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By R. W. M. Lai

INTRODUCTION Manganese nodules were first discovered by the H.M.S. CHALLENGER expedition of 1873-76 and have engaged the interest of oceanographers and natural resources engineers ever since (1-4). Manganese nodules, some of them as large as grapefruit, cover vast stretches of the ocean floor throughout many parts of the world. In some areas the ocean floor is literally paved with nodules. The Pacific Ocean alone is estimated to contain 1.5 trillion tons of nodules which are forming at the rate of about 10 million tons per year (2). An analysis of the principal metal values for Pacific Ocean nodules is shown in Table I (5). [ ] The principal ingredients of the nodules are manganese and iron, but they also contain other important associated metal values such as copper, nickel, cobalt, and molybdenum. These associated metal values,

Jan 1, 1980

By Klaus Schwerdtfeger

The rules of solution of oxygen from H2O-H2-He gas and of nitrogen from N2-H2 gas in shallow melts of liquid iron were measured at 1610o and 1600o C, respectiuely. Concentration profiles were detemined in the liquid iron. Tire rate data indicate that the solution process is controlled by diffusion in the iron melt. The diffusivities for oxygen and nitrogen in liquid iron, as calculated from the present data, are DFe-o = (12 ± 3) < 10-5 sq cm per sec and DFe-N = 11 ± 2) X 10-5 sq cm per sec at the temperatures employed. AN attempt was made by Shurygin and Kryukl to measure the diffusivity of oxygen in liquid iron. In their experiments a silica disc was rotated in liquid iron containing oxygen, and the rate of formation of liquid iron silicate was measured. Assuming that the rate of dissolution of silica is controlled by diffusion of oxygen in the iron, the oxygen diffusivity was computed from the rate data giving Dfe-0 = 6.1 X 5 sq cm per sec at 1600°C. Although this value seems to be of the right order of magnitude, there is no proof of the correctness of the assumptions involved in the interpretation of these rate data. The oxygen concentration in the iron at the iron-iron silicate interface was taken to be that in equilibrium with the silica-saturated silicate melt. That is, it was assumed that no concentration gradient existed in the liquid silicate. This is a questionable assumption, unless it is proved that the thickness of the silicate layer is very much smaller than that of the diffusion boundary layer in the iron. Furthermore, Shurygin et al.1 used the Levich equation2 to interpret their rate data. This equation was derived for mass transfer between a solid disc and a single-phase liquid. The hydrodynamic and diffusion boundary layers in the iron stirred by a disc, via coupling of the silicate melt, may be appreciably different from those predicted by Levich&apos;s derivations. In the present work the diffusivities of oxygen and nitrogen in liquid iron were measured at 1610" and 1600oC, respectively. EXPERIMENTAL METHOD Iron melts contained in high-purity gas-tight alumina crucibles were reacted with H2O-H2-He gas for the determination of the oxygen diffusivity and with N2-H2 gas for the determination of nitrogen diffusivity. At the end of the reaction period, the samples were quenched in a cold H2-He gas stream at the top of the furnace. Oxygen or nitrogen contents in the iron were determined by chemical analysis. Two different types of diffusion experiments were perforxed. To determine concentration profiles, a few rate measurements were made using 4-cm-deep melts. The solidified samples were sliced into discs and each disc was analyzed for oxygen or nitrogen. In another series of experiments, oxygen or nitrogen was diffused into shallow melts (about 0.5 to 1 cm in depth) and the total sample was analyzed to obtain an average concentration of the diffusate. In most experiments, 4- to 5-mm-ID alumina crucibles were used. Some experiments were also made in smaller (3 mm) and larger (7 mm) diam crucibles. This variation in diameter caused no difference in the reaction rate, within the limits of experimental uncertainty. To promote the establishment of a stable density profile in the melt, all the samples were suspended in the lower end of the hot zone so that the top of the melt was hotter by a few degrees. Molybdenum wire resistance heating was used. The reaction tube of the furnace was a gas-tight recrystal-lized alumina tube. In most experiments the furnace was heated by an ac power supply. To check the possibility of inductive stirring, some experiments were carried out in a dc operated furnace, with essentially the same results. The temperature of the furnace was controlled automatically in the usual manner. The temperature was measured with a Pt/Pt-10 pet Rh thermocouple and is estimated to be accurate within ±5°C. The iron used was prepared by melting and vacuum-carbon deoxidizing electrolytic "Plastiron" in a zir-conia crucible. The main impurities are: Si 0.004 pct P, S <0.002 pct Cr 0.005 pct N 0.001 pct Zr 0.002 pct O 0.003 pct Mn 0.004 pct C 0.002 pct The gas composition was controlled by constant pressure head capillary flowmeters. Oxygen was removed from the gas mixture by passing it through columns of platinized asbestos (450°C) and anhydrone. Selected H2O contents were obtained by passing the purified gas through oxalic acid dihydrate-anhydrous oxalic acid mixtures held at constant temperature in a water bath. Water vapor pressure data for the oxalic acid dihydrate-anhydrous oxalic acid equilibrium were taken from the 1iterature.3 The flow rate used was about 1.5 liters per min. The whole system was checked for tightness at regular intervals.

Jan 1, 1968

By R. E. Ogilvie, A. R. Kaufmann, S. H. Gelles

The solid-solubility limits of iron in beryllium were determined between 850o and 1200oC by analysis of differential type multiphase diffusion couples, using an X-ray absorption technique. The maximum value of the solubility limit was found to be 0.92 ± 0.02 at. pct (5.46 wt pet) at the eutectic temperature 1225°C. The solubilities of nickel and beryllium were determined between 900°and 1200°C by the same technique and the maximum solubility was found to be 4.93 + 0.01 at. pct (25.2 wt pet) at the eutectoid temperature, 1065°C. A previously unreported high-temperature phase which decomposes eutectoidally at 1065 °C was found to exist in the beryllium-nickel system at a composition of approximately 8 at. pct Ni (36 wt pet) by diffision-couple analysis. The presence of this phase was confirmed by thermal analysis and metallo-graphic analysis of the structure resulting from the eutectoid decomposition. G. V. Raynor1 has treated the solid solubilities of some of the elements in beryllium on the basis of the "Hume-Rothery" rules2 which have been modified to include ionic size and ionic distortion effects. It was predicted that the solubility of iron and nickel in beryllium should be slightly less than that of copper. The lowering of the solubility, according to Raynor, is due to a more unfavorable relative valency effect and an ionic size effect. Kaufmann and corzine3 have compiled data on the solubilities of elements in beryllium and have discussed them in the light of the Raynor paper. These authors feel that, because the elements having the greatest solubility in beryllium systematically fall in the Group VIII and IB Columns of the periodic table, the electronic structure greatly influences the maximum solid solubility of elements in beryllium. The solubility of iron in beryllium was determined by Teitel and cohen4 as part of the study of the beryllium-iron phase diagram. The determination was carried out by X-ray and thermal analysis and according to the phase diagram presented, the maximum solubility of iron in beryllium is 0.41 at. pct (2.5 wt pct) at 1225oC. However, it is estimated that the uncertainty in the position of the a-beryllium primary solid-solution boundary is about 0.5 at. pct (3wtpct). Losana and Goria3 in studying the beryllium-nickel phase diagram, determined the solid solubility of nickel in beryllium by thermal analysis. They found the maximum solubility to be between 1.65 and 2.65 at. pct (10 to 15 wt pct) at 1240°C. This value decreased rapidly with decreasing temperature. In determining approximate ranges of solubilities for different elements in beryllium, Kaufmann, et al,8 reported a value of between 1.3 and 1.7 at. pct (7.9 to 10.1 wt pct) for the solubility of nickel in beryllium. The value was obtained by metallographic examination of quenched alloys and lattice-parameter measurements. However, the authors also noted a single-phase structure for a 1.7 at. pct Ni alloy (10 wt pct) on cooling from the liquid. This would indicate a higher solubility range than was reported. ~isch,&apos; in his X-ray studies of beryllium-copper, beryllium-nickel, and beryllium-iron intermetallic compounds, reports the disappearance of a second phase (Ni,Be2) in the beryllium primary solid solution at approximately 4 at. pct (20 wt pct). THEORY The analysis of concentration gradients in diffusion couples has proven to be a useful tool in determining phase equilibria.8-14 In this particular study the diffusion couples were chosen to straddle the expected composition range of the phase boundary, then heat treated at a given temperature and the concentration gradient evaluated. The composition of the phase boundary for a given temperature appears at a point of discontinuity of the composition gradient. Examples of typical phase diagrams and the concentration gradients which should be found in such systems are shown in Fig. 1. In the present work, gradients of the form of Fig. l(c) were obtained in diffusion couples made of pure beryllium and two-phase alloys of beryllium with either iron or nickel. The composition at the point where the gradient becomes discontinuous, Cs, corresponds to the solubility limit of either iron or nickel in beryllium. The analysis of the concentration gradients was carried out by an X-ra absorption method developed and applied by Ogilvie and later used by Moll13 and Hilliard.l4 It depends on the fact that the absorption of X-rays by matter is determined by the concentration and type of the various atomic species present. The relationship for the intensity, I, of a monochro-

Jan 1, 1960

By Dorsey A. Lyon, Robert M. Keeney

Page I. Introduction..........234 I1. Chemistry oF Copper Smelting...235 1. Native Copper Ores.......235 2. Oxide and Carbonate Ores....235 3. Sulphide Ores.......236 (a) Elimination of sulphur ...236 III. Reverberatory Copper Smelting...237 IV. Substitution of the Electric Furnace for the Reverberatory Furnace..................... 238 V. Smelting of Fine Michigan Native-copper Concentrates in the Electric Furnace.................. 239 1. Conclusions.................240 VI. Proposed Process for Electric melting of Fine Native-copper Concentrates.................. 241 VII. OrdINary Copper Blast Furnace &apos;Smelting......... 245 1. Prerequisites.................. 245 2. Losses .................... 247 VIII. Experimental Data on the Electric Smelting of Raw Sulphide. Oxidized and Carbonate Ores............. 248 1. Vattier&apos;s experiment................ 248 2. Wolkoff&apos;s experiments................ 249 3. Experiments of Stephan.............. 250 (a) Conclusions of Stephan ............ 251 IX. Smelting of Sulphide Ores in the Electric Furnace...... 251 1. Ores Treated.................. 252 2. Conclusions................... 253 X. ELectric Furnace vs. the Reverberatory Furnace and Blast Furnace as a Melting Agent.............. 253 XI. Smelting of Sulphide Ores without Roasting or Partial Pyritic Smelting............254 XII. The Electric Blast Furnace vs. the Copper Blast Furnace for Partial Pyritic Smelting.......259 1. Metallurgical conditions......259 2. Mechanical considerations......259 3. Comparison of costs.......260 (a) General considerations .....260 (b) Cost of plant........260 (c) Amount of electrical energy...260 (d) Cost of smelting.......261 (e) Cost of electrodes........262 XIII. The Use of the Electric Furnace in the Smelting of PyRITIc ORES : 262 XIV. Summary...................... 264

Jan 1, 1914

By Herbert H. Kellogg

The chemistry of sulfide roasting is analyzed to show those aspects of performance which Thecan be predicted from considerations of thermodynamic equilibrium. It is concluded that equilibrium calculations can serve as a yardstick for prediction of roaster-gas composition and sulfate formation in the calcine for fluid-bed roasting practice. A method of calculation which combines the equilibrium and stoichiometric factors is outlined, and illustrated by an example from the roasting of Cu-Fe sulfides. ON first examination the roasting of sulfides would not seem to be a process to which equilibrium calculations could be applied to yield results of practical value. The burning of sulfides is an ex-tremely spontaneous reaction, excess air is always used to obtain rapid reaction, and in most traditional roaster designs (hearth roasting) the temperature and gas composition vary widely from point to point in the roaster. These are all conditions which promote a non-equilibrium state in the roasting process. However, the introduction of the fluid-bed roaster1 has brought to roasting practice new conditions of operation, so that equilibrium calculations can be used to predict certain limited aspects of practical operation. The fluid bed is characterized by uniformity of temperature, gas composition, and calcine composition from point to point throughout the bed. These relatively uniform properties of the bed make the meaningful calculation of the equilibrium state of certain reactions in roasting possible. It is the purpose of this paper to analyze roasting from an equilibrium viewpoint, to show where the method may be expected to yield valuable information, where it fails, and to outline methods of calculation. Main Roasting Reaction The reaction MS (s) + 3/2 0, = MO (s) + SO, [I] might well be called the main roasting reaction. It adequately accounts for the well known facts that metal sulfide and air are the main reactants, and metal oxide and SO, the main products. Deviations from this simple stoichiometry will be discussed under side reactions. For all metal sulfides, Reaction 1 is strongly exothermic and the equilibrium position is far to the right. Since the reaction is exothermic, the equilibrium shifts in the right-to-left direction as the temperature is raised. However, at all practical roaster temperatures (between 500" and 1200°C), the amount of this shift is small, and the equilibrium position remains far to the right. The main roasting reaction is essentially irreversible, and the application of equilibrium considerations to this reaction will not be! a fruitful field of investigation. A knowledge of the rate of this reaction. although not the proper subject of this paper, is of great importance to the roasting process. Analysis of the factors which influence this rate shows that: 1) The rate will increase rapidly with increased temperature. 2) The rate will be proportional to the partial pressure of oxygen at the surface of the reacting particle. Because of diffusion effects, the partial pressure of oxygen at the particle surface may be considerably less than in the bulk of the roaster gas. For this reason, intimate contact between gas and solid and high relative velocity between gas and solid, such as can be obtained in a fluid bed or flash roaster, will markedly increase the rate of reaction. 3) The rate will be proportional to the surface area of metal sulfide. Therefore, smaller particle size will permit a more rapid rate per unit of weight. To achieve practical rates of roasting the partial pressure of oxygen in the roaster gases must be maintained at a finite level at all times. Even where the gas-solid contact is very good, as in the fluid-bed roaster, the oxygen content of the gas must not be allowed to drop below some limiting value (perhaps 1 pct), or the rate of the main roasting reaction will become too small for practical operation. Side Reactions in the Gas Phase The gas phase during roasting of metallic sulfides contains the elements oxygen, sulfur, nitrogen, by-

Jan 1, 1957

By M. A. Grossman

THE hardenability of most steels can be predicted within 10 to 15 per cent provided the complete chemical composition is known, including "incidental" elements; and provided the as quenched grain size is known; and provided, finally, that the composition and heating temperatures for hardening are such as to result in austenite free from carbide particles. In method proposed herein, the steel is considered as having a base hardenability due to its carbon content alone (the hardenability of a "pure steel" of the given carbon content, without any other elements), and this base hardenability is multiplied by a multiplying factor for each chemical element present. After multiplying all these together, the final product is the hardenability. Grain size may be taken into account either in the base hardenability or after the multiplication. Hardenability is stated in terms of "ideal critical diameter"; namely, the diameter of bar, in inches, that will just harden all the way through (absence of unhardened core) in an "ideal" quench, and the calculation may also be related to the Jominy test. The data bring out certain features rather clearly. For example: I. It is quite useless to attempt to predict hardenability unless all elements, including "incidentals," are known; thus an "incidental" chromium content of o.-2o per cent increases the hardenability by about 50 per cent. 2. The relative effectiveness of different alloys is sometimes unexpected; thus molybdenum when calculated in this way appears to be of the same order of effectiveness as manganese, rather than much more powerful as would appear to be common experience; observe that an increase from no manganese to 0.20 per cent Mn provides a multiplying factor of 1.67, and an increase from no molybdenum to 0.2o per cent Mo provides a factor of 1.63, an increase of over 6o per cent in each case. However, if to a steel containing 0.50 per cent Mn there is added "20 points of manganese," the factor is raised from 2.65 (for 0.50 per cent Mn) to 3.35 (for 0.70 per cent Mn), or an increase of only 26 per cent. Thus the first small addition of an element has a much more powerful percentage effect than an equal further addition when some is already present, and in most cases the effect of molybdenum is considered in relation to a steel in which molybdenum is absent. 3. If two elements are equally effective, a greater harden ability will be obtained by using, for example, o.5 per cent of each than by using 1.0 per cent of either of them alone. 4. A knowledge of the as-quenched grain size is essential for precise work, since a difference of only one grain size number (say No. 7 instead of No. 6) makes a difference of almost 10 per cent in hardenability (in these units); this, however, does not apply to certain steels of high hardenability. It should be emphasized that in chromium steels (over 0.30 per cent Cr) and chrome-molybdenum and chrome-vanadium steels, undissolved carbides are likely to be present in the steel as quenched, and that in such cases the charts can indicate only a maximum possible hardenability, whereas the extent of hardening actually obtained may be much less. Thus tests on a number of chrome-molybdenum steels have indicated a degree of hardening amounting to only 50 to 65 per cent of the maximum possible, and in chromium steels from full hardening (100 per cent) down to as low as 70 per cent. On the other hand, when the amount of such elements is small (Cr under 0.30 per cent, Mo up to 0.25 per cent in the absence of Cr, and V up to 0.04 per cent), the charts provide reasonable approximations. The precise figures on the charts are suggested as tentative, subject to some modification as more data accumulate, but the fundamental concept appears to be supported by tests made on a wide variety of steels, a few of the correlations being shown in Fig. I.

Jan 1, 1942

By N. J. Grant, K. M. Zwilsky, J. T. Blucher

DISPERSION strengthening has been studied in a number of metal-metal oxide systems. To date those pure metals available as fine particles (less than about 5 p) have received most of the attention whereas solid solution alloys, generally unavailable as fine powders, have not been examined in detail. The present work was undertaken as a preliminary study to evaluate combined effects of solid solution and dispersion strengthening. The technique used was that of dry mixing of elemental powders with the oxide, followed by a diffusion treatment to obtain the solid solution. The average particle size of the powders used was: Ni, 3.2 p (1nco B); Mo, 1.3 p (Fansteel Metallurgical Corp.); and y alumina, 0.027 \i (G. L. Cabot, Inc.). The particle size of the metal powders was determined by Fisher subsieve measurements while the size of the alumina was calculated from a surface area determination. Prior diffusion studies on small compacts pressed from mixtures of nickel and molybdenum powders with and without 7.5 vol pct A1203 indicated that diffusion was complete after sintering for 100 hr at 1800" , or 50 hr at 2250", or 20 hr at 2550° F. The progress of diffusion was followed by microstructural examination and micro-hardness measurements. Two observations made during the diffusion experiments were: 1) That diffusion proceeded faster in the presence of alumina, confirming similar findings by Seith and Loepman on "SAP&apos;&apos;;4 2) That diffusion at temperatures above 1800°F resulted in coarsening of the alumina in agreement with results of other authors reporting on the growth of alumina in nickel above 1800"F. Similar experiments in the Ni-Cr system using 6 and 20 p Cr powders of 0.7 and 0.025 pct 0, respectively, did not result in significant solid solution formation even after long time exposures at 2550° F. It must be concluded that chromium oxide films around individual particles interfere with normal diffusion. A nickel alloy containing 5 wt pct Mo and 7.5 vol pct alumina was fabricated by dry mixing of powders, isostatically compacting at 30,000 psi, diffusing for 120 hr in dry hydrogen at 1800" F, and extruding at 1950"~ with a 30 to 1 reduction of area ratio. The resulting structure, presented in Fig. 1

Jan 1, 1963

By J. E. Coyne, L. E. Tanner, J. E. Boyd

In recent investigations of the isothermal transformations of the retained 0 phase in a titanium alloy having 13 pct V, 11 pct Cr, and 3 to 4 pct Al, there have been indications that the w transition phase is the initial decomposition product at temperatures below about 950°C (510 c).&apos; This metastable phase, found in a number of titanium-base alloys rich in stabilizing elements and known to have an adverse effect on their ductility and toughness, is not easily observed by the usual experimental techniques. Since wis coherent with the /3 matrix,495 distinct interphase boundaries are not revealed metallographically. However, in some cases, an etching effect (i-e., a "mottled" /3 surface) poorly resolved by light microscopy has been found to correspond with the presence of w.6 X-ray diffraction examination has generally been the most suc- cessful means for identifying this phase and crys-tallographic determinations have been made.4&apos;5 Dynamic elastic modulus7 and electrical resistivity measurementse have also aided in the identification of w. In the current study, the electron microscope was utilized to examine microstructures in aged Ti-13V-11Cr-3A1 that were unresolvable by light microscopy techniques. Using negative parlodion replicas (shadowed with chromium at 27 1/2 deg) examination was made of various transformation

Jan 1, 1962

By P. E. Fitzgerald, J. C. Gindlesperger

1. The Acidizer (Trade Publication of Dowell Inc.): No. 1. Aspects of Acidizing. No. 2. Inhibited Acid. No. 3. The Value of Acidizing. No. 4. The Chemistry of Acidizing. No. 5. Jelly Seal in Acidizing. No. 6. Loosening Stuck Drill Pipe. No. 7. The Geology of Acidizing. 2. J. C. Albright: Acid Is Being Used for Cleaning Intake Wells in Water-flooded Areas. Oil Weekly (Sept. 23, 1935) 24. 3. D. H. Bancroft: Acid Tests Increase Production of Zwolle Wells, Miller County, Arkansas. Oil and Gas Jnl. (Dec. 22, 1932) 42. 4. A. H. Bell: Possible Areas for Acid Treatment in Illinois. Pub. of Illinois- Indiana Petr. Assn. and lllinois Geol. Survey (1934) 49. 5. N. A. Best: Treating Limestone Formations Consists of More than Pouring Hydrochloric Acid in the Well. Oil and Gas Jnl. (April 5, 1934) 54. 6. L. G. E. Bignell: Sealing Off Salt Water from Oil-bearing Formations. Oil and Gas Jnl. (Jan. 30, 1936) 59. 7. L. G. E, Bignell: Acidizing Upper Limestone Formation Facilitated by the Use of Jelly Seal. Oil and Gas Jnl. (March 5, 1936) 30. 8. L. G. E. Bignell: Bailing, Shooting, Tubing and Acidizing Have All Been Tried in Deep Test at Binger, Oklahoma. Oil and Gas Jnl. (May 10, 1934) 38. 9. L. G. E. Bignell: Dowel1 Incorporated Increasing Its Field Service for Acid Treatment of Wells. Oil and Gas Jnl. (June 7, 1934) 62. 10. L. C. Chamberlain, Jr. and P. E. Fitzgerald: Oil Well Acidizing in Michigan. Mich. Acad. Sci. Papers (March 1937). 11. M. E. Chapman: Some of the Theoretical and Practical Aspects of the Acid Treatment of Limestone Wells. Oil and Gas Jnl. (Oct. 12, 1933) 10. 12. C. E. Clason and J. G. Staudt: Limestone Reservoir Rocks of Kansas React Favorably to Acid Treatment. Oil and Gas Jnl. (April 25, 1935) 53. 13. W. F. Cloud: Cleaning and Reconditioning Oil Wells—Reconditioning by Use of Acid (Petroleum Production 1937), 389. University of Oklahoma Press, 1937,&apos; 14. K. A. Covell: Developments in Eastern Production Practice. Acid Treatment of Michigan Wells. Amer. Petr. Inst. Prod. Bull. 213 (1934) 77; Oil and Gas Jnl. (May 31, 1934) 89; Oil Weekly (May 28, 1934) 17. 15. B. A. Cunningham: Acid Treatment of Lime Formations. Amer. Petr. Inst. Paper 901 (1933) 9D, 1-7. 16. 0. I. DeSchon: Pondera Production up 22,000 Barrels. Montana Oil & Min. Jnl. (Jan. 26, 1935) 1.

Jan 1, 1937

By Charles Jones

IN the winter of 1902, while occupying the position of mining and examining engineer for the Mountain Copper Co., Ltd., of Keswick, Cal., I had occasion to discuss with the General Manager, among other things, the question of the enormous waste of sulphur in smelting cupriferous, pyrite ore, and to cast about for means to utilize this waste or to render the fumes innocuous to vegetation. The decreasing rate of production of California farms, the consequent need of fertilizers, and the opportunity thus offered to utilize large quantities of sulphuric acid in the manufacture of superphosphates naturally brought the subject to the practical and commercial side of phosphates. I was informed that the company had made efforts to find phosphates in the West, without success. I advised as a start placing inquiries for phosphate deposits on the Pacific coast in the current mining ,journals. This was done, with the result that in January, 1903, from an advertisement in the Mining and Scientific Press we got track of a deposit in Rich county, Utah, through T.J . Wilson, of the Southern Pacific railroad in San Francisco. As soon as snow in the mountains would permit, in May, 1903, I undertook an examination of the property, with the result of recognizing the importance of the find, in spite of its lack of facilities for transportation at a commercial figure. On June 15, 1903, I concluded my report by saying, "This may be one of the world&apos;s deposits of phosphates." On Nov. 17, 1903, I concluded my report for the season in the following words : " I have now no hesitation in saying that I believe the Carboniferous formation so largely developed in the Wasatch mountains of Utah and the mountains of the Great Basin will prove to be one of the greatest storehouses of phosphoric acid in the world."

Jan 10, 1913

By J. R. Lucas, R. L. Frantz

Mine sealing programs for the alleviation of acid mine drainage were begun by the Federal Government in the &apos;30&apos;s. As part of an acid mine research project to The Ohio State University, a field test was undertaken to determine the effectiveness of mine seals for reducing mine acid. Details of the tests made at the McDaniels mine show the feasibility of air-sealing and flooding an acid-producing drift coal mine. Work by the University continues. The Federal Government undertook in the early thirties an extensive mine sealing program for the alleviation of acid mine drainage. In the states of West Virginia, Ohio, and Pennsylvania many general summaries have been made on this program but no critical evaluation with supporting data has been found on the effectiveness of mine seals for reducing mine acid. Because of this it was decided to undertake a field testing program as part of an acid mine research project of the Engineering Experiment Station, The Ohio State University. The objective of this phase of the project was to develop and install a field research laboratory for acid mine drainage studies. Specifications for such a mine laboratory were developed and resulted from the following considerations: 1) Size and preservation—the test mine shou1d be of limited size to permit frequent and detailed examinations; it should be in a reasonably sound condition to permit safe entry. 2) Isolation—the mine should be physically separated from all other mines. 3) Openings—a minimum number would be desirable to reduce the sealing problem. 4) Structure—the mine especially near the openings should be structurally sound. 5) Acid production—the mine should be producing significant quantities of acid. After considerable searching, the McDaniels mine was selected in the Hocking coal field in Vinton County just above Lake Hope State Park in southeastern Ohio. The small drift mine in the middle Kittaning coal consisted of two small rooms with one mine opening about 6 ft above a seasonal stream bed. Before it was finally decided to seal the mine, a short study was made of the acid drainage over several months. Fig. 1 shows a plot of acidity (milligrams per liter), flow (gallons per minute), and acid production (pounds per day) vs time over the sampling period. Because the coal was overlain with a porous sandstone, the mine drainage was consistent the year around even though the mine was

Jan 1, 1961

By Y. Nakada, E. J. Fasiska, A. S. Keh, L. Zwell

DURING an investigation of solid-solution hardening of nickel by carbon,&apos; we found that there were no reliable data on the dilating effect of dissolved carbon on the unit cell of nickel. Bernier2 reported an increase in lattice parameter of nickel (3.523A) when an unknown amount of carbon was introduced into solid solution (3.536A). Similarly KG, Tsien, and Misek3 found an increase from 3.523 to 3.531A when nickel was carburized. strong4 studied the variation of nickel lattice constant as a function of dissolved carbon. However, in his case carbon was introduced into nickel by solidifying the melt at 65 kbar. Therefore, his carbon concentrations were much above the highest equilibrium solubility limit5 (2.7 at. pct C at 1 atm). Knowledge of the lattice mismatch between solute and solvent is essential in understanding the mechanism of solid-solution hardening. Therefore, we decided to determine the dilation of the nickel lattice by carbon under the more usual conditions of atmospheric pressure. The material used was Inco Electrolytic Nickel of 99.95 pct purity, as specified by the supplier. The impurity contents in wt pct were: Fe, 0.01; Cu, 0.01; Co, 0.03; 0, 0.005; C, Si, S, N, trace. To this base material various amounts of AUC-grade high-purity carbon were added during vacuum induction casting. Ingots, 2 in. in diam, were heated in a muffle furnace at 1000°C for i hr., hot-rolled to 1/2-in. thickness, and quenched into brine. All ingots with carbon concentration greater than 0.98 at. pct were enclosed in a cast iron box during the heating to minimize decar-burization. Small pieces were then either swaged down to 0.028-in.-diam wire or cold-rolled to 0.010-in.-thick sheets. After the final cold reduction in area of 50 pct, the specimens were recrystallized by heating in evacuated capsules at 1100°C for 60 to 80 sec and then quenched into brine. The final grain size was 50 to 100 µ. The carbon contents of the alloys, Table I, were determined by a combustion method

Jan 1, 1969

By A. A. Smith, J. S. Smart, A. J. Phillips

The preparation of a sufficient quantity of high-purity copper for extensive laboratory use in the study of impurity effects has been an integral part of a research program in progress at the Central Research Laboratory of the American Smelting and Refining Co. for the past several years. Initially the degree of purity desired was that necessary to meet two essential requirements for experimental accuracy: (I) no impurity was to be present in the copper in an amount capable of producing a detectable effect on the properties to be studied, (2) even the minor changes in propertties caused by the addition of extremely small amounts of individual elements were to be measured without significant interference from contaminants. In the course of the investigation, evidence has accumulated to justify the conclusion that this copper is of exceptional purity; in fact, no impurity can be detected by any chemical method or spectrographic technique known to the authors. The present paper describes the method of production of the copper and presents some of its properties. Preparation oF High-purity Copper The elimination of impurities from commercial electrolytic copper is accomplished by a three-step purification process. The first operation consists of an electrolysis through a purified CuSO4-H2SO4 electrolyte using bagged anodes, low current density (5 amp. per sq. ft.) and low tem- perature (25°C max.). The electrolyte is prepared from commercial Cu2SO4 and is purified by heating the solution and adding I gram Fe as Fe2(SO4)3 to each 100 grams dissolved copper; sufficient hydrated oxide of copper, prepared from Cu2SO4 and NaOH, is supplied to precipitate Fe(OH)3, which carries down arsenic, antimony, tin, selenium, tellurium and other minor constituents. Silver and remaining traces of lead are removed by addition of Na2S to the acidified filtrate, which is then reatered after standing hot overnight. The cold solution is made up to 30 grams per liter Cu and 150 grams per liter H2SO4 and introduced without addition agents into the electrolytic system, which consists of a double row of hard-rubber battery boxes with the inside webbing and partitions removed, arranged in seven steps of ascending height. Electrolyte is uniformly supplied to the two highest cells by glass siphons operating from a constant-level reservoir, and flows by gravity through the bank of seven cells, entering each at the bottom and leaving at the top. Glass air lifts continuously supply the returning liquid to the constant-level reservoir for recirculation. Each cell contains three 4 3/4 by 4 3/8 by 1-in anodes cast from high-quality electrolytic wirebar copper, various lots of which have been used with equal success. The anodes are pinned to sheet copper hangars shaped to prevent crawl of the electrolyte to the multiple-wired bus bars—which would result in fouling of the contacts—-and hang inside wool flannel bags stretched over open wood frames. The deposit

Jan 1, 1941

By A. J. Phillips, A. A. Smith, J. S. Smart

The preparation of a sufficient quantity of high-purity copper for extensive laboratory use in the study of impurity effects has been an integral part of a research program in progress at the Central Research Laboratory of the American Smelting and Refining Co. for the past several years. Initially the degree of purity desired was that necessary to meet two essential requirements for experimental accuracy: (I) no impurity was to be present in the copper in an amount capable of producing a detectable effect on the properties to be studied, (2) even the minor changes in propertties caused by the addition of extremely small amounts of individual elements were to be measured without significant interference from contaminants. In the course of the investigation, evidence has accumulated to justify the conclusion that this copper is of exceptional purity; in fact, no impurity can be detected by any chemical method or spectrographic technique known to the authors. The present paper describes the method of production of the copper and presents some of its properties. Preparation oF High-purity Copper The elimination of impurities from commercial electrolytic copper is accomplished by a three-step purification process. The first operation consists of an electrolysis through a purified CuSO4-H2SO4 electrolyte using bagged anodes, low current density (5 amp. per sq. ft.) and low tem- perature (25°C max.). The electrolyte is prepared from commercial Cu2SO4 and is purified by heating the solution and adding I gram Fe as Fe2(SO4)3 to each 100 grams dissolved copper; sufficient hydrated oxide of copper, prepared from Cu2SO4 and NaOH, is supplied to precipitate Fe(OH)3, which carries down arsenic, antimony, tin, selenium, tellurium and other minor constituents. Silver and remaining traces of lead are removed by addition of Na2S to the acidified filtrate, which is then reatered after standing hot overnight. The cold solution is made up to 30 grams per liter Cu and 150 grams per liter H2SO4 and introduced without addition agents into the electrolytic system, which consists of a double row of hard-rubber battery boxes with the inside webbing and partitions removed, arranged in seven steps of ascending height. Electrolyte is uniformly supplied to the two highest cells by glass siphons operating from a constant-level reservoir, and flows by gravity through the bank of seven cells, entering each at the bottom and leaving at the top. Glass air lifts continuously supply the returning liquid to the constant-level reservoir for recirculation. Each cell contains three 4 3/4 by 4 3/8 by 1-in anodes cast from high-quality electrolytic wirebar copper, various lots of which have been used with equal success. The anodes are pinned to sheet copper hangars shaped to prevent crawl of the electrolyte to the multiple-wired bus bars—which would result in fouling of the contacts—-and hang inside wool flannel bags stretched over open wood frames. The deposit

Jan 1, 1941

By T. K. Perkins, L. R. Kern, R. E. Wyant

Many factors influence the results of hydraulic fracturing to stimulate well productivity. Most of these factors have been studied and their effects discussed in the literature. However, the movement of sand through the crack during fracturing has received little consideration. Exactly where the sand goes in the fracture, how it gets there, and how this might influence results have not been studied, judging from published reports. To get an answer to some of these questions we have studied sand movement through a simulated crack segment in the laboratory. Although the work is preliminary, we have answered some questions about sand movement in a semiquantitative sense. We are reporting our results at this time in the hope of stimulating interest and further study in this important aspect of fracturing. LABORATORY EXPERIMENTS Fig. 1 is a schematic diagram of equipment used in this study. A segment of a vertical crack 0.25-in. wide was formed between two Plexiglas plates, and movement of sand in fluid through the crack was observed under various conditions. Sand and fluid were injected into the crack at constant rates. The top and bottom of the crack were closed so that the fluid flowed horizontally. Because of gravity most of the sand settled to the bottom before moving very far through the crack. If the fluid velocity exceeded a certain critical value, then all the sand injected was washed on through the crack even though it had settled to the bottom. When the fluid velocity was less than this value when sand injection was started, most of the sand which settled in the crack remained in place until enough of the area open to fluid flow was blocked by the settled sand to increase the fluid velocity to the critical value.. At this point all injected sand moved on through the crack even though it settled to the lower portion of the moving fluid section. All sand previously settled remained in place. When fluid velocity was increased at this paint by increasing fluid injection rate (without also increasing sand injection rate), some of the previously settled sand was washed out until enough area was open to flow that the velocity dropped to the critical value. Conversely, when the fluid velocity was decreased by decreasing the fluid injection rate (without changing the rate of sand injection), additional sand would settle until the critical velocity was again attained. In short, the critical velocity is an equilibrium value which the system will automatically seek, if possible. The equilibrium (or critical) velocity was measured under several conditions. The results of these measurements are shown in Pig. 2 where the equilibrium velocity is shown as a function of the amount of sand injected per minute. The points falling in the shaded band shown in Fig. 2 were obtained with water. The higher viscosities were obtained by adding CMC (carboxymethylcellu-lose). Data indicate that the equilibrium velocity is not very sensitive to sand size or transport fluid viscosity, or to sand injection rate (except at very low rates). Points connected by the solid CUrve in Fig. 2 were obtained with water and 20- to 40-mesh steel shot. The equilibrium velocity is apparently fairly sensitive to difference in density between particle and fluid. Only one run was made with a gelled fluid and that was gelled water at zero sand injection rate as shown in Fig. 2. (Measurements of equilibrium velocity at zero sand injection rate were made by filling the crack

By Charles Will Wright

The present struggle for economic and industrial supremacy by the Communist world is against the United States, its main target, and the other Free World nations. The basis of industrial power is minerals, so the Communist nations are building up mineral production within their borders, often regardless of costs, and are acquiring quantities of mineral products from Free World countries for future use. They are contracting for additional supplies and advancing funds for the development of these quantities. A number of the less developed mineral producing countries of the Free World are being infiltrated by Communists and Communistic influence. This situation is well known, but as yet little has been done to oppose these activities or to awaken the public to their seriousness. The Communist countries are still far behind the Free World Nations in production of many essential minerals. But, shortages are being made up from Free World nations in exchange for petroleum, manganese ore, chromite and other products. Restrictions on shipments of strategic materials to Communist countries have not been maintained.

Jan 1, 1964

By W. W. Krech, T. E. Perkins

As fractures are propagated through rocks, energy is absorbed near the extending crack tip. Apparent surface energies for several rocks have been measured by cleavage under dynamic con-ditions. At nominal crack velocities from 0.5 to 500 in (min, measurements showed that fractures propagated in discrete jumps. Calculated surface energies and moduli were relatively insensitive to nominal rate of cleavage. in another set of experiments, rocks were cleaved under high confining stresses. The rocks were submerged in low leak-off fluids which formed a filter cake on the freshly cleaved surfaces (similar to the hydraulic lracturing process). Apparent surface energies were increased substantially as the surrounding fluid pressure was increased. Moduli in bending increased significantly upon application of the first 1,000 psi but were insensitive to stress level at greater pressures. INTRODUCTION For almost 20 years, hydraulic fracturing processes have been utilized effectively to stimulate oil and gas wells. During this period, some process improvements have resulted from studies of fracture orientation, mechanics of fracturing, areas generated, conductivities of cracks, etc. Yet many questions remain concerning the conditions and pressures needed during fracture propagation. In this paper we will report additional studies of the mechanics of fracture extension. It was shown previously3 that large rock samples could be cleaved under controlled conditions so as to measure the apparent surface energy (that amount of energy absorbed per unit area of new surface created). In this paper we consider the effects of two additional factors on surface energies, viz.: (1) effect of cleavage rate and (2) effect of confining stress level. THEORY Cleavage experiments were conducted on rock samples similar to that illustrated in Fig. 1. Blocks of rock several inches wide, 2- to 3-in. thick, and up to 3 ft in length were grooved longitudinally with shallow guide slots. A crack was initiated and allowed to extend along the web as the top of the rock specimen was pulled (or pushed) apart. Auxiliary equipment permits the measurement of (1) force applied at the top, (2) separation at the top and (3) crack length. (Further experimental details will be given in the next section.) The rock beams created by the crack are considered ton be cantilever beams. The deflection (or separation of the rock beams) at any point is calculated 4,5 by the beam Eq. 1.

Jan 1, 1967

By K. E. Merklin, F. D. DeVaney

Students of the modern blast furnace seem unanimously agreed that they are observing a major revolution in practice. Rather than changing construction and operation of the furnaces, most of the great advances now under way deal with the raw materials feed. Prepared burdens have gained increasing significance during recent years. Sizing, concentration, and agglomeration have all become bywords to greater production and reduced operating costs. The remarkable results with test furnaces in which self-fluxing sinter makes up part or all of the ore charge foretell even greater improvements. Other investigators have reported many of these advances, and it is enough to predict here that more and more operators will turn to the use of self-fluxing sinter. This change in the operator&apos;s requirements comes at a time when one major section of the iron ore regions, the Mesabi, is finding it difficult to maintain, let alone improve, the tonnage and grade of its shipping product. To meet the new demands with respect to quality and structure, the iron ore producers of this area turned first to screening, crushing, and concentrating intermediate ores. More recently, declining higher-grade reserves have led to multimillion-dollar investments in plants to treat the low-grade magnetic taconites.1,2 To yield a suitable concentrate, the Minnesota magnetic taconites must be ground to about 80 pct through 325 mesh, too fine to be charged directly to the blast furnace or to make acceptable feed to the sinter strand. The pelletizing process was developed to meet the structure requirements laid down by the blast furnace operators. Taconite producers are succeeding in their efforts just when the attention of the blast furnace man is being drawn to the great advantages of self-fluxing sinter.3 The taconite operators are ready with an alternative. Why not produce sclf-flus ing pellets? At the Hibbing laboratory of Pickands Mather & Co., where much of the early work was done that led to Erie Mining&apos;s 7.5-million ton taconite plant at Hoyt Lakes, Minn., preliminary investigations have been made. Present Taconite Operations: Details of this work will be clarified by a brief review of agglomeration practice in commercial taconite plants. After the taconite has been ground to about 80 pct through 325 mesh to liberate the magnetite grains, the ore is concentrated on magnetic separators, thickened, and filtered to about 10 pct moisture. This structure and moisture fit in closely with the conditions required to produce good balls. It has been found with taconite concentrates that at least 60 pct of the feed to the balling drum must be —325 mesh to obtain good ball compaction within reasonable time. The filter cake moisture usually permits a slight water addition at the balling drum, and further balling control is obtained by adding about 12 Ib of bentonite and 1 1/2 lb of soda ash per dry short ton of concentrate. These additives, and occasionally small quantities of fine coal, are thoroughly mixed with the filter cake ahead of balling. The swelling property of bentonite improves the green and dry strength of the balls so they can be handled without breakage until they are fired. Soda ash is added for pH control to insure maximum performance of the bentonite, and coal is added for more heat in the firing process when required. The balls, normally 3/8 to l-in. diam, are brought to a temperature of 2350" to 2500°F in the indurating furnace. Heat for this process comes from three sources: 1) fuel burned in the combustion chamber, 2) oxidation of the magnetite concentrates to hematite, and 3) coal added to the balls. Pellet strength is developed by grain growth and bridging and, if enough slag-forming constituents are present, by slag bonding. Conversion to hematite is usually close to 100 pct in the fired pellets. Batch Test Procedures: Initial studies of self-fluxing pellets were made by producing balls on a batch basis under carefully controlled laboratory conditions. The l-in. balls, uniform in size, were fired in an oxidizing atmosphere in a standard laboratory muffle furnace. The firing procedure consisted of

Jan 1, 1961

By Kenneth A. Gutschick, Robert S. Boynton

Lime has become a general and loosely used term to denote almost any kind of calcareous material or finely divided form of limestone or dolomite, as well as burned forms of lime. However, according to Webster, lime is only calcined limestone (known as quicklime, unslaked lime, or calcium oxide). It also embraces the secondary product of quicklime, i.e., hydrated lime or slaked lime (calcium hydroxide). There are two basic types of limestone used for lime manufacture-high calcium and high magnesium (dolomitic). The dolomitic quicklimes and hydrated limes correspond to their "high calcium" counterparts. The only difference is that the dolomitic types are a combination of the elements calcium and magnesium, in varying percents, whereas high calicum limes contain less than 5% MgO down to 1/2%. In either case a pure form of limestone of at least 97% combined carbonate content is needed to make salable lime, barring a few exceptions. The calcination of both types is diagrammed chemically as follows: [ ] In the foregoing reversible reactions the limestone is burned (calcined) in lime kilns, with the carbon dioxide content of the stone expelled as a gas. The weight loss during calcination of pure high calcium limestone is 44%; with equimolecular dolomitic limestone, the weight loss is 48%. Since lime has a strong affinity for CO_, particularly if moisture is present, it will readily revert to its original carbonate form. This occurs when quicklime "air-slakes" and adsorbs CO.. from the air. Thus, quicklime is perishable and should be stored in dry, water-tight areas. Depending upon the type of kiln used and the physical structure of the stone, the size of the quicklime may range from sandlike granules to 8 in. lumps. However, pebble sizes ranging from 1/4 in. to 2 in. are the most common. Screened quicklime fines are also com¬pressed (pelletized) into 1 in. pellets. For some purposes ground or pulverized forms of quicklime are used, ranging from No. 10 mesh to dust. Generally a pebble of quicklime is about the same size as the pebble of limestone before calcination (perhaps shrinking slightly). A more stable form of lime is hydrated lime. This is obtained by adding water to quicklime, slaking it into a dry, fine, subsieve size white powder. Because its affinity for moisture has been satisfied, it is not vulnerable to further moisture like quicklime. However, it has the same strong affinity for CO_. The following reversible chemical reactions illustrate how hydrated lime when heated (or dehydrated) can revert to the original oxide form:

Jan 1, 1975

By Enrique Hauser

Discussion of the paper of ENRIQUE HAUSER, presented at the New York meeting, February, 1916, and printed in Bulletin No. 110, February, 1916, pp. 521 to 534. GEORGE A. BURRELL, Pittsburgh, Pa. Mr. Hauser determined the influence of inert gases on the properties of fire-damp mixtures, and one might take some issue with his work on the effect of hydrogen, on those mixtures. It is valuable information to have, but as far as the practical application of it goes, it is not of much significance, because one does not find it in mines, at least not in any appreciable amount. The same Fig. 1.-APPARATUS FUR DETECTING COMBUSTIBLE GASES IN AIR. thing can be said of ethylene. Ethylene is formed in considerable quantities as a product of heat reaction, but normally, in the coal mines, as a result of a great many analyses we have made at the Bureau of Mines, we cannot say that ethylene is a combustible gas found in the fire-damp in coal mines.

Jan 5, 1916