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By Emma Smith, R. W. Guard

INVESTIGATION of the nickel-rich end of the Ni-Zr system has been prompted by an interest in the effect of small amounts of zirconium in high temperature alloys. Hansenl presents a hypothetical diagram based on the work of Allibone and Sykes,2 who investigated the system up to 44 atomic pct (55 wt pct) Zr by metallographic means. Since the alloys they used were melted in Alundum crucibles they were probably contaminated with appreciable amounts of aluminum. They found two compounds which they called Ni3Zr and Ni4Zr. Hayes, Roberson, and Paasche3 identified a compound NiZr (50 atomic pct Zr) with a melting point of 1470°C. Such a high melting compound is not compatible with the observation of Allibone and Sykes2 hat a eutectic is formed near 50 atomic pct. The alloys containing up to 60 atomic pct Zr were melted in an inert atmosphere in a noncon-sumable electrode arc furnace. Analysis of most of the alloys indicated negligible loss of zirconium. The alloys were examined in the as-cast condition and also after annealing for 6 hr at 1175°C (or 1000°C for the alloys between 25 atomic pct and 45 atomic pct). Metallographic specimens were etched using mixtures of HF-HNO3-H2O or acetic acid- HN0,-H,O. The specimens for the solidus determinations were wrapped in molybdenum foil and sealed in evacuated quartz tubes. Specimens to be heated above 1300°C were enclosed in zirconia tubes inside the quartz to prevent reaction with molybdenum. The solidus determinations were carried out by heating the specimens at various temperatures for a few minutes and then examining the interior structure metallographically to determine at what temperature melting began. The 150 mesh powders for X-ray analysis were annealed in an evacuated quartz tube for 2 hr at 540 °C to remove cold work. Alloys with 22.5 atomic pct Zr and 44.4 atomic pct Zr required annealing at 700 °C for complete recovery. The grains of both of these alloys were soft and rather ductile. Both Debye-Scherrer film patterns and spectrometer traces were made for each alloy using filtered copper radiation. Results and Discussion The presence of a substantial amount of second phase in an alloy with 1.1 atomic pct Zr indicated that the solid solubility of zirconium in nickel is somewhat less than 1 atomic pct. The a0 value for the face-centered-cubic matrix in the same alloy was equal to that of pure nickel, so that the solubility must be very small. This observation is as expected from the large atom size difference. Annealed specimens of alloys with 15.0, 16.7, 22.5, 28.6, 44.4, and 59.4 atomic pct Zr showed. micro-structures that were essentially single phase. When

Jan 1, 1958

By W. T. Isbell

In the Parkes process of lead refining, after desilverization has been completed by means of the addition of zinc, there remains in the lead from 0.5 to 0.6 pct zinc. At this stage in the refining operations virtually all impurities are reduced to refined lead specifications except zinc and antimony. All excess copper, nickel and cobalt have been removed by the zinc additions. The silver remaining should not exceed 0.1 oz per ton. It then remains only to remove the zinc and antimony in order to produce refined lead. At the Herculaneum plant of the St. Joseph Lead Co., practice formerly was to complete the refining by means of oxidation in a reverberatory furnace. This process consisted of heating the lead to 1400O to 1500OF and allowing it to come in contact with excess air passing over the surface of the lead and periodically agitating the bath of lead with air. This procedure not only resulted in the loss of zinc but of a considerable quantity of lead converted to the oxide in the form of dross and fume. The fume, collected in the baghouse, together with the dross was returned to the blast furnace, thus adding appreciably to the circulating load on the blast furnace and refinery. In addition, operation and maintenance of the reverberatory furnace were expensive and required the performance of disagreeably hot and somewhat hazardous work. The process of dezincing with high vacuum was developed at this plant as a substitute for the oxidation process and finally led to lower operating costs, the practical elimination of dross and fume and the direct production of metallic zinc, which is very suitable for re-use in the desilverization process. Essentially this process consists of subjecting the lead to high vacuum, in a kettle provided with means for heating, a bell-shaped member open at the lower end so constructed that it can be immersed in the molten lead, and means for evacuating the interior of the bell and for sealing an agitator. Metallic zinc is thus produced and collects on the upper cooled part of the bell member. This metallic zinc is periodically removed and re-used in the desilverization process. The process is quite simple. It is conducted at a low temperature, thus permitting the use of iron equipment, and shows a material saving in refining costs over the old method. Equipment and Its Uses Degincing Machine.—The equipment employed in this process consists of a dezincing machine, a steel kettle of 120 tons capacity, a vacuum system and equipment for handling the dezincing machine (Fig 1). The dezincing machine consists of a compact assembly, suitable for handling with an overhead crane, consisting of a treating chamber and the agitating mechanism. Treating Chamber.—The treating chamber is a bell-shaped shell (4), similar in shape to a hemispherical kettle with a lid

Jan 1, 1949

By W. S. Pellini, F. N. Rhines

IN many of the otherwise well established alloy phase diagrams the solidus curves (temperatures at which liquid first appears upon melting) have not been located accurately, chiefly because the experimental techniques that have been most popular for determining solidus temperatures have been both cumbersome and subject to large errors. C. E. Homer and H. Plummer1 have described a sensitive and at the same time a very simple technique for locating the solidus curves in a number of alloys of tin. Their method consists simply in noting the temperature at which a specimen of the alloy ruptures under the influence of a light bending force. Although this method is not new, it has been little used for this purpose in the past; nevertheless, it appears so attractive in its possibilities of speed, simplicity, and precision that the investigations presently to be described were undertaken to verify its merits. The redetermination of the solidus and eutectic temperatures of the lead-antimony system was employed as a medium 0f experimentation. New values obtained in this way are believed to be sufficiently precise to warrant a revision of the currently accepted lead-antimony phase diagram (Fig. I). Determinations of the solidus of the lead-rich lead-antimony alloy series have been published by R. S. Dean2 and by E. E. Schumacher and F. C. Nix.3 Their points are represented in Fig. I. In their work the solidus was not located at antimony concentrations greater than 2 per cent; the curve was presumed to meet the eutectic horizontal at an antimony concentration of about 2.5 per cent. In the present studies the solidus has been traced to 3.5 per cent, where the eutectic horizontal is intersected (Table I and Fig. I). Within the range investigated by Dean and by Schumacher and Nix the agreement with their values is good. H. Seltz and B. De Witt,4 on the basis of activity measurements, have calculated that the intersection of the solidus and the eutectic reaction horizontal should lie at 3.86 per cent of antimony. The following determinations of the eutectic temperature have been reported: Roland-Gosselin,5 228°C.; Ewen,6 228°C.; J. E. Stead,7 247°C.; Campbell,8 228°C.; W. Gontermann,9 245° to 252°C.; R. Loebe,10 245°C.; E. Heyn and 0. Bauer,11 245°C.; H. Endo,12 250°C.; R. S. Dean,2 249° to 258°C.; E. Abel, 0. Redlich, and J. Adler,13 245°C.; W. Broniewski and L. Sliwowski,14 250° to 252°C.; H. Seltz and B. De Witt,4 250°C. A temperature of 249°C- appears to have been generally accepted in recent years; it is the value most frequently obtained by the method of cooling curves. Values below 249°C. have usually been obtained by inferior cooling-curve techniques; those above 250°C., by the method of heating

Jan 1, 1942

By H. C. Chao, Y. E. Smith, L. H. Van Vlack

ThE phase relationships for the MnO-MnS system have been investigated only in the eutectic region. wentrupl reported a eutectic at 1280°C (2345°F) with approximately 50 wt pct of each component, as based on the earlier work of Andrew, Maddocks, and Fowler.2 More recently silverman3 corrected these figures to 2275° ±10°F (1246°C) with 63 wt pct MnS and 37 pct MnO. Neither Wentrup nor Silverman investigated the solid-solution limits in this system. However, Wentrup did suggest with his phase diagram sketch that the maximum solubility of MnS in MnO is 10 wt pct, and the solubility of MnO in MnS is nearer 20 wt pct, Fig. 1. These figures have been utilized in subsequent publications without further verification.4 There had been evidence found in this laboratory that the above solid-solution figures were high. An investigation was therefore conducted using the following procedure. Samples containing MnO and MnS were contained in platinum foil, heated just above the eutectic temperature in a purified nitrogen at- mosphere, and quenched into distilled water. The MnO had been prepared by calcining reagent grade MnCO3 in nitrogen at 1400°C. Only MnO was detected in subsequent X-ray diffraction analyses. The MnS had been prepared by sulfur deoxidation of reagent grade MnSO4 according to the reaction: MnSO4 + S2 ? MnS + 2SO2 The details of the sulfide preparation procedure, which utilized atmospheric sulfur pressures at 950°C, are discussed elsewhere.5 The sulfur content of the sulfide was determined to be 36.8 * 0.2 pct, which matches the stoichiometric value of 36.9 pct. No evidence of either MnO or MnSO4 was detected microscopically or by X-ray diffraction. The single-phase combinations of MnO and MnS at the eutectic temperature indicate the limit of solid solution as shown in Fig. 2(a) and (c), and the revised phase diagram is shown in Fig. 2(d). The melting temperature for MnO is that provided by Glasser.6 The eutectic temperature of 1232° ± 5°C and eutectic composition of 64 wt pct MnS are in close agreement with Silverman's correction of Wentrup's work. The maximum solubility of MnO in MnS was determined to be 1.7 ± 0.2 wt pct. The maximum solubility of MnS in MnO was determined to be 1.8 ± 0.2 wt pct. These figures represent the lower limit for the presence of liquid at 1260°C and are significantly lower than the earlier estimated solid-solution figures. The authors wish to acknowledge the support of U.S. Steel's Edgar C. Bain Laboratory for Fundamental Research. The comments and critical suggestions of Drs. A. S. Keh, R. Rickett, and R. B. Snow are appreciated.

Jan 1, 1963

By Carl Floe

THE possibility of separating the various constituents of mixed ores or metalliferous products by differentially producing their soluble sulfates has not received the consideration that it deserves. The production of the soluble sulfate may be accomplished by baking the ore with sulfuric acid or by treating it with gases containing oxides of sulfur. This paper is primarily concerned with the first method. The procedure used was to mix the material with a solution containing a definite weight of sulfuric acid, baking the mixture at various tempera-tures and then leaching out the soluble sulfate with water. The amount of acid used was calculated to produce certain desired soluble sulfates. The temperature of baking depended on the decomposition points of the sulfates involved. The object was to prevent the formation of or to decompose certain sulfates while leaving a maximum quantity of another sulfate in a water-soluble form. This procedure has two possible advantages over the ordinary method of leaching with an aqueous solution of sulfuric acid. First, a more selective leach may be obtained, and second, the higher temperatures and stronger acid concentrations involved will permit a better recovery of materials that are difficultly soluble in relatively dilute aqueous sulfuric acid solutions. The method of sulfuric acid baking followed by water leaching is not new, as it has been applied in the treatment of various metalliferous materials such as the recovery of cadmium from flue dust1, the re-treat-ment of zinc leach residues2 and the working up of electrolytic slimes3.

Jan 1, 1936

By F. C. Holden, R. D. Buchheit

In the course of studies made on the Ti-Cu alloy system,' particles identified visually as TiH were encountered unexpectedly in a specimen of high-purity Ti-0.8Cu alloy. The particles did not appear immediately after preparation, but following-storage in a dessicator for several days. Since the alloy had been vacuum annealed to a low hydrogen concentration (19 ppm by vacuum-fusion analysis), the appearance of the TiH particles was suspected as a surface contamination effect. A specimen of a Ti-0.8Cu alloy had been annealed 2 hr at 790°C and water quenched, to produce the single-phase equiaxed a structure. The metallo-graphic sample was ground through 600X paper, rough-polished on a high-speed wheel with a mixture of alumina and chromic acid, awLhish-polished on a slow-speed wheel. Disturbed metal layers were removed from the final surface by alternately etching with a 1 1/2 HF-3 1/2 HNO3-95 H2O solution and polishing on the slow-speed wheel with fine alumina. The final surface was swab-etched with a 1 1/2 HF-3 1/2 HNO3-95 H2O solution, rinsed in tap water, and dried in a stream of filtered air. The micro-structure, shown in Fig. 1, was composed of equiaxed a grains without any trace of second phase. After storage in a dessicator for 3 days, the appearance of small particles in the structure became evident (see Fig. 2). After 7 days of storage, the particles were more distinct, and in some grains, began to appear as elongated particles. The specimen then was etched a second time (without repolishing) to produce the structure shown in Fig. 4. On the basis of micro- structural observation, the line markings were identified as TiH; attempts to obtain diffraction patterns from the surface for positive identification were not successful. That the structure is a surface effect was demonstrated by grinding with 240X paper to remove the surface layer. The surface then was ground, polished, and etched using the procedures described previously. The same observations were made: the freshly prepared surface was devoid of any second phase, whereas the particles again appeared after several days storage. The same effect has been observed in other a-stabilized titanium alloys, including a high-purity Ti-0.20 alloy containing only 16 ppm H by analysis. The effect has not yet been observed in unalloyed titanium, nor does its appearance appear to be consistent with any particular alloy system. From these observations, it is concluded that certain a-titanium alloys are susceptible to the formation of titanium hydride at the surface of metallo-graphic specimens. The hydrogen probably is diffused into the surface from a residual film of etchant decomposition products. Estimates of the hydrogen

Jan 1, 1961

By O. Katz, A. Spilners, M. Simnad

THE oxidation of titanium has been studied in recent years by a number of investigators.'-W ne of the most important aspects of the reaction of titanium with oxygen, which has been observed by several of the investigators, is the extremely high solubility of oxygen in the metal. A detailed study of the Ti-0 systeml% as shown that the solubility limit of oxygen in a-titanium is approximately 14.5 wt pct 0 from 800" to 1700°C. The unusual oxidation behavior of titanium may be ascribed to this phenomenon, since part of the oxygen uptake enters into solid solution in the metal and only a part is available for scale formation. A more complete study of the kinetics of oxidation of titanium would require information on the influence of oxygen in solution in the metal upon the oxidation rate. Such information would also be of interest in regard to the theory of oxidation of metals. The authors have measured the oxidation rates of oxygen-saturated and oxygen-free titanium between 800" and 1200°C and have found appreciable differences in their rates of oxygen uptake and in the activation energies. The experimental set-up was similar to that described in a previous paper.u The course of oxida-dation was followed continously by attaching the specimen to a sensitive spring balance and observing the extension of the spring with time of oxidation. High purity electrolytic oxygen was used and precautions were taken to purify the gases employed. The specimens were prepared from iodide titanium and were ground on metallographic polishing papers to a fineness of 4/0. The furnace was first evacuated and filled with argon; the specimen was lowered into the furnace, which was then evacuated and filled with oxygen, after the specimen had reached the oxidation temperature in argon. The samples were presaturated with oxygen by means of the following procedure: Each specimen was first heated in oxygen at 1200°C until its increase in weight indicated 13.8 pct by weight uptake of oxygen. The furnace was then immediately evacuated and filled with purified argon five times. The specimen was heated in the argon atmosphere for a period of 72 hr at 1200°C in order to saturate the metal with oxygen by complete absorption of the scale. The temperature was lowered to the oxidation temperature, the furnace was evacuated and filled with oxygen, and the oxidation rate was followed by noting the extension of the spring balance. No oxide film was visible on these specimens at the start of the oxidation experiments. Titanium samples in the oxygen-free as-received condition were also oxidized and their oxidation rates measured. The results are shown in Figs. 1, 2, and 3, which indicate that the initial rates of uptake of oxygen by both the oxygen-saturated and oxygen-free specimens are parabolic at these temperatures. The rate of uptake of oxygen is, however, appreciably lower

Jan 1, 1956

By S. C. Sircar, B. K. Dhindaw

KINETICS and mechanism of sulfation reactions have been studied extensively. Qualitative and quantitative data are reported for the kinetics and mechanism of sulfation reaction of copper oxide pellets;''2 yet it remains a controversial Subject. The mechanism and kinetics of the sulfation of zinc oxide powder (B.D.H. analar quality) were studied by gravimetric technique. A spring of high-carbon steel wire with accuracy well within 0.2 mg was used for measuring weight change. The sample was taken in a Pyrex glass bucket, hanging freely inside a Pyrex glass tube reactor, with a gold wire attached to the spring, the reactor being surrounded by a movable furnace (double wound type) with a temperature control within *5"C. Flow of oxygen, 60 cu cm per min, and sulfur dioxide, 45 cu cm per min, were measured through calibrated flowmeters and the mixture dried over PZ05 before entering the reactor. The experimental results are presented in Figs. 1 and 2. The plot of weight change of zinc oxide, giving the amount of zinc sulfate formed, was found to be linear over the temperature range of 500" to 700° C and up to a conversion of approximately 60 pct, as beyond it they divert from linearity.' The rate of sulfation was found to be affected by total flow rate of gases at a fixed ratio of SO2/O2, up to 60 cu cm per min, and was found to be linearly dependent on quantity of total powdered sample taken. The plot of rate of sulfation of zinc oxide, 70 mg, with varying amounts of vanadium pentoxide, at a constant ratio of gases as mentioned above, and at a constant temperature of 625"C, was also found to be linear as seen in Fig. 2. Extrapolation to zero V2O5 addition gives the same rate of sulfation, as in case of pure zinc oxide, 70 mg. Vz05 was prepared by heating ammonium vanadate at 600°C for 15 min in a muffle furnace. The available literature points out that the theory of sulfation that has received the maximum support postulates the following mechanism for sulfation of ZnS: 2ZnS + 30, = 2Zn0 + 2SOZ; the ZnO formed acts as a catalyst for the reaction, 2 S02 + O2 = 2S03. This SO3 reacts with ZnO to give zinc sulfate. It is suggested here that the above widely supported mechanism is operative in the sulfation reaction and that the rate-controlling step in the sulfation reaction of ZnO is the formation of SO3 from SO, and 0,. The authors are carrying on further work in this line, and the details will be published later. The authors wish to acknowledge the facilities provided and the help extended by Prof. P. R. Dhar, Head, Department of Metallurgical Engineering, and the permission granted to publish this note by the Director, I. I. T., Kharagpur.

Jan 1, 1969

By M. J. Stewart, F. Weinberg

PREVIOUS investigations into convective flow in molten metals have examined systems in which the length-to-height ratio of the enclosure is large, usually using long graphite boats.''2 In normal castings this ratio is generally small and it is not possible to extrapolate the results of previous work to casting configurations. The present investigation was undertaken to examine liquid flow in systems of low length to height ratio. The apparatus used for the experiments is shown in Fig. 1. The molten zone of tin is contained in an aluminum mold of inside dimensions 6.4 by 6.4 by 0.31 cm. A heating block containing a resistance coil is attached to one end of the mold and a cooling block is attached to the other end. These are used to produce a wide range of temperature differences across the mold. The mold is contained in a stainless steel tank which is heated from the outside by resistance heaters to give the required operating temperature during the experiment. Temperature measurements in the molten tin are made with an iron-constantan thermocouple PT equal to 5 atm. The phase diagram derived by us was the same as that found by Krivsky and Schuhmann, except for a similar discrepancy in one boundary line representing the gas composition in equilibrium with SO a. The vapor pressures of the pure components can be neglected at 1700°K, 1 atm and 2500°K, 5 atm. There-fore, in these cases PT is essentially the total pres-sure in the system. This is valid for all the diagrams presented by Krivsky and Schuhmann except for the one at 2500°K and 0.01 atm, where the vapor pressure of silicon is comparable to PT.6 This work was supported in part by the American Iron and Steel Institute. 'C. E. Wicks and F. E. Block: Thermodynamic Properties of 65 Elements, U S. Bur Mines, Bull. No. 605, 1963. R. P. Elliott: Constitution of Binary Alloys, 1st supplementary, McGraw-Hill Book Co., New York, 1965. 3W. A. Krivsky and R. Schuhmann, Jr.: Trans. TMS-AIME, 1961, vol. 221, pp. 898-904. 4M. F. Ancey-Moret, M. Olette, and F. D. Richardson: Mem. Sci. Rev. Met., 1964, vol. 61, pp. 169-76. 5J . F. Elliott and M. Gleiser: Thermochemistry for Steelmaking, vol. 1. Addison-Wesley, Reading, Mass. 1960. 6 J.Drowart, G. De Maria, and M. G. Ingraham: J. Chem. Phys., 1958, vol. 29, p. 1015. probe of 0.4 mm diam inserted into the open top of the mold. Radioactive sn113, in the form of a small sphere, was gently lowered into the molten tin at point A in Fig. 1 THERMOCOUPLE> WATER ARGON END HEATER QUENCH ARGON Fig. 1—Apparatus for detecting liquid flow. Point A indicates position the tracer was introduced.

Jan 1, 1970

By Leonard Bernstein, Harry Bartholomew

MANY of the properties of metals and alloys are structure dependent. Not the least of these is the grain structure. For example, in producing alloy bonds between silicon or germanium and gold-alloy foil at temperatures above their eutectics (which are, respectively, at 370" and 356°C) but below the melting points of the individual constituents, it was found that wetting was more complete with a fine-grain structure than with a coarse-grain structure. In cases where the assemblies were held together for long periods of time at elevated temperatures, wetting was not as critically dependent on the grain structure as was the case for shorter periods of time. When liquid phase forms, it first forms in the grain boundaries due to the fact that most of the impurities are concentrated at the grain boundaries and that most impurities tend to lower the melting point of the gold alloy somewhat. In addition, grain boundaries are at high-surface free-energy levels and are thus in a more favorable condition to initiate good wetting. A fine-grain structure will have more grain boundaries per unit area than coarse-grain structure thereby making wetting somewhat easier. To make the grain structure of an alloy sys- tem visible, a satisfactory method for differentiating between the different grains of which the system is composed must be employed. For the alloy bonding system previously described, the grain structure on the foil surface is of primary interest. Conventionally used etchants for the gold-alloy system will not adequately delineate grain boundaries in their structures. For this reason an alternate method is proposed for those cases where such a determination is necessary. Foil between 0.0005 and 0.005 in. thick has been used in this determination. The equipment used in this investigation was described in a previous publication.' Essentially, it employs a resistance-type filament which can be heated up very rapidly and cooled very rapidly. A thin refractory material is placed on top of the heating element. This refractory material may be molybdenum, tantalum, tungsten, or even ceramic but its thickness should be less than 0.010 in. to minimize the temperature gradient across it. The gold-alloy foil is placed on top of this refractory material. The temperature of the heating filament is raised to 900°C and cooled down to below 100°C in 5 to 10 sec. During this cycle, the foil can be observed to plastically deform due to its own weight. With no further chemical etching, this assembly is then viewed under a standard metallurgical microscope and the grain size measured. Fig. 1 is a Au-1 pct Sb alloy at 400X magnification and Fig. 2 is a Au-1 pct Ga alloy at 130X. This technique has proved to be an adequate means for controlling processing conditions in producing large quantities of gold alloy foil.

Jan 1, 1963

While mining silver at Butte in 1882, Marcus Daly is reputed to have told his partners in The Anaconda Silver Mining Co., "It looks rather gloomy. The vein is running into copper." Recently, The Anaconda Co. authorized an $11 million expenditure to be made over the next five years to ensure their "running into copper" at deep levels for many more years. The new development program is designed to permit both deeper and more widespread selective mining in "the richest hill on earth" by means of additional shaft-sinking, reactivating closed mines, and establishing new haulageways and centralized hoisting facilities. Historically a deep level selective mining camp, Butte will now continue its uncommon use of three main types of mining methods at one camp-selective mining, block caving, and open pit mining. It was thought by some people that the construction of the Kelley shaft and the subsequent beginning of block caving operations in 1952-outgrowth of the Greater Butte Project-signaled the end of stope mining of high-grade ores at depth. This feeling became more pronounced when, in 1957, the Berkeley open pit commenced operations. During recent years, the tonnage of ore mined selectively underground has decreased, but the new program will maintain the position of selective underground mining in the overall picture of the Butte operations for many years to come.

Jan 7, 1960

By Milton H. Ward

This paper contains a review and discussion of the major design parameters of a block caving mine, with special attention given to access, development, and operational methods. Information will be provided on the economics and costs of various design alternatives, and how they can influence financial results of a particular project. Technical and economic factors that pertain to an actual operation, Freeport Indonesia, Incorporated's block caving mine in Indonesia, will be provided, and highlights of the planning, development and operation of this mine will be presented. The major thrust of the paper will be to emphasize the importance of good planning and to encourage economic analysis of various methods of layout and design.

Jan 1, 1981

By W. T. Isbell

IN the Parkes process of lead refining, after desilverization has been completed by means of the addition of zinc, there remains in the lead from 0.5 to 0.6 pct zinc. At this stage in the refining operations virtually all impurities are reduced to refined lead specifications except zinc and antimony. All excess copper, nickel and cobalt have been removed by the zinc additions. The silver remaining should not exceed 0.1 oz per ton. It then remains only to remove the zinc and antimony in order to produce refined lead. At the Herculaneum plant of the St. Joseph Lead Co., practice formerly was to complete the refining by means of oxidation in a reverberatory furnace. This process consisted of heating the lead to 14000 to 1500°F and allowing it to come in contact with excess air passing over the surface of the lead and periodically agitating the bath of lead with air. This procedure not only resulted in the loss of zinc but of a considerable quantity of lead converted to the oxide in the form of dross and fume. The fume, collected in the baghouse, together with the dross was returned to the blast furnace, thus adding appreciably to the circulating load on the blast furnace and refinery. In addition, operation and maintenance of the reverberatory furnace was expensive and required the performance of disagreeably hot and somewhat hazardous work. The process of dezincing with high vacuum was developed at this plant as a substitute for the oxidation process and finally led to lower operating costs, the practical elimination of dross and fume and the direct production of metallic zinc, which is very suitable for re-use in the desilverization process. Essentially this process consists of subjecting the lead to high vacuum, in a kettle provided with means for heating, a bell-shaped member open at the lower end so constructed that it can be immersed in the molten lead, and means for evacuating the interior of the bell and for sealing an agitator. Metallic zinc is thus produced and collects on the upper cooled part of the bell member. This metallic zinc is periodically removed and re-used in the desilverization process. The process is quite simple. It is conducted at a low temperature, thus permitting the use of iron equipment, and shows a material saving in refining costs over the old method. EQUIPMENT AND ITS USES Dezincing Machine.-The equipment employed in this process consists of a dezincing machine, a steel kettle of 120 tons capacity, a vacuum system and equipment for handling the dezincing machine (Fig I). The dezincing machine consists of a compact assembly, suitable for handling with an overhead crane, consisting of a treating chamber and the agitating mechanism. Treating Chamber.-The treating cham¬ber is a bell-shaped shell (4), similar in shape to a hemispherical kettle with a lid

Jan 1, 1947

By R. W. Hertzberg, F. D. Lemkey, J. A. Ford

The effect oj unidirectional solidification upon the microstructural, crystallographic and mechanical characteristics of high -purity A1-A13Ni eutectic a1loy specimens has been investigated. varying the growth rate caused the morphology of the Al3Ni phase to gradually change from platelets to rods. In addition, the effect of increasing the rate of planar liquid-solid interlace movement is to decrease the Al3Ni rod diameter and inter-rod spacing. The platelets and rods grow approxinlately parallel to each other and are aligned in a faulted substructure. A single unique crystallographic relationship hetween the A1 and ASNi phases was found and may he descrihed as : interface {001} Al3Ni {331} Al and growth direction <010>Al3Ni || <110> Al The platelet interlace and the aligned rows of rods are both uniquely defined by the above statements. Alignment of the Al3Ni phase by unidirectional solidification has given rise to a threefold increase in strength over that exhibited by specimens with an as-cast microstructure. These results illustrate the possible use of this eutectic alloy as a zohisker -reinforced structure. It has been demonstrated by Winegard et al.,1 Kraft and Albrighht,2 Chilton and winegard,3 Chad-wick,4 Yue,5 Tiller,6 and others that a planar liquid-solid interface may be established in binary eutectic alloys by proper control of heat flow during the solidification process. The unidirectional movement of such an interface results in a eutectic mi-crostructure consisting of an essentially parallel array of the two phases over an entire ingot. Two dominant phase micromorphologies have been produced using this technique: that of substantially parallel alternating lamellae of each phase or long thin parallel rods of one phase imbedded in a continuous matrix of the other phase. While not all eutectics can be controlled in this manner, it has been shown that a "normal" eutectic may, under properly controlled conditions of purity, liquid and solid thermal gradients, and solidification rate, be forced to solidify with essentially parallel phase particles. Kraft and Albright2 have demonstrated that, when the growth rate is too rapid or when the thermal gradient in the liquid at the growing interface is too low, a layer of constitutionally supercooled liquid, formed by impurity build-up, will stabilize a cellular rather than a planar interface. The unidirectional movement of the cellular interface forms the macrostructure of eutectic colonies.7 Chilton,3 Chadwick,4 and Tiller,&apos; studying eutectics of zone-refined Pb-Sn, A1-CuAl2, Zn-Sn, and Cd-Zn systems, noted that in the absence of impurities a planar interface is stabilized even when rapidly solidified. Although it is generally agreed that impurities break down a planar interface to form the eutectic-colony macrostructure, at present the origin of the varying micromorphologies (i.c., plates or rods) in a given normal eutectic alloy is not completely understood. Tiller8 has predicted that the micromor-phology produced may be dependent on solidification rate (i.e., at fast rates a rodlike structure is preferred whereas at slow rates a lamellar structure should form), and also suggested that rod formation may be favored at large phase-volume ratios. yue5 has experimentally verified Tiller&apos;s prediction in the eutectic Mg-Mg17Al12 by observing a lamellae-to-rod transition with increasing growth rate where the phase-volume ratio was approximately 2.3:1. However, Hunt and chilton9 more recently unidirec-tionally solidified over a wide range of growth velocities six different eutectic systems with phase-volume ratios between 12:l and 2.7:1 and observed no lamellae-to-rod transition. chadwickl0 has proposed that the change in micromorphology in some eutectic alloys is due entirely to the presence of impurities and further states that the lamellar structure is the characteristic structure of pure eutectic alloys even when the phase-volume ratio is as great as 12:l. Kraft11,12 has shown that the parallel lamellae in certain eutectic alloys assume a unique crystallographic relationship during unidirectional growth. This preferred crystallography may be developed

Jan 1, 1965

By E. Clarence Peterson, Rip V. Thompson

ALTHOUGH the metallurgical treatment for the Ophir Hill tailings was developed more than 20 years ago and in the interim a vast amount of experimental work was done with later-developed flotation reagents, it was found necessary to revert to the original oil reagents and their combinations with specific chemical reagents in order to obtain results of economic value. This is of especial interest because in present ore-treatment investigations later-developed flotation reagents are used almost exclusively and the coal-tar derivatives, except in certain patented reagents, are virtually neglected. These coal-tar oils and derivatives still have a definite field of application, especially in the treatment of some oxidized ores when modified sulphidization is desirable. The Ophir Hill tailings deposit is in Ophir Canyon, Tooele County, Utah, about 55 miles southwest of Salt Lake City and 25 miles south of the Tooele smelter of the International Smelting and Refining Co. A road good for trucks leads directly from the Ophir mill to the smelter. TAILINGS These tailings were produced in the old Ophir Hill gravity concentrator previous to 1921. The original mill ore was a heavy lead-zinc-copper-silver-iron sulphide ore formed by limestone replacement. The gangue is a highly siliceous lime. An assay of the original milling ore would approximate 1.0 per cent Cu, 6 to 8 per cent Pb, 6 to 8 per cent Zn, 8 to 12 oz. Ag, 18 to 25 per cent Fe, 2.0 per cent CaO, 20 to 5o per cent insoluble. The gold content was 0.005 oz. or less. Tailing was impounded in a fairly wide area in a canyon about ½ mile below the mill. A number of dirt embankments were thrown up to form settling ponds and the tailing was allowed to flow through them in series. Coarser sand deposited at the upper end of the ponds and the finer slimes settled gradually as they passed through the successive settling ponds. As soon as the coarser tailings filled a pond, the tailings were by-passed to the next pond and teams and scrapers were used to haul the deposited sand onto the banks to make them higher. The final dump was approximately 2500 ft. long and 250 ft. wide at the bottoms, with banks sloping up at 45°. The coarser sands predominated at the upper ends and sides of the ponds while the proportion of slimes increased steadily at the lower ends. The lower end of the pond contained over 50,000 tons of slime tailings, 90 per cent of which passed through a 200mesh screen. SAMPLING The entire tailings deposit was sampled by drilling vertical holes so placed as to give data concerning size segregation and the contour of the underlying ground as well as to secure an assay sample. Different types of auger and post-hole drills were tried. None was found satisfactory because the stratified layers of

Jan 1, 1940

By Alden Greninger

MANY recent contributions in the field of theoretical metallography have been concerned with crystallographic definitions or descriptions of various phenomena. The lattice orientation of the crystal being studied must be known before such definitions can be made. Practically all the metallic crystals used do not have crystal faces-other than the unknown "plane of polish"; hence resort must be had to other means of determin-ing orientations. Some use has been made of etch-figures, slip bands, optical properties, etc., for this purpose; but for. accurate results, and because of the fewer restrictions to their use, the various X-ray methods have been the most satisfactory.1 The methods that, have been used may be divided into two general classes: (1) oscillation methods, and (2) transmission Laue methods-2

Jan 1, 1934

By Chris Weyand

Throughout the past 40 years we have been drawing inevitably closer to the conclusion that coal will be our chief source of energy. Our reserves of natural gas and petroleum, on which we depend, are dwindling. Knowing that we will soon reach our physically extractable limit of these energy sources. The US government has created many incentives to encourage energy companies to transfer technology and resources from natural gas and petroleum to coal. Such encouragement is stimulated through taxation methods. Coal mining engineers deal mainly with production, which involves planning and design throughout the acquisition, exploration, development, production, and reclamation of a mine. Though taxation is not one of the areas analyzed by the engineer, it should be considered and understood as it is involved in the mine life cycle. This paper will discuss coal taxation in the five stages of mining and will explore certain tax saving decisions.

Jan 1, 1983

By Dale Dixon, A. R. MacPherson

The present comminution process in principle has remained more or less the same for the last 50 years. The majority of the changes have been in the fields of equipment improvements and size increases of prevailing systems. The development and utilization of completely new systems has virtually been non-existent. As energy and steeel costs increase and lower grade ores are mined, the comminution costs per ton of ore treated must improve accordingly or the operations will not be profitable. The chapter examines further developments in existing systems and touches on relatively unexplored methods of comminution appropriate for future needs.

Jan 1, 1982

By W. F. de Lange, M. I. Brittan

A mathematical model of the Western Holdings gold concentrator in South Africa was generated by multiple regression analysis of routine plant operating data. Interpretation of the computer results and application of the model indicated that significant economic improvement in gold recovery could be realized by relatively minor modifications to the circuit. The modifications made are described and the economic benefits resulting from the increased gold production are assessed. Comparison, of the projections of the mathematical model with actual plant performance is illustrated.

Jan 1, 1976

By Harold L. Hinds

The Homestake Ledge or Lead was discovered by Fred and Moses Manuel on April 9, 1876. They recovered a small amount of gold from the quartz outcrop, but, unfortunately, they were prospectors not hardrock miners. Their claim was sold to a group of experienced California mining men in 1877. These mining men incorporated their holdings as the Homestake Mining Company. This was the beginning of one of the greater gold mining enterprises of the Western World. Metallurgical operations progressed over the years from an eighty- stamp grinding circuit using amalgamation for recovery to the more complex circuit of today. The Homestake Mine, today, is an underground mine with the deepest winze reaching the 8,000-foot level. The ore is hoisted to the surface by two primary shafts, the Ross Shaft and the Yates Shaft. See Figures 1 and 2. It is crushed in three-stage crushing plants located in the headframes. From here, it is trammed to the mill fine- ore bins by electric locomotives. The Lead grinding plant milling circuit, known as the South Mill, receives ore from the crushing plant at minus 112 inch (Figure 3). This material is dumped into overhead fine-ore bins from five- ton rail cars. The fine-ore bins have a live capacity of about 3,700 tons. The South Mill bins can also be fed by a conveyor system which reclaims ore from an 80,000-ton capacity stockpile. This stockpile is built up from peak mining production periods and available surface low grade material. It is then reclaimed during low mine production periods. The grinding circuit at the South Mill consists of four rod mills in open circuit followed by four ball mills in closed circuit. Three of the rod mills are 6-foot by 12-foot Marcy&apos;s. The other is

Jan 1, 1983