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By S. W. Ross

A. A. CENTER*—This paper reminds me of the beginning of the work of the Electrolytic Zinc Company of Australasia. Early work for this Company, as some of you may know, was done at the Bully Hill Plant in California. After working there for some time Herbert Gepp, now Sir Herbert, came to the Anaconda Co. plants in Montana where I met him. He spent some months there, in fact almost a year if I remember rightly. That was when we were breaking in the new electrolytic zinc plant at Great Falls, Montana. From there he went back to Australia and started in on the plant there. I believe they had an original plant of about 20 tons per day of cathode zinc. We heard from them for some time, until they were able to go along on their own, and they have done very well ever since as you have heard. Their practice, (the flow sheet, the leaching practice), has been quite different as reported in the paper and apparently they are still following that practice to some extent. While the earliest work on selective flotation was done down in Australia, the big companies there, strange to say, were far behind the American companies in taking it up; and it is only in comparatively recent years they have gone to the alkaline circuit and dropped out thereby more of their iron and lead. The zinc plant now is up to a capacity of 100,000 tons a year and I believe that may be notched up to about 105,000. T. H. WELDON*—I would like to ask

Jan 1, 1950

By G. H. Anderson

Cadmium-coppeR precipitate, a byproduct of the purification stage of the zinc plant, is composed mainly of zinc, cadmium and copper in varying amounts depending on the efficiency of precipitation and the cadmium and copper contents of the impure solution treated. The composition range is approximately: cadmium 10-12 pct, copper 6-8 pct, and zinc 3&35 pct. As received at the cadmium plant, the precipitate is a dark gray to black press cake produced by filter pressing the flocculent cadmium-copper precipitate formed when impure zinc solution is agitated with zinc dust. The details of the cadmium-copper precipitation are to be found in the paper "Electrolytic Zinc at Risdon, Tasmania," by W. C. Snow. Transactions AIME, 121, 501. The treatment, in addition to recovering cadmium as metal, yields a copper product as a residue for realization and recovers in solution for return to the zinc plant, the excess of zinc dust used during purification. The recovery of cadmium metal is approximately 200 tons† per year, but, with the amount of copper residue recovered, varies with the zinc production tonnage, composition of the original concentrates, degree of roasting and other operational factors. The operations involved in the recovery of the foregoing materials are carried out in a building adjacent to the purification section of the zinc plant and can best be described by dividing the process into the following stages: 1. Oxidation and Grinding of the cadmium-copper precipitate. 2. Leaching and Filtering. 3. Precipitation of cadmium. 4. Oxidation and Grinding of the precipitated cadmium. 5. Leaching oxidized cadmium precipitate and Purification of leach solution. 6. Electrolysis. 7. Melting, Casting and Packing. A feature of the process is the use of two distinct solution circuits. Spent electrolyte from the zinc plant circuit containing about 10 pct sulphuric acid is used as the primary solvent of the cadmium and zinc present in the oxidized precipitate. When the cadmium is later separated by precipitation and filtering, the filtrate, originally zinc plant spent electrolyte and now fortified in zinc, is returned to the zinc plant. The precipitated cadmium, after oxidation, is redissolved in spent electrolyte from cadmium electrolysis which solution is in closed circuit within the cadmium plant except for discards to the zinc plant circuit as mentioned in Section 5 hereinafter. (Fig 1.) Details of the various operations follow. I. Oxidation and Grinding of the Cadminm-copper Precipitate The precipitate having had a preliminary drying by compressed air in Dehne filter presses in the zinc plant is trucked to a storage platform of sufiicient length to store 3 to 4 days' production separately. After 48 hr exposure to atmosphere, oxidation is sufficient to enable all cadmium and zinc but only a portion of the copper to be dissolved in dilute sulphuric acid. The oxidized precipitate is shovelled into trucks and, after weighing and sampling, is broken in size by means of

Jan 1, 1950

By G. H. Anderson

H. R. HANLEY*—I have been asked to discuss briefly the development of rotating cathodes for the electrolytic deposition of cadmium. The earliest recorded use of rotating cathodes was by Hoepfner at Frufurt, Germany about sixty years ago. He elec-trolized zinc chloride solution using diaphragms to separate electrodes. In the early experimental work of the Bully Hill Copper Mining and Smelting Co., Shasta County, Calif., rotating aluminum cathodes 4 ft in diam were used in the electrolysis of an acid zinc sulphate solution. Finished cathodes weighing up to 400 lb were produced. Because of mechanical difficulties, this type of cathode was abandoned for zinc, but was later used for cadmium because of the relative smoothness of deposit in comparison with stationary plates with comparable current densities. Cadmium sponge which forms on the cathode at moderate current densities (without special treatment) is entirely eliminated by a slow rotation. The rate of rotation of the cathode has an effect on the mechanical nature of the deposit. A high rate of rotation concentrates the adhering electrolyte on the shaft; a moderate rate appears to concentrate on the cathode a short distance out from the shaft tending to corrode the deposit in the form of a ring. At a very slow rotation (2 to 3 rpm) the adhering electrolyte gravitates nearly vertically, thus avoiding the cutting ring referred to above. The true explanation for the smoother deposits obtained on rotating cathodes may not be given definitely as the numerous factors involved are not thoroughly understood. Smooth deposits are obtained when the orderly growth of the metal crystals in the cathode lattice are disorganized. Thus the crystals form and grow for a very short interval when they are arrested and a new crystal forms. The continued growth of the original crystals provides large crystals and a rough deposit. Also if the acidity of the electrolyte is low, hydrogen gas bubbles adhere to the deposit. As the cathode is rotated the gas surface is brought into the atmosphere where they burst; thus the deposit is made on a surface relatively gas-free. An aluminum hub distance piece was riveted to each aluminum disk 4 ft in diam, slipped on a 4 1/2 in. steel shaft and pressed tight to prevent acid electrolyte seeping through to the shaft. The 9-cathode assembly was supported on insulated bearings. Electrical contact to the shaft was made through what was equivalent to a copper pulley. Sufficiently high conductivity brushes were placed on the face of the pulley to lead the current to the cathode bus bar. The assembly was driven by a link belt contacting a sprocket insulated from the shaft. The lead anodes were semicircular and supported on porcelain insulators placed on the bottom of the cell. Two anodes were provided for each cathode to permit an 8-in. space between them without increasing the ohmic resistance. This ample spacing permitted easy stripping of deposit with the assembly in place. Cathode cadmium was melted under 650 W cylinder oil. After casting, the primary slabs were remelted under molten caustic soda and cast into pencils 1 1/32 in. in diam. Rotating cathodes for deposition of cadmium are used at Risdon, Tasmania, and at Magdeburg, Germany. W. G. WOOLF*—This paper is very-interesting to me because in our work at the Electrolytic Zinc Plant of the Sullivan Mining Co. we had an exactly similar problem—that is, a method of producing cadmium from our purification residue, the recovery of the contained copper as a copper precipitate which could be sent to a copper smelter and the production of merchantable cadmium. It is interesting to me, not knowing of the work of the Risdon people, how closely we approximate them in their main metallurgy, diverging at several interesting steps which I would like to discuss for just a moment. For example, at Risdon they oxidize their purification residue. In our practice we take the current residue as it is produced in the purification department of the zinc plant and process it in the cadmium plant. The only oxidation that it obtains is the oxidation in the presses, the dumping of the presses and the collection and transportation of the residue to the cadmium plant. We find that the leaching of that residue does not necessarily require the oxidation step that the Risdon people evidently find necessary. The discussion of oxidation comes in again in the matter of the treatment of the precipitated cadmium sponge with zinc dust which again at Risdon is oxidized but which we do not attempt to oxidize except as it oxidizes itself in the storage. There is a partial oxidation which cannot be avoided, as Mr. David-sou pointed out, but we make no attempt to attain a complete oxidation and we dissolve the cadmium sponge in the sul-

Jan 1, 1950

Established as a Division November 17, 1948 T D Jones, Chairman O C Ralston, Past Chairman R R. McNaughton, Chairman-elect B. W Gonser '53, Vice-Chairman P. T. Stroup '54, Vice-Chairman Ernest Kirkendall, Secretary 29 West 39th Street, New York City

Jan 1, 1952

Established as a Division November 17, 1948 B W Gonser, Chairman J C Kinnear, Jr, Past Chairman H H Kellogg, Chairman-Elect R C Cole, Vice-Chairman, '58 A E Lee, Jr, Vice-Chairman, '59 W R Opie, Vice-Chairman, '60 R W Shearman, Secretary AIME, 29 West 39th Street New York 18, N Y

Jan 1, 1957

By W. R. Opie, O. W. Mole

By confining the electrolytic reduction of TiCl4 to the interior of a porous basket-cathode the electrolyte between the anode and the cathode can be kept free of reduced chlorides of titanium eliminating undesirable reactions of reduced titanium chlorides with brickwork or oxygen in the cell atmosphere. Semiwalls are not needed. Operating conditions are given and discussed. Titanium yields based on TiCl4 introduced are 93 pct. Cell voltage and amperage characteristics are adaptable to multiple cathode-series cell urmngements. EXPERIMENTS conducted in the National Lead Co., Titanium Division, Research Laboratories showed that titanium metal could be deposited in an electrolytic cell in which titanium tetrachloride was introduced through a tubular cathode.' The reduced chlorides as formed were confined to the vicinity of the cathode. They were reduced to form a cocoon-shaped growth of sponge titanium metal at the end of the cathodic tube below the surface of a sodium chloride-strontium chloride electrolyte. No diaphragm was needed.' Although high-purity, coarse crystalline metal was produced in the cocoon growth, the yields of metal removed from the cell, based on the amount of titanium introduced as TiCl4, were usually only 40 to 60 pct. The growth of the cocoon deposits was away from the cathodic tube for introduction of TiC14 toward the anode following the flu lines. The reduced chlorides of titanium diffused through the deposit and were reduced to titanium electrolytically within the deposit or on the outer surfaces. Deposition within the pores of the cocoon was detrimental to further deposition because it plugged the passages used by the reduced chlorides to diffuse away from the point of titanium introduction. As plugging due to deposition in the pores proceeded, the reduced chlorides in the center of the cocoon would force their way out by forming a channel and would escape from the vicinity of the cathode without being reduced to metal. Some work had been done previously1, 2 which indicated that if the end of the cathodic tube for introducing TiC14 was surrounded by a screen connected to the tube so that it too was cathodic, growth could be made to occur within the screen. Direction of growth was observed to be from the screen toward the point of TiCl4 introduction. Through a series of design modifications a cathode evolved which consisted of an introduction tube surrounded by a perforated metal basket as shown in Fig. 1. Modifications of this design are described in Refs. 3, 4, and 5. This basic cathode designa" was incorporated in electrolytic cells containing graphite anodes and constructed of high-silica brick refractory for containing the electrolyte (see Fig. 2). The inner crucible was surrounded by insulating brick and the entire unit contained in a welded steel shell. Suitable cooling chambers with inert gas atmospheres were devised so that the cathode could be removed from the cell without exposing the deposit to air while hot. Various alkali and alkaline earth chlorides were used as electrolytes.

Jan 1, 1961

By Robert D. Pehlke, Kazuhisa Okajima

ThE prediction of the activity coefficient for a solute in a multicomponent liquid metallic solution based on measurements in the related binary systems has been the object of study for some time, particularly in iron-base alloys.1,2 Very few measurements, however, have been made on nonferrous systems; furthermore, measurements made for the specific purpose of testing various prediction equations are almost nonexistent. The measurements here described of the activity coefficient of cadmium in liquid lead alloys were directed toward this goal. One of the most direct means by which the activity coefficient can be calculated has been based on the independent influences of each solute. When solute metals i,j, k, and so forth are present, the activity coefficient, y, of one of these (i for example) may be expressed as a product of factors which represent the effects of each of the other components:

Jan 1, 1964

By H. H. Kellogg

Available high-fernperatzwe equilibrium measurements and other thermochernical data have been critically reciewed for sulfation reactions of the type: Recommended values for log K and ?Fo as junctions of temperature are given for thirty-six suck reactions involving twenty-six different metals. Methods of calculation and choice between conflict-ing data are discussed. SELECTIVE sulfation now enjoys a considerable renaissance as a metallurgical unit process due, in large part, to the available of the fluid-bed re- actor—a device capable of the close control of temperature and gas composition necessary for a successful sulfation process. An additional need, however, if sulfation chemistry is to be approached in a scientific manner, is a reliable body of data on the equilibrium and kinetics of sulfation reactions in simple and complex systems. This paper offers a critical review and compilation of equilibrium data on pure sulfates (simple systems). The last such review, by K. K. Kelley in 1935:30 has proven itself most useful, but the existence of many new data makes necessary a re-evaluation of the available information. METHODS OF CALCULATION For a simple sulfation reaction such as MS04(c) = MO(c) + SO3 [1] there are two approaches one may take toward establishing the value of ?Fo or K as a function of

Jan 1, 1964

By Milton E. Wadsworth, W. Martin Fassell, William H. Dresher

A study of the rate of dissolution of molybdenite (MoS2) in alkaline solution was carried out under carefully controlled conditions. Effects of temperature, oxygen over-pressure, and KOH concentration were evaluated. Studies were made in the temperature range 100°C to 175°C and in the pressure range 0 to 700 psia of oxygen. Under these conditions molybdenite was found to leach according to a linear mechanism. Both oxygen over-pressure and KOH concentration were found to control the rate of leaching. The mechanism has been explained in terms of adsorption of oxygen at the molybdenite surface followed by configurational rearrangement of the adsorbed molecules. The hydroxyl ion dependency is believed to be diffusion-controlled. Laboratory batch studies have shown that molybdenite can be readily dissolved in alkaline solutions under moderate conditions of temperature and pressure. Commercial application of this process to the production of ferro-molybdenum and molybdenum chemicals is promising in view of the ease of dissolution of molybdenite and the relatively noncorrosive conditions involved in the process. HIGH temperature-high pressure techniques have long been used to great advantage in the organic chemical industry, the petroleum industry, and the paper industry. Only recently, however, have these methods been used to extract metals from their ores on a commercial scale. The Chemical Construction Corp., together with interested producers of nickel and cobalt, has done much to develop methods of producing nickel and cobalt powders by the use of high temperature-high pressure techniques.' The Howe Sound Co. is currently operating a plant near Salt Lake City in which the new Chemical Construction Corp. technique is applied to cobalt-arsenic sulfide concentrates. In this method the concentrates are leached in a dilute solution of sulfuric acid under air pressure at a high temperature. Near Edmonton, in Alberta, Canada, the Sherritt Gordon Co. is also using the new method in conjunction with its ammoniacal leach process.' National Lead Co., at Fredricktown, Mo., is operating a process similar to that of Howe Sound. In spite of the rapid growth of high temperature-high pressure hydrometallurgical processes, there has been very little work done on fundamental principles involved in this type of process. Conse- quently the literature lacks any significant data in this field. Oxidation of pyrite,3 phalerite,4 galena,5 and recently molybdenite" in alkaline solutions has been reported to some extent. This investigation was initiated, therefore, to broaden the application of high temperature-high pressure hydrometallurgy and also to contribute some insight into the basic mechanisms of oxidation in an aqueous medium under these conditions. In addition to the immediate consequences of these processes much valuable information remains to be learned about fundamental reactions when they are carried out at temperatures and pressures slightly higher than those normally encountered. Leaching of molybdenite by alkaline solutions under oxygen pressure has recently been reported in the Russian literature by E. S. Usataya,8 who found that leaching rate increased with an increase in the pH of the solution—the strongest effects being observed at a pH of 10—and also with an increase in temperature. Usataya also has reported the formation of a protective film on the surface of the molybdenite when the leaching solution was weakly alkaline, neutral, or acidic. He believes that this oxidation process explains the migration of Mo+ + in the oxidation zones of molybdenite ore deposits and the subsequent impoverishment of the ores. Equipment: The high temperature-high pressure reaction unit used in this work was especially designed for kinetic studies of this nature, emphasis being placed on accuracy of measurement and precision of control. As details of the unit's construction and operation have been discussed elsewhere,7 Only basic features and principles of operation will be discussed at this point.

Jan 1, 1957

By A. J. Boyer, T. R. Meadowcroft

An experimental determination has been made of the efject of surface diffusion on the vapor pressure obtained in a Kwudsen cell. The results show that a knife-edge orifice in a molybdenum lid may give erroneous results for the vapor pressure of silver. when a correction suggested by winterbottom and Hirth is made, the results become meaningful. Experimental work with cylindrical orifices in molyhdenz~m lids and knife-edge orifices in ceramic lids shows that good vapor-pressure results may he obtained when the orifice is designed so that ejfusion is the only significant source of weight loss. THE Knudsen effusion cell technique has been used by many investigators to study relative and absolute vapor pressures over metals and alloys at elevated temperatures. As a result of this interest, some theoretical studies have been made of geometric factors affecting effusion from the orifice of a Knudsen cell. A review of these studies has been made by Margrave.1 Recently, Winterbottom and Hirth2 have suggested that a physical effect, that of condensation, surface diffusion, and re-evaporation, may seriously affect Knudsen cell measurements, especially when a knife-edge orifice is used. Inasmuch as many of the present vapor-pressure data have been obtained using knife-edge orifices. it was felt that an experimental evaluation of the effect of surface diffusion on Knudsen cell measurements would be of value both in the design of the cell orifice and in establishing realistic estimates of the limits of error for existing data. The experimental program was in two sections. Firstly, measurements of the vapor pressure of silver were made in cells with cylindrical and conical orifices in which the surface-diffusion contribution to the weight loss should be negligible. These results were correlated with existing data to give a reliable measure of the real vapor pressure of silver. Secondly, measurements made using knife-edge orifices were corrected using the surface-diffusion theory of Winterbottom and Hirth2 and were then compared to the real values. The experimental arrangement is shown in Fig. 1. The balance was an Ainsworth RV-AU-1 automatic recording microbalance capable of measuring weight changes of ±0.02 mg on a total load of up to 100 g. The temperature was measured by an optical pyrometer signted on the bottom of the suspended crucible. The temperature readings were calibrated in a separate experiment in which pyrometer measurements were made while at the same time thermocouple readings were taken of the temperature at the inside top and bottom of the crucible. The range of reproducibility with this technique was ±4°C. A constant voltage source maintained a temperature of ±2°C at 1200°C over 10 hr. Silver of 99.99 pct purity was used in nonporous alumina crucibles. Lids of molybdenum, tantalum, or magnesia were cemented to the crucible using a sodium silicate cement. Table I shows the characteristics of the lids used. The vapor pressure is obtained using the basic Knudsen formula:

Jan 1, 1965

By R. G. Ward, A. E. El-Mehairy

A technique was developed to calculate the density of liquid metals from the profile of a weighed levitated drop obtained by emitted light photography and calibration. The density of liquid copper was determined over the temperature range 1370" to 2100°K and expressed with the equation: The molar volumes and thermal coefficients of expansion are calculated. Copper, upon melting, was found to expand 3.34 pct. A knowledge of the density of liquid metals is essential as it is a parameter appearing in most theories of the liquid state. The liquid density is also necessary for the calculation of the commercially important contraction occurring on solidification. Unfortunately good density values are rarely available and the present work was undertaken to establish a new method of density determination free from the limitations of the existing methods. For liquid metals with high melting points, the variation and the limited number of density values reported in the literature have been caused primarily by difficulties in technique, a matter which has been adequately discussed by Urbain.' In all the methods used for measuring the density of the high-melting liquid metals there are sources of error, and corrections have to be made. For example, in the method of balanced columns, the formation of gas bubbles in one or both arms was a great source of error. In the pyknometer3 and immersed-sinker methods, corrections of certain volumes for thermal expansion to the operating temperature exceed 4 pct. In the maximum-bubble method5 a correction for the thermal expansion of the bubble-tube is used (about 1.6 pct); a second correction takes into account the change in the volume of liquid displaced by the bubble-tube (about 3.5 pct). Another method used is based on the shape of a drop of liquid, at rest on an inert refractory plaque, obtained photographically or radiographi-cally. The volume is calculated from the shape, which is sensitive to the surface tension and supporting material, by some intricate empirical relation. The drop contour changes with 90 deg rotation of the support.7 The application of all methods is limited by reaction between liquid metal and the apparatus. The successful stabilization of a mass of liquid metal during levitation melting8 afforded a technique for measuring the density of liquid metals out of contact with solid materials. The technique was developed by using copper as the density of liquid copper has already been studied extensively, thus allowing comparison with the new method. It consists of calculating the volume from the dimensioned profile of the levitated drop obtained by emitted-light photography and careful calibration. The molten egg-shaped drop spins around the magnetic axis of the coil (lined vertically) at a speed which seems to increase with increasing temperature. The method does not involve corrections and the experimental setup allows the use of vacuum or special atmospheres. The maximum operating temperature is limited by excessive evaporation of metal from the droplet. EXPERIMENTAL Levitating System. A stabilizing ring was used in conjunction with a 3-turn levitating coil energized by a 450-kc per sec, 10-kw Toccotron with a high current-low voltage output stabilized to f 1 pet.' APyrex glass tube with a Pyrex Optical flat window inserted through the coil and ring allowed the use of special atmospheres during levitation melting, as shown in Fig. 1. Optical System. Fig. 1 shows the optical arrangement used to photograph the profiles of the drop, parallel and perpendicular to the vertical spinning axis, simultaneously on the same frame of a high-speed 35-mm film. The optical system was adjusted to avoid any distortion of the profiles and to ensure that both profiles are in focus. The length of path from object to objective was about 30 cm, and a magnification factor close to 1 was obtained with the single-lens reflex-camera system. Temperature Measurement and Control. In operation, the specimen temperature was constant to 5 C for a fixed power input to the levitating coil and fixed rate of gas flow through the tube. The temperature was measured by an automatic two-color pyrometer (Coloratio) checked at the melting

Jan 1, 1963

By P. W. Bennett, E. M. Elkin

A new technique for the recovery of palladium and platinum and sludge from go12 electrolyte eliminates many of the drawbacks of the zinc-dust cementation process. In the electrolytic refining of gold by the Wohwill method, the electrolyte contains about 100 gpl gold as AuCl, and 100 gpl free hydrochloric acid. Any palladium and platinum contained in impure gold dissolve and accumulate in the electrolyte. When their concentration approaches that of gold, it becomes desirable to discard the electrolyte and to recover the gold and the palladium and platinum. As practiced at Canadian Copper Refiners until recently, discarded electrolyte was treated with sulfur dioxide to precipitate the gold. 2AuCl3 + 6H2O + 3SO2 —2Au + 6HC1 + 3H2SO4 The gold precipitate was filtered off, washed with water, and melted with scrap anodes. The wash water was added to the gold-free filtrate and treated with zinc dust to cement the palladium and platinum. PdCl2 + Zn — Pd + ZnCl2 PtCl4 + 2Zn—Pt + 2ZnCl2 The precipitate was separated and sold as Pd-Pt concentrate, and the barren filtrate and wash water were discarded. The receipts of palladium and platinum at this company were too small to warrant their refining. The process suffered from a number of disadvantages. Handling the gold-free solution was disagreeable, as it was laden with sulfur dioxide even after blowing with air overnight. Subsequent treatment with zinc dust was also troublesome. A large excess of zinc dust was required. Spray, acid mist, and vigorous evolution of hydrogen accompanied the reaction. The Pd-Pt precipitate was grossly contaminated with occluded unreacted zinc and with yellow granules of sulfur, thought to be formed by the reduction of sulfur dioxide and its derivatives. Repeated boiling of the precipitate with hydrochloric acid and washing with water decreased somewhat, but did not eliminate, the zinc. After drying overnight, the precipitate hardened to a clinkerlike mass that was difficult to grind. Grinding, handling, and packing created heavy dusting. Even weighing in an analytical balance generated dust, indicating that the powder was easily charged with static electricity. The total palladium and platinum content of the powder varied from 56 to 72 pct, with an average of 65 pct. The powder dissolved in aqua regia with difficulty, leaving a small amount of insoluble residue. However, the main drawback of the process was the strongly hygroscopic Pd-Pt product. The powder showed a marked though erratic increase in weight and a corresponding decrease in assay on exposure to air, even during the course of laboratory analysis. Such changes were the source of disagreement between Canadian Copper Refiners and the buyers of concentrate. The cause of the change in weight is believed to be hydrogen, liberated by the action of zinc dust on hydrochloric acid. Some of this hydrogen was adsorbed so strongly by the precipitate that it could not be driven off during the drying. On exposure to air it oxidized catalytically, forming water. Experiments soon showed that the entire process of treatment of discarded electrolyte then in use would have to be changed. A small quantity of apparatus was moved into the gold room, and experimentation was held to a minimum compatible with the development of a new process. PALLADIUM AND PLATINUM PRECIPITATION It was thought that zinc could be replaced with a reagent other than a hydrogen-evolving metal. Semiquantitative experiments showed that platinum and palladium could be precipitated completely from the gold-free solutions with such well-known reagents as sodium formate, formic acid, formaldehyde, and paraformaldehyde. Sodium formate was the preferred reagent, since the other three are volatile and somewhat toxic. The gold-free solutions used were those saturated with sulfur dioxide, partly freed of sulfur dioxide by aerating overnight, completely freed by boiling, as well as strongly and weakly acidic, neutral, and alkaline solutions. The reactions are complex and can be represented by these over-all equations: PdC1, + HCOONa — Pd + NaCl + HC1 + Co2 PtCl4 + 2HCOONa — Pt + 2NaCl + 2HC1 + 2CCX, Strongly acidic solutions required as much as five times the theoretical amount of sodium formate as did solutions neutralized to pH 4.5. Increasing the pH decreased the consumption of

Jan 1, 1965

By G. R. Rigby

Used basic brick taken from copper anode furnaces, reverberatory furnace roofs, md copper converters have been examined by chemical analysis, petro-graphical examination, and X-ray diffraction tech-niques. In the anode furnace roof, brick absorb copper and copper oxides, whilst in the reverberatory roof, iron oxide with smaller amounts of alumina and silica are absorbed. Chemical reactions in the converter are most complex since the brick absorb both a copper-copper oxide slag and a fayalite-magnetite slag. CHANGES taking place in basic brick after use in copper smelting furnaces have received attention from C. Burton Clark and J. Spotts Mc Dowell,' who examined brick taken from copper converters, and by G. R. Rigby and B. G. Hamilton' who examined brick from copper anode furnaces. This paper describes the changes taking place in basic brick from all types of copper smelting furnaces and discusses the major reasons for lining wear. A) METHODS OF EXAMINING SAMPLES A variety of used brick have been taken from the following positions in various furnaces, a) roofs, reverberatory arches, and bottoms of copper anode furnaces, b) roofs of copper reverberatory furnaces and c) the tuyere line and end walls of copper converters. The method of examination was standard in that brick were first measured and the extent of wear-back computed. They were next sawn in two by a central cut vei-tical to the working face to enable the freshly cut surfaces to reveal any zonal changes and cracks, both of which usually ran parallel to the working face. Each well-defined zone was then sampled for a) chemical analyses, b) petrographical studies using transmitted and reflected light techniques under the microscope, and c) X-ray diffraction studies on powdered samples. From the data accumulated, changes taking place in a variety of

Jan 1, 1962

By Attilio Ligasacchi

The production of magnetite in copper smelting has been the subject of wide investigation in the past.1-9 It is accepted that most of the magnetite is produced in the converters and returned to the re-verberatory furnace with the converter slag. Part of this magnetite is circulated back to the converters in the matte. As a result, if proper fluxing practice is not followed, and temperatures in the converting process are not strictly controlled, an undesirable circulating load of magnetite may occur. This might ultimately lead to saturation in the reverberatory bath with consequent settling of the magnetite at the furnace bottom. This paper contains the cumulative results of a series of investigations on the formation and control of magnetite presently being conducted by the metallurgical department of the Chino Mines Division of Kennecott Copper Corp. THE METHOD USED FOR THE DETERMINATION OF MAGNETITE AND OTHER MAGNETIC MATERIALS Conventional wet chemical magnetite analyses are neither reproducible nor theoretically sound. These analyses are sensitive to time, temperature, and heating rate-all factors which tend to decrease their accuracy and reproducibility. To overcome the unreliability of the above method, H. W. Mossman1 developed an inductance bridge magnetite "meter" which completely replaces the wet chemical analysis with the advantage of quicker and reproducible results. Fig. 1 is a schematic diagram of the meter in use in the Chino metallurgical laboratory. When a pulverized specimen with magnetic properties is placed in coil A of the inductance bridge circuit, the permeability of the core is increased, causing a voltage drop across the coil. The unbalance is amplified to operate the balancing motor which adjusts the movable iron core in coils C and D to bring the bridge back to balance. An indicator arm connected to the movable iron core indicates directly the percent, or weight, of the magnetic material in the specimen. Coil B acts as a reference coil and a zero adjustment. The meter is standardized with a set of various amounts of -100 mesh pure magnetite crystals (99.8 pet Fe3O4) thoroughly mixed in a test tube with -100 mesh quartz sand. These mixtures are combined with inert transoptic powder and heated to 140° to 160°C (melting point of the powder) and subsequently cooled to room temperature, solidifying the whole mixture, in this manner eliminating any possible preferential orientation of the magnetite grains. In order to determine the effect of the particle size" of the magnetite and other magnetic materials on the meter determinations, five size fractions were obtained by wet screening of the 99.8 pet pure magnetite and several reverberatory matte and slag samples. The following size fractions were investigated: -80 + 100; -100 + 150; -150 + 180; -200 + 250; -250 + 300 and -100 mesh. The deviation of the meter readings for the various sizes of each sample never exceeded ±0.3 pet. This deviation is considered well within the accuracy needed for common smelter operation procedures. Further investigation proved also that grinding slag and matte samples in our laboratory disc grinder normally produced a powder constituted of 95 pet -200 + 300 mesh. For practical purposes a - 100 mesh size was used as standard for our studies. It was also found that for rapid determinations tamping of the sample in the test tube was sufficient and gave comparable results with those samples solidified with transoptic powder. For the purpose of this study we define the magnetite content in copper smelter slags and matte as the equilibrium amount of crystalline Fe3O4 existent in the specimen at a given reference temperature, excluding any secondary magnetite precipitated in the specimen during nonequilibrium conditions. MAGNETITE DETERMINATION IN SLAGS Since it is impossible to ascertain the status of equilibrium of a conventional slag sample, due to

Jan 1, 1965

By M. Shelef, A. W. Fletcher

The effect of alkali sulfates in promoting the sul-fation of nickel and copper in a bulk sulfide flota -tion concentrate by fluidized bed roasting has been studied in the laboratory, and it was shown that the various alkali sulfates promote sulfation to approximately the same extent. The sulfation of a mixture of synthetically prepared iron and nickel oxide and of nickel ferrite has also been studied. Nickel sulfation was promoted by high ratios of Fe:Ni and by the presence of sodium sulfate. THE work described in this paper was a continuation of earlier studies into the role of alkali sulfates in promoting the sulfation roasting of nickel sulfides1,2 in an endeavor to determine how the system was affected by the presence of compounds of iron and copper. The earlier work1 showed that, in the sulfation of NiO at 680°C, the reaction was limited by the formation of an impermeable film of nickel sulfate on the oxide surface. The relative effect of the various alkali sulfates in promoting nickel sulfation varied in the order: Li > Na >Cs > Rb > K A study of alkali sulfate/ nickel sulfate interactions at high temperatures showed that the promoting action was due to the fact that the nickel sulfate product layer sintered and agglomerated only when the more active additives were present. This resulted in the formation of discontinuities in the nickel sulfate layer so that diffusion of the sulfating gases to the NiO surface was no longer impeded and the reaction could proceed to completion. A similar explanation was used for the observation that sodium and lithium sulfates promote the oxidation of NiS to NiO at temperatures below 750°C since small amounts of nickel sulfate were formed during oxidation.2 It was of interest to study the effect of alkali sulfates on the sulfate roasting of a sulfide flotation concentrate which is typical of material treated commercially. In order to control temperature it is essential to roast sulfides in a fluidized bed and this technique was therefore used, although the batchwise operation of a small-scale laboratory reactor does not reproduce all conditions which prevail in full-scale continuous plant. The results obtained are therefore only comparative, and cannot be used for predicting the optimum conditions for metal extraction. The sulfation of synthetically prepared mixed oxides of nickel + copper and nickel + iron and of nickel ferrite was also studied to evaluate the relative effects of alkali sulfates with more complex systems. SULFATION ROASTING OF A SULFIDE FLOTATION CONCENTRATE The bulk sulfide flotation concentrate used in this work contained 7.92 pct Ni, 1.74 pct Cu, 35.66 pct Fe, and 31.28 pct S. The sulfide minerals present in order of abundance were pyrrhotite FeS, pyrite FeS2, pentlandite (FeNi)S, and chalcopyrite CuFeS2. Two samples described as coarse and fine were used. The coarse sample, which was a flotation concentrate (58 pct plus 300 mesh), was ground to 100 pct minus 350 mesh to produce the fine sample. Before roasting, the sample of sulfide concentrate was agglomerated by wetting witli a solution of the alkali sulfate (or water), thoroughly mixing, and drying at 110°C. This gave a cake which was gently crushed and screened, the -18 +100 mesh fraction being used for fluidized bed roasting. A similar-size fraction had been used by the authors in pilot plant work with a 4-in.-diam fluidized bed reactor.' In this work it was found that the molar ratio of additive to the total iron + nickel + copper content of the sulfide sample should be adjusted to a value of approximately 0.06, as this was the optimum amount necessary for nickel sulfation. Experimental. The fluidized bed reactor consisted of a quartz tube approximately 60 cm long and 30 mm in diameter resting in a vertical tube furnace. The sulfide bed (30 g) was supported on a bed of -4 +12 mesh quartz particles 3 cm high, which rested on a sintered quartz disc welded to the tube. The temperature of the furnace was controlled with a variable transformer to give a final bed temperature of 680°C. The bed was fluidized with air or mixtures of air + 10 pct v/v SO2, at a total apparent gas velocity of 60 to 65 cm per sec at 680°C. The SO2 was introduced into the fluidizing air stream only when the oxidation of the sulfides was completed. At the end of the roasting period the calcine was leached with boiling water and the

Jan 1, 1964

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 T. P. Papazoglou, N. A. D. Parlee, W. C. Phelps

HE high vapor pressures of metals such as lead, calcium, lithium, bismuth, and magnesium at steel-making temperatures present experimental problems which have thus far rendered it almost impossible to obtain really reliable data for their solubilities in liquid iron. Many investigators have attempted to equilibrate liquid iron in the same crucible with the undissolved liquid phase of such volatile metals. The published results,l-6 although useful, are often questionable on the grounds of not being obtained under true equilibrium conditions. Such information is not always of practical value in any case since the equilibrium vapor pressure is often several atmospheres and steel is not currently made at these pressures. Solubility measurements at lower partial pressures (< 1 atm) appeared to be more practical and easier to make. These data could be used to calculate solubilities at other pressures where Henry&apos;s Law is obeyed, and to estimate solubilities under other conditions of temperature and pressure with useful accuracy. This note outlines the development of a general technique for the measurement of solubilities in systems of this kind. The method involves equilibration of the high-boiling metal at known temperatures with the vapors of a low-boiling metal at known partial pressures. Analyses of quenched samples provide vapor pressure-solubility data and make calculations of alloy thermodynamics possible. The principle involved is similar to the closed-system diffusion-transport method which is difficult to apply with most high-melting liquid metals like iron, due to closure problems. A two-zone horizontal (molybdenum wire element) tube furnace, Fig. 1, provides two constant-temperature zones in a reaction tuve (3/4 in. ID by 1 in. OD by 36 in.). A high-boiling metal sample, 9, introduced in the higher-temperature zone, is maintained at temperature T1, and a volatile metal sample, 7, in the lower-temperature zone, is held at T2. A stream of argon with 10 pct hydrogen is passed via flowmeters, 1, and mixing-drying towers, 2 and 3. through the reaction tube at a favorable flowrate and acts as a carrier gas which entrains vapor from the volatile liquid metal at a specific constant partial pressure. The gas mixture passes through a system of constrictions, 6 and 8, to the higher-temperature zone, establishes intimate contact with the high-boiling liquid sample, 9, and exits through a small condenser tube. The entire system is at (nearly) atmospheric pressure and the volume of carrier gas passed is measured with a wet gas meter. When the high-boiling sample attains equilibrium with the vapor phase, both samples are removed and quenched. Analysis of the high-boiling sample for solute content reveals its solubility for the low-boiling metal at a vapor pressure calculated from carrier-gas volume and either a) weight loss of volatile metal or 6) weight gain of high-boiling sample plus condenser (which collects the undissolved vapors). To test the capabilities of the method, liquid silver at 1403&apos;C was equilibrated with lead vapor at various partial pressures. The relatively simple Pb-Ag system was particularly suitable for uncovering experimental problems and providing the first step toward investigations with more volatile metals and more difficult combinations.

Jan 1, 1965

By Wayne L. Worrell

Thermodynamic data for the stable carbides and oxides of chromium, molybdenum, and tungsten have been critically eualuuted and are used to determine the stable condensed phases at 1 atm total pressure in each metal-carbon-oxygen system. Pourbaix-Ellingham diagrams have been constructed and are used to estimate the minimum temperatures neces.see)*y to oblain these metals by corbolhermic redlrction under reduced pressures. USING recently reported data for the carbides and oxides of chromium, molybdenum, and tungsten, the stable phases in the Group VI metal-carbon-oxygen systems at specific temperatures and pressures can be determined. The thermodynamically stable regions in each metal-carbon-oxygen system can be completely pictured using a Pourbaix-Ellingham diagram in which the two coordinates are oxygen potential (-RT In Po2) and temperature. Worrell and chipman&apos; have described and constructed such diagrams for the metal-carbon-oxygen systems of vanadium, columbium, and tantalum. Downing2 has used similar diagrams to provide a description of the reactions in submerged arc furnaces. One of the examples which he cited was the Cr-C-O system; however, an improved diagram is presented in this paper using recently obtained data for the carbides of chromium. Although some of the oxides and carbides of the Group VI metals exist over slight ranges of composition,3 all solid phases are represented as stoi-chiometric compounds in the subsequent diagrams. Any error introduced into the calculated equilibria by neglecting the compositional variation of these phases is less than the uncertainty in the thermodynamic data. In a Pourbaix-Ellingham diagram for a three-component system, the phase rule requires that at equilibrium four condensed phases specify a point, three condensed phases determine a line, and two condensed phases define an area. To obtain a well-defined area at high oxygen potentials and teniperatures, the initial carbon to oxygen mole ratio is specified to be more than sufficient to reduce the oxide but not enough to convert all the metal to carbide. Thus, the lowest regions in the subse- quent Pourbaix-Ellingham diagrams represent areas in which the metal and its carbide are the stable condensed phases. Cr-C-O SYSTEM A) Thermodynamic Data. To simplify the subsequent calculations, all thermodynamic data are presented in the form of linear Gibbs-energy-of-formation* equations which were derived for the temperature range 505° to 2000°K and are reasonably accurate over the entire temperature ranges of the diagrams. The thermodynamic data for the compounds occurring in the Cr-C-O system are summarized in Table I. Several volatile chromium oxides have been reported,4 but their influence is negligible in the Cr-C-O system. Thermodynamic data for CO and CO2 were taken from the compilation of Coughlin.5 The only solid chromium oxide sufficiently stable to exist below 1800°K in the Cr-C-O system is Cr2Os.11 The Gibbs-energy-of-formation equation for Cr2O3 was obtained by revising the data tabulated in Coughlin,5 using the heat of formation of ah&apos; and the high-temperature thermal data for chromium.7 In 1953 Richardson8 re-evaluated the original work of Kelley and coworkers,9 who measured the thermodynamic properties of the chromium carbides. However, the high-temperature equilibrium results of Boericke9 yield values for ?G°f of Cr2O3 which are 2 kcal more positive than the accepted data. His results for the chromium carbides are based on the same questionable equilibria. Gleiser10 recently determined the Gibbs energy of formation of Cr3C2 and the equation in Table I was calculated from her data and the thermal data of Kelley and coworkers.9 Thermodynamic data for the Cr7C3 and Cr23C6 carbides were obtained by combining the Cr3C2 data

Jan 1, 1965

By John Chipman, Wayne L. Worrell

Using recent thermodynamic data for the carbides and oxides of tantalum, columbium, and vanadium, the stable solid phases aboue 1300°K and at 1 atm CO(g) pressure in each M-C-O system have been determined. Pourbaix-Ellingham diagrams have been constructed and are used to estimate the minimum temperatures necessary to obtain these metals by carbothermic reduction under reduced pressures. TANTALUM, columbium, and vanadium react with carbon to form the M,C and MC carbide phases. The M,C carbides have narrow homogeneity ranges, but the MC phases possess considerable ranges of homogeneity.&apos; The free energies of formation of the MC carbides were recently measured, and those of MC and M2Ccarbides have been estimated., (The notations MC and MC denote, respectively, the carbide compositions at the carbon-rich and at the metal-rich boundaries of the homogeneous MC phase.) In this paper, these carbide data are combined with thermodynamic data for the oxides to predict the stable phases in the Ta-C-O, Cb-C-O, and V-C-O systems at specified temperatures and oxygen pressures. M-C-O PHASE DIAGRAMS The solid phases in equilibrium above 1300°K for each M-C-0 system are shown in Figs. 1, 2, and 3. Above 1300°K, the vapor would consist entirely of carbon monoxide; and these phase diagrams were obtained by specifying the vapor pressure of CO(g) to be 1 atm. The temperature at which any two solid phases become unstable is specified along the two-phase equilibrium lines in each diagram. To simplify the diagrams, all solid phases in the Ta-C-O and Cb-C-O systems are represented as stoichiometric compounds. However, the homogeneity ranges of VO and VC are indicated, because they form a complete series of solid solutions.3 It also should be noted that the homogeneity range of

Jan 1, 1964

By T. Rosenqvist

From the chemical, metallurgical, and mineralogical points of view, the importance of thermodynamic data for metal-sulphides and sulphur dissolved in molten metal has long been realized. Such data will give a basis for thermo-dynamic calculations of processes such as roasting and the distribution of sulphur between molten metal and slags. Furthermore, it will help us understand the chemical forces which tie sulphur to the metal and to the melt. Thermodynamic studies of solid sulphides and the equilibria between solid sulphides and metals have been carried out in a rather large number of cases by several investigators. An excellent review of this work is given by Kelley.1 On the other hand, very little is known about the thermodynamic properties of molten sulphides and of sulphur dissolved in molten metal. The only system which previously has been investigated in this respect is the iron/sulphur system. The purpose of the present investigation was primarily to obtain further data for metalisulphur melts and the system silver/sulphur was chosen, not because this system was expected to be of special interest in itself, but because the silver/sulphur system is a typical example of a metal/sulphur system. Because of the low melting temperatures, the work could be carried out without appreciable furnace and refractory difficulties. The sulphur in this system possesses a rather high escaping tendency compared to most other metal/sulphur systems. By the method employed (reaction with hydrogen to form hydrogen sulphide) this gives ratios Ph2s/Ph2 up to 0.30, which may be determined accurately. &apos;rhermodynamic investigations of the heterogeneous reaction Ag2S +H2 —+ 2Ag + H2S have previously been carried out by Pelabon,2 Keyes and Felsing,3 Jellineb and Zakowski,4 Wat-anabe,6 and Britzke and Kapustinsky.6 Their results are in rather poor agreement. Experimental Part PRINCIPLE In the present invehgation the escaping tendency of sulphur from the condensed phase was determined by the reaction: S + H2 = H2S The ratio Ph2s/Ph2 is at constant temperature, directly proportional to the escaping tendency of sulphur and to its chemical activity in the condensed phase. (This ratio will subsequently be denoted by H2S/H2.) From this ratio and its variation with temperature and with change in the com- position of the condensed phase, thermodyrlamic quantities such as free energy and heat of reaction can be derived. The apparatus used is shown in Fig 1. The silver/sulphur alloy was placed in a crucible inside a refractory tube and mounted inside the furnace to the left. The tube was a part of a closed, gas-tight system in which the gas mixture could circulate until equilibrium was established. Circulation was maintained by the glass propeller to the right, driven by an external magnet, and promoted by the chimney action of the furnace. The gas passed up through the furnace, down into the crucible where it came in contact with the alloy, passed rapidly up through a narrow tube and out of the hot zone. After equilibrium was established between the gas and the condensed phase, the composition of the gas mixture was determined from its density. In its circulation, the gas mixture passed through a chamber (lower right on Fig 1) containing a buoyancy-balance, where the density of the gas was determined by magnetic balancing. From the measured density,

Jan 1, 1950