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By Joseph Chalmers, J. S. Desmond, R. Van A. Mills

[Because of the limited space in this volume and the fact that a description covering the oil recovery buildings and equipment (see page 4, Technical Publication No. 144) will be given in a forthcomnig U. S. Bureau of Mines publication, and because the paper was published in detail as Technical Publication No. 144, and is available as a separate at the Institute, it is impossible to include here more than the following brief extract.] Problems Being Studied The selection of the first problems to be studied was based largely upon the following very practical questions asked by several of the largest operators in the Mid-Continent field. 1. What are the causes for the formation of excessive amounts of bad thick oil on properties that are being repressured with compressed air and how can this trouble be prevented? 2. Why, in some areas, does the oxygen in the air injected into an oil sand during repressuring operations apparently change to carbon dioxide? Is carbon dioxide formed by the oxidation of some of the oil in the sand'? How can the formation of inert or low B.t.u. casinghead gas, loaded with nitrogen, carbon dioxide and oxygen, be prevented? 3. What percentage of the oil originally present in an oil sand can be recovered by the ordinary methods of operation when the gas originally dissolved in the oil is the main propulsive agent? 4. What percentage of the amount of oil produced by ordinary methods can be recovered under average conditions by repressuring a so-called depleted oil sand? 5. What are the relative propulsive efficiencies of compressed air and natural gas in repressuring operations? 6. What are the practical advantages of pressure maintenance and early repressuring as compared with late repressuring?

Jan 1, 1929

By George J. Young

The recurrence of mine-fires in Nevada during the past decade is not only a matter of interest, but also one of considerable concern to engineers and mine-managers. The more important fires may be enumerated as follows: Forman Shaft fire, Gold Hill, April 21, 1903; shaft-house, machinery, and shaft destroyed; loss estimated at $50,000; cause unknown. Union Shaft fire, Virginia City, July 14, 1904; shaft-house, machinery,and shaft in part destroyed; loss estimated at $100,-000; cause of fire stated as the careless throwing of a match in the rope-house. Sutro Tunnel fire, Virginia City, Jan. 27, 1909; 700 ft. of tunnel-timbering destroyed, and direct damage of $10,000; cause " probably electric wires." Belcher Shaft fire, Gold Hill, Aug. 9, 1910; shaft-house destroyed, machinery ruined, and upper part of shaft damaged; damage, $25,000; no cause given. Belmont Mine fire, Tonopah, Feb. 23, 1911; fire started ill winze; 17 men killed. Giroux Mine fire, Kimberly, Aug. 23, 1911; fire originated in shaft-station from unknown cause; 7 men killed, and $34,-521 damage caused. (Note 1.—On June 11, 1912, some time after the present paper was written, a fire broke out in the pump-room on the 2475 station of the Ward shaft, Virginia City. This fire was caused by the short-circuiting and exploding of the starting-switch on No. 5 pump. The oil in the switch was scattered about and set fire to the timbers and lagging of the pump-room. The pump-man in charge notified the surface, and the electric current was shut off. A hose was then turned on the fire, but the dense smoke drove the attendant out. Inspection by the electrician and the shaft-foreman, equipped with oxygen-

Jan 1, 1913

By Leo F. Reinartz

This is the fourth and final installment of Leo Reinartz' summation of the steelmaking industry over the first one hundred years. ALTHOUGH basic open-hearth steel has been in the limelight for many years, it has had to share some of its glory with two lesser but nevertheless important rivals, one 50 years old and one a rejuvenated process. Electric steel Electric-steel manufacture began in a humble way in the United States in 1906, when the first heat was tapped from a 3-ton Heroult electric furnace at the Halcomb Steel Co., Syracuse, N. Y. During the intervening years, it has completely eliminated the crucible process. For many years electric melting furnaces were located predominantly in relatively small, nonintegrated plants, making all kinds of high-grade alloy steels. Most of these furnaces, with capacities ranging from 5 to 20 tons, had basic bottoms. During and after World War II, many large basic electric furnaces, some up to 200 tons in capacity, were installed in integrated plants to make commodity steels. There were several reasons for building electric furnaces. Capital investment per annual ton of capacity was low and furnaces could be installed quickly. Since these furnaces used mostly scrap iron in the charge, no extra expensive coke- and blast-furnace capacity had to be added. The perfection of the top-charging design, and the development of large transformers and furnaces, made possible tons per hour as high as from modern 350-ton. open-hearth furnaces. Electric furnaces have high availability and repair costs are low. The process is flexible. Alloy and plain carbon steels of high quality can be made. Where power costs are low, the operating cost can

Jan 1, 1961

By J. Szekely

The paper discusses simultaneous heat conduction and diffusion. These problems may arise in slag-metal kinetics and in connection with rapid heating or cooling of specimens. A mathematical formulation is given and it is shown that the transients due to either the variability of the diffusion coefficient with temperature, or thermal diffusion, are unlikely to be important for molten or solid metals and slags. The negligibility of these effects is due to the large values of the ratio thermal diffu-sivity/molecular diffusivity. In many problems relating to transport-controlled metallurgical kinetics the diffusion process is accompanied by the transfer of heat; large-scale metal refining or metal reduction may be quoted as examples. It has been shown that due to turbulent mixing in these systems the relevant transport phe- nomena are best described in terms of "surface renewal models" which require the solution of the unsteady-state diffusion equation.1-1 Problems involving simultaneous diffusion and heat conduction are also encountered in the laboratory in connection with rapid heating or cooling of specimens. The effect of variable temperature on diffusion may be twofold: a) in the absence of "cross effects" the problem, due to variable temperature, is that of diffusion with a variable diffusion coefficient; b) if the Soret Effect is nonzero, thermal diffusion must also be taken into account in the formulation. Although information is available on thermal-diffusion transients in a steady-temperature gradient,4'5 little work has been done on the effect of unsteady temperatures on either the diffusion coefficient or on thermal diffusion.' In the following a formulation is given and solutions presented with a view of examining the areas where these effects may be important.

Jan 1, 1965

By M. Gangal, B. Paul

It may be seen from Maurer's survey (Maurer,l 1966) that most of the common methods used in rock drilling today depend upon mechanical loading of the rock. In order to predict the behavior of rock under an imposed system of stresses, it is necessary to have a valid criterion which delineates those combinations of stress which cause macroscopic rock failure from those which do not. Admittedly, the conditions under which brittle materials fracture depend very much upon statistically variable local effects such as inclusions, grain boundaries, minute imperfections, etc. However, it is extremely desirable to be able to predict fracture under complex states of stress by means of a stress-criterion of failure, where, for a given material, the effects of the other variables mentioned are incorporated into a small number of experimentally observable parameters. One of the most successful of these "macroscopic stress criteria" has been the Coulomb-Mohr theory 2,3 (also see Jaeger 4 and Terzaghi 5), modified by the addition of tension cutoffs (Paul 6). At first thought it might seem rather odd that a macroscopic theory is capable of predicting fractures which originate at highly localized random imperfections at an atomic level (or more likely at the level of individual mineral grains in rocks). However, Griffith7 showed that fractures originating near the tips of randomly oriented elliptical cracks lead to the macroscopic stress criterion of failure indicated in Fig. 1 (a). Fig. I (a) is to be interpreted in the following way: P and Q represent the maxi mum and minimum principal stresses (compression is considered positive) in a macroscopically isotropic body; fracture will not occur for any combination of principal stresses lying within the 'fracture curve' shown, but

Jan 1, 1967

By G. R. Speich, R. A. Oriani

The rate of coarsening (Ostwald ripening) of copper precipitate Particles in an a-iron matrix has been studied in three Fe-Cu alloys containing 2.3, 4.0, and 5.4 wt pct Cu in the temperature range 730° to 830°C. The kinetics of Ostwald ripening ohevs the law as deduced by Lifshitz and Wagner so that matrix diffusion controls the process. The equilibrium shape of the particles is that of rods with hemispherical caps, and is due to the BOTH Lifshitz and slyozov and wagner3 have recently solved the complex problem of determining the rate of coarsening of an assembly of precipitate particles. Both workers, although using somewhat different mathematical approaches, arrive at the same equation for the relation between the mean radius 7 of a distribution of spherical particles and the bulk solubility co, diffusivity D in the matrix, large anisotropy in the copper/a iron interfacial free energy (y,/y, = 4). A Lifshitz -Wagner equation, modified for both the nonsphericity of the diffusion field and the reformulation of the Gibbs-Thomson equation for the case of nonsphencal particles, is derived. Using the modified equation, the calculated values of the mean radius are smaller than those observed by a factor of about one -half. absolute temperature T, gas constant R, surface free energy y of the precipitate/matrix interface, and the time t of aging: where V, is the molar volume of the precipitate and v is a stoichiometric factor. Wagner's treatment is restricted to the case of spherical particles growing in a liquid matrix where strain energy is not a factor. Lifshitz and slyozov have discussed the effect of anisotropic particle shapes and internal strains on the rate of coarsening. Oriani4

Jan 1, 1965

By Howard J. Smith

At the meeting of this Institute in February 1933,I presented a paper on potash development in southeastern New Mexico1, which contained a brief review of the Geological Survey's 20-year search for commercial sources of soluble potash minerals and of the events that led to the discovery, exploration, and first development of the deposits in that state. AS a more comprehensive review of the subject was published in December2, the present paper is intended only to discuss recent developments in the domestic potash industry, particularly in New Mexico. A summary of the findings of potash in the Permian Basin, by G. R. Mansfield and W. B. Lang, geologists of the United States Geological Survey, is now in process of publication by the University of Texas and will be omitted from this paper. In New Mexico prospecting for potassium salts during the past two years has increased materially the known area of commercial deposits; development has resulted in an increase from one to three in the number of shafts and the starting of a fourth; an increase of annual productive capacity to 1,000,000 tons of crude salts, or 250,000 tons of K2O; and an increase in annual refining capacity to 140,000 tons of muriate of 62.5 per cent grade. Construction has been completed to increase still further the production of high-grade salts. In California the American Potash & Chemical Co., operating on Searles Lake brines, has increased its plant capacity from about 100,000 tons to 180,000 tons of muriate a year and is now producing soda ash and salt cake in addition to potash and borax. Potash production from cement-plant and industrial-alcohol wastes remains unimportant.

Jan 1, 1938

By Arthur F. Taggart, G. R. M. Del Giudice, M. Hassialis, A. M. Saddler

Flotation of sulphide minerals and native metals is no longer a practical difficulty. The underlying scientific principles of the method, although not explored in anything like complete detail, have been formulated with sufficient breadth so that the direction of attack on any specific case is clearly indicated. It has been known since the earliest days of flotation that certain of the nonsulphide nonmetal minerals—e.g., sulphur and graphite—would respond to flotation by oils as sulphides do; therefore they have been grouped with sulphides in many patents. But since the development of the chemical theory of collection,1,2,3 this similarity in action has been one of the facts cited in disproof of the theory by opponents of the hypothesis.4,5,6 Air levitation† of certain oxidized minerals, both metalliferous‡ and nonmetalliferous, and separation thereby from other oxidized minerals, is a recent development in practice, although the discovery of the method dates back 17 years7 This paper presents evidence that the principles underlying the air levita-tion of the nonsulphide nonmetal minerals are the same as those utilized in the flotation of the sulphides and native metals. It will be apparent immediately to anyone with knowledge of mineral concentration that separation of air-levitated minerals in water from minerals not so levitated is a simple matter of water-gravity concentration, the separating device employed depending solely upon the size of the particles treated and the buoyancy differences effected by the levita-tion. The usual separating means employed are: (1) a flotation machine for fine, slimy and for well aerated granular material; (2) shaking tables for granular less aerated material. By using large quantities of oil, it is possible to form substantially air-free oil-bound masses or granules of nonsiliceous minerals and to

Jan 1, 1939

By José Geraldo Vieira

Companhia Vale do Rio Doce (CVRD), founded in 1942, is located near the town of Itabira, Brazil. Its iron ore deposits include over 500 million metric tons (mt) of high-grade hematite and about 2 billion mt of itabirite (banded rock formed by thin alternate layers of hematite and silica). These deposits are part of the Iron Ore Quadrangle, of which the overall reserves have been estimated at more than 30 billion mt. The magnitude of the project meant difficult early years for the company, including large investments, extensive railroad rehabilitation, and construction of ore port and ore handling facilities. Companhia Vale do Rio Doce was organized as a stock company, and 91% of its shares are presently owned by the Brazilian government. Financing has been accomplished in part through loans from international monetary agencies foreign banks, and foreign industry, as well as Brazilian banks. The company markets mainly abroad. In 1967, it exported 10.6 long tons (It) of ore, principally to West Germany, Japan, Netherlands, Italy, France, United Kingdom, and the United States. The Rio Doce geographic region where CVRD operates is strategically located among the large population centers of Rio de Janeiro, Belo Horizonte, Salvador, and Vitoria (Fig. 1). It covers an area of approximately 81,000 sq km in the states of Minas Gerais and Espirito Santo, and its population in 1968 was about 3.5 million. This region presents excellent prospects for development, with factors favorable for forestry, agriculture, grazing, and exploration for minerals other than iron ore, Electric power (hydroelectric plants) is available, as well as two main systems of transportation: the Rio-Bahia highway and the Vitoria-Minas Railroad. Pioneer activities at Itabira started about 1720 as small gold mining centers. The town is about 80 km northwest of Belo Horizonte and

Jan 1, 1969

By Howard J. Smith

At the meeting of this Institute in February 1933,I presented a paper on potash development in southeastern New Mexico1, which contained a brief review of the Geological Survey's 20-year search for commercial sources of soluble potash minerals and of the events that led to the discovery, exploration, and first development of the deposits in that state. AS a more comprehensive review of the subject was published in December2, the present paper is intended only to discuss recent developments in the domestic potash industry, particularly in New Mexico. A summary of the findings of potash in the Permian Basin, by G. R. Mansfield and W. B. Lang, geologists of the United States Geological Survey, is now in process of publication by the University of Texas and will be omitted from this paper. In New Mexico prospecting for potassium salts during the past two years has increased materially the known area of commercial deposits; development has resulted in an increase from one to three in the number of shafts and the starting of a fourth; an increase of annual productive capacity to 1,000,000 tons of crude salts, or 250,000 tons of K2O; and an increase in annual refining capacity to 140,000 tons of muriate of 62.5 per cent grade. Construction has been completed to increase still further the production of high-grade salts. In California the American Potash & Chemical Co., operating on Searles Lake brines, has increased its plant capacity from about 100,000 tons to 180,000 tons of muriate a year and is now producing soda ash and salt cake in addition to potash and borax. Potash production from cement-plant and industrial-alcohol wastes remains unimportant.

Jan 1, 1938

By A. M. Saddler, G. R. M. Del Giudice, Arthur F. Taggart, M. Hassialis

Flotation of sulphide minerals and native metals is no longer a practical difficulty. The underlying scientific principles of the method, although not explored in anything like complete detail, have been formulated with sufficient breadth so that the direction of attack on any specific case is clearly indicated. It has been known since the earliest days of flotation that certain of the nonsulphide nonmetal minerals—e.g., sulphur and graphite—would respond to flotation by oils as sulphides do; therefore they have been grouped with sulphides in many patents. But since the development of the chemical theory of collection,1,2,3 this similarity in action has been one of the facts cited in disproof of the theory by opponents of the hypothesis.4,5,6 Air levitation† of certain oxidized minerals, both metalliferous‡ and nonmetalliferous, and separation thereby from other oxidized minerals, is a recent development in practice, although the discovery of the method dates back 17 years7 This paper presents evidence that the principles underlying the air levita-tion of the nonsulphide nonmetal minerals are the same as those utilized in the flotation of the sulphides and native metals. It will be apparent immediately to anyone with knowledge of mineral concentration that separation of air-levitated minerals in water from minerals not so levitated is a simple matter of water-gravity concentration, the separating device employed depending solely upon the size of the particles treated and the buoyancy differences effected by the levita-tion. The usual separating means employed are: (1) a flotation machine for fine, slimy and for well aerated granular material; (2) shaking tables for granular less aerated material. By using large quantities of oil, it is possible to form substantially air-free oil-bound masses or granules of nonsiliceous minerals and to

Jan 1, 1939

By Mark J. Klein

The precipitation of nitrogen in quenched and aged CY-35 at. pct Re was followed by internal-friction and electrical-resistivity measurements. It is shown that the precipitation kinetics of nitrogen from solution are many orders of magnitude lower in this alloy than in chromium. The reluctance of nitrogen to precipitate from sol~ction can be attributed to the composilional r-eadjustnzents required in the z.icinity of a chromium nitrade particle for gowth to occur. AS part of an investigation of the mechanism by which rhenium increases the low-temperature ductility of chromium and other Group VI-A metals, it was of interest to compare the concentration of interstitial elements retained in solution in chromium with that of a ductile Cr-Re alloy. The presence of certain interstitial elements in chromium has been shown to decrease its ductility1-5 and, of these interstitials, nitrogen is believed to have the most detrimental effect on mechanical properties. Thus, a comparison of the disposition of nitrogen in chromium with that in a ductile Cr-Re alloy might provide information about the mechanism whereby rhenium enhances the ductility of chromium. In previous internal-friction studies it was shown that the equilibrium solubilities of nitrogen in chromium and Cr-35 at. pct Re did not differ greatly.= However, results indicated that the actual concentration of nitrogen retained in solution at ambient temperatures is likely to be considerably greater in the alloy than in chromium because the precipitation kinetics of nitrogen are greatly retarded by the addition of rhenium. The objective of this investigation was to study the kinetics of nitrogen precipitation in Cr-35 at. pct Re in greater detail and to obtain information concerning the mechanism of nitrogen precipitation in this alloy. The properties studied were internal friction and

Jan 1, 1967

By R. Schuhmann, J. Th. Overbeek, P. L. De Bruyn

W. E. Ewers (Commonwealth Scientific and Industrial Research organization, Melbourne, Australia)— Any attempt to elucidate further the meaning of the contact angle, particularly if it deals with the magnitudes of those mysterious quantities rsa and YRL, is certain to receive attention from those interested in the fundamentals of flotation. It is imperative, therefore, that attention should be drawn to a serious defect in the treatment by de Bruyn, Overbeek, and Schuhmann. Their Eq. 5 dy = — ld/u — Th2odiH2n states correctly the relationship between the change in surface free energy dy, the surface excesses Tx and Th2o of the components x and H2O and the changes du, and dph2o in the chemical potentials of these components. This equation applies for constant pressure and con- stant temperature and for the interface of the solid in contact with the solution of X. The relation remains valid when 1.1 is defined so that TH2O becomes zero and the equation becomes dr = — r:dur [17] From this equation de Bruyn et al. correctly concluded that increased adsorption of the collector at the liquid solid interface would reduce the interfacial energy ysl. Their second conclusion, however, that an increase in the contact angle, corresponding to a greater decrease in rsa than in rsL can be attributed to greater adsorption of X at the solid-air interface than at the solid-liquid interface is quite unjustified. The portion of the solid surface from which the collector solution has been displaced by air must be considered in terms of the equilibria between the gaseous phase and the surface. If the pressure and temperature

Jan 1, 1955

By O. K. Riegger, L. H. Van Vlack

The median angle of relatively few two-dimensional observations provides a satisfactory value for dihedral angle determinations. Those data that contain bimodal or non-equilibriated dihedral angle distributions are readily detected. THIS paper is concerned with a statistical simplification for determining dihedral angles in solids. Microstructures are partially controlled by relative energies of interphase boundaries and grain boundaries. smith1 has shown rather conclusively that such boundaries may be equilibrated to give a vectorial balance at grain and phase edges. Thus if an accurate measurement is made of the dihedral angle associated with grain and phase edges, an accurate calculation of the relative energies may also be made. The accurate measurement is difficult because the dihedral angle involves two planes in three dimensions, and only a two-dimensional cut through the dihedral angle is usually observed. Such a two-dimensional cut is statistically random provided there is no preselection or orientation. Harker and parker2 have shown that the mode of the frequency distribution of observed angles is within 5 deg of the true angle if the true dihedral angle is 120 deg. smith1 had calculations made to show that the same generalization could apply to dihedral angles of 30, 60, 90, and 150 deg. This procedure of requiring the determination of the statistical mode is time consuming because experience has shown that more than 200 angle observations are necessary before the mode is clearly indicated. It is possible to determine the dihedral angle with equal accuracy with only 25 angle observations if the median angle is used in place of the mode angle. Use is made of Harker and parker's2 basic equation tan =2 sin x cos

Jan 1, 1961

An "international gold conference" was held at Spokane, Wash., on Sept. 5, 1918, under the auspices of the Northwest Mining Association, which was attended by a large number of engineers, mine operators, and bankers, representing the northwestern states, as well as Alaska and British Columbia. The principal subject under consideration was the doleful condition of the gold-mining industry, gripped between the diminished purchasing value of its product, and the increased cost of its operation. Two resolutions were adopted by the meeting. The first approved the resolution adopted Aug. 12 by the American Gold Conference, a permanent organization formed at Reno, Nev., in which it is stated that relief by the United States to the gold producers of this country should he made to correspond with the increased cost of production, the extent of the relief to be fixed from time to time to Meet changing conditions. The second resolution, adopted after serious and long-continued discussion, was introduced by F. A. Ross, mining engineer, of Spokane, and a member of the Institute, and was as follows: WHEREAS, it is now clearly understood and agreed by the allied nations that national gold reserves must be augmented without delay if rapidly expanding credits are to be secured; and WHEREAS, at this time, the production of gold is daily decreasing to an alarming extent because of well known conditions that, render the mining of gold-bearing ores unprofitable; and WHEREAS there is no way by which the mining of gold-bearing ores may be made profitable except by the direct' aid and influence of the government; and WHEREAS the government has appealed to the mining industry for practical information and practicable suggestions upon this subject; therefore, be it

Jan 10, 1918

By M. C. Flemings, R. Mehrabian

In a recent series of papers,1-3 it was shown that a variety of apparently different types of macrosegrega-tion result from the same basic mechanism, the inter-dendritic flow of solute rich liquid to feed solidification and thermal contractions. The important parameter governing sign and extent of segregation was shown to be (v.?T)/ where v is interdendritic flow velocity, ?T is temperature gradient, and E is rate of temperature change. It was proposed in one of the previous papers that the "negative cone of segregation" often found at the base of large ingots results from the large value of (v. VT)/E expected to pertain there. The reason why this parameter is expected to be large in that location is the high flow velocity to be expected there, as discussed earlier,' and as shown schematically in Fig. 1. In this note, two experiments are summarized which are intended as laboratory simulations of the inter-dendritic flow behavior sketched in Fig. 1 and, therefore, of the "negative cone of segregation". Experimental procedures employed are as described in the the earlier work.3 Alloy cast was nominally Al-4.5 pct Cu. Fig. 2 shows one of these ingots, a "prism" ingot, 6 cm high by 12 cm wide by 10 cm long, unidirectionally solidified from a water cooled base. In this ingot, in-terdendritic flow velocity increases near the upper portion of the ingot because of the gradually decreasing ingot cross-section. Calculations were presented earlier of the negative segregation to be expected in this type of ingot,' and results presented in Figs. 2 and 3 agree closely with those of earlier calculations. A more direct (although less easily calculable) simulation of the "negative cone of segregation" is the

Jan 1, 1970

By Stanislaus Skowronski

THE melting of cathode copper, usually containing 99.98+ per cent. Cu, would appear to be a simple matter. Owing to the well known affinity of copper for sulphur, however, so much sulphur is absorbed by the copper during the operation that a long and expensive refining process is required to remove it, and the final product is inferior in purity and lower in electrical conductivity than the original cathode copper. The furnace operation, as now conducted, includes the following stages: 1. Charging the cathodes into the furnace. 2. Melting the cathodes. 3. Oxidizing period-saturating the copper with cuprous sub-oxide and bringing the metal to the condition known as "set copper;" this is now done by blowing air into the molten bath of metal. 4. Reduction period, otherwise known as "poling," whereby the oxygen previously added is removed by covering the molten bath completely with coke or charcoal and introducing large poles of green wood; this reduces the cuprous sub-oxide, bringing the copper to the "tough pitch" stage ready for casting. 5. Casting Period. Many experiments have been made to obviate the long periods of oxidation and subsequent reduction, with no success. While the copper first, cast will be good, as indicated by the set surface of the bars, it is impossible to hold it to the "tough pitch" condition, and long before the whole charge has been poured the copper will become "overpoled," entailing numerous difficulties. Sulphur has always been blamed by the furnace men as the reason why this refining process is necessary, yet the literature on the subject does not sufficiently emphasize how small a proportion of sulphur will ruin copper for practical purposes.

Jan 3, 1918

By Reino Kurppa

The Pyhäsalmi mine of the Outokumpu Co. is located in central Finland (Fig. 1). The deposit was discovered in the autumn of 1958. Construction of the mine was started a year later, and production began on Mar. 1, 1962 (Fig. 2). At present, 750,000 metric tons (mt) of ore and a nearly equal amount of country rock are mined annually. Slightly over one-half million tons of various concentrates are produced per year, of which 85% is pyrite concentrate, 10% zinc concentrate, and 5% copper concentrate. The pyrite concentrate is treated in the company's pyrite smelter,1 yielding elementary sulfur, sulfur dioxide, and iron-oxide cinders, as well as electric power. Thus all the sulfides, totaling 75% of the ore, are of economic value. This is a fact which had a bearing on the development of the stoping method. GEOLOGY The ore is of breccia type containing fragments of country rock of varying sizes. It is coarse-grained pyrite ore containing sphalerite and chalcopyrite as a matrix between the pyrite grains. Only small amounts of pyrrhotite have been observed in certain marginal parts of the orebody. The ore averages 0.85% Cu, 3.0% Zn, 38% S, and 33% Fe. The orebody extends from the surface down to at least 500 m, which is the lowest level from which present information has been obtained. Horizontally, the orebody is 650 m long. It is 75 m wide at its central part, narrowing towards the ends (Fig. 3). At a depth of 115 m, the orebody suddenly narrows down to a thickness of 25 to 30 m. In the northern part of the orebody, the dip is 50° to 70°, but elsewhere it is almost vertical (Fig. 3).

Jan 1, 1969

By W. E. Mitchell

ELECTROLYTIC, production of cadmium at the Great Falls plant started in the first part of the year 1925. Prior to that time, an experi¬mental unit had been in operation for a few months during the year . 1922. At present the plant is the largest producer of cadmium, not only in the United States but in the world. Electrolytic cadmium is characterized by its very low content of impurities. The main impurities, which are lead, copper, zinc, and iron, do not aggregate more than 0.05 per cent., leaving a cadmium content better than 99.95 per cent. The consumption of cadmium has increased rapidly since the war, with the growing use in the plating industry. Cadmium plating has come into prominence because of its connection with the automobile and radio industries. As a protective coating, a deposit of, cadmium is superior to zinc. Cadmium is also used extensively in alloys. Important among these is the alloy with copper used in telephone and trolley wire. In proportion of 0.5 to 1.2 per cent., cadmium adds materially to the strength and wearing qualities. Practically all commercial zinc concentrates contain a small amount of cadmium. In the electrolytic zinc process, the cadmium that is dissolved in the regular leaching operation is removed from solution with metallic zinc, generally added in the form of zinc dust. This precipitate also contains the copper taken into solution and the excess of zinc dust, which is required in order to insure a complete removal of all impurities. The precipitate is treated by a separate process to recover the copper, cadmium and excess zinc. In order to make the metals more soluble and to render insoluble such impurities as iron, arsenic and antimony, the precipitate is roasted. The roasting is carried out in gas-fired McDougall furnaces, with a roasting temperature of about 700° C.

Jan 1, 1930

By J. E. Powers, T. D. Taylor, N. E. Wood

A computer program has been developed to simulate a gas transmission 1ine operating under transient conditions. The program was developed by application of the basic mass and momentum balance equations. The nonlinear equations thus obtained were transformed by the method of characteristics and finite-difference techniques to make them suitable for machine computation. Three tests of the computer program were made. In each of these, the initial condition was assumed to be steady-state flow, the inlet density was held constant at the initial condition, and the outlet mass flow rate was varied with time. One test was theoretical in nature, but the others were made for direct comparison with actual field data. Results of the tests are presented in graphical form. INTRODUCTION Theoretical investigations of the transient flow of gas in long pipelines have been limited in number. In 1951, Olds and Sage4 developed a gaphical method of solving the nonlinear flow equations describing transient gas flow in a long pipeline. Research in this field at the U. of Oklahoma was initiated by Nelson1 in 1956. Nelson utilized a computer to solve numerical equations involving a lengthy trial-and-error procedure, and his results have been reported in the literature? It does not seem likely that Nelson's method can be extended to the analysis of transient behavior of pipeline networks; therefore, the investigation to be discussed in this paper was undertaken to improve and extend the previous work. The following discussion presents, briefly: (1) the basic differential equations governing transient flow, (2) transformation of these equations into a form suitable for machine computation. (3) role of boundary conditions in the calculations, (4) a very brief description of the computer program which was developed to simulate the pipeline and (5) results of tests of the computer program.