Search Documents

By Warden A. Moller

The Fushun coal-field, now being opened up by the South Manchurian Railway Co., is connected with the main line by a branch, 30 miles long, from Sui Chia Tun, 10 miles south of hlukden, the oapital of Manchuria. The coal-field, which lies on the southern bank of the Hun river, 25 miles east of Xiukden, consists of a long narrow valley, running east and west between hills composed of granitic and volcanic rocks. At the foot of the northern hills, the Hun river flows westward towards the Liao river, which drains the greater part of South Manchuria. At the foot of the southern hills, resting immediately upon volcanic rocks, and overlain by 30 ft. of alluvium, is a thick layer of Tertiary shales, carrying near the base two coal-seams. In many places the lower of these two seams is in immediate contact with volcanic lavas and basalts, the flows of which were evidently both prior and subsequent to the formation of the coal-seams. This contact has practically destroyed the lower seam for commercial purposes. The value of the field lies in the upper seam, which, for a length of nearly 8 miles, varies from 100 to 130 ft. thick. Of this seam, 85 per cent. is good coal. The seam dips to the north conformably with the shales, at angles of from 25 to 30°, and appears to have an uninterrupted dip for about 1.5 miles before being possibly cut oft' by the northern hills. The coal-bearing shales extend east and west for a length of about SO miles, but towards the east the continuity of the seams is broken by intrusive rocks. The coal is of the sub-bituminous variety, low in ash, bright in appearance, and breaks with a conchoidal fracture. It has the usual qualities of Tertiary coal—quick steam-raising and poor stacking qualities. Pieces of resin, varging in size from

Jan 1, 1911

By D. A. Koss, R. O. Gordon

Changes in the ultrasonic attenuation in steel specimens have been observed during tensile tests. Samples of AISI 1020, 1045, and 1095 steel quenched and tempered to a spheroidized condition have been used. Both attenuation and microstrain measurements fail to reveal preyield dislocation motion in the as-tempered specimens. Upon reapplication of the load to samples in the strain-aged condition there is a large preyield attenuation increase but no observable micvostrain up to 90 pct of the flow stress. These and other observed changes can be related to changes in the free disloca-tion density and the average dislocation segment length. The results show that the increase inflow stress in the strain-aged condition is caused by increased fric-tional stresses rather than by dislocation pinning. Strain aging is found to proceed through three distinct stages. The second of these follows the t2,3 law and is characterized by an activation energy of 20 kcal per mole. RECENT years have seen substantial progress in the understanding of the sharp yield-point phenomenon in metals; the influence of solute atom pinning of dislocations, of dislocation multiplication, and of the stress dependence of dislocation velocity on the stress-strain curve has been investigated. Evaluation of the relative importance of these factors in the various metals and alloys is not complete, however, because observation of the stress-strain curve alone is not sufficient to determine the yield mechanism. In the experimental study of the sharp yield point it would be desirable to be able to directly observe the dislocations in a bulk specimen. Ideally, the observations should be made during deformation because of the rapidity of recovery processes in lightly strained metals. While direct observation of dislocations under these conditions is not possible, certain physical properties sensitive to the dislocation structure can be observed while a test specimen is undergoing deformation. Of these properties, one of the most useful is the ultrasonic attenuation. The total observed attenuation, a, is due to a number of causes: acoustic diffraction, scattering by inhomogeneities in the structure, heat flow, magnetostriction, and dislocation damping. Satisfactory measurements of the absolute a are difficult to make because all of these factors cannot be evaluated in any given specimen, but changes in a due to changes in the dislocation contribution alone can be reliably observed by appropriate technique. Because dislocation damping is sensitive to small concentrations of long, free dislocation segments, observation of acoustic attenuation due to dislocations is particularly useful in the study of the early stages of plastic deformation. In the experiments to be described, acoustic-attenuation measurements have been made on steel samples as these were pulled through their yield in tension. While many details of dislocation-damping phenomena in the bee metals remain obscure, the main characteristics of this type of internal friction are established. At low strain amplitudes and at frequencies well below that of dislocation resonance (the conditions of these experiments) the dislocation contribution to a is given by1 where A is the density of freely oscillating dislocation segments, L is the average segment length, B is the damping constant for a moving dislocation, w is the frequency, O is an orientation factor, C is the dislocation line tension, and t is a numerical factor relating average segment length to the distribution of lengths. Because of its fourth-power dependence on L, a is highly sensitive to changes in dislocation free segment length such as might arise from unpinning, recovery, jog formation, or dislocation intersection. The above equation is not valid near the resonant frequency of dislocations, but the segment length in steel is believed to be much too short to cause over-damped resonance near 10 Mc; this will be discussed later. In the present experiments plastic-strain measurements with a sensitivity of 2 x 10-5 were also made during the early stages of loading in order to detect preyield strain. The general plan of the experiments is to observe

Jan 1, 1967

By M. I. Sherman, J. D. H. Strickland

USE of a hydrometallurgical approach to the oxidation of sulfide ores and extraction of metals therefrom may have advantages over the more common smelting techniques when a low grade deposit is difficult to concentrate or the subsequent separation of metals, coexisting in the ore, is laborious by any known smelting operation. For economic reasons, the most promising oxidants are either atmospheric oxygen or electric power. The use of oxygen, or air under pressure, has recently been revised. Pyrrhotite has been converted to iron oxide and elementary sulfur' and a variety of sulfides have been treated by Forward and co-workers.2-4 Generally sulfate is the end form of the sulfur but with galena in an acid medium, elementary sulfur can be formed." For economic reasons chlorine and ferric iron salts are about the only possible alternatives to the atmosphere as oxidizing agents for base metal sulfides. If aqueous solutions of chlorine or ferric iron are employed, the reduction products can be oxidized electrolytically in situ and used again, thus acting as catalysts for electric power as oxidant. The use of ferric salts for this purpose is established hydrometallurgical practicea but, although chlorine gas has been employed in the dry state at an elevated temperature, its use in aqueous solution at or near room temperature has not found favor. The reaction of chlorine water with the soluble sulfide ion has been studied by several workers,7-9 and both sulfate and elemental sulfur are found as end products, the latter being favored by the presence of a low concentration of oxidant relative to that of sulfide in solutions of about pH 9 to 10. Of direct bearing on the work in hand are an early American patent" and a recent Austrian patent." The former advocates stirring powdered ore with an aqueous solution of ferric chloride chlorine oxides and chlorine. In the latter it is claimed that both metal and sulfur can be obtained by electrolysis, in a diaphragm cell, of a metal ore slurry in brine. Details in these patents are scant and no data or explanation is given for the mechanism of the reaction which, in the Austrian work, is attributed to the (unlikely) action of nascent chlorine at the anode surface. No mention is made of possible differences in behaviour between various ores. Apparatus A complication encountered when working with chlorine water is that a serious loss of chlorine occurs by gas partitioning unless an enclosed system is used and any air space in the apparatus is kept very small and constant. Arrangements were made, therefore, to take out samples for analysis without letting air into the system to replace the liquid removed. For convenience in studying a heterogeneous reaction the apparatus was so designed that a reproducible controlled stirring rate could be maintained and the ratio of surface area of ore to volume of solution was approximately constant throughout any experiment. The apparatus used is shown in Fig. 1. The ground ore was placed in the horizontal cylindrical vessel, A, of about 1 liter capacity, heated by a constant temperature circulating bath pumping water through the concentric jacket, B. By adding chro-mate to this water, an ultraviolet radiation filter effectively surrounded the reaction vessel, greatly reducing any possible photochemical decomposition of chlorine solutions. Stirring was effected by glass paddles, C, attached by an axle to a magnet which was rotated by another powerful Alnico magnet, D, outside the glass end, this magnet being itself rotated by an electric motor electronically controlled to constant speed. Speed could be varied from about 150 to 900 rpm and was measured and held to within 1 pct of a given value. The end of the reaction vessel remote from the stirring magnet was closed by another one-ended glass cylinder, E, connected by thin polyethylene bellows, F, clamped by screw clamps and watertight rubber gaskets to the main vessel. Through E, a glass electrode and calomel electrode projected into the solution and a hypodermic syringe pierced a small bung and allowed acid or alkaline to be added to maintain a constant pH. By pushing the fully extended bellows until the two cylinders touched, from 50 to 100 ml of solution could be forced out through a sintered disk into the three-way tap system, G, either to waste (for flushing purposes) or up into a 10 ml burette where the solution could subsequently be measured out for analysis. The ore samples were introduced at H, the tube being stoppered by a thermometer of —1 to +52ºC range, graduated to 0.1°C intervals. To prevent ore from being ground in the end bearings of the stirrer these bearings were pro-

Jan 1, 1958

By A. M. Ross

DEURING the past year the uranium milling industry has seen the installation and initial operation of solvent extraction circuits in the Climax Uranium Co. mill at Grand Junction, Colo.; the Kerr-McGee Oil Industries mill at Shiprock, N. M.; and the AEC pilot plant at Grand Junction, operated by National Lead Co. Inc. The Vitro Uranium Co. mill in Salt Lake City is converting to solvent extraction, and the new Union Carbide Nuclear Co. mill to be constructed at Rifle, Colo., will also employ this method. Solvent extraction circuits will be incorporated in two mills now under construction -the Texas Zinc Minerals Co. mill at Mexican Hat, Utah, and the Gunnison Mining Co. mill at Gunnison, Colo. In Canada the expansion of the Eldorado mill at Port Radium includes installation of a solvent extraction circuit. Solvent Extraction Process: In simplest terms, the organic solvent is mixed with leach liquor and then the organic and aqueous phases are allowed to separate. By this contact the solvent selectively extracts uranium from the liquor, leaving a raffinate, or barren leach liquor, which may be discarded. The pregnant solvent is stripped of its uranium by con- tact with a stripping solution, such as aqueous sodium carbonate, which has a greater affinity for uranium than the solvent does. The barren solvent from the stripping step is available for further leach liquor extraction, and the pregnant stripping solution, which usually contains 30 to 60 g U3O8 per liter, is ready for precipitation of a final product, or uranium concentrate.

Jan 9, 1957

[Secretary's Note.—The following official tables, prepared by the government of the State of Nuevo Leon, and presented to the Institute at the fifth seseion of its Mexican meeting, held at Monterrey, November 26, 1901, present in compact form a general summary of the mineral and metallurgical industry of the State for the preceding calendar year, together with a statement of the railway-shipments of ores from January 1 to October 31, 1901, which, in the absence of direct official mining and metallurgical statistics (not yet available for that year), clearly indicates an increased activity and productiveness of these industries. In other words, the mines are reported to have yielded in 1900 a total product of 83,960,169 kilos of ore; whereas, the railway is reported to have carried, in 10 months of 1901, a total of 108,318,647 kilos of ore. At this rate, the railway-shipments for 1901 mould amount to nearly 130,000,000 kilos. It is, of course, possible that the railwaystatistics include other freight than ores, or that they include duplications, such as would be involved in the record, first of certain ores transported, and afterwards of the " base bullion " produced therefrom. An accurate comparison is therefore impossible ; yet, after making all reasonable allowances for such items as have been named, it seems safe to infer that the mining and metallurgical industry of this State accomplished in 1901 a substantial progress.]

Jan 1, 1902

By N. Jaffee, J. E. Dorn

An activation energy of 27,400 5 1000 cal per mole was obtained for the creep of poly crystaLline aluminum over the temperature range of 273° to 350°K, at strains varying from 0.003 to 0.230. Stresses varied fronz 2250 to 6000 psi, strain rates from 0.1145 x 10-5 to 29.5 x 10~s min.-1. A strain temperature-compensated time correlatzon was oblained indicating that the structure was depedent solely upon the strain at a given stress; X-ray diffraction analyses and tensile tests on precrept specimens verified this. PREVIOUS investigations1 have revealed that the apparent activation energies for creep of high-purity polycrystalline aluminum are insensitive to the applied stress, strain, and strain rate. On the other hand, the apparent activation energy is dependent on the temperature1,2 as shown by the data reproduced in Fig. 1. Each plateau of this curve is associated with a unique rate-controlling dislocation mechanism for thermally activated creep. It is now very well-established that over the high-temperature range, where the activation energy is about 35,500 cal per mole, the creep rate is controlled by the dislocation climb mechanism. 3 Over the intermediate range of temperatures, from about 260° to 370°K, another unique activation energy, namely 28,000 cal per mole, is obtained. Current evidence strongly supports the contention that cross slip is the rate controlling mechanism for creep in this case: 1) Crystal recovery can occur by a number of mechanisms prominent among which are climb and cross slip. Astrom,4 using the Borelius micro-calorimeter, reported rapid crystal recovery of cold-worked polycrystalline A1 at 340° to 370°K giving an activation energy of 28,000 cal per mole and additional recovery at 455° to 478°K giving an activation energy of 3 6,000 cal per mole. Whereas the latter recovery process was clearly associated with the dislocation climb mechanism, the former, 28,000 cal per mole process could have arisen as a result of the cross-slip mechanism. 2) Single crystals of high-purity A1 favorably oriented for octahedral glide exhibit an activation energy for creep of about 3400 cal per mole over the range from 0° to 400°K and that of about 28,O:O cal per mole over the range from 600° to 775°K . Whereas the lower activation energy process is accompanied by the development of sharp slip-band traces, the 28,000 cal per mole process was uniquely characterized by extensive cross slip. 3) Theoretical calculations by Seeger and schoeck6 suggest an apparent activation energy for cross slip in A1 of about 24,000 cal per mole and another theoretical estimate by Friedel 7 gives the observed 28,000 cal per mole as the theoretical activation energy. It is therefore consistent with this evidence to, at least tentatively, ascribe the 28,000 cal per mole creep process in A1 to the operation of cross slip as the rate controlling mechanism. Although creep has been rather well-explored in the dislocation climb range, very little, excepting perhaps the activation energy, is known about creep in the cross slip range. This investigation was undertaken, therefore, to reconfirm the activation energy for creep in polycrystalline A1 in the intermediate temperature range and to study the strain-time relationship for creep at a constant stress in the cross slip region. EXPERIMENTAL PROCEDURE AND TECHNIQUE The three types of high-purity A1 alloy specimens described in Table I were used in this investigation. Specimens were milled from 0.100 in. thick cold-rolled sheets with their tensile axes in the rolling direction following which they were given a recrys-tallization and grain growth heat treatment at temperatures well above those which were subsequently employed in the creep tests.

Jan 1, 1962

By Nigel B. Pearson

INTRODUCTION Chingola Division of Nchanga Consolidated Copper Mines Limited is situated in the Copper- belt of Zambia, approximately 30 kilometres south of the border with Zaire. It is the country's leading producer of copper and its principal foreign exchange earner, and has been, in its time, the largest single copper producer in the British Commonwealth and the second largest in the world. There are two 'primary' copper sources at the Division, namely Underground producing approximately 100 000 tonnes of contained copper annually, and the Open Pits, producing approximately 200 000 tonnes annually. Following the commissioning of the Tailings Leach Plant in 1974, treating final tailings, metallurgical recovery increased from approximately 70% to in excess of 80%. Extensive deposits of cobalt are present and following the rise in the market price in 1979 and 1980, plans were made to extract this separately. Cobalt production commenced in 1980 from Underground and will increase substantially when the Open Pit cobalt reserves become available in approximately 2 years time. This paper covers the development of the production method of the principal Underground Orebody, known as the Lower Orebody, and the control systems applied to it. HISTORY The Lower Orebody, formerly known as the Nchanga West Orebody, was discovered by extensive Diamond Drilling operations in the late 1920's. Development of the Shaft complex and the orebody commenced and continued until 1931 when flooding from development on the 480 foot horizon necessitated closure. The mine remained closed until 1937 and production commenced from the Lower Orebody in 1939. Various expansion schemes of the Metallurgical Plants and the Underground workings were completed over the years with the result that Underground production increased steadily to in excess of 150 000 tonnes copper annually in the 1960's. In recent years production has levelled out at slightly less than 100 000 tonnes annually. This decrease is as a result of producing from a lower grade reserve, available within a more limited production front, and the application of production constraints in the light of experience gained. GEOLOGY Within the Nchanga Mining Licence Area of Chingola Division, there are a total of 14 known copper orebodies in the basal part of a thick succession of sediments of Pre-Cambrian age. The sediments rest on a Basement Complex of granites, gneisses and schists. Three superimposed stratiform orebodies, known locally as the Lower, Intermediate and Upper Orebodies, occur over a vertical interval of approximately 150m. Current annual copper production is approximately 300 000 tonnes. The ore originates from four different mining operations, one Underground mine and three Open Pits. Approximately 9 million tonnes of ore grading between 3 and 3.5% copper are milled annually. The two largest sources of ore, the Under- ground mine and the Nchanga Open Pit are on the gently dipping south limb of the Nchanga Syncline (Figure 1). In the Nchanga Open Pit the major source of ore is the Upper Orebody with lesser tonnages from the Intermediate and Lower Orebodies. In the Underground mine the Lower Orebody is the only orebody being mined, underground mining of the Upper Orebody being suspended for economic reasons in 1978. The typical stratigraphic column is illust- rated in the legend on Figure 1. The footwall rocks comprise either Nchanga Red Granite or the Arkose formation, both of which are hard competent rocks, offering generally good mining conditions for footwall development. Copper mineralisation of the Lower Orebody extends from the upper part of the Arkose to the top of the Lower Banded Shales. Overlying the Arkose is the Transition zone (0-18111 in thickness). The Transition zone comprise a lower member, Transitional Arkose, consisting of incoherent kaolinised sandstone, and an upper member, Transitional Lower Banded Shales consisting of normally soft bleached and silicified shale. The soft nature of the Transitional rocks is

Jan 1, 1981

By T. S. Fuller

The tensile properties of steels at elevated temperatures have been studied by numerous investigatorsll primarily for the purpose of determining their suitability for structural uses. Tests with this objective have been confined largely to the temperature range of approximately 20" to 500' C., because of the rapidly decreasing tensile strength at the latter temperature. Tests have also been extended to higher temperatures2 for the purpose of determining the most suitable forging range. This temperature range is also of theoretical interest. The phenomenon if increased tensile strength and decreased ductility .in the so-called "blue-heat" range, approximately 200" to 300" C., is well known. It was suggested to the Bureau early in 1926 that a study of the properties of rail steels in this temperature range might prove of interest in relation to the studies of transverse-fissure failures in rails. There were available from another investigation3 several new rails from several different heats made to approximately the same specification requirements, the complete manufacturing histories of which were known. It was decided to determine, at least in the preliminary series of tests, only the ultimate tensile strength, elongation and reduction of area, because of the difficulty, time and cost involved in determining the proportional limit and yield point at elevated temperatures.

Jan 1, 1930

By M. R. Geer Olds, H. F. Yancey

THE increase in the number of coal-cleaning plants employing dense-medium processes occurring since 1946 is especially interesting when viewed historically. Both sand and magnetite were introduced as material for heavy mediums at about the same time, sand in the Chance process in 1921 and magnetite in the Conklin process in 1922, but from that point on their records diverge. The Chance process enjoyed a steady growth from its inception, whereas no additional magnetite plants were built in the United States for over 20 years. Then, following the close of the World War 11, magnetite was again introduced, this time with marked success. During the following years some 47 plants employing magnetite medium were built. This rapid growth of dense-medium cleaning has been concurrent with widespread adoption of full-seam mining on one hand and a return to a more competitive market on the other. At a time when changing mining practice has provided cleaning plants with dirtier coal and changing market conditions have simultaneously demanded a cleaner product, the industry has through necessity turned to improved preparation. The inherently greater sharpness with which dense-medium processes can separate coal from impurity is thus helping to hold the line against ever-increasing mining costs and at the same time assisting materially in retaining badly needed markets. Although dense-medium cleaning unquestionably offers a distinct advantage when the washing problem is difficult, other methods can provide almost equally high efficiency when the coal is easy to wash. Moreover, fine coal cannot yet be treated by heavy medium in a proved process, although the Driessen cyclone is in the pilot-plant stage. Most of the present types of dense-medium equipment have been in use only a few years, and the dearth of information in the literature concerning their performance characteristics is entirely understandable, Nevertheless this information is necessary if the process is to be intelligently applied to individual cleaning problems. Without data on the efficiency of a process in a particular type of separation, it is difficult to assess the advantage to be expected from it. Similarly, the role of particle size in heavy-medium separation is important in some cases, yet there is little published information on this aspect. To mention only one more of the numerous points on which essential information is lacking, the bearing of medium characteristics on performance has been discussed only in qualitative terms. It was with the hope of providing information on some of these points that the Bureau of Mines built a dense-medium pilot plant for cleaning coal at its Northwest Experiment Station in Seattle in 1950. The plant has been operated continuously since that time, and over 50 runs have been made on 7 coals exhibiting a wide range of washability characteristics. An idea of the magnitude of this work will be gained from the fact that examination of the plant products has involved some 600 float-and-sink separations and about 2500 ash determinations. A laboratory pilot plant is especially well adapted to investigate many aspects of performance because close control over test conditions can be exercised and because a large number of tests can be made rapidly. On the other hand, factors such as consumption of medium and other cost items can be investigated satisfactorily only in a commercial plant. Actually, the two forms of investigation should be complementary, with the laboratory work pointing the way for confirming tests in commercial units. Pilot Plant The dense-medium pilot plant employed for this work comprises a 24x30-in. drum-type separating vessel, a 12-in. densifier, a 12-in. magnetic separator, a 26-in. x 9-ft vibrating screen, and the necessary pumps and conveyors for handling materials. Arrangement of these units corresponds with the flowsheet used in most commercial plants, except that a thickener is not provided for the feed to the magnetic separator. Coal and refuse discharge from the separating drum to the vibrator, which is divided longitudinally down the center. Medium draining through the first 3 ft of the screen is recirculated directly to the drum. Sprays on the middle 3 ft of the screen rinse medium from the products, and the last 3-ft section is for dewatering. Dilute medium from the rinsing and dewatering sections is pumped to the magnetic separator, where magnetic solids are recovered. These are pumped to the densifier, from which they return to the medium-drainage sump by gravity through a demagnetizing coil. The drum-type separating vessel is a scale model of a commercial unit. Feed enters axially at one end of the drum just below the surface of the bath, and float material overflows through a circular opening at the other end. Particles sinking to the bottom of the bath are picked up by lifting flights bolted to the inner wall of the drum, elevated out of the bath, and sluiced to the vibrating screen. Baffles suspended in the bath prevent float material from entering the sink-lifting flights. About 8 gpm of medium is used to sluice the feed into the drum. An additional 15 to 24 gpm, depending upon operating conditions, is added through two pipes dipping into the bath behind the baffle on the side where the sink-lifting flights enter the bath. The bath available for separation is 2 ft long and 13 in. wide, giving an area of 2.08 sq ft. Depth of bath from the surface to the top of the sink-lifting flights, measured vertically below the axis, is 6 in.

Jan 1, 1954

By D. G. Westlake

Levine has concluded that prismatic slip in titanium is controlled by two distinct thermally activated processes in the temperature ranges O° to 220°K and 220" to 300°K. We feel that his evidence is not convincing. He has drawn a curve through the plot of activation volume u* vs temperature T with a break at 22WK, Fig. 2. We believe that considering the possible experimental error apd the fact that there are only four points between 4° and 196°K, one could be justified to represent these points by the smooth curve we show as a solid line in Fig. 12. Levine believes that (8rf/aT)+ is discontinuous at 220°K, but he hints that he recognizes the possible doubt. We doubt seriously that the results shown in Fig. 1 indicate a discontinuity at 220°K. Certainly if he were justified in drawing a straight line through only four points on Fig. 2, we are even more justified in drawing straight lines through the seven points between 77" and 297°K on Figs. 1 and 3, because a straight line touches all seven of his circles. If his measurements of t were so precise that his decision to draw two curves in Fig. 3 is valid, then the experimental technique is deserving of further mention. We have drawn a smooth curve through the results in Fig. 3 and used this and the curve in Fig. 12 to obtain the values of shown in Fig. 13. Finding continuous is a contradiction of Fig. 4, which deserves comment itself. Levine obtained the solid circles in Fig. 4 using Fig. 2 and a smooth curve through the points shown in Fig. 3. We have replotted these solid circles in Fig. 14. We submit that there is no basis for his two straight lines (dashed in Fig. 14), and therefore have superimposed the curve that we believe is indicated. Certainly, there is reasonable doubt that a discontinuity in dH/dT exists at 220°K, even when one accepts Fig, 2. Furthermore, if one draws the straight lines indicated by least-squares treatments of the results presented as open circles in Fig. 4, their intersection is at about 196°K, not at 220°K as shown. Thus the alleged agreement between Figs. 3 and 4 with respect to "break" temperatures is not apparent. In Fig. 9, Levine has drawn a straight line very much different from the one indicated by a least-squares treatment. Obviously, he has done this because he recognized that H = 0 when T = 0. He is consistent in drawing the low-temperature curves in

Jan 1, 1968

Jan 1, 1936

By Charles Legrand

The power plant of the Burro Mountain Copper Co. is located near Tyrone, N. M., at 5,950 ft. elevation. It is interesting because it uses the largest stationary Diesel engines in the United States. The general layout of the plant is shown in Figs. 1 and 2. The building is of steel-frame construction with hollow-tile walls and corrugated-iron roof; the floor is concrete and all windows have steel frames, making a fireproof building. The building is made to accommodate one more electric generator unit which is being installed; also two 4,000-cu. ft. air compressors direct-connected to three-cylinder engines with the same size cylinders as the electric units. The power plant consists of two 815-kva., 60-cycle, three-phase, 6,600-volt, 180-r.p.m. generators, direct-connected to five-cylinder, two-cycle Diesel engines; the necessary circulating pumps for cooling water, cooling towers and oil storage tanks. A portion of the power is used in the same building as the generating units to run one of two 200-kw. rotary converters delivering 260-volt direct current, and a 2,500-cu. ft. air compressor, delivering air at 95 lb. pressure, direct-connected to a 400-hp., 6,600-volt, three-phase synchronous motor. The remainder of the power is used to run a 1,500-ton concentrating mill 3 ½ miles from the power house, various hoists, pumps and motors about the mines and the lighting and water-works of the town of Tyrone. The direct current is used for electric traction in the mine, also to run hoists and motors that were installed before the new power plant was built. Due to the large hoists that take starting peaks of 250 kw. and the 10-ton electric locomotive that takes at times 150 kw., the load is extremely variable, as shown in. Fig. 3.

Jan 9, 1916

By J. P. Witherow

Mr. BIRKINBINE's description of the bad working and sudden chilling of the Warwick Furnace last summer, seems to me quite phenomenal in blast-furnace practice. During my connection with the management of blast furnaces for many years, in the anthracite, cake, and raw-bituminous coal districts, such a me has never come under my notice. It seems to me that the scaffold in the zone of fusion could not have established itself (so that the whole power of the engines, with a blast pressure ranging from 12 to 18 pounds per square inch, could not penetrate), unless the furnace was previously badly scaffolded and bridged in some of its upper sections, from 15 to 30 feet above the hearth level, either in the zone of chemical action or the beginning of the zone of fusion. Such a 'barrier to the upward gaseous currents would cause a suspension of combustion which may have superinduced the scaffold that is assumed to have existed in the zone of fusion. It is difficult to conceive that an agglomerated mass almost across the section of the lower bosh should have formed all at once, if the furnace was free and porous for the reaction of the gaseous currents, above. Therefore, I think, the Warwick furnace was by no means in as free, open, and fine internal condition at the time the trouble commenced as the description of its regular working would lead us to suppose. I do not question the accuracy of the statements relative to the output of iron, etc I merely claim that the phenomenta, as described, could not have been almost instantaneously developed, unless there had been an incipient scaffold and lodgment above the zone of fusion. I would state my hypothesis to be that the stock lodged or collected in some of its upper sections during fie process of blowing-in, on which, as a basis, more or less accumulations may .have lodged, and thus, indirectly, caused the formation of the suc-

Jan 1, 1881

By C. E. Sims, G. A. Lillieqvist

The following presentation is more a summation of observations made and conclusions drawn over a period of three to four years than a report of experiments made with a definite objective. Tests were made, of course, to verify hypotheses, but the large mass of data was collected incidental to other work. Lack of space forbids the presentation of more than a few of those data. On the other hand, it is unfortunate that some of the most illustrative examples are not available. The work is admittedly incomplete but is presented in the hope that it will stimulate further work along a similar line. The answers to many of the questions left unanswered by this paper will be found only after long and painstaking study, such as is scarcely possible in a commercial laboratory where problems of immediate importance consume most of the time. It is also hoped that this study will in some measure throw light on some of the dark spots of steelmaking, and help to clarify some of the annoyingly vague references to "good practice." Only cast steel was studied and the conclusions apply to cast steel, but it is the authors' opinion that the quality of rolled or forged steel is largely dependent on the quality of the ingot, which is a casting. Although this discussion deals almost entirely with inclusions, it is important that the reader understand that only those natural inclusions which owe their origin to the solubility of certain materials in steel and to the chemical reactions of steelmaking are considered. Accidental inclusions, such as entrapped particles of slag, nozzle, runner, or mold arc outside its province and are not considered. Such accidental inclusions are always dangerous and should be carefully avoided. Relation between Ductility and Nonmetallic Inclusions In September, 1928, an acid electric furnace was put into operation at the Indiana Harbor plant of the American Steel Foundries. Following the usual practice, aluminum additions were made in order to obtain sound castings. Immediately the problem of low ductility arose. Previous to that time, only open-hearth steel was produced, and, aluminum being used rarely in that steel, there had been practically no trouble with low ductility in the lower carbon range.

Jan 1, 1932

By R. R. Morral, F. M. Krill, K. T. Aust

THE occurrence of twins in aluminum ingots cast under certain conditions has been well established. For instance, Herenguel and Lacombe1,2 have observed a twinned columnar structure in 99.5 and 99.9 pct Al, solidified by continuous or semicontinuous casting. The structure, which they termed "basaltic," appears in the form of elongated crystalline fibers of alternating luster. Herenguel and Lacombe2 found sections of flat crystalline plates composed of twin crystal plates, in which the twin plane (111) and a [I121 direction contained in the twin plane are parallel to the principal heat gradient. This orientation relationship is different from the usual orientation texture observed in cast face-centered cubic metals, where the [100] direction is approximately parallel to the long columnar axes lying in the direction of growth." A large amount of twinning has also been observed by the present authors in semicontinuous cast aluminum of 99.2 pct purity, Fig. 1. Mondolfo4 has suggested that the twinned cast structure forms by a deformation twinning process as the result of uneven cooling stresses within the ingots. This suggestion appears to be inconsistent with the large amount of twinning observed in the semicontinuous cast ingots and with the large difference of dendritic

Jan 1, 1953

By C Woodward

AT the Belmont mine of the Anaconda Copper Min-ing Co., Butte, Mont., a new double-drum electric main hoist and a single-drum electric auxiliary hoist have recently been placed in operation to meet the demand for greater hoisting depths, and to replace a steam-driven hoist which was inefficient and inade-quate for the new conditions. The ease of operation, safety, and efficient consumption of power of the main hoist all tend toward improved hoisting conditions. It consists of a 2150 hp. d. c. motor, directly connected to a shaft carrying two drums equipped with clutches and brakes. The auxiliary hoist consists of a 1000 hp. d. c. motor directly con-nected to a single drum provided with the necessary brakes. Fig. 1 shows the general arrangement of the complete installation, with idler towers and head-frame. The headframe is built of Bethlehem H and G rolled sections, equipped with two 12-ft. dia. cast steel sheaves for the main hoisting ropes. The sheaves are mounted 129 ft., 6 in. above the collar of the shaft; and above them trolleys are provided for their removal or repairs. Between the headframe and the hoist machinery are two idler towers for supporting the ropes on 3-ft. man-ganese steel idlers provided with roller bearings which fleet on a fixed shaft. As a city street is between- the mine headframe, and hoists, it is necessary to protect the thoroughfare from drippings of the preservative compound used on the ropes which pass over the street. For this purpose, at the point where the hoisting ropes cross the street ,a bridge of trusses, supported on columns and roofed with corrugated iron, is provided from property line to property line. Since one of the idler, towers is set close to the street, a part of the structure is built in-tegral with the street covering. The general design of idler tower as well as the covering over the street is shown at A in Fig. 1.

Jan 9, 1927

By Lucien Eaton

THE accuracy with which detachable bits can be sharpened and gauged by grinding has made possible refinements in design that cannot be attained by forging. This applies particularly to gauging, where accuracy is of the greatest importance. Gauge changes of 1/32 in., and even 1/16 in., are now feasible, whereas 1/16 -in. changes are about the limit in accuracy for forged bits. These small changes in gauge are feasible, however, only with certain fundamental principles of design. Within reasonable limits the depth to which a hole can be drilled is dependent primarily upon the loss in gauge, or diameter, of the bit. In drilling there is always a certain amount of " overdrilling "-the bouncing of the bit in the hole and its rotation wear away the walls-so that the actual diameter of the hole is greater than the diameter of the bit. The amount of this overdrilling depends upon the kind of rock being drilled, and is sometimes sufficient to permit the use of the same diameter of bit for the full depth of the hole. If, for instance, a 1 ¾ -in. diameter bit overdrills 1/32 in. in diameter, and loses 1/64 in. in gauge in drilling 2 ft. of hole, the bottom of the hole is still 1/64 in. larger than the original diameter of the bit, and a new 1 3/4-in. bit can be used to drill the next 2 ft. If, however, the overdrill is less than the loss in gauge, the next bit must be of smaller diameter. The first requisite in reducing loss of gauge in drilling is to have the points and corners of the wings of the bit cut circles of the same diameter, thus producing a reaming surface with continuous bearing on the wall of the hole. The greater the reaming surface, the less the gauge loss that is to be expected. If the reaming surface, however, is too large, it may impede the rotation of the bit through excessive friction. This principle was probably first recognized in the design of the Carr bit (Fig. 1)) a bit with a single cutting edge of wide angle, a definite reaming edge, and a large center hole. The same idea was followed in designing the double-taper cross bits, which are now standard almost everywhere Fig. 2 illustrates a typical double-taper cross bit. The taper from the point of the wing A to a little behind the corner B is 5° and from

Jan 1, 1939

Jan 1, 1936

By Basil B. Zavoico

Production of crude oil in the U.S.S.R. during 1940 is estimated at 222,600,000 bbl., as compared with the revised figure for the preceding year of 220,866,000 bbl., an increase of 0.79 per cent, and at least 20 per cent under the planned output. Baku district (Transcaucasus) remained the outstanding producing area, accounting for 72.33 per cent of the country's production, and this relative importance of the Apsheron Peninsula oil fields will probably continue lor many years to come. In the North Caucasus the Grozny-Dagestan area further extended its decline and the current production in this district amounts to only 6.06 Per cent Of the total U.S.S.R. output, as compared with 35.76 Per cent in 1931 The production of the Kuban-Maikop district remained virtually ullchanged from the preceding year, amounting to 7.64 per cent of the country's total. The Caucasian Province, therefore, in I940 produced 86.03 Per cent of the total production of the U.S.S.R. The promising new Permian Basin district, called Ural-Volga, and referred to sometimes as the "Second Baku," increased its production appreciably during 1940 and the output of that area amounted to 8.31 per cent of the country's total, as compared with 6.35 per cent during 1939. Production in the secondary districts of Russia continued unchanged though some new discoveries in the Emba salt-dome basin suggest that the output of that district may show some increase during the coming years. The drilling operations in Russia during 1940 reached a stage of virtual breakdown and the original footage plans were fulfilled only to the extent of around 50 per cent, while exploration drilling schedules were fulfilled only to the extent of about 25 to 35 per cent. This condition was due to the obvious lack of modern drilling rigs, of good drill pipe and of strong casing. Extended breakdowns occurred with deeper drilling, particularly because with few good drilling rigs available speed was being attempted with rotation rates that the inferior drill pipe and bits could not stand. Twistoffs and endless fishing jobs resulted. The difficulties became so serious a handicap to the Soviet oil industry that it was found imperative to order American drilling equipment and pipe in quantity, a policy that was never favored in the U.S.S.R because it requires expenditure of gold or exchange in major volume, and such exchange usually is reserved only to less bulky key items or to orders directly connected with armaments; also, oil-field machinery is very bulky and the prevailing lack of bottoms makes the transport of the indicated equipment a major problem. But little progress was similarly made in other branches of the Soviet oil industry. NO new major pipe-line construction work was undertaken, consequently the railroads continued to be overburdened with traffic that could be much more effectively handled by pipe lines. No new refinery construction of any magnitude was initiated, though several plants that had been in course of construction for several years were completed during 1940 and operation 554

Jan 1, 1941

By Basil B. Zavoico

Production of crude oil in the U.S.S.R. during 1940 is estimated at 222,600,000 bbl., as compared with the revised figure for the preceding year of 220,866,000 bbl., an increase of 0.79 per cent, and at least 20 per cent under the planned output. Baku district (Transcaucasus) remained the outstanding producing area, accounting for 72.33 per cent of the country's production, and this relative importance of the Apsheron Peninsula oil fields will probably continue lor many years to come. In the North Caucasus the Grozny-Dagestan area further extended its decline and the current production in this district amounts to only 6.06 Per cent Of the total U.S.S.R. output, as compared with 35.76 Per cent in 1931 The production of the Kuban-Maikop district remained virtually ullchanged from the preceding year, amounting to 7.64 per cent of the country's total. The Caucasian Province, therefore, in I940 produced 86.03 Per cent of the total production of the U.S.S.R. The promising new Permian Basin district, called Ural-Volga, and referred to sometimes as the "Second Baku," increased its production appreciably during 1940 and the output of that area amounted to 8.31 per cent of the country's total, as compared with 6.35 per cent during 1939. Production in the secondary districts of Russia continued unchanged though some new discoveries in the Emba salt-dome basin suggest that the output of that district may show some increase during the coming years. The drilling operations in Russia during 1940 reached a stage of virtual breakdown and the original footage plans were fulfilled only to the extent of around 50 per cent, while exploration drilling schedules were fulfilled only to the extent of about 25 to 35 per cent. This condition was due to the obvious lack of modern drilling rigs, of good drill pipe and of strong casing. Extended breakdowns occurred with deeper drilling, particularly because with few good drilling rigs available speed was being attempted with rotation rates that the inferior drill pipe and bits could not stand. Twistoffs and endless fishing jobs resulted. The difficulties became so serious a handicap to the Soviet oil industry that it was found imperative to order American drilling equipment and pipe in quantity, a policy that was never favored in the U.S.S.R because it requires expenditure of gold or exchange in major volume, and such exchange usually is reserved only to less bulky key items or to orders directly connected with armaments; also, oil-field machinery is very bulky and the prevailing lack of bottoms makes the transport of the indicated equipment a major problem. But little progress was similarly made in other branches of the Soviet oil industry. NO new major pipe-line construction work was undertaken, consequently the railroads continued to be overburdened with traffic that could be much more effectively handled by pipe lines. No new refinery construction of any magnitude was initiated, though several plants that had been in course of construction for several years were completed during 1940 and operation 554

Jan 1, 1941