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By C. H. Green

Recently two papers on the structure of eutectics were read before thc British Institute of Metals, one by F. L. Bradyl and the other by A. Portevin. 2 In the preparation of photomicrographs of laboratory specimens as illustrations for student work, and in connection with various researches, the writer has been working along the same line for some time and, therefore, was much interested in the results given, especially those of Portevin. He has correlated the information on the subject and has outlined a broad scientific basis for future work along this line. It is largely with the idea of adding something by way of illustration to the work already done that this paper is written. The eutectic is generally understood to mean the alloy having the lowest melting point of any alloy of the system, and in which the melt becomes supersaturated with reference to the two or more phases which separate out simultaneously or alternately. From the surface pattern, the eutectic means a structure with no primary crystals. It is not always easy to get the exact eutectic composition in making any alloy; in fact, a slight range in composition may be expected and some of the coarser particles may be richer in either of the composing metals. Even when chemical analysis shows the right composition, differences in density of the pure metals may cause segregation, or supercooling resulting from different crystallizing properties may give a surface pattern the appearance of being on one side of the eutectic point. However, the alloys on either side of the eutectic composition, showing primary crystals of one of the phases and the effect of the proeutectic form on the surface pattern of the eutectic make a very interesting study, which is slightly touched upon in this paper. The phases, or microconstituents, of a eutectic may be pure metals (Pb-Sb, Tl-Au), solid solutions (Cd-Zn), a pure metal and a solid solution (Ag-Sn), or a pure metal or solid solution with an intermetallic compound (Sb-Mg, Sb-Cu, As-Cu, Bi-Te, etc.).

Jan 1, 1925

By P. Wynblatt, J. E. Dorn

The effect of strain rate, temperature, and interstitial impurity concentration on the flow stress was investigated in a poly crystalline Fe-2 pct Mn alloy. The temperature dependence of the flow stress was found to be independent of the interstitial impurity concentration, removal of interstitials merely decreasing the atherma1 stress level. Below 160%, the large temperature dependence of the flow stress was interpreted in terms of the Dorn-Rajnak theory of the Peierls mechanism of plastic deformation. Above 160"K, some other thermally activated intrinsic meckanisln seems to be operating. The strong temperature dependence of the flow stress of bcc iron at low temperatures has been studied by numerous investigators and interpreted in terms of several different thermally activated dislocation mechanisms. The following mechanisms have been proposed: a) interaction of dislocations with interstitial impurity atoms1 or with solute atoms in general; b) interaction of dislocations with clusters of impurity atoms;3 c) resistance to the motion of dislocations due to jogs on screw dislocations;* d) resistance to the motion of dislocations due to the Snoeck effect;' and e) interaction of dislocations with the intrinsic resistance of the bcc lattice or Peierls "hills". In previous work: it was found that the plastic behavior of polycrystalline iron containing 2 wt pct Mn (and 100 ppm C + N) is in good agreement with predictions based on the Peierls mechanismlo from 77" to 160°K whereas from 160" to about 370°K another as yet unidentified thermally activated mechanism is operative. The purpose of the present investigation was to determine the effect of interstitial impurities on the temperature dependence of the flow stress, by purifying the previously investigated alloy with respect to interstitial impurities. The investigation revealed that removal of the interstitial impurities to a level that eliminated Cottrell locking and the Portevin-LeChatelier effect (i.e., serrations in the stress-strain curve due to dynamic strain aging) affects neither the strong temperature dependence of the flow stress at low temperatures nor the general characteristics of the higher-temperature behavior. Such purification, however, lowered the athermal yield stress. I) EXPERIMENTAL PROCEDURE AND RESULTS The material used in this investigation consisted of an Fe-2 wt pct Mn alloy having the following additional elements present: 0.004 pct C, 0.006 pct N, 0.05 pct 0, 0.004 pct S, 0.003 pct P, and 0.001 pct Si. This was the same material as that studied in the previous investigation and will henceforth be referred to as the "impure material". The as received -jr by 2 in. hot-rolled bars were cold-rolled to 0.08-in. thickness, recrys-tallized under argon for 30 min at 80O0C, and further cold-rolled to 0.053-in. thickness. Flat tensile specimens 3 in. wide having a 1.625-in.-long gage section were machined from the sheet and then purified. The first stage of purification consisted of holding the specimens at 850°C for 24 hr in a stream of hydrogen saturated with water vapor at room temperature. Material prepared in this fashion will subsequently be referred to as "wet hydrogen purified material". It has been estimated that this type of treatment reduces the carbon content of iron to less than 5 ppm.' The second stage of purification consisted of holding the specimens in a closed system through which hydrogen was circulated past the heated specimens as well as a zirconium hydride getter. The specimens were held at 850°C and the getter at 800°C. The process was continued for 212 hr. The starting material for this process was wet hydrogen purified material. Material prepared in this manner will be referred to as "ZrHz purified material". The technique used was substantially the same as that described by Stein et al.' who have shown that the carbon level at the end of the process is reduced to about 5 parts per billion. The mean grain size of the three materials was found to be -60 C( for the impure material and -160p for the wet hydrogen and ZrHz purified materials. All tensile testing was carried out on an Instron Testing Machine either in controlled-temperature baths or at fixed-point baths such as boiling liquid nitrogen, and so forth. When fixed points were used the temperature was kept constant to better than *1°C whereas in all other tests the temperature variation was less than *2°C. Test of Purity. It was not possible to determine the purity achieved at each stage of the purification because the resulting level of interstitial impurities was near or below the limit of sensitivity of standard analytical techniques. It was possible, however, to obtain a qualitative measure of the extent of purification from the appearance of the stress-strain curves for the three materials. Fig. 1 shows the stress-strain curves for the three different materials tested at 300°K. It can be seen that the effect of wet hydrogen

Jan 1, 1967

By Wieland D. R., Hurst R. E., A. R. Hendrickson

Analysis of acidizing techniques, in correlation with reservoir data and a backlog of past treatments, has resulted in the development of a valuable engineering guide for planning acidizing treatments. Such treatments fall into three categories: (I) acid injection into the pores of the matrix; (2) acid injection into natural formation fractures at less than parting pressure; and (3) combination acidizing-fracturing treatments in which acid solutions (without propping agents) are injected at treating pressures sufficient to open and extend fractures through which the acid flows. Because the spending tirnze of acid during a specific well treatment does not change appreciably, maximunl penetration is attained when the first increment of injected acid is completely spent. Additional acid injection cannot be expected to further extend the benefits of the treatment. Depth of penetration will depend upon the reaction rate of the acid under treatment conditions, the injection rate of the acid into the matrix or fractures and the area-volume relationship existing in the flow channels. Based on Darcy's flow formula, extremely low injection rates must be used in order to keep bottom-hole injection pressures below formation fracturing pressure. As a result, only limited penetration of unspent acid will occur. Treatment records indicate that, in most acidizing treatments, formation parting pressures are exceeded, greatly extending acid penetration. Under these conditions, stimulation benefits are limited to the fracture area produced during the spending time of the first increment of acid injected into the formation. This area may be calculated from laboratory and well data to estimate depth of penetration. This, in turn, may be correlated with productivity data to assist The art of gas and oil well acidizing has been characterized by many changes in treating materials and techniques since its inception. These developments have been designed to provide greater production increases. prolong production declines and shorten payout time. Such improvements have been based primarily on data derived from laboratory research and field experience. As more of the variables influencing these treatments have been recognized and evaluated, acidizing has become less of an art and more of a science. Recent studies of fracturing treatments,' in light of individual well conditions and the results of thousands of fracturing treatments, made possible the formulation of an engineering guide that is now being used to select optimum treating techniques and to forecast probable results of such treatments. A similar analysis of the factors controlling acidizing treatments has been made and is the basis for this paper, The findings herein can be used as a guide in the selection of acidizing solutions and techniques, tailored to fit specific well conditions and to provide optimum stimulation per dollar cost. Acidizing treatments may be classified into three basic categories—(1) treatments in which the acid is injected uniformly into the pores and flow channels of the matrix, (2) treatments in which the acid enters natural fissures and fractures in the formation at less than fracturing pressures and (3) injection of acid into the formation at a pressure sufficient to open and extend fractures into the rock through which the acid penetrates (without the inclusion of a propping agent). TYPE 1 — MATRIX ACIDIZING This category consists of treatments in which acid solutions are injected into a homogeneous carbonate

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970

By J. Murray Riddell, Arthur E. Anderson, Grover J. Holt

OUT of the depths of each business cycle we emerge with a stimulus for greater efficiency and a realization of progress in industrial technique. The recent years have not been an exception to this rule, and in retrospect we sense that significant advances have been made in mining practices on the Mesabi Range, with emphasis on speed of production and economy of operation. A survey of such changes does not reveal anything that is fundamentally or basically new, but it is apparent that ideas and machines that have found wide application in industry elsewhere, with necessary modifications, are being put to productive use in mining iron ore. Innovations in transporting and conveying ore from the open pits stand out with greatest prominence, so that, in the main, it will be with these features that this paper will treat, in the review of progress made to this date. The late years have witnessed the development of small deposits for open-pit operation, the cleaning up of pits nearing exhaustion, and in two large operations the approach to the limit for haulage by steam locomotives, owing to grades, making it necessary to recover the ore in slopes, track benches and below the power-shovel limit by a milling and scram operation. For several years small standard-type dump trucks have been used in clean-up or scram operations within the pits, but during the seasor1 of 1936 the first effort was made to use heavy truck haulage in the transportation of stripping and iron ore. The new units in use today vary from 15 to 20 tons in capacity, and are equipped with powerful gasoline or Diesel engines and supplemented by trailers or wagons on rubber or crawler treads. The Diesel engine finds favor with most operators because of fuel economy. These units develop upward of 125 hp. in the 15-ton trucks. The speed of these trucks varies from 2 miles per hour in low gear to 35 in high gear. The transmission affords 12 speeds forward and 3 speeds in reverse.

Jan 1, 1938

By T. H. Alden

Using alternating tension-compression straining, the hardening of magnesium oxide single crystals was studied up to large stresses and strains. At 0.25 pct plastic strain amplitude, the hardening curve is approximately linear with slope 25,000 psi from the shear yield stress, 7 to 8000 psi, to 35,000psi. Above this stress, the slope decreases. The strain hardening behavior of MgO is considered qualitatively similar to that of metal single crystals. The relatively high stress attainable by strain hardening is associated apparently with the high yield stress on the cross-slip system, (001) <110>. Cleavage fracture during testing is uncommon. It is argued that the centers of high internal stress at glide band intersections, at which cracks tend to nucleate, are dispersed by cyclic strain. Special features of the glide band structure produced by cyclic strain and revealed by dislocation etch pits, support this view. Strain hardened MgO has mechanical properties greatly superior to the as-received material: yield stress, greater than 100,000 psi; elongation to fracture about 1 pct. A material is said to strain harden if the yield stress increases with an increment of plastic strain. This definition is usually applied for straining done in one direction, but is also applicable when the strain direction is periodically reversed, Fig. 1. For certain metal single crystals, data are available which permit a comparison of the hardening behavior for cyclic straining and for tension straining.&apos;-4 With certain qualifications, these data show that the same processes of hardening are operative in each type of test.5 Despite this fact, the importance of the technique is not immediately evident, although tension-compression studies of the common metals appear to suggest some deficiencies in theories of strain hardening developed exclusively on the basis of tensile tests. However, a recent observation suggests that the cyclic straining method may be very useful for studying semibrittle crystals in which large plastic strains are not accessible in unidirectional testing. The observation is that zinc crystals, when strained in tension-compression at -52°C, do not fail by cleavage at low stress (-500 psi)6 as they do in tension, but harden to a limiting stress of more than 5000 psi over a total plastic strain of about 600 pct.2 An important characteristic of the behavior of zinc crystals is the high stress, relative to the yield stress, attainable by strain hardening. By comparison, the hardening of aluminum single crystals tested by an identical technique saturates at 1100 psi. This difference is best explained by the cross-slip hypothesis of dynamic recovery.7,8 In zinc, cross slip is difficult because of the high yield stress for glide on planes other than the basal plane in the < 1120 > zone. The present work was undertaken in order to test whether these methods and ideas are applicable to other materials. Magnesium oxide single crystals, in common with most crystals of the rock-salt structure, deform plastically but fail by cleavage after a small strain when tested in tension. It was hoped that larger strains would be attained using tension-compression. There is, in addition, evidence 8a which shows that slip on the probable cross system, (001) < 110>, is difficult in magnesium oxide; it may therefore be possible to attain high stresses by strain hardening. 1) EXPERIMENTAL PROCEDURE Experimental methods used in this study were based in part on techniques reported in papers of Stokes, Johnston, and Li.&apos; MgO blocks, purchased from Norton Co., were used without further annealing. Specimens were cleaved to dimensions approximately 0.125 in. sq and 1 in. in length. The gage section, formed by chemical polishing, was sprinkled with 280 mesh silicon carbide particles in order to introduce fresh dislocations. The crystals were then cemented into cylindrical aluminum adapters and clamped in an Instron testing machine. One of two alternating straining programs was used. In the first, total cross-head travel was established and increased in steps after various numbers of cycles. In the second, a capacitance gage was used to directly measure the elongation of the specimen and the crosshead was controlled so as to keep the plastic strain amplitude constant. The straining was always symmetrical with respect to the initial, zero strain condition. While both procedures produce strain hardening, only the latter permits a measure of the total plastic strain so that hardening curves may be drawn. Constant plastic strain amplitude tests were done

Jan 1, 1963

By Frank Harwood Riddle

Products made from the ores of the sillimanite group, and synthetic substitutes for them, have unique properties, and service tests prove that they are playing, and will continue to play, a major part in improvements in metallurgical processes. Also, they are examples of how comparatively useless, rare, or "museum" minerals have been made available in commercial quantities. Until research developed uses for them, the minerals had been observed only in small, scattered masses, or crystals. Improvements in ceramic processes and compositions have come about very gradually. In the last 10 or 15 years, several outstanding developments have been made. Of greatest interest to the mining and metallurgical engineers are those affecting the composition, necessitating the mining of new minerals, and improvements in ceramic products used in metallurgical processes. Sillimanite and Mullite Sillimanite (Al2O3.SiO2), then known as Fibrolite, according to Dana,&apos; was first reported in 1792. In 1796, cyanite (Al2O3.SiO2) was first described.2 In 1873, Behrens published results of his microscopic study8 of porcelain. He showed that it contained what appeared to be crystals, which dissolved less readily than quartz and glass in hydrofluoric acid. Hussak4 described the presence of needlelike crystals as occurring in some instances in small amounts in certain porcelains. In 1908, Plenske and Zoellner both studied porcelain structure.5 Plenske observed that "sillimanite" might be present in masses of exceedingly minute grains which apparently were amorphous. Zoellner advanced the theory that the needlelike crystals were formed by a molecular change in the clay,

Jan 1, 1932

By R. E. Crockett

MAGNETITE mining and milling in the Eastern States was at capacity during 1940, owing to the heavy industrial, defense, and war demand, coupled temporarily with the almost total stoppage of imports of foreign ores other than those under American ownership. In New York State the active mines were those operated by the Republic Steel Corp. at Mineville and Lyon Mountain. Republic&apos;s new concentrating mill at Port Henry, handling the Mineville ores, went into production about the first of April. Other large magnetite deposits in New York are currently being prospected and studied and may be under development in the near future. In New Jersey the Mt. Hope and

Jan 1, 1941

By Jack R. Elenbaas, Donald L. Katz

A radial turbulent flow formula has been developed which permits the computation of the pressure drop for radial flow in gas wells whether the flow is laminar, turbulent, or partially laminar and partially turbulent. Using the formula a complete back-pressure curve has been calculated and analyzed by a comparison with existing back-pressure curves. A procedure is presented for computing the permeability of the porous media when the porosity, sphericity and average particle diameter are known. Introduction Recent studies on the flow of fluids through porous media have provided new methods for computing flow under turbulent conditions. The present study was initiated as an investigation of pressure -drop computations for flow through porous sands in gas wells and as an analysis of present-day back-pressure tests. Although laminar flow exists in the producing formation of gas wells under normal flow rates, turbulent flow does take place adjacent to the well bore. As the flow rate is increased, turbulent flow exists further and further into the producing formation and under open-flow conditions in relatively deep wells a considerable portion of the flow through the sand will be turbulent. In order to calculate a complete backpressure curve for relatively deep wells without resorting to extrapolation of curves drawn from data obtained in back-pressure tests or calculations made in the completely laminar region, it was necessary to have some means of calculating pressure drop for flow through porous sands under turbulent conditions.4.6 As a result, a radial turbulent flow formula was derived from the equations recently presented by Brownell and Katz.1 A rigorous solution of this equation involves a trial and error graphical integration. By making the assumptions of average viscosity, temperature and compressibility factor of the flowing gas, it was possible to graphically integrate the relation for the general case of radial flow of gases through the porous sand of gas wells. A chart has been prepared relating the value of the integral to the radius of the well bore and to the Reynolds number of the flowing gas. The chart actually performs the graphical integration and the solution of the formula is then reduced to a simple trial and error calculation. A typical back-pressure curve is calculated to show the transition from laminar to turbulent flow. This paper will briefly describe the pro-cedurcs for computing flow through porous media, including a method for predicting sand permeability from porosity and grain size and shape. Flow through Porous Media Brownell and Katzl recently reported a correlation for computing flow of fluids through porous media by the use of an enlarged Reynolds number and friction factor in which the porosity of the bed is included as an additional prime variable. The enlargcd Reynolds number com-

Jan 1, 1948

By Donald L. Katz, Jack R. Elenbaas

A radial turbulent flow formula has been developed which permits the computation of the pressure drop for radial flow in gas wells whether the flow is laminar, turbulent, or partially laminar and partially turbulent. Using the formula a complete back-pressure curve has been calculated and analyzed by a comparison with existing back-pressure curves. A procedure is presented for computing the permeability of the porous media when the porosity, sphericity and average particle diameter are known. Introduction Recent studies on the flow of fluids through porous media have provided new methods for computing flow under turbulent conditions. The present study was initiated as an investigation of pressure -drop computations for flow through porous sands in gas wells and as an analysis of present-day back-pressure tests. Although laminar flow exists in the producing formation of gas wells under normal flow rates, turbulent flow does take place adjacent to the well bore. As the flow rate is increased, turbulent flow exists further and further into the producing formation and under open-flow conditions in relatively deep wells a considerable portion of the flow through the sand will be turbulent. In order to calculate a complete backpressure curve for relatively deep wells without resorting to extrapolation of curves drawn from data obtained in back-pressure tests or calculations made in the completely laminar region, it was necessary to have some means of calculating pressure drop for flow through porous sands under turbulent conditions.4.6 As a result, a radial turbulent flow formula was derived from the equations recently presented by Brownell and Katz.1 A rigorous solution of this equation involves a trial and error graphical integration. By making the assumptions of average viscosity, temperature and compressibility factor of the flowing gas, it was possible to graphically integrate the relation for the general case of radial flow of gases through the porous sand of gas wells. A chart has been prepared relating the value of the integral to the radius of the well bore and to the Reynolds number of the flowing gas. The chart actually performs the graphical integration and the solution of the formula is then reduced to a simple trial and error calculation. A typical back-pressure curve is calculated to show the transition from laminar to turbulent flow. This paper will briefly describe the pro-cedurcs for computing flow through porous media, including a method for predicting sand permeability from porosity and grain size and shape. Flow through Porous Media Brownell and Katzl recently reported a correlation for computing flow of fluids through porous media by the use of an enlarged Reynolds number and friction factor in which the porosity of the bed is included as an additional prime variable. The enlargcd Reynolds number com-

Jan 1, 1948

By W. R. Crane

The support of roof in mines is dependent largely on the character of the top rock and its occurrence. The formations overlying the orebed in the Birmingham district are sandstone and slate. The sandstone occurs in beds of sufficient thickness to constitute important elements in the system of support; the slate is relatively strong but is thinly stratified. The alternating beds of sandstone and slate furnish an excellent combination in that the former are strong and the latter are impervious; however, their continuity is broken by jointing or slip planes. The thickness, and consequent weight of the overlying formations, is important in so far as the size and arrangement of pillars are concerned but may be relatively unimportant, temporarily at least, as a factor in the support of the roof. The position of the respective beds of sandstone and slate, their relative thickness, and their inherent weakness because of the occurrence of faults and folds, may exert a predominating influence on their support in place, which is affected only in part by the weight of the cover. The only mine in this district that operates under a cover of approximately 2000 ft. (1900 ft.) is the Shannon so-called "twin slope," situated 14,000 ft. southeast of the portal of the No. 7 mine of the Tennessee Coal, Iron, & Railroad Co. The conditions in this mine indicate what may be expected in other mines when the same depth is attained; in fact, the effect of pressure is shown in several mines that have not reached two-thirds of that depth. It is evident, then, that conditions affecting the support of roof will not improve but rather will become more difficult with the extension of the mines into the valley under a constantly increasing weight of cover. Further, the difficulties will probably be augmented by the occurrence of faults and other disturbed ground, the presence and position of which at present are largely unknown.

Jan 1, 1925

By R. H. SWEETSERS

A Discussion of Mr. Hall&apos;s paper, read at the Washington Meeting, February, 1905. R. H. SWEETSER, Sault Ste. Marie, Ont. (communication to the Secretary*) :-The recent work of furnace No. 1 of The Algoma Steel Co., at Sault Ste. Marie, Ontario, using for fuel, first all charcoal, then a mixture of charcoal and coke, and finally all coke, upsets the estimates given by Mr. Hall in the last two paragraphs of his interesting paper. Furnace No. 1 was built for a charcoal-furnace, and it was run on charcoal fuel from March 6 until July 9, 1905; then, for a week, a mix¬ture of charcoal and coke was used, and, since then, coke alone has been the only fuel. It is to be regretted that at no time, during the run on charcoal, was there enough charcoal on hand and " in sight&apos;" to warrant blowing the furnace as hard as it should have been blown. During April and May the supply of charcoal limited the amount of blast to about 8,000 cu. ft. of air per min.; in June, to about 8,900 cubic feet. For a few days in July about 10,000 cu. ft. were blown, and then the maximum daily output, 173 tons (2,240 1b.), was reached. At all times the quality of the charcoal was poor, having been made from wood cut from three to four years before. Doubtless this furnace could have produced .200 tons of iron per day, even with this quality of charcoal, provided the supply had been sufficient; and with good charcoal I see no reason why it could not make as much iron per day as it has made with coke. During the one week of using a mixture of charcoal and coke, the changes in quantity of blast and composition of the slag followed each other so closely that no single condition could last long on the furnace. The results were good, and the pig-iron was about the same in quality as that made during the

Jan 1, 1906

By Dorsey Hager

DORSEY HAGER, Tulsa, Okla.-I have been asked why the Dexter region is dry. I would like to know myself. I drilled two dry wells on that same anticline which has production to, the east and west and south and north. That same anticline has seven domes that have been drilled and only three have proven productive out of the seven. In other words, you have one well-defined anticline with seven domes along its length, all mappable, all checked carefully, and only three of those seven have produced. Now why the others did not produce, I do not know. L. L. HUTCHISON, Tulsa, Okla.-There is no appreciable difference in the character or structure of the sands? DORSEY HAGER.-NO, I have never .been able to see any appreciable difference in the sands where we got production and Where we did not get it. I will say that the deepest wells we drilled, in the vicinity of Dexter, had no sands at the point where they got the production farther north. I. N. -KNAPP, Ardmore, Pa.-I first went to the Kansas-Oklahoma field in 1895 and occasionally thereafter until 1899, when I began drilling; for oil. I started close to wells that were producing gas and which, when drilling, lead made good oil showings in a sand above the gas. I thought these showings were good enough evidence of an oil pool for me without looking for any anticline and I made a paying development in what&apos; I was afterward told by geologists was probably a synclinal trough. If I remember correctly Dr. Orton&apos;s report on the lola gas pool (about 1898) he distinctly said that he found no evidence of anticlinal structure and concluded from the records of a large number of wells that there, was a terraced structure of the productive gas sands, of which there were no surface indications.

Jan 4, 1917

The cut-and-fill mining methods in use in the mines of The International Nickel Company of Canada Limited, at Falconbridge Nickel Mines Limited, in the Sudbury District, and at McIntyre Porcupine Mines Limited in the Porcupine District, are typical of the general cut-and-fill practice in both the base-metal mines and the gold mines in the Province of Ontario. The following pages describe the methods in use. HORIZONTAL CUT-AND-FILL STOPING Horizontal cut-and-fill stoping, as practiced in the mines of The International Nickel Company of Canada Limited, falls into two main classifications: (I) transverse cut-and-fill stoping, where ground conditions or wide ore bodies make it necessary to establish narrow stopes and pillars across the width of the ore body; (2) longitudinal cut-and-fill stoping where ore body widths and ground conditions allow safe mining of the full width of the ore body. Transverse Cut-and-fill StopiNg General Stoping Layouts Stope and Pillar Widhs.—For transverse stoping, experience has indicated that narrow stopes and pillars, 27 ft. 6 in and 22 ft., respectively, are most satisfactory. These width are multiples of 5 ft. 6 in., which is the center to center dimension of a standard square set. Thus any stope, or block of stopes, can be readily converted from cut-and-fill to square set-and-fill if ground conditions should warrant such a change. Formerly, pillars were established 16 ft. 6 in. wide, the equivalent of the width of three square sets, with the intention of extracting the pillars by mining an overhand slice one set wide and an underhand slice two sets wide. However, subsequent experience proved that a two-set overhand slice is so much more efficiently mined than a one-set slice, and 22-ft. pillars have been made standard. Chutes.—The stope chutes are established on the longitudinal center line of the stopes and are spaced at 22-ft. centers from footwall to hanging wall. The footwall chute is usually from 5 to 10 ft. from the footwall of the stope. It is offset where necessary at an angle of 52° and vertical chutes are carried from the offset at 22-ft. centers. In sections of the ore body where a flat hanging wall would cut off hanging-wall chutes within a few cuts of the sill, it has proved to be more economical to use temporary chutes or slushers, or both, for these first few cuts, rather than to el permanent chutes at the hanging wall.

Jan 1, 1946

By W. R. Jones

composition, he cannot reasonably asl; the manufacturer to guarantee that this composition shall give certain physical results. W. R. Jones, Pittsburgh, Pa. : The question that naturally occlun to me is this: Has Dr. Dudley in his inveatigtrtions been aiming to prove a theory, or has he been guided by an earnest desire to discover what are the proper elements in the composition of a good-wearing steel rail ? Unfortunately, for correct chemical information, he has omitted in his analgw two very importnat elements,—sulpbur and copper. Now, before we will even begin to admit the correctnets of Dr. Dudley&apos;s conclusions and the formula he prescribes, we will at the start question the propriety of any chenlist or scientist prescribing a formula for making steel when he has iguared such important ele ments as sulphur aud copper. I, for one, will not accept any such formula. Are we sure, or is Dr. Daclley sure, that the chemical analyseq embodied in his paper are correct? This may seem a presumptuous queetion, yet, with my experience with chenlists, I naturally doubt the correctness of the analyse8, and, before I will accept them as correct, I will ask that comparative tests of phosphorous and man ganese be made by the Pennsylvania Railroad chemists and the chemists of the leading steel-works in the country. Let us first verify the correctness of the analyses before we consider the con clusions. I can enumerate a great number of instances in which chemists have differed very widely in their determinations of phos phorus and manganese. A prominent iron fitam made a contract with the Edgar Thomson Steel Company to deliver pig metal guar anteed to be between 0.07 and 0.08 phosphorus; an analysis by our chemist resulted in phosphorus 0.148 and 0.152,—a rather startling difference! Again, a sample bar of steel, in which our chemist reported phosphorus 0.11, was tested by a chemist of au other Bessemer works, ana his determinations were phosphorus between 0.07 and 0.08: A leading engineering establishmeut of Pittsbdrgh bought iron claimed by a chemical analysis to contain 0.08 phosphorus; our Mr. Ford found phosphorus 0.145. A chem ist connected with an open-hearth works reported manganese in a piece of steel, 1.14; in a second determination from the same piece of steel, by the same chemist, manganese was reported 0.43. The chemist was kept in ignorauce of the fact that both samples were from the saw piece of steel. Two dctermiuations for manganese

Jan 1, 1881

By C. W. Hamilton

In the production of crude petroleum for the year 1930 Venezuela again maintained its world rank of second only to the United States. During the latter months of the year Russia forged ahead of Venezuela in daily rate of production, and from the present outlook will probably rank second among the oil-producing countries of the world in 1931. It is now believed Venezuelan production has passed its peak. The producing fields of the Bolivar coastal district, along the east side of Lake Maracaibo, continue to be the principal source of supply. Such fields as Lagunillas, Mene Grande and La Pas were not produced to capacity throughout 1930. During the year, three new fields entered into commercial development: The Rio Tarra field in the District of Colon, State of Zulia; the El Mene de Salto field in the District of Acosta, State of Falcon and the Quiriquire field, District of Piar, State of Monagas. For 1930 the total crude production of Venezuela was approximately 137,424,000 bbl. as compared to 137,739,000 bbl. during 1929, or a decrease of 315,000 bbl. During the past year exports of crude from Venezuela amounted to 137,8.50,000 bbl. as against 130,044,000 bbl. exported during 1929. Crude oil in field storage decreased from 14,138,-000 bbl. at the beginning of 1930 to 9,600,000 bbl. as of Jan. 1, 1931— present crude stocks fill about one half the field tankage available. Drilling operations in the producing fields declined sharply but throughout the year there was maintained a satisfactory potential reserve in the shape of proven territory and closed-in producers. Of the 2555 wells which have been completed in Venezuela, 1571 wells were in production on Jan. 1, 1931. The status of these present producing wells may be classified as follows: flowing 652, pumping 170, gas-air lift 193 and closed-in 556. The ratio of closed-in wells to producing wells declined in 1930, while the number of wells produced by artificial means increased materially. In the Bolivar coastal fields there is now observed a marked decline in gas volume and well-head pressure. Lago Petroleum Corpn. and Venezuela Gulf Oil Co. each have under construction electric plants designed to generate power for pumping wells by electricity. The first units of these plants will be placed in conlmission in the early spring of 1931. Venezuela Oil Concessions, Ltd. has installed a number of stenrn-

Jan 1, 1931

By R. S. Shoemaker, F. W. McQuiston

RECOVERY METHODS AS RELATED TO PROPERTIES OF SILVER ORES The earliest metallurgical process for treating silver ores was amalgamation with mercury which was in use in the early 1500&apos;s. Closely following was the development of the Patio process for treating ores at Pachuca, Mexico. Silver often occurs as the native metal and in deposits associated with other metals such as gold, copper, lead and zinc. Principal silver minerals include compounds of sulfur, antimony, arsenic, and copper. Silver chloride, and argentiferous galena are also prime sources. Native silver and the chloride characterize the oxide zone of most deposits. Native silver can be concentrated, amalgamated, or cyanided. Silver chloride rapidly dissolves in cyanide without oxygen. Oxidized silver ores containing the higher oxides of manganese are generally refractory to metallurgical treatment. A refractory compound of manganese and silver is formed, probably a manganite , which is insoluble in cyanide solutions. Argentite is the predominant silver mineral. Other important economic minerals are native silver, argentiferous galena, cerargyrite, pyrargyrite and tetrahedrite. Argentite dissolves slowly, the reaction being reversible requires an excess of cyanide. Argentite with disseminated fine grained gold occurring in quartz veins with minor amounts of chalcopyrite and galena may be cyanided directly after grinding, with high silver and gold recoveries. The flowsheet adapted to this type of ore is the Minas de San Luis, Tayoltita cyanide plant, which treats a high grade argentite silver ore containing gold. It is a standard cyanidation circuit but employs long periods of agitation contact for the high silver sulfide content ore. Lead-zinc base metal semi-oxidized- complex ores in pipes and chimneys are replacement deposits in limestone at El Mochito Mine, Honduras. Mineralization comprises galena, cerussite, anglesite, sphalerite, calamine, and smithsonite. The silver content of about 16 ounces occurs as native silver globs, wire and the mineral argentite. After flotation of lead and zinc concentrates, the tailings are cyanided for gold and silver recovery. The great Huelva pyritic copper deposits of Spain carry a little gold and silver. At Cerro Colorado, these metals were concentrated above the chalcocite zone into the so-labeled Mgossanu. This material with a content of 0.08-0.09 oz Au and 1.50 oz Ag per ton is amenable to cyanidation. See Cerro Colorado flowsheet. In southern Hidalgo, about 60 miles north of Mexico City, are the great silver--gold veins of one of the foremost silver mining districts of the world. Real del Monte, Pachuca. The mineralized area consists of flow rocks and intrusives. The oxidized ores carry pyrites with oxides of iron and manganese and other minerals, silver and gold. The first zone contains

Jan 1, 1975

By A. F. Brock

The cut-and-fill mining methods in use in the mines of The International Nickel Company of Canada Limited, at Falconbridge Nickel Mines Limited, in the Sudbury District, and at McIntyre Porcupine Mines Limited in the Porcupine District, are typical of the general cut-and-fill practice in both the base-metal mines and the gold mines in the Province of Ontario. The following pages describe the methods in use. HORIZONTAL CUT-AND-FILL STOPING Horizontal cut-and-fill stoping, as practiced in the mines of The International Nickel Company of Canada Limited, falls into two main classifications: (I) transverse cut-and-fill stoping, where ground conditions or wide ore bodies make it necessary to establish narrow stopes and pillars across the width of the ore body; (2) longitudinal cut-and-fill stoping where ore body widths and ground conditions allow safe mining of the full width of the ore body. Transverse Cut-and-fill StopiNg General Stoping Layouts Stope and Pillar Widths.—For transverse stoping, experience has indicated that narrow stopes and pillars, 27 ft. 6 in and 22 ft., respectively, are most satisfactory. These width are multiples of 5 ft. 6 in., which is the center to center dimension of a standard square set. Thus any stope, or block of stopes, can be readily converted from cut-and-fill to square set-and-fill if ground conditions should warrant such a change. Formerly, pillars were established 16 ft. 6 in. wide, the equivalent of the width of three square sets, with the intention of extracting the pillars by mining an overhand slice one set wide and an underhand slice two sets wide. However, subsequent experience proved that a two-set overhand slice is so much more efficiently mined than a one-set slice, and 22-ft. pillars have been made standard. Chutes.—The stope chutes are established on the longitudinal center line of the stopes and are spaced at 22-ft. centers from footwall to hanging wall. The footwall chute is usually from 5 to 10 it. from the footwall of the stope. It is offset where necessary at an angle of 52° and vertical chutes are carried from the offset at 22-ft. centers. In sections of the ore body where a flat hanging wall would cut off hanging-wall chutes within a few cuts of the sill, it has proved to be more economical to use temporary chutes or slushers, or both, for these first few cuts, rather than to el permanent chutes at the hanging wall.

Jan 1, 1946

By C. C. McDermond, Others, B. N. Zavoico

Production of crude oil in Venezuela for the year 1937 was 186,653,-916 bbl., as compared with 155,228,982 bbl. produced in 1936, an increase of almost 31,500,000 bbl. The oil-field strike, which started on Dec. 14, 1936, lasted until Jan. 25, 1937, when a total of 10,000,000 bbl. of production was estimated to have been lost (5,750,000 bbl. in 1936 and the remainder in 1937). Of the total of 186,653,916 bbl. produced in 1937, the Lagunillas field contributed 45.6 per cent, a drop of approximately 10 per cent under 1936 figures; Tia Juana, 14.3 per cent; Quiriquire field, 14.0 per cent; La Rosa area, 13.3 per cent and the Mene Grande field, 7.6 per cent, these being the five leading contributors of the year&apos;s production. Detailed production figures by fields and by companies are shown on Table 1. Of the total production of Venezuela the following four major companies accounted for the greater proportion: Standard of Venezuela and Lago Petroleum Corporation together, 48.3 per cent; Shell, 37.4 per cent and Venezuela Gulf, 14.0 per cent. At the end of the year, 2491 wells were producing, of which number 1291 wells were pumpers, or about 52 per cent. Production The accompanying statistics of production record of Venezuela oil fields from its inception to Dec. 31, 1937, shown on Table 2, indicates that the cumulative production reached 1,503,283,228 bbl. In order of importance, Lagunillas leads all the fields, having accounted for about 50 per cent of Venezuelan production, followed by La Rosa area, which produced about 21 per cent; Mene Grande, which produced about 13 per cent, and Quiriquire, which produced 6.5 per cent. Table 3 shows the age of each field. Mene Grande is the oldest, having produced for 24 years; Tarra is next, 21 1/2 years old, and El Mene third, 171/2 years. The gravity of the crude from the various fields, also shown on Table 3, remained unchanged, the only addition being the new Tia Juana field, which averages 15.6&apos; A.P.I. Table 4 shows the method of production by fields as of Jan. 1, 1938; at the end of the year 2491 wells were producing, compared with 2005 wells producing at the beginning of 1937, or an increase of 486 wells. The completions for the year reached 512. Of the 2491 wells producing crude oil, 1291 wells, or approximately 52 per cent, were pumping, 925 (approximately 37 per cent) were flowing naturally and 275 (11 per cent) were on

Jan 1, 1938

By A. C. Heath, E. H. Dix

The importance of aluminum-silicon alloys in thc light alloy field is now generally recognized. Where silicon was once considered detrimental to the properties of aluminum, useful alloys now contain as much as 13 per cent. Furthermore, silicon is present as an impurity in aluminum of even the purest grades. Iron is another important impurity in aluminum and likc silicon is found in small quantities in the purest aluminum available today. The manner of occurrence of iron and silicon in commercially pure aluminum and many of its alloys exerts a very pronounced influence upon their properties. The effect of large quantities of iron in aluminum-silicon alloys is so detrimental that it has been found necessary to limit the allowable maximum percentage to as low a figure as is consistent with quantity production and satisfactory physical properties. The intimate relation existing between iron and silicon in aluminurn and alurninum alloys makes necessary a more complete knowledge of this ternary system. This paper is divided into two principal parts, dealing respectively with aluminum-silicon solubility equilibria and the constitution of aluminum-iron-silicon alloys. Part I.—Aluminum-silicon Alloys The early work on this system was largely confined to determining the general form of thc diagram and in this respect the results are in good agreement. Briefly, alulninunl and silicon form a simple eutectifcrous series without compounds. Until recently the solid solubility of silicon in aluminum had not been accurately determined. Ipraenkell (1908) stated that aluminum will retain 0.5 per cent. silicon in solid solu-t&apos;ion. Rosenhain, Archbutt and Hanson2 gave the maximum solubility as about 1.50 per cent. and added that "this solubility docs not appear