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By J. E. Forciea, R. W. Salmi

Operating and research personnel at U.S. Steel's Pilotac concentrator have evolved specifications for magnetic separators suitable for use in beneficiation of Mesabi magnetic taconite. These specifications were developed as a result of investigations into the operating and metallurgical problems at the Pilotac concentrator over a period of some ten years. This paper discusses primary magnetic separator specifications, but much of the information is also pertinent to secondary and tertiary magnetic separators. Of particular significance is the proposal of a new basis for specifying the magnetic field which includes consideration of both field strength and field gradient. Other areas of discussion include flowsheet service, operating conditions such as drum speed and pulp level, tank configuration, magnet strength including field gradient, sleeve bearing design, designation of corrosion resistant materials and other details that result in a metallurgically efficient machine that has low repair and maintenance requirements. The mechanical and metallurgical design of magnetic separators at the Pilotac plant, Minnesota Ore Operations, United States Steel Corp. has progressed from the original single-stage, belt-type machines to the present multiple-stage, drum-type units, In ten years of operation, the plant operating and research personnel, in consultation with manufacturers, have made design changes to improve mechanical, electrical and magnetic features to meet process needs as flowsheet designs developed and to incorporate new permanent magnet materials as they became available. Separator specifications were an end result of this effort. The latest specifications which were developed for the Minnesota Ore Operations' plant were not all-inclusive listings; they covered only those points indicated by plant experience. For example, magnet pole piece design and dimensions were not stated. The manufacturers were given freedom to use any design as long as the field pattern was as specified. The selection of the items to be included in the specifications was undoubtedly influenced by particular conditions at Pilotac such as composition of plant water, concentration characteristics of the ore, repair and maintenance facilities available and the research programs that had been conducted. Only the developments leading to the specifications for a primary or "cobber" magnetic separator are presented in this paper. However, the same general concepts are used to specify secondary and tertiary separators. The areas of interest discussed include flowsheet service, tank configuration, magnet design, bearings, materials of construction, and some minor mechanical details. FLOWSHEET SERVICE The primary magnetic separators in the Pilotac Concentrator flowsheet are located after the first stage of grinding as shown in Fig. 1.

Jan 1, 1965

By Luther B. Eames

The recovery of dissolved gold from slime pulp in the cyanide process was first accomplished by intermittent decantation. This simple process consists in mixing with the pulp containing the values in solution, a solution of lower gold content, settling the mixture in a tank and decanting the clear supernatant fluid. The thick pulp remaining in the tank is pumped to a second tank together with more barren-solution and again settled and decanted. After several repetitions of this operation, values are so far reduced that further washing is not profitable. The gold recovery of this process is high, but the plant required is bulky, labor cost is high and the amount of solution to be precipitated is excessive. Historical Data As early as 1901, a plant was built in the Black Hills of South Dakota by John Randall, employing the same principles but attempting to make the process continuous by substituting for flat-bottomed tanks, cones which operated continuously, receiving a constant feed and discharging a steady stream of thickened pulp. These cones were operated in series, the thick underflow of the first one forming, with a stream of diluting solution, the feed to the second cone of the series. Barren solution was added to the tank immediately preceding the discharge tank and, after being slightly enriched by the low-grade pulp in this tank, overflowed to form a diluting solution again for the richer feed entering the third tank from the end of the series, and so on back to the richest tank of the series. Clear water was used for the wash in the final tank. This is the principle on which all successful countercurrent decantation plants operate at the present time, but Randall's plant was not successful because of mechanical difficulties in getting a continuous thick discharge from his cone tanks. A similar plant was built in South Africa although there the washes were not repeatedly used, as in Randall's case, but were precipitated after each contact with the ore. This also was abandoned

Jan 1, 1918

By Philip Drinker, R. M. Thomson

In the mining and metallurgical industries, numerous problems arise requiring determinations of solid and of liquid particles suspended in air. Frequently, these problems are of local interest and involve purely economic or operating difficulties; on the other hand, toxic effects from the inhalation or the ingestion of substances suspended in air may be a matter of concern. For checking up process improvements or for determining ventilation efficiencies in problems of this general nature, various methods have been proposed. .Although simpler methods are often satisfactory, precipitation combines several desirable features that other methods lack. This paper is, therefore, devoted to a discussion of portable alternating-current precipitators applicable to the quantitat,ive determination and microscopy of dusts, fumes, and smokes in air. Small particles suspended in gases, such as air, are referred to in the literature as aerosols or suspensoids. Of these terms, suspensoids appears to be preferable as being more nearly in keeping with words in everyday use. In problems involving air pollution by small particles—by suspensoids—it is convenient to distinguish three broad classes based on the size, the degree of dispersion, and the method by which the particles become suspended in the air. Dusts.—Particles or aggregates of particles, 150 to 1 micron in diameter, that are thrown into the air by mechanical agencies are termed dusts. Examples are silica; talc; cement; organic dusts such as hard rubber, starch, and cocoa; flocculated fume, and smoke products. Fumes.—Particles, 0.2 to 1 micron in diameter, formed from chemical or physicochemical reactions are known as fumes. Examples are ammonium chloride, lead and mercury, zinc and magnesium oxides, fogs, and acid mists. Smokes.—Particles, less than 0.3 micron in diameter, formed by incomplete combustion of carbonaceous and other substances are called smokes. Examples are the particles formed from burning tobacco, oil, tar, and gas.

Jan 1, 1925

By Gilbert H. Cady

Extensive sampling of coal in Illinois during the past 10 or 12 yegrs by the State Geological Survey, in cooperation with various organizations, such as the U. S. Bureau of Mines, the University of Illinois, and the Illinois Cooperative Mining Investigations, has made possible the delineation of two areas of low-sulfur coal in the southern part of Illinois. The $fur content is less than 1.25 per cent., so that, if otherwise suitable, these coals can be employed for metallurgical uses and for the manufacture of water-gas and retort gas. One of these areas is small and lies in Jackson County, near Murphysboro, the other is much larger and includes a large part of the famous Franklin County field. A small area of No. 2, or Murphysboro, coal has been worked for many years near the town of Murphysboro, Jackson County, Ill. In two mines, at least, the coal has a sulfur content of less than 1.25 per cent. It is doubtful, however, whether this field will ever be a source of large tonnage as the total area underlain by low-sulfur coal in workable thickness is probably less than 15 sq. mi. (24.14 sq. km.), and a large part of it has already been worked out.. The location of the area of low-sulfur coal in the Franklin County field is shown in the accompanying map. The small area underlain by the Murphysboro low-sulfur coal is shown near the town of that name in the central part of Jackson County. The larger area lies in the west side of Franklin County, extending also about 6 mi. (9.65 km.) south into Williamson County, about 4 mi. (6.43 km.) west into northern Jackson and western Perry County, and northward an undetermined distance into Jefferson County. All but the northern limit of the area is fairly well defined by sampling in numerous mines. The inner cross-lined area is underlain by coal having less than 1 per cent. sulfur, the outer boundary surrounding the area underlain by coal having less than 1.25 per cent. sulfur. The coal mined in the district is No. 6 or Herrin coal, commonly known as the Carterville or Franklin County coal. The bed has a thickness varying from about 8 ft. (2.4 m.) on the border of the low-sulfur area up to more than 10 ft. (3 m.) in the central portion, locally having a thick-

Jan 1, 1920

By H. M. Bannerman

IN the past few years the mineral raw materials problem has risen from comparatively obscurity to great national significance. The transition has come so rapidly that the nature of the problem and what it portends have not yet been clearly defined. Consequently there are wide differences of opinion among the experts as to what can, or should be done to insure a continuing, adequate supply. On one hand, it is said that depletion of known domestic deposits and lagging discovery of new ones will bring about an imminent decline in mineral production and that in future U. S. mineral-using industries will have to depend on foreign sources for more and more of the major mineral raw materials. On the other hand, it is argued that domestic mineral resources are virtually inexhaustible if there were only legislation against competition from low-cost foreign producers. Still another school of thought contends that within the past decade the country has entered a period of unprecedented mineral discovery that is rapidly solving the problem of depletion. Only time will tell which of these widely divergent views is nearer the truth, for mineral resources are, after all, the result of research and adaptation, and resources of the future, as of the past, will depend largely on human response to circumstances. Of 18 substances now termed reactive metals, only 8 (boron, thorium, potassium, sodium, lithium, uranium, molybdenum, and tungsten) were considered of commercial significance at the turn of the century. Except for hafnium and rhenium, the other ten were well known to chemists and mineralogists before this, but none had found use commercially, whereas some, like titanium, were looked upon as nuisances in the ores of other metals. Of these ten elements, several were considered of minor commercial significance, and others were of little more than academic interest as recently as 15 years ago. Before World War II only a few of the better informed spoke of uranium or thorium as possible sources of power. Still fewer talked of converting boron into a super- fuel, and the time had not yet arrived when hafnium, rhenium and even zirconium were thought of as critical to U. S. welfare and safety.

Jan 10, 1957

By A. M. Talbot

THERE is still considerable controversy as to the exact nature of the mechanism of the age-hardening process, in spite of the many experi-ments already performed. For this reason, a considerable number of typical systems should be examined in detail and their results correlated. Furthermore, the methods of studying the changes that take place during the progress of this phenomenon should be carefully scrutinized and the different methods applied to the same samples, preferably at the same time. The experiments described in this paper are presented to supply information on one age-hardening system and to show how the various methods of investigation compare in their indication of what has hap-pened. In brief, the progress of aging in magnesium alloys containing about 10 per cent by weight of aluminum at different temperatures was followed by measurements of hardness, electrical resistivity, dilation and lattice parameter, and by microscopic observation.

Jan 1, 1936

Asarco's Sacaton unit, located near Casa Grande 45 miles southeast of Phoenix, Arizona, was brought into production in February 1974. Ore reserves are estimated at about 48 million tons at 0. 95% copper. Mineralization of the orebody comprises both sulfides and oxides, with minor amounts of gold and silver. Preproduction stripping of 40 million tons of overburden was accomplished in less than two years. Mining is done at the rate of 9,000 stpd ore with a waste-to-ore ratio of 4.6 to 1. 0. The mine is developed on 40 ft benches, and by 1979 underground block caving methods will be phased into the operation at a depth of 1900 ft. Ore is mined in the pit two shifts per day, seven days per week, and is hauled with 85 ton trucks to the primary crushing plant. Trucks dump their loads into a 165 st capacity pocket equipped with a chain curtain which controls the ore flow to the 6 ft 6 in. wide by 24 ft long (2. 0 m by 7.3 m) vibrating feeder grizzly. Grizzly bars are inclined at a 5º slope on the first half of the 0 feeder and 10 on the last half. Grizzly bars have 6 in. (13 mm) spacings which pass 60 to 70% of the mine-run ore. Grizzly oversize drops into a 48 in. by 60 in. (1. 22 m by 1.52 m) primary jaw crusher. Both the grizzly undersize and crusher discharge are collected on a 42 in. (1.07 m) conveyor belt operating 398 fpm (121 mpm) for transport to the 4,750 live st intermediate storage area ahead of the fine crushing plant. Capacity of the Sacaton primary crushing plant is a nominal 1150 stph (1043 mtph). It operates fourteen shifts per week with an effective crushing time of four hours per shift. An economical open-air design was used, but an enclosed operator control room is located on the upper level where the dump truck, feeder, and crusher can be observed. Oversize rocks are removed from the feeder or crusher by the 10-ton bridge crane. Direct operating costs for this well-designed primary crushing plant, including labor, power, supplies, and maintenance, are only 3.5 cents per short ton. An interesting feature of the Sacaton orebody is that the non-sulfide copper minerals, which are less prevalent than the sulfides, occur intimately with the sulfide minerals; for this reason, all ores as mined are treated in a flotation circuit using sulfidizing methods for oxide recovery.

Jan 1, 1978

By Nicholas J. Grant, Henry Epstein, John Chipman

FOR many years steel producers have been concerned with the presence of hydrogen in steel. Hydrogen dissolved in excess of its solid solubility at the melting point may cause bleeding and gross unsoundness in ingots and castings. In fully-killed steels the hydrogen content which results in these difficulties varies with the alloy content1 and may range from 6 to 12 ppm.* It is also recognized that of heat treatment diminishes rapidly as the size of the part to be treated increases. The time required to remove a specified fraction of the hydrogen present increases as the square of the diameter of the section, making the treatment of large- sizes impractical. In order to prevent bleeding and gross unsound-ness due to hydrogen, and in view of the inadequacy of corrective heat treatment alone to provide optimum properties when hydrogen is present in large sections, efforts to limit the hydrogen content must be taken while the steel is still liquid. The main subject of this paper is a discussion of the factors affecting the hydrogen content of liquid steel in acid and basic open hearth, and basic electric two-slag furnace processing. Literature Survey Because many of the results and observations reported in the literature have a direct bearing on the results obtained in the present research and form part of the basis for conclusions reached, a brief summary of the findings is necessary. Methods for experimental investigation of hydrogen in liquid steel are comparatively new. The first attempts to estimate the hydrogen content of the open hearth bath were by Kalling and Rudberg,6 who devised a method for collecting gases evolved from the boiling bath of a basic open hearth furnace. They found that the gas during the boil generally contained 2 to 3 pct H2. From this and Sieverts'

Jan 1, 1958

By McDowell Graves, Joseph Jensen, W. D. Goold, M. L. Gwin

During the present year the Santa Fe Springs field has proved to be the most important oil field under development in the United States. Its production will exceed that of any other field. Without it, production within the United States would have been almost in even balance with consumption. The field has forced to the fore and compelled, more than any other single factor, serious consideration of the need of conserving the gas resources of the State of California. It has also contributed much to the development of new drilling equipment and important features in petroleum engineering, and has required the development of a new cement for use in deep wells because of high temperatures. It is the only oil field in the country where wells 8000 ft. deep have become so common that they are no longer given special attention. The field covers a little less than 1400 acres of land. In it 196 producing deep-zone wells have already been drilled since August, 1928, and 224 wells are still drilling. More wells may be drilled. This indicates that probably more than 450 wells will be drilled in the present campaign. In hardly a year's time it has caused the outlay of nearly $50,000,000 for drilling, and within the year 1929 it will have returned practically an equal amount of revenue. In contrast to conditions during 1923, when this field was once before an important and disturbing factor, the California oil industry has not been forced to refinance to meet the situation presented by such a large production of oil. There has been practically no unit selling and no stock selling, nor any unnecessary development far beyond the limits of the field. Production is now about 275,000 bbl. per day and gives every promisc of remaining at this figure or increasing to nearly 300,000 bbl. Six months from now the field should still be capable of producing more than 200,000 bbl. per day. Obviously, the need for conserving gas that would be blown to the air, incident to the production of so large a volume of oil, represents a problem in conservation deserving of the most serious attention of the oil industry and of the State of California.

Jan 1, 1930

By D. O. Noel, E. B. Gebert, J. D. Shaw

It is, of course, expected that manufacture of the various metal powders should involve numerous methods adapted to the specific characteristics of the metals themselves. Several methods for powdering a particular metal may be available, but the physical characteristics of the powder as produced by the several methods may be quite different. In the final analysis, the application of the metal powder dictates the physical properties that the metal powder must possess, and this, to a very large extent, will determine which of several possible methods of manufacture is adopted. Sometimes powder of the same metal is produced by more than one method, each powder having specific characteristics that make it particularly suitable for its application. For purposes of classification, the various methods of powdering metals can be grouped as follows: 1. Machining. 2. Milling (by ball mills, stamps, attrition mills, etc.). 3. Shotting (pouring molten metal into water, air, etc.). 4. Granulation (by stirring molten metal while solidifying). 5. Atomizing (disintegrating by steam, compressed air, etc.). 6. Condensation of metal vapor. 7. Reduction of oxide powders. 8. Chemical precipitation. 9. Electrolytic deposition. 10. Sintering (for producing alloy powders). 11. Formation of an alloy followed by dissolving or otherwise removing one of the alloying constituents. Machining Turnings, cuttings, filings, etc., are, of course, produced as by-products in machine work of various kinds, and find certain applications as metal powders. For example, iron powder so produced is used extensively as a reducing or precipitating agent and as a constituent in certain hard-finish cements.

Jan 1, 1938

By A. A. Smith

THE application of indium to bearings was first advocated by C. F. Smart1 in 1938, for the purpose of inhibiting corrosion of certain bearing alloys. He found that a thin layer of indium plated on the bearing surface and subsequently diffused at elevated temperatures markedly decreased the rate of corrosion in lubricating oils containing organic acids. Most of his work was done on cadmium-base alloys, although he showed that the treatment was also effective for copper-lead and other alloys. Since their introduction, indium-treated bearings have found considerable use, particularly in connection with airplane motors.2 Undoubtedly the effectiveness of the indium plate in resisting corrosion depends upon its relative concentration on the bearing surface after the diffusion treatment and after prolonged use at operating temperatures. Some diffusion experiments were made on a number of base metals that had been indium plated, to determine the depth of penetration after heating for reasonably long periods of time. As only the practical aspects were desired, no attempt was made to determine actual diffusion constants. METHOD Two different series of experiments were carried out to study the diffusion of indium into several base metals. Cast cylinders about one-inch diameter were machined to 0.750-in. diameter, so as to remove all surface imperfections, and final finishing was done with oo sandpaper. The original diameters were very carefully measured with micrometers and then a layer of indium was electrolytically deposited on the cylinders from a sulphate bath.3 In the first series of experiments, about 0.002 in. of indium was deposited and diffusion was carried out at 150°C. for a period of 500 hr. in a thermostatically con- trolled electric oven. After the diffusion period, successive layers were machined from the specimens and analyzed spectro- graphically and chemically-the spectro- graphic examination serving as a guide to determine which layers should be chemically analyzed. Apparently there was some surface oxidation of the indium plate due to the long heat-treatment, which may have had some effect on the diffusion. Also, the cadmium alloy showed some incipient melting, which probably was caused by the formation of a low-melting eutectic. To eliminate these factors, another series of experiments was carried out by heating in an electrically heated, automatically controlled, oil bath at lower temperature. The preparation of the specimens was the same as before except that a layer of indium 0.0001 in. thick, an amount commonly used commercially on bearings, was deposited on each of the cylinders. Diffusion was carried out at l20°C. for a period of 200 hours.

Jan 1, 1944

By E. W. Krampert

Oil development continued active in Wyoming during 1937, with much of the interest centered on prospecting for new pools, following a rather intensive seismograph play in 1936. The new fields of Lance Creek, Medicine Bow and Quealy were the centers of intense development in proved areas, but routine drilling occurred in most old fields. A new and deeper producing horizon (the Leo sand) was opened up in the Lance Creek field, which materially extends the producing area and will probably account for continued active development during 1938. The discovery in 1936 of oil in the Eocene Wasatch formation at Powder Wash, northwestern Colorado, tended for a time to stimulate prospecting in the deep Tertiary basins, but this interest as rapidly declined after drilling several failures on similar structures and the discovery that the productive area at Powder Wash itself was very small. A most interesting test of an Eocene structure was drilled by The Superior Oil Company of California on the South Baggs structure, in Carbon County, Wyoming. South Baggs is one of the best developed structures in the Red Desert Washakie Basin, which also includes the Hiawatha gas field and the Powder Wash gas and oil field. The Superior Oil Co. started the well according to contract but apparently had lost interest in the prospects, for it abandoned the well at 2138 ft. without testing the oil-bearing and gas-bearing horizons of the Wasatch. The Sinclair Wyoming Oil Co. drilled a deep test in the Muskrat gas field to the Madison. This well discovered oil in the Embar formation, which was considered noncommercial because of depth. After the Embar was abandoned the hole was plugged back and the various sands tested by shooting through the casing with the Lane-Wells gun. The Lakota sand was found to contain high-pressure gas and the well was completed as a 24-million-foot gasser at 2175 lb. rock pressure. This is a new sand in an old field. The Italo Petroleum Co. drilled a flank well on the Hamilton dome structure and found oil in the lower Embar formation. Thk merely extends the limits of production, as one old well on the structure had

Jan 1, 1938

By J. C. Martin

A theoretical analysis of steam stimulation is presented for single sands. The analysis includes the effect of steam production and most of the effects of heat conduction. The results show the effects of a number of important variables on the performance of an idealized well under steam stimulation. Calculated responses are presented which indicate the effects of steam production, amount of steam injected, water production, formation thickness and formation damage. Results indicate that steam production can cause large reductions in the heat contained in the formation. This effect can be eliminated by drawdown control. Water production reduces the amount of oil produced during stimulation. The optimum amount of steam to inject depends on economic factors as well as the well response. In many cases, the increased temperature resulting from stimulation reduces oil viscosity near the well sufficiently to overcome the effects of formation damage even if the damage is not removed during steam injection. Calculated responses for thin sands are more favorable than anticipated. INTRODUCTION Little has been published on the theory of steam stimulation'-' despite the interest it has created and the wide variation in well responses. The results of the present analysis provide an insight into steam stimulation, and the methods employed provide a foundation for future work. Analyses presented in Refs. 1 and 2 are very limited and apply to gravity drainage conditions. Ref. 3 contains an analysis similar to the one presented here. The idealized models used and the assumptions made in Ref. 3 are different from those used in this paper. The analysis assumes that after steam injection has heated a small portion of the volume within the radius of drainage of a single uniform sand, a shut-in soaking period is allowed before returning the well to production. The effects of gravity, capillarity, transient pressure and water-sensitive sands are neglected. The injection and soaking times are assumed short compared to the stimulated production time. The initial temperature is assumed uniform; thus, the results apply primarily to first-cycle stimulation. The effects of gas production other than steam are neglected, and the water-oil ratio during production is assumed constant. Steam stimulation involves the simultaneous variation of the temperature, pressure and saturations. General mathematical equations for these variations are complicated and can be very difficult to solve. Simplified equations based on idealized models are used to reduce the mathematics sufficiently to allow approximate solutions to be obtained. DISCUSSION INJECTION An idealized model for which heat conduction is neglected is used to represent the behavior of a well during steam injection. The mathematics for this model is presented in Appendix A which also contains an approximate solution for the behavior of the no-conduction model. SOAKING During soaking, the well is shut in. Only the temperature, pressure and saturation distributions at the end of soaking are needed in the analysis. During soaking the heat is considered to be conducted in a uniform medium from an initially uniformly-heated circular cylinder confined to the producing interval. At the end of soaking the saturations and pressures are assumed to correspond to the cold zone. Analysis of the heat flow during soaking is included in the next section. The radius of the heated cylinder is calculated from the following heat balance (for constant quality steam injection). At the end of the soaking period it is assumed that little or no free gas is present near the well, and that the soaking period has been sufficiently long that the steam zone has had time to expand and the steam has condensed. The condition where there is no soaking is considered in the next section. PRODUCTION In this section, an approximate method is presented for solving the equations of heat and fluid flow associated with the production of oil and water during steam stimulation. Where initial pressure drawdowns are sufficient to cause steam to flow into the wellbore, steam production is assumed to occur within a short initial adjustment period (Appendix B). The production practices followed soon after the well is returned to production can have a large influence on the amount of oil produced during the stimulation cycle. Under most conditions, there is a period of time in which some or all of the water in the heated zone is converted into steam and produced. This flashing of hot water is caused by the pressure in the heated zone dropping below

By S. G. Gibbs

A new method for predicting the behavior of sucker-rod pumping systems is presented. The pumping system is described by a flexible mathematical model which is solved by means of partial diflerence equations with the aid of computers. Polished rod and intermediate-depth dynamometer cards can be calculated for various bottom-hole pump conditions. The technique permits simulation of a wide variety of operating conditions, both normal and abnormal. The data generated with the new technique are useful in refining the criteria for design and operation of sucker-rod systms. INTRODUCTION Sucker-rod pumping systems are used in approximately 90 per cent of artificially lifted wells. In view of this wide application, it behooves the industry to have a fundamental understanding of the sucker-rod pumping process. Oddly enough, our understanding has been rather superficial. This is evidenced by the semi-empirical formulas which have been used as the basis for design and operation of sucker-rod installations. Thoueh- we have realized the limitations of our methods for many years, it has not been computationally feasible to use more refined techniques. With the advent and widespread use of digital computers, it is now possible to handle the mathematical problems associated with sucker-rod pumping. This paper summarizes a computer-oriented method which can provide greater insight into the sucker-rod pumping process. It is hoped that this technique, and techniques which may evolve from it, will prove to be the tool needed by industry to obtain the most efficient use of rod pumping equipment. THE MATHEMATICAL MODEL Prediction of sucker-rod system behavior involves the solution of a boundary value problem. Such a problem includes a differential equation and a set of boundary conditions. For the sucker-rod problem, the wave equation is used, together with boundary conditions which describe the initial stress and velocity of the sucker rods, the motion of the polished rod and the operation of the down-hole pump. Of these items, the wave equation, the polished rod motion condition and the down-hole pump conditions are of primary importance. Discussion of the mathematical model centers about these factors. ROD STRING SIMULATION WITH THE WAVE EQUATION The one-dimensional wave equation with viscous damp- is used in the sucker-rod boundary value problem to simulate the behavior of the rod string. This equation describes the longitudinal vibrations in a long slender rod and, hence, is ideal for the sucker-rod application. Its use incorporates into the mathematical model the phenomenon of force wave reflection, which is an important characteristic of real systems. The viscous damping effect postulated in Eq. 1 yields good solutions, even though nonviscous effects such as coulcomb friction and hysteresis loss in the rod material are present. Fortunately, the nonviscous effects are relatively small, so the viscous damping approximation used in the wave equation is adequate. The coefficient v is a dimensionless damping factor which is found in field measurements to vary over fairly narrow limits. For mathematical convenience the gravity term is omitted in Eq. 1. The effect of gravity on rod load and stretch can be treated separately, as will be noted later. Since Eq. 1 is linear, the legitimacy of this procedure is easy to demonstrate. POLlSHED ROD MOTION SIMULATION The motion of the polished rod is determined by the geometry of the surface pumping unit and the torque-speed characteristics of its prime mover. By determining the motion of the polished rod, we formulate an important boundary condition. From trigonometrical considerations it can be shown that the position of the polished rod vs crank angle 0 is given by (see Fig. 1) These equations are obtained from the general solution of the "four-bar" linkage problem and can be used to describe the kinematics of any modern beam pumping unit.' If prime mover speed variations are disregarded, the angular velocity of the crank is constant, and Eq. 2 can be used to predict the position of the polished rod vs time. However, the constant-speed condition leading to constant crank angular velocity is only approached in practice: hence, it is better to make provisions for prime mover speed variations in the mathematical model. The speed at which the prime mover runs is determined by its torque-speed characteristics and the torque imposed upon it. The torque that the prime mover "feels" is the net torque arising from the polished rod load and the opposing torque from the counterbalance effect. The

By N. A. Gokcen, J. Chipman

A revised treatment of the authors' published data eliminates the complex relation previously proposed between concentration of silicon and activity coefficient of oxygen in liquid iron. Revised values of the thermodynamic properties of the liquid solution are presented. IN a recent experimental study of the reaction SiO2 (s) = Si+ 2O; Kf, = [% Si] [% O]² [1] the authors' found a substantially constant equilibrium product in liquid iron at 1600°C of 2.8x10-5 They also reported extensive data on the reactions: SiO2 (s) + 2H2 (8) = Si- + 2H2O (g); K'2= [% Si] (H2O/H2 [2] and H, (g) + 0 = H2O (g); K'3 =( H2 O [31 (H2) [%O] From the results on reaction 3 and earlier data of Dastur² on this same reaction in the absence of silicon, they determined the activity coefficient of oxygen, f0, on the basis of the definition K3 = (H2O)/ (H2)f0 [% 0] where K, is the equilibrium constant and f0, is taken as unity in the pure Fe-0 system. Similarly values of fsi were deduced from results on reaction 2. In a more recent study" of analogous reactions in the system Fe-A1-0, it was found impossible to reconcile the results on reaction 3 with Dastur's data; accordingly the latter were ignored and the equilibrium results were extrapolated to find a value of K, at zero concentration of aluminum. This procedure failed to locate the cause of the discrepancy but it did yield reasonable values of activity coefficients. It also avoided introduction of the complex empirical relation between the oxygen activity coefficient and the concentration of the added element. The same type of discrepancy exists for system Fe-Si-0.' In the earlier paper an attempt was made to fit both sets of data by a single curved line (Fig. 6 of ref. l), the form of which is contrary to the theoretical requirement of a finite slope at infinite dilution. In the light of experience on the Fe-A1-0 system the discrepancy must be recognized as one which can be resolved only by more refined measurements. Accordingly Figs. 6 and 10 are retracted. It is pointed out also that until the discrepancy is resolved Figs. 7, 8, and 11 are subject to some uncertainty. Qualitatively the following conclusions still appear valid: 1—The activity coefficient of oxygen is reduced by addition of silicon. 2—In dilute solutions the activity coefficient of silicon increases with its concentration. 3—With respect to equilibrium in reaction 1, the above effects are approximately compensating. The discussion of K'1 in the previous paper requires no revision. It was pointed out that the constancy of the product [% Si] [% 0]² ndicated a compensating effect of the activity coefficients of silicon and oxygen. Therefore, as a very good approximation, K1 = K'1 and the following average values are suggested both for K, and K', at the temperatures 1550°, 1600°, and 1650", respectively, 1.0x10-", 2.8~10-" and 5.5 ~lo-'. Revision of the thermodynamic treatment is necessitated by the recent appearance of new data, based on a combination of combustion and solution calorimetry,' which yields for the heat of formation of low-cristobalite from the elements, the value —209,330 ±250 cal per mol at 25°C. This is about 4000 cal larger than the value previously accepted. The new value for cristobalite is used, together with Kelley's tables of high-temperature heat contents" and entropies and with Korber and Oelsen's' heat of fusion of silicon to obtain the following equation for the standard free energy of cristobalite in the temperature range 1700" to 2000°K: Si (1) + 02 (g) = Si02 (crist.); ?F° = -217,700 + 47.OT [4] The free energy of solution of 0, in liquid iron is:8 O2 (g) = 20 (in Fe); AF° = -55,860 - 1.14T [5] and these two equations are combined to give: Si (1) + 20 = SiO, (crist.); AF° = -161,840 + 48.14T [6] ?F°1873 = -71,700 cal. From the experimental value of K, = 2.8x10-5, Si + 20 = SiO2 (crist.); ?F°1873 = -39,000 cal. [7] The combination of Eqs. 6 and 7 yields the free energy change when liquid silicon dissolves in iron to form the dilute solution of unit activity (1 pct). Si (1) = Si; ?F°1873 = -32,700 cal. [8] The heat effect in this process according to Korber and Oelsen' is an evolution of 28,500 cal per gram

Jan 1, 1954

By P. Marier, T. R. lngraham

The kinetics of the sulfation of MnO,, MnzO3, and Mn3O4 in SO,, SO3, and O, mixtures was examined and the descending order of sulfation rates at temperatures near 400°C was found to be Mn,O3 > MnO, > Mn3O4. The respective activation energies for the thermal decomposition of MnO, and Mn,O, are 39 and 47 kcal. At 900°C, the thernzal decomposition of MnSO, to MnzO3 is slower than that to Mn3O4. The respective activation energies are 62 and 51 kcal, respectitely. MANGANESE is used to improve the hot workability of steel in the proportions of approximately 13.5 Ib of ferromanganese per ton of steel produced. This requirement accounts for about 95 pct of its large industrial market in North America. The remaining 5 pct is used in the battery and chemical industries. In most North American ores, the percentage of manganese and the manganese-to-iron ratio are not suitable for the direct production of ferromanganese. Hence, most of the North American requirements for manganese are satisfied by importing ore. Typically, many of the studies done on native low-grade manganese resources have been directed toward the production of ore substittes" and the recovery of manganese from open-hearth slags.3 Of the wide variety of processes which have been proposed, the most popular involve compounds in the Mn-S-O system. The thermodynamic properties of manganese and its compounds were reviewed by ah' in 1960 and, more recently, lngrahams discussed the thermodynamics of some of the reactions involved in the Mn-S-O system at normal roasting temperatures. The conditions for producing manganese sulfide during the reduction roasting of manganese sulfate are discussed by Fuller and Edlund.9 A novel scheme was proposed recently by zimmerley7 for utilizing waste sulfur dioxide from stack gases to recover the manganese from ocean-mined manganese nodules. Very little of the work published on manganese compounds has been related to reaction rates. Singleton et 1.' studied the rates of reaction on the MnO-C and Mn7C3-3MnO systems and observed linear and parabolic kinetics respectively in the systems. Tatievskaya et a1.' studied the low-temperature reduction of Mn&, MnO,, and Mn& in HZ and CO, and reported activation energies in the range 1628 kcal. In this paper, the rates of some of the decomposition and formation reactions involving MnSO4, MnOz, MnzO3, and Mn3O4 will be examined after the conditions for the thermodynamic stability of the individual compounds have been designated. CONDITIONS FOR STABILITY OF MnSO4, MnOz, MnZO3, AND Mn3O4 The areas of stability for h'hSO4, MnOz, MnzO3, and h3O4 were established from the data of Mah4 and Ingrahams and are shown in the predominance area diagram,10 logpq- logpsq, in Fig. 1. The diagram is drawn for two temperatures, 700°K, solid lines, and 1100°K, dotted lines. The sketch for the lower temperature includes the conditions likely to prevail in the Zimmerley patent7 when manganese nodules react with sulfur dioxide and that for the higher temperature indicates the conditions for recovering MnzO3 or Mn3O4 from MnSO4 during a roasting reaction. From the fact that the boundary between the areas of stability of MnOz and hSO4 at 700°K is parallel to the abscissa, it is evident that MnOz and SO2 should react together to produce MnSO4, irrespective of the oxygen pressure in the system. If the source of sulfur for sulfation were from waste flue gases, it is likely that the oxygen content of the gas stream would be more than sufficient to oxidize any MnsO4 to MnzO3, 107 atm of O2 required, or even to convert any Mnz03 to MnOz, 0.02 atm OZ required, prior to sulfation. At 1100°K, the diagram indicates that MnSO4 may be converted directly to either MnzO3 of Mn3O4, depending upon the prevailing partial pressure of oxygen. When the gas stream contains more than 1 pct Oz, logpq = -2, only MnzO3 would be recovered from an experiment done under equilibrium conditions. At oxygen partial pressures of less than 1 pct, one would expect to bring about a reversible exchange between MnSO4 and Mn3O4 by appropriate adjustment of the partial pressure of SOZ. These various reactions will be described in the subsequent kinetic experiments.

Jan 1, 1969

By Gen-ichi Nakazawa, Masamichi Fujimori, Ichiro Doi

INTRODUCTION The major products of nickel industry in Japan are the electrolytic nickel (E-Ni), the ferro-nickel (Fe-Ni) and the nickel oxide sinter (NOS), totaling in production to 90 - 100 X 103 tons/year in terms of pure nickel content, and sharing approximately 20% for E-Ni, 10% or less for NOS, and the rest for Fe-Ni. The raw materials to be used for these products depend totally upon the imports from abroad; matte, the raw material used in the production of E-Ni, is from Australia and Indonesia; matte for the NOS is from Indonesia and New Caledonia; and garnierite, the raw material for Fe-Ni, is imported in bulk from New Caledonia, Indonesia and the Philippines. This report describes the acquisition of these raw materials, the production processes, trend of supply-demand, problem areas and the measures taken, and so forth, in connection with the nickel industry in Japan. ELECTROLYTIC NICKEL 1. History The nickel industry in Japan began in 1939 by the Sumitomo Metal Mining Co., Ltd. (SMM) with 150 tons E-Ni/month under the blast furnace smelting and the metal anode electrolysis. This was temporarily suspended after the end of last World War II, but resumed in 1952 with garnierite of New Caledonia as a raw material. Since then, the production capacity has been expanded incidentally to the increase of demand. However, since the smelting of garnierite contained about 2% nickel was very costly, the smelting process had been changed to smelt the briquettes of the mixture of garnierite and nickel sulfide concentrate imported from Canada and Australia. In 1975, this smelter was closed, and SMM's nickel refinery used matte from Western Mining Corp. (WMC) of Australia as a raw material. The refining process had been the metal anode one. The crude nickel oxide derived from roasting the high grade nickel matte had been melted and reduced with coke in the electric furnace and cast into anode. In 1970, a new refinery was constructed to produce 1,000 tons Ni/month under the matte anode process, and further, in 1976, the old refinery was converted into the matte anode process, thus resulting in the total production capacity of 1,690 tons Ni/ month today under the matte anode process. Shimura Kako Co., Ltd. began in 1951 to produce a little of E-Ni, and in 1955, they constructed a shaft furnace smelter with New Caledonian garnierite as a raw material to begin the E-Ni production under the matte anode process. Its monthly production reached 400 tons in 1957, but it discontinued its smeltering in 1963 to convert to the matte import from S. Le Nickel, WMC and INCO. In 1979, it changed the source of matte supply from the above to Soroako of Indonesia, discontinued the E-Ni production in 1979, and arranged for a total toll contract with Sumitomo. Meanwhile, SMM has developed a very distingished process for nickel-cobalt separation and refining by a solvent extraction method from nickel and cobalt mixed sulfide generated as a by-product in the nickel recovering process from laterite under the Nicaro process. Nippon Mining Co., Ltd. (NMC) also developed its own process. The plants under these processes were constructed in 1975, with the monthly production capacity of 230 tons of nickel and 130 tons of cobalt by SMM, and 300 tons of nickel and 110 tons of cobalt by NMC. SMM imports the raw material from the Marinduque Mining and Industrial Corp. of the Philippines, while NMC from Queensland Nickel Pty. Ltd. of Australia. 2. Production Status 2.1 Producers and materials Today, SMM holds the monthly E-Ni production capacity of 1,920 tons, of which 1,690 tons are from the matte anode process, and the remaining 230 tons are from the mixed sulfide as a raw material. NMC also uses the mixed sulfide as a raw material for a production of 300 tons/month. [ ] Most of SMM's raw material, nickel matte, is imported from WMC of Australia, partly using the Soroako matte from Indonesia. An example of material analysis is given in Table 1. Matte is a raw material for E-Ni and NOS, and the major sources of matte supply at present are Australian WMC and Indonesian Soroako, occupying ap-

Jan 1, 1982

By W. E. Newell

The basic principles and assumptions involved in D. L. White's solution5 for ultrasonic wave amplification in piezoelectric semiconductors are summarized. If the gain per unit length is maximized at each frequency without regard to the drift-field power density, the gain increases linearly with frequency. Therefore it would seem that very high gain is possible for frequencies in the hundreds of megacycles. However, with present materials the resulting drift-field power density is so large that cooling would be a major problem. For this reason it is more realistic to judge the capabilities of an ultrasonic amplifier in terms of the gain that is possible with a specified power density. In this case the gain at high frequencies does not increase but decreases more rapidly than l/f. The calculation of the gain us frequency at a fixed power level involves the repeated solution of a cubic equation and is therefore tedious. For the purpose of comparing different materials, sufficient accuracy can be obtained from an approximate solution which may easily be plotted as three straight lines on logarithmic graph paper. This method is used to compare the Performances of CdS, CdSe, ZnO, and GaAs. At a power density of 10 watts per cu cm, the first three give maximum gains of 60 to 80 db per cm in the vicinity of 60 megacycles. At the same power density, GaAs has a maximum gain of about 7 db per cm near 150 megacycles. Although an increase in mobility usually increases the gain for a specified power density, it is shown that there is an optimum mobility above which the gain decreases. An approximate expression for the optimum mobility is established. It is well-known that the interaction between charged particles and an electromagnetic wave can be used to amplify the electromagnetic wave. In the conventional traveling-wave tube, electrons in a vacuum are caused to move at a velocity slightly greater than the wave velocity, thereby coupling energy into the wave.' In principle, the same effect would be evident in a solid except that repeated collisions prevent the electrons from attaining a velocity greater than that of an electromagnetic wave. However, the drift velocity of electrons can be made to exceed the velocity of an acoustic wave in a solid. In 1956, weinreich2 noted that the deformation potential provided one means of coupling energy from drifting electrons into an acoustic wave, but the effect was too small to be of any practical importance. It was not until 1961 that White and Hutson demonstrated that the piezoelectric effect provided a strong enough coupling to give appreciable amplification of an ultrasonic wave and awakened wide interest in electron-phonon interactions. Qualitatively the effect may be described as follows. An acoustic wave propagating through a piezoelectric material is accompanied by an electric-field wave. If the vibrations are properly oriented with respect to the crystal axes of a piezoelectric semiconductor, the local electric fields are axially directed and cause bunching of the charged carriers. If the drift velocity of these carriers is less than the velocity of sound, they tend to be dragged along by the acoustic wave. This electroacoustic effect results in a net potential appearing across the material and in attenuation of the acoustic wave. On the other hand, if the drift velocity exceeds the velocity of sound, energy is coupled from the carriers to the acoustic wave and amplification results. Related effects which are being investigated include ultrasonic amplification in semimetals6-' and semi-conductor, and current saturation in piezoelectric semiconductors.'4-17 There are various problems which hinder the immediate application of ultrasonic amplification in a practical device. First, there are the traditional problems encountered in the design of acoustic delay lines: attenuation in the electromechanical transducers, reflected waves (which can lead to self-oscillation in an amplifying device), mechanical tolerances necessary to maintain phase coherence, and so forth. Second, there is the problem of selecting the best material from among the various piezoelectric semiconductors as they become available in single-crystal form with a wide choice of characteristics. This paper deals with the second problem and shows that the most realistic way of comparing materials is in terms of the gain which can be obtained without exceeding a

Jan 1, 1964

By H. O. McLeod, J. M. Campbell

This paper presents the results of the data obtained in the first stage of a long-range study at high pressures of the system, vapor-hydrate-water rich liquid-hydrocarbon rich liquid. The data presented are for the three-phase systems in which no hydrocarbon liquid exists. Tests were performed on 10 gases at pressures from 1,000 to 10,000 psia. One of these was substantially pure methane, and the remainder were binary mixtures of methane with ethane, propane, iso-butane and normal butane. Several conclusions may be drawn from the data. 1. Contrary to previous extrapolations, the hydrocarbon mixtures tested form straight lines in the range of 6,000 to 10,000 psia which are parallel to the curves for pure methane, when the log of pressure is plotted vs hydrate formation temperature. 2. The hydrate formation temperature may be predicted accurately at pressures from 6,000 to 10,000 psia by using a modified form of the Clapeyron equation. The total hydrate curve may be predicted by using the vapor-solid equilibrium constants of Carson and Katz' to 4,000 psia and joining the two segments with a smooth continuous curve between 4,000 and 6,000 psia. 3. The use of gas specific gravity as a parameter in hydrate correlations is unsatisfactory at elevated pressures. 4. The hydrate crystal lattice is pressure sensitive at elevated pressures. INTRODUCTION Prior to 1950 many studies had been made of the hydrate forming conditions for typical natural gases to pressures of 4,000 psia.""'"'"" Most of these attempted to correlate the log of system pressure vs hydrate formation temperature, with gas specific gravity as a parameter. One of the more promising correlations was made by Katz, et al, which utilized vapor-solid equilibrium constants. The only published data above 4,000 psia are those of Kobayashi and Katz7 for pure methane to a pressure of 11,240 psia. In the intervening years, most published charts for the high-pressure range have represented nothing more than extrapolations of the low-pressure data, with the methane line serving as a general guide. The reliability of these charts has become increasingly doubtful (and critical) in our present technology as we handle more high-pressure systems. The portion of our high-pressure hydrate research program reported here was designed to: (1) investigate the reliability of existing charts; (2) obtain actual data on gas mixtures to 10,000 psia; and (3.) develop a simple hydrate correlation that was more reliable than those which simply used specific gravity as a parameter. Binary mixtures of methane and ethane, propane normal butane, or iso-butane were injected into a high-pressure visual cell containing an excess of distilled water. Hydrates were formed and then melted to observe the decomposition temperature of the hydrates at pressures from 1,000 to 10,000 psia. EQUIPMENT The equipment consisted of a Jerguson 10,000-lb high-pressure visual cell, a 10,000-1b high-pressure blind cell and a Ruska 25,000-1b pressure mercury pump. The visual cell was placed in a constant-temperature water bath controlled by a refrigeration unit and an electric filament heater. A Beckman GC-2 gas chromatograph was used in analyzing the gas mixtures after each run was completed. EXPERIMENTAL PROCEDURE After evacuating the gas system, the heavier hydrocarbon was injected into the high-pressure mixing cell to that pressure necessary to give the desired composition. This cell then was pressured to 1,100 to 1,200 psia by methane from a high-pressure cylinder. The mixing cell holding the gas contained a steel flapper plate and was shaken intermittently over a period of 15 minutes. After mixing, the valve to the high-pressure visual cell containing excess distilled water was opened, and the gas mixture was allowed to flow into the cell. The temperature in the water bath was lowered 10" to 15'F below the estimated hydrate decomposition point. As a first check, the temperature was increased at a rate of 1°F every six minutes to find the approximate point of decomposition. It was again lowered 1.5° to 5°F to form hydrates. The temperature was raised to within l° of the estimated decomposition point and then increased 0.2F every 10 to 15 minutes until the hydrates decomposed. This procedure was repeated at various pressures to obtain 7 to 13 points for each mixture between 1,000 and 10,000 psia. After completion of the hydrate decomposition tests, the gas mixture composition was analyzed with a calibrated gas chromatograph. These gas analyses have an estimated error of ± .1 per cent.

By Chas. A. R. Lambly

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

Jan 1, 1950