Search Documents

By R. P. Smith

The carbon content of a number of alloys ranging (before carburitation) from purified iron to purified cobalt was determined for fixed activities of carbon by equilibration with each of six CO-CO2 atmosphere at 1000°C. At fixed carbon activity, the carbon content decreases continuously with increasing cobalt content. For concentrations of 0 to 50 at. pct Co the results can be expressed by a linear interaction equation of the type In y2 = 7.76n2/(n1+n3)+ 2.30 n3/(n1+n3) where y2 is the activity coefficient of carbon (standard state such that y2— 1 as n2 — 0 in the Fe-C binary) and n1, n2, and n3 are the number of atoms of iron, carbon, and cobalt per unit weight of alloy. The results are also given in terms of the a functions. In the same concentration range, the results are in accord with the regular solution theory of Alcock and Richardson. THE carbon content of Fe-Si, Fe-Mn, and Fe-Ni alloys at fixed activities of carbon has been reported previously.'-, The present work extends R. P. SMITH, Member AIME, is Research Chemist, E. C. Bain Laboratory for Fundamental Research, United States Steel Corp., Research Center, Monroeville, Pa. these measurements to a series of Fe-Co alloys. These alloys, like the Fe-Ni alloys, are fcc ("aus-tenitic") over the entire range of composition at 1000°C. EXPERIMENTAL PROCEDURE The alloys were made from electrolytic cobalt and Plast-Iron in a small vacuum-melting furnace. A small amount of graphite was added to each melt to keep the oxygen content low. The ingots were analyzed for iron and carbon: the amount of cobalt was obtained by difference. Analysis of the top and bottom of the ingots indicated that they were uniform in alloy content. Except for the pure cobalt the ingots were hot-rolled to approximately 0.15 cm thickness, cleaned by sandblasting, then cold-rolled to about 0.045 cm. We were not successful in cold rolling the pure cobalt; therefore it was hot-rolled to the final thickness. The last traces of surface oxide were removed by acid pickel, then the samples were treated in a hydrogen-2 pct water vapor mixture at 1100°C to remove carbon and most of the oxygen. The carburizing procedure was similar to that previously described.1,2,4 The alloys in the form of strips about 1.7 cm wide, 7.5 cm long, and 0.04 cm thick, together with a similar specimen of purified iron, were equilibrated with each of six CO-CO2 mixtures of fixed composition, then quenched in the

Jan 1, 1965

By C. A. Herbert, C. M. Young

WithIn the past few years, considerable development has been made in the coal-mining industry in Illinois and Indiana and it is the purpose of the authors to record its most important phases. Perhaps the two most striking features are the entry into the producing fields of certain large consumers of coal and the magnitude of some of the new operations. Mines are now being equipped by the Chicago & Northwestern R. R., the Chicago, Burlington & Quincy R. R., the Standard Oil Co., of Indiana, and the Union Electric Light & Power Co. of St. Louis. Besides, the U. S. Steel Corpn. has increased its coal-mining activities by commencing operations on its large holdings northeast of Benton, Franklin CO., Ill. Until recent years, the supply of coal has commonly been adequate and prices have generally been favorable, especially to large consumers, as the magnitude of their purchases enabled them to obtain satisfactory quotations from producers. In fact, most consumers of coal have believed that they could buy their fuel more satisfactorily in the open market than they could produce it themselves. The entry of the United States into the war was accompanied by various disarrangements of industrial conditions, and a demand for coal in excess of the supply seemed likely to be experienced for an indefinite period. Under these circumstances it seemed that the greatest assurance of a supply of fuel was its production by the operation of mines. In the case of the Chicago & Northwestern R. R. and the U. S. Steel Corpn. the production of coal has been carried on for many years with satisfactory results, but the C. B. & Q. R. R., the Union Electric Light & Power Co. and the Standard Oil Co. of Indiana had not been producers of coal, at least in this district. Apparently the most impressive feature of the newest mines is their capacity, for in some cases it is planned that production shall be in the neighborhood of 1000 tons per hr. There are mines now in this field having nearly that output, but these were not planned for such output; in fact, when the shafts were sunk and the machinery installed it was generally supposed that the daily outputs would be only in the neighborhood of 4000 to 5000 tons. Incidentally, there has risen

Jan 1, 1920

By B. Egeberg

IN A previous paper1 the authors dealt with the physical properties of cold-rolled and heat-treated Britannia metal of the approximate composition Sn, 91 per cent., Sb, 7 per cent. and Cu, 2 per cent. It was found that cold rolling to more than 50 per cent. reduction produced a continuous softening of the metal when measured by the Brinell machine, whereas the Shore instrument indicated increased hardness; that the strength of the material followed the Brinell hardness and that the metal therefore became weaker the more it was cold-rolled; and further, it was found that hardness and strength could be largely restored to the cold-rolled metal by "baking" for about one hour at any temperature, between 200° and 450° F., the melting point. The higher the baking temperature, the more effective was the heat treatment as far as increased hardness was concerned. It seemed to be of little importance whether the metal was soaked or not at the baking temperature for a great length of time and the rate of cooling seemed to be immaterial. The present paper deals with the same alloy and more particularly with the resistance towards various forms of stress - tensile, bend and deep drawing. The metallography of Britannia metal has in part also been covered and an explanation attempted for the abnormal behavior of this alloy.

Jan 1, 1929

By John B. Campbell, George G. Binder, Loren H. Jenks

Conventional methods of producing crude oil generally leave in the depleted reservoir 1 to 3 bbl of oil for every barrel brought to the surface. This paper describes a field test which explores the possibility of using gas-driven liquid propane to improve oil recovery efficiency. Laboratory theoretical and experimental studies leading to the field test indicated that low residual oil values could be obtained using limited amounts of propane driven by gas. The field test involved (I) repressuring the reservoir to the vapor pressure of liquid propane, (2) injecting the liquid propane, and (3) injecting dry natural gas to drive the propane and oil to the producing wells. Considerable preliminary work, including new drilling, work-overs, and a pressure survey was necessary. During the test, oil production rates have averaged about nine times the final primary rates. The field average GOR is still below the pre-test value. Oil recovery to date is 112,000 STB. About one-third of the 143,-000 bbl of propane injected has been produced to date. Results of the C-2 block test are encouraging and larger, more definitive field projects ore indicated. INTRODUCTION The cost of finding new crude oil reservoirs continues to increase. Today the oil industry is drilling more and deeper wells and making fewer big discoveries. Nevertheless, the demand for oil and oil products increases each year. It is natural that oil companies, faced with this situation, concentrate on recovering more of the oil existing in our known crude oil reservoirs. There is room for much improvement. For every barrel of oil brought to the surface by present day methods of producing crude oil, generally 1 to 3 bbl are left behind in the depleted reservoirs. The "unrecoverable" oil in petroleum reservoirs is largely "trapped" by capillarity. Since flushing with a solvent is exceedingly effective in combatting oil loss due to capillarity, much effort in this direction has been made in recent years1,2,3,4,5. Liquefied petroleum gas (LPG) has been suggested as a relatively inexpensive. abundant, easily recoverable solvent for crude. Some operators have observed striking responses in oil recovery at production wells offsetting input wells in LPG storage reservoirs6,7,8. In spite of the low unit cost of LPG, however, it is uneconomic in general to inject a volume of LPG into a reservoir to remove a like volume of oil. A modification of solvent flushing in which a limited quantity of liquid propane is gas driven through a reservoir appears capable of recovering much more than 1 bbl of oil per barrel of propane injected. Also, most of the propane injected should be recoverable for sale or re-use. Since such a process could well have widespread and profitable application, The Carter Oil Co. is making an extensive evaluation of the technique's possibilities. Laboratory development of the gas-driven propane recovery technique was sufficiently encouraging to justify testing in the field. Since 1952, the Production and Research Departments of The Carter Oil Co. have cooperated in a field test in the C-2 Block leases of the

Jan 1, 1958

SAMUEL L. HOYT,* Minneapolis, Minn. -I was privileged to see the apparatus used by Mr. Scott, which is quite unique. It is, as the author says, a modification of the apparatus designed by Dr. Rosenhain, but it is evidently quite an; improvement. The author is to be congratulated on the fact that lie has given us, with his apparatus, a quantitative thermal analysis of these critical points. I notice the author refers to Ac3, Ac2, and Ac1. His designation of the Ac2 point is all right, but naturally in a steel with as much carbon as this, which would mean a hypereutectic steel, the other point might better be considered as an Ac, point with possibly another point above the Ac, point. Another important point brought out is the fact that, in Fig. 3, the lines representing the Ar1, transformation and the Ac,1_3 transformation do not intersect at the zero rate curve,

Jan 5, 1919

By G. E. Warner

This paper presents a review and analysis of a highly stratified Burbank sand waterflooding project in Osage County, Okla. Permeability values in this reservoir range from less than 0.1 md to nearly 3 darcys, and 90 percent of the permeability capacity is contained in approximately 26 percent of the reservoir volume. To date, the project has recovered 103 bbl/flood-acre-ft since waterflood initiation, which represents 50 percent of the total preflood cumulative and 12 percent of the original stock-tank oil in place (OSTOIP). The outstanding feature of this project has been the economical recovery of a sizable quantity of oil at water-oil ratios higher than those normally considered prohibitive. Nearly 70 percent of the recovery to date has been achieved at water-cut values greater than 95 percent. An ultimate waterflood recovery of 138 bbl/flood-acre-ft, or 67 percent of the preflood recovery and 16 percent of the OSTOIP, is predicted. The ultimate primary and secondary recovery will be 302 bbl/acre-ft and approximutely 40 percent of the OSTOIP. The actual performance is compared with the predicted performance by both the Stiles and the Dykstra-Parsons methods, and applicabiliry of these methods to this particular reservoir is evaluated. The performance of this project indicafes highly stratified reservoirs can be successfully and economically waterflooded when either reservoir or economic conditions preclude the use of mechanical and chemical means of mdbility control. Introduction The concept of the stratified reservoir has been used by reservoir engineers for many years to describe the vertical variations in permeability within a reservoir. Numerous techniques have been advanced to predict the expected waterflood performance of the stratified reservoir, and in most cases they prove reasonably accurate. The advisability of undertaking a prospective waterflooding project is determined, based on these predictions. This paper presents a review of a waterflood project in a highly stratified reservoir that normally would be considered a marginal prospect, at best, when evaluated with any prediction technique properly adjusted for reservoir conditions. This kind of project can be successful if the predictions are used to design a waterflooding system capable of economically recovering an oil cut of 1 to 2 percent. Experience gained in the early stages of the project's development was profitably applied in the later development of the entire reservoir and could prove beneficial to other projects with similar reservoir stratification or fracture characteristics. The East Burbank pool waterflooding operations have been developed in three separate stages, each of which has been a separate project: Mid-Burbank Unit, Flood A, and Stanley Waterflood. When referred to collectively, these projects will be called the Composite Project. The Skelly Oil Co.'s Barber lease — the only portion of the pool not included in the three projects — is operated as a separate facility with a cooperative injection-well agreement. It contains less than 80 productive acres and would not significantly affect the over-all performance picture.

Jan 1, 1969

By J. A. Morgan, R. E. Lund, D. E. Warnes

The design, operation, and performance of an integrated pilot plant for recovering zinc and copper from a complex sulfide ore are described. Metallurqical processing comprised selective sulfate roasting in a fluid bed, leaching in a weak sulfuric acid solution, recovering copper by cementation with scrap iron, air oxidation to remove soluble iron, more-or-less conventional Purification for removal of impurities deleterious to zinc electrolysis, recovering zinc by electrolyzing solutions, and recycling spent electrolyte to the sulfating roaster. The effects of several feed preparation and roasting variables on the sulfation response of zinc, copper, and iron are set forth, together with the electrolyzing characteristics of the zinc solutions. SEVERAL years ago, St. Joseph Lead Co. undertook a research program to develop procedures for winning values from a complex ore. The deposit was a massive sulfide ore body containing upwards of 65 pct pyrite. Gangue, composed principally of quartz, calcite, dolomite, and alumina, comprised only 10 to 20 pct of the mineralized zones. Zinc, lead, and copper values were variable, but averaged about 6.7 pct Zn, 2.6 pct Pb, and 0.4 pct Cu. Early in the exploratory test work, it was established that the intimate physical association of the minerals in the ore made beneficiation by conventional flotation procedures difficult. Accordingly, in addition to a comprehensive program to investigate physical methods of ore beneficiation, a research program was established to develop a direct metallurgical method for processing the ore. After surveying a number of potential methods for recovery of both metal and sulfur values, effort was concentrated on the development of a sulfate roasting process. It is the pilot-plant development of the sulfate roasting-solution purification—electrolyzing procedures which forms the subject of the present paper. PREPILOT PLANT INVESTIGATIONS From a thermodynamic viewpoint, it is feasible to sulfate selectively a host of metals, including Zn, Cu, Pb, Cd, Ni, and Co, in an iron sulfide ore without sulfating iron. Of course, a number of conditions such as roasting temperature and gas composition must be controlled. In addition, as first disclosed by Hybinette,' an alkali metal salt is frequently beneficial in promoting the sulfation reactions. The application of fluid bed roasting to the sulfate roasting of nonferrous metals was pioneered by Stephens of Battelle Memorial Institute.' Subsequently, Falconbridge Nickel Mines, Ltd. developed an unique process for selectively sulfating and recovering nickel, copper, and cobalt from their nickeliferous pyrrhotite, which recently has been described in an excellent paper by Thornhill.3 One of the unique features of the Falconbridge process is that the feed is introduced into the fluidized roasting furnace in a manner such that the feed slurry is broken up into droplets and the droplets lose their moisture before contacting the surface of the fluidized bed. The dried droplets do not tend to cake or fuse, but rather remain as discrete agglomerates of concentrate particles and roast without significant degradation. St. Joseph, aware that Battelle Memorial Institute had participated in the development of the Falconbridge process, obtained permission from Falcon-

Jan 1, 1962

By Harold Vance

The Lafitte field, the largest oil reserve in South Louisiana, is in Jefferson Parish, some 25 miles due south of the City of New Orleans. The discovery well, The Texas Company's No. I, Louisiana Land and Exploration Co., in the southeast corner of sec. 19, T. 17 S., R. 24 E., was completed on May 15, 1935. The initial production of this well was 960 bbl. of 34.9º A.P.I, gravity oil per day, through 1/4-in. choke, with a tubing pressure of 1600 lb. per sq. in. Producing depth was 9558 to 9572 ft., the total depth being 9572 feet. The development of the field has been at the rate of three to eight new producing wells per year. Up to Jan. I, 1944, 60 producing wells and three dry holes had been drilled, while two of the original producing wells had been abandoned, Although the field is not unitized, The Texas Company was the sole operator in the field until the Lafitte Oil Co. completed its NO. I Jefferson Parish, near the center of the west line of the NW/4 sec. 20, T. I7 S., R. 24 E., on July 27, 1941. No additional wells have been drilled by the Lafitte Oil Company. Drilling and WeLl-completion Practices All drilling operations have been conducted with the steam-driven rotary equipment mounted on submersible barges. A canal is dug to the well location and the drilling unit is floated into position. The seacocks are then opened and the barge settles to the bottom. After the well is completed, seacocks are closed and water is pumped out of the barge. This procedure permits the barge, with drilling equipment still in place, to be floated and moved off the location. The conventional open-hole and the gun-perforation method of completing wells has been used in this field. Where possible, the initial completion is from the deepest 'sand and recompletions are made progressively up the hole, by plugging back with cement to the upper sand and gun-perforating opposite the new sand. The oil string usually is 7 5/8 in., 26 to 34-lb. casing. Three wells have been completed using 9 5/8-in. 44-lb casing set below 10.000 feet. Fig. I shows the completion and recom-pletion practices on one well in the field. This figure also shows the relative locations of the different producing sands as shown on the electric well log. Structure The structure is an elongated dome, cut by two north-south faults, which divides the field into three segments. The west segment is the highest structurally and contains most of the recoverable oil and gas. The center segment is less important and the east segment is of little value. Figs. 2, 4, 6, 8 and 10 are depth contour maps on top of the five main producing sands in the field. These maps show the relative size of the five reservoirs and the

Jan 1, 1945

By C. O. Lee, Z. W. Bartlett, R. H. Feierabend

DESPITE the fact that American brimstone production has increased 2 ½ times since the prewar period 1935-1939, the demand for sulphur exceeds the supply. To alleviate this situation efforts are being made to increase the production of sulphur and bring the shortage to an end. The most important developments are being undertaken on the Gulf Coast. One new brimstone mine has been put into operation and three others are being developed. These projects are expected to account for a large part of the additional sulphur production now anticipated for the free world.1,2 Except for a small amount of byproduct sulphur from sour natural gas and refinery gases, American brimstone is produced by the same process devised by Frasch 60 years ago, consisting essentially of the injection of superheated water at about 325°F into sulphur-bearing formations through wells which are drilled from the surface by standard oilfield methods. Sulphur deposits vary in depth from a few hundred feet to 2500 ft from the surface; in a typical deposit the sulphur occurs as crystalline aggregates in porous limestone and gypsum in concentrations varying from 5 to 40 pct, the thickness of the formation ranging from a few feet to several hundred.3 The deposits are usually surrounded by formations impermeable to the flow of hot water, the top and sides being enclosed by impermeable sediments and the bottom by dense impermeable anhydrite. As the hot water leaves the well bore, it percolates through the formation and transmits its heat to the rock and sulphur. When the sulphur reaches 240°F it melts, separates from the rock, and flows through the hot channels downward to the bottom of the well. Here it is freed from incoming water and from formation water, which saturates the porous rock, by gravity separation, liquid sulphur being 1.8 times as heavy as water. It then is pumped to the surface in a relatively pure state, 99.5 pct or better. Chief contaminants are organic impurities, such as petroleum residues, which are soluble in sulphur and tend to discolor it.4 Sixteen deposits of commercial magnitude have been discovered in the Gulf Coast area. Five of these deposits have been depleted and abandoned with production totalling more than 26,886,000 tons. Eight properties are operating and have produced 65,875,000 tons to Jan. 1, 1952, for a total Frasch-type production of 92,761,000 tons. Three new deposits are under development. Aforementioned deposits are listed in Table I. Each of these deposits has been found in the caprock overlying shallow intrusive salt plugs or domes and buried in several hundred to as much as 2500 ft of sediments. All salt domes do not contain sulphur; only 16 sulphur deposits have been found in a total of 231 salt domes which are known to exist in the Gulf Coastal states of Texas, Louisiana, Mississippi, and Alabama. Ten of the salt domes lie in the Gulf of Mexico offshore from the Texas and Louisiana coasts. Additional offshore domes may be found, but their value as a sulphur potential is questionable owing to the great hazard and high cost that can be visualized at these exposed locations where depths of water reach 50 ft. Marsh operations of Freeport Sulphur Co. are in the southeast section of Louisiana. In general, the country is low, flat, relatively uninhabited marshland broken by many shallow lakes and bayous which extend for irregular distances inland. Geologically, the coastal area is of sedimentary origin, having been built up gradually under the sea by deposition of material carried down by large streams to their deltas. Offshore bars were formed enclosing

Jan 1, 1952

By R. G. Hall

The method of crushing rock by throwing with violence against a resisting surface does not differ in principle from the method of dropping it from a height upon a specially arranged surface. The former is applied in various mills now in use, especially those for reduction of the softer rocks and minerals. The latter is the principle of the Hadsel mill. Probably not new, but Hadsel's method of approach is at least novel, and applicable to hard as well as softer rocks. An elevating wheel raises the rock and drops it from a height upon a hard metal plate or series of plates. The first commercial mill consisted of two units, one on each overhung end of a horizontal shaft, driven by a gear train from a 100-hp. motor. A description of this machine, then in operation at a mine at Georgetown, Calif., was printed in the Engineering and Mining Journal (April, 1932) and need not be repeated here. That machine took mine run ores of a size up to 12 in. and successfully produced a pulp of approximately minus 60 mesh, suitable for flotation. The mills were each 24 ft. in diameter and 3 ft. in width of face. The elevating buckets were formed of steel plates, extending inward from the periphery, inclined to the tangent at 75". This plate was 3 ft., the width of the wheel, and 2 ft. deep. One side of the wheel towards the carrying shaft was closed, the other was open except for the rim plate enclosing the ends of the buckets and extending inward about 5 ft. on the radius. About a 90" arc was immersed in a concrete tank, which acted as a classifier. The breaking plates, seven in number, were carried cantilever through the open side of the wheel. The lowest plate was placed just above the water level. The illustrations (Figs. 1 and 2) make plain the general arrangements. The plant set-up was somewhat crude but worked well enough to demonstrate the weak points of the mill. In this operation it was early apparent that, considering all the factors from delivery of the mine run ore to the final pulp for flotation or cyanida-tion, the Hadsel mill offered possibilities of great simplicity of plant and savings in power. Continuous operation developed mechanical weaknesses, some of which were readily discernible in details of design, some became more apparent as operations progressed.

Jan 1, 1935

By L. G. Austin, R. R. Klimpel

This article demonstrates that the Gilvarry and Klimpel-Austin equations for the random fracture of solids are incorrect by deriving intuitively correct expressions for simple cases and showing that the foregoing equations do not reduce to right answers. It is the purpose of this article to point out serious errors in some existing treatments of the random fracture of solids. Gilvarry ' gives the equation where B(y,x) is the cumulative fraction by weight of particles which fall below size y when particles of size x are broken. The constants i, j, k contain shape factors and densities of flaws leading to propagating fracture surfaces. The authors believe this equation to be incorrect. Klimpel and usttin' give where r~ is the number of fracture planes originating from edge flaws (per L where L is the edge length of the original particle), rs is the number of fracture planes originating from flaws on the particle surface (per S where S is the surface of the original particle) and rv is the number of fracture planes originating from flaws within the particle volume. This equation is also incorrect. The failures of these equations will be demonstrated by deriving intuitively correct expressions for simple cases and showing that the Eqs. 1 and 2 do not reduce to the correct answer. In this process, the flaws in statistical reasoning inherent in the derivations of Eqs. 1 and 2 become apparent. THE KLIMPEL-AUSTIN EQUATION Eq. 2 was derived by considering a set number of fracture planes occurring spatially at random within a given particle. This case will be treated here in its simplest form: a line with 1, 2, 3, etc., fractures occurring at random. This is illustrated in Fig. 1, with the original length of the line as Vo. The breakage of a long, thin, brittle fiber is a corresponding physical case, with the volume of a fragment proportional to its length. Assuming independence of the occurrence of the fractures, the probability of any one fracture occurring in length dV is dV/V0. The probability of the first fracture occurring at a length V to V + dV is dV/Vo. For the case of only one fracture per particle, formation of a number-size distribution of the left-

Jan 1, 1969

By R. D. Rubright, Owen J. Hart

In the Bingham district over a span of more than 90 years, 43,947,104 tons of "non-porphyry" copper, lead, zinc, gold, and silver ore have been mined from a folded and faulted alternating series of Pennsylvanian and Permian limestones and sandstones and from some Tertiary intrusives. This ore came from six major mine areas located east, south, and west of the Bingham stock, site of Kennecott's porphyry copper deposit. The sediments were affected by varying degrees and types of metamorphism within 2000 feet or more of the Bingham stock. The intrusives consisting of the mineralized Bingham and the unmineralized Last Chance stocks and numerous dikes and sills of granite and monzonite were intruded into the asymmetrical northwest-trending and plunging Bingham syncline. Extrusives, although present, contain no ore. The bulk of the non-porphyry ore in the district was found above the Midas thrust. The primary ores are found in three major types of deposits: concordant bedding fissures, crosscutting fissures, and replacement deposits. These deposits consist of both copper and lead-zinc ore bodies that contain various amounts of gold and silver. These are in places roughly zoned both laterally and vertically and have been mined for up to 10,000 feet outward from the center of the Bingham stock and up to 5000 feet vertically below the surface. Secondary ore was of minor importance and was found in and under the oxidized zone at shallow depths. Genetically the Bingham stock, the ores, and related phenomena have had a common source.

Jan 1, 1968

By W. F. Boericke

AT Kellogg, Idaho, J. Fred Johnson is driving the 5000.-ft Hooper Tunnel under contract for the Bunker Hill & Sullivan Mining Co. This was visited by a group of engineers during the recent meeting of the A. I. M. E. at Spokane, who were impressed with several features of the work. An excellent record of advance is being made by the contractor, over 47 ft. being driven in 3 days during October, and 223 ft. advanced in 29 shifts. The size of the tunnel is 8 by 9 ft. The ground is Wallace quartzite and, save for the first hundred feet, stands well and does not require timbering. The grade is held uniformly at 0.5 per cent. Two shifts are employed, with two sets of machine men to the shift. Ingersoll-Rand S-70 drills are used. each machine being set up independently in the heading; 26-30 holes are drilled to the round; using 1 1/4: in. round hollow drill steel. A 7-hole cut is required, and these holes are blasted simultaneously, using 60 per cent gelatine, 12 sticks to the hole. Instantaneous and fire delays of electric detonators are used for the complete round. A feature of the drilling is the use of long carriage drills, necessitating fewer changes of steel, and more time spent on drilling, with less time consumed taking steel back and forth from the face. Of course, this means better blacksmithing so that the drill steel shall stand up for the added footage.

Jan 1, 1930

By H. B. Pulsifier

THE following report on the structure of copper is the result of work done in the laboratory of the Rome Wire Co. early in 1925. Previous work had indicated to the author that excellent results might be expected if a suitable technique in surfacing and etching could be developed. The resources of the Rome Wire Co. enabled this metal to be studied more intensively than previously, although none of the materials were investigated exhaustively. The material presented is hardly more than a preliminary summary of the most obvious character of the metal; a much longer research would have been necessary to solve many of the pressing problems connected with the casting and working of copper. The Rome Wire Co. has generously released the paper; to Dr. E. H. Darby, technical supervisor of the company, are due particular thanks for his appreciation of the importance of metallographic study of copper and for his effective support in the prosecution of the work and in making its publication possible.

Jan 1, 1926

By P. K. Koh

Isothermal salt bath annealing of 0.014-in. thick 3 pct Si-Fe sheet was conducted at temperatures ranging from 927" to 1260°C in order to investigate the grain-growth behavior. Within the temperature range studied normal grain growth did not take place; instead nucleated grain growth prevailed. The nucleation frequency was found to increase with the annealing temperature, with corresponding decrease in grain diameter of ultimate secondaries after long-time annealing. Large secondaries grown by low-temperature isothermal annealing are essentially of (110)[001] orientation, and the (110)[001] texture is strong. The relatively small secondaries grown by high-temperature isothermal annealing also have orientations near (110)[001] as an average but widely scattered; thus the amount of (110)[001] texture is found to decxease with isothermal annealing temperature. Grain growth and texture development at various annealing temperatures can be described by families of sigmoidal curves with the time axis. ThE cube-on-edge (110)[001] texture and predominating large secondaries in fully annealed 3 pct Si-Fe are customarily produced by a 1100°C annealing of a primary recrystallized sheet. On annealing the primary recrystallized sheet above the temperature at which secondary recrystalliza-tion starts May and turnbull' have demonstrated that secondary grains decrease in size as the temperature is increased. The phenomenon was attributed to a decrease in the (G/N) ratio where G is the growth rate of secondary grains and N is their nucleation frequency. In addition, the growth of grains in orientations rather far from (110) [0011 occurred. The present investigation is devoted to detailed rate studies of the same phenomenon. Mi-crostructure, magnetic anisotropy, and texture data were obtained on strips of 3 pct Si-Fe isothermally annealed for various lengths of time in a BaC12 salt bath at temperatures ranging from 927' to 1260°C. The texture results are reported elsewhere.2 EXPERIMENTAL PROCEDURE Commercial 0.014-in. cold-rolled silicon-iron strip (3.16 pct Si), prepared by two stages of cold rolling with an intermediate short anneal, was given a decarburizing 3-min anneal at 800°C. Metallo- graphic studies indicated complete recrystallization. In order to provide a protective atmosphere and to obtain a rapid rise to an elevated temperature, a BaC12 and NaCl fused salt bath was used to heat samples isothermally at temperatures from 927" to 1260°C for various lengths of time. Torque magnetometer measurements were made on 1-in.-diam disc specimens before and after annealing, since magnetic torque curves supply information on textures and texture changes. Surface layers of the strip samples were polished and etched to reveal the grain structure. The diameters of grains were measured from micrographs at X100. The average value of diameters of matrix grains was established. Individual grains in the micrograph were cut out separately along grain boundaries. These grains were classified either as matrix grains or growth grains. The percent weight of growth grains in the cut-out patterns was reported as percent growth. Manv of the individual grains in the annealed samples were large enough to be X-rayed with a 5-mil beam for Lauegrams. Complete (110) pole figures were obtained for the primary recrystallized structure and short-time annealed structure with CoKa radiation at 30 kv in the back-reflection range, such as (220) for (110) poles, as described recently by Newkirk and ~ruce.~ Both the X-ray optics and the technique for pole figure differed slightly from those adapted previously.4

Jan 1, 1960

By A. O. Dady

PRODUCERS of both bituminous and anthracite coal have for many years been worrying about the gradually decreasing consumption of their product in the United States. Twenty years ago production had climbed to a peak of 569,000,000 tons of bituminous and around 90,000,000 tons of anthracite. In 1939 it was about 393,000,000 tons of bituminous and 51,000,000 tons of anthracite. This does not mean that the country is using less polver than it did twenty years ago, or that we are living in colder buildings. As a matter of fact, about three times as much electric power is now being consumed as in 1920. In general two factors are responsible for the drop in coal consumption: (1) greater efficiency in the use of coal, and (2) increased competition from hydro power, internal-combustion engines, and substitution of oil and gas for coal in conventional furnaces. As to the first factor, it took 3 Ib. of coal to generate a kilowatt-hour of electricity in central stations in 1920 and it takes 1.4 Ib. now; it took about 175 Ib. of coal in railroad locomotives to haul 1000 long tons a mile in 1920 where it talces about 120 lb. now: and marked economies have also been made in the use of coal in steel plants: in by-product coke ovens, and in other large uses. As to the second factor, three times as much electricity is produced from water power now as was produced in 1920; three times as much oil is being consumed in various ways; and likewise, the consumption of natural gas has tripled- whereas coal consumption in the same period has dropped by 40 per cent

Jan 1, 1941

Aamot, Olav Crone, (Member '29) Met , Det Norske Aktieselskab for Elektrokemisk, Industry, Inc , Oslo, Norway Aase, Glenn D , (Junior Member '40) Met Engr , Western Research Dyv , American Smelt & Refin Co , Selby, Calif Abadye, Henry G., (Member '43) Petr Engr , Sr Prod Engr , Long Beach Oil Development Co, 255 So Santa Clara Ave, Long Beach, Calif Abbott, Clarence E , (Member '04) Civil Engr , 'P , Tennessee Coal, Iron & R R Co , 1542 Brown Marx Bldg Birmingham, Ala Abbott, Fred E , Jr , (Member '44) Petr Engr , c/o Arabian American Oil Co , A P 0 816, c/o Postmaster, New York, N Y Abbott, John H , (Member '46) Resident Mgr , Hickman, Williams & Co, 230 No Michigan Ave, Chicago 1, Ill Abbott, R W R., (Member '47) Asst Power Supt, Cananea Consolidated Copper Co , Cananea, Son , Mexico Abbott, Robert H? Jr., (Junior Member '47) Petr Engr, Producing Dept, The Texas Cc, P 0 Box 951, Houma, La Abel, Raymond L., (Member '28) Chem Engr, Prof of Chem Engrg, Univ of Pittsburgh, Pitts¬burgh Pa Abel, Walter D , (Member '32) Min Engr, Div of Corporations, 403 State Bldg , Los Angeles 12, Cal if

Jan 1, 1952

By Albert Mendelsohn, Charles F. Jackson

The system of stoping described in this paper was first introduced at the Champion mine of the Copper Range Co., Painesdale, Mich., in 1929, and since that time has been developed to a high state of efficiency. The method was devised to overcome some of the objections to the system of horizontal cut-and-fill stoping formerly employed at this mine and has done so successfully. The objections to the old system became more pronounced as the depth of mining operations increased and a change became a matter of necessity. While the combination of conditions encountered at the Champion mine is rather unusual, it is believed that the sublevel system developed there could be employed advantageously at other mines, though certain modifications might be necessary. This paper describes the method and the results obtained through its use, and suggests modifications for making it applicable under quite different conditions from those at the Champion mine. Champion Lode The lode being mined at the Champion mine is the heavily brecciated top of what is known in the district as the Baltic lava flow. The brecciated portion of the lode varies from 6 to 50 ft. in width. Under this breccia is the thinner, tighter, amygdular portion of the lode; under the amygdular portion is the trap footwall. Native copper is found deposited in the rock fragments of the breccia, in the cementing material of the breccia, in the amygdules and in the trappy rock under the amygdular part of the lode. In some places masses of copper many tons in weight extend far into the footwall of the lode. The Baltic lode on Champion property is 8000 ft. long, very straight along the strike, and dips uniformly 70' to the west for 3000 ft. in depth, at which point the dip begins to flatten slightly. The width of the copper-bearing portion of the lode varies from 10 ft. to as much as 80 ft. in a few places; the average width now mined is 17 ft. The lode rock is hard—the standard one-man drilling machines used in the mine drill from 10 to 14 in. per minute when mass copper is not

Jan 1, 1932

By Margaret G. Scroger, Luke L. Y. Chang, Bert Phillips

Using the sealed-system technique, isothermal sections at 1000°, 1400°, and 1700°C for the system Co-W-O hare been determined. The equilibrium data were obtained by microscopic and X-ray diffraction examination of samples winch had been equilibrated in pellet form fit sealed capsules and subsequcelly quenched. The only ternary phase in the system is CoWO, which mells congruentlly at approximately 1280°C. At 1000°C, this ternary phase exists in equilihrinm with all the allows and oxide compounds in the hounding binary systems. At 1400°C, liquid forms in thin oxygen-rick region near the CoWO4 composition and only one phase assemblage in the system (W + WO2 + H'5CV7) contains no In order for tungsten to be used to its full potential as a refractory metal. it is necessary that coatings be developed which will preclude oxidation of the metal. It is true that coatings now in use do protect tungsten for short periods at temperatures near 1700 °C. but no coating has been developed which protects adequately at or above this temperature where the use of tungsten is most required. This situation exists despite the great effort in recent years to develop such coatings. and it has become evident that success in this area can be achieved only after a more fundamental knowledge of the thermochemical and therrnophysical propertics of tungsten oxidation and tungsten -metal oxidation is obtained. In our 1aboratories. a project liclrriel. At 1700°C, tungsten is the ouly solid phase which is stctble, and a large range of compositions melt to two liquids, one metallic and one oxidie. The sealed-system method of etjuilibra lion prevents the determination of the partial pressure of oxygen and the total gas pressure orer the condensed phases at equilibrium. This data is not available and can be obtained only by some other method. We can generalize on these unknowns and indicate that both partial pressure of oxygen and total pressure of the gas phase must increase as compositions change from the W-Co binary toward the oxygen apex and as temperatures increase. has been initiated which is directed at providing some of this fundamental information through a study of metal-tungsten-oxygen reactions and equilibria and the properties of any intermediate phases which result. This paper is based on a part of this program in which the system Co-W-O was studied between 1000° and 1700°C. I) RELEVANT INFORMATION FROM THE LITERATURE A) The Notable Lack of High-Temperature Work on Metal-Tungsten-Oxygen Systems. It should be pointed out that there has been no published report (of which we are aware) which deals with the study of a metal-metal-oxygen system at 1700°C. Also, there has been no report on a complete metal-luugsten-oxgen system at any temperature. As a matter of fact. before this program started. The

Jan 1, 1965

By Janshi Sen

Vertical-intensity magnetometers, for instance 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. I, lean ore, rich ore and a single pole produce the same magnetic field. The depth of the pole, 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, 1946