production in East Texas, and the discovery of Wilcox sand production in west-central Kansas. Engineering technique has progressed both along the lines of development and production, with a net result that costs have been reduced and extraction increased in notable degree, although it is impossible to give a close measure of either. In the field of re- fining crude oils variations to the end of promoting reduced cost and greater flexibility has been more important than genera! acceptance of innovations in practice.

Jan 1, 1930

By R. L. Templin

In treating a cast metal by any working process such as rolling, drawing or forging, variations in the conditions present in the remelting, casting, chilling and preheating of the initial ingot will certainly have some effect on the physical properties of the final product. In considering the effects of cold working on the physical properties of metals, however, it will be assumed that these factors, being controlled according to best practice, will remain constant and therefore may be neglected. Starting with such a uniform metal ingot, the next operation in the production of wrought metal involves hot working by rolling, forging or extrusion so as to produce a slab, billet or bloom. This working must be sufficient to produce substantial grain refinement and minimize any defects occurring in the initial ingot. Even such a hot-worked product is not suitable as a starting material for determining the effects of cold working because it contains internal strains analogous to those resulting from cold working. It is necessary, therefore, to supplement the hot working with a considerable amount of cold working, followed by a proper annealing in order to obtain a material suitable for use in defining the effects of cold working on the physical properties of a metal. That is, if the effects of cold working are to be evaluated accurately it is necessary to have other effects at least constant if not absent. Such material, unfortunately, cannot always be obtained in commercial practice but the requirements just indicated certainly represent the desired goal; this goal has been attained in the treatment of some metals, one being aluminum. Factors That Must Be Considered Simultaneously with Cold Working So many other factors affect the physical properties of a metal that it is necessary to consider these factors simultaneously with cold working; their effects vary in magnitude with the amount of cold work present. The most important of these factors are: (I) chemical composition, (2) structure, (3) temperature of working, (4) character of working, (5) rate of working, and (6) methods used for measuring results. For

Jan 1, 1930

By M. I. Sherman, J. D. H. Strickland

USE of a hydrometallurgical approach to the oxidation of sulfide ores and extraction of metals therefrom may have advantages over the more common smelting techniques when a low grade deposit is difficult to concentrate or the subsequent separation of metals, coexisting in the ore, is laborious by any known smelting operation. For economic reasons, the most promising oxidants are either atmospheric oxygen or electric power. The use of oxygen, or air under pressure, has recently been revised. Pyrrhotite has been converted to iron oxide and elementary sulfur' and a variety of sulfides have been treated by Forward and co-workers.2-4 Generally sulfate is the end form of the sulfur but with galena in an acid medium, elementary sulfur can be formed." For economic reasons chlorine and ferric iron salts are about the only possible alternatives to the atmosphere as oxidizing agents for base metal sulfides. If aqueous solutions of chlorine or ferric iron are employed, the reduction products can be oxidized electrolytically in situ and used again, thus acting as catalysts for electric power as oxidant. The use of ferric salts for this purpose is established hydrometallurgical practicea but, although chlorine gas has been employed in the dry state at an elevated temperature, its use in aqueous solution at or near room temperature has not found favor. The reaction of chlorine water with the soluble sulfide ion has been studied by several workers,7-9 and both sulfate and elemental sulfur are found as end products, the latter being favored by the presence of a low concentration of oxidant relative to that of sulfide in solutions of about pH 9 to 10. Of direct bearing on the work in hand are an early American patent" and a recent Austrian patent." The former advocates stirring powdered ore with an aqueous solution of ferric chloride chlorine oxides and chlorine. In the latter it is claimed that both metal and sulfur can be obtained by electrolysis, in a diaphragm cell, of a metal ore slurry in brine. Details in these patents are scant and no data or explanation is given for the mechanism of the reaction which, in the Austrian work, is attributed to the (unlikely) action of nascent chlorine at the anode surface. No mention is made of possible differences in behaviour between various ores. Apparatus A complication encountered when working with chlorine water is that a serious loss of chlorine occurs by gas partitioning unless an enclosed system is used and any air space in the apparatus is kept very small and constant. Arrangements were made, therefore, to take out samples for analysis without letting air into the system to replace the liquid removed. For convenience in studying a heterogeneous reaction the apparatus was so designed that a reproducible controlled stirring rate could be maintained and the ratio of surface area of ore to volume of solution was approximately constant throughout any experiment. The apparatus used is shown in Fig. 1. The ground ore was placed in the horizontal cylindrical vessel, A, of about 1 liter capacity, heated by a constant temperature circulating bath pumping water through the concentric jacket, B. By adding chro-mate to this water, an ultraviolet radiation filter effectively surrounded the reaction vessel, greatly reducing any possible photochemical decomposition of chlorine solutions. Stirring was effected by glass paddles, C, attached by an axle to a magnet which was rotated by another powerful Alnico magnet, D, outside the glass end, this magnet being itself rotated by an electric motor electronically controlled to constant speed. Speed could be varied from about 150 to 900 rpm and was measured and held to within 1 pct of a given value. The end of the reaction vessel remote from the stirring magnet was closed by another one-ended glass cylinder, E, connected by thin polyethylene bellows, F, clamped by screw clamps and watertight rubber gaskets to the main vessel. Through E, a glass electrode and calomel electrode projected into the solution and a hypodermic syringe pierced a small bung and allowed acid or alkaline to be added to maintain a constant pH. By pushing the fully extended bellows until the two cylinders touched, from 50 to 100 ml of solution could be forced out through a sintered disk into the three-way tap system, G, either to waste (for flushing purposes) or up into a 10 ml burette where the solution could subsequently be measured out for analysis. The ore samples were introduced at H, the tube being stoppered by a thermometer of —1 to +52ºC range, graduated to 0.1°C intervals. To prevent ore from being ground in the end bearings of the stirrer these bearings were pro-

Jan 1, 1958

By R. Wachtell

IN order to study better the phenomena at work in various phases of diffusion of the Fe/Si system when compounded and alloyed by powder metallurgy methods, several attacks have been planned. Electrical, hardness and other measurements have been discussed e1sewhere.l,2 The metallographic phenomena will be considered here. Concerning the metallography of the ferrosilicon alloys, some initial discussion is in order. As this laboratory discovered to its chagrin, there are many traps that lie in wait for even the experienced metal-lographer when he deals with alloys of high silicon content. Chief of these is probably the anomalous behavior of these materials under conventional etching techniques. We have before us, for instance, Corson's paper" of 1928, wherein is mentioned and illustrated (probably for the first time) the peculiar "barley shell" structure sometimes found in these alloys. Again in 19412 he same author reports the structure in 5-14 pct silicon irons. Houghton and Becker mention its presence' but make only tentative effort to explain it. In 1943, Hurst and Riley6 discussed the "barley shell" and declared it to be a false structure, resulting from the peculiar and characteristic film-forming propensities of the alloy. At about the same time, Wrazej7 proved correct the contentions of Hurst and Riley, establishing pretty well the mechanisms of formation and identifying the constituents involved. Hurst and Riley6 a1so describe in their paper (1943) a "cracked film" structure, also a pseudo-morph, the precise nature of which they do not suggest. Careful examination of such a film reveals that it not only cracks but begins to curl away from the surface of the metal and peel, much in the manner of flaking paint. The careful microscopist will be able to focus on the topmost edge of the flake, and, by optical sectioning, follow it down to the metal surface. We have been able, by oblique illumination of such specimens, to observe the raised segment of the film, the shadow that it casts, and the mating segment on the metal into which it fits. As late as 1946, Hurst and Riley8 found occasion to correct misinterpretation of this "cracked film" structure in a work by others on a 12 pct Si/Fe alloy. The "barley shell" and "cracked film" pseudo-morphs are not the only ones encountered in studies of these materials. A third pseudomorph, which may, however, have some structural significance is a fine striation of the surface. Closely spaced and parallel striae extend entirely across single grains. Each grain shows a different direction for the striae, and different closeness of the spacing. The regularity of this structure, and the fact that the striae do not cross from one grain to another, suggest that the structure is a film cracking controlled in some way by the crystal plane orientation of the surface being etched. In general it is wise for the metallographer dealing with these alloys to cultivate a feeling of uncertainty not only of the more complex questions of interpretation, but also of the basic question of whether that which he sees is indeed a true representation of the structure. We have taken, as a first test of the validity of a structure, its reproducibility "in situ" after repolishing and re-etching of the samples. Thus we knew the "barley shell" to be false (fig. 1) even before encountering the definitive works on the subject. The Metallography of Silicon Diffusion in Laminate Bars: As a first step in the study of the metallography of the diffusion process, we have pressed a series of laminate bars. These bars, or "sandwiches" as we have termed them, were pressed with covering layers of iron powder, and a "filler" of the Fe/Si master alloy under study. Bar dimensions were approximately %, x 3 x 1/4 in. The total

Jan 1, 1951

Jan 1, 1889

By Harrison R. Cooper

The key to concentrator automation is practical means for obtaining continuous assay data from concentrator streams. The technique most successfully used is x-ray fluorescence analysis. The practical reality of on-stream x-ray analysis has been achieved through effort over the past twenty years, marked by an occasional breakthrough, but more typically a step-by-step development to which many individuals and organizations have contributed. The net result of this progress is the present capability to install on-stream x-ray analyzer systems on a turn-key basis. It is now possible to provide all components of an x-ray analysis system as a completely engineered unit which can be put into immediate use, minimizing needs for specialized application knowledge to benefit from continuous assaying in process control. The economic factors which influence the choice of system and limitations of the x-ray fluorescence technique for slurry analysis are reviewed.

Jan 1, 1976

By Thomas W. Dawson

The H. C. Frick Coke Co. has used belt conveyers for handling coal for the last eighteen years but, until recently, only for small tonnages and over short distances. The first installations were outside the mines for distributing coal to various parts of plants. A belt conveyer was put in service at Colonial No. 4 Mine in November, 1906, to carry coal from the main hoisting shaft to a storage bin serving two blocks of beehive ovens on the other side of the valley; it was in constant service until the Colonial No. 4 ovens were abandoned less than a year ago. This conveyor was in an overhead gallery that spanned our sidings and yard tracks and the main line of the Pennsylvania Railroad. It was a crude installation, from the viewpoint of present conveyor engineering practice. It had no mechanical feeding device, the coal simply ran through a gate in the side of a bin on to the belt; but one belt on this conveyor lasted eleven years. This conveyor is about 300 ft. long center to center of pulleys. It operated at a speed of 250 ft. 'per min-and carried about 400 tons per day. On this basis, the belt life of eleven years means that not over a million and a quarter tons was carried by the belt. whi1e: this is not a large tonnage, the cost per ton was much less than by any previous method in use for supplying coal to ovens located at a distance from the main shaft. The next belt conveyor was installed at Bridgeport Mine in 1907. It carried about the same daily tonnage as the first, delivering from the hoisting shaft to a coke-oven bin. Very few data were kept on this conveyor. It also operated until the coke ovens at the plant were put out of blast and the coal taken to the Clairton byproduct plant by river. In 1916, work was started to equip the Bridgeport, Palmer, Gates, and Ronco mines to load their entire output on the river for transportation to the Clairton byproduct plant, which was then being built. These mines were all operating at that time and had a large area of coal available for river shipment. The Monongahela Railroad was located between the hoisting shaft and the river at each of these plants and, therefore, had to be crossed at each location to deliver coal to the river. After careful study, it was decided to transport this coal from the hoisting shaft to the river at each of these plants by belt conveyors.

Jan 1, 1925

By Walter H. Bruckner, George B. Clark

Since the end of World War 11 interest has been increasing in the use of shaped charges in the mining industry and in other industries using explosives for blasting purposes. Shaped charges employ the principle known as the "Munroe effect," which was discovered by Charles E. Munroe more than 50 years ago (in 1888). Details of their design have been explained elsewhere.1,2,3 Fig I shows the essential features of design of two types of shaped charges (with conical and hemispherical cavities) and a schematic sketch of their action upon detonation. The following discussion deals with the behavior of the metal in the cavity liners when they are subjected to intense Pressures exerted when the explosive charge is detonated. Among the conclusions reached in research on shaped explosive charges, the following have been established conceming cavity liners:1'2 I. The optimum wall thickness of a conical cavity liner is dependent upon the apex angle of the cone as well as On the base diameter of the charge. Acute apex angles require thinner walls for optimum perfonnance and more obtuse angles require thicker walls for the same base diameter. '. Cones were more effective in forming a penetrating jet when the walls were tapered; that is, the thickness of the wall of the cone increased from the apex down. 3. The physical and mechanical properties of metals have a marked effect upon their performance as cavity-liner material. Boiling point, ductility, malleability, ten-sile strength, and hardness are among the properties that influence the effectiveness of a metal used as a cavity liner. Lead, for example, makes a wide, flat crater in steel plates. Aluminum makes a deeper crater than lead, and an aluminum alloy having a high tensile strength makes a deeper hole, but slightly smaller in diameter. Cast iron makes a deep, narrow hole. These findings, together with the following analysis under Metallographic Survey of slug, offer solutions to many of the problems involved in solving the mechanism of the formation of Munroe jets. It has been fairly well established that cavity liners collapse in a manner similar to that indicated in Fig 2, Conical liners are known to collapse upon themselves while hemispherical liners are believed to turn inside out in the process of jet formation. The first has been definitely proved by recovery of collapsed portions of cones, while liners from partly detonated charges show that hemispheres may turn inside out. Fig 3 shows a sketch of a recovered slug from a 6-in. shaped charge using a 45" cone made of cast iron. (This size of charge drilled holes up to 3 ft in depth in solid granodiorite.)

Jan 1, 1949

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By James Elton

THIS paper covers, besides laboratory work, a study of actual operation at the Washoe Smelter over a considerable period of time, together with the results of a visit to the Midvale plant of the United States Smelting- & Refining Co., near Salt Lake. For the latter privilege, as well as permission to publish the resulting data herein given, I am indebted to George W. Heintz, General Manager of the Midvale plant, and here take pleasure in. expressing my appreciation for the many courtesies extended to me by Mr. Heintz and members of the technical staff at the smelter. Almost all of the world's supply of arsenic is recovered as a by-product in the smelting of other metals. It is seen on the market in various forms, the most common being arsenic trioxide, the ? white arsenic " of commerce, which is sold as a heavy white powder, or in white, glassy, translucent masses. The trioxide is used in the manufacture of pigment, glass, and insecticides. About 50 per cent. of the domestic consumption is supplied by imports.

Jan 8, 1913

By Frank C. Appleyard

The two calcium sulfate minerals-gypsum and anhydrite-occur in many parts of the world, and gypsum has long been of economic importance in the family of industrial minerals. Gypsum, the dihydrate form of calcium sulfate (CaS042H20) and anhydrite, the anhydrous form (CaSO4) are frequently found in close association, and it is seldom that a calcium sulfate deposit will consist exclusively of one mineral or the other. Although known gypsum deposits are extensive, anhydrite makes up the largest part of total calcium sulfate reserves. However, it has very minor economic use, and most of the following discussion will be devoted to gypsum. Calcium sulfate is one of the principal con- stituents of evaporite deposits, and when pure, has the following composition: [ ] Deposits of pure gypsum or of pure anhydrite which are large enough to be considered commercial have never been found because of both the metastable relationship between the two minerals, and the presence of impurities such as calcium or magnesium carbonates, chlorides, other sulfate minerals, clay minerals, or silica. As a result most mine production of gypsum will range between 85 and 95% pure. Often it is used as mined, although in certain cases, one or more forms of mineral beneficiation are employed to upgrade the product.

Jan 1, 1975

By R. A. Hadfield

The present tribute to a great pioneer in the steel industry, with the accompanying remarks upon the town of Sheffield, which has remained to this day what his invention made it, an important center of that industry, originally formed a part of the paper on "Iron Alloys," prepared, at the invitation of the Council of the Institute, for the Chicago meeting of August, 1893. That paper, begun with the intention of dealing with many alloys of iron, became, of necessity, through the limitations of time on the part of the writer and of space in the Transactions of the Institute, mainly a discussion of the principal modern commercial alloy of iron, manganese-steel; and much introductory and general matter was consequently omitted. The Secretary suggested, however, that this particular portion would be interesting and appropriate as a separate contribution, and I therefore present it as such, offering the foregoing statement simply as an explanation of its incomplete character. If I had undertaken to write a paper upon this subject alone, I should have felt myself bound to bestow upon it more elaborate research and to treat it with more abundant details. It was in the immediate neighborhood of Sheffield that the first successful process for the fusion of steel on a commercial scale saw the light. The late Dr. Percy, a leading authority in general metallurgy, says: " Formerly, so far as I am aware, steel was never melted and cast after its production; and in only one instance, viz., that of wootz was it ever molten during its production.....Now, by the fusion and casting of steel after its production its heterogeneousness is remedied, and ingots of the metal may be procured of perfectly uniform composition throughout; and for the practical solution of this important problem we are indebted to Benjamin Huntsman, of Sheffield." As a recent American journal appropriately remarks: " Huntsman's patient efforts, at last rewarded with success, entitle him to an elevated niche among the heroes of industry. The invention of cast-steel was second

Jan 1, 1895