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By Luiz C. Correa da Silva, Robert F. Mehl

A. D. Le Claire and R. S. Barnes (Atomic Energy Research Establishment, Harwell, Didcot, Berks., England)-—This much awaited paper admirably confirms that the Kirkendall effect is a true diffusion phenomenon and likely to be found in all diffusing systems. The most likely explanation put forward so far to account for the shift of interface is that the two components of the system, diffusing in opposite directions, do so with unequal rates so that there is a net loss of material from that half of the diffusion zone rich in the more rapidly diffusing component. This loss is accompanied, and at least partly compensated for, by a closing up or decrease in volume of this latter half of the diffusion zone and an expansion of the other half so as to tend to keep constant the overall volume per atom. This is manifest as a movement of the original interface. If this is a true picture of what is taking place then there are at least two important matters requiring further study."1—The quantitative relating of the observed shift with the independently measured separate diffusion coefficients of the two diffusing species."2—A detailed study of the mechanism and characteristics of the closing up process."We would like first to make a few remarks pertinent to these topics. In connection with the first of these, it is a pity that more and detailed measurements were not made on the Au-Ag system, the only one for which the separate diffusion coefficients of the two species have been measured.'" One of us (A. D. Le C.) has calculated from Johnson's results the shift to have been expected in samples D.2 and D.3. Darken4 showed that the velocity v with which the markers move is related to the separate diffusion coefficients Dau, Dag, by the eauation:"dN"v = (Dag — Dau "dx"dN"where .is the concentration gradient at the mark-ax"ers. Under certain conditions,4, 20 it can also be shown that:"Xm"v = "2t"where x, is the total marker shift after time, t:"dN". . xm = 2t (Dag — Dau) [9]"dX"Johnson measured Dag and DAu in a chemically homogeneous alloy and so his reported figures represent the true or ideal diffusion coefficients"; the same quantities in eq 9 represent the actual diffusion coefficients as would be measured in the presence of a chemical concentration gradient, so we need to multiply Johnson's reported values of Dag and Dau by the"1 before insert-dlogN "ing them in eq 9. 7 is the activity coefficient of Ag at the concentration prevailing near the markers which we shall assume to be - 50 pct. The values of the thermodynamic factor for N = 0.5 and the temperatures at which D.2 and D.3 were annealed were obtained from thermodynamic data quoted by Darken.' aN/ax was computed from the coefficient of inter-diffusion of Au and Ag given by Johnson on the assumption that in the samples D.2 and D.3, the concentration distance curve could be represented by the standard equation:""No is the original difference in concentration between the two halves of the couple, and + is the error or probability function."aN/ax at x = 0 is then given by:"adN/dx "2/nDt"and we assume that this represents the concentration gradient at the markers, although it is probable that the figure is slightly high."When all the appropriate quantities are inserted in eq 9, we find that the expected shifts are: for D.2, 0.0134 cm and for D.3, 0.0119 cm."The agreement between these results and those

Jan 1, 1952

By K. J. Brondyke, P. D. Hess

Lack of correlation between densities of aluminum alloy samples solidified under reduced pressure (vacuum gas test) and hydrogen content of the metal is explained on the basis of inclusions serving as nuclei for hydrogen bubble formation. As a result of these findings, the vacuum gas test is suggested as a practical tool to indicate metal cleanliness and the potential effects of hydrogen. Visual obseroation of the sample during solidification under vacuum provides more useful information than that obtained either from density meas -urements or sectioning of the solidified sample. A rapid, practical method of detecting and measuring gas in molten-aluminum alloys has been sought for many years. The development of the vacuum gas test appeared to solve this problem. However, the vacuum gas test is subject to certain limitations, and the results must be interpreted mindful of these limitations, as described herein. Since the only gas which is appreciably soluble in aluminum is hydrogen, the terms gas and hydrogen will be used interchangeably in this paper. In the early days of the aluminum industry, when techniques were not as refined as they are today, it was customary to pour small "pancakes" of alumi- num on a metal plate. These were allowed to solidify under atmospheric pressure. Excessive gas in the metal was evidenced by the appearance of small bubbles on the surface of the "pancake".' It is obvious that such a method would only show very high gas levels far above any which could be tolerated by today's standards of quality required for most applications. Various attempts were made to develop methods which would be more sensitive. Among these were measurements of density of aluminum alloys solidified under controlled conditions, radiography of cast slabs, and examination of fractures. The method which showed the greatest promise and appeared most practical was the solidification of a sample of the molten aluminum alloy under vacuum. The samples so solidified are evaluated either by observation for bubble emission during solidification, by sectioning the solidified sample and examining for porosity, or by determining the density of the solidified sample. This test is known internationally as the Straube-Pfeiffer test or the reduced-pressure solidification test, but in America it is commonly referred to as the vacuum gas test2-' and will be so designated in this paper. The determination of density appeared an attractive and convenient means of providing a numerical rating for the samples. Attempts were made in some instances to correlate densities to hydrogen contents. Unfortunately, the visible loss of gas during solidification of samples with high gas contents precluded a correlation of vacuum gas test densities with the observed gas evolution. This led to Alcoa's use of two different pressures; 50 mm of

Jan 1, 1964

By D. H. Gurinsky, D. G. Schweitzer

The empirical equilibrium constantsd the heat of reaction for the reduction have been determined from 300° to 500°C. The mechanisms of the oxidation of uranium and magnesium from dilute bismuth solutions have been investigated. The kinetics of some of the reactions were affected by air adsorbed on the uranium oxides. THE Liquid Metal Fuel Reactor studied at Brook-haven National Laboratory uses a solution of uranium in bismuth as the fuel. Corrosion and stability problems require that the fuel solution contain corrosion inhibitors and deoxidants. Of the additives tested, magnesium has been found to be the best deoxidant and zirconium the best corrosion inhibitor in bismuth. Some experiments1 indicate that magnesium may play a secondary role in increasing the effectiveness of zirconium in inhibiting corrosion. The work reported here was intended to determine the relative concentrations of Mg and U required to prevent oxidation of the uranium fuel in the event of an air leak. In addition, an attempt was made to identify some of the mechanisms of the oxidation and reduction reactions occurring in the liquid bismuth solutions. EXPERIMENTAL Apparatus and Procedure—The equipment shown in Fig. 1 was designed so that the reaction kinetics could be studied as functions of temperature, melt composition, pressure, and flow rate. At the start of a run, the 10-liter bell jar (F) is removed by opening the bottom Dresser fitting (L), and the bismuth and additives are placed in a pyrex crystallizing dish (I). The bell jar is replaced, the system is evacuated to 10-5 to 10-6 mm Hg and then the heater (J) is turned on. When the selected temperature is reached, the oxidizing gas is admitted to the system to the desired pressure as read on the manometer (D). Gas is introduced through the slide tube (Q) which has a pyrex frit (R) sealed to its bottom. Stirring is accomplished by immersing the end of the tube in the melt. To maintain a constant pressure under flow the input and evacuation rates must be equalized. To obtain this condition stopcock (C) is closed and flowmeters (B) are matched. The reaction is "stopped" by closing the inlet flowmeter and quickly evacuating the system. To obtain a liquid metal sample, the system is evacuated with the sampler positioned near the surface of the liquid. The sampler, Fig. 2, is then immersed until its thermocouple reading matches thermocouple (G) Fig. 1. The system is then pressurized with helium which forces the solution through the frit. Twenty to 30 min elapsed between the time the system was evacuated and sampled. During evacuation and pressurization of the system, the temperature variations did not exceed ± 5°C. While bubbling occurred the temperature remained constant to within 1/2 oC. The volume of the equipment was determined by adding a known volume of air at atmospheric pressure to the evacuated system and noting the resulting pressure.

Jan 1, 1962

By Milton E. Wadsworth, John N. Ong, W. Martin Fassell

The temperature and oxygen concentration dependence on the reaction of sphalerite in oxygen at pressures from 6 to 640 mm Hg have been investigated in the temperature range 700° to 870°C. Sphalerite has been found to oxidize linearly over this entire temperature and pressure range. At higher rates, considerable self-heating was observed. The dependence of the oxidation rate on the oxygen concentration may be expressed by: Rate = k' where k' is the specific rate constant and 0 = K1[O2]/(1 + KI[O2]), where K1 is the equilibrium rate constant for the adsorption of oxygen on sphalerite. Values of the activation energy and the enthalpy of adsorption were found to be 60.3 and —59.6 kcal per mol, respectively. Data have been presented that indicate conditions under which either a sulphatizing or an oxidizing roast may be obtained. ROASTING processes have been employed in the metallurgy of a great number of basic metals in use today. Until recently, however, little attempt has been made to deduce the fundamental mechanism by which oxidation takes place. Structural changes upon oxidation have been investigated1 = and thorough work has been performed with regard to the reaction products and the thermodynamics of the oxidation reactions. The objective of this investigation was to determine the rate of oxidation of sphalerite and to apply the absolute reaction rate theory in the theoretilcal interpretation of the experimental results." "O Description of Apparatus The apparatus contains the following features: 1—a furnace with gas system, 2—an autographic weighing system, and 3—a temperature control system. Furnace With Gas System—The furnace was mounted by a yoke on each end and was free to travel up and down on steel guide rods, permitting the introduction and removal of samples from the apparatus with minimum difficulty. The furnace was heated by three sections of Kanthal wire wound externally on a zirconia tube and insulated with Fiber-frax. Pure gases or mixtures of gases were introduced into the lower end of the furnace through two flowmeters, a mixing chamber, and a gas-tight sparger, Fig. 1. The gaseous products of the reaction and the unused gases were prevented from escaping to the atmosphere by maintaining a small negative pressure on the system. This procedure permitted air from the outside to enter through the small an-nulus around the quartz suspension fiber and prevented the escape of the internal gases. Autographic Weighing System—The autographic weighing system consisted of an adapted gold balance, a photoelectric circuit, and a weight recorder, Fig. 2. The sample was suspended from one end of the balance beam by means of a platinum chain outside the furnace and a long thin quartz fiber on the inside of the furnace. On the other end of the balance beam, the necessary weights were added to counterbalance the sample. In addition, a calibrated chain was added to the balance which was linked directly to the weight recording unit. Control was effected by means of a photoelectric circuit." A parallel beam of light from a strong source was directed onto a small front-surfaced aluminum mirror mounted in the middle of the balance beam and then to a front-surfaced mirror mounted on the ceiling of the room. The beam was then reflected onto a front-surfaced right prism at a distance of 9 ft from the balance

Jan 1, 1957

By H. M. Katz, J. L. Speirs, F. B. Hill

Jan 1, 1961

By C. S. Samis, D. P. Seraphim

In the presence of oxygen, galena is oxidized in an aqueous medium containing ammonium acetate in accordance with the following reaction: PbS + 1/2 0, + 2 NH~Ac -» PbAc, + So + 2 NH: + H2O. This reaction was studied in an autoclave under oxygen pressure by polarographic measurement of the concentration of lead ion in solution. Oxidation rate is parabolic. The reaction products are lead acetote and elemental sulfur, the latter forming a film on the galena crystal. Reaction kinetics appear to be determined by the diffusion of lead from the galena phase to the solution interface through the film of sulfur. RECENT applications of pressure leaching on a commercial scale have emphasized the metallurgical importance of the oxidation of sulfide minerals in aqueous media by oxygen under pressure. Research experience has shown that sulfates, thio-nates, and metal oxides are the products of humid oxidation of sulfide minerals in basic solutions. The metal oxides may be solubilized in the presence of a complexing ion.' Certain oxide minerals are also amenable to treatment by this type of oxidation.' Several kinetic studies have been made of reactions of this type. Stenhouse' found that pyrite (FeS?) oxidized in a sodium hydroxide solution to produce a soluble sulfate and an insoluble layer of iron oxide and that the kinetics of the reaction were controlled by oxygen diffusion through the oxide layer. Andersen' studied the oxidation of galena (PbS) in a sodium hydroxide solution under oxygen pressure. He found the lead and sulfur to be simultaneously oxidized to produce plum bite ions and sulfate ions, and that the reaction rate was proportional to the surface area of the galena. The kinetics appeared to be determined by the reaction of oxygen and water with the galena surface. In each of the above reactions the metal and sulfur components of the mineral were oxidized simultaneously and at equal rates. In the present study the alkaline medium has been replaced by an approximately neutral solution of ammonium acetate in which lead sulfate is soluble. Under the conditions no sulfate was produced and no sulfur could be detected in the solution. The metal component of the mineral was solubilized and the liberated sulfur remained as a film of elemental sulfur on the surface of the galena crystal. A kinetic study of this reaction was undertaken to establish the kinetic mechanism and to determine the rate controlling step. Specific reaction rates were determined by oxidizing crystals of galena of measured surface areas in an autoclave maintained at the desired temperature and oxygen pressure. The rate of reaction was followed by measuring the concentration of lead in the ammonium acetate solution with a cathode ray polarograph, employing stationary platinum electrodes incorporated in the autoclave. Experimental Materials and Equipment Crystals of galena obtained from Violamac Mines, Retallack, B. C., were selected for oxidation because

Jan 1, 1957

By J. E. Andersen, J. Halpern, C. S. Samis

In the presence of oxygen, galena is oxidized in an aqueous medium containing sodium hydroxide, in accordance with the following reaction: PbS + 2O2 + 3OH ? HPbO2 + SO4 = + H2O A novel method was devised for following this reaction which takes place in an autoclave under oxygen pressure, by measuring the concentration of HPbO2- in the solution with a cathode ray polarograph employing stationary platinum electrodes. Using this procedure a study was made of the kinetics of the reaction, in which the effect of a number of variables, including temperature, oxygen pressure, NaOH concentration, and agitation, on the rate, were determined. The results of this study are described and discussed, in terms of possible mechanisms for the reaction. IT is known that sulphide minerals can be oxidized in aqueous media in the presence of oxygen under pressure. Reactions of this type can be classified in two categories depending upon whether the products of oxidation are insoluble or soluble in the aqueous medium. An example of the first type of reaction is the oxidation of pyrite in alkaline solutions, giving rise to an insoluble iron oxide. The kinetics of this reaction have been investigated and its mechanism established.' It has been applied in the oxidation of iron sulphides present in refractory gold ores to improve the subsequent recovery of gold by cyanida-tion.' The second type of reaction finds application where it is desired to recover the metal by leaching during oxidation. In this case a medium is selected in which the oxidized mineral is soluble. For example, nickel, copper, and cobalt sulphide ores can be treated by oxidation in the presence of a solution of ammonia dissolving the nickel, copper, and cobalt as the metal ammine sulphate salts.:' A similar treatment has also been reported to apply to zinc sulphide ores.' It is known that lead sulphate is soluble in solutions of sodium hydroxide or ammonium acetate. It should therefore be possible to dissolve the lead from galena ores by aqueous oxidation with oxygen under pressure using either of these solutions. The applicability of the ammonium acetate treatment has been investigated and confirmed in a recent study.' The present paper describes the results of a similar study in which the kinetics of the oxidation of galena in sodium hydroxide solutions have been investigated. The object of this investigation was pri- marily to obtain information of a fundamental nature relating to the kinetics and mechanism of this reaction. The use of pulps of comminuted ore in a study of this type is undesirable because of the difficulty in controlling and measuring the surface area. The measurements were therefore made using galena crystals of measured surface area and in this way absolute reaction rates were obtained. The reaction was found to proceed as follows: PbS + 2O2 + 3OH ? HPbO2 + SO4 + H2O [I] The products are sodium plumbite and sodium sulphate salts which dissolve in the aqueous solution. The reaction studies were carried out in an autoclave in which a desired pressure of oxygen was maintained. The reaction was followed by measuring the concentration of lead in the solution with a cathode ray polarograph. The results obtained in this investigation are presented and discussed in the present paper. Experimental Materials: The crystals of galena were obtained from Violamac Mines, Sandon, B. C. The following impurities were indicated by spectrographic analysis carried out by the Provincial Assay Office, Victoria, B. C.: Sn, 0.02 to 0.2 pct; Sb, 0.07 to 0.7; Zn, 0.01 to 0.1; Si, Fe, Mg, As, less than 0.05 pct; Ag, 126 oz per ton, the latter determined by standard fire-assaying procedure. After cutting a specimen measuring approximately 5x7x12 mm, the crystal surfaces were ground parallel to the 100 axes using No. 2 emery, washed with water, and the dimensions were measured with a micrometer. Oxygen gas was of commercial grade and supplied in cylinders by Canadian Liquid Air. Solutions were made up with chemicals of chemically pure grade and distilled water. Equipment: Autoclave: The autoclave used in these studies was constructed of stainless steel and

Jan 1, 1954

By Milton E. Wadsworth, R. Ted Wimber

Cobalt sulfate solutions containing ammonium acetate and chloroplatinic acid were reduced by hydrogen in a pyrex-glass lined autoclave in the temperature range of 170o to 232°C and hydrogen partial pressure range of 115 to 830 psia. The reduction rate was directly proportional to the hydrogen partial pressure and surface area of the pyrex glass and was independent of the quantity of chloroplatinic acid added initially. Experiments involving the variation of the relative concentration of ammonium acetate indicated that the reducible cobalt complex was probably the diacetate complex of cobalt, Co(AC)4H20, or a new mononcetate complex Co Ac, which was in solubility equilibrium with a pink precipitate of CO(AC)-4HzO. THE reaction in which a metal is dissolved by an acid to produce gaseous hydrogen and a salt solution was discovered early in the history of chemistry. In 1859 Beketoff found experimentally that this reaction could be reversed; i.e., a salt solution could be reduced by gaseous hydrogen to produce a metal and an acid. A review of work done on this phenomenon since that time may be found elsewhere., The hydrogen reduction of a cobalt salt solution is facilitated by complexing the cobalt ion. An ammonia complex of cobalt has been reduced commercially in the recovery of cobalt metal. A new reducible complex of cobalt was discovered5 when it was found that a co-baltous sulfate solution containing ammonium acetate could be reduced by hydrogen at temperatures in the region of 200°C. When a cobalt sulfate-ammonium acetate solution was heated to a temperature below its normal boiling point, a violet color became apparent, indicating complex formation. The nature of this complex was investigated5 by the addition of NH4Ac to a CoSO4 solution maintained at 85o. During the first additions of NH, Ac, the pH of the solution remained fairly constant at about 5.85. However, as the ratio of NH,Ac to CoSO, approached two, the pH rose and then leveled off at about 6.05. The absorption spectra of a Co(Ac), solution and a CoSO,, NH Ac solution were obtained at 85°C and were compared and found to be the same. These findings suggested that the diacetate complex of cobalt, Co(Ac),.4H20, was formed at 85°C. When a cobalt sulfate-ammonium acetate solution was heated to a temperature above about 165o, a finely divided pink precipitate appeared. The X-ray diffraction pattern of this precipitate indicated that it was Co(Ac), - 4H,O. In addition, it was discovered that when chloroplatinic acid, H,PtCl;, was added initially to the cobalt sulfate-ammonium acetate solution, a faster reduction was obtained. The roles of the solution complex, pink precipitate and chloroplatinic acid in the reduction process were then investigated. APPARATUS The experimental work was carried out in a two-liter stainless-steel autoclave. Adetailed description of the autoclave and the auxiliary equipment used in maintaining constant temperature and pressure may be found elsewhere.= Because the stainless steel was corroded, and also because it acted as a hydrogena-tion catalyst, an all-glass liner was fabricated such that the solution came only into contact with flame-polished pyrex glass. EXPERIMENTAL PROCEDURE The solutions used in the experimental work were prepared by dissolving reagent grade chemicals in distilled water. Although variation of the brand of ammonium acetate appeared to have no effect on the experimental results, CoSO, 7H O from the J. T. Baker Chemical Co. of Phillipsburg, N. J.,was found to give faster reductions than that prepared by Allied Chemical and Dye Corp., N. Y. The former was used throughout the course of the experimental work and was weighed up at the outset of each experiment. The ammonium acetate was dissolved to form a 6M stock solution, which was stored under refrigeration. A 10 pct solution of chloroplatinic acid (J. T. Baker Chemical Co.) was diluted to a 1.15 x 1Q2 M stock solution, which was standardized by precipitation of K,PtCl, as outlined by Scott. The appropriate volume of the chloroplatinic acid, H,PtCl,, solution was pipetted into the clean, dry glass liner. The cobalt sulfate-ammonium acetate solution, which had previously been saturated with

Jan 1, 1962

By P. Marier, T. R. Ingraham

When anhydrous cupric sulfate is heated in a stream of nonreactive gas, cupric oxysulfate is formed. When this reaction is complete, the cupric oxysulfate then decomposes to cupric oxide, which is the normal end product of reaction. The kinetics of each of these reactions has been studied using pellets prepared from finely divided cupric sulfate and from finely divided cupric oxysulfate. In each case, the reactant-product interface within the pellet is well-defined, and by normalizing the decrease in inter facial area with the weight fraction of the pellet decomposed it has been shown that the interface migrates into the pellet at a uniform rate at constant temperature. The activation energy estirnated for the decomposition of cupric sulfate is 57 ± 7 kcal per mole and that for cupric oxysulfate is 67 ± 8 kcal per mole. The rate of interfacial reaction is flow-sensitive, increasing with increasing flow of a sweep gas in the range from 50 to 2000 cu cm per min. The rate of decomposition is retarded by sulphur trioxide in the sweep gas. The relationship between sulfur trioxide Partial pressure and rate is consistent with the Langmuir Adsorption Isotherm governing the retention of sulfur trioxide in the interfacial layer. MOST of the work reported on the thermal decomposition of cupric sulfate and of cupric oxysulfate is related to the thermodynamics of the individual decompositions.'-4 No information is available on the rates of the individual reactions, on their activation energies, or on the effects of gas flow or of partial pressure of sulfur trioxide on the reaction rates. Some dubious conclusions based on relative reaction rates have been drawn from studies on the oxidation of sulfides. For example, on the basis of the products observed in a quenched system, it has been suggested5" that, when copper sulfide is roasted, the cupric oxide which is formed arises from the decomposition of either cupric sulfate or cupric oxysulfate. This is very difficult to reconcile with the thermodynamics of the C-S-O System,3 which support the probability that both cuprous oxide and cupric oxide precede cupric oxysulfate and cupric sulfate in the sequence of products formed during the oxidation of cuprous sulfide. In this paper the individual sulfate decomposition reactions will be examined and suggestions will be advanced for the mechanism of reaction. This will be based on the effects on reaction rate of changes in interfacial area, flow rate, and back-pressure of sulfur trioxide. EXPERIMENTAL Materials. The copper source material used in all experiments was Fisher Certified Reagent Grade anhydrous cupric sulfate, for which the following analysis was supplied by the manufacturer: chloride, 0.002 pct; alkalies and earths, 0.2 pct; insoluble, 0.002 pct; iron, 0.020 pct. Cupric oxysulfate was prepared by roasting cupric sulfate for 48 hr, with frequent rabbling, in an open tray in a muffle furnace maintained at 725°C. At this temperature, cupric sulfate develops about 0.16 atm pressure of combined sulfur trioxide, sulfur dioxide, and oxygen. Cupric oxysulfate develops only about 0.05 atm pressure, and is stable until

Jan 1, 1965

By R. Bainbridge

THE Consolidated Mining and Smelting CO. of Canada Ltd. has operated a lead smelter at Trail, B. C., for many years. In order to take advantage of metallurgical advances, as well as to improve materials handling methods, this company, commonly known as "Cominco," commenced planning a program of smelter revision and modernization some years ago. The first stage of this program involved the design and construction of a new blast furnace gas cleaning system. The selection of equipment, the design of facilities, and preliminary operating details of this system will be dealt with in this paper. The essential problem was to clean and collect 100 tons of dust daily from 153,000 cfm* (12,225 lb per min) of lead blast furnace gas which varied in temperature from 350º to 1100°F. Because it was desired to collect the dust dry, either a Cottrell or a baghouse cleaning plant was to be selected. Comin-co's many years of experience with both systems provided a background for choosing the most satisfactory installation. All information pertinent to the two methods of dust recovery was carefully investigated, and it was decided to replace the existing equipment with a baghouse. Very briefly, the reasons for this decision were as follows: 1—A baghouse installation would be practical because the SO2 content of the gas was low and corrosion would not be a problem if the baghouse operating temperatures were held sufficiently above the dew point. 2—Variations in the physical characteristics of fume and dust, which are inherent in this blast furnace operation, should not substantially affect the operating efficiency of a baghouse. 3—For the same capital cost, metal losses (stack and water losses) would be appreciably less in a baghouse. 4—A baghouse would be easier to operate, and would not require the use of highly skilled labor. 5—Operating and maintenance costs of a bag-house would be lower. 6—The only available space for reconstruction was relatively small, and not suited to a Cottrell installation. Once the baghouse system was decided upon, detailed design of the installation was begun. Baghouse Design Gas Cooling: Before the required capacity of the baghouse could be determined, the method of cooling the gas to the temperature necessary for bag-house operation had to be chosen. The problem confronting the design engineers was how best to cool 153,000 cfm of gas from a temperature ranging from 350°F to brief peaks of 1100°F, down to 210°F, the maximum safe baghouse inlet temperature. A survey of existing blast furnace gas temperatures in the outlet flue showed that the normal range was as given in Table I. The obvious choices of cooling method were: 1— cool completely by the addition of tempering air; 2—utilize a heat exchanger; 3—cool by radiation; and 4—cool with water spray in conjunction with the admission of tempering air. The advantages and disadvantages of the various cooling methods were: Air Addition: To cool completely by the admission of tempering air involved tremendous volumes, Fig. 1. For example, to cool 1 lb of blast furnace gas at 450°F requires 1.84 lb of air at 80°F or 1.60 lb at 60°F. As it is necessary to design for peak conditions, it can readily be seen that volumes of tempering air in the order of 1,500,000 cfm would have to be handled. Using the normal design figure of 2.5 cu ft per sq ft of bag area, a baghouse installation comprising some 600,000 sq ft of filter cloth would be necessary. Such design requirements would be prohibitive, not only from a standpoint of capital expenditure, but also because of space limitations. Heat Exchanger: The utilization of a heat exchanger was given serious consideration. A horizontal tube unit using air as the medium to cool the required volume of blast furnace gas from 400" to 250°F was investigated. Cooling above 400°F would be done by water spray, and below 250°F by admission of tempering air. The estimated capital cost of such a unit was found to be prohibitive. From an operating standpoint, there was considerable doubt as to whether the soot blowing equipment provided would effectively keep the dust from building up on the tube surface. The performance of heat exchangers operating on dusty gas in other company operations had not been too favorable. Radiation Cooling: Although somewhat cumbersome, gas cooling by radiation from 'trombone' tubes or other similar equipment (cyclones) is employed in many metallurgical operations. Such an installation was also considered. However, calculations showed that an installation much larger than the space available would be required to handle the gas volume involved. For example, to cool 153,000 cfm of blast furnace gas from, say, 600' to 250°F (i.e., remove in the order of 58,500,000 Btu per hr with heat transfer rates varying from 1.1 Btu per sq ft per hr per OF for the higher temperature ranges to 0.88 Btu per sq ft per hr per OF for the lower ranges) would need a cooling area of some 175,000 sq ft.

Jan 1, 1953

By John Wulff, David A. Stevenson

The solubility of iron and cobalt in liquid lead, the solubility of chromium in liquid tin, and nickel in liquid silver were determined up to 1300°C in an atmosphere of hydrogen. Liquid samples were taken at different temperatures and analyzed. The data obtained were plotted to show the relationship between log mole fraction of solute and the inverse of the absolute temperature. Resulting values for the differential heat of solution (AH) and the excesspartial molar entropy (S)are: for the iron-lead system = _24,200 cal per mole and AT = 5.0 e.u. for the a phase, and AH = 18,300 and S = -0.1 e.u. for the y phase; for the cobalt-lead system AH = 16,500 cal per mole and AS =0.7 e.u.; for the chromium-tin system AH = 15,600 cal per mole and AS = 8.5 e.u.; for the nickel-silver system AH =18,000 cal per mole and AS = 4.0 e.u. RECENT developments in coating, brazing, and heat transfer have focused attention on specific cases of solid-liquid interaction in metal systems, particularly those involving dilute liquid solutions. One fundamental quantity of interest in such systems is the equilibrium phase distribution between solid and liquid phases. In any effort to fully understand such interactions, one must be able to describe the equilibrium state precisely. Furthermore, in order to develop genera! relations for such systems, one must know the equilibrium state in a number of different cases. This investigation was devoted to a study of the equilibrium phase distribution in four binary systems: solid iron-liquid lead; solid cobalt-liquid lead; solid chromium-liquid tin; solid nickel-liquid silver. The iron-lead system was previously studied by Shepard and parkman' and later by Fleischer and Elliott2 with differing results. Both investigations were carried out by sampling the saturated liquid, and analyzing it chemically. The cobalt-lead system was previously reported by Pelzel,3 who noticed no discontinuity near the transformation. A limited investigation of the chromium-tin system was made by Wlodek and wulff4 using a weight loss method. This work was done in conjunction with the development of a process for coating molybdenum with chromium from tin solutions. To more fully under- stand the kinetics of the process, a more accurate determination of a solubility was desirable. No significant data were found in the literature on the nickel-silver system. Data on the latter system were, however, relevant to the investigation of the corrosion kinetics of copper-nickel alloys in liquid silver as carried out by Harrison and wagner,5 EXPERIMENTAL APPARATUS AND PROCEDURE Of the techniques available for the study of such systems the most effective appears to be that of liquid sampling,2,6,7 which is the method used in the present investigation. A schematic diagram of the sampling furnace is given in Fig. 1. Large pieces of solute metal, in excess of saturation solubility, were added to 200 g of the solvent metal, and the entire charge placed in an alundum crucible. Analytical reagent or materials of chemically pure grade were used in all investigations. The crucible was contained in a zircon muffle tube positioned vertically in the uniform hot zone of a Globar resistance furnace. The tube was provided with a number of ports which permitted sampling at temperature and in controlled atmospheres or vacuum; the ports were also useful for evacuation and introduction of gases as well as for thermocouple measurements. A chromel-alumel thermocouple Sheated in a porcelain tube was used for temperature measurements. A purified hydrogen atmosphere was employed in all experiments reported in this paper. The hydrogen was purified by passing it first over hot copper gauze to remove oxygen and then through a molecular sieve immersed in

Jan 1, 1962

By M. J. Spendlove, H. W. St. Clair

The problem frequently arises, particularly in refining metals or smelting scrap metals, of separating metals in the metallie state. Many metals may be separated by taking advantage of their difference in vapor pressure. Such separations can be made at atmospheric pressure, but the separations are much more selective and can be carried out at considerably lower temperatures if the distillation is done at pressures of a few millimeters or less in an evacuated enclosure. Until recently, this has not been considered feasible as a metallurgical operation, but the recent improvemcnts that have been made in vacuum technology have broadened the applicability of vacuum processes and have prompted re-examination of low-pressurc distillation of metals as a practicable process. The distillation of zinc from lead is one separation that has already been reduced to practice.l This paper is the first of a series of studies being made on separation of nonferrous metals by distillation at low pressures. Although these experiments were confined to the separation of zinc from aluminum, the significance of the results is by no means confined to these two metals. The purpose has been to investigate a metallurgical technique rather than merely to devise a means of separating specific metals. The experimental work on distillation of zinc from zine-aluminum alloys at reduced pressure grew out of earlier work on distillation at atmospheric pressure.2 The earlier work indicated that it would not be practicable to decrease the zinc in the alloy much below 10 pct owing to the high temperature required. Therefore attention was turned to distillation ah low pressures, at which lower temperatures are required. After preliminary tests were made in a small, evacuated tube furnace, a larger furnace having a capacity of 100 to 150 Ib of metal per charge was constructed. Distillation tests were first made on pure zinc and then on aluminum-zinc alloys of various composition. Particular attention was given to the limit to which zinc could be reduced in the residual metal. Data were also taken on the rate of evaporation, and heat balances were made for both the crucible and the condenser. Distillation Furnace The vacuum-distillation unit is illustrated schematically in Fig 1. The major components are the induction furnace, the condenser, the vacuum system, and the power-conversion unit. Power is supplied to the induction furnace from a 50-kw 3000-cycle motor-driven alternator. The pressure in the furnace is reduced by a vacuum pump having a nominal pumping speed of 10 liters per sec. When in operation, the metal vapors travel upward from the furnace to the water-cooled condenser where they are collected in amounts of 50 to 100 lb. The condenser is removed with aid of an electric hoist. When the system is under vacuum, the condenser is made self-sealing by a rubber gasket between the smooth-faced, water-cooled flanges at the top of the furnace and the bottom of the condenser. The pressure of the atmosphere is more than sufficient to insure sealing. At the conclusion of an experiment, the residual metal can be removed from the furnace by removing the condenser and tilting the furnace with the electric hoist. The metal was cast into the molds carried on a mold truck. A photograph of the furnace and auxiliary equipment is shown in Fig 2. The details of the vacuum furnace are illustrated in Fig 3. The furnace proper is made vacuum-tight with rubber gaskets placed at each end of a fused quartz cylinder. A clamping plate at the bottom and a ring at the top are made to squeeze the rubber between the metal and the end of the quartz tube. A large graphite crucible placed inside the quartz cylinder is thermally insulated and physically supported by refractory insulating bricks. A thermocouple in a quartz protection tube is located at the bottom of the crucible: the leads pass through a rubber seal in the bottom plate. The supporting structure for the furnace is an angle iron frame with transite board sides. The condenser is made in the form of a water jacketed cylinder with an opening to the vacuum line at the top. The bottom has a projecting skirt inside the machined flange to provide additional cooling for the rubber gasket. Condenser sleeves are made in the form of two semicylindrical pieces of sheet metal that fit snugly inside the cooling jacket. The split sleeve facilitates removal of the condensate. Measurement of Temperatare and Pressure The metal temperature was measured by a platinum-platinilm rhodium thermocouple inserted in a well extending up into the bottom of the graphite crucible. During rapid evaporation there is a wide difference in temperature between the surface and the main body of metal in the crucible because of the large amount of heat that must be conducted to the surface to supply the heat of evaporation. The heat of

Jan 1, 1950

By A. Steiner, E. Miller, K. L. Komarek

Equations have been derived to calculate the chemical potentials of the components of liquid binary alloys from liquidus and enthalpy data. The equations are applicable to systems with intermetal-lic compounds of limited solid solubility. The liquidus curve of the Mg-Sn system was accurately redetermined, and the melting point of the compound Mg2Sn was found to be 770.5° * 0.3°C. The thermo-dynamic properties of the liquid alloys were calcu -lated from the liquidus data. The activity of tin shows both positive and negative deviations from Raoult&apos;s law and the relative integral entropy is positive. THE phase boundaries in an equilibrium phase diagram are functions of the thermodynamic properties of the coexisting phases. Specifically, the shape of the liquidus curve near an intermediate compound is determined by the properties of the compound and the liquid phase. Hauffe and wagnerl have derived an equation for the chemical potential differences (ui) of the liquid alloys near the congruent melting point when the heat of fusion of the compound and the liquidus curve are known. In their derivation the temperature dependence of the ui value was neglected, and liquid of stoichio-metric composition was chosen as the reference state. To obtain chemical potentials over the ntire composition range of the phase diagram with the pure components as the reference state, equations have been derived which incorporate the temperature dependence of the chemical potentials. Such a thermodynamic analysis can be carried out when the liquidus curve and the enthalpy values are known. The equations have been applied to the redetermined liquidus curve of the Mg-Sn system to obtain the thermodynamic properties of liquid Mg-Sn alloys. The Mg-Sn phase diagram has been studied by various investigators and the results have been compiled and critically assessed by Hansen.2 Preliminary thermal analyses showed that the basic features of the diagram were correct. The system has one congruent melting compound, Mg2Sn, two eutectics, and some solid solubility of tin in magnesium. However, the published liquidus temperatures show considerable scatter. In order to perform an accurate thermodynamic analysis, the liquidus curve between the magnesium-rich and tin-rich eutectic compositions was therefore redetermined by careful thermal analysis using highest-purity starting materials. EXPERIMENTAL PROCEDURE The magnesium metal (Dominion Magnesium Ltd, Toronto, Canada) had a purity of 99.99+ pct with the following impurities (in ppm): 20 Al, 30 Zn, 10 Si, <1 Ni, <1 Cu, <10 Fe; tin (Cominco Products, Spokane, Wash.) contained 99.999+ pct Sn. The graphite crucibles were machined from graphite rods (United Carbon Products Corp., Bay City, Mich.) which had an average total ash content of less than 30 ppm. All the graphite parts were baked out in vacuum at 950" to 1000°C for a minimum of 24 hr to remove volatile organic materials. The graphite crucibles (3 in. long, 1-5/8 in. OD, 1-3/8 in. ID) were tightly closed with a threaded cover (1/2 in. thick) which had a central thermocouple well (2-1/4 in. long, 5/16 in. OD, 3/16 in. ID). Cover and thermocouple well were machined from one piece of graphite. A thin sheath of quartz was inserted in the well to protect the thermocouple

Jan 1, 1964

By H. W. Mossman

Three aspects of magnetite smelting are discussed. The first is the working out of equilibrium conditions for eliminating sulfur. The second is the influence of magnetite solubility on the difficulty of tapping the reverb matte. The third is an approximation of the equilibrium conditions in the reverb gases which govern whether magnetite is mode or reduced in the reverb slag by these gases and by any iron sulfide in the slag. MAGNETITE has had a varied history in the Hurley Smelter since its start in 1939. Magnetite determinations on the smelter products are made regularly only on the monthly composite samples. Variations on the monthly averages are shown in Table I. Magnetite which drops from the slag and matte in the reverb has some slight bottom buildup which comes and goes, but no substantial accumulation from this source has been found at the end of a normal nine months&apos; furnace campaign. However, there has been some low grade magnetite bearing material mixed with considerable A1,0,, which has slid down from the bottom of the sloping flue between the reverberatory furnace and the waste heat boilers. This accretion has required drilling and blasting near the skimming end of the furnace. The magnetite has interfered with tapping at times. When the smelter was first started, tapping trouble from magnetite was extremely severe. Increasing the reverberatory furnace temperature by putting in an air preheater and a Dutch oven has helped greatly, although there still is occasional tapping trouble. When the present series of physical chemistry articles on copper smelting started coming out in 1950, they were read with interest, but no immediate application was seen for them. Results of some laboratory work in 1952 aroused a much stronger interest in this physical chemistry. A series of melts was made on some converter slags, which had magnetite in very large grain sizes, with the object of reducing the grain sizes in the slag, as it was known that it was easier to handle in the reverb in that condition. Anything done in the tests greatly reduced the grain sizes—even in the controls, where nothing was done except melt the slag and cast it. There was more magnetite in the slags after the tests than before, and with wide variations. There were no obvious reasons lor much of what happened in these tests. Much of the base material published in English in this field was made available for study. Recalculations were made on many of the type problems, and part of the data was reduced to local temperatures and compositions. Explanations were found for what happened in the 1952 series of tests on converter slags, and the same principles turned out to be a description of much of what magnetite does in the reverb. This article is to present the results of that study, from the viewpoint of applying the technical material in definite numerical form to the operating conditions in both the converters and the reverberatory furnaces at the Hurley smelter. Table I. Magnetite Variations on Monthly Averages, 1939 to 1955 Pet Magnetite Lowest Highest Average Converter slag 13.6 43.3 25.4 Roverb slaa 2.7 20.9 8.7 Matte 28 15.9 98 In general it was found that magnetite is made or reduced in both the converters and the rever-beratory furnace, depending on variations of temperature, matte composition, and reverb gas composition occurring in ordinary plant operation. Within reasonable limits, the field conditions for formation or reduction can be predicted, and probably can be set up and maintained. Converter conditions affecting magnetite formation can be put into numerical values better than for the reverb from purely technical calculations. The converter can be operated so as to keep the magnetite in the slag down to between 12 and 14 pct and still give satisfactory life for the converter brick. This depends upon having converter flux available which will make a slag with a good separation without raising the temperature too high. In the Hurley reverb and others with similar conditions, it is likely that a compromise of conditions will give a reasonably good control of combustion and still keep the magnetite from building up on the bottom. This discussion consists of three main parts. The first is the working out of the equilibrium conditions in the converter for determining in which direction the reaction 3 Fe,,O, (s) + FeS (1) F? 10 FeO (1) + SO, will go under actual converter operating conditions. The second deals with the influence of the solubility of magnetite in the slag and matte in the reverb on the difficulty of tapping matte. The third is an approximation of the equilibrium conditions in the

Jan 1, 1957

By Roshan B. Bhappu, Ronald J. Roman, Dexter H. Reynolds

The reaction between manganese dioxide and molybdenite in a water- sulfuric acid medium was studied at atmospheric pressure and from 25° to 103°C. Both solids are dissolved to give, as final products in solution, Mo(VI), S(VI), and Mn(II). Stoichiornetric relationshifis are most simply represented by the equation MoSz + YMnO2 + 15H+ - HMoO-4 + 2HSO-4 + 9Mn++ + 6H20 The rate of this reaction was found to be dependent on temperature, manganese sulfate concentration, acid concentration, and, in a complex way, the surface areas of both manganese dioxide and molybdenite. Both the total amount of manganese dioxide and tnolybdenite and the ratio of manganese dioxide to molybdenite played a part in determining the over-all rate of the reaction. The effect of acid concentration on the rate was also dependent on the ratio of manganese dioxide to molybdenite. SEVERAL reports in the literature describe reactions using manganese dioxide in water-sulfuric acid media as an oxidizing agent. Nearly all previous work describes interactions between an ion and manganese dioxide, such as the oxidation of ferrous to ferric iron,&apos; or the interaction between a molecule in solution and manganese dioxide, such as the oxidation of toluene to benzaldehyde.2 Only a few references describe the use of solid manganese dioxide to oxidize another solid in sulfuric acid. One report3 states, "Manganese dioxide in dilute sulfuric acid solutions had no beneficial effect (in leaching copper from chalcopyrite), but in more concentrated acid solutions the rate of solution was about the same as that for permanganate." In 1902 a German patent4 described a process which used manganese dioxide in strong (50°Be&apos;) sulfuric acid for oxidizing chromite ores to soluble iron and chromium salts. The process consisted of mixing the sulfuric acid with the finely ground ore and manganese dioxide. The pulp was then heated to about 120°C and an electric current passed through it. The manganese dioxide apparently dissolved in the strong, hot acid solution, oxidized the iron and chromium and itself was reduced to manganese 11! then was oxidized at the anode back to manganese dioxide, and the cycle repeated. An extensive study5 of the kinetics of the oxidation of pyrite with manganese dioxide indicated that in dilute sulfuric acid solutions (about 0.1 N) and at elevated pressures the dissolution of the pyrite was brought about by air oxidation of the iron to ferrous sulfate, followed by manganese dioxide oxidation of the ferrous sulfate to ferric sulfate. The mechanism proposed by the authors involved the oxidation of the iron at the surface of the manganese dioxide. Also! the oxidation of sulfide sulfur to sulfate was credited to the oxygen in the air, not to the manganese dioxide. Since the metallurgical literature is lacking in information concerning the use of one solid to oxidize another solid. the current investigation was under-

Jan 1, 1965

By Robert D. Pehlke, Arden L. Bement

A mass transfer coefficient for the removal of hydrogen from liquid aluminum by inert flush degassing has been determined experimentally at 700°C. A mathematical model has been derived, assuming transport control, which adequately describes the experimental data. The removal of hydrogen is shown to increase with decreasing bubble size, and with increasing flow of flushing gas. The presence of hydrogen in liquid metals in amounts which exceed the solid solubility at one atmosphere hydrogen pressure is highly undesirable, causing unsound ingots or castings, or resulting in difficulties during further fabrication. The removal of hydrogen from liquid metals can be accomplished by several methods, the principal two being vacuum treating and inert flush degassing. The design of a process involving either of these techniques is based on the mass transfer coefficient for hydrogen between the liquid metal and a hydrogen-dilute gas phase. In an effort to extend our knowledge in this area, an investigation of the rate of removal of hydrogen from liquid metals by inert gas flushing was undertaken. In view of the difficulties encountered with gas porosity in the casting of aluminum and aluminum alloys, and because of certain experimental advantages, liquid aluminum was selected for the study. PREVIOUS INVESTIGATIONS The removal of gases from liquid metals, particularly steel, has received considerable attention in recent years, The first definitive work of a theoretical nature was presented by Geller1 who considered the case where equilibrium exists between purging gases and metal bath during flushing, and also deviations from this ideal case. Hulme2 has surveyed a number of experimental studies and discussed some of the practical limitations of removing dissolved gases from molten metals. More recently a number of studies have been reported on degassing of steel melts by flushing with an inert gas.3&apos;4 Considerable work has also been reported on aqueous and organic liquid-gas systems, and an excellent summary prepared by Hovis.5 The treatment of liquid aluminum by a flushing gas to remove dissolved hydrogen has been known for several decades.6 A number of gases have been used for this purpose, including nitrogen, argon, and chlorine, or mixtures of these gases. Tikkanen and Erkko7 have studied the relative density changes of aluminum castings poured from melts degassed with chlorine, nitrogen, and mixtures of the two gases. Their

Jan 1, 1962

By Carl Wagner

AT elevated temperatures. most metals react with oxygen, sulphur, or halogen rather rapidly, although a coherent layer of the reaction product is formed and separates the two reactants from each other. The reaction proceeds, since one of the reactants or both dissolve and diffuse across the reaction product. In many cases the rate is diffusion-controlled as is indicated by the validity of the parabolic rate law of Tammann&apos; and Pilling and Bed-worth. wott and Gurney&apos; have suggested that diffusion processes also play an important part in the case of reduction reactions. This is confirmed by observations reported below. In general, however, the rate of reduction reactions is determined not exclusively by diffusion processes. Reduction of Silver Sulphide Above 179°C silver reacts with liquid sulphur very rapidly. Since the volume of silver sulphide is much greater than that of the equivalent amount of silver, a coherent layer of silver sulphide is formed in accordance with a rule established by Pilling and Bedworth.&apos; According to Wagner&apos;,&apos;&apos; silver ions and electrons migrate from the Ag-Ag,S interface across the silver sulphide layer to the Ag,S-S interface where the following chemical reaction takes place: 2 Ag +2 e + S Ag,S If silver sulphide is reduced by means of hydrogen, the equivalent volume of the reaction product, silver, is less than that of the initial substance. Accordingly, no coherent layer of silver is formed. Instead, fine filaments of silver grow from the silver sulphide surface into the ambient gas. This phenomenon has been investigated by many authors, especially by Kohlschiitter," who has also reviewed previous literature. Kohlschiitter has interpreted the observed phenomena by migration of "silver" dissolved in silver sulphide. According to Wagner,&apos; silver ions and electrons are the migrating particles. The sequence of consecutive steps involved in the reduction of silver sulphide is shown in Fig. 1. Every- where at the silver sulphide surface, hydrogen reacts with sulphur so that excess silver ions and electrons are formed, which migrate at random until they are captured by a nucleus of metallic silver. If such a nucleus lies on the surface and material is furnished only from the base, the original nucleus is pushed outward and thus a filament grows. Removal of sulphur and formation of metallic silver may also take place at different points when silver sulphide is reduced by copper. The reaction is usually formulated as: AgL,S + 2 Cu = Cu.S + 2 Ag but the actual reaction products are silver and a solid solution of the two sulphides. Primarily, excess silver and electrons dissolved in silver sulphide are formed. After a certain degree of supersaturation has been reached, nuclei of metallic silver may form. If, however, a silver foil is in contact with the silver-sulphide phase, as is shown in Fig. 2, silver ions and electrons will migrate to the Ag,S-Ag interface where they are captured. To confirm this scheme, a composite sample, with the geometry shown in Fig. 2, was made by pressing silver sulphide tablets 0.6 cm in diam and 0.4 cm high together with a copper cylinder and a silver foil in crucibles of iron, which does not react appreciably with silver sulphide at 400°C. After heating for 3 to 7 hr at 400°C. the specimens were cut normal to the phase boundaries. It was found that the thickness of the silver foil had increased from 0.01 cm as initial value to a maximum value of 0.06 cm. In the interior of the silver sulphide tablets only traces of metallic silver were observed. These observations show clearly that silver ions and electrons may migrate over distances of several millimeters before they form metallic silver. Reduction of Cuprous Sulphide by Iron The behavior of cuprous sulphide is very similar to that of silver sulphide, since the lattices of the high temperature modifications are identical, but in Cu.S there is a variable deficit rather than an excess of metal. To investigate the displacement of copper by iron, cuprous sulphide powder covered by a copper foil was pressed in iron crucibles and heated up to 800 °C for 3 to 7 hr. The thickness of the copper foil increased from an initial value of 0.005 cm to a nlaximum value of 0.029 cm. In the interior of the

Jan 1, 1953

By Carl Wagner

J. Pearson (British Iron and Steel Research Association, London, England)—Dr. Wagner has referred to the work of Richardson and Dancy and Gellner and Richardson on the reduction at 900 °C of wiistite films on iron. He has assumed that iron ions and electrons migrate backwards to the iron-wustite interface and that the reduced iron grows on the metal substrate. This would be a possible explanation of the results of these workers were it not for the fact that, as stated by Gellner and Richardson but not stressed, this growth in the center of the specimen can occur even

Jan 1, 1953

By Tennyson Smith

Jan 1, 1962

By Harry M. Mikami, A. Gene Sidler

Chemistry of the evolution of materials in contact with copper converter tuyeres is delineated by means of analyses of periodic punch rod samples taken during a converter cycle. Lining samples from known peripheral and end locations from three converters were investigated chemically and petrographically. One set of eight specimens representative of all lining parts was removed for study before completion of the campaign. Purely surface removal by slag action is minor. Slag filling of interconnecting cracks extending to the hot face results in a "stoping&apos;&apos; action which shapes the hot face contour and is a significant factor. Major wear mechanism is believed to be copper metal slabbbing caused by metal first penetrating joins and invading fractures or impregnating brick structure laterally from the joints. Importance of lining design and joint fit are emphasized. New laboratory tests are indicated. In the copper converter, molten matte, which is essentially a solution of iron and copper sulfides, is blown with air to produce blister copper. During the first stage of the process the FeS component is oxidized to SOz and FeO, the latter combining with a siliceous flux to form a predominantly ferrous silicate slag. Successive increments of matte and flux are charged, blown, and skimmed of slag until sufficient CuzS ("white metal") has accumulated. This white metal is then blown to copper metal in the final or blister-forming stage by the oxidation of sulfur to SOz and the reduction of combined copper to metal. The reactions of both stages are exothermic so provide the required heat. Temperature is controlled by cold scrap copper, flue dusts, smelter reverts, and so forth. Considerable vari- ation in practices occur depending on the type of converter, grades of the matte and flu, and other factors. Lathe and Hodnett&apos;s&apos; survey of world converter practices gives much information on these variations and their effects on the refractories. Principles of converting chemistry are presented by Newton and wilson,&apos; Ruddle,&apos; and schuhmann. Most copper converting occurs within the temperature range 2100" to 2370°F although individual smelters report1 ranges from as low as 1700" to 2240°F to a high of 2200" to 2400°F. Temperature intervals reported from beginning to end of blow varied from 45" to 540°F. Average temperature range was 2030" to 2260°F and the average interval 250°F During the matte-blowing stage varying amounts of magnetite (FeO.Fe,O,) are formed in the slag as suspended crystals and also the liquid portion of the slag will contain, in addition to ferrous silicate, some Fe20s which adds to the potential magnetite formation if the temperature falls. This is a component of the so-called "dissolved magnetite" often referred to in the literature. Although the slag may contain minor amounts of CaO, A1209, MgO, and alkalies, the principal components are Fe0-Fez03-SiO,. The phase equilibria relationships of this system have been elucidated by Muan~ and Schuhmann, et a1.&apos; These data and estimates of converting atmospheres by schuhmann4 are useful tools in studying slag behavior. It would appear that in many cases reported magnetite contents of slags are apparently higher than actually is present at operating temperature. This is probably because much of it is formed during slag solidification. The quaternary systems which have Mg-0,1,0,&apos; and addition to iron oxides and silica are also of interest in considering the reactions of the slags with basic refractories. The formation of magnetite has an important interrelationship throughout the pyrometallurgy of copper (Mossman,&apos; Gronningsater and Drummond," Ruddle,&apos; Elwood and Henderson") but its presence in the converter particularly relates to operating efficiency and lining life in certain instances. Archi-

Jan 1, 1963