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By E. F. Kingsbury

One of the well-known and important uses of the noble or precious metals has been for electrical contacts. In fact, the elements of this group, comprising gold, silver and the six platinum metals, have been called "noble" because of their ability to resist heat and corrosion, properties which were early found to be of importance in good contacts. Of these, platinum and its iridium alloys, palladium, and silver, have been extensively used. In recent years the electrical industry has been responsible for about one-eighth of the total consiimption of the platinum metals in the United States. Information is not available as to what part of this one-eighth has been used for contacts but it is certainly an important one. Substitutes for Platinum During the past 20 years a considerable number of contact alloys have been proposed in the patent literature largely with the aim of more or less replacing platinum by gold or silver, and some of these alloys are finding increasingly widespread application as the reliability of their performance becomes established. From even the partial data at hand it appears that this substitution by the less precious metals has replaced at least one-half of the amount of platinum that the entire electrical industry would otherwise be consuming at the present time. While the subject is of considerable interest to the metallurgist it is even more so to the electrical engineer and it is surprising that the literature records no systematic and comprehensive investigation along these lines. Practically all the scattered work on noble metal contacts has centered around platinum as a standard and has had two more or less distinct aims. One of these has been the replacement of platinum by cheaper matcrials, and the other, the dcveloprnent of materials of superior electrical performance. The merit of a contact metal has been assumed to be dependent on its infusibility, and since thc more precious metals have the higher melting points there has naturally been an association of the two qualities in the sense that the more precious were

By Michael A. Gustafson, Fernando B. Sotelino

INTRODUCTION The degree of success of any new project will ultimately depend on two factors: (i) the underlying economic strength of the project; and (ii) how successfully the parties involved can coordinate their interests to establish a financing structure which indeed capitalizes on the project's inherent economic attractiveness. A project's underlying economic attractiveness is determined by its inherent comparative advantage in accessing the factors of production, combining these factors of production to create a product and marketing this product. Initially, the decision regarding whether to go ahead with the project - the go/no go decision - has to be based on its total ability to generate cash flow returns to project participants to compensate them for resources committed and risks undertaken. The extent to which project sponsors will actually benefit from the project's inherent economic attractiveness is fundamentally dependent on their ability to establish the long-term financing arrangement which best meets the objectives and constraints of each project participant and sources the most advantageous funding. The decision regarding the financing of the project, the "how-to-go" decision, is then an attempt to find the capital structure, consistent with market terms and conditions and sound financial management policies which will best: (i) match planned outflows with cash inflows; and (ii) maximize the return for project sponsors. This article makes use of a case study to illustrate the analytical process involved in dealing with the go/no go and the how-to-go decisions on a project. Though most commonly applied to project finance situations, the guideline described herein is equally applicable to other financing situations, such as leveraged buyouts and debt restructurings, where timely repayment of debt depends largely on the future cash flows to be generated by the entity being financed. Our case study is of a hypothetical expansion of an existing copper mining operation in Latin America. The company, Latin American Copper (LAC) currently has 147 million metric tons of proven sulphide ore reserves. The existing mine complex consists of an open pit mining operation and a conventional flotation plant with a processing capacity of 20,000 metric tons of ore per day. LAC sells copper concentrates primarily to Japanese custom smelters. The mine has been in operation for 5 years, fine copper production last year was 108,240 metric tons, and LAC currently has assets of $175 million and net worth of $135 million. Due to a recent increase in proven reserves, LAC is considering doubling the treatment capacity of its flotation plant to 40,000 tons per day, and expanding its mining capacity to meet the additional concentrator capacity (the "Project"). Project capital expenditures and incremental working capital have been estimated at approximately $210 million. The construction period is expected to cover one year, including a three-month start-up phase. The Project is to be funded by a combination of equity, senior debt, including both supplier credits and/or commercial bank term loans, and subordinated debt. LAC's shareholders are willing to commit a maximum of $50 million to the Project. A government development agency, which is also a minority shareholder, might be willing to lend approximately $40 million to LAC in the form of subordinated notes, depending on the ability of the Project to generate foreign exchange reserves. Commercial banks cannot be expected to offer loans with maturities longer than ten years. The Project is expected to increase the value of the firm because: (i) It will solidify the Company's position as a low-cost copper producer by taking advantage of economies of scale. As a result of the expansion, cash production costs are expected to decrease 7.0[c]/lb. and all-in costs 11.0[c]/lb; (ii) The potential earnings stream from the sulphide ore reserves will be realized

Jan 1, 1985

By E. J. Schmitz

The following abstract of an unpublished treatise, prepared by me, on the geological formations and minerals of the State of Alabama, is deemed of interest to the Institute. This State, called after the river of the same name (which, formed by the junction of the Coosa and the Tallapoosa, waters, with its tributaries, more than half the State), is divided by its main watershed into a larger southern district, sloping south and southwest, and drained into the Gulf of Mexico, and a smaller northern district, sloping generally north and west, and drained into the Tennessee. Geologically the State may be divided (according to historical geology) according to the different formations it exhibits, among which I include the Azoic, the existence of which in Alabama has been doubted by some; or it may be divided (according to structural geology, as closely related to topography) into a north and northwestern, a middle, and a southern zone.

Jan 1, 1884

By A. K. Sinha

The market for thermal coal has recently been growing rapidly and it is projected to continue to do so in the foreseeable future. Both the national and international markets often demand high-Btu low-sulfur thermal coal in order to minimize the cost of utilization and meet the environmental restrictions. High quality thermal coal can be prepared to compete in the utility market by desulfurizing the high-Btu midwestern coal and blending it with low-sulfur western coal which is usually low in Btu content. The methods that can be used for optimizing the coal preparation plant operating parameters for physically desulfurizing coal are presented.

Jan 1, 1985

By Nicholas J. Grant, Italo S. Servi

THE creep and stress rupture properties of three grades of aluminum have been reported in a previous paper.' It was found that the stress coefficient of the creep rate and of the rupture time changes at the transition from low temperature-type to high temperature-type deformation. The transition corresponds to the equicohesion for grains and grain boundaries. In order to support the conclusions obtained from the examination of creep and stress rupture data, the deformed specimens were observed metallographically and a few special tests were conducted. The results of these qualitative observations are given in the present paper. Experimental Procedure and Results Table I gives the types of aluminum specimens.' The creep and stress rupture tests were conducted in simple tension at constant stress. Further details on the experimental techniques have been given.' The test bars were electropolished in a perchloric-acetic mixture before testing and were observed after testing without any further preparation unless otherwise specified. Observations of specimen cross-sections were made after electrolytic polishing and etching. The samples were extended along a direction parallel to the longer side of the micrographs of Figs. 12 through 27. The log-log plots of stress vs. minimum creep rate for the three grades of aluminum appear in Figs. 1 and 2. The reference numbers on each curve help relate the photographs to the conditions of testing, indicating where high temperature-type deformation and failure prevailed and where low temperature behavior prevailed. Deformation of Single Crystals at Very High Temperature: An annealed, high purity aluminum specimen was slightly strained at one end and re-annealed. A large single crystal was formed at the strained end of the specimen during re-annealing; the other end remained polycrystalline (see Fig. 3). A stress of 50 psi was applied to the cold specimen to avoid loading at 1100°F. The temperature was then raised to 1100°F, kept at 1100°F for 3 hr, and lowered again to room temperature. Fig. 4 shows the appearance of the specimen after testing. The single crystal deformed by gliding along parallel planes, while the polycrystalline part deformed mainly by grain boundary flow. Some grain growth occurred in the polycrystalline part. Another high purity specimen, which had two large single crystals at the two ends and a polycrystalline zone in the middle portion, was tested at 1000 °F under a stress of 100 psi. The appearance of the sample after test is shown in Fig. 5. Even though considerable slip occurred in the single crystals, the fracture occurred in the weaker polycrystalline portion. Evidence of Grain Boundary Migration during Creep Testing: A high purity, coarse grained specimen, which had a large crystal in the middle portion (see Fig. 6) was tested at 700°F under a constant stress of 175 psi for 30 hr. The appearance of the specimen after testing (Fig. 7) indicates that the large crystal deformed by gliding along parallel planes, while the rest of the specimen deformed mainly by grain boundary flow. Microscopic examination showed clear evidence of grain boundary flow and grain boundary migration, an example of

Jan 1, 1952

By Norman Weiss

Most papers on xanthate have dealt with principles rather than practice. On the assumption that many millmen are interested in knowing where and in what manner the xanthates are being used in mills other than their own, this paper will summarize the results of a recent survey of North American plants. A brief background story and a description of the xanthates will be followed with the results of that survey; then a few observations will be made on storage, feeding, and ore testing. The first xanthate was made in 1822 by Zeise of the University of Copenhagen. With commendable scientific detachment (or perhaps a bad cold) Zeise ignored a more pervading quality of the substance and named it for its color, after the Greek word "xanthos," meaning yellow. During the first 80 years of its recorded existence xanthate found use only in small quantity as an insecticide; after the turn of the century there was some further demand for it in the vulcanization of rubber. It was not until 1922, by coincidence the centennial of Zeise's discovery, that C. H. Keller of San Francisco, seeking sul-phidizing agents, recalled xanthate in connection with his experience years earlier when he had seen this material used in Europe to combat grape phylloxera. His successful tests with the new flotation reagent culminated in the basic xanthate patent (U. S. 1,554,216) in 1925. In the minds of many metallurgists the introduction of xanthate as a flotation reagent is identified with the beginnings of alkaline circuits, the first use of chemical collectors, and the beginning of selective flotation. Actually, these preceded xanthate by a number of years. The alkaline circuit in lead-zinc flotation was introduced by Lyster as early as 1912. In 1922 Sheridan and Griswold made their momentous discovery (Timber Butte) that alkali cyanides and zinc sulphate effectively inhibited the flotation of pyrite and sphalerite in pulps made alkaline with sodium carbonate or bicarbonate. The use of lime to retard pyrite in copper flotation began within two years after Ralston directed attention to its possibilities in 1917. These were the beginnings of modern selective flotation, which was practiced successfully on copper-iron ores at Nacozari in 1922, and at Inspiration, Cananea, and others only shortly afterward. In the lead-zinc mills, selective flotation was of even greater economic importance at that time. The Sunnyside plant of U. S. Smelting, Refining, and Mining Co. used the process earlier than 1920, and Consolidated, Timber Butte, and others soon followed. As a chemical collector xanthate was relatively a latecomer, for the Corliss patent of 1917 covered the first of these, alpha-naphthylamine, and the Perkins patent of 1919 the second, thiocarbanilid. Both of these were in extensive use before Keller's xanthate made its commercial debut in 1923. There is general agreement, nonetheless, that the introduction of xanthate was

Jan 1, 1949

By Joseph A. Holmes

Iv all new movements unavoidable misapprehensions arise, which should be cleared away, lest they retard the progress of the movement itself. An impression has gone abroad that the movement for conservation of resources is a new movement. No one knows better than the engineer that this is not the case. Men like Douglas and Raymond and John Fritz and Edison and McDonald, and hosts of others, have preached and practiced conservation as applied to mineral resources throughout a long and useful life-time; and, in general, the work of the engineer and the chemist, both in this and in other countries, has been along the line of continually greater efficiency in the use of all available materials. And the results already accomplished through the efforts of these experts are now saving annually many millions of dollars to the people of the United States in materials and labor. It is true, however, that the need for conservation, that is, for a wiser and more efficient use of the natural resources, has taken on a new life, and is being brought home to the people of the United States in a manner which is awakening public interest and even the public conscience. It is, therefore, both appropriate and fortunate that the engineers, the chemists, and the thinking men of all classes are availing themselves of the present impetus to the movement to emphasize anew the lessons which heretofore they have found it more difficult to impress upon the American public or even upon the captains of industry. Still another of these misapprehensions is to the effect that many of the leaders in the present movement hope to bring about conservation of resources through legislation or execu-

May 1, 1909

By Carlton D. Hulin

ARE we a "have" or a "have-not" nation in our domestic supply of metals and minerals? Impinging on the ears of a people weary of war and faced with the problems of reconversion to peace, the import of this question has built up with the momentum of an avalanche. Stemming from statistics cited by Pehrson and accelerated by the pessimistic and possibly biased statements of Ickes, hardly an issue of any technical periodical dealing with mining has appeared during the past year but within whose pages was to be found some comment, some discussion of the question: Are we a "have" or a "have-not" nation? We have proved commercial reserves of high-grade iron ore in the Lake Superior region sufficient for 22 years, based upon the prewar rate of consumption; commercial copper reserves to last only 34 years, one fourth being inferred ore; zinc sufficient for nineteen years, over half being inferred; lead for twelve years, two thirds of which is inferred; mercury for only three years. So read the figures. Although admitting that these figures may be extended by new discoveries, by the use of marginal and submarginal ore, and by further technological advances, are we nevertheless to assume complete pessimism in the acceptance of the meaning of these statistics; are we to expect a rapidly increasing dependence upon foreign sources of mineral supply; are we to believe that we are at or near the peak of the "metal age" as has been held by Theodore J. Hoover? Or is some more optimistic viewpoint possible?

Jan 1, 1947

By B. Ladanyi

The problem of rock behavior under concentrated loadings has been the subject of a number of studies in the past and has been receiving increasing attention by rock mechanics investigators in recent years. Its application covers various fields, but main interest in its solution has arisen in connection with the problems of drilling, bearing capacity, and rock testing. These widely different purposes and fields of interest probably have been one of the reasons why, in spite of the large amount of published experimental data and several theoretical studies, there are still many points which need further clarification. Another reason for this situation undoubtedly is due to the fact that the rock behavior in indentation is complicated by a great number of influencing factors. Experimental evidence shows that, depending on the rock type (mechanical properties, porosity), environment (ambient effective pressure, pore fluid, temperature), and the type of loading (indenter shape, initial embedment, load level, loading rate), the rock response to indentation may vary within very large limits. The widely different rock behavior has made it impossible up to now to treat the problem with a consistent single mathematical theory, but has required instead a series of simplified models covering only certain most important events in the failure sequence. As far as the theory of rock indentation is concerned, this chapter does not add much new information. Its main purpose is, however, to state the indentation problem in a general manner and to indicate which points need to be studied in more detail before its definite solution is attempted. Also, since excellent reviews of the state-of-art of rock indentation in drilling have recently been given,1-4 this will not be repeated here, and only references pertinent to the points of the problem under study will be quoted. The theoretical part of this study attempts to give a quantitative evalu-

Jan 1, 1972

By C. H. Lieb

ONE astounding fact in the production of petroleum is the comparatively recent realization by producers that flowing production is the cheapest crude produced. About 1910 or even later, operators actually endeavored to get a producing well to pumping as quickly as possible after completion. A well was not really considered to be completed until it was operating on the pump. Remarkable progress has been made since then. It was not until about 1920 that producers began to take a serious interest in pressures and to think a little about proper methods of handling oil wells. Pressure gauges were installed, and records of pressures maintained, which was quite an innovation. Desultory experiments were made with tubing of different sizes, as well as experiments with placing the bottom of the tubing string at different levels in the well. Probably the first important departure from customary procedure was about 1926, during the Seminole development, when gas lift

Jan 1, 1936

By J. W. Evans

ONE of the most important and frequent calculations that the extractive metallurgist is called upon to make is that of the standard free energy change of a reaction (?F°). For many reactions of metallurgical interest, the lengthy arithmetical calculations have been eliminated by the use of free energy-temperature diagrams which have been constructed by Ellingham,1 Richardson and Jeffes,2 Kellogg,3 and Osborn4 (e.g., for oxides, sulphides, halides) from critical surveys of existing data. The speed and simplicity of making calculations with these graphs make them of great practical value and, this should be stressed, in the majority of cases no greater accuracy is obtained by the individual calculations from specific heat data, entropies, and heats of reaction. With the increasing industrial application of vacuum methods to the winning of metals, notably at present magnesium and the alkaline earth metals, it is necessary to calculate low metal vapor pressures above reaction mixtures. The existing free energy diagrams usually give values where the solid and liquid metals are at unit activity. To obtain the free energy change associated with the reaction Metal (solid or liquid) = Metal (gas) it is customary to calculate this from either vapor pressure equations or from the appropriate free energy functions. This paper presents graphically the free energies of vaporization of a number of metals of economic importance which can be used in conjunction with other free energy diagrams. Certain metals have been excluded on the following grounds: titanium, tungsten, molybdenum, and platinum have very high heats of vaporization and boiling points and the uncertainty of these values arising from difficulties of high temperature measurements makes free energy values of doubtful accuracy; the data for cobalt, vanadium, and zirconium are uncertain; finally, arsenic, antimony, and bismuth could not be conveniently included. because of the presence of mixtures of polyatomic and mona-tomic vapors in the temperature range considered. Free Energy Diagram In Fig. 1, the standard free energies of vaporization per mol (?F°T) have been plotted for the temperature range 0" to 2000°C for the reaction M (solid or liquid) = M (gas) The data have been obtained from the free energy functions — (F-H298)/T for the solid, liquid, and gaseous elements and the heats of vaporization at 298°K, i.e., ?H298 (vapor). The values of these functions have been taken without change from the recent critical survey by Brewer;' the bulk of this data is from Kelley7 with adjustments in the light of more recent work. Values of the function — (F-H298) /T at T = 298°, 500°, 1000°, 1500°, and 2000°K were employed. Where possible, a further independent value was obtained by using the boiling point, i.e., where ?F°T = 0. All calculations are based on the assumption that the metal vapors are monatomic and that the boiling points refer to those temperatures at which the partial pressure of the monatomic vapor above the metal is 760 mm. For the alkali metals, e.g., sodium, potassium, and lithium, diatomic gas molecules are present as minor species; details of these are discussed by Brewer." On the basis of the above values and allowing for the change in slope at the melting point, straight line graphs were constructed and extrapolated to 2000°C. The deviation from a straight line did not exceed ±500 cal, which in all cases appears less than the probable error in the ?F°T values.

Jan 1, 1954

By F. L. Wolf

THE tests here described were made to obtain information regarding costs, efficiency, etc. of baking cores in an oil-fired oven, and two electric ovens, which were installed, early in 1920, in the core room of The Ohio Brass Co. and were used almost continuously during the rest of that year. Operating costs were rather high so a complete test was undertaken early in 1921. The ovens will be referred to as Nos. 1, 2, and 3. Nos. 1 and 2 are electric ovens of different makes while No. 3 is an oil-fired oven made by the firm that made No. 2. The three are equipped with similar ventilating devices. They have full-height doors at front and rear. The cores are placed on iron racks, which are lifted by Cowan trucks and rolled into the ovens through the front doors. When baked, they are removed through the rear doors. Fig. 1 shows the general plan of the electric ovens while Fig. 2 shows the plan of the oil-fired oven. Oven No. 1 is 91 in. (231 cm.) wide, 83 in. (210 cm.) high, and 79 in. (200 cm.) deep, inside measurements. The walls are constructed in, panels about 18 in. (46 cm.) wide and 4 3/16 in. (10.6 cm.) thick, packed with blocks of insulating material and covered with sheet iron. The concrete floor on which the ovens set is 13 in. (33 cm:) thick. A recess about 5 in. (13 cm.) deep was cut in the floor to provide room for the heating units. Channels and plates supported. on I beams serve as' supports for the trucks and racks. The ventilating system consists of a Sirroco No. 1 ½ fan, connected so as to draw air from the top of the-oven or from the room, through damper C, and to deliver it either to the bottom of the oven or to the floor; the delivery is controlled by the vane A. The fan and pipes are covered with ¼ n. (6.3 mm.) asbestos cement lagging.

Jan 4, 1922

By Hans W. Schreiber, David W. Neuhaus

INTRODUCTION The Sunshine Mining Company's 16-to-1 silver deposit and mine project are located 384 kilometers (240 miles) southeast of Reno and 360 kilometers (225 miles northwest of Las Vegas, near Silver Peak in Esmeralda County, Nevada, at an elevation of about 2,300 meters (7,000 feet). The deposit was being developed, in 1980, with an anticipated mining and milling rate of 150,000 tons per year (500 tons per day). Mining is by underground, blast-hole open stoping with rubber-tire haul-age out of the mine to a straight cyanide plant about 5.5 kilometers (3.5 miles) from the mine. The product is a gold-silver dore which may be further refined at Sunshine's new refinery in Kellogg, Idaho. The cash required for investment was obtained through the sale of $25 million worth of 6 l/2h silver-indexed bonds. Some $15.0 million were directed toward development of the 16-to-1 Mine. These bonds, which were underwritten by Drexel Burnham Lambert of Chicago in April of 1980, are redeemable on or after April 15, 1985 or payable at maturity on April 15, 1995 at the greater of either $1,000 or the market price of 50 ounces of silver. Behre Dolbear acknowledges the kindness of the Sunshine Mining Company (Sunshine) for allowing the presentation of the subject material for this Case Study. APPLICABILITY AS CASE STUDY Reasons for Choosing 16-to-1 The reasons for choosing the 16-to-1 Mine as a case study follow: 1. The project is simple and compact. 2. The mine is small -- thus relevant to other contemplated precious metal operaions in the western states. 3. The areas of risk are readily identifiable. 4. A change to project financing was under consideration. 5. A substantial decline in the price of silver was experienced during development. 6. Behre Dolbear undertook one of the first feasibility studies of The 16-to-1 and thus has some familiarity with the project. Historical Summary The following history serves as a setting: 1. Late 1920s -- minor "production"(?) via shallow pitting. 2. Mid-1930s -- neighboring Nivlok Mine successfully put into production. 3. Late 1950s (?) -- Callahan Mining Company acquired rights and drilled -- results inconclusive; Callahan turned property back to Arthur Baker, who had originally interested Callahan in 16-to-1 on the basis of its similarity with Nivlok Mine.

Jan 1, 1985

By T. O. King, Y. P. Gupta

The self-diffusion of sodium in two sodium silicate liquids was measured in the temperature range 850" to 1500°C by the capillary-reservoir technique. Radioactive Na 22 was used as the tracer. The total count and autoradiographic methods were used for determination of the total Na 22 depletion and diffusion profiles of the dqy-used specimens. The diffusion coefficients obtained by the autoradiographic technique are slightly smaller than those obtained by a total count method. Error analysis of the two methods suggests that more confidence be placed in the results of the total count method. The experimental results were analyzed in terms of the existing activated rate process theory. The activation energy for diffusion was shown to decrease markedly as the temperature increased. This was attributed to a variation in the heat of activation with temperature, probably related to a change in the distribution of anions associated wilh the cation. In terms of a model, suggested for diffusion in liquid silicates, differences in the aclivation energies for diffusion and electrical conduction may arise from the effect of the electric field applied in conduction measurements. It is generally accepted that liquid silicates consist of cations and an equilibrium distribution of complex anions determined by temperature and composition. Transport properties in molten silicates are of interest, not merely because of their relevance to the kinetics of pyrometallurgical reactions and glassmak-ing processes, but also because they are useful to development of the theory of such ionic liquids. Self-diffusion is one such transport process that may, if studied for liquid silicates of widely varying compositions, indicate structural changes through changes in the activation energy for diffusion. The self-diffusion of calcium, silicon,' sulfur: aluminum, 4 and oxygenS in lime-alumina-silica melts and of iron6 in molten iron silicates have been measured. Unfortunately, the errors in some of the reported activation energies for diffusion are too large to allow firm conclusions concerning structure and the mechanism of the diffusion process to be drawn. The research reported here was a study of the self-diffusion of sodium in liquid sodium silicates. The soda-silica system was chosen since: i) a reasonable composition range can be covered in the binary system at moderate temperatures; ii) suitable isotopes of sodium (radioactive Na 22 and NaZ4) can be obtained; iii) data on the electrical conductivity and viscosity of sodium silicate liquids are available. However, the range of composition actually used in diffusion experiments was limited, to 20 to 35 wt pct soda, by the relatively high viscosity of silica-rich compositions and by evaporation of soda from basic melts. EXPERIMENTAL PROCEDURE The capillary-reservoir technique was used, wherein a radioactive tracer, Na 22, incorporated in the capillary melt, was allowed to diffuse out of the platinum capillary tube into a large reservoir of silicate liquid containing a chemically identical melt, but without radioactive tracer. After a specific diffusion time, both the total depletion in the tracer concentration and concentration-distance profiles of the diffused samples were measured by procedures to be described later. The diffusion cell assembly was similar to that used by Koros and King.' The temperature of the diffusion cell was measured with a calibrated Pt, Pt-10 pct Rh thermocouple located at the center of the liquid reservoir. The same thermocouple was used to obtain the temperature profile in the reservoir. The temperature at the top of the reservoir was slightly higher (2°C) than that at the bottom, to minimize convection in the capillary tubes, which were placed, open end up, in the reservoir. Convection was not expected to be a problem since the length-to-diameter ratio of the tubes was about 12:1 and the maximum capillary diameter was 1 mm. The diffusion cell was heated in a molybdenum-wire resistance furnace, previously used by Koros and King, but somewhat modified. A Pt, Pt-10 pct Rh thermocouple enclosed in an alumina tube and located near the furnace winding was used in conjunction with a preamplifier and Micromax proportional controller for temperature control, within 4°C, at temperatures near 1500°C. Before starting a diffusion run, the furnace was heated to the desired temperature and an alumina guide assembly for the capillaries was slowly lowered to within 2 cm of the liquid reservoir. The capillaries were thus heated to about the temperature of the bath. The run was started by lowering the assembly till the open ends of the capillaries were about 1 cm below the surface of the liquid. The run was ended by raising the assembly to about 15 cm above the liquid bath. Later, the sample assembly was slowly withdrawn from the furnace. Six diffusion samples were usually run at the same time. Three runs made with the open ends of the capillaries down were discarded because in the majority of these samples air bubbles trapped at the open end were observed. Some glass adhered to the outside of the capillary on withdrawal. This was carefully removed and used to check that the Na 22 activity in the sink remained at a low level. Materials. Nonradioactive sodium silicates were prepared by melting, in a platinum crucible, weighed quantities of sodium carbonate and powdered silica. To prepare radioactive sodium silicates, Na2 obtained in HC1 solution, was first checked for radioactive impurities by obtaining a y-ray energy spectrum, then the chloride ions were removed by anion-ex-

Jan 1, 1968

By E. T. Turkdogan, P. Grieveson

It is shown experimentally that, by making use of the coupled S-O reaction in ionic melts, it is possible to transfer slclJur or oxygen from a lozv to a high chemical potential through an ionic nzenzbrane. This uphill transfer of sulfur is accompanied by a counter downhill transfer of oxygen and vice versa. The Materials balance shows that the amoulits of sul -fur and oxygen lost or gained by one system are the same as those gained or lost by the other systenz — the two systejns being separated by an iofzic membmne. The systems used in these experi?nents were solid Mn-MnS-Mno and liquid Fe-S-0. Sufficient results were obtained Lo study the kinetics of the transfer and the indications are that the counter diffusion of sulfide and oxide iojzs in the membrane is the rate controlling process. WHEN an ionic oxide-melt reacts with an atmosphere containing sulfur and oxygen-bear ing gases the following reaction occurs: 1/2 S2 (gas) +O2- (melt) = S2- (melt) + 1/2 0, (gas) [1] Under the restrictions that i) electric conductance in the melt is purely ionic and ii) each constituent element in the melt does not change valence, the solution of an atom of sulfur in the melt will result in the evolution of an atom of oxygen, so that the electric neutrality is maintained. For a given temperature and melt composition, the equilibrium concentration of sulfur in the melt is determined by the ratio, po2 /ps2 , in the gas phase and not by the individual partial pressures of oxygen and sulfur. This has been demonstrated by a number of investigators, for example by Richardson and withers', Fincham and Richardson,la by St Pierre and Chipman,' and by Turkdogan and Darken.3 In his Howe Memorial Lecture for 1961, Darken discussed the possibility of uphill pumping of sulfur through an ionic membrane at the expense of oxygen. TO test the validity of this concept associated with coupled reactions, experiments were carried out in the following manner. A presintered mixture of Mn-MnS-MnO and prefused mixture of Fe-S-0, contained in platinum and iridium crucibles, respectively, and separated by a thin layer of soda-glass (to act as an ionic membrane), were placed side by side in an evacuated silica capsule and sealed. This was then kept at 1400°K for a required time to allow Reaction [11 to occur via the ionic membrane. When partial equilibrium is reached, the ratios Po2 /pS2, for systems on either side of the membrane, are equal and the flux of sulfur and oxygen through the membrane becomes zero, although the individual partial pressures po2 andp,, in these two systems may differ considerably.

Jan 1, 1962

By Willard C. Lacy

Survival of the mining industry as a viable economic entity in the United States is being seriously threatened by declining grades of ore reserves, rising operational and capital costs, and increased governmental taxes, restrictions, and subsidizing of third world competition. Discovery of major high-grade deposits is becoming increasingly unlikely. International competitiveness of the U.S. deposits will depend upon 'Yankee Ingenuity, ' innovations in mining methods, metallurgical treatment, and financial arrangements, and a viable national strategy. This will require not only effective management for control of operations, but management that encourages breakthroughs in technology, financing, labor relations, and in public and governmental understanding. The mine geologist and the exploration geologist both have important roles to play in all aspects of this fight for survival of the industry.

Jan 1, 1985

By Royal Preston Jarvis

minerals of different specific gravities. It is simple in construction, easily operated, capable of treating large quantities in a short time, and highly efficient under various conditions. The question, whether the material to be jigged has first been sized, determines the two principal methods of jigging. Jigging preceded by close sizing, generally known as the Continental or German system, involves a more or less elaborate series of screens or trommels, with attendant cost for installation, operation, and repairs. Jigging without sizing, known as the English system, is, according to Munroe,1 a development of the hand-jigging formerly employed in Cornwall . and introduced by English miners to this country." In its simplest form, the method consists in jigging an ore-mixture previously crushed to some maximum size (although, in some cases, even this preliminary is omitted) oil a relatively coarse sieve, and then jigging again on a filler sieve, the material passing through the first sieve and bedding. While many modifications have been necessary to adapt it for use in mills of large capacity, where hand-work was necessarily, replaced by machines, the principle remains the same; the fact that the English system has been successfully employed, both in this country and abroad, is well known; and arguments have been made for its efficiency and applicability in a wider sphere than it has occupied hitherto. II. REVIEW OF PREVIOUS INVESTIGATIONS. The fact that treating a mixture of minerals under jigging conditions increased the amount of mineral saved, or, as Professor Richards aptly terms it, 1 ?the extra jig-catch," has long been known. To account for this fact a number of theories have been proposed. The work of Rittinger2 in this field has, for many years, been a classic in the literature of ore-dressing. For the purposes of my present paper, however, the work of two American investigators, Prof. H. S. Munroe and Prof. R. H. Richards, is chiefly concerned. Professor Munroe has given the results of an elaborate series of experiments,3 and his deductions, based largely oil theoretical grounds, of this work. After reviewing briefly the two systems of jigging, followed by a discussion of Rittinger's formulas and

Sep 1, 1908

By Albert F. J. Bordeaux

INTRODUCTION. THE following general survey of the character and present condition of the silver-mines of Mexico, though not offered as a detailed treatise, has been prepared with care, is believed to he generally up-to-date, and is offered as a summary which may be found useful by mining engineers. Mexico has ranked for several centuries first among the silver-producing countries. It is now second as copper-producer, and seventh as gold-producer, with a fast-increasing output every year. The total amount of capital invested in its mining- and smelting-works is more than $250,000,000, while the yearly metallic production is $160,000,000. There are more than 20,000 recorded mining properties. The silver-mines of Mexico represent deposits of three principal classes: (1) veins and stock-works in eruptive rocks; (2) fissure-veins crossing the crystalline schists, granite, diorite, etc. ; and (3) veins and replacement-deposits in limestone. The true fissure-veins in some districts contain (in value) more gold than silver. In other deposits, the silver is decreasing in depth, leaving galena and zinc-blende as the principal ores.

Jan 9, 1908

By F. O. Blackwell

It is roughly estimated that some three hundred companies in the United States engaged in mining and the kindred arts now employ electricity in their operations. As all these plants have been installed within the last few years, it is evident that mining engineers have been quick to appreciate the great advantages of the use of electricity and adopt it for their many requirements. In

Jan 1, 1894

By Fred H. Klaer

Water supplies dependent upon induced infiltration can be developed by vertical wells, by infiltration galleries, or by horizontal water collectors. Each method is described and the advantages of water obtained by induced infiltration with respect to cost, chemical quality, and temperature are shown by specific water supplies. THE subject of ground water in industry is assuming increasing importance as new processes and new industries are developed. Industrial specifications of chemical quality and temperature for process water and for cooling purposes are becoming more rigid, and water must now have many properties to meet these requirements. As ground water is usually more constant in quality and temperature, many industries are turning to ground-water supplies to satisfy their growing need. As many published articles have stated during the past several years, the ever increasing demands have caused shortages of water supply in many localities throughout the United States. Although such shortages are generally of a local nature, they have given rise to many attendant problems, such as increased pumping costs, salt water encroachment, and rising ground-water temperatures. One partial solution to the problems of overdevelopment of the ground-water reservoirs in many areas can be accomplished by developing ground-water supplies that are dependent upon induced infiltration and by planning the use of underground reservoirs. These reservoirs can be filled during periods of excess precipitation and stream flow and can be emptied during periods of drought, in a manner similar to the use of surface reservoirs. In this way water that otherwise would be wasted into the oceans is salvaged during periods of excess. The development of water supplies dependent on induced infiltration is a relatively new concept and should be defined. When a well is pumped, the water level around the well is lowered in the shape of an inverted cone or cone of depression. As pumping is continued, the cone of depression will continue to expand, horizontally and vertically, until the quantity of water crossing the perimeter of the cone plus the recharge within the area of the cone is equal to the quantity of water being pumped.

Jan 1, 1954