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By P. A. Beck

ACCORDING to the traditional definition,1 re-x crystallization is a process taking place upon annealing of cold worked metals, characterized by the appearance of new strain-free grains, growing at the expense of the cold worked grains until ultimately the new crystals completely replace the cold worked structure. In metals like brass this process was observed long ago to take place simultaneously with the return of the mechanical properties to those characteristic of well annealed material. The above definition implicitly assumes that the matrix in which the new grains are growing is essentially still in the cold worked condition at the time it is absorbed by the new grains, i.e., that the energy stored in the metal during cold working is not greatly decreased by recovery, prior to recrystallization. Accordingly, the bulk of this energy is assumed to serve as the driving force for the process of recrystallization. Recently, a great deal of evidence accumulated, indicating that the traditional concept may be considerably in error. Suzuki' found that by far the largest portion of the stored energy of cold work in compressed polycrystalline copper is released in the early stages of annealing, well in advance of recrystallization. Thus, the new recrystallized grains are undoubtedly growing in a matrix which is, at least energetically, no longer in the cold worked condition, but has nearly returned to the annealed state. It is true that in Suzuki's experiments the major hardness changes due to annealing were observed to take place concurrently with the formation and growth of the new grains, but this process apparently involved the release of only a minor residual fraction of the stored energy of cold work. Confirming in this respect the views of Wood," several recent investigations by various techniques lead to the conclusion that the grains of many, if not all, metals in the cold worked condition consist of small, slightly disoriented "crystallites" or "sub-grains".* (After appreciable deformation at room temperature, the subgrain size in copper appears to be about 7.104 cm and in high purity aluminum about 2.10-4 cm.) Recent investigations which demonstrated this phenomenon most convincingly are those of Heidenreich8 with hammered high purity aluminum, of P. B. Hirsch and collaborators7,8 with cold worked polycrystalline aluminum, copper, nickel, iron, gold, etc., and of Warren and Averbach9 with a rolled copper plus 2 pct Si single crystal. In addition to the existence of slightly disoriented sub-grains, the cumulative evidence from these investigations also shows that in the cold worked condition the subgrains themselves are severely strained. With regard to the early stages of annealing, Heiden-reich's results indicate that within the subarains of hammered high purity aluminum the strain gradually disappears, without discernible migration of the subboundaries. This recovery process presumably accounts for the major portion of the energy release

Jan 1, 1953

By C. R. Dale

IT IS the purpose of this paper to point out a number of the most common causes of crooked holes; to outline methods of drilling and straightening which to my personal knowledge have proved successful in numerous cases; to describe some of the characteristics of holes which, though not crooked in an exaggerated sense, may still cause serious mechanical trouble if not recognized and properly remedied; and to define clearly the part of oil-well surveying in the concerted effort operators generally are making to keep their wells straight from spudding to completion

Jan 1, 1931

By T. F. Archbold, W. H. King

THE diffusivities of gold in polycrystalline copper have been determined in the temperature range from 700° to 1000°C by the utilization of the neutron-activation method. Martin, Johnson, and ~saro' used this method for their investigation of the diffusion of gold into copper at 375° and 550°C for 59 days in order to circumvent the experimental difficulties imposed by the 2.7-day half-life of Au-198. They employed the conventional technique of depositing Au-198 on copper specimens for short-time dif-

Jan 1, 1965

By D. R. Parrish, V. J. Berry

Reverse combustion is one thermal method of recovering hydrocarbons from porous undergrortnd formations containing oil or tar. In applying this method, air is introduced via an injection well and the mixture of air and hydrocarbons is ignited in the production well. A combustion zone then recedes toward the injection well, counter-current to the air flow. If the combustion-zorie temperature is sufficiently high, the oil or tar irz place is distilled arid cracked. The hydrocnrhon flows as a vapor to the production well and subsequently is condensed at the surface. Maximum temperature and velocity of morcment arc the two dependent variubles defining the progress of the combustion zone. A theoretical analysis has been made of heat flow in the reverse-comhustion process as.srtrning linear flow in a homogeneous system. The differentirll equatious, which include the oxygen-hydrocar-hon rctrction rate, have been solved numerically. Rescllts indicate that the maximum temperature reached and he cornbustion-zone relocity both increase with an incrense in air-injection rote. Heat loss to surroundings has little effect on the maximum comhustion-zone ten?perature achieved, but it is reflected in a reduced com-bustion-zorzc, velocity. It is also predicted that an increase in the oxygen-hydrocarbon reaction rate results in a reduction in the maximum temperature reached. The calculated results are in agreement with results from reverse-combustion experinments using samples of a tar sand. INTRODUCTION Reverse combustion is one thermal method of recovering hydrocarbons from porous underground forma-tions containing oil or tar.' In applying this method, air is introduced into the underground formation via an in-jection well. In one or more adjacent production wells, the mixture of air and hydrocarbons is ignited at the sand face. The combustion zone thus formed recedes toward the injection well, counter-current to the air flow (Fig. 1). Under certain conditions, the rate of recession of the combustion zone is sufficiently rapid that only a small fraction of the hydrocarbon at any point in the formation is consumed by combustion. If the combustion-zone temperature is sufficiently high, the uncon-sumed oil or tar in place may be partially distilled or cracked. This material then can flow as a vapor with the gaseous combustion products to the production well and subsequently be condensed at the surface. Under other conditions, the combustion-zone temperature may be so low that little, if any, of the hydrocarbon is vaporized. Nevertheless, the viscosity of the hydrocarbon may be reduced enough so that other recovery methods subsequently can be applied to produce oil at attractive rates. The maximum temperature achieved in the combustion zone and the velocity of movement are the two dependent variables defining the progress of the combustion zone. Presumably. knowledge of these factors will permit estimates to be made of oil recovery efficiency, oil and gas producticn rates and average amount of injection air required per barrel of oil recovered. The maximum temperature and combustion-zone vclocity are, in turn, expected to be dependent upon air flow rate, heat losses to surrounding formations, physical properties of the rock and the hydrocarbon. the extent to which the porous formation is saturated with the hydrocarbon, the rate of oxidation of the hydrocarbon and the amount of heat released. It is desirable that the relative importance of each of these factors be understood so that the results of laboratory tests of oil recovery by reverse combustion can be properly interpreted in terms of expected field performance. It is particularly desirable to assess the importance of heat lcsses because these are difficult to control or eliminate

By E. Cortese

Phosphate occurs in many places in Egypt, in two main zones: one in Upper Egypt, along the Nile Valley, principally on the right side, and one near the Red Sea coast. In the Nile zone, the principal layers of phosphate rock are to be found: in the mountains east of Mahamid (right side of the valley); in the hills near Sibaiya, also called Sebaieh (both sides); in the hills of Agula, Hagaza and Gurn (right side); and further on, along the camel road leading from Kena to Kosseir, a few miles east of Legheita's Oasis. Following that track, after having crossed the watershed, underlain by rocks like granite, serpentine, porphyry, and diabase, one comes again to the Cretaceous formation containing considerable quantities of phosphate, at Gebel Duwi, Gebel Nakheil, and Gebel Kosseir El Kadim, not far from the coast. More to the north, in the hills near Safaga Bay (Um El Uettat El Wasif) is another rich bed of the same formation. There are reports of phosphate existing near Suez and in the Sinai peninsula, and possibly some other zones can be traced or discovered in the future. The mines are well developed on the right bank of the Nile at Sharawna (near Sebaieh Station) and along the Red Sea coast, at Safaga and near Kosseir. Prospecting work has been done at Gebel Duwi, Gebel Gurn, Agula, Hagaza, and near Mahamid. The phosphatic formation in Egypt is quite variable, not only from place to place, but even within small areas. The phosphate is all of Cretaceous age, as is proved by the fossils found in it. It is considered to belong to the Campanian or, more appropriately, to the Maestrichtian series. The principal fossils are: Alectryonia Villei Coq. sp.; Roudereia Aures-sensis Coq. sp. = R. Drui Mun. Chalm.; Trigonoarca multidentata, Newton.; Cardita libyca, Zittel; Libycoceras Ismaely, Zittel; Nautilus sp.; Ancystrodon libycum Dam.; Corax pristodontus Ag.; Lamma biauriculata Zittel sp. Above the Maestrichtian beds, we have the Danian series, unconform-ably covered by Lower Eocene beds. Under the phosphate is a grayish marly clay, and beneath this, bluish-gray scaly shales resting on the Nubian sandstone (Santonian).

Jan 1, 1918

By V. F. Zackay, W. A. Goering

ALLOYS of iron and aluminum up to 35 wt pct aluminum are single-phase solid solutions, and are of potentially wide applicability.1-3 In spite of early and continued interest1-4 little progress has been made until recently in the preparation and evaluation of sound alloys containing more than 6 wt pct aluminum. Vacuum-melting techniques for the production of ductile Fe-A1 alloys have been described recently.1-7 A. procedure for air melting these alloys is presented here. Low-carbon iron is induction melted without a slag in a rammed magnesia crucible. At the beginning of melt-down, aluminum pig (99.95 pct Al), charged in a clay-graphite bottom-pouring crucible is placed in a pot furnace at 1800°F. The primary deoxidation of the molten iron after melt-down is effected by the addition of 0.1 pct aluminum and 0.5 pct manganese. (Hilty and Crafts" have reported a significant increase in the deoxidation efficiency of the aluminum and manganese combination over that of the aluminum alone.) A more drastic deoxidation designed to reduce the oxyen content to the lowest possible level is accomplished by plunging metallic calcium to the bottom of the melt. This is done by wiring small cubes of the metal to a steel rod. A circular shield larger than the diameter of the crucible opening is attached to the rod so that any spa'ttering of the molten metal will not endanger the operator. Since the temperature of the molten metal is above the boiling point of calcium, the bath is vigclrously purged by calcium vapor. It is believed that the calcium-vapor treatment permits a homogeneous distribution of calcium in the melt. Owing to the vigor of the reaction the temperature of the molten metal should be kept below 2900°F prior to the calcium addition. A total of 0.05 pct calcium is added in two stages in this manner. The second calcium deoxidation is made just before charging the molten aluminum into the iron, in order that an excess of calcium be present for the remainder of the melt. The aluminum, which has been removed from the holding furnace, is then hydrogen degassed by bubbling chlorine through a quartz tube immersed in the molten aluminum. The hydrogen-chlorine reaction is an exothermic one preventing the solidification of the aluminum during the 5-min chlorination. Approximately 0.1 pct calcium, based on the amount of aluminum, is then added to the aluminum. A further excess of calcium is introduced into the melt in this manner. The oxide dross is removed, fluorspar is added to the molten iron, and the molten aluminum is poured through the fluorspar slag. The fluorspar should be dried thoroughly prior to its use, as any water present will react with the aluminum. Aluminum oxide formed during the pouring operation reacts with the fluorspar slag to form gaseous aluminum fluoride and calcium oxide. A forced-draft ventilating system is required for this operation as aluminum fluoride is toxic. As soon as the molten aluminum has been added, vigorous manual stirring of the melt is required because the slag-aluminum oxide reaction is highly exothermic and tends to take place near the top of the melt. The combination of high temperature and the slagging action of the fluorspar quickly erodes the crucible at the slag line if the aluminum is not stirred uniformly into the melt. It has been found that at least 4 min of manual stirring combined with induction stirring are necessary to ensure homogeneity. The power is shut off 1 min prior to pouring to allow metal and slag to separate. As much slag as possible is removed from the melt, which is then poured directly into cast-iron molds. A mold wash of aluminum oxide is used to prevent ingot sticking. For slab ingots which are to be rolled into sheet, a carbon-tetrachloride vapor atmosphere or a chlorinated-pitch mold wash is desirable, as the aluminum oxide formed in the pouring operation is subsequently removed by the chlorine in the presence of carbon." As in vacuum melting, a pouring temperature of about 2900°F is recommended. Adequate hot-topping is important as iron-aluminum ingots are subject to very deep piping. Ingots are removed from the molds and buried in vermiculite, where they are allowed to cool slowly to room temperature. The ingots are radiographed,

Jan 1, 1959

By M. E. Wadsworth, C. H. Pitt

Nickel corrosion in sulfuric acid solutions at elevated temperatures and oxygen wer-pressures was imestigated. Weight loss was 1inear with time and varied directly with oxygen concentration. Independent variation of the possible active species in solution indicated that the undissociated sulfuric acid molecule plays an important part in the corrosion mechanism. The rate determining step was considered to be the adsorption of oxygen on a site containing adsorbed H2SO4. THE tendency for a metal to go into solution depends on a number of factors some of which are: 1) the position of the metal in the electrochemical series; 2) the type and concentration of corroding species in solution; 3) the concentration of oxygen in solution; 4) the type of film formed at the liquid-metal interface; and 5) the physical state of the metal. The corrosion of a metal in an aqueous solution may be considered to be a heterogeneous reaction that occurs at the liquid-metal interface. The reaction is generally considered to occur as a result of electrochemical potential differences within the system. Little work has been published in the literature concerning the corrosion of pure nickel. A great deal has been publistled dealing with the experimental and theoretical aspects of the over-all corrosion process. Evans,1 Uhlig,' and recently Todt 3 and Tomashoff,4 have published books which review the results of experiments which have been conducted on corrosion, and have developed some of the basic ideas of the more widely accepted theories of corrosion. Friend5 states that. nickel does not readily discharge hydrogen from any of the common nonoxidiz-ing acids and that a supply of some oxidizing agent is necessary for appreciable rates of corrosion. He further notes that at temperatures between atmospheric and close to boiling, corrosion rates tend to vary with concentration, temperature, and aeration in acid concentrations up to 25 pct. Whitman and Russel6 studied the effect of dissolved oxygen on a number of metals in acid solutions and found that oxygen accelerates the rate of corrosion of nickel in sulfuric acid. Flowers and Kelley7 in studying the action of dissolved CO, on electroplated nickel postulated that nickel dissolves as NiHCO3 in the presence of CO, and air. It was found that oxygen was necessary for corrosion to take place and that the rate increased with increasing amounts of oxygen in solution. Flowers and Kelley attributed the oxygen effect to cathodic depolarization. MacGillavry, Rosenbaum, and wensson' determined metal solution potentials of nickel in various electrolytes. They found that nickel has no tendency to corrode in alkaline solutions in the presence or absence of air. In acid solutions in the presence of air, metal solution potential increases in the order of 1/2 v were observed. This increase was attributed to probable chemisorption of oxygen on the nickel surface and the corrosion mechanism suggested was the interaction of H+ ions with the nickel oxide formed at the surface. Delahay9 observed the reduction of oxygen on various metals. A polarograph was used to measure the course of the reaction. With some metals hydrogen peroxide was produced as an intermediate compound. However, nickel was not considered to be a hdrogen peroxide producer. Van Rysselberghe 10,11 and associates attributed the accelerating effect of carbon dioxide on the corrosion process to the presence of percarbonic acid produced in solution reacting with the hydrogen peroxide. The percarbonic acid can be reduced at less negative potentials than hydrogen peroxide and as a result accelerates the reaction. EXPERIMENTAL Since a complete description of the apparatus used can be obtained elsewhere,12'13 no attempt will be made here to describe the equipment other than to note that it consisted essentially of an autoclave unit from which samples could be withdrawn periodically without disturbing the interior state of the autoclave. A titanium autoclave liner was used to protect the autoclave from the solution.

Jan 1, 1961

By D. W. Fuerstenau

Recently R. J. Charles1 showed that comminution of brittle or semi-brittle materials in batch operations is described more appropriately by a variable energy relationship than by the specific relation- ships proposed by Rittinger,2 Kick,3 and Bond.4 That is, the value of n in the empirical differential operation which applies to crushing and grinding (Eq. 1) is variable. Independently, Holmes6 developed an identical generalized equaton for size reduction. The experimental data of Charles and Holmes show that the value of n is determined by the material and the comminuting machine and is constant for a given combination of these two conditions. The general objective of the present article is to extend the Charles-Holmes concepts to continuous ball milling. The specific objective is to analyze the operation of a vibratory ball mill dry-grinding chromite in a continuous open circuit and to develop an equation for the comminution of chromite in this mill.

Jan 12, 1959

By Shiou-Chuan Sun

THE effects of oxidation on the flotation behavior of sulphide minerals have been extensively studied,1-3 but no similar study has been made of coals. Coals of bituminous and lower rank undergo atmospheric oxidation during mining and storage, and their degree of oxidizing increases with temperature4,5 and time of exposure.6 The problem of recovering fine oxidized coal from culm piles, rivers, and streams by froth flotation has become increasingly important to the coal industry as well as to those concerned with air and stream pollution. This paper describes the flotation behavior of oxidized and unoxidized coals and carbons in the presence of oily collectors and presents the reagents found effective in the flotation of oxidized coals and carbons. Except where otherwise stated, flotation tests were performed in a 50-g Denver flotation cell. Solid feed for each test was 30 g, and the pulp dilution was kept at 8.5. The -65 mesh coal sample was wetted by agitation with water in the flotation cell for 6 min, then conditioned with a predetermined amount of collector for 2 min, and finally floated with pine oil as frother for 3 min.

Jan 4, 1954

By D. W. Fuerstenau, D. A. Sullevan

In the experimental work for this comparison of wet and dry grinding, it was found that the size distributions for wet grinding operations are characterized by a constant value of the distribution modulus, (a; whereas, for dry grinding the distribution modulus decreases slightly with increased size reduction. Wet grindability seems independent of the weight of material in the mill while dry grindability depends on weight. Experimental data and eqmtions for each process are presented for noncleavable isotropic material, for cleavable or material in which there is some form of internal stress, and for energy consumption. The equation ED/EW=BK-? (where ED and Ew are, respectively, energy for wet and dry grinding, k is the size modulus, and B and ? are constants) was found to give the ratio of energy. Dry grinding is generally considered to require more energy than wet grinding. Taggart' states that power consumption in wet grinding ranges from 60 to 90 pct of that required for dry grinding. Bond' states that dry grinding requires one third more power than wet grinding. It is interesting that Rose and Sullivan' note this same difference in power required for wet and dry grinding, but they state that on purely mechanical grounds it is difficult to see any great difference in fundamental principles between wet and dry milling, since dry milling may be regarded as wet milling with a fluid of low viscosity (i.e., air vs water). Yet it must be the viscosity of the fluid that is responsible for the differences between wet and dry grinding. Probably the most comprehensive investigation comparing wet and dry grinding has been that made by Coghill and Devaney. They compared the grind-ability of both chert and dolomite in a 19x36-in. batch ball mill under both wet and dry conditions. In these tests the weight of ore in the mill was varied from 35 to 200 lb and each grinding test was carried out for such period of time that the amount of -200-mesh material in the product was essentially the same for each test with a given material. Cog-hill and DeVaney's data for the grindability of dolomite, which are reproduced in Fig. 1, show that a comparison between wet and dry grinding is com- plex. The comminution of a material by a wet grinding method is characterized by a constant grindability over a wide range of mill loadings. Since grindability is the weight of material comminuted per unit of energy expended, it is equivalent to (w)/ (p)(t), where W is the weight of material being ground, P is the power, and t is the grinding time. In the case of wet grinding, the grindability of dolomite in the Coghill and DeVaney experiments was 0.546 * 0.009 ton per hp-hr when the load of dolomite in the mill ranged from 50 to 200 lb. On the other hand, for dry grinding no such simple relationship was found to exist between power, time, and weight of material in the mill. Dry grinding efficiency appears to depend greatly upon the amount of material within the mill. When the mill contained

Jan 1, 1961

By K. Sassa, I. Ito

The dynamic principal stresses induced within rock by an explosion under con fined conditions are analyzed for the case of blasting with one free face by applying the values of measured radial displacement, particle velocity and so on to equations derived from the theory of elasticity. The dynamic stresses on the free face are also calculated from the strain measured by wire strain gauges affixed on the free face. Some results obtained are briefly: 1) The maximum stress on the free face appeared at the intersection between the free face and the burden, and its magnitude was about twice as large as the magnitude of the tangential tensile stress imparted by the incident stress wave within the range related to the breakage by blasting. 2) The maximum value of the compressive stress near the free face was decreased due to the existence of the free face, while that of the tensile stress was increased. 3) At any point near the free face, the directions of the principal stresses in the r? plane using spherical coordinates with the origin at the center of the spherical charge, changed with time except in the direction of the burden. However, at another point near the explosion point where the principal stresses were mainly those stresses caused by the initial longitudinal wave, the directions of the principal stresses coincided with those of the coordinate used. 4) The maximum value of the tensile stress in the F direction was larger than those in the re plane at and near the free face. 5) The tensile stress caused by the Hopkinson effect appeared only near the burden. 6) It is presumed from the results of stress analysis that radial cracks will be produced in the rock mass near an explosion point, but with the approach of the free face, the directions of cracks curve towards a direction parallel to the free face. INTRODUCTION In order to investigate the mechanism of rock breakage by blasting, it is considered to be very im- portant to know how the stresses are generated and distributed within the rock by explosion of a confined charge. As part of a research program concerning the mechanism of rock breakage by blasting, the authors have already published papers on the detonation pressure produced at the inner surface of the charge hole1 and on the behavior of the dynamic stresses within rock caused by an explosion in the case where no free face existed.' In this study the dynamic stresses on and near the free face have been investigated for the case of blasting with one free face. The mechanism of rock breakage in blasting is discussed in the light of these studies.

Jan 1, 1967

By Gray L. L., Max Littlefield, Godbold A. C.

The West Edmond pool of Central Oklahoma, a limestone reservoir, has an area in excess of 29,000 acres and as of Sept. 15, 1946, had produced 53 million barrels of oil from 731 wells at an average depth of 6900 ft. Water has encroached into the reservoir along the west side of the pool and although the area of water invasion is large the net volume of water influx is small, because encroachment has been primarily into a fracture system. The most significant part of this study concerns the character of the Hunton reservoir. Geological study of cores of the producing section was made to supplement core-analysis data. It was determined that approximately 90 pct of the pore volume was contained in intergranular or sand-like porosity and 10 pct in fractures (intermediate porosity). The porous limestone is divided into blocks by the fractures, although in the strongly dolomitic parts of the producing section, fractures are less well developed. The low-permeability intergranular porosity, largely subsidiary to the high-permeability fractures, will produce oil into the fractures by evolution of solution gas. The fractures in effect serve as drainage channels for production of oil from the inter-granular porosity. Because of such a large difference in permeabilities of the two com-ponents, a high degree of by-passing will occur, and, accordingly, the economics involved in undertaking full-scale high pressure gas injection operations have been seriously questioned. Introduction The West Edmond Hunton pool, which covers portions of Canadian, Logan, Oklahoma, and Kingfisher Counties in Central Oklahoma, is the largest reservoir producing oil from a limestone in the state. The total area of the pool comprises about 29,240 acres and development is considered to be essentially complete. In all, 731 producing wells have been completed in the Hunton reservoir at an average depth of approximately 6900 ft, and have yielded a cumulative oil production in excess of 53 million barrels as of Sept. 15, 1946. In addition to wells completed in the Hunton limestone, 21 wells have been completed in the Bartlesville sand at a depth of approximately 6500 ft, which have produced a total of 730,000 bbl of oil. One small producer has been completed in the Cleveland sand at a depth of 5700 ft. However, a study of only the Hunton limestone reservoir will be presented in this paper. The most significant phase of this study concerns the special type of geological core examination that was made in an effort to ascertain the type and degree of porosity in the rock in accordance with lithologic types. Results of this geological study showed the west Edmond pool to be of such a complex nature as to appreciably affect the behavior normally associated with homogeneous reservoirs. Recognition of this condition in the

Jan 1, 1948

By Max Littlefield, Godbold A. C., Gray L. L.

The West Edmond pool of Central Oklahoma, a limestone reservoir, has an area in excess of 29,000 acres and as of Sept. 15, 1946, had produced 53 million barrels of oil from 731 wells at an average depth of 6900 ft. Water has encroached into the reservoir along the west side of the pool and although the area of water invasion is large the net volume of water influx is small, because encroachment has been primarily into a fracture system. The most significant part of this study concerns the character of the Hunton reservoir. Geological study of cores of the producing section was made to supplement core-analysis data. It was determined that approximately 90 pct of the pore volume was contained in intergranular or sand-like porosity and 10 pct in fractures (intermediate porosity). The porous limestone is divided into blocks by the fractures, although in the strongly dolomitic parts of the producing section, fractures are less well developed. The low-permeability intergranular porosity, largely subsidiary to the high-permeability fractures, will produce oil into the fractures by evolution of solution gas. The fractures in effect serve as drainage channels for production of oil from the inter-granular porosity. Because of such a large difference in permeabilities of the two com-ponents, a high degree of by-passing will occur, and, accordingly, the economics involved in undertaking full-scale high pressure gas injection operations have been seriously questioned. Introduction The West Edmond Hunton pool, which covers portions of Canadian, Logan, Oklahoma, and Kingfisher Counties in Central Oklahoma, is the largest reservoir producing oil from a limestone in the state. The total area of the pool comprises about 29,240 acres and development is considered to be essentially complete. In all, 731 producing wells have been completed in the Hunton reservoir at an average depth of approximately 6900 ft, and have yielded a cumulative oil production in excess of 53 million barrels as of Sept. 15, 1946. In addition to wells completed in the Hunton limestone, 21 wells have been completed in the Bartlesville sand at a depth of approximately 6500 ft, which have produced a total of 730,000 bbl of oil. One small producer has been completed in the Cleveland sand at a depth of 5700 ft. However, a study of only the Hunton limestone reservoir will be presented in this paper. The most significant phase of this study concerns the special type of geological core examination that was made in an effort to ascertain the type and degree of porosity in the rock in accordance with lithologic types. Results of this geological study showed the west Edmond pool to be of such a complex nature as to appreciably affect the behavior normally associated with homogeneous reservoirs. Recognition of this condition in the

Jan 1, 1948

By AIME AIME

A separate room in the suite occupied by the American Institute of Mining Engineers on the ninth floor of the United Engineering Society Building, has been equipped with furniture and telephone extension for the temporary use of members of the Institute or of sister societies, or visitors suitably accredited. .In addition to this general accommodation, another room has been similarly furnished as a business headquarters for a single person, and for a limited period. For the conditions of this privilege, inquiry should be made at the office of the Institute. . In order that the benefit of this arrangement shall not be restricted to a few members only, the period of continued occupancy will be limited, and applications will be considered in the order of their receipt. A small. fee will be required for the use of the facilities furnished. ANNOUNCEMENT. THE BIRMINGHAM MEETING. - The XCIVth Meeting of the Institute for the reading and discussion of professional papers will be held at Birmingham, Ala., beginning about October 1. After the Birmingham sessions and excursions, it is proposed to return by special train via Chattanooga, Ducktown, etc., to Washington, D. C., where additional sessions will be held. Further particulars will be given in a special circular, to be issued at an early day.

Jan 5, 1908

By E. D. Peters

It frequently occurs in the establishment of reduction works, in an entirely new and untried mining district, that the metallurgist in charge finds considerable difficulty in determining the process best adapted to the ores which he receives for treatment. At the first glance it would seem easy enough to decide what style of furnace is best adapted for beneficiating any one class of minerals. If the ore possessed a quartzose gangue and was comparatively free from base metals, while salt could be obtained at a reasonable price, one would naturally resort to a chloridizing roasting, and pan or barrel amalgamation for the extraction of its silver contents. If galena or carbonate of lead was the prevailing mineral, and charcoal could be obtained at moderate figures, the blast-furnace would beneficiate the ores most advantageously. But if, as is usually the case in Colorado, the ores consisted of an intimate mixture of galena, zinc-blende, copper-pyrites, and noble silver minerals, associated with an overwhelming mass of siliceous heavy spar, or limestone gangue, the common reverberatory furnace can be used to the greatest advantage ; and, although the product is only a copper matte, I do not hesitate to affirm that it can be treated, and the silver, gold, and copper produced at nearly the same price for which silver and lead can be separated. In Germany, blast-furnaces are frequently used for this same purpose, both for argentiferous and non-argeutiferous copper ores; but any one who will take the trouble to examine the statistics of smelting at Fahlun, the Oberharz, and other Continental works of this description, will see that the expenses are far too great, and the production much too small to think of employing this method in our mining districts, where only the softest and most miserable kind of charcoal is obtainable. I propose in this paper to give an accurate account of the expenses incurred at the Mount Lincoln Smelting Works in treating the same kind of ore, and producing the same end-product in both blast and reverberatory furnaces. I have taken for comparison a favorable campaign of the blast-furnace, and a good average month's running of the reverberatory. In the blast-furnace estimate I have included the calcining and concentration of the matte produced during that campaign, as I always, when possible, concentrated the regulus made

By G. M. Schwartz, J. N. Gundersen

John W. Gruner (Professor Emeritus, College of Science, Literature, and the Arts, Dept. of Geology, University of Minnesota, Minneapolis) — It is surprising that two men as well acquainted with the taconite rocks of the Mesabi District as Drs. J. N. Gundersen and G. M. Schwartz should have concluded that the taconite of the Eastern Mesabi before the intrusion of the Duluth gabbro consisted largely of quartz and magnetite. These minerals were "locally reconstituted" into "a wide variety of silicate mineral assemblages" by the gabbro. It is well known, of course, that the silicate mineralogy of the Eastern Mesabi ('about 14 miles in length) is quite different from that of the Main Mesabi which is over 60 miles long. The change from one to the other is abrupt and cannot be attributed to any other cause than the thermal effects of the gabbro. Are we to assume or believe that the taconite of the Eastern Mesabi was radically different from that of the Main range before the gabbro metamorphosed it?

Jan 1, 1962

By Janshi Sen

VERTICAL-INTENSITY magnetometers, such as the Hotchkiss Superdip and the Askania vertical field balance, are now [ ] widely used, because vertical-intensity charts give definite information for the location of upper pole, dip and strike, and approximate shape of the deposit. Again, a slow and steady increase of dip over a large area indicates a large deep-lying deposit, while a large dip limited to a small area and falling away rapidly suggests a small and shallow one. The complication comes more from the richness of the ore. A large body of lean ore near the surface might produce the same vertical-intensity chart as a rich ore at depth. This paper presents a method for calculating the maximum depth of a single-pole deposit and another for a two-pole mineral body, by utilizing data of both vertical and horizontal intensities as obtained by a Thalen-Tiberg magnetometer. As shown on the left side of Fig. 1, lean ore, rich ore and a single pole produce the same magnetic field. The depth of the pole, [ ] h, can be calculated from the following relation: [h=a=tan¢=G-F h=aXFy] where G is the vertical intensity at point A due to the pole; F, the horizontal intensity of the same; and a, the distance from the point of maximum vertical intensity or the vertical projection of the pole on the surface to the point of observation A. If an entirely covered deposit is indicated by magnetic measurements and boring cost estimated from the calculated depth is so reasonable as to encourage drilling, the ore should be struck somewhere within the limit of the calculated depth by the borehole. A number of such points as A should be taken for calculation, erratic results should be rejected and the average figure used for cost estimation.

Jan 1, 1944

By Charles White Merrill

Mankind's progress is measured in minerals. Man's emergence from prehistory is marked by passage through a Stone Age and a Bronze Age and into the present era, sometimes called the Iron Age or the Age of Metals. With the evolution of civilization and the development of a modem industrial economy, minerals become more and more important. Structures, roads, and machines are built largely of minerals; the energy to heat the structures and to drive the machines is mostly mineral; agricultural fertility is maintained by mineral fertilizers and water; and military strength is provided by fortresses, armaments, and munitions, largely of mineral origin. Thus in the material world, in the productivity of the economy, access to minerals has paramount importance. Every evidence indicates increasing dependence on minerals in the modem economy, particularly on the mineral fuels as energy sources to provide the power base for advancing industrialization. Both the total and the per capita consumption of almost every mineral is increasing in the industrialized nations. Moreover, nations with predominantly agrarian economies call themselves underdeveloped and universally strive to industrialize. Most mineral consumption trends in the more advanced economies have been greatly exceeding population growth-which itself has been advancing at an accelerating rate-and sometimes have surpassed the growth in real income. Some minerals have only recently been adopted or grown to importance in use. Titanium metal did not appear as a commercial product until 1948, and even coal seriously displaced wood as a source of heat less than two centuries ago. On the other hand, minerals have always played an important role in construction. The prehistoric megaliths that form the dolmens of Salisbury Plain are stone; the pyramids are andesite faced with limestone; and many other ancient structures are built of mineral materials. In modem times the whole course of architecture has been redirected by new building materials, mostly mineral. Most conspicuous is the sky-

Jan 1, 1964

By Arthur H. Wilson

INTRODUCTION The stratified deposits of the coal measures are strongly laminated and exhibit a high degree of anisotropy. The stronger rocks in the sequence possess joints and shear planes, the coal itself has a closely spaced cleat pattern, and the softer rocks are capable of flowage. To calculate the general distribution of stress in such a medium by analytical means based on the theories of elasticity is difficult, and has not yet been satisfactorily achieved. Nevertheless, some estimate is required to aid the design of supporting pillars and permit the correct siting of main roadways. This is important when laying out districts for longwall extraction, particularly if the results of previous experience are not available in the area to be exploited. Although elastic stress analyses, including finite element analysis, may not be very appropriate, the basic equilibrium conditions will apply. These, combined with logical deductions based on observed fact, can be used to give appropriate solutions which experience has shown to be capable of giving acceptable results in the design of pillars associated with longwall workings. Constancy of cover load is one of the fundamental equilibrium conditions. When mining below ground, the cover is not removed, therefore the average vertical load after mining must equal the initial cover load. If mining produces a local decrease in load in any particular area, it must be balanced by an equivalent increase in load nearby. Bending moments in the strata above are capable of compensating for the turning moments which this redistribution of load may produce, and the existence of a "rear abutment", much discussed between 1950 and 1960, is not an essential equilibrium requirement. If the distribution of vertical stress is plotted three-dimensionally, then the load on any small element of area 6S will be o.6S, where o is the average stress. This also represents the volume of the "stress envelope" which lies above 6s. By making the individual areas very small and summing together, it is possible to show that the total load on any particular area can be found by calculating the volume of the stress distribution envelope which lies above it. A similar consideration applies to the stress distribution in a two-dimensional situation. If the stress distribution curve is drawn, then the area below any portion will represent the load per unit depth of section. For example, Figure 1 represents the possible stress distribution across a caved waste and associated ribside. If AW is the load deficiency [ ] These loads must balance, therefore area AW = AS, and if one can be deduced, the other will then be known. Attempts have been made to measure stress distributions in the ribside and in the waste. Stress meters in soft rocks such as the coal measures have not proved very effective. Those based on the measurement of strain have experienced difficulty in strata capable of flowage, high modulus stress meters suffer from cross sensitivity, and overcoming to obtain absolute stress is almost impossible. Load measuring instruments left in the waste have also proved largely ineffective, mainly as a result of the difficulties of protecting the leads in a caving environment. Certain logical deductions, however, can be made about the conditions in the waste. CONDITIONS IN THE WASTE Over the area of the waste the surface trough which develops is remarkably consistent, despite the geological variations which occur from coalfield to coalfield. In Britain, an accuracy of 210% is claimed for subsidence prediction based on width of extraction and depth of cover (National Coal Board, 1975). The subsidence in the USA is more variable (Adamek and Jeran, 1981), particularly with respect to the point of inflexion, but the same general pattern occurs. At the shallower depths the somewhat harder roof strata involved in the caving in the USA may suffer less compaction, giving reduced values of subsidence. Subsidence at intermediate horizons below the surface

Jan 1, 1982

By Ernest Burchard

FLUORSPAR is found in most of the states from the Rocky Mountains westward, and commercial production of the mineral has been reported from Arizona, Colorado, Nevada, New Mexico, Utah and Washington. The map in Fig. 1 indicates the general distribution of many of these deposits but it is by no means complete. In the summer of 1926 the writer examined certain deposits of fluor-spar in Arizona, California, Colorado, New Mexico and Washington; in November, 1927, revisited the Jamestown, Colo., district, and in earlier years examined deposits in Colorado and New Mexico. Based on notes of his own and. on descriptions and other data by H. A. Aurand, H. W. Davis, R. B. Ladoo, V. C. Heikes, W. D. Johnston, Jr., and by producers of fluorspar, he prepared the following brief descriptions of deposits and estimates of reserves of fluorspar, originally for the use of the subcommittee on fluorspar of the Joint Committee on Foreign and Domestic Mining Policy and on Industrial Preparedness respectively of the Mining and Metallurgical Society of America and the American Institute of Mining and Metallurgical Engineers. Descriptions of many fluorspar deposits in various western states and districts have been published. It is not feasible to summarize this literature here, but the appended bibliography will indicate to the interested reader what papers are available.

Jan 1, 1933