Search Documents

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

From the inception of zinc retorting on a commercial scale in the United States in 1890,' the retort employed has undergone wide variations in its composition and manufacture, facilitating in part equally remarkable improvements in furnace capacities. The early day hand made clay retort was charged with carbonates or silicates or with coarse dead roasted concentrates mixed with a large proportion of charge fuel resulting in a relatively low zinc burden and fired 24 hr in direct coal fired furnaces. Its modern counterpart is fabricated in hydraulic presses from clay mixtures containing sizeable amounts of either silicon carbide or silica flour, charged with sintered flotation concentrates to more than three times the early day zinc burden and fired 24 to 48 hr in gas fired furnaces. This paper does not attempt to describe in detail the early day clay retort practice as it is well outlined in treatises by Ingalls,2 Lodin,3 Liehig,4 Hofman5 and others. A brief review of clay retort practice is presented together with a description of the major developments since 1912. Clay Retorts The Belgian type retort, both in the circular and elliptical forms, has been used almost exclusively. Typical dimensions of press made clay retorts around 1910 are shown in Table 1. Variations in these dimensions were used at different plants according to local conditions to a maximum inside diameter of 9 in. and inside length of 54 in. However, the effective heat penetration in a 24 hr firing cycle and the tendency of the retort to bend limited the retort size. Use of the elliptical vessel was an attempt to present a stronger cross-section resisting the tendency to bend and to increase the burden without increasing the depth of heat penetration. One exception to the 48-54 in. length was the 60 in. retort used as early as 1905 at Palmerton by means of supporting the last 12 in. at the butt end with a specially designed furnace back-wall. This backwall construction with the 60 in. retort had been developed and used at Bethlehem by G. G. Con-vers and A. B. DeSaulles. An attempt was made at Blende, Colo. to use even larger retorts of the Rhenish type based on European practice and requiring much higher furnace temperatures. Satisfactory plastic clays capable of withstanding these temperatures were not available, and the plant never operated successfully. PREPARATION OF BATCH Composition of the clay retort by weight was 40 to 50 pct raw clay and the balance "grog." Generally speaking the mix consisted of 7 parts plastic clay to 9 parts grog by volume. Principal source of the clay used was the Cheltenham vein—sometimes referred to as "St. Louis city clay." A typical analysis was A12O3-31.0 pct, SiO2-50.0 pct, Fe2O3-2.5 pct, MgO-0.3 pct, CaO-1.5 pct and loss on ignition 14.0 pct. At the smelter the clay was weathered whenever possible and then crushed to 0.08 in. or finer. Grog consisted of calcined adobies, cleaned retort scrap and cleaned refuse fire brick such as old furnace brick, blast furnace linings, and others. Saggers from ceramic plants and calcined flint clay were later used. The grog materials were ground to 0.12 m. or finer. Occasionally coke dust up to 10 pct of the mix was substituted for a part of the grog following European practice.² Particle size of the grog has a major influence on the retort properties—the larger the grain, the better can the retort withstand thermal shocks, resist bending at furnace temperatures and resist corrosion from slag; the smaller the grain, the lower the loss of zinc vapor through the retort walls. Grog forms the skeleton of the retort, and the clay shrinks around its grains to act as a binder. In the drying process, the grog has a stabilizing effect on the drying rate, decreasing shrinkage and giving up previously absorbed water to the surrounding clay to minimize the danger of cracking or checking.² Grog and clay were mixed through a horizontal pug mill with 10 to 20 pct water added, depending on whether the retort was to be formed by hand or mechanically, more water being required for the hand process. The batch or "mud" extruded from the pug mill was cut in convenient lengths for handling, stacked in piles or in special rooms, covered with wet burlap and allowed to "rot" or age from 1 to 8 weeks to increase plasticity. HAND MOLDING If the retort was to be molded by hand, the mud was repugged after the rotting period and given to the molders. Their molds consisted of 3 sheet iron or wood cylinders, each one third the retort length and defining the outer shape of the retort. Beginning with the bottom section, mud was placed in the form and tamped with a rammer

Jan 1, 1950

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

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

Jan 1, 1950

By F. A. Forward, J. Halpern

A new process for extracting uranium from ores with carbonate solutions is described. Leaching is carried out under oxygen pressure to ensure that all the uranium is converted to the soluble hexavalent state. By this method), alkaline leaching can be used successfully to treat a greater variety of ores, including pitchblende ores, than has been possible in the past. The advantages of carbonate leaching over conventional acid leaching processes are enhanced further by a new method which has been developed for recovering uranium from basic leach solutions. This is achieved by reducing the uranium to the tetravalent state with hydrogen in the presence of a suitable catalyst. A high grade uranium oxide product is precipitated directly from the leach solutions. Vanadium oxide also can be precipitated by this method. The chemistry of the leaching and precipitation reactions are discussed, and laboratory results are presented which illustrate the applicability of the process and describe the variables affecting leaching and precipitation rates, recoveries, and reagent consumption. THE extractive metallurgy of uranium is influenced by a number of special considerations which generally do not arise in connection with the treatment of the more common base metal ores. Perhaps foremost among these is the very low uranium content of most of the ores which are encountered today, usually only a few tenths of one percent. A further difficulty is presented by the fact that the uranium often occurs in such a form that it cannot be concentrated efficiently by gravity or flotation methods. In these and other important respects, there is evident some degree of parallelism between the extractive metallurgy of uranium and that of gold and, as in the latter case, it has generally been found that uranium ores can best be treated directly by selective leaching methods. It is readily evident that this parallel does not extend to the chemical properties of the two metals. Unlike gold, which is easily reduced to metallic form, uranium is highly reactive. It tends to occur as oxides, silicates, or salts. Two ores are of predominant importance as commercial sources of this metal: pitchblende which contains uranium as the oxide, U3O51 and carnotite in which the uranium is present as a complex salt with vanadium, K2O-2UCV3V2O5-3H2O. These ores may vary widely in respect to the nature of their gangue constituents. Some are largely siliceous in composition, while others consist mainly of calcite. Sometimes substantial amounts of pyrite or of organic materials are present and these may lead to specific problems in treating the ore. Further complications may be introduced by the presence of other metal values such as gold, copper, cobalt, or vanadium whose re- covery has to be considered along with that of the uranium, or whose separation from uranium presents particular difficulty. In general, there are two main processes for recovering uranium in common use today.'.2 One of these employs an acid solution such as dilute sulphuric acid to extract the uranium from the ore. A suitable oxidizing agent such as MnO, or NaNO, is sometimes added if the uranium in the ore is in a partially reduced state. The uranium dissolves as a uranyl sulphate salt and can be precipitated subsequently by neutralization or other suitable treatment of the solution. The second process employs an alkaline leaching solution, usually containing sodium carbonate. The uranium, which must be in the hexavalent state, is dissolved as a complex uranyl tricarbonate salt, and then is precipitated either by neutralizing the solution with acid or by adding an excess of sodium hydroxide. The latter method has the advantage of permitting the solutions to be recycled, since the carbonate is not destroyed. This is essential if the process is to be economical, particularly with low grade ores. With each of these processes, there are associated a number of advantages and disadvantages and the choice between using acid or carbonate leaching is generally determined by the nature of the ore to be treated. In the past, more ores appear to have been amenable to acid leaching than to carbonate leaching and the former process correspondingly has found wider application. With most ores, acid leaching has been found to operate fairly efficiently and to yield high recoveries. One of the main disadvantages has been that large amounts of impurities, such as iron and aluminum, sometimes are taken into solution along with the uranium. This may give rise to a high reagent consumption and to difficulties in separating a pure uranium product. Excessive reagent consumption in the acid leach process also may result

Jan 1, 1955

By G. Derge, R. E. Grace

Diffusivity of bismuth in liquid Pb-Bi alloys has been measured by the capillary reservoir method as a function of temperature and composition. Fair agreement between theory and experiment is found for the measured diffusion coefficients and activation energies of diffusivity and viscosity in dilute lead alloys. The measured diffusion coefficients in the high bismuth alloys describe mass transport in special concentration gradients. THE study of diffusion in liquid metals has been restricted largely to the amalgams over small ranges of temperature and composition. In contrast to this, diffusion and transport of ions and molecules in aqueous and organic solvents have been studied extensively for the past 50 years. The general difficulty of measuring diffusion coefficients in both gases and liquids at high temperatures has been that of eliminating convection currents. Measuring techniques based on freezing and sectioning after the diffusion period1 are not particularly satisfactory because of the volume changes and segregation which occur during freezing. The most promising technique which is currently being developed for liquid metals is the capillary reservoir method. This eliminates convection currents from the diffusion zone and requires measurements only of the length and initial and final compositions of the alloy within the capillary. Anderson and Saddington' have measured the diffusion coefficient of tungstate anions in aqueous Na,W04; Wane4' and Hoffman- ave studied structure and self-diffusion in water and self-diffusion in mercury, respectively, by similar techniques. Selection of the Pb-Bi system for initial study was made for several reasons: I—It afforded a low temperature system for development of techniques which would be suitable for high temperature measurements in liquid iron. This research on diffusion in liquid iron is now in progress as a part of a general program on the rates and mechanisms of slag-metal reactions. 2—Activity data on liquid Pb-Bi alloys of all compositions are available from the emf meas- urements of Strickler and Seltz.9 3—Some data are available on the diffusion of bismuth into solid lead. The measurements on solid alloys were made by von Hevesy and Seith7 for the concentration range of 0 to 2 pct Bi. An appropriate solution to Fick's law (unsteady state) for the capillary reservoir has been found by considering the mass flow which passes through a unit area of surface on a slab of infinite thickness as a function of time. This is the first application of this solution to diffusion measurements, although a similar solution is available for identical boundary conditions.' By analogy with heat conduction, the problem reduces to a perfectly insulated infinite slab at temperature C,, one surface of which is quenched to C, for all times, t. Heat is permitted to flow from quenched surface of the slab to an infinite supply of coolant at C.' (see Appendix). For the corresponding solution for diffusion in liquids, these terms become Where Cu is the initial concentration of component A" in entire capillary at t - 0, C, is the final con- In this case. component A is bismuth, and C, = 0. centration of component A in entire capillary at t = t, C, is the concentration of component A in bath for all t, 1 is the length of the capillary in centimeters, t is the time of diffusion period in seconds, and D is the diffusion coefficient in square centimeters per second. By comparison with solutions to Fick's law used in the solid state, this solution for liquids is similar to the Grube solution for solids, where the change in D with chemical composition is not taken into consideration. Difficulty is to be expected if large concentration gradients are employed and if the concentration gradients extend over an appreciable length of the capillary.

Jan 1, 1956

By R. E. Hudrlik, G. W. Preckshot

The diffusion coefficients of silver from solid silver in molten lead were measured to within ± 0.8 pet in a columnar type diffusion cell ower, the temperature range of 326° to 530°C. Fick's law describes the process up to 530°C where the laminar mechanism appareltly breaks down. These is negligible resistance at the interface as shown by mathematical analyses. The diffusion coefficients are found concentration independent. IT would seem that diffusion in liquid metals would be free of such effects as molecular structure, dissociation. polarization. and compound formation. This view was taken by Gorman and preckshot in their study of diffusion of copper from solid copper into molten lead. They reported diffusion coefficients which were independent of the concentration over the range of 478° to 750°C. They found that the Stokes-Einstein equation with constant radius of the diffusing specie represented the diffusion data better than Eyring's rate theory equation and Sheibel's correlation. The radius of diffusion was found to be that of the doubly charged copper. There appeared to be no resistance across the solid-liquid boundary. In the present work the diffusion coefficients for silver in liquid lead were measured over a range of temperatures of 350° to 505°C. The solubility of silver in lead over the range of 303° to 630°C was also obtained. These results are compared with calculated or correlated values or with data in the literature. EXPERIMENTAL Procedure—The experimental equipment techniques and procedures were those reported in detail by Gorman and preckshot9 and will not be repeated here. Measured values of WT, Co, A. L were obtained for various diffusion times and the diffusion coefficient was computed for the case of no resistance at the interface9, 11 by: WT/CoAL = 1- 8/p2 n=1 1/(2n - 1)2 exp[-(2n - 1)2p2 Dt/4L2] [1] or where there was resistance at the interface by: WT = 1- ?n=1 2h2/ap2L [sxp [-Dan2t]/[(h2 + an2) L + h] The roots an are those of the transcendental equation3 tan (an L) = Iz/cun. The diffusion coefficient is that defined by Hartley and Crank.7 The total silver in the lead cylinder and equilibrium slug was determined by a cupellation technique' with proper correction for losses. Analysis of known samples showed that this method is surprisingly accurate. The amount of silver in the lead adhering to the silver cylinder was obtained in the same fashion as shown by Gorman and preckshot.9 The small errors involved in this determination are not critical since the silver in this adhering lead layer is only 3 to 15 pet of the total diffused. Materials—Electrolytic silver containing 99.9+ pet Ag obtained from General Refineries of Minneapolis, Minn. was used for all but runs 7 and 8. For the balance of the runs this silver was reduced with hydrogen at 1100°C and its oxygen content was found to be about 0.017 pet. For the runs. 7 and 8, phosphorous-reduced silver of the same purity was obtained from Handy and Harman Co. of Chicago, Ill. The densities of the phosphorus-reduced silver and the hydrogen-reduced electrolytic silver were 10.484 and 10.487 g per cm3, respectively. These values agree with those reported for pure silver. Silver which was reduced at 900 C had an average density of 9.998 g per cm3, indicating porosity. This silver was used for a number of runs which were not tabulated in Table I. These are indicated by crosses on Fig. 2. The 99.999 pet Pb was obtained from the National Lead Co. Research Laboratory of Brooklyn, New York. DISCUSSION OF RESULTS The diffusion and solubility results are reported in Table I for eleven runs using either phosphorus-reduced electrolytic silver or hydrogen-reduced silver at 1100° C. The solubility data shown in Fig. 1 show the excellent agreement with that reported by Heycock and Neville.8 The data of Friedrichs5 apparently are in error. The experimental solubility data of this work are reported to 0.3 pet. The experimental diffusion coefficients computed from Eq. [1] are reported within 1.2 pet of the mean and are plotted in Fig. 2. These are expressed within +0.8 pet of the experimental values over the entire temperature range by: D= 8.26 x 10 -5 e-1925/RT . [3] There appears to be little difference due to the

Jan 1, 1961

By M. T. Simnad, G. Derge, Ling Yang

A method is described for using radioactive tracer for the direct determination of ionic mobility in a molten electrolyte. The ionic mobility of Fe" in silica-saturated iron silicate is found to be 0.0009 (*0.0002) cm2 sec ' volt L at 1250°C. eLECTROCHEMICAL studies', ' of molten iron eE oxide and iron silicates have shown some ionic conductivity at all compositions. The current efficiency in electrolysis increases with the concentration of silica to a maximum of 90 pct at about 34 pet SiO,. In the silica-saturated melts the transport number of ferrous ions is near unity, indicating that

Jan 1, 1957

By W. J. Kroll

W. J. Kroll (Belgium)—Those who are acquainted with the metallurgy of our forefathers may have occasionally a good time when glancing at recent publications in which aged discoveries are unconsciously reported as novelties. We know very well, that oxygen blowing of iron has been described by old Bessemer. The author of the present article is, however, not aware of the fact that the process he describes, the fused salts scrubbing, was already used in the middle of the last century by St. Claire Deville in his attempts at making higher-purity aluminum from raw AlCl3 by sodium reduction. The same fact escaped the three investigators to whom the author refers. The technical grade AlC13 produced by St. Claire Deville from impure oxide, chlorine, and carbon (Oersted process) was contaminated with Sic4 and FeC13, which he removed on an industrial scale by various methods.' He sublimed, for instance, his chloride in contact with iron nails, or aluminum chips, or in hydrogen, which reduced the FeC13 to the less volatile FeCl2. He used also alkali-alkaline earth chlorides as a solvent for A1C13 and FeC13 in presence of iron as a reducing agent, and sublimed his purified AlC13 from solvent salts. The successors of Deville used also zinc or lead as a reducing agent. These methods were later used currently for the purification of BeCl2 and for collecting this compound in a salt bath, suitable for fusion electrolysis. More recently, the same Deville process has been described as applied to the purification of ZrCL4, as part of a method of extraction of ZrCL4 from silico zirconium, which later was reacted with FeC12 in a separate part of the equipment. According to the patent referred to3 the ZrCL4 is dissolved as a gas in a carrier salt such as alkali-alkaline earth chloride, from which it can be reclaimed in a purified form by distillation. A bath of NaCl/KCl is recommended in the specifications. The FeC12-FeC13 go in the salt solution in which raw ZrCl4 is introduced by bubbling, while Sic4 and Tic4 pass through the bath and escape in the atmosphere. purification by this salt-solvent method is thus well described. It may be worth while insisting on some minor details about which the author did not bother too much. One of these concerns the material of construction for the container. If iron is used, and if no other reducing agent is added, then it is the iron of the pot that acts as a reagent in the breakdown of FeC13 to FeCl2, and corrosion can be expected. This shows the importance of using extraneous large surface-reducing agents, of which there is a choice. Zir-

Jan 1, 1963

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 F. F. Poland

The purpose of this paper is to describe the recent applications and improvements made in the process and equipment for the recovery of metallic zinc from secondary metals by means of high temperature electric resistor furnace distillation. The process and high temperature furnace, as developed by Revere Copper and Brass Incorporated, Research and Development Department, was described in a previous paper. † Since the publication of that paper the process and equipment have been licensed to others. Consequently its field of application has been expanded, particularly for the treatment of scrap brass, galvanizers dross and scrap zinc. Other applications are being considered, particularly the melting of metallic titanium and special alloys requiring a controlled atmosphere. Consideration is also being given to modification in design that would allow the melting of copper cathodes for the production of oxygen-free and electrolytic tough pitch copper shapes without the necessity of blowing and poling. In the design of a furnace for melting copper cathodes it is intended that the fundamental principle of charging the solid material into a large pool of molten metal through a double door charging vestibule as shown in Fig 3 of the previous paper, † would be adhered to and, in addition, either dehydrated and desulphurized producer gas or commercial nitrogen, that is, approximately 96 pct nitrogen, 2 pct hydrogen 2 pct CO, also dehydrated, would be used to provide a protective atmosphere for the resistors and to flush out more soluble gases from the molten copper. The operation of a small unit producing approximately 1000 lb of copper per hr has shown that the commercial nitrogen will allow the production of tough pitch electrolytic copper when the metal is transferred from the meltdown unit contemplated to a special type induction unit where the oxygen absorption and temperature can be controlled. Standard practice in treating low grade scrap brass has been to oxidize the zinc by blowing air through the molten metal. This operation results in the recovery of approximately two-thirds or less of the zinc content of the brass in the form of an impure zinc oxide having only a fraction of the value of the zinc in metallic form. The balance of the zinc is lost to slag, fume and drosses. Fig 1 shows the revised flow sheet of the process for the treatment of scrap brass. For this application and as shown on the flow sheet a low frequency induction furnace has been substituted for the Wilkins-Poland type as a melting unit. This type of furnace has been found the most suitable for the purpose as will be shown by the following discussion. Serap Brass The melting of refinery scrap brass preparatory to distillation presents a melting problem quite different from the usual melting of brass mill scrap for recasting because refinery scrap is a heterogeneous mixture of alloys each having different melting and boiling points, and coated with a variety of corrosion products and foreign materials. In addition, brass and copper plated pieces of sheet iron, stainless steel, ceramic materials, and others are present that do not melt except at temperatures above the boiling point of the liquid bath. In the previous paper it is noted that difficulty was encountered in the melting of refinery grade scrap brass without excessive loss of zinc by volatilization and that 5 to 10 pct razorite flux was used to alleviate this condition. While this flux was successful in preventing the excessive loss of zinc by volatilization it interposed a low heat conductivity layer between the source of heat and liquid charge and resulted in lowering the melting rate of the furnace. This same problem presents itself in all types of furnaces in which the metal is heated by radiation and conduction from above the charge. It has been found in melting this type material that the melting rate per hour or per square foot of hearth area

Jan 1, 1950

By F. F. Poland

S. ROLLE*—As one who has been privileged to read Mr. Poland's paper before its publication, I wish to express high praise for it and the record which it gives of the work done by the Research and Development Department of Revere Copper and Brass Inc. in developing a satisfactory furnace for the recovery of metallic zinc from secondary metals by distillation. I have also had the opportunity of studying the operation of the furnace when running on scrap brass at the Detroit plant of Revere, and I can therefore say from knowledge gained first-hand that the unit is efficient and economical when used in conjunction with distillation. However, when I read what Mr. Poland has to say in reference to the possibility of using the furnace for melting copper cathodes for the production of oxygen-free and tough-pitch copper shapes, I am constrained to be somewhat skeptical. No doubt the furnace can readily be modified so as to make it possible to melt cathodes, but that is only a part of what is required to produce refinery shapes. Because of its construction the atmosphere in the furnace is highly reducing and even without protective gases such as nitrogen the copper cathodes which are

Jan 1, 1950

By Douglas J. Harvey

The surface tension of lead-tellurium alloys (in the range 0 to 6.70 at. pct Te) ad lead-arsenic alloys (in the range 0 to 10.53 at. pct As) has been examined by the maximum bubble pressure method. The surface tension of high purity lead can be expressed over a range of temperatures (M.P. to 750°C) by the relationship = 512 — 0.133t. Arsenic and lellurium were both found to be surface active in lead. The surface excess concentration at 700°C was .found to be constant at 0.74 x 10-10 moles per sq cm for tellurium. The surface excess of arsenic at 700°C varied between 0.26 x 10-10 moles per sq cm at 0.16 at. pct to 1.5 x 10-10 moles per sq cm at 8.3 at. pet. The surface tension of lead has been investigated sporadically over the last 90 years. The first recorded study was made by Quincke1 who used a rather crude application of the drop-weight method. Hogness2 obtained the first reliable value for the surface tension of lead by measuring the pressure necessary to force the liquid metal from the end of a small capillary of known radius. Hogness gives the empirical formula: ?(surface tension)= 444 - 0.077 (l-327) which describes the relationship between surface tension and temperature ("C); This work was followed by that of Bircumshaw3 who applied the two-capillar maximum bubble pressure method of Sugden.4 The values published by Bircumshaw are in good agreement with those found by Hogness. Matuyama5 examined the surface tension of lead by the drop weight method. However, the values found are somewhat higher than those previously reported. More recent investigations by Greenaway6 and by Melford and oar' have resulted in data which are not in close agreement, but which are well within the expected range at most temperature levels. The study of the surface tension of lead alloys began with Drath and sauerwald8 who examined the lead-bismuth system. They found that the surface tension values of the alloys in this system deviated only slightly from the ideal mixture rule. Matuyarna5 examined the lead-antimony system. Bircumshaw the lead-tin, and the lead-indium system was studied by Hoar and Me1ford.10 The results of these investi- gations are summarized in Fig. 1. It may be seen that the data vary only slightly from the ideal mixture rule. As none of the metals studied and reported in the literature has a very interesting effect on the surface tension of lead,*the next logical step might be to examine the effect of metals of low solubility or nonmetals having some liquid solubility. Baes, and Kellogg11 have pointed out that the elements most likely to be surface active in liquid metals are ones of limited solubility in the liquid state and possessing weak intermolecular bonding forces in the solid state. Therefore, Group V, VI and VII nonmetals are most likely to be surface active. It was decided to examine the effect of a borderline metal, arsenic, and a nonmetal tellurium. Arsenic forms a eutectic with lead having a melting point of 290°C and a composition of 6.9 at. pct As.13 Pure arsenic sublimes when heated above 610°C. Even though alloying with lead lowers the vapor pressure, fumes containing arsenic are evolved at a very noticeable rate when lead-arsenic alloys are heated above 650°C. Tellurium raises the liquidus temperature of lead sharply to 904°C,14 the melting point of PbTe (lead-telluride). The solid solubility of tellurium in lead was determined by Greenwood and worner15 to be 0.0007 at. pct at room temperature. EXPERIMENTAL METHOD Surface Tension Measurement—The two capillary maximum bubble pressure method for measuring surface tension developed by Sugden,4 and used later

Jan 1, 1962

By C. A. Winkler, V. Hospadaruk

PREVIOUS papers from this laboratory have discussed the effect of chloride ion on the cathode polarization during electrodeposition of copper from copper sulphate-sulphuric acid electrolytes, in the presence and absence of gelatin. The steady state polarization'" was found to decrease sharply and pass through a minimum with increasing chloride ion concentration in the presence of gelatin. The minimum shifted to higher chloride ion concentrations and to higher polarization values with increase in current density or gelatin concentration, while an increase of temperature shifted the minimum toward lower halide concentrations and lower polarizations. Since these observations were made in acid-copper sulphate electrolytes that contained no other addend than gelatin, there was obviously the possibility that they were not applicable to deposition of copper from commercial electrolytes that contain a variety of other substances in relatively small amounts. In particular, it was of interest to determine whether the presence of arsenic, antimony, or bismuth in the electrolyte would materially alter the behavior. Experiments have now been made under a variety of conditions with systems containing these cations, and the results are summarized in the present paper. Experimental Polarization measurements were made at 24.5oC in a Haring cell in the manner described previously.' Electrolytes were made with doubly-distilled water, and contained 125 g per liter of copper sulphate and 100 g per liter sulphuric acid, both of reagent grade Eimer and Amend gelatin from a single stock was used throughout. Chloride ion was introduced as reagent grade sodium chloride, and arsenic, antimony, and bismuth by dissolving the chemically pure metal in hot concentrated sulphuric acid and adding appropriate amounts of the solutions to the electrolyte. Each cathode, of 1/16-in. thick rolled copper, was first etched in 40 pct nitric acid and washed thoroughly with distilled water. The surface was then brought to a standard condition4~9 by electrodeposition from an acid-copper sulphate electrolyte containing no gelatin, at a current density of 3 amp per sq dm for 30 min, followed by deposition at a current density of 2 amp per sq dm for l hr. As in previous studies, the cathode polarization eventually attained a steady-state value (15 to 75 min) such that further change in polarization did not exceed 0.2 mv per min. The polarization values recorded are those for the steady states. "Excess weights" were determined with arsenic and antimony present in the electrolyte, as the difference between the weights of the deposits obtained in the presence of these cations and those obtained in their absence, with the two cells connected in series. When gelatin was present along with the arsenic or antimony, it was also added to the electrolyte in the cell in series. Results and Discussion The results of the study are summarized in Figs. 1 to 6. From Fig. 1, top, it is evident that the presence of arsenic or antimony alone results in an increase of polarization, while bismuth alone causes a decrease. The presence of gelatin (25 mg per liter) rather drastically modifies all three cation effects, as indicated in the lower panels of the same figure. The addition of chloride ion, when no gelatin is present, causes comparable decreases in polarization in the presence of antimony and bismuth, but a relatively larger decrease when the electrolyte contains arsenic. It is interesting to note that the decrease in polarization brought about by addition of chloride when both arsenic and antimony are present parallels the behavior with arsenic alone, while the polarization in the electrolyte containing the cation mixture, without chloride added, corresponds to that for an electrolyte containing only the antimony cation. Similarly, the polarization at zero concentration of chloride in electrolyte containing arsenic and bismuth is that corresponding to an electrolyte containing arsenic alone. From Figs. 3a, 4a and 4b, it is clear that, in the presence of gelatin at a level of 25 mg per liter, the effect of chloride in the presence of arsenic and antimony, or a mixture of the two, becomes quite analogous to that observed in the absence of added cations. When both bismuth and gelatin are present (Fig. 5), the decrease in polarization with increased chloride concentration is virtually absent. This is perhaps a reflection of the large decrease in polarization brought about by the bismuth itself in the presence of gelatin. The shifts of the minimum in the polarization-chloride concentration curves brought about by changes of temperature (Fig. 3b), gelatin concentration (Figs. 3c and 4c) and current density (Fig. 3d) when the metal cations were present are all similar to the corresponding shifts observed in their absence." The approximately linear "excess weightv-anti-mony concentration relation recorded in Fig. 6 would seem to indicate that antimony is codeposited with copper to a considerable extent. On the other hand, only very limited amounts of arsenic appear to be adsorbed or codeposited.

Jan 1, 1954

By C. F. Baes, H. H. Kellogg

The effect of dissolved sulphur on the surface tension of liquid copper has been measured by a sessile-drop method at 1120°C. Sulphur is highly "surface active" in liquid copper; the surface tension lowering is about 650 dynes per cm for a sulphur content of 0.6 pct. The degree of adsorption of sulphur on the liquid copper surface has been calculated, and various models of the surface layer have been tested against the results. THE science and technology of surface chemistry and physics at room temperature are relatively advanced, and the engineer possesses a background of information that allows him to predict and control many phenomena—such as frothing, emulsifica-tion, wetting of solids by liquids, flow through capillaries, adsorption, mineral flotation, penetration of porous bodies by liquids—which are related to the behavior at interfaces. Surface phenomena in high temperature systems are probably of equal importance to the metallurgist and ceramist, but at present are incompletely understood. Frothing of liquid slags, penetration of refractories by liquid metals and slags, infiltration of powder metallurgy "skeletons" by liquid metals, bonding of metals to ceramic oxides in the manufacture of "cermets," wetting of one metal by another for purposes of soldering and welding, are but a few of the many phenomena of metallurgy that are controlled by surface behavior. Concerning the most elementary of surface properties—the gas-liquid surface tension—the existing data on liquid metals are inadequate. For the surface tension of "pure" metals the results of different investigators differ by 10 or 20 pct. The role of surface-active agents in liquid metals is relatively unknown, and it is not unlikely that the discrepancy in the values for the surface tension of pure metals is due to the presence of traces of surface-tension-lowering impurities in the samples. Based on analogy to room-temperature systems, the highly surface-active impurities in liquid metals should be elements or compounds of limited solubility in the liquid metal and possessing considerably weaker in-termolecular bonding forces than the metal itself. The nonmetals of the V, VI, and VII groups of the periodic table might be expected to act in this way. The experimental work in this paper represents a study of the surface tension of pure liquid copper in hydrogen at 1120°C, and the effect of dissolved sulphur (which is a strongly surface-active agent in liquid copper) on the surface tension of copper. The Gibbs equation has been applied to the results and an attempt made to correlate the surface tension behavior with two models of the interfacial layer. The sessile-drop method for surface-tension measurement at high temperatures which was used in this investigation has been described by several investigators.1,2 It involves the measurements of the dimensions of a stationary liquid drop on a horizontal surface. From these dimensions and the liquid density, the surface tension of the liquid may be calculated either by means of various approximate formulas, or more accurately by means of the Bash-forth and Adams tables.- or liquid drops which exhibit a contact angle greater than 90°, by far the simplest dimensions to measure are those indicated in Fig. 1, i.e., the diameter of the maximum horizontal cross section (2X) and the maximum height (Z) above this plane. In order to calculate the contact angle and the volume of the drop, the diameter at the base of the drop (2a) and the total height (h) also must be measured. Upon investigation of the literature, the authors failed to discover a concise statement of the accuracy of the method when used in conjunction with the Bashforth and Adams tables, hence a careful examination of the tables was made and the curve shown in Fig. 2 was constructed. The "error factor," plotted as the ordinate, is the factor by which the

Jan 1, 1954

By Walter R. Hibbard, Harold Margolin

The effect of ten ternary additions on the aging of a Cu-Ag alloy was measured by X ray, microstructure and hardness. A supersaturated copper-rich transition structure was found. The effect of the ternary element was related to differences between transition and precipitate lattice parameters and analogies with solution and strain hardening. IN addition to the prerequisite of diminishing solid solubility with falling temperature, the theory of coherency hardening1 proposes that the principal source of hardening during precipitation is strain caused by coherency between the matrix and precipitate. For a given system greater coherency hardening would result from lengthening the period of coherent growth and/or increasing the number of nuclei, other things remaining equal. However, the hardness change during aging is a function of a number of phenomena in addition to coherency hardening, such as dispersion hardening, loss of coherency, solute depletion, matrix recovery or recrystal-lization, coalescence, and their relative time sequences. The present investigation was undertaken to study the effect of ternary additions on the age-hardening of Cu-Ag alloy. For this purpose a series of alloys were made containing as ternary constituents Mg, Al, Si, P, Mn, Ni, Zn, As, In, and Sb. These elements were added to a master 5 wt pct Ag in Cu alloy in such amounts as to obtain alloys listed in Table II containing the following approximate atomic percentages: 96 pct Cu, 3 pct Ag, 1 pct ternary addition. A 97 atomic pct Cu, 3 atomic pct Ag alloy was used as a comparison standard. There have been numerous investigations of age-hardening in Cu-Ag alloys2-lo but few of ternary Cu-Ag alloys. Pfister and Weist" found that 1 wt pct Ni reduced the solubility of Ag in Cu above 600 °C and that the rate of precipitation was slower than in polycrystalline Cu-Ag alloys of similar silver content. More recently Hodge, Jaffee, Dunleavy, and Ogden12 investigated the effect of ternary additions of Mg, Li, P, Cd, Zn, Zr and Be on strength properties of Cu-Ag alloys. Mg, Cd, and P were found to increase strength in the aged and cold-worked condition. Experimental Procedure Charges weighing 7.5 lb were melted in a closed carbon crucible under a reducing atmosphere* of nitrogen and hydrogen. A Durville-type casting procedure with a hot-top arrangement was used to obtain a sound casting, 2x2x43/4 in. Prior to rolling and annealing of the ingots, a trial solution anneal of rolled sections of the hot-top portions indicated that at 780°C melting was avoided except for the Cu-Ag-Sb alloy. This alloy could not be produced as a single phase within the limits of temperature control. i 3" to 5oC, since melting occurred at 64O °C, and below this temperature precipitation was

Jan 1, 1952

By B. B. Kunkle

The history of this refinery has been one of expansion. It now has the capacity to produce 240,000 tons of cathodes annually. A general flowsheet and descriptions of the plant layout, operations involved, and equipment used at one of the world's largest and most modern electrolytic copper refineries are presented. THE El Paso copper refinery of Phelps Dodge Refining Corp. was originally constructed by the Nichols Copper Co., in association with Phelps Dodge Corp. and the Calumet & Arizona Mining Co., in order to refine anode copper produced at the smelters of the two latter companies. Initial production was obtained from the plant in January 1930 and its annual refining capacity was 100,000 tons of anodes. In order to keep pace with Phelps Dodge Corp.'s increasing copper production in the Southwest, it became necessary in 1937 to increase the El Paso plant's refining capacity to 160,000 tons of cathodes annually. This was accomplished by adding two sections to the electrolytic tank house, increasing the number of anodes placed in each commercial electrolytic tank from 33 to 37, and making changes in the dc circuits which made it possible to put three generators on each of the two circuits instead of two, thus increasing the amperage on each circuit from 10,000 to 14,800 amp. The capacity of the two furnaces from which refined shapes were cast was increased by raising the sidedoor bays on the furnaces 6 in. An increase of 70 pct in the productive capacity of Phelps Dodge Corp.s Morenci Branch in 1942 again made it necessary to expand refining capacity. In order to accomplish this, Defense Plant Corp. set up a project to increase the capacity of the El Paso plant to 240,000 tons of cathodes annually. This project consisted of enlarging the tank house approximately 50 pct, the construction of another wire- bar furnace, and corresponding increases in other departments where necessary. The engineering work on this expansion was begun in March 1942. Actual construction was under way by August 1942, and the work was completed by March 1944. These additional facilities were purchased from Defense Plant Corp. by Phelps Dodge Refining Corp. in August 1949. The plant is in El Paso about eight miies east of the center of the city and near Highway 80 and the main lines of the Southern Pacific and Texas & New Orleans railroads. The plant site covers approximately 494 acres of company-owned land that is ideally adapted for its use, being well-drained sandy soil and having enough slope so that a minimum amount of excavation was required when the plant was constructed. There is still ample room on the plant site for any future plant expansion. The plant is constructed on four different ground levels and so arranged that the unrefined copper is unloaded on the highest level at the north side of the plant and the refined copper is shipped out from the lowest level on the south side of the plant. The receiving or unloading platform, weighing and sampling rooms, main floor of the tank house, acid and condensate storage tanks, and the slimes and copper sulphate plants are on the first or highest level. The tank house basement, yard scales, charging and furnace bays, balloon flue and stack, brick-storage building, substation and power house, boiler feedwater treating plant and pump house, water storage tanks, and change houses are on the second level. This second level also contains ample yard area for setting-up furnace charges, and for storage of custom scrap, cathodes, scrap anodes, blister, coke, and poles. The casting wheels, inspection conveyors, cathode shears, wire-bar scalping machine, slag

Jan 1, 1952

By T. J. Woodside

Prior to 1927 the lead blast furnace charge at El Paso consisted principally of direct-smelting carbonate ores, very low in zinc, and the resulting slag seldom carried more than 4.0 pct. With the exhaustion of the mines in Northern Mexico which supplied these ores, they were supplanted with a much smaller tonnage of sulphide concentrates containing appreciable amounts of zinc, and the zinc content of the slag rose to 10.0 pct. The Anaconda Company's slag fuming plant at East Helena, Montana, had then been in successful operation for several years, and it was immediately recognized that, while at that time the production of high zinc slag was small, if the grade held, a plant for the recovery of zinc from the accumulated slag might some day be practical. The grade not only held but became increasingly higher so that by 1946 suffcient slag had been accumulated to warrant the construction of a fuming plant regardless of future slag production from the blast furnaces, and with the removal of Government restrictions on construction at the end of the war the project was undertaken. The general design of the plant follows that at Bunker Hill, with waste heat boiler, unit coal pulverizers and even more extensive instrumentation and automatic control. The area around the lead blast furnaces is rather congested, being cut off from the nearest feasible site for the fuming plant by the copper plant, so that the most practical route for hot slag transfer is through the converter aisle. For this reason and also to make use of the existing converter cranes to serve the fuming furnace, an unused reverberatory at the north end of the converter building was dismantled and the fuming furnace and boiler located in its place. A flue at right angles to the furnace and boiler runs northward to the U-tube coolers and baghouse (Fig 1). From the bag-house a steel flue runs eastward to an existing and long unused 300 ft con- Crete stack. The coal handling system and de-leading plant lie north of the furnace and east of the cooling tubes and baghouse. The furnace is conventional, 8 X 21 X 33 ft high inside and completely water jacketed (Fig 2). There are 21

Jan 1, 1950

By T. J. Woodside

A. E. LEE*—One question of the flue leading to the tube coolers. What gas velocity is maintained to prevent settling of the finely divided fume ? T. J. WOODSIDE (author's reply)— 30 fps. Of course some fume settles out, but is immediately carried away by the screw conveyors. H. R. BIANCO —How effective are the soot blowers ? Do you have to use any hand lancing? T. J. WOODSIDE Only one point, which is right at the top of the economizer and which the soot blowers do not reach, we occasionally hand lance. But everywhere else the soot blowers have been very effective and no hand lancing has been done. L. V. DAVIDSON*—Do you have considerable variation in the lead content of the fume collected if the lead is higher, and is there greater tendency for the fume to stick on the boiler tubes ? T. J. WOODSIDE—There is very little variation in the lead content. From 5 to 7 pct is the widest range variation we have had and we have noticed no difference in the behavior of the fume. C. R. KUZELL—Mr. Woodside has been quite generous with data pertaining to metal and fuel economics. I think he stated that the coal consumption of the furnace was 21/10 21/10 for a pound of zinc recovered. I take it that that is the gross coal input to the furnace. I wonder what the net fuel consumption in terms of coal might be after crediting the fuel equivalent of the waste heat steam recovered? T. J. WOODSIDE—The fuel value of the waste heat steam is approximately 42 pct of the heat input to the furnace. A very large part of the heat, of course, goes out in jacket water—44 pct. A. D. TERRELL*—I would like to know how much lead is in the deleaded material. T. J. WOODSIDE—It is a little less than 2 pct now.

Jan 1, 1950

By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4'/2x2'/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By G. Osuch, G. Derge, G. M. Pound

Using a new dternating-current potentiometer circuit and a specially designed four-terminal cell, the specific conductance of molten Cu2S-FeS mattes was measured as a function of temperature, from the liquidus to 1500°C, over the complete range of composition. The high conductivities, about 1500 ohm-I cm-l for FeS and 100 ohm-l cm-l for Cu,S, indicate that the conduction is electronic rather than ionic. Molten FeS has a negative temperature coefficient of specific conductance, resembling metallic conduction. Molten Cu,S has a positive temperature coefficient, resembling semiconduction. The binary roughly follows an additive rule of mixtures with respect to both magnitude and temperature coefficient of specific conductance. Metallic bonding in the liquid is postulated to explain these phenomena. MUCH has been learned in the past about the nature of liquids and the ionic or molecular species in solution by means of electrical measurements. Thus, dielectric constants','2 have given information about molecular liquids such as water and benzene. Measurements of dielectric constant usually are impossible in electrically conducting liquids, such as aqueous solutions of ionic salts and molten ionic salts. However, measurements of electrical conductance and ionic transference have provided much knowledge about the latter systems.a-" In recent years, the ionic nature of certain molten metallurgical slags has been established by Derge and Martin7 through electrical conductance and electrolysis measurements. Chipman, Inouye, and Tom-linsonq ave studied the electrical conductance of molten FeO and report a high specific conductance of about 200 ohm-' cm-' (compared with 4 ohm" cm-' for an ordinary ionic liquid such as molten NaCl) and a positive temperature coefficient of conductance. They interpret these results in terms of p-type semiconduction by analogy to the situation in solid FeO.Y imnad and Derge" have studied cell efficiency in the electrolysis of molten FeO-SiO, systems and conclude that ordinary ionic conductance increases with SiO, content. Very recently, interest has been revived in the electrical conductance of liquid metals and liquid metallic solutions. Scala and Robertson1' report a close resemblance between the liquid and solid states with respect to thermal, structural, and compositional relationships. Molten sulphides have not received a great deal of attention. Bornemann and von Rauschenplat" measured the specific conductance of molten Cu2S as a function of temperature with a four-terminal cell using direct current. A high specific conductance and a positive temperature coefficient were found in that investigation." Using a two-electrode apparatus, Savelsberg" electrolyzed various molten sulphide mixtures. He concluded that pure molten Cu,S and FeS were electronic conductors but that the mixtures exhibited some ionic conduction. In the present investigation, the specific conductance of the industrially important Cu-Fe sulphide mattes was measured as a function of temperature and composition in order to investigate the mode of electrical conduction and the structure of these molten mattes. An alternating-current circuit was used to eliminate the effect of any possible electrode reactions. Apparatus The Conductance Cell: Due to the high specific conductance of the systems studied (10' to 10" ohm-' cm-'), the classical two-terminal cell and Wheat-stone bridge apparatus could not be used. A four-terminal cell was developed in order to eliminate lead resistance, and an ac potentiometer circuit was designed to give rapid and sufficiently accurate measurements of the cell resistance. A diagram of the conductivity cell is given in Fig. 1. The molten matte is contained in a dense alundum crucible, and spectrographic graphite rods that dip into the molten matte serve as the four conductance terminals. Two of the graphite rods on opposite sides of the cell serve as current-carrying leads, and the other two graphite rods are null-current probes that detect the potential drop across the cell. These graphite rods are contained in silica tubes, and the lower constricted portions of the two silica tubes define the column of liquid whose electrical resistance is being measured. The electrical resistance of the broad ex-

Jan 1, 1956

By L. J. Howell, H. H. Kellogg

AN a previous paper' the phase diagram and vapor pressure of melts in the systems NaC1-ZrCl,, KC1-ZrCl,, and NaC1-KC1 (1:1 molar)-ZrC1, were reported. This note supplements the earlier paper with the results of electrical conductivity measurements on NaC1-ZrC1,and NaC1-KC1 (1:l molar)-ZrC1, melts. The design of the conductivity cell is shown in Fig. 1. Tungsten electrodes, A, were sealed into the Pyrex cell, C. That portion of the electrodes in contact with the melt was covered with a tight-fitting platinum cap. The electrodes were sur-

Jan 1, 1960