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By E. M. Haynsworth

The disposal of phosphatic clays has presented a problem to the phosphate industry since the mining of phosphate rock first began. As the flotation process was introduced in the late 1920's, the problem of clay disposal was increased significantly since clay is present in proportionally greater quantities in reserves yielding flotation concentrates. Because of their physical character and the tremendous surface area per unit volume, these clays have the characteristic of retaining large volumes of water as they are slurried during processing. The phosphatic clays with which we are dealing consist primarily of apatite, wavellite and montmorillonite. In the beneficiation process, we remove the -150 mesh material as waste. However, the particles are essentially smaller than -325 mesh, and more than 70% are smaller than three microns. The disposal of clay and the recovery of a portion of the water require settling areas with a capacity ranging up to 2 cubic feet for each cubic foot of phosphate matrix mined. According to a recent study by the Bureau of Mines, the average cost for disposal of these wastes is 24.5 cents per ton of, product (U. S. Bureau of Mines Circular 8404). This is a major cost of phosphate rock production.

Jan 1, 1970

By R. S. Shrode, R. F. Bunting

Transportation is one of several important increments in the total market cost of most commodities. On a national average it has been reported that transportation accounts for about 25% of the total cost of a product at the market place. And industrial minerals are no exception. In fact, the proportionate cost frequently exceeds the 25% average. Further, it may actually determine the economic value of a deposit. Transportation possibly may be thought of as a specialized subject, of concern only to a traffic manager and of only nominal interest to the mining engineer, the geologist or other technically oriented member of a mineral producers' staff. However, in the industrial minerals field, interest is not commonly confined to exploration, development, production or marketing, but is in contrast an overlapping one. In a very real sense this broad area of interest in the total economics of the business is the specialty of the industrial minerals man. And transportation is a vital part of the total picture.

Jan 1, 1970

By Frank Arthur Forward

FRANK ARTHUR FORWARD was born in Ottawa and attended the University of Toronto, where he received a B.A.Sc. in chemical engineering in 1924. From 1924 to 1929, he was a smelter operator and research metallurgist with the Consolidated Mining and Smelting Company of Canada Ltd. (Cominco) at Trail, B.C. Moving to Australia in 1929, he was assistant smelter superintendent, Mount Isa Mines Ltd., Queensland, until 1934, when he returned to Canada, spend-ing one year as metallurgist with B.C.

Jan 1, 1970

By R. L. Ash, R. R. Rollins, C. J. Konya

An investigation was performed to determine conditions for optimizing the spacing of simultaneously initiated multiple explosive columns. This was done by using models of mortar, dolomite, and Plexiglas with 10-grain mild detonating fuse as the explosive charge. It was desired to simulate blastholes with multiple primers initiated by detonating fuse or when high-velocity explosives are used in low-velocity materials. It was found that optimum spacing between multiple charges was strongly influenced by charge length. At less than optimum charge length, the spacing at which complete shearing was possible between adjacent charges decreased exponentially with a subsequent loss of broken material volume. For charges fired simultaneously, larger burdens and spacings were possible as compared to those necessary for single-crater charges. For each material studied, there was a characteristic optimum charge length and a maximum attainable spacing at any given burden. Proper selection of the spacing distance between charges is fundamental to successful blasting. Its value directly affects the cost of drilling and explosives used per unit of broken material. In addition, the choice of a spacing that is Compatible with a given set of blasting conditions aids in the control of fragmentation sizing, ground vibrations, overbreak, and throw which in turn, influence other production costs. For example, normally loaded blastholes that are spaced too closely invariably promote overbreak and usually give coarse fragmentation. Unless care is taken, airblast and violent flyrock will occur and under certain conditions cutoffs and misfires may result. Too large a spacing, on the other hand, frequently leads to conditions that form bootlegs or toes. The choice of a particular spacicg to use, however, is largely a matter of individual experience and judgment, usually based on trial and error. Very little is known or can be found in the literature with regard to how the spacing between charges is related to field conditions and charge geometry. As a general rule, the firing time sequence of adjacent charges and properties of a material are thought to have the most significant influence on the spacing distance best suited for any given field condition. For example, delayed initiation of adjacent charges usually always requires a closer spacing than when charges are fired at the same time. This should be expected if one considers that the energy normally dissipated and lost in the surrounding ground from charges fired independently would be captured and utilized for breaking material between charges when they are initiated together. Spacing can be extended also when charges are aligned with structural planes of a material, such as jointing, along which shearing is relatively easy. It is customary to relate the spacing (S) between charges to their common burden (B) in the form of a spacing ratio, or SIB. The burden normally is considered as the optimum depth or distance from any single charge perpendicular to the nearest free or open face at which the desired fragmentation and maximum crater yield are obtained. For production blasting, value of the ratio is generally considered to vary from 1 to 2, depending on conditions.1-6 When adjacent charges are fired independent of one another, the value varies from 1 to about 1.4, the closer amount being employed to square corners or produce craters having the ideal 90" apex angle. The larger ratio is the geometric balance value for craters having an apex angle of 135". The basic ideal crater forms in the plane of the charge diameter for charges fired independently are shown in Fig. 1. In the event charges are fired simultaneously, geometric balance in the plane of their charge diameters suggests that a spacing ratio near 2 would be appropriate, as illustrated by Fig. 2. In practice, however, some compromise ratio value must be selected to conform with the specific ground conditions. An example would be where the jointing planes tend to produce 60° or 120° crater angles, the appropriate geometrically balanced charge arrangement being given by Fig. 3. In this condition, the spacing ratio is 1.15, not 1 or 1.4 as suggested for the 90° cratering of independently fired adjacent charges. In view of the foregoing, it would seem logical to assume that whenever charges all having the same burden are fired at the same time, spacing distances always can be greater than those permitted by charges fired independently. In practice this is not the case, however.

Jan 1, 1970

By H. J. de Haan, L. Schenk

Scope for Thermal Recovery in Shell's Heavy Oil Fields in Venezuela The main heavy oil reservoirs on the East coast of Lake Maracaibo (Fig. I), known as "Bolivar Coast", initially contained some 20 billion bbl of oil in place with gravities in the range of 10 to 15° API. The current total production rate is about 400,000 B/D. Since the recovery to date is on an average only 12.5 percent of the initial oil in place, these fields offer a vast potential for secondary recovery methods. The reservoirs are characterized by moderate depth (generally 1,000 to 3,000 ft; maximum 5,000 ft), good formation properties such as net oil sand thickness (50 to 300 ft), high porosity (30 to 40 percent) permeability (1 to 3 darcies) and oil saturation (initially about 80 percent) and high oil viscosity (100 to 10,000 cp in situ), all of which are favorable for the application of thermal recovery processes. Since 1957, steam drive, steam "soak" and underground combustion have been tested or are being tested in this area.1-3 This paper deals with a major steam drive test in the Tia Juana field. The test was commenced in Sept., 1961, on the basis of encouraging results of laboratory experiments and pilot field tests, which will be discussed briefly. Early Laboratory Investigation of the Steam Drive Process In 1956 a series of model experiments was carried out in the Koninklijke/Shell Laboratorium, Amster- dam, to investigate the displacement of heavy oil by steam.' The prototype studied was a horizontal, un-consolidated heavy oil reservoir, subjected to a linear steam drive. The model, representing a slab from this reservoir along the main flow direction, consisted of a steel tube filled with oil-saturated sand, provided with layers of rock at the top and bottom to simulate conductive heat losses to cap and base rock. Lateral heat losses, perpendicular to the main flow direction, were minimized by insulating the sides of the model. Heat flow and viscosity and gravity forces were scaled by taking the steam injection rate and sand permeability inversely proportional to the geometric reduction factor of the model, and using materials (oil, steam, cap rock, etc.) identical with those of the prototype. The main feature of the process was a frontal displacement mechanism that was observed within wide ranges of injection pressure and rate, oil viscosity, initial oil saturation and sand permeability. The high stability of this front can be attributed to the high volumetric steam flow rates, combined with the reduction in oil viscosity. High recoveries were obtained due to the low residual oil saturation of about 15 percent in the zone swept by steam. The latter value pertains to the specific test conditions (heavy oil, uncon-solidated sand) and may be lower for light oil, due to the effects of steam distillation. The applicability of this experimental approach is limited, as it does not take into account the fact that

Jan 1, 1970

Location & Permit Status Three major Mesozoic Tertiary basins exist along the continental shelf area of southeastern Australia, known as the Bass Strait area. They are the Gippsland (east of the longitude of Melbourne) Otway (west of Melbourne) and the Bass (underlying central Bass Strait between Victoria and Tasmania). The total area, both onshore and offshore, is approximately 100,000 square miles. In March 1960, the Broken Hill Pty. Co. Ltd. acquired exploration title to the continental shelf areas of the Gippsland, Bass and Otway basins from the States of Victoria and Tasmania. The title to that part of the Otway Basin in South Australian waters was granted in July, 1961. Other exploration permits, being 8 miles wide in Gippsland, held by Woodside :E.) Oil Co. N.L., and 15 miles wide along the southwest coast of Victoria, held by Mobil-BP-Interstate Oil (Frome-Broken Hill Co. Pty. Ltd.), were in existence prior to granting the B.H.P. title. Subsequently, Shell, B.H.P., Esso and Magellan have been granted title to the remaining areas of the continental shelf in the basins under discussion. In three separate agreeme negotiated between 1964 and-1966, Esso obtained farm-in rights to B.H.P.'s Gippsland, Bass and Ott' Permits.

Jan 1, 1970

By Barry D. Lichter, Pierre Sommelet

High-temperature heat contents of InAs and GaAs were measured over the temperature range 400°K to temperatures above the melting points using a di-phenyl ether drop calorimeter. Smoothed values of the thermal properties have been derived and tabulated at even temperature interuals . The heats of fusion of InAs and GaAs were determined as, respectively, 8790 ± 200 and 12,590 ± 300 cal per g-atom at the respective melting points of 1210º ± and 1513º ± 1°K, yielding entropies of fusion of 7.26 * 0.16 and 8.32 ± 0.20 cal perºK per g-atom, respectively. A general discussion of the thermal properties of III-V com-pounds is given. THE low-temperature heat capacities of InAs and GaAs have been determined by Piesbergen.1Using a dynamic temperature modulation technique, Miller2 has determined the heat capacities of the compounds as well as GaSb and Ge between room temperature and 650°K; however, it has been observed that values determined by drop calorimetry for GaSb3 and for germanium4 are consistently about 5 pct lower than Miller's results. Cox and Pool10 report values for the heat content of InAs and GaAs and the heat of fusion of InAs. No value for the heat of fusion of GaAs appears to have been previously reported. This paper reports results of high-temperature heat content investigations of InAs and GaAs measured in the range 400" to above 1500°K employing a diphenyl ether drop calorimeter. These results, together with our previous results on III-Sb compounds3 as well as heat-content measurements5 of solid InP and Gap, are used in a general discussion of these compounds. EXPERIMENTAL PROCEDURES High-purity samples of the compounds, stoichio-metric to ±0.1 at. pct, were supplied in the form of crushed crystal fragments by Dr. Carl Thurmond of the Bell Telephone Laboratories. The sample preparation, description of the calorimeter, and details of the heat content measurement have been previously described.3 Samples were sealed in evacuated silica capsules and suspended by a Pt-Rh wire in a Pt-wound vertical tube furnace, heated to temperature

Jan 1, 1970

By D. P. Schlick

The German experience in dealing with the health hazard of respirable dust in underground mines is reported from first-hand observation in this Bureau of Mines paper. The report covers-- 1. The control of dust by water infusion, water spraying, and ventilation. 2. The sampling devices (tyndalloscope , konimeter , and BAT-I) and methods used to determine the dust concentrations in mines. 3. The systems of classifying hazardous and safe dust levels in mines. 4. The research facilities devoted to theoretical and applied work on mine dust and pneumoconiotic disease. Tile major research projects on dust and disease now underway in West Germany are outlined in an appendix.

Jan 1, 1970

By B. Chalmers, L. Bäckerud

The growth temperature of dendrite tips has been measured in a binary alloy system, Al-Cu, as a function of rate of formation of solid phase. Metallographic examination has rerealed the dendritic structure dereloped during different stages of the freezing process and factors influencing the undercooling, such as solute buildup, surface tension, and kinetic effects, have been estimated. Practical implications of the results are discussed. THE development of a dendritic growth pattern during the solidification of metals and alloys has an influence at least equal to that of grain size on the properties of the casting or ingot being produced. and consequently much interest has been devoted to the study of this phenomenon. A review of the knowledge gathered up to 1964 has been given by Chalmers.' Among recent publications in the area the works of Jackson et al 2 and Flemings and his coworkers3,4 are of special interest. This work is an experimental study of the factors which influence the development of a dendritic pattern during the solidification of a typical binary alloy system, A1-Cu, when there is no heat source in the liquid. This means that the temperature and the temperature gradient are controlled by the heat extraction from and latent heat evolved by the growing crystals. EXPERIMENTAL PROCEDURE Samples were prepared from a master alloy made up of aluminum, 99.995 pct, and copper. 99.99 pct, of the concentrations given below and of sufficient size to fill cylindrical graphite molds of 35 mm ID and 35 mm height. Bottom. wall. and lid of the graphite molds were all of 6 mm thickness, Fig. 1. Thin ceramic thermocouple shields could be introduced at different positions in the sample through holes in the lid. The assembly was heated in a small vertical resistance furnace to a suitable temperature. 700° to 720°C. which is well above the melting point of the sample. It was thereafter taken out of the furnace and chilled at selected rates, and the temperature changes in the sample were recorded. For high sensitivity in the temperature recordings, a reference thermocouple was held in a solidifying pure aluminum melt. which was very well insulated and therefore was at a constant temperature. 660°C, for nearly half an hour. The thermocouples were connected so that the difference voltage was fed into the recorder of which the most sensitive range was 1 mV. Temperature variations could be read to within 0.l°C. The thermocouples used were of the Chromel-Alumel type (28- or 38-gage wires) and the recorder a two-pen Moseley strip chart recorder. The different rates of chilling were obtained in the following way. a) The graphite crucible was wrapped in one or two layers of an insulating. feltlike material (Fiberfrax paper from the Carborundum Co.). The drop in temperature per unit time after complete solidification could thereby be varied from 8" to 18°C per min. This corresponds to a rate of heat extraction in the given samples from 0.08 to 0.17 cal cm-' sec-I. 6) The bare graphite mold was placed free on a piece of firebrick. The measured temperature drop was around 50°C per min (-0.5 cal cm-" ' sec- I). c) By blowing compressed air through a tube with many holes surrounding the mold in a ring shape. the rate of heat transfer could be raised to 150' to 200°C per min (-1.7 cal cm-3 sec- I). d) Cooling rates around 1000 C per mill (-9.4 cal cm-3 sec-l) were obtained by substituting the air in the device used in c by water. The range of cooling rates used in these experiments correspond to technical casting processes from slow solidification as obtained in sand molds to very rapid freezing which occurs in some continuous casting and welding processes. EXPERIMENTAL RESULTS The A1-Cu system was chosen for the experiments: the relevant part of the phase diagram is reproduced in Fig. 2. Observation of the Center Temperature. When a thermocouple is inserted in the center of the melt. the type of cooling curves obtained at slow solidification rates differs little from numerous published curves from similar systems. and the solidification pattern

Jan 1, 1970

By B. A. Eaton

The subject of many discussions and technical papers in the last 20 years has been the prediction of the well-bore pressure gradients that are required to induce or extend fractures in subsurface formations. The subject merits this attention because of the frequently recurring problems that arise from an inability to predict fracture pressure gradients. Encountered in several common types of operations in the oil industry are problems associated with the prediction of formation fracture pressure gradients. When wells are being drilled in both new and old fields, lost circulation is often a very troublesome and expensive problem. Complete loss of circulation has been disastrous in some cases. Many times, such disasters could have been avoided if techniques for calculating fracture pressure gradient had been employed in the well plans, and if casing strings had been set, and mud weight plans had been followed accordingly. In areas of abnormally pressured formations, the prediction of fracture gradients during the well-planning stage is extremely important. In fact, it is as important as the prediction of formation pressure gradients, which has received a great deal of attention in recent years. There are several published methods used to determine fracture pressure gradients. However, none of these methods appears to be general enough to be used with much reliability in all areas. In 1957, Hub-bert and Willis published a classical paper that included the development of an equation used to predict the fracture extension pressure gradient in areas of incipient normal fau1ting.l Overburden stress gradient, formation pore pressure gradient and Poisson's ratio of rocks were the independent variables that were shown to control fracture pressure gradient, the dependent variable. In 1967, Matthews and Kelly published another fracture pressure gradient equation that is different from that of Hubbert and Willis in that a variable "matrix stress coefficient" concept was utilized.3 Later the same year, Costley wrote about a similar idea.5 Goldsmith and Wilson used a least-squares curve-fitting technique and field data from the Gulf Coast area to correlate fracture pressure gradient with formation pore pressure gradient and formation depth.' They noted that the fracture pressure gradient increased with increasing depth while the pore pressure gradient remained constant. In each of these cases, the problem for which a solution was sought was to determine the bottom-hole pressure gradient required to initiate or extend a fracture. Results of the previous work show that fracture pressure gradient is a function primarily of overburden stress gradient, pore pressure gradient, and the ratio of horizontal to vertical stress. There is argument for a fourth variable in that in many cases breakdown fracture pressure gradient is greater than the fracture extension pressure gradient. However, if the fracturing fluid is able to penetrate the formation through the pores or existing cracks, there is very little

Jan 1, 1970

By E. C. Baker

Costs of disposing of steel slag are estimated based on the present method and a proposed alternate method for a steel mill in the Pittsburgh area. Estimated cost data were calculated by the Bureau of Mines and do not reflect the investment and operating costs and credits of a specific steel mill and processing plant. The alternate method of steel slag disposal would result in an appreciable cost savings to the mill and concurrently supply the industrial and commercial markets with low-cost, high-quality construction materials.

Jan 1, 1970

By A. R. Rule

Laboratory tests were conducted to study removal of magnesium impurities from western phosphate rock concentrates by selective leaching. The systems evaluated included leaching with dilute sulfuric acid, with ammonium sulfate, and in an aqueous medium with sulfur dioxide gas. Set standards for the magnesium impurity level of phosphate rock concentrates are not available, but industrial practice has indicated that magnesium in excess of 0.3 percent MgO causes problems in phosphoric acid manufacture. Removal of magnesium was accomplished with all three systems. With dilute H2S04 the magnesium level was reduced to less than 0.1 percent MgO with an accompanying P205 loss of 6 percent. Magnesium oxide levels from 0.3 to 0.2 percent were obtainable with ammonium sulfate. However, the system requires concentrate roasting at temperatures of 900° C or higher, heated pulps (100° C), and large quantities of ammonium sulfate. Selective leaching was obtained using S02 gas in an aqueous concentrate slurry. Magnesium oxide content was reduced to less than 0.1 percent without loss of phosphate.

Jan 1, 1970

Up to 19 chemical parameters were studied in flotation pulps in six Australian lead-zinc concentrators with the aim of obtaining a better understanding of the nature of, and changes in, the chemical environment. This paperdiscusses pH, redox potential, and oxygen concentrations, which were all measured continuously.Two plants operated at natural pH in slightly acid circuit and one at natural pH in alkaline circuit, and pH was stable at all these plants. In two plants varying amounts of cement fill were present in mill feed and pHwas variable. At the other plant pH was controlled with soda ash and lime.Redox potentials became more positive as the pulp passed through the plant. The minimum change was 20 mV and the maximum about 250 mV. Potentials tended to be variable at any point early in the circuit, but were more stable later on. In some plants reagent additions had no noticeable effect on redox potentials. However in the two plants operating in acid circuit the redox potential of .zinc rougher feed was markedly affected by changes in copper sulphate addition rate. Sodium sulphite additions had an effect at other plants.Oxygen concentrations in grinding circuits tended to be low (30 per cent of saturation) and very variable. Flotation was usually conducted in pulps at least 90 per cent saturated with oxygen. Addition of sodium sulphite had a marked effect on oxygen content in one plant, but less effect in two others.Statistically significant relationships were found in some plants between pH and redox potential, redox potential and oxygen saturation, and oxygen saturation and pH.Improved control of copper sulphate and sodium sulphite additions could result from applications of this work.

Jan 1, 1970

By J. E. Lawver

On the basis of both laboratory and pilot-scale tests, a newly developed high-tension electrostatic process is technically and economically feasible for the production of iron ore superconcentrates containing 0.02 to 0.20% silica at iron recoveries of 85 to 90%. The process can be used to recover both a superconcentrate and a pellet-grade concentrate, from a feed containing middlings, and the system is not unduly upset from changes in feed structure and composition. The adverse affect of excess quantities of material finer than 30 microns, i.e., above 10% in quantity can be avoided by a desliming step. Finally, because of the cost of drying, the best feed for the process appears to be dry-ground ores as well as spiral, flotation, or magnetic concentrates.

Jan 1, 1970

By D’Arcy A. Shock, J. G. Davis

Continental Oil Co. has spent more than four years of research on the fundamentals of recovery of potash by solution mining. This work included laboratory investigations of salt solution, brine flow, selective extraction of KC1 from sylvinite, and control of cavity geometry. Laboratory results were expanded by means of a series of field tests, and subjects such as well completion and hydraulic fracturing were added to the investigation. Both single-well and multiwell mining systems were studied in the field work. This paper discusses one field test in which potash was solution mined from a thin sylvinite ore zone in the Carlsbad Potash Basin. A two-well mining system was used for this test.

Jan 1, 1970

By Frank G. Snyder

INTRODUCTION Numerous deposits of lead and zinc sulfides, often accompanied by barite and fluorite, occur in carbonate host rocks throughout central and eastern United States. So similar are the deposits in broad general characteristics that they long have been classified as Mississippi Valley type. They differ widely, however, in details such as metal ratios, age of host rock, type of ore-bearing structure, and probable time of mineralization. The theories developed to explain the source and transport of the metals are many and are known to all. The deposits have been termed telethermal on the basis of their presumably shallow occurrence and the attempt to relate the metals to the classic hydro- thermal sequence. 1t is now recognized that deposits of this type can occur at any depth; the only depth limitation is thickness of the sedimentary section. Ideas on the source of the metals also have changed, A rapidly growing body of evidence suggests that the metals were brought by concentrated brines from nearby sedimentary basins. The hypothesis preferred today by many geologists postulates that the metals were deposited during the course of normal sedimentation in depositional basins of miogeosynclinal or orthogeosynclinal type. At some time after deposition of the sediments, the metals were carried out of the basins by migrating brines; the brines moved laterally and upward into structurally high areas; they entered any reservoir-like structure, either primary or secondary in origin, that happened to be available; in this reservoir, possibly in a sulfur-rich environment, the metals were precipitated as ore minerals.

Jan 1, 1970

By D. U. Deere

The title of this chapter is "Indexing Rock for Machine Tunneling" or, a simplified approach to a very difficult problem. I do not think that we can divorce the considerations for conventional tunneling from that of the requirements in terms of geologic information that is necessary for machine tunneling. Experience has shown that in conventional tunneling, the rate of progress achieved in driving a tunnel and the cost involved can be almost directly related to the general rock quality and to the water conditions actually encountered. The very method selected for driving the tunnel or perhaps adapted during the construction period, that is, full-face, top-heading and bench, or multiple-drift, is largely determined by the degree of severity of the adverse geological features. The amount and type of support required to stabilize the rock behind the advancing face are also directly related to the ground conditions. The support may range from bald-headed to rock bolts placed as needed- perhaps pattern rock bolting, light steel sets, all the way to closely spaced heavy steel sets with full lagging and invert struts. At times crown bars, spiling, and breast boarding must be used to advance the face. Shotcrcte applied immediately to the arch and upper walls is a fairly recent advance in providing both temporary and permanent support. Ground-water inflows can further reduce the rate of advance, causing stability problems and reducing the overall efficiency of the mining crew. Required pumping may also be very expensive depending on the grade of the tunnel and the amount of water. And in shafts, of course, there is the pumping problem. To offset the adverse effect of the ground-water inflows, one must often resort to grouting ahead of the face. But this very operation which helps cut down the water inflow also increases greatly the time required for mining, and of course, there is additional cost involved in the actual operation. The conditions described above do not occur uniformly throughout the rock mass but rather they are concentrated

Jan 1, 1970

By R. M. Kleine

In the fall of 1968, Sheridan Geophysics leased the Renzy Mine property 50 miles northwest of the town of Maniwaki, Quebec. A 500-tpd flotation mill was dismantled at Belleterre, Quebec, trucked into the mine over a winter road and re-erected during extreme winter conditions. Approximately one billion gallons of water were removed from Renzy lake to expose the orebody. A 3,000-hp diesel generating plant and a self-supporting community were set up during the early part of 1969. This was a remark-able achievement considering that most of the work, including the pouring of foundations and the draining of the lake, had to be done during the winter months. Despite this, the project was completed ahead of schedule and one million dollars below budget. The mill is now treating an average of 800 tpd.

Jan 1, 1970

By C. W. Haworth, M. G. Hall

THE diffusion of copper in silver has been studied by Sawatzky and Jaumot1 and Cahoon and Youdelis.2 Sawatzky and Jaumot measured tracer diffusion in single crystal specimens but the 12.9 h half-life of CU64 limited the time of the diffusion anneal to about 50 hr, and in consequence the minimum temperature of the investigation was 716°C. Cahoon and Youdelis employed polycrystalline diffusion couples, and measured the interdiffusion coefficient over the range 0 to 2 at. pct Cu. These experiments were performed at temperatures down to 630°C and a discontinuity in the Arrhenius plot was observed at -700°C. The high D-values at lower temperatures were attributed to the effect of grain boundary diffusion. The present experiments were undertaken to investigate this anomaly and study the low temperature region of the diffusion of copper in silver. The method employed single crystal samples and a plated source of copper. Silver single crystals 2½ by 4 in. sq were grown from zone-refined silver bar (>99.99 pct Ag) by the Bridgeman technique, and $ in. thick slices were cut by spark machining. The specimens were electro-polished, and a 2 µm layer of copper was electroplated on to the planar surface at one end of each specimen using a potentiostatic voltage source. The following solution was used for plating: CuSO4 . 5H2O 125 gm H2SO4 (S.G. 1.8) 53 ml H2O 1000 ml The diffusion couples were sealed in capsules, in an argon atmosphere, and were annealed at either 701, 597, or 498°C. The temperature measurements were made to an accuracy of *l°C After the anneal, the penetration of copper into the silver was determined using an electron probe micro-analyzer. 100 sec scaler measurements of the intensity of Cu Ka! X-radiation were performed at 2 to 5 µm intervals along a line normal to the source, with a beam potential of 25 kv. Allowances were made for the background radiation, and the dead-time of the counting system in the normal way, and the intensity ratios were corrected to weight concentration by the method of Duncumb & de-Casa.3 For the analysis of the diffusion profiles, it was first observed that the copper layer did not completely dissolve in the host crystal for the specimens annealed at 498 and 591 whereas for the specimen annealed at 701°C, a single phase solid solution was formed. For the 498" and 597°C specimens, the concentration of copper in the silver at the interface was therefore constant throughout the anneal and the analysis for a

Jan 1, 1970

Jan 1, 1970