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By K. Osseo-Asare

The presence of carbonaceous matter in gold ores presents a two-fold problem: (1) The poor release of gold from the carbonaceous matrix and (2) the uptake of dissolved gold by the carbonaceous leach residues. A critical review is presented of the published information from various parts of the world pertaining to the nature of the carbon-containing components of these ores and their different roles in the gold extraction process. Included in this paper are new results based on the authors' research on carbonaceous ores from the Prestea goldfields of Ghana. Methods for isolating and characterizing the carbon-containing materials are given and a discussion is presented of the interaction of the carbonaceous ores as well as the fractionated components with aurocyanide solutions. Emphasis is on basic physico-chemical phenomena, investigated with the aid of X-ray diffraction, infrared spectroscopy, adsorption, and electrophoretic mobility studies.

Jan 1, 1984

By R. Simon

Maurer's review1 summarizes quite thoroughly the various theoretical developments and experimental findings that contribute to the knowledge of rock mechanics in drilling. This discussion supplements Maurer's paper by including generalities that either have been incompletely treated in his review, or formulated from new considerations by the discussant and from hitherto unpublished experimental data. These generalities are associated with the observation that all mechanical methods of rock drilling essentially comprise a sequence of indexed, concentrated loadings on the bottom of the hole being drilled. A concentrated loading is defined as a mechanical loading by a tool that contacts at any one time only a small fraction of the surface of the body being worked and penetrates only a small fraction of the body thickness. The surface in this application is the bottom of the hole. The deformation of ductile material by concentrated loading has received considerable attention because of the importance of indentation tests for measuring hardness of metals. Some of the generalities involved in concentrated loadings on ductile materials apply also to concentrated loadings on more brittle materials such as rock. Among these generalities is the applicability of similitude considerations to scaling laws and to force-displacement and energy-volume characteristics for tools of various shapes, as summarized by Tabor.2 Vork on the flow and fracture behaviors of glass loaded with pyramidal and spherical indenters is reviewed by Douglas.3 Maurer refers to the work of Evans and Murrell on wedge loadings on coal and on mixtures of pitch and sand. It is evident from Maurer's paper that a considerable portion of the studies of rock mechanics in drilling, such as crater formation experiments, has involved concentrated loadings on rock.

Jan 1, 1967

By W. S. Morris

PETROLEUM'S importance in World War II can perhaps be better realized by the recitation of a few facts and figures: Gasoline needs in this war are already eighty times greater than in the last war. It takes nearly three pounds of gasoline to deliver one pound of bombs, filled with petroleum explosives, on an enemy objective. Every American soldier overseas requires an average of more than fifty gallons of petroleum products per week, with the oil used by Navy and supply ships and America's Allies boosting the consumption of oil per soldier to an even higher figure. Every oil well in the United States now must produce each day an average of four barrels of oil for war and combat jobs. More than fifty million gallons of gasoline, fuel oils, lubricants, and other petroleum products are now going directly to the fighting forces every day.

Jan 1, 1945

By Walter Renton Ingalls

Metallurgical literature has no record of any ore beneficiated for cadmium alone, and the cadmium of commerce is derived from zinc ore, with which cadmium is generally associated. Zinc ores free from cadmium—e.g., the ores of the Franklin and Stirling mines, New Jersey, and of Broken Hill in Rhodesia—are rare. Blende concentrates of 50 to 60 per cent grade have contained (in production of important tonnage) as much as I per cent Cd, which is unusually high. The concentrates from the Tri-State district average from 0.3 to 0.4 per cent Cd, which is high. The concentrates from mines west of the Rocky Mountains seldom are higher than 0.2 per cent. In the periodic system of the elements, cadmium is in the same group with zinc. Its properties and compounds are similar. Its metallurgy is also similar. Cadmium has a melting point of 320°C. and a boiling point of 778"C., while the melting point of zinc is 415'C. and the boiling point gos°C. These conditions, together with the lower reduction temperature of cadmium oxide, indicate means for. separating cadmium from zinc. Cadmium in spelter is viewed, if not as an impurity, at least as an alloying element that may be objectionable, its special effect being hardening. However, for some purposes, a little cadmium in spelter is desired, and 0.4 per cent may be permissible, even in spelter for rolling. Cadmium in spelter for brassmaking is completely volatilized at the temperature of that process, and in bygone days an immense quantity of cadmium must have been lost in the fumes from these furnaces. Pyrometalltjrgy—Early Cadmium compounds, sulphide and oxide, being more volatile than the corresponding compounds of zinc, loss of cadmium begins to occur from the first furnace operation. In normal roasting, even with Tri-State ore, this loss may not be very high. Raw ore, assaying 0.4 per cent Cd, may be expected to give a roasted product of the same assay, which would imply a cadmium loss of about 15 per cent. Roasted ore, assaying 0.4 per cent Cd when distilled at ordinary furnace temperature, yielded spelter containing about half of that in the ore charged. The first draw of spelter under these conditions may assay 0.8 per cent Cd (accounting for 22 per cent of the cadmium recovery) and after the several draws of spelter have been equalized the cadmium content may average 0.35 per cent. If, therefore, 2000 lb. of blende averaging 0.35 to 0.4 per cent Cd is roasted and distilled, there may be obtained 1000 lb. of spelter, assaying 0.35 per cent Cd or 3.5 lb. content, which entering into Prime Western spelter will be sold at the spelter price. It will be observed from the outline hereinbefore (summarizing conditions prior to the introduction of sintering) that the main loss of cadmium occurred in the distillation process and by failure to condense, which could be ameliorated by attaching a prolong to the condenser, thus collecting

Jan 1, 1944

By Joseph B. Story, John T. Clarke

A modification, of the Kelvin bridge using an inductor was used to measure the conductivities of molten sodium chloride, calcium chloride, and mixtures thereof. A capillary-type four-lead fused quartz dipping cell was constructed. The effect of a small amount of potassium chloride on the conductivity of the melt was determined. IN the electrowinning of sodium and chlorine from sodium chloride in Downs cells, calcium chloride is added to the sodium chloride in an approximately equimolar amount to permit cell operation at a lower temperature. This lower temperature improves current efficiency and ease of operation, but decreases voltage efficiency by decreasing the electrical conductivity of the melt. A knowledge of the electrical conductivity of the Downs cell melt is necessary to the understanding of Downs cell operation, for the IR voltage drop across the melt is responsible for a significant part of the electrical power consumed by the cell. The work of previous investigators on fused sodium chloride-calcium chloride mixtures was reviewed from the standpoint of consistency with the results of others on the conductivities of the pure constituents. The electrical conductivity of pure fused sodium chloride has been established with accuracy by Van Artsdalen and Yaffe,1 by Edwards, Taylor, Russell, and Maranville,2 and by Lee and Pearson.3 It has also been studied by a number of other investigators.4-13 The best values for calcium chloride are believed to be those of Lee and Pearson,3 since the sodium chloride and potassium chloride conductivity values given by them check the best literature values. A number of other investigators have also reported values for calcium chloride.'-"."' Electrical conductivity data for fused sodium chloride-calcium chloride mixtures have been reported by Sandonnini,12 Vereshchetina and Luz-hnaya,18 Ryschkewitsch,11 Barzakovskii,5 and Alaby-shev and Kulakovskaya.15 Sandonnini reports no data at temperatures lower than 850°C. The conductivity values reported by Vereshchetina and Luz-hnaya, by Ryschkewitsch, and by Barzakovskii for pure sodium chloride and for pure calcium chlor- ide vary somewhat from the literature values believed to be best and the sodium chloride-calcium mixture data of these authors and of Alabyshev and Kulakovskaya (who give no data for the pure compounds) show significant variations. In view of these facts, it seemed that an investigation of the electrical conductivities of fused sodium chloride-calcium mixtures would be worthwhile. Apparatus Bridge System—The bridge used,in this investigation is diagrammed in Fig. 1. In general, the elaborations of this bridge over the simple Wheatstone bridge were for the purposes of a) eliminating lead resistances or making them accurately measurable, and b) eliminating major capacitative and inductive effects or balancing them against each other. Lead resistances in the lower arms of the bridge were kept as low as possible by making the leads short and of low resistance wire. This could not be done in the cell arm of the bridge because of the remoteness of the cell from the other bridge components and because of the high-resistance nickel or chrome1 wires used in the cell. Since maximum sensitivity is attained when the resistances of the four arms of the bridge are equal, 50-O resistors were used in the lower arms of the bridge when fused salt conductivities were being measured, and 1000-O resistors were used when the cell was being calibrated with normal aqueous KCl solution. The cell resistance was of the order of 50 ohms with fused salts and of the order of 1000 ohms with normal aqueous KC1 solution. The resistors used were non-inductive and accurate to 50.05 pct (General Radio Co. Types 500-C and 500-H). The lead resistance between point A in the cell and point E at the measuring resistor was eliminated by using the kelvin bridge principle; that is, the voltape drop across the lead from point A to point E was divided between the cell arm of the bridge and the measuring arm of the bridge in proporation to the resistaxes in the lower arms of the bridge.

Jan 1, 1958

By D. T. Peterson, T. Carnahan

The diffusion coejTicients of carbon, nitrogen, and oxyget were determined in $ thorium over the tempernilcre range 1440" io 1715°C. The diffusion coyfiicir?zls are given by: D = 0.022 exp (-27,000/RT) jor carbo)~, D = 0,0032 exp(-l7,00Q/RTj for nitrogen, and D = u.0013 expt(-11,UOU/RT) for oxygen. Cavl~orz was found to increase the hardness of thoriunz nearly linearly with concentration over the range 100 to 1000Ppm carbon. ThORIUM has a fcc structure up to 1365°C and a bcc structure from this temperature to its melting point at 1740°C. Diffusion of carbon, oxygen, and nitrogen in bcc thorium was of interest in connection with the purification of thorium by electrotransport.' In addition, it was possible to measure the diffusion of all three of these interstitial solutes in the same bcc metal. Only in niobium, tantalum, vanadium, and a iron have all three interstitial diffusion coefficients been measured in a given bcc metal. Diffusion coefficients have been measured for carbon and oxygen in a thorium by Peterson2, 3 and for nitrogen by Gerds and Mallett.4 Activation energies for diffusion are reported by the above authors to be 38 kcal per mole for carbon, 22.5 kcal per mole for nitrogen, and 49 kcal per mole for oxygen. Values of the diffusion coefficients of carbon and nitrogen in 3 thorium have been reported by Peterson et al.' However, these were secondary results of their investigation of electrotransport phenomena in thorium and it was hoped that the present study could provide more precise data. EXPERIMENTAL PROCEDURE The specimens used in this study were the well-known pair of semi-infinite bar type. The couple was formed by resistance butt welding two 0.54-cm-diam by 3.0-cm-long bars of thorium together under pure helium, the concentration of the solute being greater in one cylinder than that in the other. The finished couple then contained a concentration step at the weld interface and diffusion proceeded only along the axis of the rod. The thorium used in this study was prepared by the magnesium intermediate alloy method.5 The total impurity content was less than 400 ppm. The major impurities were: carbon, 100 ppm: nitrogen, 50 ppm; and oxygen. 85 ppm. The total metallic impurity content was less than 150 ppm. The high solute concentration portions of the diffusion couples were prepared by adding the solute to the high-purity thorium in a non-consumable electrode arc melting procedure. Carbon and nitrogen were added in the form of spectroscopic graphite and nitrogen gas while a Tho2 layer was dissolved by arc melting to add oxygen. High-purity thorium formed the low concentration portions in the carbon and nitrogen couples. The low oxygen portions were obtained by deoxidizing high-purity thorium with calcium for 3 weeks at 1000°C according to a method reported by Peterson.3 The high C-Th contained 400 ppm C, the high N-Th contained 400 ppm N, the high 0-Th contained 220 ppm 0, and the low 0-Th contained 25 ppm O. The high O-Th was brine-quenched from 1500°C to retain most of the oxygen in solution at room temperature. These concentration levels were all below the solubility limits in 0 thorium at 1400°C. A resistance-heated high-vacuum furnace was used to heat the couples. The samples were mounted horizontally on a tantalum support which had small grooves near each end. Spacer rods of thorium, 0.4 cm in diam, were placed in these grooves to prevent contact between the sample and the tantalum support. This arrangement should have prevented contamination of the sample by contact with the support. In further effort to reduce contamination, the oxygen diffusion couples were sealed inside evacuated outgassed tantalum cylinders lined with thorium foil. Thorium rings around each end of the samples acted as spacers in this case. Pressure during diffusion runs was about 10-6 torr after an initial outgassing stage. Temperature measurements were made by sighting on black body holes in the sample support adjacent to the samples with a Leeds and Northrup disappearing-filament optical pyrometer. Temperatures were constant during a diffusion anneal to ±5C. The observed temperatures were corrected for sight glass absorption after each diffusion run. The pyrometer was checked against a calibrated electronic optical pyrometer and a calibrated tungsten strip lamp with the electronic pyrometer being taken as the standard. All temperature readings agreed to within ±3C over the temperature range 1450" to 1690°C. Time corrections due to diffusion during heating and cooling were necessary because of the short diffusion times. The diffusion times ranged from 6 min for the oxygen sample run at 1690°C to 90 min for the carbon sample run at 1500°C. A series of temperature vs time plots were made for heating and cooling of the samples to the various diffusion temperatures. This data was then used in a method according to shewmon6 to determine the time corrections. The corrections amounted to

Jan 1, 1970

By J. E. House, J. L. Drobnick, R. R. Swanson, D. W. Agers

Since the commercial acceptance of the liquid ion exchange process in the mineral processing industry, it has been predicted that eventually the hydrometallurgist would have a wide selection of commercially available chemical reagents to make separations efficiently and economically of virtually all metals in solution under a wide range of conditions. Several years ago, General Mills began a research and development program to determine if it were possible to develop a reagent to extract a specific metallic ion from solution at a cost that would allow large scale commercial application. A giant step toward reaching this objective was announced in late 1963, when development quantities of LIX®-63 reagent, a highly selective reagent for the extraction of copper, were first offered to industry. The properties of this reagent were de- scribed in a previous paper.1

Jan 12, 1965

By S. Lipson, H. W. Antes, H. Rosenthal

A study was made of the effect of ingot structure on the strength and ductility of high-strength wrought-aluminum alloys. It was found that a fine-cast structure facilitated complete homogenization which, in turn, resulted in significant increases in ductility and strength. A completely homogenized 7075-T6 alloy developed tensile properties of 85,000 psi UTS, 75,000 psi YS, with 40 pct RA. Completely homogenized 7001-T6 alloy tensile properties were 102,000 psi UTS, 99,000 psi YS, with 19 pct Ra. A method was devised for making small ingots having secondary dendrite arm spacing of less than 10 u. This method involved multiple-pass arc melting of commercial rolled plate with a tungsten electvode. This material could be completely homogenized after 3 hr at 900°F; homogenization of the original plate material was not complete after 120 hv at 900°F. Degree of homogeneity was determined by use of metallographic and electron-microprobe analyses. The electron-micro-probe study also showed the preferential segregation of solutes in the microstructure. HIGH-strength aluminum alloys, such as those of the 7000 series, usually freeze by the formation and growth of dendrites. The dendrite arm spacing (DAS) depends on the rate of solidification.' Commercial ingots are usually direct chill-cast to promote more rapid solidification, but, due to the large mass of the ingot, localized solidification times are long and a large DAS results. During solidification, solute elements are rejected by the solid as it forms, causing enrichment of the liquid and ultimately solute-rich interdendritic regions. In order to attain a homogeneous ingot, the segregated solutes must diffuse across the dendrite arms. The larger the DM, the longer the time for complete homogenization. In the case of commercial ingots, the DAS is so large that the time for complete homogenization is prohibitively long and, therefore, second phases or compounds are always present. These un-dissolved phases are carried over to the wrought material during processing, resulting in an impairment of strength and ductility. In addition, the mechanical fibering of the undissolved second phases or compounds during working results in mechanical property anisotropy. If complete homogenization could be attained, higher ductility could be expected. The realization of higher ductility at current strength levels is a desirable objective; however, if higher-strength alloys were wanted, it might be possible to sacrifice some of this ductility by adding more solute elements and produce even higher-strength alloys than are currently available. Further, if complete homogenization leads to more efficient utilization of solute elements, then more dilute alloys should have relatively high strengths with very high ductility. In all instances, it would be expected that the degree of mechanical property anisotropy due to mechanical fibering would be reduced. Therefore, it was the purpose of this investigation to produce cast structures that would facilitate homogenization and to determine the effect of homogenization on the properties of high-strength, wrought-aluminum alloys. MATERIAL CLASSIFICATION Commercial Alloys. In order to illustrate the non-homogeneous condition that exists in commercial high-strength, wrought-aluminum alloys, typical micro-structures of 7001, 7075, and 7178 are shown in Fig. 1. The chemical compositional specifications of these alloys are given in Table I. It can be seen in Fig. 1 that a considerable amount of undissolved second-phase material is present in each of these alloys. The solute elements associated with the undissolved phases were identified by electron microanalyses. Back-scattered electron images and characteristic X-ray images of the three commercial alloys are shown in Figs. 2, 3, and 4. These data indicate that the second phases are regions of high copper and high iron-copper concentrations. The second-phase material also was analyzed for magnesium, zinc, manganese, chromium, and silicon, but no significant enrichment above that of the matrix was found. Therefore, the problem of homogenization resolved itself into one of dissolving the copper-rich and the iron-copper-rich second phases. In order to accomplish this objective, two approaches were made. The first was to reduce the iron as low as possible since this element has a maximum solid solubility of 0.03 pct in aluminum. The second was to produce cast structures with finer DAS to facilitate dissolving the second phases. Commercially Produced High-Purity Alloys. A special high-purity, 2000-lb ingot of 7075 alloy was made by a commercial producer. This alloy contained the following weight percentages of solutes: 5.63 Zn, 2.48 Mg, 1.49 Cu, and 0.21 Cr. All other elements combined were less than 0.02 pct by wt including iron and silicon at less than 0.01 pct each. The ingot was cast and processed into rolled plate using standard commercial techniques. Microstructures of standard commercial 7075 and the special high-purity 7075 are shown in Fig. 5. It can be seen from this figure that the high-purity alloy has less undissolved second-phase material, but a significant amount was still present. The second phase in the high-purity material did not contain iron but it was found to be enriched with copper. The slight effects of the increased purity and de-

Jan 1, 1968

By A. B. Johnson, J. S. Browning

The USBM has developed new and improved physical beneficiation techniques for economically recovering potash minerals from high-clay and low-grade ore from the Permian Basin of New Mexico. With the depletion of the high-grade ores a need for changes in industry's standard methods of recovery is a must. Technology may be improved in the beneficiation of the ore prior to mineral flotation and in brine clarification from the desliming phase to effect a reduction of losses of fine potash and brine in the disposal of the clay slimes. This report summarizes the results of these tests and demonstrates the feasibility of producing products that meet the requirements of the agriculture market, namely, 95% KCl equivalent to 60% K2O with recoveries ranging between 80 and 90%.

Jan 1, 1973

By A. M. Gsudin, F. W. Bloecher, C. S. Chan-s, P. L. De Bruyn

M ANY elements can now be obtained in radioactive form. The radioisotopes have the same chemical properties as the corresponding inactive forms, differing from them only by their nuclear instability. This combination of properties permits radioelements to act as atomic tracers, the tracer atoms revealing themselves at the time of their decay with emission of electrons and positrons (beta radiation), helium nuclei (alpha radiation), or electromagnetic radiation (gamma radiation). The tracer technique is particularly attractive for the study of surface chemistry problems in general and flotation problems in particular because of its extreme sensitivity. A large proportion of the ionizing events, initiated by the characteristic decay emissions in a radioactive sample, can be recorded by radiation probes of suitable design and geometry. An idea of this sensitivity can be gathered from the following example: assume that 0.001 mg of pure RaD (radiolead), a beta-ray emitter, is used in a flota¬tion tracer experiment. This is 1210 X 10-1 mole; since the fraction trans¬formed per second (the so-called decay constant), is 9.92 X 10-10, the number of ionizing events per second would be (6.023 X 1023 X y,,, X 10-1) X 9.92 X 10-10 = 2.845 X 106. (The decay constant is expressed in sec-'. Its product with the half life, expressed in seconds, is the pure number, natural logarithm of 2 = 0.693. The half life is the length of time required for decay of one half of the total number of radioactive atoms of one particular type.) (See reference (1) at the end of this article.)

Jan 1, 1948

IN the decade beginning 1850, the development by William Kelly in I the U.S.A. and Henry Bessemer in England of the pneumatic method of refining pig iron, known as the Bessemer process, gave the world the first method of producing low-cost steel in large volumes. Briefly, the Bessemer process consists of blowing air through liquid blast furnace iron in a refractory-lined vessel known as a converter. The air enters the iron in the furnace bottom through tuyeres, and during its passage through the iron removes the carbon, manganese, and silicon by oxidation. The heat produced by oxidation is sufficient to raise the temperature of the liquid metal to normal steel casting temperatures and, in addition, to melt some scrap. In the 1860's, the open hearth process was introduced which in the early part of the twentieth century surpassed the Bessemer process in steel production. In the early 1870's, Thomas and Gilchrist discovered that phosphorus could also be removed if lime were added and the converter lined with basic refractory; the modification is known as the basic Bessemer process and is used in Europe. Bessemer steel usually contains between 0.010 and 0.020 per cent nitrogen and between 0.05 and 0.10 per cent phosphorus which influences the properties. Henry Bessemer,1 in 1856, recognized the possibility of using oxygen in the converter process, but the cost was prohibitive. This interest was associated with the theoretical balance of the Bessemer process. Experimental work' has indicated that bottom blowing with pure oxygen results in serious converter bottom and tuyere erosion. Dilution of the oxygen by steam has been observed to be an effective means of maintaining low-nitrogen content of the steel and at the same time reducing bottom and tuyere erosion. A further advantage is that the fume emission characteristic of Bessmer operation is reduced. The addition of steam to the oxygen, however, reduces scrap-melting capacity due to the endothermic reaction characteristics of steam. The possible advantages of refining methods for steel using high-purity oxygen have been recognized for over a hundred years, but the develop-

Jan 1, 1964

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By M. D. Everell

A study has been made of the relationships between grinding selection functions and physical properties of rocks, particularly the failure loads of irregular particles and the compressive and tensile strengths of large regular specimens. The rocks used were limestones of various structures and impurity contents, a carbonatite, and a high-purity crystallized calcite. The selection functions were obtained from carefully performed grinding experiments on selected narrow size ranges in various grinding machines-a laboratory batch ball mill, a Hard grove grindability machine, and a mill of the type suggested by Yang, Mempel, and Fuerstenau. The specific rates of breakage during milling and the failure loads of irregular specimens correlated well for the rocks studied. The testing of irregular particles is suggested as a means of predicting the behavior of commercial mills.

Jan 1, 1973

By William P. Blake

The literature of the occurrence of quicksilver-ore in the United States does not contain, so far as the writer is aware, any mention of the locality herein described. In the preliminary report * upon the resources of the transPecos region of Texas, von Streerwitz gives a long list of minerals observed as occurring there, but cinnabar is not mentioned. In the second report upon the same region † mention is made of the reported existence of cinnabarite in one of the mountain ranges north of the Sierra Carrizo and the Bofecillos, but the author adds: " In spite of my careful examination of the float, I have not yet found any traces of this metal (quicksilver) up to the present time." Prof. Dumble, also, in * Geol. Surv. of Texas, E. T. Dumble, 1889, p. 225 (first report). † Geol. Surv. of Texas, E. T. Dumble, 1890, p. 713 (second report)

Jan 1, 1896

By H. J. Rahilly

MINE FIRES, in the Butte District, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the top of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "Horsetail Ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 2, 1922

By K. N. Weaver

In the early 1950's the cement industry began putting a new emphasis on geology. This article points up some of the industry's raw materials problems that geologists are uniquely qualified to handle. Portland cement can be made from relatively abundant industrial minerals and rocks, and this may explain why cement producers placed little emphasis on geology during the early days of the industry. After World War 11, however, the industry began to recognize the need for geological exploration and some of the larger companies began to build a permanent geological staff. This development has its roots in: 1) The rapid expansion of the industry. After World War II and continuing into the early 1950's the demand for cement far exceeded the production capacity. This placed a great deal of pressure on the producers to enlarge capacity at their plants and to develop new sites. 2) The closing of many cement plants due to inadequate raw materials. Producers feared lack of raw materials due to inadequate exploration and evaluation. 3) An ever increasing pressure for upgrading of the quality of portland cement, necessitated upgrading the quality of raw materials. 4) The increasingly large capital investment needed to build new cement plants required a higher degree of accuracy in reserve calculation. These reasons are more or less self explanatory and will not be elaborated on. They are outlined to illustrate the probable evolution of thought among cement management to justify the development of a geological staff. In the author's personal experience the geologist must continually readjust to changing conditions within the industry. For example, cement production capacity grew so rapidly that today production is only about 70% to 75% of rated capacity (73.8% in 1963). Therefore, whereas the geologist's function during the active expansion of capacity was primarily one of exploration, one of his major functions today is the development of existing reserves. This assumes of course that reserves at existing plants have been proven to be adequate in terms of the company's reserve policy. Through the proper development of raw materials, the geologist can markedly influence quality control, production and long range planning. Examples of several of the operations of Medusa Portland Cement Co. will illustrate how proper attention to geologic detail pays off in the above areas. Because limestone is the major raw material of port-land cement, the examples will be concerned with it. DEVELOPMENT OF GEOLOGIC DETAIL One of the prime objectives of an exploration or development project should be the establishment of the stratigraphy and structure of the deposit. The basic tool for exploration and development is still the core drill, and the geologist must gain a maximum amount of information from a minimum amount of drilling. In order to obtain this objective the following is essential: 1) Adequate detailed geologic logging of cores. The core is first logged and separated into lithologically distinct units. In order to make the lithologic breaks as meaningful as possible, representative pieces of core (not all) are etched in dilute HCI. This removes the "drill polish", brings out the structure of the rock and indicates the proportion of dolomite and argillaceous material because dolomite and argillaceous material etch in relief. In this way the lithologic breaks become clearer and one gains a general idea of the chemical makeup before analysis. This method of logging lessens the probability of including two types of stone, e.g. limestone and dolomite into the same sample for chemical analysis. 2) Representative sampling for chemical analysis. After logging, the lithologic units are sampled for chemical analysis. A representative analysis can be obtained by extracting a 1-in. sample for every foot of core. For example, a lithologic interval of 10 ft to 20 ft would yield ten 1-in. samples taken at 11 ft, 12 ft, 13 ft... and 20 ft. These samples are then crushed, pulverized and split into a sample for chemical analysis. This sampling method represents a significant saving of time over splitting the entire core. 3) Preparing the core as a record. Because of the expense of storage and the difficulty of handling large amounts of core, only representative samples of each lithologic unit are saved for reference. An

Jan 1, 1965

By H. OSBORG

THE patent literature of alloys for the last two decades or so indicates that the number of liatents referring to smaller and smaller percentages of essential alloying constituents is on the increase, whereas the number of patents relating to considerable percentages of alloying additions becomes smaller. The claims regarding physical properties indicate consistently higher and more rigid values and show the possibility of modifying the physical properties by only slight changes in the percentage of minor constituents or by applying different methods - of treatment. Alloys of the same base have been made suitable for a number of different applications by intelligently and rigid- ly controlling their composition, melting, casting, working, and heat-treatments. Other trends indicated in the development of modern alloys are higher corrosion resistance, higher fatigue limits, and "stability" in volume.

Jan 1, 1937

By P. H. Mell

AMONG the minerals exhibited at the Atlanta Exhibition of 1881, soapstone, kaolin, and asbestos were well represented. The first two occur in large quantities, of very pure quality, throughout the South. The last is somewhat limited in amount, yet in certain localities it exists in masses sufficiently large to pay for working; in North Carolina, Georgia, and Alabama it has been mined with some profit. I would, in this paper, call special attention to the minerals soapstone and kaolin, because their many valuable qualities seem to have been al most entirely overlooked by our Southern manufacturers. Steatite, particularly, occurs in such large quantities, and generally in such close proximity to the great iron regions of Alabama and Georgia, that it is amazing that the iron companies will persist in sending abroad under. heavy expense for firebrick. It would require a much smaller sum of money to quarry this rock and place it at the various furnaces ; and the results obtained, I am sure, would prove its superiority over the majority of firebrick brought into this region. Soapstone, as well as kaolin, is found throughout the metamorphic regions of Alabama, Georgia, South Carolina, and North Carolina. Below-will be found seven analyses that I have made of this mineral, taken from various points in the four States above mentioned. I have also subjected these same specimens to severe fire-tests, and the results were very satisfactory. The tests were conducted in the following manner: Small portions of the compact variety were carefully weighed and placed in a Fletcher's new crucible furnace for gas, and subjected for half an hour to a heat that will in 12 minutes fuse iron. The only effect upon these specimens was a slight change in color .to a darker shade, a small loss in weight, due evidently to the evaporation of water, and, in one or two specimens, a slight degree of fusion on the sharp edges. The first specimen subjected to investigation was taken from a large outcropping in Tallapoosa County, Alabama. Specific gravity, 2.88 ; hardness, 1.9; color, blue gray. Composition Water, - 2.68 Silica, - 65.46 Iron oxide, - 2.10 Alumina, - 15.02 Magnesia, - 15.35 100.61

Jan 1, 1882