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By B. A. Wilcox

ThERE are several reports in the literature which indicate that both solid-solution hydrogen and hydride precipitates can promote low-temperature em-brittlement of tantalum.1-3 For example, Imgram et al.2 showed that hydrogen additions in the range of 140 to 580 ppm promoted a ductile-to-brittle transition at slightly above room temperature in both wrought and recrystallized tantalum. This contrasts with the observation that tantalum of norma1 purity is very ductile, even at liquid-nitrogen temperature. Imgram et al. also found that hydrogen additions caused a deterioration of notch tensile properties of tantalum. The purpose of this communication is to present some limited evidence that hydrogen can also cause fatigue embrittlement of tantalum. The material used in this investigation was wrought, stress-relieved (1000°C, 1 hr), are-melted tantalum rod (5/8-in. diameter) containing the following impurities in weight percent: 0.005 C, 0.007 02, 0.0005 H2, 0.003 N2, 0.08 Cb, 0.01 W, and 0.01 Fe. Hydrogen charging was done in an atmosphere of flowing dry hydrogen for 3 hr at 600°C, followed by an homogenizing anneal in purified argon for 15 min at 600°C, and cooling in an argon atmosphere. Typical micro-structures of the hydrogen-free and the hydrogen-charged materials are shown in Fig. 1(a) and Fig. 1(b), respectively. Fig. 1(b) shows the appearance of hydride precipitates which, in many cases, formed preferentially parallel to the rod axis. The hydride dispersion was uniform across the diameter of the specimen. The fact that the precipitate was a tantalum hydride was tentatively confirmed by X-ray diffraction on a metallographically prepared surface. After hydrogen charging, chemical analyses by vacuum fusion revealed the hydrogen content to be 0.0295 wt pct. In addition, the oxygen and nitrogen contents were slightly raised to 0.0095 O2 and 0.005 N2.

Jan 1, 1964

Pulverized Fuel Combustion The apparatus shown in the illustration is a complete equipment f r supplying coal to boiler and metallurgical furnaces, cement and lime kilns, dryers, and to all furnaces requiring a temperature and quality of fame under prompt and easy control. It is claimed that the Aero Pulverizer makes practicable the highest efficiency obtainable from burning coal. It makes coal burn like a gas, a flame, the physical and chemical character of which is regulable- a flame that may be elongated or shortened, thus placing the zone of highest temperature where needed- a flame that may be made oxidizing, reducing, or neutral, as occasion may require. The coal is burned as pulverized, and there is no storage of the powder with its attendant hazard. Artificial drying before pulverizing is not necessary if the coal supply be sheltered from rain and snow. Where the Aero is used, it is wholly a furnace question whether a dryer should be installed; it is not at all a pulverizing or storage question. Labor is reduced to a minimum. Slack coal at low cost yields its last B.T.U.

Jan 11, 1915

By F. Petkovich, M. P. Sudbury

Development of a new method of treating flotation concentrates for the preparation of feed suitable for direct charging to blast furnaces or converters is described. The method takes advantage of the fact that pyrrhotite contained in the concentrates can be made to undergo exothermic reactions which, under properly controlled conditions, result information of oxidation products that serve to cement the concentrate particles together. Thus green concentrates either in the form of pellets or briquettes containing 5 to 8 pct moisture not only were hardened, but also completely dried autogenously by continuous countercu,vrent shaft treatment with air. The process was developed through pilot plant operations ranging from 35 1b per hr to 2.5 tph, culminating in products evaluation tests on a railway-car scale. Falconbridge Nickel Mines Ltd., Falconbridge, Ont., have practised sintering, on a travelling grate machine, for many years to prepare a pyr-rhotite-bearing Cu-Ni sulfide flotation concentrate for blast furnace smelting. In recent years the development of new mines, and the construction of concentration plants some 40 miles from the smelter has increased the quantity of concentrate shipped, and encouraged the search for a method of producing dry, hard pellets or briquettes suitable for direct charging to a blast furnace or converter, and capable of withstanding considerable handling. The cementing properties of pyrrhotite oxidation products were discovered during a search for a suitable agglomeration method for sulfide concentrates, and were exploited in a process subsequently developed for autogenously drying and hardening pellets and briquettes of pyrrhotite-bearing Cu-Ni sulfide flotation concentrate. EARLY TEST WORK The production of hard, dry pellets of concentrate filter cake was attempted in the drying plant of one of the new concentrators. While soft pellets were produced containing 5 pct moisture, attempts at further drying resulted in disintegration of the pellets and intolerable dusting. The plant therefore was controlled to produce a concentrate containing about 6 pct moisture. The partially dried concentrate, which was shipped to Falconbridge for sintering, tended to heat spontaneously in transit. An at- tempt to briquette the moist concentrate at Falconbridge, in a plant incorporating a muller, a 20-in. roll briquetting press, and a steam-heated travelling grate oven, was more successful, although the equipment corroded rapidly due to the acid nature of the concentrate after spontaneous heating had occurred. LABORATORY STUDIES OF SPONTANEOUS HEATING OF CONCENTRATE The spontaneous heating of partially dried concentrate was studied in the laboratory. Moist concen-

Jan 1, 1961

By Edwin Berry

THE consideration of the present value of a mine is of great importance when planning for the operation of a property,, and it is particularly so in the case of mines that have large ore reserves. At various times during the past 15 years, the writer has had occasion to make rather thorough studies of the subject, and has been greatly impressed by certain important points which these studies have brought out, points which, though they may be well known, are often lost track of in the rush to obtain low mining costs and high daily tonnage output. . It is the object of this paper to present briefly, and in a general way, some of these points, and especially those which have to do with the relationship which (1) ore reserves and plant capacity, (2) grade of ore mined, and (3) ore losses in mining, bear to the present value. Present value, sometimes called present worth or actuarial value, means the value today of the future earnings of an enterprise. Great industrial, manu-facturing or railroad enterprises, if properly handled, go on indefinitely and their value is directly dependent on their estimated annual earnings. Mines, or oil properties, on the other hand, have an earning capacity only until the ore or oil is exhausted, and their present value is dependent not only on their annual earnings but also on their life, which is the length of the period over which they can produce at a profit. A mine, or an oil property, is a "wasting asset," while an estab-lished manufacturing business is not, or should not be. In other words, the present value of a mining property. represents a sum of money on which there could be paid from operating profits a determined rate of interest, say 8 per cent., and in addition an amount which, if invested at compound interest as received, would at the end of the life of the mine pay back the present value. In his "Wealth and Income of the American People," Walter Renton Ingalls says: "The fundamental principle of mine valuation is the present worth of an annual dividend accruing during the period of years corresponding to the life of the mine."

Jan 7, 1922

By S. F. Shaw

The theory that maintaining a high back-pressure on the oil sand lowers the viscosity of the oil has been generally accepted. The theory has also been advanced that lower viscosity permits the oil to be moved through the sand to the well at an expenditure of less gas per barrel of oil than would otherwise be required. It seems reasonable to assume that production of oil at the minimum gas-oil ratio will result in a greater ultimate recovery of oil. During the past five years the writer has acquired data on thousands of tests made on wells to determine the best rate at which to produce them. In the majority of cases reduction in back-pressure did not definitely increase the gas-oil ratio. The information obtained from these tests has raised doubt in the mind of the writer as to the correctness of the theory that lowered viscosity results in lower gas-oil ratios and in higher ultimate recovery of oil. The reasons that explain the results obtained in these tests cannot yet be definitely stated, but the hypothesis is being tentatively assumed that there are other factors which outweigh the influence of viscosity. Energy in Gas Moves Oil to Well The general opinion is that in most cases it is the energy in the gas that moves the oil to the well. In wells flowing naturally it is also the energy in the gas that lifts the oil to the surface. In a flowing well, there are two very distinct operations: 1. Moving the oil through the sand to the well. 2. Lifting the oil to the surface. The more important function of the gas is that of moving the oil to the well. Means for lifting the oil can always be found once it has been delivered to the well. Principles Involved in Lifting Oil A well that flows naturally is the same as a gas-lift well except that all of the gas is supplied by the oil reservoir. This natural flow is subject

Jan 1, 1931

By J. L. Green

Virginia-Carolina Chemical Corp., in its Florida Operations, uses both roller mills and ball mills in dry grinding of phosphate rock. The roller mills have been in operation since 1954 in the Mining Department. The ball mills were installed and put in operation in 1960 in the Harding Concentrate Superphosphate Plant. The roller mills in the Mining Department are used to grind phosphate rock for outside customers. This involves short notice orders of varying grinds, grades and types of rock. The ball mills grind phosphate rock for use in V-C's own phosphoric acid and concentrated superphosphate plant. These mills grind generally two types of rock for one grind. In 1959, a 10% speed-up was made in the operating speed of the roller mills. This paper deals with the operational gains and resulting cost savings which evolved from this speed-up, and compares the present operations of the roller mills and of the ball mills in grinding phosphate rock. In January 1959 it was decided to determine the feasibility of reducing the grinding cost on the roller mills, per ton of rock ground, by speeding up the mill 10%. At this time, one mill was completely reconditioned so it would be in top condition. Extensive tests were then run both before and after the mill speed-up to determine the resultant cost savings. These tests showed that, due to this 10% speed-up, the mill capacity was increased approximately 20% and grinding power costs per ton were reduced approximately 10%. Table I shows the results of these tests. Each test run actually represents several of the same type rock and grind, averaged together. Each test run lasted approximately three hours. The validity of these runs has been confirmed by subsequent operation in grinding over 1,000,000 tons of phosphate rock. Accurate records have been kept on power consumption and repairs as well as tons produced. Two types of feed rock were used for the tests — pebble phosphate rock and flotation phosphate concentrate rock. The pebble phosphate rock size is approximately 80-90% plus 20 mesh. The flotation concentrate phosphate rock size is approximately 15% plus 20 mesh and 95% plus 100 mesh. The same two types of rock feed were used for all the test runs. Test Nos. 1, 2, and 3 were run at the original mill speed of 85.5 rpm on the vertical shaft. Test No. 1 was run with a feed of high grade flotation concentrates; the desired fineness attempted was 70% passing 200 mesh. Tests No. 2 and No. 3 were run with a feed of low grade pebble. Test No. 2 was run for a desired fineness of 70% passing 200 mesh. Test No. 3 was run for desired fineness of 50% passing 200 mesh.

Jan 1, 1963

By J. Mason

In order that the use of continuous mining machines be most economic, the equipment must be operated at rated capacity as much of the time as possible and downtime for repairs kept at a minimum. A large factor influencing any proposals for maintenance programs is the complexity of the machines—often the intricacies of the precision-built equipment requires engineering knowledge byond the scope of mine maintenance personnel. A training program is part of the answer; there are others. In addition to investigating a joint company-manufacturer equipment-design improvement program, Mountaineer Coal Co. has launched a detailed preventive maintenance program to combat extensive downtime often encountered when continuous miners are used. Everyone in the coal industry knows the importance of operating continuous miners at their rated capacity as much of the time as possible. The potential capacity of the miner is such that l-min. delay means from 3 to 8 tons of coal lost. Today's miner is a well designed, precision-built machine, complicated with hydraulic and electrical equipment and complex mechanical assemblies with all possible combinations of gears, levers, and rods. As the machine becomes more complex, the problem of maintenance becomes more acute. The transition from conventional to continuous mining has taken place so rapidly that the mine mechanic has been unable to keep pace. When there is a machine failure, the maintenance man usually employs either the trial-and-error or the play -it-by-ear method of repair, but this represents an expensive proposition. Mine mechanics, maintenance superintendents, and mine managers complain about the complexity of the machines and how difficult it is to keep them running. There is no shortage of skill in the design offices. Machines are designed by good engineers, but for engineers. What's the answer? Better trained maintenance men—better maintenance instruction—better field service. These are all good corrective measures, but the real cure is to provide for the men at the do it level. Designers have got to cooperate with men to simplify maintenance on the increasingly complex machine. The designer must be aware of how his machine can be maintained. Sometimes a simple design change in a bearing, oil seal, or a bushing will correct recurrent breakdowns. Parts such as filters, hydraulic cylinders, hoses, and pivot pins can be better arranged on the machine to make them accessible so that the mine mechanic can easily keep the machine producing. Therefore, to get our maintenance off on the right foot, the program must begin back in the design office, before the machine is delivered to the mine property. Realizing the importance of this aspect, our company set out on a joint company-manufacturer equipment-design improvement program. This is a formal program in which quarterly reports are written covering machine defects, design weakness, and general problems relating to machine maintenance and performance. The manufacturer in turn comments on each and every item appearing in these reports, indicating the action he proposes to take. In addition to the written reply, meetings are scheduled with manufacturer representatives to discuss further the outlined problems. The benefits realized from this organized program are both substantial and far reaching. Good maintenance begins with a good design. The subject of preventive maintenance is not a new one. Volumes have already been written on the subject. It consists primarily of doing the maintenance work when you want to rather than when the machine demands it. However, a preventive maintenance program can be overdone. Theoretically, if carried to ridiculous extreme, all working time could be spent inspecting and repairing the machines and no time would be left for production output. Our goal in maintenance is to maximize running time for all machines, and to prevent or at least minimize production downtime. INSPECTION AND REPORTING The center of any preventive maintenance schedule is regular and systematic inspection, but it must be

Jan 1, 1961

By N. R. Morrow, C. C. Harris

The experimental points which describe capillary pressure curves are determined at apparent equilibria which are observed after bydrodynamic flow has ceased. For most systems, the time required to obtain equalization of pressure throughout the discontinuous part of a phase is prohibitive. To permit experimental points to be described as equilibria, a model of capillary behavior is proposed where mass transfer is restricted to bulk fluid flow. Model capillary pressure curves follow if the path described by such points is independent of the rate at which the saturation was changed to attain a capillary pressure point. Aemodified suction potential technique is used to study cyclic relationships between capillary pressure and moisture content for a porous mass. The time taken to complete an experiment was greatly reduced by using small samples. INTRODUCTION Capillary retention of liquid by porous materials has been investigated in the fields of hydrology,l soil science,2 oil reservoir engineering,3 chemical engineering," soil mechanics,5 textiles: paper making: and building materials In studies of the immiscible displacement of one fluid by another within a porous bed, drainage columnsl and suction potential techniques2 have been used to obtain relationships between pressure deficiency and saturation (Fig. 1). Except where there is no hysteresis of contact angle and the solid is of simple geometry, such as a tube of uniform cross section, there is hysteresis in the relationship between capillary pressure and saturation.2 The relationship which has received most attention is displacement of fluid from an initially saturated bed (Fig. 1, Curve R2), the final condition being an irreducible minimum fluid saturation Swr. Imbibition (Fig. 1, Curve A), further desaturation (Fig. 1, Curve R), and intermediate scanning curves have been studied to a lesser but increasing extent.2 ,3,9-12 This paper first considers the nature of the experimental points tracing the capillary pressure curves with respect to the modes and rates of mass transfer which are operative during the course of measurement. There are clear indications that the experimental points which describe these curves are obtained at apparent equilibria which are observed when viscous fluid flow has ceased; and any further changes in the fluid distribution are the result of much slower mass transfer processes, such as diffusion. Unless stated otherwise, this discussion applies to a stable packing of equal, smooth, hydrophilic spheres supported by a suction plate5 with water as the wetting phase and air as the nonwetting phase. The discussion can also be applied, in many

Jan 1, 1966

By R. N. Zitter

PRESENT work was done in connection with a program for obtaining a low energy gap semiconductor suitable for photodetection. The inter-metallic compound InSb, which is now being in-

Jan 1, 1959

By R. L. Caldwell, E. E. Glenn, L. J. Norton, A. J. D. Straus, S. V. Holcomb, J. Zemanek

A new and unique logging tool, called the Borehole Televiewer (BHTV), has been developed to inspect boreholes and to evaluate formations. Even though geologists and engineers have had only about 3 years' experience with the BHTV, the thinking of many of them has been radically influenced by the fascinating and revealing "log-picture" recorded with the tool. The BHTV fills a need for a logging tool that can produce a direct and descriptive record of the physical properties of the borehole environment. Although primarily developed to evaluate fractured reservoirs, the BHTV has been used successfully to solve a variety of problems related to formation evaluation and borehole inspection. Geologists and petroleum engineers have been confounded for many years by the problems of locating and evaluating fractures. Interpretation of the usual resistivity, acoustic or radioactivity logs, as well as the analysis of cores, for evaluation of fractured reservoirs is extremely difficult and often not definitive. This is understandable since analysis of cores from fractured formations has reported variously: (a) low porosity, (b) no effective permeability, (c) lack of any matrix oil saturation and (d) oil stains on fracture planes as the only indications of potential production.14 In recent years acoustic amplitude logs have been introduced and used with varying degrees of success to locate fractures.8'9'11"13 Conventional amplitude logs at best give only qualitative answers where quantitative ones are desired. The logs reveal nothing of the orientation of the fractures. In addition to these logs, various types of cameras, television, and rubber impression packers have been used for borehole inspection. (See Appendix.) The BHTV logging tool supplies both the quantitative and the orientation answers about fractures. The borehole can be filled with any homogeneous, gas-free liquid such as fresh water, saturated brine, crude oil or drilling muds. This tool takes an "acoustic picture" of the borehole wall. The resulting log is a representation of the borehole wall as if it were split vertically and laid flat. The log is oriented with respect to magnetic north. Any physical changes in the borehole wall are seen as changes in picture intensity. In this way, fractures, deformation or pits are reflected on the log. Tool Description The fundamental parts of the BHTV are shown in Fig. 1. A piezoelectric transducer probes the borehole wall with bursts of acoustic energy. A flux-gate magnetometer senses the earth's magnetic field and provides the means for determining the orientation of the log. A motor rotates the transducer and mag-

Jan 1, 1970

By E. F. Adkins, R. W. Douglas

The variation of the spectral emissivity of osmium and ruthenium with temperature can be expressed by the following relations: DURING a study of the sintering characteristics of osmium and ruthenium, discrepancies in the previously accepted melting points of the metals were noted. The melting temperatures had not been measured at the time of the work, but were estimated in the literature as 2700°C for osmium and 2450°C for ruthenium.' A recent investigation by Baird2 has shown the melting temperatures of osmium and ruthenium are 3000" and 2250° C, respectively. In the present work, osmium bars sintered at temperatures as high as 2900° C showed no evidence of melting, whereas ruthenium showed definite evidence of melting at temperatures as low as 2250°C. In view of the apparent discrepancies, the melting temperatures of both metals were determined. The spectral emissivities of both metals were determined as an aid to studies of the sintering characteristics of these metals. MATERIALS AND PROCEDURES The materials used in these studies were metal powders supplied by Englehard Industries, Inc., through the courtesy of The International Nickel Co. Analytical data from three laboratories, although not in complete agreement as to the purity of the material or the major impurity elements, indicated that the maximum impurity levels were 0.5 and 1.0 wt pct in osmium and ruthenium, respectively. The metal powders were pressed to 1/4 by 1/4 by 6-in. bars at 20Tsi using carbon tetrachloride as a lubricant. The pressed bars were presintered 1 hr at 1200°C after heating 1 hr at 400°C. The presintered densities of the several bars of the metals ranged from 65 to 80 pct of ideal density. Specimens were self-resistance heated in a vacuum sintering bell. Temperatures were measured with an L&N, disappearing filament optical pyrometer. True temperatures were read at the base of an 0.30-in.-diam, 0.150-in. deep hole drilled into the center of a presintered bar. Apparent temperatures were read from the surface of the bar adjacent to the hole. MELTING POINT DETERMINATION The melting temperature of the metals was determined as the temperature at which a slight additional increase in current through the specimen resulted in a reduced temperature at the base of the hole. The apparent reduction in temperature is a result of the liquid metal filling the hole, thus altering the black-body conditions. The melting points of osmium and ruthenium, based on three deter-

Jan 1, 1962

By J. C. Detweiler

THE Jacksonville plant of Humphreys Gold Corp., operating under leases from National Lead Co. and Rutile Mining Co. of Florida, recovers rutile, ilmenite, zircon, and monazite from an ancient beach deposit containing about 2.5 pct of these minerals in combination. The location of the plant is unique, being situated about ten miles from the center of Jacksonville, Fla., surrounded by a thriving suburb. Two four-lane highways pass within 1 ½ miles of the plant. An expressway now under construction will pass over the tailing pile, within 500 ft of the mine office. The elevation at the plant site is approximately 50 ft above sea level. This is one of two mines producing titanium minerals and zircon in Florida. Another mine is also operated by Humphreys at Starke, about 50 miles southeast of Jacksonville. The Jacksonville deposit extends about six miles in a north-south direction, averaging approximately one half mile wide and twenty ft thick. Values begin at the surface over most of its length and ter-

Jan 1, 1952

By N. O. Taylor

THE failure, by swelling, of several cast aluminum-zinc, spiral, pump rods, used to circulate water in a constant-temperature bath, brought up the question as to whether the presence of tin in varying quantities would have any appreciable influence in controlling this action. Previous workers had found the same difficulty with aluminum-zinc alloys, due to their swelling in water or solutions of electrolyte^.^ It was known that pure aluminum was practically insoluble in concentrated sulfuric and nitric acids, presumably due to the formation of an oxidized film on the surface. Concentrated sulfuric acid had little or no effect on pure zinc and tin and concentrated hydrochloric acid would attack all three metals. These facts aroused curiosity as to the probable reaction of the ternary alloys of these metals, and tests were made to determine their resistivity to the action of common laboratory reagents. Samples, covering the entire range of the ternary series of aluminum, zinc and tin alloys, were made up from good grades of the commercial metals.2 The materials were melted, at as low temperatures as possible, in graphite crucibles in a gas-fired crucible furnace. Graphite crucibles were used in preference to clay, to prevent the absorption of silicon by the aluminum, as found by previous workers. The melted materials were stirred thoroughly with pine sticks in order to ensure complete mixing and the greatest possible freedom from oxidation, and were cast in sectional, iron molds giving bars % in. in diameter and 5 in. long.

Jan 1, 1927

By D. L. King, C. D. Rubert, R. W. Christy, B. A. Schepman

R. W. Christy (Manager of Sales, Sanitary Engineering Equipment, Link-Belt Co.) and C.D. Rubert (Barrett, Haentjens & Co.) — Messrs. King and Schepman have presented an excellent paper which outlines the general practice in the design of thickener mechanisms. Lifting devices, controls, special applications and underflow pumping arrangements have been reviewed in detail. Pumping arrangements were illustrated which eliminate the need for an expensive tunnel beneath the thickener floor. Most of those illustrated have the common feature of a relatively long underflow suction line extending from the hopper floor to the pump located on the bridge at the periphery of the tank. The authors limit these arrangements to comparatively dilute or thin slurrys. It is implied that dense and coarse concentrates, which are difficult to pump, cannot be handled unless a tunnel is constructed. Fig. 1 shows an underflow pump arrangement which negates the need for a tunnel but which permits the operator to convey a heavy dense slurry. The pump is mounted directly on the fixed hoist platform and the impeller is located at the base of the steel center column. Priming water is not needed. The column

Jan 1, 1963

By H. J. Rose

In a paper1 presented before this Institute in 1926,I briefly discussed the evaluation of coking coals by means of ultimate analysis. The paper contained several graphic studies in which coal analyses calculated free of moisture, ash, sulfur and nitrogen were plotted on triaxial diagrams in which carbon + hydrogen + oxygen = 100 per cent. From a study of such data for 600 coking coals of the United States, and 150 coking coals of Wales, France and Germany, it was concluded that it was usually possible to predict coking properties and by-product yields with fair success from ultimate analysis. The coals which have proved exceptions to the general rules were found to have physical peculiarities. Run-of-mine Coal It must not be forgotten that run-of-mine coal is a composite material consisting of many constituents differing in physical properties and chemical analysis. A coal sample usually contains: 1. Glossy black coal known as anthraxylon or vitrain. This constituent usually has the strongest coking properties, and the lowest ash content. 2. Dull black coal consisting of attritus containing thin layers of anthraxylon. It is also known as clarain. This constituent, in a coking coal, usually has good to excellent coking properties. 3. Mineral charcoal or fusain. This is a porous black material which is soft and crumbly (unless impregnated with mineral matter) and which has absolutely no coking properties. It is different in every respect from the other constituents in coal. The volatile matter content of mineral charcoal is less than 20 per cent. even when it is associated with high-volatile coals. As would be expected the hydrogen content is low, and the carbon content is high. 4. A coal sample may contain tough gray canneloid material which has a high-volatile matter and a high hydrogen content, but is poor in coking qualities.

Jan 1, 1930

By Fred D. Wright, Philip B. Bucky

FOREWORD By E. D. Gardner[$] THE present known petroleum reserves are limited, and unless important new fields are discovered the Nation will be dependent, in the not too distant future, upon imports or upon synthetic liquid fuels to supplement domestic petroleum supplies. As a national defense measure, it would appear desirable to develop a substitute source of liquid fuels within the borders of the country if new discoveries of petroleum should prove disappointing. Congress undoubtedly had this in mind when it passed the Synthetic Liquid Fuels Act (78th Congress, approved April 5, 1944). The act directed the Bureau of Mines to construct and operate demonstration plants for producing synthetic liquid fuels from coal and oil shale, and from forestry and agricultural products. As a part of this program, the Bureau of Mines has built a demonstration plant at Anvil Points, near Rifle, Colo., for retorting the oil shale, and has developed a mine to supply the plant with oil shale. At the beginning of mining operations, the physical qualities of the oil shale, which would have a bearing upon the selection of mining methods and practices for exploiting the deposits, were largely unknown. A comprehensive mining-research program has been set up to determine the best methods and practices to follow to obtain the lowest overall practicable mining cost for producing the oil shale on a large scale. Unusually low mining costs will be necessary for shale oil to be produced commercially. Consideration now is being given to exploiting a series of flat-lying beds 70 ft thick by a room-and-pillar method of three benches with the advance heading at the top. This report is one of a series giving the results obtained in research on mining problems. INTRODUCTION The structure and physical qualities of the oil shale indicate that the 70-ft interval of the oil-shale measure can be safely mined in open stopes; studies indicate that this method would be the most economical one to use.1 As labor will be a major item of expense, the largest equipment that could be operated efficiently in the workings would be required to obtain the lowest practicable mining cost. It was therefore desirable to know the maximum safe span of unsupported rooms and the size of pillars necessary to support the overburden. This paper describes the work done at the Columbia School of Mines Barodynamic Laboratory in determining the maximum safe unsupported roof spans and the minimum safe pillar dimensions for the oil-shale mine at Rifle, Colo.

Jan 1, 1947

By C. E. Gerity

Near Bunker, Mo., in the New Lead Belt, Ozark Lead Co. operates a 6000 tpd lead-zinc mine. Mined ore is transported to the surface in three mechanized and automated steps. The mine is developed by a rectangular-shaped sublevel haulage drift 15 ft high by 28 ft wide, driven on strike below the ore zones. Ore transfer raises were driven from the sublevel at locations near the shaft, 1500 ft distant, and on subsequent 1000 ft centers. These 6 x 5-ft raises are slashed to obtain maximum storage below dump points. These dump points are located on each mining level and are of steel and concrete construction. Spiral 12 X 24-ft inclines driven from the sub-level provide access to these levels for equipment and mine utilities.

Jan 11, 1972

By H. S. Price, J. C. Cavendish, R. S. Varga

A numerical formulation of bigh-order accuracy, based on variational methods, is proposed for the solution of multidimensional diffusion-convection-type equations. Accurate solutions are obtained without the difficulties that standard finite difference approximations present. In addition, tests show that accurate solutions of a one-dimensional problem can be obtained in the neighborhood of a sharp front without the need for a large number of calculations for the entire region of interest. Results using these variational methods are compared with several standard finite difference approximations and with a technique based on the method of characteristics. The variational methods are shown to yield higher accuracies in less computer time. Finally, it is indicated how one can use these attractive features of the variational methods for solving miscible displacement problems in two dimensions. INTRODUCTION The problem of finding suitable numerical approximations for equations describing the transport of heat or mass by diffusion and convection simultaneously has been of interest for some time. Equations of this type, which will be called diffusionconvection equations, arise in describing many diverse physical processes. Of particular interest here is the equation describing the process by which one miscible liquid displaces another liquid in a one-dimensional porous medium. The behavior of such a system is described by the following parabolic partial differential equation:

Jan 1, 1969

By C. Crussard, R. Tamhankar

It is the policy of The Metallurgical Society to provide, in the TRANSACTIONS OF THE METALLURGICAL SOCIETY OF AIME, a prompt and accurate medium for publication of reports of significant new research in both the scientific and engineering aspects of metallurgy. Papers submitted to the Society are judged by competent reviewers accord-ing to established criteria for technical merit, accuracy, originality, and presentation. STAFF: Editor, Gerhurd Derge Carnegie Institute of Technology Schenley Park Pittsburgh 13, Pa. Editorial Assistant, M. A. Redmerski Staff Coordinator, James J. Burke THE METALLURGICAL SOCIETY: W. R. Hibbard, Jr., President; J. C. Kinnear, Jr., Past-President; John Chiprnan, Vice President; T. D. Jones, Treasurer; R. W. Shearman, Secretary. PUBLICATIONS COMMITTEE OF THE METALLURGICAL SOCIETY: A. W. Thornton, Chairman; Dennis Carney, John Chip-man, J. C. Fulton, J. H. Hollomon, H. H. Kellogg, A. E. Lee, R. Maddin, J. D. Sullivan, and D. Swan. Review and selection of manuscripts is under the supervision of the following Publications Committees: Extractive Metallurgy Division: T. B. King, Chairman; S. J. Dickenson, R. P. Hermsdorf, R. Schuhmann, Jr., and P. T. Stroup. Institute of Metals Division: Robert Maddin, Chairman, F. Lincoln Vogel, Jr., Vice-chairman, W. A. Backofen, R. W. Balluffi, J. H. Bechtold, M. B. Bever, R. S. Busk, C. G. Goetzel, Paul Gordon, C. D. Graham, .Jr., Robert F. Hehemonn, R. I. Jaffee, F. E. Jaumot, Jr., Irving R. Kramer, A. J. Lena, Harold Margolin, C. J. McHorgue, Russell A. Meussner, J. B. Newkirk, H. W. Paxton, G. E. Pellissier, I. S. Servi, Alan U. Seybolt, A. J. Shaler, and Jack Washburn. Iron and Steel Division: J. C. Fulton, Chairman; T. B. King, F. C. Langenberg, W. 0. Philbrook, and T. 0. Winkler. Published bimonthly by the American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc., 29 West 39th Street, New York 18, N. Y. Telephone: PEnnsylvania 6-9220. Subscription $20 per year for non-AIME members in United States and North, South, and Central America; $25, foreign; $5 for AIME members. Single copies, $4.00, single copies foreign, $5.00. The AIME is not responsible for any statement made or opinion expressed in its publications . . . Copyright 1958 by the American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc. . . . Registered cable address, AIME New York ... Indexed in Engineering Index, Industrial Arts Index, and Chemical Abstracts... Second class postage paid at New Yark, N. Y., and Ann Arbor, Mich. Volume 212, Number 6.

Jan 1, 1959

By Milton F. Goodrich

The idea of using water-only cyclones to scalp clean coal from the feed to other devices has recently been gaining in popularity. 1,2,6 An indication of this popularity is that water-only cyclone scalping is already employed in a few commercial plants. Water-only cyclones scalp the feed to a heavy media vessel in the Warwick plant,3 to Deister tables in the Rowland plant,4 and to a heavy media cyclone circuit in the Amherst Coal Company's McGregor plant.5 Despite this trend, we found surprisingly little information in the literature on the performance of water-only cyclone scalping. Therefore, we undertook a study of its technical merits, limiting our investigation to a water-only cyclone scalper in conjunction with heavy-media cyclones. We felt this to be a reasonable combination of equipment, since both processes operate well on the same size coal. First, we developed simple mathematical relationships revealing general characteristics of the system. Next, we compared the clean coal yield and quality obtainable by the system on four different coals with that obtainable by heavymedia cyclone circuit alone and water-only cyclones middlings recirculation circuit alone. This was done by using computer program to simulate the operation of the three systems. Finally, we considered the effect that implementation of the system would have on a typical preparation plant circuit. The System The system chosen for study [(shown schematically in Fig. 1)] comprises two stages: a primary stage (PI) consisting of water-only cyclones, and a secondary stage (P2) consisting of heavy media cyclones. The water-only cyclones separate raw coal at a low specific gravity, producing a relatively small amount of high quality clean coal and a refuse containing much good coal. This refuse is then recleaned in the secondary stage, where the sharper separation of heavy media cyclones, operated at a higher specific gravity, reclaims additional clean coal. Finally, the clean coal from both stages is combined as the system's clean coal, while the refuse from the secondary stage becomes the system's refuse output. Theoretical Analysis Two assumptions lie behind the following theoretical analysis. The first assumption is that partition curves can represent not only the performance of each stage but also the performance of the system as a whole. These curves show partition number as a function of specific gravity, where partition number is either the percent or fraction of a raw coal with a given specific gravity reporting to clean coal. The second assumption is that these curves are independent of the raw coal characteristics. At any specific gravity the system's partition number is solely a function of the partition numbers of each stage. For convenience, we use the following variables in describing this relationship: g = specific gravity, PS(g) = the partition number of the system at specific gravity g, P1(g) = the partition number of the primary stage at specific gravity g, P2(g) = the partition number of the secondary stage at a specific gravity g. For this system, we can show algebraically that the relationship is: [Ps (g) = PI (g) + (1- P1(g)) P2 (9)(1)] With this equation the system's partition curve can be determined if the partition curves of each stage are known. By definition, the partition numbers P1(g) and P2(g) must each be less than, or equal to, one. Placing these constraints on Eq. (1), we arrive at the relations: [Ps(g) % P1(g) and Ps(g) % P2(g)(2)] Thus, at any specific gravity the system's partition number is always greater than that of both stages. [Figure 2] shows typical partition curves. Curve I represents the primary stage, which consists of water-only cyclones operating at a separating gravity of 1.3. (The separating gravity or the

Jan 1, 1979