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By C. D. Cochran, F. H. Deily, M. M. White, Victor Horner

A laboratory study of the effects of physical and easily measured rheological properties upon dynamic filtration was carried out. The purpose was to determine the factors affecting dynamic filtration and thus learn how to control this important property. The recommended procedure for determining the amount of fluid loss prescribed by API Code RP 29 "Field Procedure for Testing Drilling Fluids,"' is very useful for mud conditioning purposes, but inadequate for comparing filtration properties of different muds under dynamic conditions. Where dynamic filtration control becomes important for purposes such as logging and coring, other means of evaluating filtration properties are needed. A special device and testing technique were developed to provide reliable data on filtration characteristics of different fluids under dynamic conditions. Liquid saturated rock samples were drilled with a microbit, and the amount of connate liquid displaced by the fluid being investigated was measured. The empirical testing procedure was considered reliable. A large number of laboratory-prepared and field-mixed dritling fluids with a wide range of physical properties were studied. These properties included: (I) shape of consistency curves, (2) apparent viscosity, (3) plastic viscosity, (4) yield value, (5) density, and (6) API filter loss. Test results indicate that very little, if any, trend could be established between changes in these fluid properties and dynamic filtration. INTRODUCTION During the process of drilling, excessive loss of liquid from the mud to the formation is undesirable for many reasons. Filtration can reduce production-zone permeability, cause sloughing of formations, affect electric log curves, and render the fluid data obtained from cores unreliable. Thus, it is very important that we gain a knowledge of the factors that control filtration. It is generally agreed that in any type of filtration (static, dynamic, or combinations thereof), the mechanics of filtering depend in great part upon the composition of the fluid. Moreover, it is also true that the physical and rheological properties are influenced by the composition of the fluid. Indirectly, then, it was hypothesized that a relationship existed between dynamic filtration and physical-rheological properties, two effects of a common cause—fluid composition. This paper discusses an attempt to relate filtration-while-drilling to physical and easily measured rheological properties. It is believed that the results of this study may prove useful to others in the industry who are working on the problem of understanding filtration. The industry accepted practice for determining fluid loss from drilling mud is described in API Code RP 29. The practice is to measure the amount of filtrate that passes during a given interval of time from a quiescent drilling fluid under applied pressure through a piece of filter paper. This is primarily a test of static filtration properties. Other investigators have studied filtration under dynamic conditions. Williams5 and Pro-kop5 studied radial filtration and filter-cake buildup by circulating fluid through a porous cylinder under conditions similar to actual mud flow. Nowak,4 Krueger,3,4 and Vogel3 studied the effect of filtration upon permeability by flowing mud past the face of rock Samples. In their tests, the bit action was simulated by mechanical scrapers that kept the filter cake from building up. Ferguson and Klotz2 studied three types of filtration: static, dynamic, and beneath-the-bit. They investigated each type separately, using a model well in which bottom-hole hydraulics and mechanics were closely approximated.

Jan 1, 1958

By John N. Dew, William L. Martin, `

This paper describes the results of a laboratory investigation conducted to obtain data for an evaluation of the in situ combustion process as a method of producing crude oil from reservoirs. Air and fuel requirements, rates of advance, com-bustion temperatures, and coke and fluid distributions are presented. The mechanism of oil recovery by in stiu combustion is discussed. Five crude oils ranging in gravity from 10.9 to 34.2"' API were prodrrced from a semiadiabatic, uncon-.solidnted sand pack by in situ combustion. Experimental conditions were varied over a wide range in order to determine the inter-relation-.ships of process variables. The mini-mum air flux requirement for self-sustained combustion was found to he less than 10 scf/hr-ft'l. The rate of advance of a self-sustained combustion zone was found to be nearly proportional to the air flux at the combustion front. The effects of presrure and injected air flux were studied in a series of experiments using a 21.2" API crude. A minimum air require-ment was observed at an air .flux of 20 scf/hr-ft'. The oil saturation con-sumed as fuel averaged 5.5 per cent of pore volume. The effect of air pressure was, found to be .small for experiments having high combustion eficiencies. This study should promote a better understanding of the problems and mechanisms involved in labora-tory investigations and field applications of the in situ combustion pro-cess. The data presented will be use- ful in the interpretation of results of field tests. When tempered with volu-metric sweep efficiencies, the data can be used in making preliminary economic appraisals of the process as applied to reservoirs containing high porosity unconsolidated media. INTRODUCTION The purpose of this work was to obtain laboratory data for an evaluation of the in situ combustion process as a method of producing crude oil from reservoirs. In situ combustion basically consists of (1) injecting air into a reservoir through selected input wells to create an air sweep through the reservoir, (2) igniting the crude at the injection well, and (3) propagating the combustion front through the reservoir by continued air injection. By this means, oil is swept toward producing wells in the area. The fuel for combustion is supplied by heavy residual material (coke) which has been deposited on the sand grains during distillation and cracking of the crude oil ahead of the combustion front. Recovery of petroleum by a combustion or heat wave process is not a new idea. F. A. Howard was granted a patent in 1923 on a process in which air and a combustible gas were pumped into an injection well and ignited.' Russian engineers reported on field experiments with a crude oil gasification process in 1935'. Other known early field tests include those conducted near Bar-tlesville and Ardmore, Okla., in 1942." More recently completed field tests include an inverted seven-spot by Sinclair Oil Co. in the Delaware-Childers field, Nowata County, Okla.'; a three-well and an inverted five-spot test by Magnolia Petroleum Co. in Jefferson County, Okla."."; and a test by California Research Corp. in the Irvine-Furnace field in Kentucky. The Worthington Corp. has completed a test in cooperation with the Forest Oil Corp. in Clark County, 111.l Several field tests are now in progress. Sinclair has acquired a 600-acre lease in the Humboldt-Chanute field, Allen County, Kans., and is under way with a large-scale operation. Three field tests are under way in California. The General Petroleum Corp. is conducting an inverted five-spot pattern test in the South Bel-ridge field in Kern County under a cooperative agreement with 11 other companies, including Continental Oil Co. California Research Corp. is conducting a four-well pattern test in Midway-Sunset field near Maricopa, and Richfield Oil Corp. is under way with a test in the Ojai field in Ventura County. Although the published information is valuable concerning results of field tests, little laboratory data other than that reported by Kuhn and Koch5 are available to aid in appraising the tests from a technical or economic standpoint. Engineering data needed to evaluate the process include the amount of air required per barrel of oil recovered, minimum injection rates which will support combustion, rates of advance, air required per unit volume of reservoir cleaned, amount of oil recovered, and amount of oil consumed as fuel. The present laboratory investigation was undertaken to provide these data for that portion of the reservoir swept by the combustion zone. The first phase of this investigation was exploratory in nature and was conducted at injection pressures less than 100 psig. Data were ob-

JOHN H . NELSON,*Worcester, Mass.-In reply to the question "Is there any final information in regard to the relation between the temperature or the rate of cooling of a piece of heat-treated steel from the drawing temperature, and the so-called notch toughness?" I might say that we have conducted a large series of tests in the study of impacts on chrome-nickel and carbon steels and have gone into the subject of quenching from drawing temperature quite thoroughly but have found that one cannot lay down any definite rule or law with respect to the effect of quenching from the drawing heat. We conducted experiments on the effect of cooling by quenching in water, cooling in air, and burying the coupon in lime. Steels that have practically no impact value when air cooled may show very marked results when they are quenched in water, depending on the temperature at which the drawing is done.

Jan 6, 1919

Bureau of Mineral and Topographic Survey, University of Kentucky, Lexington, Ky. Arthur C. McFarlan, State Geologist The Kentucky Geological Survey was abolished by a recent legislative act and the Bureau of Mineral and Topographic Survey at the University of Kentucky was created in its place Publications of the former Survey are distributed by the present Bureau Several series of papers and volumes have been issued but only those of Series 3, 4 and 6 are now to be had in whole or in part Of Series 4, (1912-18) only one group of papers is available Vol: 1, pt. 2, Miscellaneous papers on fire clays, limestones and coals of parts of the state (1913), $125. Of Series 3 (1904-1912), the following bulletins are pertinent and may be had. Bulletin 3, Coals, clays, etc., of Kentucky (1905), 50 cents, 14, Coals of the Pineville Gap region (1912), $1.25, 16, The Waverlain formation of east central Kentucky (1912), 50 cents, 17, Coals of the Tradewater region (1912), 50 cents, 18, Coals of Quicksand region (1912), 50 cents; 19, Geology and coals of the Central City, Madisonville, Calhoun, and Newberg Quadrangles (1912), 50 cents

Jan 1, 1933

By H. D. Kessler, M. Hansen, R. J. Van Thyne

The phase diagram of the titanium-rich portion of the system Ti-Cr-Fe to 70 pct Ti was established by means of isothermal sections at 900°, 800°, 750°, 700°, 650°, 600°, and 550°C, using arc-cast alloys. An isotherm at 800°C was determined for the section bounded by Ti, TiFe2, and TiCr2. Micrographic analysis was employed as the principal method of investigation, supplemented by X-ray diffraction and incipient melting studies. A ternary eutectoid, ß ? a+ TiCr2 + TiFe, occurs at approximately 8 wt pct Cr, 13 wt pct Fe, and about 540°C. AMONG the first titanium-base alloys in commercial production were the Ti-Cr-Fe alloys. For this reason, the Materials Laboratory, Wright Air Development Center, sponsored an investigation of the system. No information on the constitution of the titanium-rich corner was available. Vogel and Wenderott studied titanium-poor ternary alloys.' As part of this program, the binary systems Ti-Cr and Ti-Fe have been determined previously by the authors.' Other investigators have also studied the binary systems Ti-Cr,3-7' Ti-Fe.³,8-10 Both of the binary systems are characterized by the eutectoid decomposition of the ß phase. Most of the effort of the present work has been concentrated on the titanium-rich corner with compositions of less than 30 pct total chromium and iron content. The determination of isothermal sections was the experimental approach used to obtain the ternary phase diagram. Vertical sections were used to check the consistency of the isothermal sections. While this investigation was in progress, a study of the Ti-Mo-Cr system at Armour Research Foundation disclosed the existence of a high temperature modification of TiCr2 above 1300 °C.11 This allotrope has a hexagonal structure of the MgZn2 type, iso-morphous with TiFe2, whereas the low temperature modification of TiCr2 is cubic," of the MgCu2 type. The existence of allotropy in TiCr2 of course means a more complicated set of phase relationships in this general region than if the phase were monomorphic, as was first assumed. Additional studies were made late in the investigation in order to clarify the part that the hexagonal TiCr2 phase plays in the ternary phase relationships. These will be discussed in a separate section. Because they are of greatest practical significance, the phase relationships in the titanium-rich corner will be presented first, omitting the equilibria involving the hexagonal modification of TiCr2 in an effort to simplify the presentation somewhat. Experimental Procedure Materials: The titanium used in the preparation of the alloys was iodide crystal bar (99.9 pct pure) produced by the New Jersey Zinc Co. High purity chromium and iron were obtained from the National Research Corp. Table I gives the analyses of these materials. Melting Practice: Over 100 alloy ingots weighing 10 to 20 g were melted in a nonconsumable electrode arc-melting furnace. As the techniques are identical to those reported previously,12,13 they will not be detailed here. The ingots were melted in the cavity of a copper melting block insert under a slight positive pressure of helium. No measurable hardness

Jan 1, 1954

By W. F. Rogers, J. A. Stewart, P. J. Kalish, E. O. Bennett

Detailed descriptions are given of materials, apparatus and the experimental procedure used to study the effect of bacteria on sandstone permeability. The factors affecting permeability during injection of bacterial suspensions which have been investigated are: (I) concentration of bacteria; (2) core permeability and median pore size; (3) species of bacteria, mode of aggregation and relative size; (4) injection rate or pressure differential; (5) mean pressure; and (6) depth of penetration of bacteria. The investigation demonstrated that bacteria cannot reduce core pertneability to zero and that their effect on permeability is subject to definite limitations. Remedial or permeability restoration studies also were made. Acidization in combination with reverse flow war found to be an effective method for restoring permeability In cores partially plugged with bacteria. The relationship between the bacterial quantities in the laboratory tests and those found in field operations is discussed. The linear laboratory flow data have been translated into terms of field radial systems; these data indicate the most practical methods of maintaining injection rates in the presence of bacteria are to increase injection pressures or hydraulically fracture the formation. INTRODUCTION It has long been known that various bacteria flourish in the above-ground components of oil field water-injection systems. The presence of these bacteria has led to the suspicion that they might enter into the causes for reduction of injection well permeability, but only recently has an attempt been made to describe the effects quantitatively.1 Biocides are used in many cases without a firm knowledge as to whether the bacteria are killed, or if they are, whether the dead cells affect formation permeability. This project was undertaken as part of a study to determine the effect of bacteria, residual oil and precipitated solids such as iron sulfide or calcium carbonate on the permeability of sandstones and to learn whether any of these materials adversely affect brine injectivity in secondary recovery or disposal operations. This report deals with that portion relating to the effect of bacteria. MATERIALS CORES Three different permeability ranges of Berea sandstone cores were used in the study. The range of initial, absolute brine permeabilities kj were: high — 278 to 400 md: medium—130 to 162 md; and low—17.7 to 48.3 md. The permeability data for the cores used in the tests are included in Table 1. All cores were cut to a nominal 1 in. diameter. Following air permeability determinations, they were molded with an epoxy resin in 1.5 in. ID aluminum sleeves. The sleeves had drilled and tapped holes for intermediate pressure connections along the length of the cores. After molding, the high- and medium-permeability cores were trimmed to 4 in. in length and the low-permeability cores to 2 in. These lengths resulted in pore volumes of approximately 10.9, 10.4 and 4.4 ml, respectively, for the high-, medium- and low-permeability cores. Pressure-transmitting channels were drilled through the plastic into the core proper, using the sleeve holes as drill guides. Width and depth of channel penetration into the core proper were approximately 0.03 in. and 0.05 in., respectively. Three pressure taps were used with all cores; for the high- and medium-permeability cores the taps were spaced 1 in. apart; for the low-permeability cores the taps were spaced 0.4, 0.8 and 1.2 in. from the inlet end. Pore size distributions were calculated from restored-state capillary pressure curves for representative cores in each of the three permeability ranges. In each case the pore size range with the greatest percentage of the pore space included the very small pores up to 0.5 micron in radius. The median pore radii were in the 5.5 to 6 micron range for the high-permeability cores; in the 4.5 to 5 micron range for the medium-permeability cores; and in the 3.5 to 4 micron range for the low-permeability cores.

Jan 1, 1965

By H. Tabachnick, N. Hedley

The extraction of gold and silver from ores with alkaline cyanide solutions is well known. Cyanide solutions are also good solvents for many base metal minerals, particularly most of the copper minerals. Leaver and Woolf,1 who determined the solubilities of copper minerals in 0.10 pct sodium cyanide solution, showed that malachite, cuprite, chalcocite, and bornite are readily soluble in cyanide; least soluble are chalcopyrite and chrysocolla. The Mining Chemicals Laboratory of American Cyanamid Co. has studied the possibilities of using cyanide solutions for leaching base metals. In this article the authors present the results of their experiments with cyanide leaching to reduce the copper content of a zinc concentrate from 0.050 pct to the desired specification of less than 0.02 pct. Zinc Ore Sample: The sample of commercial zinc ore used in this investigation to produce zinc concentrates was ball mill feed ranging from 1/2 in. to fines. The ore was crushed to —10 mesh and riffled into 600-g charges for zinc flotation. Analysis revealed 4.54 pct Zn and 0.005 pct Cu. Examined microscopically, the ore consisted essentially of cream-colored sphalerite and carbonate gangue. It contained a small amount of pyrite, partly tarnished. The only copper mineral identified was chalcopyrite, mostly smaller than 20µ, some of which occurred as attachments on sphalerite. Flotation Procedure: Batches of zinc concentrates were made as required according to the following general procedure. A 600-g charge of —10 mesh ore was ground for 4 min at 60 pct solids in a laboratory steel rod mill. Screen sizing showed it to be all —35 mesh with 42 pct —200 mesh. The pulp was transferred to a Fagergren laboratory flotation machine, diluted to 25 pct solids and conditioned for 5 min with 0.6 lb per ton CuSO4, . -5H2O. It was then conditioned with 0.10 lb per ton Sodium Aerofloat promoter and floated for 5 min with pine oil as frother and creosote as froth stiff-ener. The pH during flotation was 8.5. A typical analysis of zinc concentrate subjected to leaching is 44.41 pct Zn and 0.051 pct Cu. The ratio of concentration was close to 10 to 1. Leaching Procedure: Aero Brand cyanide, technical calcium cyanide analyzing 49.5 pet NaCN equivalent, was used for all cyanide leaching tests. For each test three rougher concentrates floated as described were combined (total weight close to 180 g) and the water content adjusted to 150 ml. The percent solids in the pulp was thus 55. After a weighed amount of cyanide flakes had been added, the pulp was agitated on rolls for a given period, then filtered and washed. Filtrate and washed concentrate were assayed for copper content. The procedure for ammonia leaching was similar to the cyanide process. A series of tests was run in which the effects of ammonia as a solvent for copper in the zinc rougher concentrate were investigated. In each test 180 g of zinc rougher concentrate was leached with 150 ml of ammonia solution. Various strengths of ammonia and times of leaching were investigated. The results are shown in Table I. It is apparent that ammonia was a poor solvent for copper in the zinc concentrate; 60 min of agitation with a relatively strong NH:, solution extracted only 2 pct of the copper. CYANIDE LEACHING TESTS Effect of Solution Strength: The solvent effects of 0.125, 0.187, and 0.250 pct NaCN equivalent solutions were compared for leaching periods of 15 min. Residue and filtrate were assayed for copper content. The copper content of the filtrate plus wash was calculated back to the original volume of cyanide solution (150 ml). The effects of cyanide strength are given in Table II, It will be noted that 15 min leaching with 0.25 pct NaCN equivalent solution extracted 74.6 pct of the copper in the zinc concentrate. The leached product assayed 0.013 pct Cu, appreciably below the required specification of 0.020 pct. Effect of Leaching Time: Portions of zinc concentrate were leached with 0.25 pct NaCN equivalent

Jan 1, 1961

By J. A. Herbst, G. A. Grandy, T. S. Mika, D. W. Fuerstenau, G. D. Gumtz

As a basis for interpretation of grinding behavior, the discretized batch grinding model is briefly presented and described, with solutions of the resultant equations for linear systems. Techniques for obtaining relevant experimental data are discussed for first order grinding systems, with methods for estimating other systems1 parameters from these data. Computer methods are described for conducting a sensitivity analysis which can be used to obtain an improved estimate of the system parameters. Examples of the use of the computer in conjunction with the grinding model for the simulation of batch grinding, locked-cycle grinding, and closed-circuit grinding are presented.

Jan 1, 1969

By Eckley B. Coxe

SINCE our last meeting, the Institute has lost, by the death of Mr. Gowen, one of its most distinguished members. I shall only attempt in this place to give a brief account of his many accomplishments and achievements; but our records would certainly be incomplete without a recognition of the important influence which he exerted upon the mining and metallurgical industries of this country, and a testimony to his private virtues of character. Although Mr. Gowen was best known as a great lawyer and as the head of a great railway corporation, he began his career, before he was of age, as the manager of a small anthracite blast-furnace at Shamokin, Northumberland County, Pennsylvania. He left this position, to engage on his own account in the mining of anthracite coal at Mt. Laffee, near Pottsville. This proved an unprofitable business, and, after the panic of 1857, he began the study of the law. As soon as he was admitted to the bar, he at once entered upon the active practice of his profession, and soon achieved an extraordinary success. In 1869 he was made President of the Philadelphia and Reading Railroad Company, and he afterwards organized the Pliladelphia and Reading Coal and Iron Company, the largest coalmining corporation in the world. He became a member of. this Institute during the first year after its organization, and remained a member until his death. He always took a deep interest in the work, and was prompt to co-operate by all means in his power in furthering the objects of the Institute. Mr. Gowen was one of the first to recognize the great importance to both miner and consumer, of the proper utilization of the small

Jan 1, 1890

By Wilbur C. Helt

The United States is self-sufficient in bituminous coal and lignite resources to meet the high range of energy demands forecast for the remainder of this century and beyond. The known or proved recoverable reserves of energy from coal in the United States are estimated to be 88% of all fuel reserves, whereas the energy reserves for oil are estimated to be 276, gas 376, shale oil 5% and uranium 2% (as now used). The demand for all forms of energy used in the United States is increasing at a rate of 3% to 4% per year. Presently, coal is competitive on a supply and cost basis with other fuels in conversion to electrical energy and as a source of energy for converting raw minerals to metals and nonmetals, but it is not competitive in transportation and other uses that demand the high flexibility that oil and gas enjoy because of their form. Coal conversion to fluid forms is essential to make it interchangeable with gas and oil to meet the markets they now occupy. Essentially, the chemical constituents that make up coal are the same hydrocarbons that make up oil and gas; it is a matter of reconstituting the proportions of hydrocarbons to make coal interchangeable with the other fuels. The production of coal has fluctuated widely between the peak production of 631 million tons in 1947 to a low of 403 million tons in 1961. Throughout this period, the total value of the production, f.o.b. mines, has averaged more than $2 billion annually. Beginning with 1962, there has been a steady upward trend in coal production with only minor fluctuations that were mainly caused by labor disputes. Output for the latest typical year, 1970, was 588 million tons. Production in 1971 was affected by a major strike and dropped to 548 million tons. Exports of coal from United States to foreign countries have increased steadily and developed into a substantial annual market for over 71 million tons per year. About 56 million tons of this is suitable for metallurgical purposes, and the remaining 15 million tons is exported for electric power generation. The largest single-industry demand for coal is in the electric generating industry, Fig.1. Roughly, the production of electricity has doubled every

Jan 1, 1973

By J. H. Fletcher

ALTHOUGH some mines have been completely mechanized, on the whole mechanical equipment has made slow progress in the coal fields. Coal mining is particularly well adapted to hand labor. The skillful hand miner can keep his working place safe under difficult roof conditions, remove impurities from the coal at the point of origin, and produce a maximum percentage of large-size coal, so as to command the highest price on the market. Hand mining has also the advantage of low capital investment per ton of production and moderate carrying charge when the mine is idle. Though animal haulage gave way to electric locomo-tives and hand cutting to mining machines, beyond these concessions most of the tonnage is still produced by manual labor. But factors over which producers have no control are speeding the change to mechanical mining. The additional investment in underground equipment and surface preparation plant must be offset by a great increase in tons per man-day, and to do this the employees should be able to utilize as much of their time as possible on production. In the installation of mechanical mining, the only change in method usually has been to add loading machines to a mine already developed for hand loading. Because of a lack of suitable equipment, few mobile loading machines have been installed in seams of coal less than 4 ft. thick; the mechanization in the thinner seams has been confined largely to shaking conveyors and hand loading onto chain conveyors. Mechanical mining requires an organization that is prepared to study its particular problem from production territory to market in recognition of the changed factors arising from the use of loading machines at the working face, and other causes, such as safety of men and property; high wage scale on hourly basis; concentrated working territory and rapid extraction from a given area; change in the percentage of sizes of coal produced and effect of this on average realization. There are also the coal-cleaning problem; increased capital investment, maintenance, obsolescence, and the relation of all these to the spread between realiza-tion and production cost. If coal seams were uniform, the answer would

Jan 1, 1939

By John P. Harmon

This paper was written with two objects in view: (1) To describe in detail what has been done toward the ventilation of the main unit of the Portovelo mines and the results; (2) to give information that may be useful to anyone who may have to make a preliminary mine ventilation survey, with recommendations for the improvement of the existing system and finally the maintenance of the improved system. General Description The mining operations of the South American Development Co. at Portovelo, Province of El oro, Ecuador, are described here only briefly, as a full discussion does not fall within the scope of this paper. The orebodies are of the steeply dipping, gold quartz vein type, averaging about 2.0 meters wide. As shown in Fig. 1, the ore chutes are relatively far apart, making considerable development work necessary before connecting raises can be driven between the levels. Faulting' on a large scale and a parallel vein system have made long dead-end drifts and many crosscuts necessary in the exploration of veins. The ventilation of these dead ends is described under Auxiliary Ventilation. The method of mining2 is the filled rill stope method, similar to the method that is largely employed at the Butte mines. The main shaft of the mine is in the Amarillo Valley, on either side of which mountains rise abruptly to a height of several hundred feet above the collar of the shaft. Entrance is made to the mine: (1) Through the American shaft, which extends vertically from the surface to a depth of 1100 ft., of which only the first 700 ft. are in active use at the present time; (2) levels are driven from the American shaft at intervals of 30 m. (100 ft.), of which only the third, fifth and seventh levels need be considered in the ventilation problem; (3) adits at intervals of 30 m. (100 ft.), with a few exceptions, above the collar of the American shaft.

Jan 1, 1931

By Donald L. Katz

The equilibrium constants developed by Souders, Selheimer and Brown1' have been very useful tools in predicting vapor-liquid equilibria during the past five years. Their equilibrium constants were computed by assuming ideal solutions and recognizing the deviations of vapors from ideal gas laws. The same fundamental concepts were used by Lewis and co-workers8 for predicting vaporization phenomena by means of fugacities. The equilibrium constant is defined by equation 1: y\x =K [1] in which K = equilibrium constant, y = mol fraction of a constituent in the vapor phase, x = mol fraction of the constituent in the equilibrium liquid phase. The equilibrium constants are a function of temperature and pressure and might be considered as replacing the ratio of the vapor pressure to the total pressure of Raoult's law. They may be used to compute the vapor or saturation pressure of a liquid or the dew point of a vapor for any pressure or temperature from the analysis of the mixture much as Raoult's law is used. Likewise, the quantity of a mixture that will vaporize at any equilibrium temperature and pressure and the composition of the vapor and liquid may be calculated. The purpose of this paper is to review the available equilibrium data of interest in petroleum production, to discuss the auxiliary physical properties of the hydrocarbons, and to present typical uses of the equilibrium constants. The particular applications chosen are the computation of the results of stage separation of crude oil and natural gas, the calcula: tion of a pressure-volume curve of a crude oil-natural gas mixture, and information gleaned from a consideration of reservoir equilibria.

Jan 1, 1938

By S. Andrew Kulin, L. Kaufman

A general treatment of the broadening of X-ray double-crystal spectrometer rocking curves due to distributions of tilts and strains has been carried out. The results of this treatment have been applied to the case where dislocations in a crystal give rise to both tilts and strains. On this basis it is found that the rocking-curve halfwidth due to imperfections in the crystal, ßC , increases with increasing Bragg angle, 8. These considerations predict linear relationships between & and tan2 , and ßCand tan 8, for Gaussian and Cauchy distributions of dislocations, respectively. The zero intercept of the ßC us 8 curve, which is a measure of the degree of imperfection of the crystal, is found to depend directly on the distribution rather than the number of dislocations. In addition, the slope-intercept ratio for the ßc - ? curve has been calculated for both the Gaussian and Cauchy distributions. Experimental determination of the - 8 curves for a series of copper single crystals prepared by the Bridgeman and Czochralski methods yields results in agreement with the calculated slope-intercept ratio for a Cauchy distribution of dislocations. CERTAIN X-ray techniques have been used to study the perfection of single crystals. Gay, Hirsch, and Kelly have considered the spreading of spots into arcs on Debye-Sherrer rings due to imperfections. The Double-Crystal Spectrometer has also been applied to this problem.2 The characteristics of Double-Crystal Spectrometer Rocking Curves have been described in the literature by various workers.2-4 By assuming a random spacing of dislocations which give rise to a Gaussian distribution of misorientations relative to some mean orientation in a crystal Gay, et al.,1 arrive at the conclu- sion that the most probable misorientation is given by where h is the spacing between dislocations, and 8 and are the relative misorientations of neighboring regions. If ßC is the halfwidth of the rocking curve (the width at one-half the peak intensity) due to dislocations in the crystal then5

Jan 1, 1960

By Joseph T. Singwald

INTRODUCTION THE purpose of this chapter is to present the diagnostic features (geologic, mineralogic, chemical, and physical) of the principal types of hypogene ore deposits recognized in current genetic classifications, and to classify the ore deposits of the western United States into those types. The genetic divisions used and the criteria determining them are essentially those followed by Lindgren in his Mineral Deposits. Deviations from Lindgren represent either changed conceptions since 1928 (the date of the third edition of Mineral Deposits) due to advance in knowledge or the inevitable differences of opinion between experts in active fields of investigation. Though nineteen economic geologists have been given free hand in presenting their particular topics, there are no fundamental departures in the chapter from the principles set forth by Lindgren. It would be entirely beyond the capacity of a single chapter to give a comprehensive geologic description of all the ore deposits of the western United States. Descriptions of individual deposits have been restricted to those features that are of most diagnostic genetic value. All of the more important deposits receive that much attention. Deposits of lesser economic importance and those inadequately known geologically are treated more briefly or merely cited under what seems to be their proper genetic assignment. The geographic limits of the descriptive matter are the Black Hills and the United States between the Rocky Mountains and the pacific Coast. Capacity of the chapter has made it necessary to impose these geographic limits.

Jan 1, 1933

By H. R. Barnhurst

It would be a difficult matter to trace from the beginning the very few improvements made in the burning of fuels prior to 1860. Donbtless the crossing of the sticks of wood in building a mood fire early appealed to our savage ancestry as a means of expediting its effectiveness for their crude culinary purposes. Stones and andirons came later in assisting the admission of air to the under part of the fuel bed, while beacons and cressets, utilized metal bars as supports for the same purpose. It is probable that Tuba1 Cain owed his success to the use of bellows making possible the use of mixed fuel and charge upon a hearth. With the development of working in cast iron, grate bars came in, and they are still with us doing duty in every household and almost every art where heat is a factor. Little was done beyond these crude methods prior to 1860. With the advancement of knowledge came the fuel-gas producer, a natural child of the retorts employed in supplying illuminating gas. It is only within the last 15 years that the conclusion became general that complicated systems for saving heat would not remedy deli-. ciencies in its primary development, and within that time there has been great advancement in the means of producing initially a fire of high effciency, so that to-day we are moving rapidly toward both high development and high absorption, with the resultant of high efficiencies. In studying the subject of the availability of pulverized coal as a fuel, it may occur to ask, why has this not come up earlier? This, question would be naturally prompted by recalling the fact that Thomas Crompton's English patents upon his methods were issued to him nearly 40 years ago. A conclusive answer to this, however, is that Crompton at that time had very crude facilities for making what was at best a very imperfect pulverization of the coal, arid the expense of preparation mas very great. He does not seem to have been aware of the advantage of a thorough drying of the coal in the

Jan 1, 1914

By John D. Nelson

Jan 1, 1985

By E. H. Crabtree, Neil S. Parker

The custom milling plant of the Eagle-Picher Mining and Smelting Co. is at Sahuarita, Arizona, approximately 20 miles south of Tucson. It is connected by a 2-mile railroad spur to the main line of the Southern Pacific Railroad between Tucson and Nogales. The capacity of the plant is 500 tons per day. The mill is conventional except perhaps in the rather wide variety of ores treated. These vary from straight copper ores, from which only one concentrate is produced, and lead-zinc or copper-zinc ores from which two concentrates are produced, to complex gold-silver-copper-lead-zinc ores, which require the simultaneous production of three or lour concentrates. Custom ores are received in lots varying in size from a few truckloads to several carloads. Similar ores, after crushing and sampling, may be commingled, or individual lots may be milled individually. The ores are first dumped either into one of two 200-ton railroad bins, or into one of the two 75-ton truck bins. From these bins the lot of ore is individually crushed in a 24 by 16-in. jaw crusher to 3-in. size, and then to ½-in. size in a 3-ft Symons shorthead crusher. After crushing to ½-in., the ore is conveyed to the automatic sampling plant. This consists of three Vezin samplers in series, each successively cutting out 10 pct, 10 pct and 5 pct of its feed. After No. I sampler, the material passes through a mixing barrel and is then crushed to -in. in a set of rolls before passing to No. 2 sampler. Between No. 2 and No. 3 samplers, the ore is again mixed in a mixing barrel. The final sample, consisting of one pound per ton of ore, is taken to the laboratory, where it is further crushed to 1/8-in. in a coffeemill, and then riffled down to the size required for grinding in a pulverizer. The capacity of the crushing and sampling plant is up to 100 tons per hour, depending upon the character of the ore. After sampling, the ore is stored in five ore-storage bins, consisting of four 200-ton bins and one 175-ton bin. Each of these bins is discharged by means of belt feeders and conveyor to the ball mills, so that ores may be composited for milling in any required ratio, or milled separately by themselves. Fine grinding is done in one 6 by 4-ft Allis-Chalmers ball mill and one 8-ft by 36-in. Hardinge mill. These two mills both discharge to one 54-in. Akins Simplex High weir classifier. The classifier sands can be returned to either or both mills, so that flexibility in grinding for different ores is easily obtained. Flotation equipment consists of four 66-in. Fagergren machines for copper-lead roughing and six similar machines for zinc roughing. Lead-copper bulk concentrate is cleaned and recleaned in eight No. 18-S Denver machines, and zinc concentrate is cleaned in three 66-in. Fager-grens. Copper-lead differential separation

Jan 1, 1949

By Albert Wright

VEINS of asphalt fill and seal vertical fault fissures at the surface of a large domal structure near Bejucal, Havana Province (about 19 miles south of Havana), so this structure was chosen, by Barnabas Bryan, for the drilling of a well on the theory that oil had risen from below the Cretaceous and had entered the porous sands of the Upper Cretaceous, which outcrop in three places within a radius of 12 miles. The asphalt in these fault fissures, he contended, had sealed them and thereby prevented further dissipation of the oil. Furthermore, he found, in the asphalt, specimens of sand and lime that were indications of its upward migratory character. Ralph Arnold, who discovered the Upper Cretaceous sand saturated with oil in the outcrop near Mariel, in Pinar del Rio Province, has compared the Bejucal structure with the Santa Maria field in California. The surface rock of the structure is Oligocene limestone and shales, thin, bedded and well stratified. This formation has, heretofore, been assigned to the Eocene with a thickness of 3000 ft. and locally called the Nazareno formation. Paleontologic examination of well cuttings, however, has disclosed Oligocene fossils at a depth of 1142 ft. and Upper Cretaceous fossils at a depth of 1541 ft., leaving approximately 400 ft. of very parenthetical Eocene. Whether there is any Eocene in this structure has not been determined.

Jan 3, 1925

By N. W. Lord

The process of molten-zone refining is analyzed for long ingots and many zone passages. Formulas are derived which give the resultant impurity distribution in terms of finite series. A comparison with the approximate procedure of Hamming is given. HE physical principles and applications of an extremely physicalprinciple efficient form of metallurgical refinement has been described by Pfann. The purpose of the present paper is to describe a method of analyzing exactly the particular program used which enables the segregation effect to be predicted for any number of molten-zone passages in a long ingot. The method is applied to the particular case of refinement of an ingot whose impurity initially is uniformly distributed throughout its length. A number of molten zones of equal length are passed through the ingot effecting a radical redistribution of impurity. Pfann has indicated an approximate method, due to R. W. Hamming, of calculating the resultant concentration after each successive zone pass for a particular value of the segregation constant defined in his paper. Here a solution will be presented in terms of the number of zone passes and the segregation constant. The expression, though cumbersome, is exact and susceptible to ordinary numerical computation procedures. The results of a similar computation using the procedure of Hamming are presented in a table together with the exact results of the present method. The discrepancy in terms of absolute concentrations is tabulated for the first eight zone-lengths. To establish the notation (which follows that of Pfann1 as closely as possible) and physical basis of the analytical equations, the physical model and principal assumptions may be reviewed. An alloy of two elements, where there is formed a continuous range of solid solutions, usually does not melt as a simple compound. Rather, a temperature is reached where the solid solution is in heterogeneous phase equilibrium with a liquid solution of different composition. The temperature dependence of these equilibrium compositions forms part of the phase diagram. For very small concentrations of a solute B in a solvent A, this usually takes the form of Fig. 1. Sometimes the solidus and liquidus slope upward. This corresponds to a segregation constant (defined below) which is greater than unity. The segregation constant is now defined as k = Cn(x)/CnI(x) [1] where C,,(x) is the impurity concentration in the solid ingot at distance x during the nth passing of a molten zone and Cnl (x) is the impurity concentration of the liquid zone from which the solid at dis- tance x is formed (see Fig. 4 of ref. 1.) C (x) remains the same after passage of the zone. The constant k may be either greater or less than unity in general. Purification in the former case is effected only in a finite ingot and in the portion that is melted last. For k less than unity purification is effected even in an infinite ingot. The method which follows gives, in the former case, the successive increases in impurity concentration and, in the latter case, the successive decreases in concentration. The general case of impurity redistribution will be considered first, and purification will be discussed later on. The analysis rests on the following assumption: The movement of the zone is too rapid to allow appreciable atomic rearrangement in the solid sections and too slow to disturb the uniform impurity distribution in the liquid zone characteristic of equilibrium. Hence, the composition in the solid at the left solidifying interface will be determined by Eq. 1 while the impurity concentration of the liquid zone will be uniform throughout its length. The reasoning which follows closely parallels that of Appendix 11 in Pfann&apos;s paper. It is reviewed here for the case of the nth zone pass in order to make clear the meaning of an operator essential to the present method. Fig. 4 of ref. 1 shows the movement of a molten zone of length 1 in an ingot of total length d. Each Cn(x) can be determined from the condition that the amount of solute added to the zone during an incremental advance, dx, is due to the melting in of a solid portion C(x)dx and the freezing out of kCnl(x), that is d I —r- CnL (x) dx = Cn-1 (x+l)- kCnL (x) dx or, in terms of Cn(x) d k k —— C,(x) +—c.(x) =— Cn-1 (x + l). [2] dx l l This, of course, is derived from the main assumption, the fact that 1 is constant, and that the total impurity content previously present up to x + 1 is constant . A correction has to be made for the region (d — nl) < x < d. This is due to the zone length changing during the passage of the solidifying interface beyond x = d — 1. Since the general solution would be too complicated otherwise, only the region 0 < x < d — nl is considered. The general solution of Eq. 2 is

Jan 1, 1954