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By John P. Bryant

Plant nutrients are obtained by plants from both the air and the soil. Carbon dioxide, a gaseous form of carbon and oxygen, supplies the carbon which usually makes up 50% or more of plant structure. Plants are capable of absorbing carbon dioxide directly from the air. Most of this nutrient comes from the air, although some carbon dioxide is absorbed by soil water and taken into the plant as liquid. Growing plants use more water than any other substance. Plant nutrients in the soil dissolve in soil water; soil water entering the plant carries nutrients into the plant structure. Water is then given off into the air as the plant breathes or transpires. The inorganic or mineral portion of the nutrients which plants require comes from the soil. Inorganic or mineral composition of plants is usually less than 5% of their weight, and this mineral requirement, usually lacking in the soil, must be provided by fertilizers. Proper fertilization of a soil to improve its mineral level and productivity is the basis for the fertilizer industry (Thomas, 1960). There are 13 essential plant nutrients which often must be added to soils or plants and are generally considered as fertilizer minerals. These nutrients are divided into the categories: primary minerals (nitrogen, phosphorus, potassium); secondary minerals (calcium, magnesium, sulfur); and micro or trace minerals (boron, iron, manganese, copper, zinc, molybdenum, chlorine). Further, cobalt is often added to fertilizers. Although it is not considered to be an essential plant nutrient, it is essential to animal health and is often applied to forages as a fertilizer constituent. One point of major significance to the fertilizer industry is that not all nutrient carriers supply that nutrient in a chemical form which will be available to the plant. Thus, a product which may be considered a fertilizer mineral may require substantial processing or treatment before it becomes useful as a nutrient carrier. Fertilizer Industry Scope Until about 1900 demands for higher crop yields were met simply by bringing new agricultural lands into cultivation. Since that time great progress has been made in increasing yields through fertilization and improved varieties and cultural practices. The importance of the fertilizer industry to the economy and welfare of the U.S. is widely recognized. [Fig. 1] illustrates the relationship of lime and fertilizers to crop yields. In terms of tons of fertilizer products, the growth of the industry has been phenomenal, particularly between 1950 and 1970. Table 1 shows the increase in fertilizer usage. Primary Plant Nutrients in Fertilizers Primary minerals (nitrogen, phosphorus, potassium) are designated as such because of their significance to plant growth. Practically all soils are deficient in both nitrogen and phosphorus, and most of them are also deficient in potassium. Consequently, these nutrients

Jan 1, 1975

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By A. W. H. Morris, G. H. Laurie, J. N. Pratt

Using- galvanic cells of the form Sn(liq)/Sn" (LiCl-KC1-SnCl,)/Sn-Ag (alloy), measurements have been made of relative thermodynamic properties of the a, C, E, and liquid phases of the Ag-Sn alloy system. Partial heats of solution of the components in the liquid alloys lzave also been observed by direct cal-orimetric measurement in an isoperibol calorimeter. The thermodynanzic quantities are disczlssed in relation to structural and other properties and the existence of anomalous minor fluctuations in the partial heats and entropies of solution in liquid alloys is tentatively suggested. In the course of a recent electro motive-force study of the thermodynamic properties of Sn-Ag-Pd liquids,' some measurements were also performed on the Ag-Sn binary system. Most previous thermodynamic studies of this system have been concerned with the liquid state. Measurements confined to the limiting heat of solution of silver in liquid tin have been made by many calorimetric workers2 while high-temperature calorimetric measurements of the heats of formation of the full range of liquid alloys are reported in the early work of Kawakami~ (1323°K) and more recently by Wittig and Gehrin~(1248°K). Electromotive-force studies on liquid alloys have been made by Yanko, Drake, and Hovorka' (606" to 686°K; 86 to 99.4 at. pct Sn) and by Frantik and Mc Donald' (900°K; 30 to 90 at. pct Sn). The only known measurements on the solid state are of heats of formation of the a, £, and c phases; these values were obtained using tin-solution calorimetry, at 723"K, by Kleppa,~ whose investigation also yielded heats of formation of liquid alloys containing more than 64 at. pct Sn. The present experiments on the Ag-Sn alloys include a re-examination of the liquid phase and, because of the dearth of free-energy data for the solid state, attempted measurements on the a, c, and E phases. The principal new feature of electromotive-force results obtained for the liquid phase was an indication of anomalous fluctuations in the partial heats and entropies of solution of tin at certain compositions. However, since the values for these thermodynamic quantities were determined from the temperature coefficients of cell potentials, which are commonly subject to considerable error, confirmation by calorimetric measurements was considered desirable. A detailed investigation of the partial heats of solution of the components in the binary liquids was made using a liquid metal solution calorimeter. I) GALVANIC CELL STUDIES a) Experimental Details. Measurements were made, as a function of alloy composition and temperature, of the potentials of reversible galvanic cells of the form: ~n(liq)/~n++/~n-Ag (solid or liquid alloy) Details of the apparatus and experimental techniques have been given elsewhere.' so that a brief account will suffice here. Molten salt, 58 mole pct LiC1-42 mole pct KC1, containing small amounts (1 to 2 mole pct) of stannous chloride was used as the electrolyte. The salts were dehydrated by pre-electrolysis and vacuum -drying techniques. Cells were established under an argon atmosphere by immersing tin and alloy electrodes in the molten salt contained in a large silica tube, heated in a vertical resistance furnace. The tube was sealed at the top by a head plate provided with openings permitting the simultaneous insertion of six electrodes, a central thermocouple sheath, and connections to vacuum and argon lines. Temperatures were controlled to *0.2"C over prolonged periods, with maximum variation over the electrodes at any time of 0.l°C. Temperatures were measured with a standardized Pt/13 pct Rh-Pt couple. The electromotive force of this and the cell potentials were measured on a Cambridge Vernier potentiometer and short-period galvanometer. Alloys were prepared from Pass "S" tin (99.999 pct) and Johnson-Matthey high-purity silver (99.999 pct) by melting in evacuated silica capsules and quenching in cold water. For liquid phase experiments, pieces of the resulting alloys were remelted into prepared silica electrode units, while solid electrodes were prepared by remelting into 3-mm bore tubing, inserting a cleaned molybdenum lead wire, and quenching to produce uniform rods about 3 cm in length, with soundly attached leads. In all cases remelting was done under an argon atmosphere. The solid electrodes were subsequently annealed in evacu ated silica tubes for 14 days at about 20°C below the solidus and quenched. Analyses showed that these techniques produced uniform electrodes with no significant change from weighed out compositions. b) Results and Discussion. Measurements were made on about forty alloys in the solid and liquid states, over varying ranges of temperature between 550" and 1050°K. Stable, mutually consistent, and reproducible electromotive-force data were obtained with most liquid alloys and these are shown in Fig. 1. Investigations were extended below the liquidus tem-

Jan 1, 1967

By Robert A. Melton

Since many of the easily mined, continuous, thick coal seams in the Appalachian Region have been exhausted, new exploration and mining is being directed toward less continuous seams that are under more cover and are more difficult to mine. The direct results of this are that more money and time and more expertise and data are needed to plan and open new coal mines today. It is also becoming apparent that the old methods of assimilating and preparing information needed for mine planning are inadequate to meet the needs of today's modern mining engineer, especially in the areas of data gathering, storage, retrieval, preparation, and interpretation. The system described here allows assimilation of old and new core data in a relatively quick, accurate, and efficient manner. The system is a set of computer assisted procedures consisting of the following four basic steps: 1) raw data collection, computer input, storage, and graphic preparation, 2) geologic profile correlation and establishment of seam continuity, 3) computerized data posting of important seam parameters, 4) geologic and planimetric analysis for mine planning.

Jan 1, 1977

By J. H. Swisher

The solubility of sulfur and rate of solution of sulfur in pure Lron were measured in H2S + H2 and H2S + H2 H2O gas mixtures. The solubility and diffusivity of sulfur at 1410°Care 0.13 pet S and 1.0 x 10-5 sq cm per sec, respectively. The solubility iS the same, but the rate of sulfurization is slower in the presence of H2O in the reacting gas. Under these conditions, the over-all rate is controlled jointly by a slow surface reaction and by solid-state diffusion; the mechanism for the surface reaction has not been identified. KNOWLEDGE of the behavior of sulfur in solid iron is desirable for the metallurgy of such products as free machining steel, where a high sulfur level is required, and inclusion-free high-strength steels, where the sulfur specifications are very low. The present investigation was undertaken to check previously reported values for sulfur solubility and diffusivity in 6 iron, and to study the poisoning effect of chemisorbed oxygen on sulfurization kinetics in H2-H2S-H2O gas mixtures. All of the experiments were performed at 1410°C. The thermodynamic behavior of sulfur in 6 iron was the subject of a paper by Rosenqvist and Dunicz.' The sulfur solubility at 1400" and 1500°C was determined by equilibrating pure iron specimens with H2-H2S gas mixtures. The maximum solubility of sulfur in 6 iron was alsc determined by Barloga, Bock, and parlee2 by reacting iron wires with sulfur in sealed capsules. In another investigation, the diffusion coefficient of sulfur in 6 iron at temperatures up to 1450°C was measured by Seibel.3 The method used was to measure sulfur concentration profiles in diffusion couples containing radioactive sulfur EXPERIMENTAL Apparatus. A vertical resistance furnace wound with molybdenum wire and containing a recrystallized alumina reaction rube was used for the experiments. The hot zone in the furnace was approximately 2 in. long with a temperature variation of ±3oC. The hot zone temperature was automatically controlled to within ±2°C, and the test temperature was measured with a pt/Pt-10 pet Rh thermocouple before and after each experiment. Flow rates of the reacting gases were obtained using capillary flow meters. Materials. The source of H2S in the gas train was a premixed cylinder containing 5 pet H2S in H2. This mixture then was diluted with additional hydrogen and argon. In some experiments, water vapor was introduced by passing hydrogen and argon through a column containing 10 pet anhydrous oxalic acid and 90 pet oxalic acid dihydrate. The vapor pressure of water above this mixture is well-known.4 Argon was used as a diluent to minimize thermal segregation of H2S in the furnace5 and to reach higher H2O:H2 ratios than could be obtained in mixtures of H2 and H2S alone. Argon was purified by passage over copper chips at 350°C and subsequently over anhydrone. Hydrogen was purified by passage over platinized asbestos at 450°C and then over anhydrone. The H2-H2S mixture was purified by passage over platinized asbestos and then over P2O5. The specimen stock was made by melting and vacuum-carbon deoxidizing electrolytic "Plastiron" in a zirconia crucible. The principal impurities are listed in Table I. In some of the equilibrium experiments, six-pass zone-refined iron was used to minimize impurity side effects. This zone-refined iron had a total impurity level of about 25 ppm. Procedure. Specimens were annealed in hydrogen for a period of at least 2 hr at the beginning of each experiment. The specimens were held in the reacting gas for times varying between 10 min and 17 hr, and cooled to room temperature in a water-cooled stainless-steel block at the bottom of the furnace. The pH2S/pH2 ratios reported are those for gas equilibrium at 1410°C. Calculations based on available thermodynamic data8 showed that the only other gaseous8 species that formed in significant amounts in the furnace were S2 and S. Even when water vapor was introduced into the gas mixture, the concentrations of SO2, SO, and so forth, were negligible. The initial partial pressure of H2S was therefore corrected for its partial dissociation to S2 and S in determining the equi-

Jan 1, 1968

By L. S. Richardson, N. J. Grant

Iodide-grade titanium, two oxygen alloys, and two aluminum alloys were studied by means of creep-rupture tests from 1000° to 2000°F. From the test information an evaluation was made of, 1) the relative creep-rupture strengths of the u (hexagonal-close-packed) and 0 (body-centered-cubic) forms of titanium and some of its alloys; 2) the effects of solid-solution strengthening by oxygen and aluminum on the u and 0 structures; and 3) the extent and effect of atmospheric oxygen alloying on rupture life in tests above about 1000°F. THE use of titanium alloys at temperatures above 1000° F has been considered unlikely by most investigators because of low creep strength (based on a requirement for room-temperature ductility, formability, and weldability) and low oxidation resistance. The results of this testing program do not entirely uphold this pessimism in terms of strength and oxidation resistance. The observations described below are based entirely on stress-rupture results, although some creep tests were also carried out on these alloys. Three items of primary importance are covered: 1) Stress-rupture data in the form of the Larson-Miller parameter curves for several titanium alloys at temperatures up to 2000°F; 2) The relative strengths of the hexagonal-close-packed a structure and the body-centered-cubic 0 structure; 3) The effects of solid-solution alloying on the creep-rupture properties of titaniu'm at elevated temperatures. Only two investigatons have reported the creep- rupture properties of titanium at temperatures above 1000° F. Lunsford and rant' tested iodide titanium; their values are also utilized in the paper. EXPERIMENTAL PROCEDURE After a few tests had been made in a conventional creep-rupture test machine, it became clear that some sort of atmospheric protection would be necessary for short-time tests conducted at temperatures above 1600°F, or for times greater than 100 hr at 1200°F or above. Several furnace units and test machines were constructed, the most successful of which is shown in Fig. 1. Certain experimental difficulties peculiar to titanium were encountered; for example, the use of molybdenum holders was dictated because titanium forms low melting-point eutectics with conventional high-temperature alloys containing appreciable amounts of iron, nickel, or cobalt. This same difficulty made the use of chrome l-alumel thermocouples in direct contact with a test specimen impossible, above about 1500" F. The high-temperature high-vacuum furnace shown in Fig. 1 was capable of maintaining a vacuum of better than 10"5 mm of mercury at 2000°F (pressure measured in cold part of system) and 5 X lo-' mm at room temperature. Most of the tests eonducted in vacuum were at temperatures

Jan 1, 1960

By John B. Malcom

The discovery of tungsten-bearing quartz outcrop- pings in Vance County, N. C., by the Hamme brothers in 1942, made international news in the mining industry and was, in fact, one of the major ore discoveries made in the U.S. during the last 25 years. Operations began at the Hamme mine in 1942, but after 16 years of successful operation, the mine was closed down in 1958 under adverse market conditions. By that time, it had produced nearly 20 million lbs of metallic tungsten. Because of the square-set timber mining method then in use, mine operating expenses were high, actually representing 78% of the total operating expenses at the property. In timber work alone, 160 men were employed. Approximately 500,000 bd ft of timber were placed monthly at a direct supply cost of $1.00 per ton of ore mined.

Jan 4, 1962

By L. P. Teas

ALTHOUGH 1927 in the Arkansas-Louisiana territory will probably pass into history as a year of small profits and little new production, nevertheless during this period two large gas fields were developed; two new oil fields were discovered, and two new oil horizons, hitherto of minor Annual Production in Millions of Barrels FIG. 1.-PRODUCTION OF NORTH LOUISIANA AND ARKANSAS BY FIELDS FOR 1926 AND 1927. importance, were developed; In spite of the "hard times" some of the leading, supply houses did more business this year in this. district than in any year since 1920.

Jan 1, 1928

By C. T. Liu, J. Gurland

The deformation behavior in tension of spheroidized carbon steels was studied at room temperature as a function of carbon content, 0.065 to 1.46 wt Pct, and carbide particle size, 0.88 to 2.77 p. It was found that the Hall-Petch strength-grain size relation is directly applicable to the yield and flow stresses of the two lower-carbon steels , 0.065 and 0.30 pct C. The strength data for the medium- and high-carbon steels, 0.55 to 1.46 pct C, also satisfied the Hall-Petch relation, provided that these data are based upon the particle spacing. Beyond 4 pct strain, the flow stress data of all the steels studied could be represented by the same Hall-Petch relation with dinerent spacings for grain boundary and particle strengthening. The behavior of the higher-carbon steels was consistent with the postulated formation of a dislocation cell network during processing and initial deformation (up to 4 pct strain). The cell size was assumed to be equal to the planar particle spacing. The true stress at the ultimate tensile strength was also found to be a function of the particle spacing. At a given temperature and strain rate, the yield and flow stresses of carbon steels depend on the type and dimensions of the microstructure. Starting with the work of Gensamer et al. in 1942,' experimental studies on pearlitic and spheroidized carbon steels revealed that the strength of steels is a function of two main parameters: the ferrite grain size2'3 and the carbide particle spacing;1'4'5 on this basis, two different strengthening mechanisms have been developed to apply to steels of low and high carbon contents, respectively. In polycrystalline iron and mild steels the grain boundaries are regarded as the major structural barriers to slip. The relation between strength and grain size is generally represented by the Hall-Petch equation which is based on a linear proportionality between strength and the inverse square root of the average grain size.2'3y677 However, Gensamer et al.' and Roberts et related the yield strength of medium -and high-carbon steels to the carbide particle spacing alone, and they found a linear relation between the logarithm of the mean free path in the ferrite and the yield strength in both spheroidized and pearlitic steels. By means of the electron microscope, Turkalo and LOW' extended the study to finer structures; they concluded that the logarithmic relation is not valid for the entire range of microstructures unless grain boundaries are also included in the measurement of the mean free path. For the specific case of spheroidized steels, Ansell and aenel' found that the yield strength data,4'5 when plotted as a function of mean free path, fit the Hall-Petch equation; however, T'ysong found that the same data fit the 0rowanl0 relation if a planar inter-particle spacing is used. Recently Kossowsky and ~rown" studied the strength of prestrained spheroidized steels, 0.48 and 0.95 pct C, and concluded that the strength due to the carbide dispersions varies linearly with the reciprocal of the square root of the mean free path between carbide particles and dislocation networks. Such networks were first observed by Turkalo." The conclusion common to all these studies is that the available slip distance in the ferrite is the most important variable in determining strendh. Previous work on carbon steels is restricted to limited composition and strain ranges. The mechanism which governs the flow properties is not clearly understood, and, in particular, little is known about the composition dependence of the transition between grain boundary strengthening and particle hardening. The purpose of the present work is to investigate the strengthening mechanism in spheroidized steels over a wide range of carbon content, 0.065 to 1.46 wt pct, and plastic strain, yielding to necking. The spheroidized structure was chosen because of its relative simplicity and the relative ease of control and measurement of the structural parameters. The experimental work is limited to tensile testing at room temperature at constant extension rate. The effects of the carbide particles on the fracture behavior of spheroidized steels are discussed elsewhere.13 EXPERIMENTAL PROCEDURE Eight different grades of vacuum-cast carbon steels were supplied in the form of forged and rolled plate by the Applied Research Laboratory of the U.S. Steel Corp. The compositions furnished with these steels are given in Table I; the carbon content ranges from 0.065 to 1.46 wt pct, or from 1.0 to 22.3 vol pct of carbide. The steel plates were cut transversely into rods a little larger than the test specimens, 1 in. gage length, i in. diam. The rods were austenitized in air (enriched with CO by a consumable carbon-rich muffle) at 50° C above theA, orA., temperature for 2 hr and then quenched in oil with vigorous stirring. The as-quenched rods were tempered in two stages in order to obtain the desired distributions and sizes of carbide particles. The rods were first tempered at 460° C for 10 hr and then at 700" C for periods ranging from 4 hr to 3 days, in vacuum. After final machining, all specimens were vacuum-annealed again at 650°C for 1 hr in order to relieve residual stresses. The tension tests were carried out in two steps. The initial part of the load-strain curve, up to about 2 pct strain, was determined on a Riehle testing machine with an extensometer of small strain range, 4 pct strain, in order to obtain the yield and initial flow piopertiesi As soon as the first part of the test was finished, the specimen was placed in an Instron testing machine equipped with a strain gage extensometer with a maximum strain range of 50 pct. The load-strain curve to fracture was

Jan 1, 1969

By C. F. Cockrell, K. K. Humphreys, H. E. Shajer

It is quite conceivable that the question of using coal or an alternate fuel in the future by large industrial consumers may be decided on the basis of possible profitable utilization of flyash. Accordingly, the Coal Research Bureau at West Virginia University has been attempting to develop new uses for this waste material. This effort has resulted in the development of a process for producing flyash-based face brick and other structural materials which generally exceed published specifications for competitive clay-based materials. These structural materials, containing up to 97% waste coal ash, can be produced from flyash derived from most coals without regard to the physical or chemical properties of the coal ash. Preliminary production cost estimates indicate that face brick can be manufactured at a rate of 9000 brick per hr or more with a resultant gross profit of over $24.00 per thousand brick. Plant payout time for a plant of this size would be only about 1.2 years assuming a market price of approximately $55.00 per thousand, a well established figure for comparable clay brick. With flyash production expected to more than double in the next 15 years, there is an increasing need for a practical and Profitable method for the utilization of large tonnages of this material (see Table I listing total flyash production of utilities in each state of the U.S.). Present industrial and governmental physical and chemical restrictions, particularly for use as a lightweight filler or pozzolanic material, have limited the use of flyash considerably. For those ashes which have met the required standards, utilization is not always assured because of lack of promotion and poor public acceptance. Less than 6% of the total flyash production in the U.S. was sold in 1962. The remainder was dumped in disposal areas at costs as high as $2.42 per ton.' It is quite conceivable that in the near future the question of using coal or an alternate fuel by large industrial consumers may be decided on the basis of the possible utilization of flyash. To help insure future coal demands, the Coal Research Bureau at

Jan 1, 1968

By R. P. Rothwell

Jan 1, 1885

By Zay Jeffries

AMONG the various metals with low specific gravity, two, aluminum and magnesium, are used in the nearly pure state or as bases for light-weight alloys. Although the production of magnesium is only about one-twenty-fifth that of aluminum, a higher percentage of its output is used in the transportation industry than that of any other metal. The per-centage varies from year to year but, on the average, more than three-fourths of the magnesium production has been used for this purpose. In the United States, about 40 per cent of the aluminum production is used for transportation purposes. It seems obvious that equipment demanding repeated movement with need for rapid acceleration and deceleration should be as light in weight as the economics of engineering materials will permit. There are, however, differences of opinion with reference to both the economics and the engineering aspects of this subject. Tables 1 to 7 inclusive relate to the principal light-metal materials currently used in the transportation industry. These, with the accom-panying titles and footnotes, are in the main self-explanatory. Although the variations in mechanical properties can be easily seen from the tables, it should be kept in mind that the values given are for ordinary tempera-tures and that the changes in properties with temperature are not the same for the different alloys. Also, the corrosion properties cannot be deduced accurately either from the compositions or from the physical properties. Again, certain of these alloys are used because they possess desirable combinations of properties not listed in tine tables, including heat conductivity, coefficient of expansion, machinability, ease of casting or working, light reflectivity, anodic coating characteristics, etc. These, together with variations in prices, must be given careful analysis to effec-tively utilize the light alloys. Some of the alloys listed in the tables are new and some are old, but all are standard. Even the older alloys are available in new shapes and larger sizes than heretofore, and all are used in the transportation industry in structures of new design. Certain of the applications of specific alloys will be touched upon in the following brief history of the use of the light metals in transportation.

Jan 1, 1936

By A. L. Marshall, F. J Norton

The object of this investigation was to make a comprehensive study of the degassing of molybdenum in order to determine how rigorous a treatment was necessary to completely remove sorbed gases from molybdenum electrodes in vacuum tubes. Some work has also been done with tungsten, nickel, iron and carbon. As a logical development of this work it has been shown that the gas is present throughout the body of the metal and a study has been made of the solubility of nitrogen in molybdenum and tungsten and the rate, of diffusion of nitrogen through molybdenum has been calculated. Tests were made on molybdenum from all the commercial sources and no difference was discovered in the nature of the gases present, the amounts or the ease of removal. The reason for this was apparent when a study was made of the manner in which gases are absorbed by gas-free molybdenum. It is necessary to heat molybdenum to 1760°C. in a vacuum of the order of 0.001 micron for a time which varies lineally with the thickness, in order to obtain a condition in which no further gas is evolved by the sample. A sample so degassed can be handled and subsequently assembled in a tube and then degassed readily. The gases obtained from molybdenum are carbon monoxide and nitrogen, the nitrogen being the more difficult to remove. Apparatus and Technique The apparatus is shown in Figs. I and 2. It consisted of two units, a degassing system and an analytical system. Each had a bake-out oven, as shown surrounding the systems in the sketches. The system , will be traced, starting from the sample in the degassing system. A water-cooled quartz chamber surrounded the sample. The high-frequency coil, which heated the sample, extended only around this chamber and was cooled by an air blast. High-frequency power of 1800-meter wave length was obtained from a self-rectifying oscillator circuit with a maximum power input of 20 kw. A graded seal connected the quartz to the rest of the apparatus, which was of hard glass. The temperature of the molybdenum sample was observed by an optical pyrometer through an optical window about 40 cm. distant from the sample. Since the object viewed under these conditions was not a black body, true temperature was not observed, but the so-called "brightness temperature." The correction of this to true temperature is discussed in a later section. The apparatus was taken apart just above the graded seal; the quartz chamber was slipped off, washed with nitric acid, water and alcohol, and dried by an air stream. The old sample was removed by unscrewing the lower nut, and the new

Jan 1, 1944

By A. L. Marshall, F. J. Norton

The object of this investigation was to make a comprehensive study of the degassing of molybdenum in order to determine how rigorous a treatment was necessary to completely remove sorbed gases from molybdenum electrodes in vacuum tubes. Some work has also been done with tungsten, nickel, iron and carbon. As a logical development of this work it has been shown that the gas is present throughout the body of the metal and a study has been made of the solubility of nitrogen in molybdenum and tungsten and the rate, of diffusion of nitrogen through molybdenum has been calculated. Tests were made on molybdenum from all the commercial sources and no difference was discovered in the nature of the gases present, the amounts or the ease of removal. The reason for this was apparent when a study was made of the manner in which gases are absorbed by gas-free molybdenum. It is necessary to heat molybdenum to 1760°C. in a vacuum of the order of 0.001 micron for a time which varies lineally with the thickness, in order to obtain a condition in which no further gas is evolved by the sample. A sample so degassed can be handled and subsequently assembled in a tube and then degassed readily. The gases obtained from molybdenum are carbon monoxide and nitrogen, the nitrogen being the more difficult to remove. Apparatus and Technique The apparatus is shown in Figs. I and 2. It consisted of two units, a degassing system and an analytical system. Each had a bake-out oven, as shown surrounding the systems in the sketches. The system , will be traced, starting from the sample in the degassing system. A water-cooled quartz chamber surrounded the sample. The high-frequency coil, which heated the sample, extended only around this chamber and was cooled by an air blast. High-frequency power of 1800-meter wave length was obtained from a self-rectifying oscillator circuit with a maximum power input of 20 kw. A graded seal connected the quartz to the rest of the apparatus, which was of hard glass. The temperature of the molybdenum sample was observed by an optical pyrometer through an optical window about 40 cm. distant from the sample. Since the object viewed under these conditions was not a black body, true temperature was not observed, but the so-called "brightness temperature." The correction of this to true temperature is discussed in a later section. The apparatus was taken apart just above the graded seal; the quartz chamber was slipped off, washed with nitric acid, water and alcohol, and dried by an air stream. The old sample was removed by unscrewing the lower nut, and the new

Jan 1, 1944

By A. M. Skov, L. F. Elkins

First response to large-scale water flooding in the fractured very low permeability Spraberry sand has led to a new unique cyclic operation. Capacity water injection is used to restore reservoir pressure. This is followed by marly months production without water irzjection and the cycle repeated. Expansion of the oil, rock and water during pressure decline expels part of the fluids but capillary forces hold much of the injected water in the rock. At least with reservoir pressure restored and with partial water flood development, field performance has proved this cyclic operation is capable of producing oil from the nzatrix rock at least 50 per cent faster and with lower water percentage than is imbibition of water at stable reservoir pressure. INTRODUCTION The Spraberry Field of West Texas presents unusual problems for both primary production and water flooding. Extensive interconnected vertical fractures in the fractional-md sandstone permitted recovery of oil on 160-acre well spacing, but they made capillary end effects dominant. Primary recovery by solution gas drive is less than 10 per cent of oil in place. The concept of displacement of oil from the sand matrix by capillary irnbibition of water has led to field techniques which promise greatly increased oil recovery. Free exchange of laboratory research, reservoir information and results of field pilot tests among the various companies has been very important in development of this technology. Five units covering a total of 170,000 acres have been formed for water flooding, and 10 other areas covering an additional 175,500 acres are in various stages of unitization. Part of the Driver Unit reaching fillup first has demonstrated very unusual waterflood behavior and indicated numerous operating problems that will develop within and among the various units. SPRABERRY ROCK AND PRIMARY PERFORMANCE The Spraberry, discovered in February, 1949, is a 1,000-ft section of sandstones, shales and limestones with two main oil productive members: a 10-15 ft sand near the top and a 10-15 ft sand near the base. In part of the field some thinner intermediate sands are oil productive, and others are water bearing. All sands have permeabilities of 1 md or less and porosities of 8-15 per cent. Ordinary core analysis and electric and radiation logs are ineffective in differentiating between oil productive and nonprcductive sands. Sands capable of containing producible oil are best identified by mercury injection capillary pressure measurement and, in some cases, by core water saturation. About 3,500 wells have been drilled in the 500,000-acre trend. Vertical fractures were observed in practically all Spraberry cores. Continuity and interconnection of fractures were confirmed by pressure interference among wells during early development.' Major fractures trend northeast-southwest as indicated by oriented cores and confirmed by five fluid injection tests, by analysis of the pressure transients observed during development,''' and by three interference tests in the Driver Unit Water Flood reported herein. Fracture spac- ing probably averages inches to a few feet. Spraberry wells typically produced 100-400 BOPD initially after hydrauLic fracture treatments. By 1962 oil production had declined to an average of 12 bbl/well/day, near the economic Limits of operation. Reservoir pressure had declined from 2,300 psi initially in the Upper Spraberry and 2,500 psi in the Lower Spraberry to 500-1,000 psi. Partial closing of the fractures with declining reservoir pressure is believed to be the cause of such low oil production rates at these relatively high reservoir pressures. Cumulative recovery of 208 million bbl of oil is 80 to 90 per cent of that recoverable by primary means. Performance of the entire reservoir is summarized in Fig. 1. IMBIBITION WATER FLOODING By 1952 reservoir performance indicated low primary recoveries. Most engineers, expecting serious channeling of injected fluids through the fractures, held little hope for secondary recovery. With its extensive background of research on the fundamentals of fluid flow within reservoir rocks, Atlantic's Research and Development Division on short notice in 1952 conceived that displacement of oil by capillary imbibition of water into the rock might significantly increase Spraberry recovery. Laboratory data reported by Brownscombe and Dyes scaled to probable reservoir conditions showed potential waterflood recovery equal to or greater than primary recovery with a 10-15 year flood life.= A pilot test using three 40-acre injection wells, one central producing well and 18 surrounding observation wells demonstrated technical feasibility of the process. Injection of 1.5 million bbl of water from November 1952 to August 1955 proved water entered the rock and displaced oil

By W. J. Helfrich, R. A. Dodd

Binary aluminum solid-solution alloys containing various amounts of silver, magnesium, and zinc were prepared by careful directional solidification, and the hydrostatic and X-ray densities were compared. With the exception of the Al(Ag) alloys, the X-ray densities were consistently greater than the hydrostatic measurements, in agreement with earlier observations by Ellwood. In contrast to Ell-wood's interpretation in terms of vacant lattice sites associated with Brillouin zone effects, a tentative explanation based on the existence of solidification microshrinkage was favored. This hypothesis was confirmed by an examination of Al(Zn) alloys prepared by vapor diffusion of zinc into aluminum. The hydrostatic and X-ray densities were now in very close agreement, and it was concluded that the filling of Brillouin zones in aluminum solid-solution alloys does not necessarily result in the formation of defect structures containing an excess of vacant lattice sites. ThE existence of defect structures of the vacancy type in alloys in which the excess vacancies have an electronic rather than a thermal or mechanical, and so forth, origin is well recognized. Examples of incomplete lattices of this type are to be found in the Ni-Al,1-3 Fe-Ni-A1,4 c~-Ni-Al,5 Fe-Cu-Al,= and Co-A17 systems. These defect structures are of a special kind in that the intermediate phases possess an ordered atomic arrangement or superlattice, and in some instances the vacancy concentration may be unusually large, e.g., at 45.25 at. pet Ni in NiA1, approximately 8.8 pet of the lattice sites are unoccupied. Ellwood8-10 has reported similar defect structures in the aluminum solid solution alloys of the Al-Zn and A1-Mg systems and in alloys of the Au-Ni system." In Al(Zn) the (apparent) vacancy concentration rose, somewhat irregularly, to a maximum of about 2 pet vacant sites at 25 at. pet Zn, while in Al(Mg) the (apparent) vacancy concentration increased continuously to 1.7 pet at 15 at. pet Mg. An explanation in terms of Brillouin zone overlap was attempted, although Pearson12 has pointed out the difficulty of reconciling the observations with zone theory. However, the possibility of the effect being caused by the Fermi surface just touching a plane of energy discontinuity inside a prominent Brillouin zone has, in general, been accepted. In fact, Massal-ski13 has interpreted Ellwood's8 observations as confirmation of Leigh's14 theoretically predicted zone overlap occurring at approximately 2.67 electrons per atom. Unfortunately, Massalski was apparently unaware that Ellwood9 had revised his earlier results considerably, and the revised data did not confirm Leigh's analysis. Ellwood's clata were reexamined by the present authors who noted a possible correlation between the percentage defects as a function of alloy composition and the temperature interval of solidification measured from the respective equilibrium diagrams. This suggested an explanation in terms of shrinkage porosity rather than vacant lattice sites, and pointed to the desirability of reexamining appropriate alloy systems using: both Ellwood's method of specimen preparation (casting followed by wrought fabrication) and alternativ'e methods, i.e., diffusion, which might be expected to minimize, or even completely obviate, microporosity. ALLOY PREPARATION 1) Cast Allolys and Aluminum Single Crystals. Al(Ag), Al(Mg;l, and Al(Zn) alloys of various compositions up to 20 at. pet silver, 13.5 at. pet mg, and 30 at. pet Zn were prepared by melting under helium and casting into graphite molds. In the first two systems, the maximum alloying addition was quite close to the limit of solid solubility, but the possibility of transformation to a' during quenching somewhat restricted the suitable Al(Zn) composition range. The alloys were prepared from high-purity aluminum, a lot analysis showing 0.002 wt pet Cu, 0.002 wt pet Fe, and 99.996 wt pet A1 by difference. The silver, magnesium, and zinc were of 99.99+, 99.98+, and 99.998 wt pet respectively. Each composition was analyzed chemically. The as-cast ingots measured 7/16 in. diam and 5 in. length. One in. was removed from the top of the ingot, and the bottom 3 in. was machined to 0.275 in. diam; a point was also machined on the smaller diameter end. The remainder of the original ingot served as a top riser during subsequent remelting and controlled solidification. The machined ingots were now remelted using a Bridgman soft-mold technique to ensure directional solidification and, therefore, a minimum of micro-shrinkage. Alumina powder was used as mold material contained in an alundum thimble, and this crucible was placed in a helium-filled Vycor tube. The assembly was lowered through a suitable temperature gradient at approximately 0.5 in. min-l, and the risered portion of the casting was subsequently removed by sawing.

Jan 1, 1962

By G. S. Rice

Some form of longwall mining is generally used in Continental Europe; also in Great Britain where the coal is weak and friable, or the coal bed provides material for pack walls and filling, or where the bottom is soft and squeezes up easily, or the roof is pliable, or the bed is thin and brushing provides building material, or the thick multiple seams are mined in layers, such as the 24-ft.seam at Weymss, Scotland, or the 10-yd. pitching bed near Coventry, England. Pillar methods have been retained in British fields where the coal beds are from 5 to 9 ft. (1.5 to 2.7 m.) thick, and free from thick partings or binders or without draw slate, which would provide waste rock for pack walls, or where the roof is hard and requires shooting to bring down, and also the bottom or floor is comparatively hard. The room-and-pillar system, by which probably 95 per cent. of the coal of the United States is produced, is now found in only a few mines in Wales, where it is known as the pillar and (single) stall method. The American room-and-pillar system is not equivalent to the bord-and-pillar or stoop-and-room system. The essential difference between the bord-and-pillar and the average American room-and-pillar system is probably due to natural conditions. The majority of the coal mines in the United States have a depth of less than 700 ft. (213 m.). The bulk of the coal mined in Europe comes from a depth of more than 1200 ft., while the deepest mines in Great Britain, Belgium, and France reach 3500 to 4000 ft. In deep workings wide spans are impracticable, even with the use of abundant timber; accordingly, as the mines in the United States become deeper the rooms will become narrower and the pillars thicker, and the bulk of the coal will be extracted on the retreat. It is generally conceded, without reference to the cost of production, that the larger the pillars left on the first mining, the more thoroughly can the coal be extracted. In Europe complete extraction is generally compelled either by the lessors or by the governments; whereas in this country it has been more largely a question of competitive cost of production, or the support of the surface in the flat farming districts of the

Jan 1, 1922

By Granville S. Borden

The fruits of industry are divided between capital, labor, and governments. Capital takes its redemption and remuneration through profits or dividends; labor takes its share through wages; governments take their share through taxes. Each must have a just share, and the question of what is a just share and the question of who-shall make the allocation are vital current political and economic issues. Certain political philosophies con- tend that government should take all the fruits and that some dictator or politician should have irrevocable discretion and power to distribute the fruits. Believers in democracy and free enterprise adhere to the view that capital is entitled to redemption and fair remuneration; that labor shall have a fair wage; and that governments shall make equitable levies. But since all human beings, both individually and in groups, are inherently selfish, there are many variable views in regard to the question of what ratios shall be applied in a division of the fruits and who shall have the power to fix these ratios. Perhaps all parties should devote more energy to means of increasing the crop, and not so much time to quarreling over a division of the existing crop. Moreover, all three parties must guard against killing the goose that lays the mineral eggs. For example, if the tax burden becomes oppressive and the taxpayer refuses to carry it, then business stagnates, with consequent unemployment, reduction of profits, and curtailment of government revenue. If a producer assumes the increased tax burden, marginal operations must be closed down, and again the consequences are increased unemployment and diminished government revenue. Quite often an increase in a rate of tax results in a reduction of government revenue. Perhaps the worst detrimental feature of selfishness and greed on the part of any one of the three distributees of the fruits of industry is the retardation of expansion of the enterprise. Growth means new jobs, more government revenue, better trade, and better standards of living. Explorations and discoveries have resulted in tremendous contributions

Jan 1, 1959

YUGOSLAVIA is one of the smaller countries in Europe-about twice the size of Pennsylvania -but one of the richest nations in variety and quantity of minerals on the Continent. The country, just now beginning the pangs of a real industrial birth, has a mining history that goes back as far as the sixth century, B. C. Among others, the Romans exploited Yugoslavia's lead and zinc deposits, and there was intensive mining during the Middle Ages. But until now, the country's industrial development was such that it made no great demand on the mining industry for development or expansion. Conversations with Yugoslavian officials in New York give the impression that they are interested in two things. They need money and they need trained engineers to solve their mining problems. Yugoslavia appears to be attempting to attract both foreign money and foreign technical skill. They want financial aid in the form of loans. The U. S. Government has given aid to Yugoslavia in the shape of armaments, foodstuffs, and equipment. Recently, American banks have advanced funds.

Jan 5, 1954