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By R. J. Hopkins, L. B. Haney

On the basis of limited experimental work conducted at the Port Pirie smelter, it would appear that, by increasing the specific surface of sinter, and possibly that of coke as well, a marked increase in the rate and degree of sinter reduction in the furnace shaft can be achieved. Fusion-point tests have shown that this increased reduction means a much higher sinter fusion point and narrower plastic zone, while the more efficient utilization of coke in the furnace is inherent in the intensification of reduction by the provision of a more reactive sinter surface. Investigations aimed at increasing the ability to operate the blast furnace with a higher lead tenor of sinter reducing fuel cost proportionately, and at eliminating the loss in reduction potential in coke entering the furnace, as shown by the presence of appreciable proportions of CO in the top gases, are reported. Several lines for future investigations that could result in improved smelting methods are proposed. A SOMEWHAT informal presentation of ideas on lead blast-furnace smelting will be given in this paper. These ideas have been developed during observations of furnace behavior at Port Pirie and earlier at Mt. Isa, and have been guided by the results of laboratory experiments at Port Pirie. The authors do not intend to propound any particularly new theories, but present their ideas as a basis for discussion which may help toward a better understanding and improved control of blast-furnace operation. Performance variations of considerable magnitude occur from time to time at most lead blast-furnace plants. While Port Pirie is fortunate in having a very regular supply of high grade concentrates, it is no exception in this respect. Such irregularities are of considerable economic importance and they constitute a challenge to present-day metallurgists in their efforts to obtain an ever greater degree of control over furnace operation. In operating as Port Pirie does with a relatively high lead charge, it has been recognized for a long time that shaft conditions are ,of the greatest importance. A practical, or engineering, attack on the problem of accretions consisted of changes in furnace design, e.g., by widening the shaft to 10 ft above the bottom row of tuyeres, and by the provision of curved ends and vertical walls.' That this attack has had a great measure of success is shown by the fact that barring down of these wide furnaces is rarely necessary, shaft conditions have greatly improved, and sensitivity to charge variations has been markedly reduced. These modifications, followed by some metallurgical changes (chiefly the incorporation of pig iron into the furnace charge in 1948), have brought furnace operation to a point where some variables which were previously masked can now be studied. Accordingly, over different periods of time during the last five years, such variables as could be given numerical values have been recorded and subjected to statistical examination. The objectives of this work were: 1—to reveal the true importance of the different variables and 2—to show the proportion of total furnace variation which is unexplained and therefore referable to causes not measured at present? e.g., shaft accretions, sinter structure, coke structure, and others. Statistical Study of Variables In this work factors relating to furnace speed and lead reduction have been examined: 1—slag com-

Jan 1, 1955

By E. M. Elkin, J. H. Schloen

All known methods of treating and recovering the various components of copper refinery slimes are discussed. The slimes treatment processes presently used by five copper refineries are described and flowsheets are given. A bibliography is appended. THE recent increase in anode copper production at the Noranda smelter placed an additional load on the Montreal East plant of Canadian Copper Refiners Limited. The consequent increase in the amount of anode slimes led to a careful consideration of the various methods of slimes treatment. A critical review of these methods is presented here in the belief that it would be of interest to copper refiners. No such review has yet been published, although a number of individual flowsheets have been described. To make the information of greater value, a questionnaire was sent to all copper refineries. While the overall response to the questionnaire was not entirely satisfactory, the foreign refiners gave much detail in their answers and expressed willingness to cooperate further if needed. Although the data below do not cover all refineries, it is thought that examples of all flowsheets currently in use are included. A comparison of these flowsheets should prove of interest. The paper is divided into two parts. One part deals with a discussion of all known methods of treating and recovering the various components of the slimes. The second describes the processes now in use at various copper refineries for the treatment of raw slimes up to the production of Doré bullion. A brief description of the two main electrolytic methods of parting Doré bullion is appended. Since the first commercial application of electrolytic refining of copper by J. B. Elkington in 1865, the treatment of anode mud or slimes has undergone many changes. The object of treating slimes has always been the recovery and separation of precious metals. With the passage of time, increasing attention has been paid to improvement in recovery and quality of byproducts, mainly selenium and tellurium. The original methods of treatment were crude, although relatively direct. For example," the slimes were fused on a dolomitic clay bed in a reverbera-tory furnace; the slag was removed and the resulting metal cupelled to silver and gold. In another process, the dried slimes were oxidized and scorified on the lead bath of a cupelling furnace. Later, these pyrometallurgical methods were modified by sulphuric acid leaching, followed by fusion with niter in a reverb eratory furnace. A preliminary oxidizing roast improved the leaching step. Straight hydro-metallurgy has not found practical application. No matter what method is employed, the result is Dorb metal—a silver-gold alloy, with or without the metals of the platinum group. Characteristics of the Slimes Anode slimes are made up of those components of the anodes which are not soluble in the electrolyte. They contain varying quantities of copper, silver, gold, sulphur, selenium, tellurium, lead, arsenic, antimony, nickel, iron, silica, etc. The color of raw slimes is grayish black and the particle size is —200 mesh. However, Baraboshkin and Gaev (Ref. 1, p. 37), describe light gray colored slimes from the electrolysis of secondary copper which are composed predominantly of sulphates and of basic salts. In anode furnace practice primary and secondary metals are not usually segregated and such slimes do not commonly occur. Copper in the slimes originates to a large extent in the cuprous oxide of the anodes and is therefore dependent on the oxygen content of the latter: Cu2O + H2SO4 + CuSO4 + H2O + Cu [I] Opinions have been expressed that some of the copper is formed by decomposition of cuprous to cupric sulphate Cu2SO4 ? CuSO4 + Cu PI The product of either reaction is very finely divided. Depending on the slimes composition, much of the copper is combined with sulphur, selenium, tellurium, etc. The balance of the copper is in the free metallic state.

Jan 1, 1951

By R. M. Caves

High-purity dense tungsten coating is obtained by means of a modified de Boer-van Arkel iodide process using tungsten bromides. The all-glass reaction system is pumped, baked, and sealed (pinched-off) at about 10-7 mm Hg pressure. The bromides then formed in situ are decomposed, depositing tungsten on heated tungsten surfaces. Heating of the specimen is done inductively in a novel arrangement that avoids electrodes of glass-to-metal seals. The tungsten, deposited at specimen temperatures from 2000° to 2800°F and bulb (halide) temperatures of 225" to 725°F, was adherent and uniform, and ranged in thickness to 1/8-in. REFRACTORY metals are of interest at NASA for use in hypersonic aircraft, in rocket engines, and in other thrust devices where very high temperatures are encountered. Tungsten, with a melting point about 6100° F, is potentially very useful for some of these applications. Considerable research on tungsten and its alloys is currently being conducted at the NASA Lewis Research Center. In this tungsten research program, there is need for a method of depositing high-purity, nonporous tungsten coatirlgs. Metallurgical bonding of the coating both to itself and to the substrate is required. In addition, it is necessary to be able to deposit thin, uniform coatings on substrates of a variety of shapes and sizes for laboratory test purposes. A survey of the methods previously used for depositing tungsten coatings indicated that none would likely meet these requirements for high purity, good uniformity, and high density except thermal decomposition of the halides. The method for coating that was selected for investigation was the thermal decomposition of tungsten bromide. This is a variation of the van Arkel and de Boer iodide process established years ago and recently reviewed.1,2 This process is applicable to the transition metals in general. More recently, the application and techniques were extended with chlorine3 and fluorine,4 and included the coating of tungsten. This is believed to be the first report on the use of the bromide for tungsten coating. Bromine was chosen over the other halogens because it promised a lower temperature for tungsten

Jan 1, 1962

By R. E. Steiner, A. F. Weinberg, R. J. Van Thyne

Uranium solubility in molten bismuth was determined in the temperature range 350" to 600°C, varying from 0.09 wt pct to near 2 wl pct, respectively. Zirconium and magnesium, simulated fission products, and sodium were found to increase or decrease uranium solubility in bismuth in the temperature range 370° to 420°C, depending upon the addition or temperature. Data were obtained by both chemical analysis of samples obtained by filtration from saturated melts and by measurement of the variation of electrical resistance with temperature. IN a study for the United States Atomic Energy Commission, a Liquid Metal Fuel Reactor Experiment (LMFRE) was proposed in which a solution of uranium dissolved in molten bismuth served as both the fuel solution and the primary coolant. It was proven necessary to add other elements to this solution for certain desirable properties: 1) zirconium acts to prevent reaction of the fuel solution with the graphite core (which would result in the formation of uranium carbide) and also to inhibit mass transfer of iron and chromium from the walls of the reactor tubing to cold spots resulting in blockage; and 2) magnesium acts as a deoxidant and prevents loss of uranium. In addition to these intentional additives, other elements entering the solution are: 1) fission products which are a by-productof reactor operation, 2) sodium (the secondary coolant) through possible small leaks in the heat exchange system, and 3) iron and chromium from the container walls. Each of these elements may affect the solubility of uranium in the fuel solution. Since the concept of this reactor is based upon a liquid solution, the relatively low solubility of uranium in bismuth will dictate the minimum temperature at which the reactor heat exchanger can be operated. The solubility of uranium in liquid bismuth and the effect of some additives on this solubility have pre- viously been investigated1-' (more recent results6" became avaliable after initiation of this investigation). However, some discrepancies between the various results have been noted, and also the cumulative effect of the several additions or impurities was not known. This investigation was commenced to obtain definitive results. MATERIALS Analyses of the materials used are presented in Table I. Although ostensibly of extreme purity (99.999 wt pct), the bismuth was found to contain a significant amount of impurities, probably oxides. Except for determination of the binary system, Bi-U, prior to use all bismuth was hydrogen-refined and

Jan 1, 1962

By R. W. Mancantelli, J. R. Woodward

IN 1947 a large low grade deposit of uranium was located in the northwest corner of Saskatchewan, in the Beaverlodge property of Eldorado Mining & Refining Ltd. Most of the values occur as thin seams and as coatings on other minerals, and the ore, which is light and friable, is not amenable to the usual gravity methods of concentration. The acid leaching process developed to retreat the company's Port Radium tailings was also impractical, as the percentage of carbonates is high and the percentage of sulphides relatively low. Construction of the mill building and installation of equipment, both scheduled for 1952, awaited selection of a satisfactory ore dressing process on or before Oct. 1, 1951. An extensive research program on dressing of ores from the Ace mine therefore was undertaken by the Mines Branch in Ottawa, a research team at the University of British Columbia, and an ore dressing group at the company's Port Hope refinery. Pilot plant operations were conducted at the Ottawa plant of Sherritt Gordon Mines Ltd. under the general direction of C. S. Parsons, consultant on metallurgy and ore dressing for Eldorado. In mid-September, having compared results of the several research programs, Dr. Parsons recommended a process employing a carbonate or basic leach. He reported that, while the chemistry of the process had been proved, engineering of the flowsheet was incomplete, and he suggested that under normal conditions further pilot plant work at the Ace mine might be profitable. However, since this would involve a delay of 12 to 18 months in bringing the property into operation, he recommended that design of the concentrator proceed immediately. Concentrator capacity was determined by two considerations: 1) the amount of ore then available and 2) the expectation that continued improvements would be made in the ore dressing process as a result of the intensive research being carried out by the Mines Branch of the Department of Mines and Technical Surveys. As these improvements could affect both the chemistry and the en-

Jan 1, 1956

By A. W. Schlechten, R. F. Doelling

Parkes' process crusts were vacuum distilled using a shortened Pidgeon retort. Zinc was effectively removed below 800°C and recovered as a zinc sheet easily stripped from the furnace liner. Lead dasdistillationazinc required about 950°C and the resulting lead condensate tillationtended to stick to the thin metal liner. Final purity of residue is limited only by nonvolatile metals such as copper. THE successful commercial application of vacuum distillation for the removal of zinc from lead, as described by W. T. Isbell,' raises the question of whether or not a similar method can be used for other metal separations. Preliminary laboratory experiments on the vacuum distillation of Parkes' crusts by one of the authors' gave such encouraging results that it was decided to make a series of tests with large scale equipment to determine the results that might be expected in plant operation. The pilot-plant test results are reported in this paper. The great advantages of vacuum dezincing are its simplicity, the relatively low temperature of operation, and the fact that the zinc is recovered in the metallic state ready for use again in the desilveriza-tion of lead bullion. These same advantages and others would be realized from the vacuum distillation of the zinc-lead-silver crusts made in the Parkes' process. Zinc readily will distill over at temperatures of 800" to 900 °C as contrasted with the 1200°C or more used in the Faber du Faur furnace. Furthermore, the recovery of zinc as metal is practically perfect since there is no loss by oxidation or fuming. If the vacuum distillation is continued and hastened by raising the temperature to around 1000°C, the lead will distill over at an appreciable rate and deposit in a warmer zone of the condenser. This lead will contain a small amount of silver, but it can be returned to the desilverizing kettles and there is not the chance of loss that there is in disposing of the rich litharge from cupelling. The lower temperatures required will mean a saving of fuel and less wear on the equipment. One further point is that the vacuum distillation could be run on a small scale practically continuously, so as to avoid the accumulation of large amounts of silver-rich material and the need of assembling a large crew for special cupellation runs. It should be pointed out here that the final purity of the silver residue will depend on the copper or iron content of the crusts, as these and other metals with low vapor pressure will not be eliminated. The idea of treating the silver-zinc-lead crusts from the desilverization of lead bullion by vacuum distillation is not a new one, but such experiments have not been reported, as far as can be determined. In 1912, T. Turner presented a paper3 before the Institute of Metals on the vacuum distillation of brass, and a member of the audience suggested that he try his method on Parkes' crusts. In 1935, W. J. Kroll, noted for his discovery of debismuthiz-ing lead by calcium additions, published a paper' on vacuum distillation of metals and pointed out possible applications in the lead industry. Principles of Separation The successful separation of two or more metals by distillation depends upon an appreciable difference in the effective vapor pressures of the metals in the alloy to be treated. The vapor pressures of the common metals have been determined with considerable accuracy; for example, at 800°C the vapor pressures of the individual metals which are of particular interest in this paper are about .as follows: Zn, 241 mm; Pb, 0.0553 mm; and Ag, 0.0000276 mm. Unfortunately,, data are not available on the activities of the components in the ternary system Zn-Pb-Ag, but it can be estimated that on the basis of the vapor pressures of the individual metals a good separation can be made by distillation. The preliminary laboratory tests previously mentioned' supported this assumption. The pilot-plant size vacuum distilling furnace is shown in Fig. 1. It consisted of an abbreviated Pidgeon retort, 5 ft long with an ID of 8 in., heated by six silicon carbide elements placed symmetrically

Jan 1, 1952

By F. Habashi

"A short account will be presented on the extraction of rare earths from monazite sand, bastnasite ore, and phosphate rock of igneous origin. This includes mineral beneficiation, leaching methods, fractional crystallization [of historical interest], ion exchange, solvent extraction, precipitation from solution, and reduction to metals. INTRODUCTIONOriginally, the term rare earths was only used for the oxides, R2O3, which are similar to one other in their chemical and physical properties and are therefore difficult to separate. Within the rare earth group, the elements scandium, yttrium, and lanthanum differ in their atomic structure from the elements cerium to lutetium (the lanthanides, Ln). Scandium occupies a special position with respect to this classification and its other properties, and therefore does not belong to either of these groups. The rare earth elements always occur in nature in association with each other. The isolation of groups of rare earth elements or of individual elements requires costly separation and fractionation processes owing to the great similarity of the chemical and physical properties of their compounds, which explains why the history of their discovery has extended over such a long period.The word “rare”, when used to describe this group of elements, originates from the fact it was thought that these elements could only be isolated from very rare minerals. Considering their abundance in the Earth’s crust, the term rare is now inappropriate. These elements are lithophilic and are therefore concentrated in oxidic compounds such as carbonates, silicates, titano-tantalo-niobates, and phosphates.The abundance of the rare earth elements taken together is quite considerable. Cerium, the most common rare earth, is more abundant than cobalt. Yttrium is more abundant than lead, whereas Lu and Tm are as abundant as Sb, Hg, Bi, and Ag. For all practical purposes promethium does not occur in nature. It forms only in nuclear reactors."

Jan 1, 2012

By C. D. Anderson

A variety of processing technologies exist for recovery from both primary and secondary sources of rhenium. Currently, there are no known primary rhenium deposits, thus, the method in which primary rhenium is produced is dependent on the commodity of which it is a byproduct, e.g., copper, molybdenum, uranium, etc. In addition, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing. This paper presents a review of both primary and secondary processing technologies for the recovery of rhenium. Key Words: Rhenium, Ammonium Perrhenate, Extractive Metallurgy

Feb 27, 2013

By C. D. Anderson

A variety of processing technologies exist for recovery from both primary and secondary sources of rhenium. Currently, there are no known primary rhenium deposits; thus, the method in which primary rhenium is produced is dependent on the commodity of which it is a byproduct, e.g., copper, molybdenum, uranium, etc. In addition, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing. This paper presents a review of both primary and secondary processing technologies for the recovery of rhenium.

By Wenming Wang, Edgar E. Vidal, Scott A. Shuey, Patrick R. Taylor, Judith C. Gomez

Introduction In modern society, vanadium is defined as a strategic metal. Some 90% of the global consumption is used as an alloying agent for carbon steels, tool steels, and high-strength low-alloy steels (particularly for pipelines). Advanced titanium-aluminum-vanadium alloys are seeing application in the aerospace industry. Small but ever growing amounts are finding application as catalysts or in electronics. New uses are continually being discovered for this metal. One example of a new application is the vanadium-redox battery for use in generation plants and back-up power sources. Vanadium is found in a large number of minerals, of which the most important are carnotite, roscoelite, vanadinite, mottramite, and patronite. Just over a third of the world's vanadium is produced as a primary product; the balance is a by-product of the iron and steel, oil refining, power generation and uranium enrichment industries. Vanadium can also be sourced by recycling spent catalysts from the petrochemical industry and ash produced by the combustion of oil emulsion in power stations. However, recycling activities are believed to account for only a small amount of total world supply. Vanadium is recovered from these ores largely as the pentoxide (V2O5) [1]. The major sources of vanadium-bearing magnetite ores are scattered throughout Australia, China, Russia, and South Africa. Current estimates of world’s Proven & Probable vanadium reserves are of the order of 41.3 million tones [2, 3]. Commercial vanadium-bearing titaniferous magnetite deposits can be found in Kachkanar, Russia; Pan Zhihua, China; Bushveld, South Africa; and Windimurra, Western Australia [4]. Other deposits have been found in New Zealand, Canada, and India, among other countries. Vanadium consumption in the steel sector has increased some 7% per ton of steel over the last 10 years as new alloying applications have been found to improve steel's strength-to-weight ratio. Future demand from the aerospace and nuclear industries for non-ferrous vanadium alloys such as titanium and super-alloys is also likely to grow. Vanadium is usually added in the form of ferrovanadium, a vanadium-iron alloy. Vanadium compounds, especially the pentoxide form, are used in the ceramics, glass, and dye industries, and are also important as catalysts in the chemical industry. While established approaches may be used at an industrial scale to treat vanadium-bearing titaniferous magnetite ores in several countries, new alternatives are being developed worldwide. This review will give short overview of the treatment of vanadium-bearing titaniferous magnetite ores, an overview of direct reduction and direct smelting processes and an approach to determine technical alternatives for the property.

Jan 1, 2005

By D. R. Spink

An historical background is presented to provide a basis for the lack of normally expected developmental effort over the first 20-25 years of viable zirconium production operations. This is followed by a description of developing technology, which is the result of years of research and piloting effort directed to the development of a more economical method to produce high-grade metallic zirconium. It is predicted that many of the process improvements described will have a significant impact on future zirconium production methods and consequently on future prices and markets.

Jan 1, 1977

By R. E. Allen, A. C. Jephson

ELECTROLYTIC zinc plants of the American Smelting and Refining Co. are located adjacent to the present city limits of Corpus Christi, Texas. The original plant commenced operations during 1942, and is designed for the production of special high grade zinc from zinc sulfide concentrates.' Subsequent changes include the erection of an additional acid plant during 1950, and additions for increased generating and production capacity during 1951. New Electrolytic Plant—The new electrolytic plant is a separate unit erected for the treatment of densified fume produced at Asarco plants located in El Paso, Texas and Chihuahua, Mexico. Construction was completed and operation commenced during June, 1953. Decision for the erection of a separate plant for the treatment of densified fume is based on several factors. One of the principal reasons is the assurance of continued production rate of the original plant, which might have been jeopardized by combined treatment of calcine and fume. Of equal importance is the advantage of locating a plant on a site with adequate space for eventual expansion. General Description—The new plant for the treatment of densified fume consists of four main buildings for the following operations: storage and grinding of fume, leaching and purification, electrolyzing and casting, and rectifier units. Auxiliary buildings include a baghouse, reagent storage, and a lunch room. Leaching of ground fume with electrolyte from the cells is a batch type operation in tanks equipped with mechanical agitators and weightometers. Completed leaches are pressure filtered, the resultant residue being recovered for shipment to El Paso by conventional use of thickeners, drum type vacuum filters, and a direct-fired dryer. Filtrates from the filter section are purified in mechanically agitated tanks, filtered after completion through plate and frame presses, and then pumped to an evaporative type cooling tower. Solution discharged from the basin of the cooling tower is collected in storage tanks for use in the cells. The cool purified solutions are pumped as required from storage to the solution circuit for circulation of electrolyte through the cells. Except for the addition of purified feed solution and withdrawal of electrolyte for leaching, the circulating system for the cells is essentially a closed continuous circuit that includes the use of evaporative cooling towers for control of solution temperatures. Aluminum cathodes and Ag-Pb anodes are used for electrolysis, which is maintained at a current density of about 82 amp per sq ft. Cathode zinc produced is melted in a gas-fired reverberatory furnace and then cast into slabs on a straight line casting machine. Unloading and Storage—Fume produced by the Chihuahua plant is shipped in box cars, and fume by the El Paso plant in hopper bottom dump cars. Shipments as received are weighed on a 7 5-ton capacity railroad scale, sampled, and then moved over one of the two unloading hoppers with a car puller. The cars are then unloaded with mechanized equipment, a car scoop being used for unloading of box cars and a car shaker for the hopper bottom cars

Jan 1, 1958

By J. F. Mitchell, R. Bainbridge, E. A. Melvin

IN the fall of 1953, The Consolidated Mining and Smelting Co. of Canada Ltd. put into operation a completely new and modern plant for sintering the rather complex assortment of materials which comprise the feed to the Lead Smelter at Trail, British Columbia. This Lead Smelter is one section of an integrated metallurgical plant producing lead, zinc, silver, gold, cadmium and other metals. Since most of the incoming lead concentrate contains appreci- able zinc values, and most of the zinc concentrate similarly contains appreciable lead values, the cross-routing of residues or tails from the respective lead and zinc plants offers obvious advantages in recovery of metals, and has been practiced for many years. A disparity has existed, however, between the output of zinc plant residue and the capacity of the Lead Smelter to treat it, with the result that over the years a stock-pile of residue had been accumulating, and by the end of World War II had reached the formidable total of half a million tons. To treat this stock-pile, in addition to all current zinc plant residues and, of course, the regular quota of lead concentrates, became a prime objective of the Lead Smelter following World War 11. As zinc plant residue plays an important part in the operations to be discussed in this paper, some consideration of its

Jan 1, 1958

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958

By L. M. Pidgeon, P. G. Thornhill

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. lengtn of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of ¼ in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10o to 15oC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press Operating at a temperature of 140°C

Jan 1, 1958

By W. A. Krivsky, R. Schuhmann

PREVIOUS papers in this series on the thermo-•t dynamics of copper-smelting systems have presented a survey of the field and an outline of the overall program: thermodynamic studies of iron silicate slags, 2-4 and thermodynamic data for simple mattes in the Cu-S system.5 Since this program was started, an excellent summary of the whole field of physical chemistry of copper smelting was published by Ruddle. The present paper represents a further step in the study of liquid mattes, and deals specifically with the Cu-Fe-S system. The importance of the Cu-Fe-S system to copper smelting is evident from the fact that these three elements constitute 90 to 99 pct of the totalweight of most commercial mattes. In copper mattes oxygen is an important fourth component. Oxygen was rigorously excluded in the work to be described in this paper. However, the behavior of oxygen in mattes was the subject of a companion research by Lander' which will be described in a subsequent publication. Several studies have been made previously of phase equilibria in the Cu-Fe-S system,8-10 so that the ternary phase diagram is fairly well established in most regions. Also, separate studies have been made of the phase diagrams and thermodynamic properties for the three binary systems: CU-S,5,12-15 CU-Fe,11,14,17 and FeS.11,14,16,17 However, no work has

Jan 1, 1958

By Ramin Motamed, Joseph A. W. Toth

Evaluation of the capacity of large-diameter drilled shafts using current load test methods present challenges when considering the magnitude of loads required to fully mobilize their capacities. Often, these full scale tests do not fully mobilize the capacity through side resistance. Therefore, the need for innovative solutions to extrapolate side resistance and interpret nominal capacity are of key interest. In this study, we evaluate three methodologies for extrapolating unit side resistance based on bi-directional load test data. Fellenius (2001) presented these methods for interpretation of pile capacity using static load tests. Our interpretations are based on the evaluation of 116 load tests. Unit side resistance was calculated and reviewed for each load test and categorized using three criteria of mobilization; (A) full, (B) partial and (C) linear. This study specifically focuses on load tests that showed fully mobilized unit side resistance behavior (Criteria A). The three methodologies were applied assuming Criteria (A) load tests were mobilized to 50% of their measured values. Subsequent comparisons of each extrapolation were then compared against the full measured values for each test and generally showed good agreement. INTRODUCTION It is widely accepted that conducting load tests such as standard top-down tests or bi-directional (i.e. O- cell) tests can help designers understand load-settlement behavior and the nominal resistance of drilled shaft foundations. These load testing methods provide the advantage of confirming design assumptions as well as improving the reliability of the foundation design. The potential cost savings from the application of these methods for mid to large size projects is self-evident. For example, in a recent project in Northern Nevada, an O-cell test on a 6.9 ft diameter, 114.8 ft length drilled shaft imposed an initial cost of $125k to the project, resulting in a 15% reduction on the total foundation cost; equivalent to $1.3 million in overall savings (Pease, 2014). However, the costs of implementing such methods on smaller projects can be prohibitively expensive (Hagerty et al, 2016). Despite the clear advantages of drilled shaft load testing, the current practice is not without limitations. Performing load tests on large-diameter drilled shafts with high capacities is a challenge considering the significant loads required to fully mobilize the capacity of such foundations. In some cases, the feasibility of these tests is impractical because the associated initial cost is not guaranteed to outweigh the final savings. As a result, large scale load tests are sometimes ruled out from the design phase of projects. Furthermore, in the majority of load tests the magnitude of loads applied are specifically selected to not approach failure, so the shafts can be used in the future foundation system (i.e. production shafts). This practice of field testing restricts our knowledge of the true capacity of the drilled shaft at failure; hence the nominal resistance remains unknown.

Jan 1, 2018

By J. Drimmer

The past several decades have seen the rapid expansion of multinational mining companies? exploration and production activities around the globe. This trend has resulted from a number of factors, including increased demand for resources as a result of worldwide economic growth, the diminishment of known and proven reserves, rising commodity prices, technology improvements, and the liberalization of investment regimes in developing countries with significant natural resources. With those opportunities, however, come associated risks and significantly increased legal compliance obligations that may be imposed by various bodies. That includes domestic regulations, as well as foreign civil and criminal laws with extraterritorial reach. A failure to adhere to those laws can have dire consequences, including substantial monetary fines and penalties, severe damage to a company?s reputation, and even jail time for individual company employees.

Jan 1, 2010

By Gabrielle Luana Jimenez

Aparelhos eletroeletrônicos como computadores e televisões são constituídos por telas de cristal líquido, que ao se tornarem obsoletos são classificados como resíduos, e dispostos em aterros sanitários. Entretanto, na composição das telas de cristal líquido há metais de valor agregado, como o índio (In). Com a fundamentação no processo de reciclagem de telas de LCD, este estudo propõe a recuperação e quantificação desse metal que é encontrado na forma de ITO (óxido de índio e estanho). Para tal feito, foram empregadas rotas hidrometalúrgicas, tendo como principal etapa a lixiviação. Como parâmetros de controle, os ensaios de extração sólido-líquido foram realizados em temperaturas que variam de 50 a 80°C. Com isso, foi obtido resultado de extração do índio em cerca de 60% em massa, nas condições de 60°C e após 4 horas de lixiviação.

Aug 17, 2017

By Alexandre Silva Guimarães

A extração por solventes sinérgica dos metais cálcio, magnésio e níquel, presentes em soluções aquosas sulfúricas em níveis de concentração similares aos dos licores industriais HPAL (High Pressure Acid Leaching), foi avaliada neste trabalho empregando-se os extratantes comerciais Cyanex 272 e Cyanex 301. O uso de Cyanex 301 (10%v/v) sozinho não se mostrou eficiente, em decorrência do baixo percentual de extração e da baixa seletividade entre os metais. Quando misturado ao Cyanex 272, porém, observou-se a possibilidade de remoção seletiva de magnésio, deixando cálcio e níquel em solução. A melhor condição operacional foi obtida utilizando-se 5-10%v/v de Cyanex 301 e 20% de Cyanex 272, na faixa de pH entre 5,4-5,7, na qual se extraiu aproximadamente 70-80% de magnésio, 15-35% de cálcio e 10-15% de níquel.

Aug 16, 2017