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By Liuyin Shi, Kang Liu, Fengqi Zhao, Yan Song, Zhiping Yang, Xing Fan

Due to the high energy consumption, long technical process, low usage rate of vanadium during the traditional smelting vanadium steel process, a new method for decarburization by CO2 to directly smelt vanadium-containing steel in blast furnace was proposed. In this paper, the thermodynamics of CO2 decarburization in molten iron process was calculated out by FactSage. The thermodynamic calculations indicated that in standard state, the transformation temperature of C and V was 1357 °C; the oxidation order of element was Si, C, V, Mn and S when using CO2 as oxidant. The experiment of blowing CO2 temperature showed that decarburization reaction temperature was 1550 °C, carbon content could be decreased to 0.676%, the decarburization rate was 88%. There was no effect on vanadium content in molten iron during the decarburization process, vanadium content was 0.323%, which meets the request of vanadium mould steel.

Mar 1, 2018

By A. Dadgar

Extraction of gold and silver from two refractory concentrates using cyanide and bromine reagents was reported at the 12th International Precious Metal Conference (1). The objective of this paper is to present the gold extraction, recovery, and economics for two refractory concentrates using cyanide and bromine reagents. Gold extractions for cyanide leaching (24- 48 hours) and bromine leaching (6 hours) were the same and ranged from 94% to 96%. Gold recoveries from bromine pregnant solutions using carbon adsorption, ion exchange, solvent extraction, and zinc and aluminum precipitation methods, were better than 99.9%. A preliminary economic analysis indicates that chemical costs? for cyanidation and bromine process are $9.50 and $10.50, respectively per ton of calcine processed.

Jan 1, 1989

By P. L. Crouse, S. J. Lubbe, M. J. Kabangu

A selective extraction and separation of zirconium by solvent extraction with 1-octanol is proposed. The process is based on the dissolution of the metal basic carbonates in hydrochloric acid in the presence of potassium fluoride, using 1-octanol as extractant. The effect of several process parameters, including hydrochloric acid and potassium fluoride concentrations, reaction time, and phase ratio between the aqueous and the organic phase was investigated. Zirconium is enriched in the organic phase. The best extraction and separation results were obtained with the basic carbonate dissolved in 10% hydrochloric acid and 1.5 M potassium fluoride. The optimum reaction time was 15 min with an organic to aqueous phase ratio of 2:1. Based on the selectivity of zirconium over hafnium in the organic phase, a McCabe-Thiele diagram was constructed. A near complete stripping of zirconium from the organic phase was achieved using a 3 M sulphuric acid solution. Additional studies are required in order to determine the reaction mechanism and the chemical speciation of the current investigation.

By T. Tamagawa, N. S. Fu, I. Iwasaki, M. Kobayashi

A 3-stage fluidizing bed process of chlorination, selective oxidation and sulfur chloride conversion was tested for processing of a chalcopyrite concentrate. It was found that the use of excess oxygen was essential to a stable operation in the oxidation reactor. The oxygen-excess gas will inevitably cause the oxidation of the sulfide concentrate in the chlorination and sulfur chloride conversion reactors. A new flowsheet is proposed to separate the chlorination and the selective oxidation steps with separate flow paths of the fluidizing gases, thereby preventing the passage of oxygen-excess gas in the chlorination and sulfur chloride conversion reactors. The feasibility of the proposed flowsheet was demonstrated in a laboratory-scale fluidizing bed setup.

Jan 1, 2000

By R. A. Pyper, J. L. Hendrix

Laboratory leaching experiments were conducted to ascertain the effectiveness of acidic solutions of thiourea, CS(NH2)2, for dissolving gold from finely disseminated gold ores. The ore tested was from the Chipmunk ore body of the Northumberland mining district in Nye County, Nevada, USA. Extractions of up to 90% were attained from the ore. Extraction was shown to be dependent on thiourea concentration, ferric sulfate oxidant concentration, acid concentration, and leaching time. Optimum conditions were an ambient temperature leach near pH 1.0 utilizing 100 to 200 kg/ton thiourea and 50 kg/ton ferric sulfate for 8 to 24 hours.

Jan 1, 1981

By Peter Fischer

Reducing mining costs has become such an important factor in the German coal industry that operators are forced to work only those areas where production costs can be kept low. These areas will normally be located close to the mine shafts because they are easily accessible. Even the shaft pillar is now being considered as part of the winning zone. Before working coal in shaft pillar zones it is essential to assess the impact this will have on the shaft lining. There is a huge difference between sliding and strata-anchored shaft systems. However, if the overburden comprises lightweight soil both types of shaft will have a tendency to protrude from the surrounding ground. This paper presents the geomechanical reasons underlying this reaction and the methods used for predicting and measuring it.

Jan 1, 2007

By K. Osseo-Asare

LIX63 in combination with DNNS selectively extracts nickel and/or cobalt from acidic leach liquors with pH values as low as 1.0. Nickel/cobalt and cobalt/nickel selectivities are achieved by varying the relative concentrations of LIX63 and DNNS. Increasing LIX63 concentration retards both extraction and strip-ping rates while DNNS has the opposite effect. Slope analyses of equilibrium data indicate that both LIX63 and DNNS are present in the extracted complexes. Examples of possible flowsheets are presented.

Jan 1, 1978

By G. Adachi, K. Machida, T. Ozaki, K. Murase

The chlorination rate with CCl4 was unexpectedly slower than carbochlorination using chlorine under the same reaction conditions. Chlorination with CCl4 was topochemical and the amount of monazite reacted after different periods suggested that the kinetic control step was the reaction between CCl4 and the monazite core at the interface. The kinetic order of the reaction was 0.5. The presence of a complex former, AlCl3, promoted the chemical vapour transport of rare-earth chlorides as complexes

Jun 18, 1905

By T. Ogasawara, J. P. Barbosa, A. J. Monhemius, F. T. Da Silvia

A thermodynamic analysis of the sulphation of rare-earth phosphates and the water-leaching of rare-earth sulphates was used to guide experimental work on the sulphation and leaching of concentrates from Pitinga. It was concluded that the sulphuric acid digestion of xenotime is technically feasible and could be used industrially to produce yttrium and other rare-earth oxides. Yttrium extractions of more than 90% were obtained by digestion at 503 K for 4 h at an acid:concentrate ratio of 2:1 with particle size -0.053 mm. The apparent activation energy at temperatures less than 473 K, as determined from the Arrhenius plot, was 11.0 kJ/mol, indicating that the reaction is controlled by the diffusion of the reactants or products through the sulphate layer

Jun 19, 1905

By Theophile S. Yameogo

Les ressources naturelles ont en général été le moteur du développement économique et social des nations. Les pays aujourd?hui développés sont pour la plupart d?anciens eldorado, charbonniers ou pétroliers. D?autres le sont toujours. L?Afrique fait cependant figure d?exception : même en étant comblé de ressources naturelles exceptionnelles, les pays de ce continent croulent sous le poids de la pauvreté et de ses acabits que sont les maladies, la famine, la guerre et l?insurrection sociale. On reconnaît néanmoins qu?il y a des cas particuliers comme le Botswana, la Namibie, l?Afrique du Sud. Face à cette situation, beaucoup de scénarii ont été proposés sans pour autant porté les fruits espérés. Le rapport Brundtland suggère que le développement durable est la clé du succès car elle veille à la prospérité économique, à l?équité dans la société et à la pérennité dans l?environnement. Pour l?Afrique, beaucoup de facteurs doivent être considérés et les solutions se doivent d?être pratiques et éthiques. Le futur ingénieur des mines qui a écrit ce rapport, apporte sa vision de la situation non sans avoir auparavant analysé les faits. Cette démarche, qui se veut avant tout éthique, inclut les aspects sociaux, économiques et politiques. Introduction In the 40s, 50s and the 60s, many economists, among which Harold Hotelling from South Carolina University, thought that developing countries could reach socio-economical success by extracting their natural resources such as oil and minerals. The current economical disaster seems to have contradicted that belief and critics like Davis Graham (Colorado School of Mines) and Michael Ross (University of California at Los Angeles) even suggested that extractive activities stop in Africa. But the experts of the World Bank (Mining Department) and some mining and oil companies are committed to reverse the tendency by implementing a new approach. Considering the principles of sustainable development defined in the Brundtland report (1987), we decided to analyze the possibility of giving the extractive industry a catalytic role in creating a socially, economically and environmentally development in Africa. In other words, can we continue to mine and extract oil in Africa for its blossoming future under the ruling of the sustainable development? In the following, we start by explaining the African socio-economic and political context. After a brief definition of the sustainable development, we explain why, for Africa, some believe in the extractive industry and some don?t. In the end, we draw out conciliatory arguments.

Apr 1, 2005

By R. Schuhmann

Available thermodynamic data applicable to copper smelting systems are collected and tabulated, and the important gaps are pointed out. A few examples are given of estimations which can be made from the available data. An experimental research program is proposed to supply the thermodynamic data that appear most essential to better quantitative understanding of the chemistry of copper smelting. The proposed program is designed also to shed specific light on the practical problems of slag losses and magnetite behavior. OPPER smelting, from flotation concentrates to V-/ blister copper, is conspicuous among the large scale chemical processes which are conducted with only relatively incomplete knowledge of the physical chemistry involved. For example, no common metallurgy text explains adequately why copper enters the matte phase while iron enters only to the extent that sulphur is left over after satisfying the copper. Explaining this important phenomenon as a manifestation of the greater affinity of sulphur for copper than for iron is unsound because the affinities of copper and iron for sulphur are about the same at smelting temperatures. As will be shown, other affinities are really decisive in the relatively clear-cut separation of copper in the matte. In contrast, thermodynamic studies have contributed much to our understanding of copper refining, zinc oxide reduction, magnesium production, steelmaking, etc. For each of these processes, the important reactions are clearly recognized, and fair to good quantitative values of the free-energy changes and equilibrium constants are available. Such data have proved to be of constantly increasing practical value in process development and improvement. Work in other fields has furnished much thermodynamic data applicable to copper smelting systems. In fact, several promising starts have been made in applying such data to specific smelting problems. Kelleyl made an appraisal of the possibilities of recovering elementary sulphur from low grade matte. His calculations and compilations of thermodynamic data have represented the starting point for much of the work in this field, including the present survey. Huang and Hayward2 nd Aksoy3 used thermodynamic methods in the study of copper losses in reverberatory slags. Peretti4 used thermodynamic data in explaining the chemistry of converting. The recent publications of Darken and Gurry"." and of Darken,' dealing with the iron-oxygen and iron-silicon-oxygen systems, respectively, present equilibrium data that are applicable to copper smelting systems. Also additional data were reported recently on the affinity of sulphur for copper, manganese, and iron8 and on the sulphur pressures of iron-sulphur melts." The survey presented in this paper was made as the basis for planning an experimental program on the thermodynamics of copper smelting. Few researches on the chemistry of copper smelting have been reported in recent years, so that a reappraisal and coordination of old and new data are essential if further work is to take the directions of maximum value. The experimental program is now in progress, and the plan of attack is outlined at the end of this paper. Progressive Oxidation and Desulphurization of Copper-bearing Liquid Phases: The chemical activities of sulphur and oxygen are two of the most important thermodynamic yardsticks to be applied to copper smelting processes. Virtually the entire smelting and refining sequence involves a series of systems characterized by decreasing sulphur activity and increasing oxygen activity. In this section, therefore, an attempt is made to define and explain these activities in terms of equilibrium partial pressures of SO2, s2, and O,. Also, estimates of these quantities are presented, the estimates being based largely on calculations presented in a later section of this paper. The sequence of steps from raw flotation concentrate to fully oxidized copper ready for poling involves progressive and controlled oxidation. Iron is oxidized and enters the slag. Sulphur is oxidized and leaves in the gas. Table I summarizes several important features of this oxidation and desulphurization sequence, starting at the beginning of the matte blow in the converter. The top line gives in order the principal smelting and refining stages up to fully oxidized copper, plus the additional step, not used commercially, of oxidizing all the way to Cu,O. In the second line are shown the principal copper-bearing liquid phases which characterize the process. Through most of the sequence the copper

Jan 1, 1951

By A. W. Schlechten, R. P. Abendroth

Jan 1, 1960

By T. R. A. Davey

The fundamental principles by which bismuth may be removed from lead by pyrometallurgical processes are enumerated. Qualitative discussion of the phase diagrams concerned is followed by presentation of quantitative diagrams. Brief mention is made of the practical aspects. Data presented show how chemical lead (<0.005 pct Bi) may be produced by the Jollivet, Dittmer, and Kroll-Betterton processes. BISMUTH is an element whose properties are very similar to those of lead, and its minerals are associated to a greater or lesser extent with nearly all the major lead ore deposits of the world. In consequence, the lead produced contains bismuth unless a special debismuthizing process is practiced. There does not appear to be any published comprehensive review of the effect of bismuth as an impurity on the properties of lead, and there is some uncertainty as to the beneficial or detrimental effects of small amounts of bismuth in lead used for particular applications. Standard specifications of most countries stipulate a maximum bismuth content of 0.005 pct for chemical lead. With one exception, this specification is met only by lead producers whose ores are relatively bismuth-free, or whose refineries are electrolytic. In lead for certain applications a higher bismuth content is preferred by some fabricators.&apos; In contrast to the refining for silver, copper, and antimony, the refining for bismuth does not usually result in a large overalil economic gain from the sale of byproduct. Thus, debismuthizing is customarily practiced only if market conditions demand a lead with a bismuth content lower than that produced by classical refining; i.e., softening, desilverizing, and zinc refining. Until this century, debismuthizing was carried out only a:; an incidental result of desilverizing, as later described. This century has seen the development of the Betts electrolytic process—the only successful one of many electrolytic processes proposed and patented —and a number of precipitation processes based on the pioneering work of Kroll,&apos; although only one of these, the Kroll-Better ton process,&apos; is in wide commercial use today. It is proposed in this paper to deal only with the pyrometallurgical processes, excluding those based on aqueous electrolysis. Historical Background Prior to 1833 silver (and incidentally bismuth) could not be removed from lead. To recover silver from rich bullion, the only course available to lead refineries was to cupel the bullion, forming a slag of litharge. The first slag tappings or skimmings, on reduction, yielded soft lead of very low silver and bismuth contents. Intermediate skimmings yielded a lead of somewhat higher silver and bismuth contents, which could if desired be recupelled. The last skimmings yielded a litharge very high in bismuth, which could by successive oxidation and reduction cycles be separated into high purity bismuth, and lead returns. This oxidation reduction process, with modern operating techniques, can still be used today to produce a portion of a lead refinery&apos;s output of very low bismuth content, as discussed later. In 1833 Pattinsone discovered that lead bullion could be purified of silver and bismuth by fractional crystallization, the primary lead crystals formed by partial freezing being purer than the melt. Repeated processing in pans produced a lead containing as little as 0.02 pct Bi. The simple explanation of this

Jan 1, 1957

By W. A. Krivsky, R. Schuhmann

Jan 1, 1962

By G. E. Johnson

E. D. HYMAN*—How much sorting of scrap is done ? G. E. JOHNSON (author&apos;s reply)—We do practically no sorting. We charge "run of mine" scrap to the furnace. The unmeltables, mostly iron, are in such demand today that there is no difficulty in disposing of them. It may soon be desirable to sort out from the unmeltables as much of the brass as possible. J. J. BRUGMAN†—We have somewhat similar problems in the secondary aluminum business. What is your method of removing the unmelted material from the furnaces? Why have you such an apparently small space in which to charge your materials? Do you find that you have to seal that opening, or can you have it open and continuously charge at one end and pull out the other ? G. E. JOHNSON—Our means of melting scrap is efficient only to the extent that we use the waste heat from the vaporizing chamber to do the job. It is a batch process. We open the charge door and place the scrap on the hearth by hand shoveling. The door is then closed during the melting down period. After the melting is complete, the opposite door is opened and the unmeltables are raked out. The doors are approximately 3½ X 4½ ft and are not sealed during the process only closed. They are nominally tight. Some metal is oxidized in the process. We have visualized a means of conveying the materials through this melting unit with the metals that are melted trickling out during its travel. T. H. WELDON‡—Mr. Johnson, in line with the last question, is it necessary to seal the furnace between the melting down and the vaporizing unit, or have you got an inverted syphon in the bottom of the chamber? G. E. JOHNSON—That was one of the first things we encountered. We had to have a sealed opening, and it is a molten metal seal. You have indirectly asked me another question, which was: "Do you have to seal up the melting unit?" I would say we should exclude as much air as possible, although we are not too efficient in doing that. We allow the melting unit doors to be open when we charge and when we remove unmeltables. You can readily see that that would lead to the idea of having a controlled atmosphere in the melting unit, and I think this would do a more efficient job of melting the scrap. T. H. WELDON—How often do you charge the furnace? G. E. JOHNSON—We charge the melting unit, and rake out the unmeltables, about every hour. H. R. HANLEY*—Are any provisions made for controlling the rate of oxidation for the production of various size particles for certain characteristics of the zinc oxide product? G. E. JOHNSON—Yes, there are many. You are getting pretty much into the fine points of zinc oxide manufacture. Some of us still think we have something to learn about that. In general, this muffle furnace as I have described it to you produces a rounded particle of zinc oxide which is generally formed by a rapid oxidation of the zinc vapor, followed by rapid cooling. We have gone to the other extreme in some of our experiments. We have changed the furnace to produce a type of zinc oxide, such as we thought was peculiar to American process zinc oxide, by controlling the temperature at the point of oxidation and maintaining that temperature for a much longer period of time than we do when we make the rounded shape. There are other relationships that this furnace readily provides. One of the important factors is the ratio of air to zinc vapors. We can vary that by varying the air supply to the baghouse, or vary the rate at which we are vaporizing the zinc by the simple expedient of regulating the temperature over the carborundum arch. We have a number of variables that permit us to produce all of the grades of French process zinc oxide from lead-free up through the highest grades of seal oxides. There are many controls that we can apply to the operation. What I have said is but a brief condensation. K. MORGAN*—Can Mr. Johnson give us some idea of the fuel consumption of the furnace ? How much oil does he use per ton of zinc distilled? I am also interested to know what sort of heat transmission he gets through the arch? What is the thickness of the tiles used to construct the arch? Some time ago we built a small furnace for a different purpose, using a carborundum arch, and we found that the reflectivity of the molten zinc surface was so great we had to use a very high arch temperature. We found we made an improvement by having a layer of carbon on the surface of the zinc. Has Mr. Johnson had any experience on these points ? Does he make any sort of insolu-bles which he leaves in the furnace which he cannot tap out ? G. E. JOHNSON—I believe your first question was the fuel consumption. If I recall, somewhere in this paper there is a test that I quote. I believe we used 800 gal in a given period of time. Offhand I cannot translate that into tons of metal. I might also state that we have this understanding; that the carborundum arch, as the temperature becomes higher, becomes more efficient in heat transmission. As a matter of fact, I believe it is at about 2600°F or higher that the highest efficiency of heat transmission becomes available. We have calculations that I cannot quote from memory which indicate that the carborundum arch does a really very fine job for this type of furnace. Another point that we had considered was the fact that this furnace could be readily constructed so as to furnish an ideal source of heat for waste heat boilers. If a carborundum arch was used over the melting unit, we would have no zinc vapors in the gases at all— just clean combustion gases from which we would have removed some of the heat. L. P. DAVIDSON†— The insolubles

Jan 1, 1950

By H. R. Page, A. E. Jr Lee

A. E. LEE, JR. (author)—In addition to the paper we should like to make a few remarks. First, the seriousness of bending of the clay retort cannot be overemphasized. Not only did bending limit the length of the vessel, but it permitted corrosive fluid slags to concentrate in the belly of the retort and thus increased the possibility of boring. Bending also practically prohibited the use of cleanout machines and increased the danger of excessive fume losses by allowing furnace gases to escape around the top of the retort mouth. It will be noted that the Blackwell pottery practice places considerable emphasis on careful drying of retorts. Perhaps it is another smelter superstition, but traditionally the aging of retorts beyond a reasonable drying period has been considered important to long life. We are unable to give a good explanation of why aging should improve retorts. During the past fall in an attempt to prove a theory that the combined water in the clay decreased during aging, about 50 samples of wet retort collars were placed in a dry room when it was closed. Loss of weight, combined water and free moisture were measured each week over the 100 day drying period. The affinity of the dried retort for moisture makes difficult the accurate measurement of free water; however, there was a clearly defined free moisture change from about 12.5 pct when freshly made to about 2 pet at the end of 18 days, then a gradual loss to about 0.25 pet at 70 days and an apparent equilibrium near this figure through the 100th day. There was no appreciable change in the combined water. The change in weight of the average sample was, however, rather unusual, there being consistently a loss of weight throughout the entire 100 days. From the 63rd day on this leveled off to about 0.03 pet per week. These were small samples and certainly should have dried

Jan 1, 1950

By Edna A. Dancy, Gerhard J. Derge

Jan 1, 1963

By Ivan B. Cutler, Milton E. Wadsworth

Coprecipitation of anionic and cationic polyelectrolytes has been applied to floccula-tion of several mineral systems. Results obtained in a study of the flocculation of kaolinite and hematite suspensions of polycationic and polyanionic electrolytes are presented. Greatly increased settling rates were observed following precipitation of positive and negative polyelectrolytes on the surface of finely divided minerals in aqueous suspension. The ratios of polycationic to polyanionic electrolytes required to produce maximum sedimentation have been shown to correspond closely with the equivalence points obtained by light scattering studies of systems containing the positive and negative polyelectrolytes by themselves. DISPERSION of colloids has been of interest to surface chemists for many years. Flocculation studies have been employed to evaluate the limits of colloid stability. In water suspensions, colloids become unstable and flocculate when the 5 potential, the potential of the kinetic solvation sheath, falls to a value at which van der Waals forces may act to draw the particles together. This may happen through an adjustment of the pH of the suspension or by addition of a salt. Optimum conditions for rapid formation of large flocs are found in the region of minimum < potential. In addition to these well known considerations, colloids in water suspensions are known to be flocculated by certain large organic macromolecules. These polyelectrolytes in some cases minimize the 5 potential and in other instances increase the attractive forces by supplying sites for hydrogen bonding.&apos; Use of organic polyelectrolytes in flocculation of mineral colloids has increased rapidly in recent years. These materials are now recognized as indispensable aids to fluid-solid separations in many extractive metallurgical operations. They have been found to increase both the settling rate in thickeners and the filtration rate of filters. In the elucidation of the characteristics of these organic reagents as floc-culants in this laboratory, it was discovered that interaction of certain polyelectrolytes had a surprisingly good effect on flocculation of mineral colloids. Experimental Procedure: The anionic flocculants used in this study were Lytron 886 and 887, produced by Monsanto Chemical Co. Both reagents are copolymers of vinyl acetate and maleic anhydride. Lime is added to Lytron 886, whereas Lytron 887 is all organic. These reagents possess two carboxyl groups for each acetate group and are therefore anionic in character. The cationic reagents used were the Peter Cooper Co. 1-X and 2-X glues. The structures are not definitely known, but the re- agents are amines and are cationic in character. The glues were prepared in the chrome-alum form. Both types of reagents may be classified as polyelectrolytes in that they are high molecular weight polymers and ionize in water. All data presented are in terms of sedimentation rate, measured in 250-ml glass-stoppered graduated cylinders. With suitable precautions such measurements are fairly reproducible and provide a sensitive means for selecting important parameters such as PH, reagent concentration, and floc size and density. Agitation was standardized with five complete end over end cycles of the suspensions in the stoppered graduated cylinders.

Jan 1, 1957

By M. H. Caron

D. C. Ralston—The fact that none of the organizations that have worked on these ammoniacal leaching processes have contributed discussion of Mr. Caron&apos;s papers today is a matter of some disappointment. Adaptations, modifications and improvements have been made by several organizations or individuals and to complete the record I want to document a number of them. The Nicaro Nickel Co. operated a Government owned plant at Nicaro, Oriente, Cuba, during the past war. A report on the project was written for Reconstruction Finance Corp. and is in the document file collected by the Office of Technical Services, Commerce Department and filed with the Library of Congress, where it can be consulted or bibliofilm or photostat copies obtained by asking for copies of PB 97271. Nicaro Nickel Co. has obtained a series of U.S. Patents, my list (probably incomplete) of which bears the following numbers: 2,400,098; 2,400,114; 2,400,115; 2,400,461; 2,400,612; 2,458,902 and 2,473,795. International Nickel Co. has also studied the problem and U.S. Patent 2,478,942 shows that some of the complications involving occasional low extractions have been unraveled. Forward, Samis, and Kudryk, of the University of British Columbia, have also published a study of copper-nickel sulphide concentrate.&apos; Others are interested and it is evident that this type of hydrometallurgy is likely to find a happy application in the not-too-far distant future. Caron&apos;s two papers practically amount to a summation of his life&apos;s work on the important subject of nickeliferous iron ores and the American Institute of Mining and Metallurgical Engineers is fortunate to have been chosen the medium of publication. M. Rey—The ammonia leaching process has been the subject of certain studies in New Caledonia, but at the present time, the conditions are not very favorable to its adoption. The process is a combination of a thermal and a hydrometallurgical process. As a result, the first costs and the operative costs are rather high. To counteract this unfavorable situation, it would be necessary to make a very high extraction. But unfortunately, the process is delicate and in spite of all of Caron&apos;s valuable work, there are still many unknown factors. It is difficult to obtain regularly a high extraction, exceeding 75 to 80 pct. It would be interesting to have the opinion of Inter-national Nickel on the process and to know what the Dutch are doing in Celebes. M. H. Caron (author&apos;s reply)—I am quite familiar with the different types of nickel and cobalt ores as found on New Caledonia, and it may be pointed out that the Celebes ores are quite the same in their properties. Their behavior when treating the ores by the ammonia-leaching process is also alike and in my opinion, there is no reason why a high extraction of the New Caledonian ores could not be obtained if the process is properly applied. For the sake of completeness I may add that one of my patents has not been discussed in the papers. This patent deals particularly with the treatment of serpentine nickel ores. (Netherlands No. 62693 40a, French No. 965.786 April 21. 1948. and Cuban No. 13.625 Jan. 1950.)

Jan 1, 1951

By R. F. Rolsten

VAN Arkel 1 prepared ductile tantalum by the thermal decompoiition of tantalum pentachloride on a resistively heated wire (2000° C) in an evacuated bulb maintained at 100°C. Burgers and Basart2&apos;3 observed the formation of a mixture of tantalum carbide and free metal when a carbon deposition base was used. campbel14 suggested the thermal decomposition of the iodide if a lower decomposition tem- perature was desired. To date, the details for preparing high-purity iodide tantalum have not been disclosed. EXPERIMENTAL The tantalum to be purified was symmetrically placed around the clear quartz deposition surface centrally located within a 64 mm cylindrical Vycor bulb. The deposition vessel usually employed5 was modified6 by incorporating a reentrant fin,mer in place of the customary metal filament in the head

Jan 1, 1960