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By R Smart

This paper examines the longer-term predictive capacity of the GAMSPATH computer program (developed by the Alberta Research Council) for a waste rock (sample MS) obtained from the Kaltim Prima Coal (KPC) mine, Kalimantan that contained significant framboidal pyrite. The model was refined by comparison with the net acid generation (NAG) and column leach tests to provide a basis for calibration of the model for both short term and longer-term reaction trends. The oxidation of pyrite during the kinetic NAG test using H2O2 provides oxygen directly at the mineral surface, generating a measured oxidation rate for pyrite of 1.16 + 10-6 mol m-2 s-1, whereas the oxygen diffusion limited rate in the leach column air equilibrated reaction was of the order 10-8 mol m-2 s-1 as calculated from the GAMSPATH thermodynamic database. Results indicated that the oxygen derived from hydrogen peroxide had to be included as a reactant (eg a reacting mineral) in the simulation to account for the direct availability of oxygen at the sulfide mineral surfaces. Optical investigations identified the proportion, particle size, and surface area, of framboidal and euhedral pyrite in sample MS. The reactive surface areas of the two morphologies were very different (eg ratio: 5:1). To manage these two different pyrite morphologies the GAMSPATH thermodynamic database was updated to include framboidal pyrite, with similar thermodynamic properties and oxidation rate to euhedral pyrite but with a particle size based on the size of the individual internal framboid microcrysts. This modification, with framboidal pyrite reaction based on the size of its microcrysts produced results that correlated well with the kinetic NAG and column leach test results when the measured (kinetic NAG test rate) and calculated (column test rate) pyrite oxidation rates were employed. This provided some confidence that the model could be extrapolated to indicate longer-term geochemical reaction trends.

Jan 1, 2003

By Richard McCaffery

CONTENTS PAGE 1. Research on Blast-furnace Slags (A Progress Report). By Richard S. McCaffery 5 2. Determination of Viscosity of Iron Blast-furnace Slags. By Richard S. McCaffery, Clarence H. Lorig, Ira N. Goff, Joseph F. Oesterle and Oscar O. Fritsche 27 3. Effect of Magnesia on Slag Viscosity. By Richard S. McCaffery, Joseph F. Oesterle and Oscar O. Fritsche 55

Jan 1, 1931

By K. R. Long

Until it was defunded in 1995, the U.S. Bureau of Mines maintained a Cost Estimating System (CES) for prefeasibility-type economic evaluations of mineral deposits and estimating costs at producing and non-producing mines. This system had a significant role in mineral resource assessments to estimate costs of developing and operating known mineral deposits and predicted undiscovered deposits. For legal reasons, the U.S. Geological Survey cannot update and maintain CES. Instead, statistical tools are under development to estimate mining costs from basic properties of mineral deposits such as tonnage, grade, mineralogy, depth, strip ratio, distance from infrastructure, rock strength, and work index. The first step was to re-estimate ?Taylor?s Rule? which relates operating rate to available ore tonnage. The second step was to estimate statistical models of capital and operating costs for open pit porphyry copper mines with flotation concentrators. For a sample of 27 proposed porphyry copper projects, capital costs can be estimated from three variables: mineral processing rate, strip ratio, and distance from nearest railroad before mine construction began. Of all the variables tested, operating costs were found to be significantly correlated only with strip ratio.

Jan 1, 2011

By Paul Worsey, Gillian Worsey

The basic blasting class at Missouri S&T blast in the quarry at the experimental mine in preparation for the International Intercollegiate Mining (Mucking) Competition in April 2014.

Jan 1, 2015

By J. C. Rae

A Japanese mining engineer was given a year's leave of absence by his company and dispatched on a tour of inspection of all the principal coal mining regions of the world. He visited the New Castle region of England, Wales, Scotland, France, Germany and Alsatian mines, the underseas col¬lieries at Nova Scotia and Vancouver, and our own important fields of Illinois, Pennsylvania, West Virginia, Kentucky, Colorado, Utah and Wyoming. You are probably surprised to know that after seeing all these, he pronounced the Gebo mine of the Owl Creek Coal Company as the most interesting operation he had seen. The Gebo Mine is located about three miles west of Kirby on the Chicago, Burlington & Quincy Railroad in the Big Horn Basin region of Wyoming, 500 miles north of Denver and 200 miles south of Billings, Montana. The mine is working on a coal bed five or six to eleven or twelve feet' thick, dipping 22 to 23 degrees toward the northeast. The coal occurs in what is called the Eagle formation; and is found between two massive sandstones from 25 to 65 feet thick, which, weathering into bold', gray-colored cliffs, serves as a guide to the location of the coal bed. The deepest workings have an overburden of 800 feet. Cleavage planes of the coal extend at an angle of 45 degres to the pitch, and perpendicular to roof and floor. The coal itself is generally clean and not difficult to cut. It ignites readily from spontaneous combustion, however, which circumstance makes an ever-present fire hazard which must be carefully watched and guarded against. The coal is classed as sub-bituminous with analysis about as follows: Moisture 17% Volatile Matter 33% Fixed Carbon 46% Ash 4% Sulphur, about ½ % B. t. u. value a trifle under 11,000. It is known as Owl Creek coal, is one of the best domestic and steam fuels in the West, and is very popular in Montana, Idaho, Washington, Wyoming, South Dakota and Nebraska. While the coal slacks somewhat rapidly when exposed to the weather, it stores readily for months if placed under cover. On account of the somewhat isolated position of the mine as regards manufacturing establishments, there was a decided difficulty in disposing of slack, especially during the early years of the mine (1907 to 1914), and approximately 10 per cent of the entire product had to be wasted as slack during those years. This difficulty has been largely overcome since 1914, due partly to growth of manufacturing establishments in tributary territory, and also to enlarged transportation facilities. Also largely due to its isolated position, the mine has been compelled to use an extremely large proportion of box cars, much over 50 per cent of the product (in some years 75 per cent) being shipped in box cars, which, of course, entails extensive loading equipment and increases cost of loading. The output of the Gebo Mine for 1920 was 348,000 tons, for 1921, 273,000 tons; the rate of production during the year 1922 was over 1,700 tons daily, and 2,000 tons daily in January, 1923. The mine was developed by driving three openings straight down the pitch, the main slope being 12 to 14 feet wide and 3,000 feet long. A little distance from the surface a pair of level entries were turned right and left, known as the first East and first West. From these, rooms were driven up the pitch and chute mining attempted with only indifferent success: The failure of this rise-room method of mining was attributable to several

Jan 1, 1923

No review of underground mining should begin with anything but the new world safety record set at the Grace underground iron mine of Bethlehem Steel at Morgantown, Pa., on Friday, November 12. There had not been a disabling injury (an injury which keeps an employee away from his next scheduled shift) for 550 working days. In all, 2,660,080 man-hours had been worked by nearly 950 employees without a disabling injury. The record was wrested from Johns-Manville's Jeffrey mine at Asbestos, Que. At 3 p.m. Friday, December 3, miner Stanton E. Friend was completely buried under several feet of coarse rock, fortunately sustaining only cuts and multiple abrasions. But with his fall, the accident- free period came to an end at 2,778,600 man-hrs worked without a disabling injury-the new record!

Jan 2, 1966

By Chas. A. R. Lambly

In the extreme northeast corner of the State of Washington, on the Canadian border, lies the Metaline mining district. This district is old in history, but young in production. Geology The Metaline district is a zinc-lead area of the replacement type in dolomite and limestone. The ore bodies of the Josephine horizon are in many ways similar to the ore bodies of the famous Tri-State zinc fields. The beds are faulted and folded and have varying low dips in varying directions, and underlie large areas of the district. History Production started in 1927 on a very limited basis. The property is now mining and milling 700 tons per day. The mine is opened by adit tunnels and a vertical shaft. As the ore horizons gained depth, it was necessary to sink inclines to follow the ore horizon (see Fig 1). From 1927 to date, approximately 600,000 ft of diamond drill was put down This work indicated that suficient tonnage existed to justify a redesigning of the whole operation, surface and underground. After four years of general study, the following program was planned: 1. A new mine entrance, which would be an incline, that could follow the ore body down at whatever pitch was necessary. The incline will be equipped with conveyors for the moving of ore and waste to the surface and with tractor-type locomotives for man and supply transportation. 2. The new incline also required a new type of mining which was developed and is now in use. It is called contour mining and will be described in a future paper. 3. The new incline exit would necessitate the moving of the mill and mine shops across the Pend Oreille River. This part of the program is now underway. The Incline The sinking of the incline was to start as soon as World War II ended and was as follows: The first leg of the incline was to be sunk from the surface 1600 ft on a 17" slope. The collar and first level at elevation 2180 ft, the second level at elevation 2000 ft, the third level at elevation 1875 ft, and the fourth level at elevation 1700 feet. From the 1700 ft elevation the incline was to flatten out to 12" for 400 ft to give the necessary depth for the ore pockets below the 1700 ft level and the necessary clearance for future sinking (see Fig 1 and 2). Due to lack of manpower in 1946, the program was changed and was as follows: A drift was driven from the old mine workings on the 1700 ft elevation in an easterly direction. At 1300 ft the drift was turned N 50" E and at this point a raise was driven 180 ft on a 50" slope. This raise intersected the Josephine horizon and commercial ore was encountered. At the 2000 ft mark, a main raise was driven, 245 ft on a 50" slope, and the 1875 ft elevation was cut. Exploration drifts were started on this level and production followed on a limited basis. The main drift at the 2500 ft point was turned N 35" E and ran parallel to and 10 ft east of and under the proposed incline line. At the proposed intersection of the drift and incline on the 1700 ft elevation, it was planned to raise the incline to intersect the 245 ft raise and to continue on to the surface, a distance of 1600 ft. When this proposed intersection point was reached, a heavy flow of water, approximately 800 gpm, was encountered and all work on the main drift face was stopped. This water flow flooded the main pump station in the old mine and the two lower levels with approximately 20,000,000 gal of water. The water was controlled and finally drained from the cave areas and lower levels after six months of pumping. After the heavy flow of water was encountered in the main heading, it was decided that the incline would have to be started from the surface, as originally planned, so that too much time would not be lost. The surface overburden had to be removed, a total of 6000 yards. A temporary dry house for 6 men was built. An 8 in. churn drill hole was intersected in the first raise driven from the 1700 foot elevation tunnel. Air and water lines were placed in this hole, and air and water were delivered to the collar of the incline from the mine working. The incline started down at 15 ft wide and 7 ft high through the Leadbetter slates. After sinking 4 sets, it was

Jan 1, 1950

By Gi-Chun Han

This study focused on stabilization of heavy metals and removal of chlorine in MSWI bottom ash by artificial carbonation with washing. Artificial carbonation was carried out by applying CO2(g) injection in washing for fractions above and below 4mesh for reuse as construction materials like aggregate, As a result, carbonation with washing could satisfy Korean environmental regulation and remove chlorine in bottom ash more effectively by resolving residual friedel?s salt(3CaOAl2O3CaCl210H2O).

Jan 1, 2006

By Dennis C. Parker

The Porters Creek clay is mined for its absorbent properties. The formation is part of the Upper Mississippi embayment and is recognized in Alabama, Mississippi, Tennessee, Kentucky, Illinois, and southeast Missouri. Major minerals in order of abundance are Montmorillonite, Quartz, Illite, Kaolinite, and Chlorite. Mining consists of open pit operation utilizing track loaders or scraper. The clay is dried or hardened in rotary kilns and crushed to size by roll crushers. The clay granules are sized and air classified prior to storage for shipment or packaging. Granules are sold for pet absorbents, oil absorbents, anti-skid material, fertilizer, and pesticide carriers.

Jan 1, 1979

By Joel K. Waterland

The Homes take Mining Company at Lead, South Dakota has a yearly production of 1,868,741 tons of ore produced from horizontal cut-and-fill stopes and timber cut-and-fill stopes. This production, in addition to the necessary development (drifts and raises), requires 5,168,271 feet of drilling annually. The rock for the most part consists of quartz; mica schist; and cummingtonite schist; and must be considered as relatively hard. Penetration rates average about 15 inches per minute using an operating air pressure of 100 psi. The rock is tough, rather than brittle, and dulling occurs quite rapidly. The average regrind period is approximately 19 feet per regrind using a 1-7/16 inch carbide bit. At the present time, all drilling is accomplished with 2-5/8 inch bore airleg mounted drills, and 2-3/4 inch and 3-inch diameter bore -stopers (raising only). These machines are all direct striking, with 4-1/4 inch chucks. 7/8-inch hexagonal carbon drill steel with forged collars are used in conjunction with 1-7/16 inch multi-use, threaded, carbide bits. 62,793 rods, and 16,234 bits are used to maintain the yearly production. The number of rods noted are the rods that were sent: underground, both new and reconditioned.

Jan 1, 1963

By D. W. Gillmore

Numerous orphaned strip mines and old refuse piles scar the landscape, contributing to stream pollution and degradation of the natural environment. Field experiments to reclaim acidic surface-mined spoil and refuse banks with fly ash from bituminous coal-fired powerplants were conducted at several sites in West Virginia. Fly ash served as a neutralizing agent and soil amendment, permitting vegetation of waste lands and disposal of large quantities of ash. Reclamation of land by this method is estimated to cost about $300 per acre of leveled site, depending upon the topography of the land and quality and availability of the fly ash.

Jan 1, 1972

By Dennis H. Laubscher

INTRODUCTION The selection of mass mining methods is the situation facing not only planners of new operations, but also those involved with cur- rent operations, including the eventual change from open-pit to underground. Increasing ground control problems on many mines, plus a better understanding of dilution and ore loss problems have meant a review of current mining methods and support techniques. Experiments are only successful if conducted in the right mining environment and on the right scale. Mining personnel have been known for their conservatism and will tend to persevere with an established method rather than change; often, hidden costs will obscure the correct economic assessment. An established mining sequence is not necessarily correct in the long term. However, recommendations to change are usually met with opposition because of inconvenience, temporary production problems and the desire for a trouble-free guarantee. In any planning exercise, care must be taken that time and money are not wasted in trying to achieve the unattainable. During the decision-making period, production personnel must be involved, and, if possible, they should visit similar omrations to that proposed. Co-operative production personnel will help ensure success. Descriptions of mining methods are available in technical papers and books, and it is not intended to discuss this aspect, but, rather, the various factors which must be recognised before arriving at a final selection. In another paper presented at this Conference (Heslop and Laubscher, 1981) factors affecting ore recovery and dilution are discussed in greater detail for various cave mining methods. PRELIMINARY CONSIDERATIONS Comprehensive Geological Data. Geological information is available during the exploration stage, but, often, core from exploration drilling is not analysed with the selection of a mining method in mind. Drilling costs are high, therefore the maximum benefits must be derived from this operation. Holes which have not intersected ore may provide data on possible shaft sites, and 35 mm colour slides of core, sound logging and geomechanics classification, will provide a record which can be used with confidence at a later date. Intersections of the various rock types should be kept for future study; for example, is weathering likely to be a problem, bearing in mind that this can be accelerated in the warm, humid atmosphere of an underground mine. The accumulation of data required to select a mining method starts with initial geological work in the target area, whether this be surface mapping or a diamond drill hole. Publications. As technical papers, reports and books are used in the selection of mining methods, an appeal is made to authors to describe the geological environment so that readers can relate to their own conditions. Not enough emphasis is given to the problems experienced and the techniques employed to overcome them. Follow-up papers are seldom written and we should be prepared to discuss our failures. Diagrams are often misleading, as the ideal and not actual situations are depicted. The caved ore/waste interface is not a straight line, and whereas this may not be significant in a diagram of a 200-117 draw height, it certainly is in sub-level caving with draw zones only 1 - 2 m width but up to 20 m high. SUMMARY OF FACTORS AFFECTING THE SELECTION OF UNDERGROUND MASS MINING METHODS (a) Regional rock stresses and mining geometry. (b) The geomechanics rock mass classification of the orebody and surrounding rock mass. (c) Ratio of the surface area of the ore/unpay interface to the contained ore which determines ore losses and dilution. (d) The cavability and fragmentation data of the ore and hangingwall. (e) The mineral distribution in the orebody and its dilution zone.

Jan 1, 1981

By John Tyssowski

IN 1925, Harry Howard Stout, then metallurgist for Phelps Dodge Corporation, while investigating the cleaning of cathode copper by various gases at elevated temper-atures below the melting point of the metal, observed in the laboratory that particles of pure copper heated in a closed tube filled with hydrogen seemingly "stuck together." By further investigation he found that there was actual crystal growth across the boundaries of the particles of pure metal; in other words, there was a complete new crystal realignment and growth throughout the newly formed copper mass. It was coalescence, not adhesion; the several parti-cles had become one unit. From this dis-covery has been developed the coalescence process, whereby copper of cathode purity is squirted out in semifabricated form from an extrusion press, by the appli-cation of about 30c0 tons pressure on a 10-in. piston. Stout decided that cathode copper pro-duced by electrolytic deposition was the logical source of the pure copper required for the coalescence process, but as there was no way of reducing the ordinary tough commercial cathodes to the small particles required he was confronted by a serious problem at the outset. Under his direction, this problem was solved by William H, Osborn and Harry Howard Stout, Jr. They developed the process and technique for producing cathode copper in brittle, friable form, and to them must be given the greater part of the credit for this phase of the development program as well as for that relating to the gas cleaning of the cathode particles.

Jan 1, 1940

By I. Fafoutis

Laboratory column studies were conducted in order to assess the effectiveness of slag and red mud reactive barriers for cleaning up leachates generated at mining and waste disposal sites and loaded with copper and manganese. Solutions containing the metals of concern were prepared by dissolving the required quantities of CuSO4?5H2O and MnSO4?H2O in distilled water so that Cu concentration was 100 ppm and Mn 20 ppm. Flow rates of 25 mL.h-1 for Cu solution and 75 mL.h-1 for Mn solution were used in order to simulate aquifer water velocities of 1 ft/d and 3 ft/d respectively. Solutions were pumped upflow, in Plexiglas laboratory columns filled with each reactive medium, using peristaltic pumps. Measurements of middle and outflow column pH, redox potential, conductivity and elements concentration were performed to provide information regarding removal mechanisms for both elements and exhaustion/inactivation rates for each reactive medium. Both reactive media showed fairly good removal efficiency over the period studied. Copper was mainly removed by precipitation and formation of Cu(OH)2 while removal of manganese was due to precipitation of Mn(OH2).

Jan 1, 2006

By K. L. Moazed, G. M. Pound

An investigation was made of the deposition of silver from a thermal beam onto tungsten field-emitter tips at 300°K. "Island"-type nuclei were observed to form and grow. The nucleation of silver crystals occurs preferentially on non-close -packed planes of a tungsten tip at a critical volume free-energy change of about —2900 cal per cc. The adatom population which is critical for appreciable nucleation rate of silver crystals is about 0.2 of a monolayer and is essentially independent of beam flux, time, and source temperature. This result is strong evidence that a metastable equilibrium exists between adatoms and critical nuclei, as assumed in the present theory. SOME of the more significant studies of nucleation in deposition of metals from vapor beams have been conducted by measuring the thermal beam flux which is critical for appreciable rate of nucleation on a substrate surface held at a given temperature.'2 The observed critical supersaturations may be described in terms of Eq. [7] which was derived3'3a on the model of an adsorption, surface migration, and statistical fluctuation mechanism. However there usually made with a low-power optical microscope, a fairly insensitive method of detection. The object of the present study was to overcome some of these difficulties by measuring the critical supersaturation for condensation of a thermal beam of silver onto a tungsten field-emitter tip. Such a tip may be cleaned by electrically heating it to a high temperature in situ, and the high vacuum obtainable in a field-emission tube (10-10 torr) precludes adsorption of residual gases for convenient periods of time. Furthermore, the degree of cleanliness, the orientation, and the state of imperfection of the tip may be ascertained from the field-emission pattern. Detailed explanations of theoretical considerations and experimental procedures relating to the field-emission microscope may be found in a number of excellent review articles12'13 and thus only a brief description will be given here. The electronic work function of a metal is defined as the energy that an electron needs to escape from the potential field of the solid. By applying a large field between the surface of the solid and a suitable collector, it is possible to free the electrons by the tunnel effect. Fields of approximately 30 million v per cm are needed to produce this effect, and such fields are obtained in practice by using microscopic surfaces of high curvature (tips). Because the work function of a metal is a function of surface orientation, the emission from the various planes is different, and thus it is possible to obtain a representation of the surface in the emission microscope. In general, the work function is higher for the more densely packed planes. Fig. 1 shows a pattern for "clean" tungsten in which the central dark region represents the (110)

Jan 1, 1964

By F. O., Garrabrant

ELECTRIC power requirements for Climax are similar to those of most metal mines, except that large blocks of power are used underground and there are a number of other unusual applications. Power is furnished to the Company by the Public Service Co. of Colorado over two 100,000-volt lines to a bank of three 3333-kva. transformers 100/13.8 kv. These transformers are designed so that with the addition of forced cooling their capacity may be increased a third at 12,000-ft. altitude. On the secondary side of this bank is a 15-kv., 1200-amp. oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. (A more complete description of the electrical distribution system at Climax, with detailed drawings, is contained in T.P. 1734, A.I.M.E., Mining Technology, July, 1944.)

Jan 1, 1946

By William R. Chedsey

IN the last two years the E.C.P.D. committees having to do with the inspection of engineering schools for possible accrediting have been concerned with the engineering content of some of the mineral industries courses. Originally the discussion arose with regard to engineering geology, or geological engineering, afterward broadening to metallurgical engineering, ceramic engineering, agricultural engineering, and other branches not pertinent to this discussion. Some members of visiting committees took the attitude that they probably would have to accredit these types of training as included in the engineering group if they met the average standard set by the mining schools, though they questioned -whether they were really engineering.

Jan 1, 1939

By W. C. Hardy, B. W. McArthur

The purpose of this work is to show application of laboratory data in calculating solution gas-drive performance of the Cisco K-I reservoir. Included herein is a diagram showing the graphical relationship between six variables common to all solution gas-drive mechanisms. With an iterative process one may assume existence of any of these six variables at a particular pressure and calculate the production performance likely to occur in a reservoir produced by solution gas drive. Four unique pressure values are described at which the material balance equation alone is sufficient to predict the cumulative volumes of gas produced. Information derived from these pressure points establishes a criterion for determining compatibility of the pressure-volume-temperature (PVT) analysis of the produced crude with the gas production history. Two averaging procedures are used to arrive at a relative permeability-saturation relationship which describes the field-wide performance. INTRODUCTION The Cisco K-1 reservoir is a lenticular-type sand structure lying at approximately 6,300-ft subsurface. It is located in the H. and T. C. Railroad Co. Block 97, SW 1/4 S 392 and NW 1/4 S 385, Scurry County, Tex. Correlation of data from logs indicated that the Cisco K-1 reservoir occupied approximately 3,218 acre-ft of space with a closure of from 15 to 20 ft. Analysis of the production data indicates that the initial formation pressure and temperature were 2,450 psig and 121°F, respectively. The bubble point occurred at 1,715 psig after production of 12,256 STB oil and 11,520 Mcf gas. Only traces of water have been produced since initiation of production. Successive material balance calculations from June, 1952, to June, 1957, showed that no initial gas cap or water influx has contributed to production of oil and that the Cisco K-1 reservoir originally contained approximately 1.44 X 10" STB of oil. THEORY Methods for calculating solution gas-drive perform-

By AIME AIME

THE revised program for the San Francisco meeting, with assignment of the various papers to the several sessions is now available and is printed below : MONDAY, OCT. 7 . General Meeting, 11-12 A.M. Chairmen: Frederick W. Bradley and E. A. Hersam. Opening of the session by the Chairman of the San Francisco Section-E. A. Heraam. Address of welcome by the President of the Institute- Frederick W. Bradley. Address: The Mining problems of the West--Charles W. Merrill.

Jan 1, 1929

By Richard Borg

The partial molal free energy of Cd in each of the four intermediate phases, Cd3 Cu, Cd8 Cu5, Cd3Cu4, and CdCu2 is determined using the Knudsen vapor pressure technique. Measurements are made also within the two-phase regions, and thus the variation in FCd is now known across the entire phase diagram. The absolute values are all less the 2.0 kcal and increase regularly with increasing Cd concentration. The small absolute value of the free energy precludes the possibility of obtaining highly accurate values; nevertheless, the results are compatible with those previously reported for the liquid phase. Additional results are reported concerning the lattice parameters of the 6, E, and b phases. THE elements Cu and Cd combine to form four intermediate phases which are designated* and E and which include respectively the following inter metallic compounds: CdCu2, Cd3Cu4, Cd3 Cu5, and Cd3 Cu. The extent of each phase with respect to temperature and composition is shown in Fig. 1 which is the equilibrium diagram drawn according to Hansen.1 All four phases are brittle, and most probably are true intermetallic compounds. Only the crystal structure of the 6 phase has been thoroughly investigated.2,3 The structure is cubic and isotypic with y brass. The structure of Cd3 Cu has not been previously reported, It is hexagonal with a = 8.11 and c = 8.76 0.02Å and the structure is presently being determined.* The compounds Cd3 Cu4 and CdCu, have been each previously the subject of a single investigation. Cd3Cu, is presumed to be cubic4 with the space group Fd3m, whereas CdCu2 is hexagonal2 with the C14-type structure. Cu2 Cd is found by X-ray analysis to be hexagonat with lattice parameters a = 5.00A and c = 16.17Å. The material appears to be pure Cu2Cd as all reflections could be indexed on the basis of the single hexagonal structure. These values disagree with those of Kripyakevich and coworkers5 who report a = 4.96Å and c = 7.98Å. Their investi-

Jan 1, 1962