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By H. W. Weart, D. J. Mack

SINCE the first recognition of eutectic solidification as the simultaneous formation of 2 solids from one liquid,1 many complex structures thus produced have been observed. Despite many attempts to ascertain it, however, the detailed mechanism by which these structures develop remains obscure. In an attempt to clarify the origin of eutectic structures, the study described in this paper was undertaken. The broad objective of the study was to investigate the effect of different solidification conditions on the structure of single eutectic grains. It was expected that the eutectic structure would thereby be isolated for study, but preliminary experiments revealed another structure within each grain. When a literature surveye revealed only 1 previous mention of the subgrain structure,' it was decided to determine the origin of the structure and to study its response to various growth conditions. Before presenting the details of the investigation, it is apropos to distinguish carefully among the 3 structures produced during eutectic solidification. This is especially desirable in view of the apparent confusion existing in the literature concerning the structures in a eutectic specimen. The confusion is due largely to the existence of the subgrain structure noted above which is easily mistaken for the grain structure. In the following paragraphs, an effort is made to distinguish the 3 structures by briefly describing their origin during solidification, even though the results are thereby anticipated slightly. The 3 structures produced during eutectic solidification are the grain structure, the colony structure, and the eutectic structure, each one contained by the one preceding it. The grain structure is analogous to the polycrystalline structure of a metal or inle-phase alloy in that each grain grows from a single nucleus. A somewhat broader interpretation is needed here, however, because of the presence of 2 phases in each grain. Since in most eutec-tics particles of 1 phase are dispersed in a matrix of the other, a grain may be taken to be the region in which the matrix phase is monocrystalline.*

Jan 1, 1959

By Bailey Willis, George Otis Smith

I. Summary of General Features. Location.—The Clealum iron-ore deposits occur on Clealum river, a tributary of the Yakima river; in the eastern spurs of the Cascade range, Washington.† Mount Stuart, 9470 feet in altitude, is a dominant peak of the district, which will therefore be referred to as the Mount Stuart district. By way of the valley and cañon of the Clealum, the ores are 18 miles from Roslyn, the colliery-town, which is reached by a short branch from the Northern Pacific Railroad. Where they occur the valley has an elevation of 2900 to 8000 feet above sea; that of the end of the branch railroad at Ronald is 2345 feet. The grade of the valley from near Ronald to the junction of the Clealum and Salmon Lasac rivers, 144 miles, is 20 feet per mile; thence to the iron-ores it is 130 feet per mile, through a cañon 241/2 miles long. Abreast of the deposits the western wall of the valley is formed by precipices of sandstone; the eastern slope is comparatively gentle, being cut in easily eroded

Jan 1, 1901

A. L. FIELD, Cleveland, Ohio (written discussion*.-)In equation 2, B is used to denote the ratio of bases (lime plus magnesia ) to acids (alumina plus silica) it being stated that this ratio gives more concordant results than the ratio either of bases to silica, or of lime to silica. I find, however, that, by making use of the ratio between true CaO (CaO by analysis minus sulfur calculated to CaS) to acids, equation 2 may he modified to agree with the observed facts within an average error of 0.0041 per cent. sulfur instead of 0.0045 per cent. The revised equation is s = 0.072 + 0.01835 - 0.00061 (t - 1400) - 0.0403B where B is the ratio of true CaO to Al2O3 Plus Si02, and the other symbols have the same significance as before. I would not appear to stress too much this slightly better agreement of the revised equation with the observed sulfur. Yet I believe that the use of this ratio of true CaO to acids is to be preferred, if for no other reason than the fact that it is more in harmony with the previous findings of the Bureau of Mines with regard to the effect of magnesia and other impurities on slag viscosity.1

Jan 12, 1919

By Van Dyne Howbert, Fred E. Gray

The Presidio mine of The American Metal Co. of Texas is situated 45 miles south of Marfa, Texas, a town on the Southern Pacific R. R., and lies approximately 20 miles north of the Mexican border. Communication with the railroad is maintained over a state highway. All supplies for mine and mill are hauled in by truck from Marfa and precipitates and concentrates are freighted to the railroad by the same means and shipped to Selby and El Paso, respectively. The mill is at the town of Shafter, 1 mile from the mine and at approximately 350 ft. lower elevation. The ore is conveyed to it by an aerial tramway which requires practically no power on account of the difference in elevation. The mill is a countercurrent decantation cyanide and gravity concentrating plant of 200 tons daily capacity. Coarse and fine crushing are done at the mill. Two neal-by shallow wells in the bed of Cibolo Creek furnish an abundance of water. Power is generated for the mill and town by a Diesel engine plant. Ore Treated The ore is oxidized and siliceous. The greater part contains approximately 16 oz. of silver per ton, but mixed with this are many particles of ore of much higher grade, assaying in many instances 100 oz. or more. In the 16-oz. material there is an almost constant amount of silver which is difficult to recover by cyanidation. The high-grade material contains a much greater proportion of silver which is easily amenable to cyanide treatment. Thus, as the grade of ore becomes higher, the percentage of recovery of silver increases rapidly, under equal milling conditions. The silver minerals contained in the ore are principally argentite and cerargyrite, the former predominating The lead minerals, all of which are argentiferous, are chiefly cerussite and galena, with occasionally a little anglesite. The gold is free; most of the ore contains merely a trace

Jan 1, 1935

By Kenneth K. Humphreys, Joseph W. Leonard, Robert L. Llewellyn, William F. Lawrence

INTRODUCTION Cleaning fine coal sizes utilizing air currents in machines as the primary separating medium is called dry concentration or pneumatic cleaning. In 1947 approximately 18 million tons (16.38 million mt) of bituminous coal were cleaned by air machines; in 1965, 25.4 million tons (23.1 14 million mt); but in 1975, only 7 million tons (6.37 million mt). Table 11-1 shows the tonnages cleaned of bituminous coal and lignite in the United States by various types of equipment from 1938-1963. Table 11-2 shows only tonnages cleaned by pneumatic methods from 1964-1975. The decrease in tonnage to approximately 7 million tons (6.3 million mt) is due to a higher moisture content in the mined coal and the greater difficulty in cleaning the coal by air. There also was a definite trend to load raw coal at the mines, especially the slack which is the most expensive to clean due to the dewatering and drying requirements. While there have not been any applications of pneumatic cleaning in the anthracite field, its use has not been fully explored. The same is true for lignite and subbituminous coals.

Jan 1, 1979

By Frank R. Van Horn

It is customary to ascribe two general modes of occurrence to natural gas, namely, shale gas which, as the name indicates, is found in shale, and reservoir gas, which occurs in sandstone, conglomerate and dolomitic limestone. Both of these types are found in the Cleveland district. The shale gas has been produced on a small scale for a long period of years, probably dating back as far as 1883. Shale gas is a low-pressure gas, small in volume with variable pressure, and is limited to no definite horizon. It is generally independent of oil, may last for a long time, and is not dependent to any great extent on the structure of the rocks. There are many such wells in the district which range from 400 to 1,840 ft. in depth, although most have been completed at about 800 ft. This indicates that the wells pass through the Cleveland and Chagrin shales into the Huron or Portage shale of the Upper Devonian. During the last 2 years great interest has developed in a deeper-seated reservoir gas. What is commonly called reservoir gas is generally a high-pressure gas occurring in large volume at a definite horizon. Oil may, or may not, be associated with the gas, and in most regions in which reservoir gas is found, the geological structure is of great importance. This type of gas occurs in all large fields such as those of West Virginia which have been supplying Cleveland until the recent discovery of local gas at great depths in commercial amounts. Historical Outline The oldest well producing reservoir gas in the district was drilled in 1886 in Newburg for what was then known as the Cleveland Rolling Mill Co. This well was 3,000 ft. deep, and was thought to have reached the "Clinton" formation. Two "sands" showing small amounts of oil were recorded in the Niagara limestone at depths of 2,658 and 2,686 ft. This is about the horizon of the so-called Newburg sand which has proved to be of some importance as a source of gas and oil by recent drilling operations. The rock pressure was reported to be 400 to 500 lb. per square inch, but since the volume was only 14,000 to 16,000 cu. ft. daily, the well was not only considered unsuccessful but was thought to prove the absence of high-pressure gas in commercial quantities in the Cleveland district.

Jan 1, 1917

By Roswell H. Johnson

The principle of sliding or graduation in royalties is accomplished either by the block, period, cumulative, or class method. The block method calls for a very low royalty rate on all oil produced up to a certain amount. This amount would normally be that which is believed to suffice to repay the cost of the well. A higher rate would be charged after the well is making a profit. This method, common in many industries, is peculiarly inadaptable to any industry such as this where the productivity per day of the units gradually and inevitably decreases. This, for reasons given later, is because it forces an early abandonment of the well and so offends against conservation. In the period method, the royalty rate is changed by some definite amount when the well produces less than a specified quantity per day or other given unit of time. While simple to explain, it operates badly because a great deal is made to depend upon a small reduction in production, whereas the reduction as a matter of fact declines rather erratically owing to the exigency of lease management, connecting of tankage, and vagaries of the gaugers. These disadvantages are obviated by the cumulative method. In this method, all the production less than a certain amount per week has one rate of royalty, or none at all, and the oil produced in excess of that amount pays an additional rate of royalty. In the long run the calculation is probably no greater than with the period method, for simple tables will be published and printed in the oil papers that can be cut out, enabling quick readings to be made. Even without the table the calculation is easy with the few royalty rates, two or three, which would be used. In this method, the amount paid in royalty declines gradually, rather than suddenly as in the period method. The following reasons might be advanced for the adoption oi cumulative royalties: 1. To distribute the burden in accordance with the capacity to pay. This principle is a unique one in business, although familiar in taxation. This is because it does not distribute reward in proportion to efficiency.

Jan 1, 1916

By E. G. Bailey

PLANTS that normally burn coal now able too obtain a substantial increase over their normal supply for their greater power needs, and also additional tonnage for extra storage against the uncertainties of the future. Many plants previously burning oil or gas have already been converted to coal, and those which must yet change to coal will undoubtedly be able to obtain ample tonnage of coal of standard quality at reasonable prices. What a contrast this is to the conditions of 25 years ago, when the shortage of coal and transportation led to prices of $5 to $6 per ton at the mines, and degradation of quality, until the already overburdened railroads were called upon to haul 100,000 ton per day of extra ash alone. This was the result of a runaway market and the loading of gob, slate, and sub-quality coal amounting to an estimated increase in ash of 5 per cent of the total tonnage hauled. The problem of getting increased steam output from inferior coal caused unbelievable reductions in the low normal boiler efficiency of that time and untold labor in firing the hand- and stoker-fed furnaces of that day.

Jan 1, 1942

The Jeffrey operation of Canadian Johns -Manville Co. , Ltd., at Asbestos, Quebec, is considered to be the world's largest asbestos mine. It accounts for approximately 37% of Canadian production of chysotile asbestos and about 13% of world production. Rated capacity is over 600,000 stpy (540,000 mtpy) of fiber, of which well over 90% is exported to the United States and seventy other countries. During the early 1970's the company completed a five-year modernization and expansion program which included the installation of a unique 72 in. by 93 in. gyratory crusher to handle the extremely blocky serpentinized peridotite ore. Because of the large diameter of this crusher the top ring of the shell was made in two halves which were bolted together. Hundred-ton trucks dump from two sides of the crusher into the 200 st (180 mt) dump pocket which is equipped with dust collecting ducts in the walls of the other two sides. The installation was designed to accommodate the eventual use of 200 -ton trucks. Crushed ore at a. size of 7 ½ in. (191 mm) falls into a uniquely designed crushed ore pocket equipped with three sloping steel sides and one vertical side, which make it totally live. Infrared heaters installed in back of the hopper sides prevent the ore from sticking to the steel. Another interesting feature is an eccentric removal bridge which is permanently mounted inside the crushed ore pocket. The two halves of the bridge are let down by electric winches to form a platform on which the eccentric cart travels and from where maintenance is done. Level indications in the crushed ore pocket are by radioactive indicators. Under the crushed ore pocket is a 72 in. by 28 ft (1.83 m by 8.5 m) apron feeder which operates at 7. 2 to 50 fpm (2. 3 to 15. 3 mpm). Crushed ore is discharged to a picking belt for tramp removal. A crossbelt magnetic separator is also mounted over this belt. Permanently mounted belt reels are also situated over the picking belt to facilitate the changing of this unit. A 60 in. (1. 52 m) wide takeaway belt then elevates the crushed ore to t.he surface facilities. Instead of discharging to a crushed ore stockpile, the takeaway conveyor delivers the ore to a 300-ton distribution bin. The ore is then screened and secondary crushed; and the plus ¾ in. (19 m) is "stored in a 3,000 ton bin while the minus ¾ in. is dried and then sent to storage. This system of directly screening primary crushed ore is similar to that used at both the Endako molybdenum and Gibraltar copper operations in British Columbia. While it would seem connecting processing equipment directly to a primary crusher would have an adverse effect on the crusher availability, the 94% of scheduled time this crusher

Jan 1, 1978

By G. P. Boomslitter

In previous discussions on stresses in hoisting ropes, little has been said concerning the effect of the elasticity of the rope itself on the stresses due to acceleration. Laschinger1 has calculated acceleration stresses for several cases but he has ignored the elasticity of the rope as a factor, except as it compresses the drum and stretches itself as it winds thereon. He also deals with steam plants and he shows the effect of the "jerky" start of the engine. This effect would actually be reduced by the elasticity of the cable. With the advent of the motor-driven hoist, new conditions have arisen. The jerk mentioned by Laschinger has been eliminated and smoother starting made possible. But the elastic properties of the rope seem to introduce considerable periodic acceleration stresses in the rope. The following discussion of acceleration stresses is based on the assumption of a constant acceleration at the drum. To be exact, the mathematical work is based upon the assumption that the top is moving upward with a constant acceleration, and ignores the fact that the cable is shortening as it winds. However, the effect of the shortening would be inconsiderable during the first part of the acceleration period. Most hoisting engineers make the assumption of constant acceleration at the drum. Under the older method of calculating acceleration stresses, this would be true with a constant torque motor. But if stresses vary, as this paper attempts to show that they do, constant torque would not give constant acceleration, although the velocity-time curves obtained in the full size test herein described showed that they approach rather closely to this condition, at least as nearly as could be measured. Theoretical Discussion According to elementary physics, the force required to accelerate a Wf body of weight W at the rate of f ft. per sec.2 is wf . To raise the body Wf with acceleration of f ft. per see.2 the total force required is I + Wf, in if

Jan 1, 1927

By Daniel R. Hull

In 1926 Gendersl reviewed the existing theories of this subject and stated his views in support of the gas-pressure theory. Again, in 1937, the subject was thoroughly reviewed by N. B. Vaughan3 and by Phillips and Brick,2 so that no résumé is called for here. The purpose of this paper is to describe certain observations made in the course of mill work, and to point out that they mostly tend to confirm the gas theory of Genders. Vapor Segregation There is, however, one type of inverse segregation that can be accounted for by considering the volatility of one of the constituents of an alloy. The most familiar case is ordinary brass, which is subject to a zinc enrichment of a few tenths of a per cent at the surface of a chill casting. This has been noted in 70-30 brass by Genders, and a typical example of it is given in Table 1, which represents a cylindrical billet 7 in. in diameter by 55 in. long. After lathe-turning 1/32 in. from the surface, the remaining distance to the center was sampled at six equally spaced intervals. the lead, which is essentially the same throughout, will be mentioned later. Consider a brass-mill mold, either round or rectangular in cross section, standing several feet high. It may require a minute or two for pouring. Before it is full, the metal in the lower end has partly solidified and contracted somewhat away from the mold. The initial shell of solidification follows not far behind the top of the column, even if the central portion is liquid for a long way down. There is, then, a space between metal and mold that is filled with zinc vapor. The vapor has no exit, because the liquid at the top of the rising column is not only in contact with the mold but keeps UP a fresh supply of zinc vapor, which has, perhaps, already begun to condense in the lower portion. The cycle is short. Within a minute or two after the mold is full the bar has cooled enough to condense all the zinc. Some of it may condense on the mold; it might be expected that most of it would condense there, but the mold at that time is well coated with a velvety carbon residue from the oil dressing, overlain by fluffy zinc oxide that commenced covering the inside of the mold as soon as pouring began. The clean skin of the freshly cast brass offers a preferred surface for condensation. Reabsorption is quick and inevitable. A simple experiment indicates that a surprisingly large quantity of zinc vapor is trapped in the space between mold and

Jan 1, 1941

By Daniel R. Hull

In 1926 Gendersl reviewed the existing theories of this subject and stated his views in support of the gas-pressure theory. Again, in 1937, the subject was thoroughly reviewed by N. B. Vaughan3 and by Phillips and Brick,2 so that no résumé is called for here. The purpose of this paper is to describe certain observations made in the course of mill work, and to point out that they mostly tend to confirm the gas theory of Genders. Vapor Segregation There is, however, one type of inverse segregation that can be accounted for by considering the volatility of one of the constituents of an alloy. The most familiar case is ordinary brass, which is subject to a zinc enrichment of a few tenths of a per cent at the surface of a chill casting. This has been noted in 70-30 brass by Genders, and a typical example of it is given in Table 1, which represents a cylindrical billet 7 in. in diameter by 55 in. long. After lathe-turning 1/32 in. from the surface, the remaining distance to the center was sampled at six equally spaced intervals. the lead, which is essentially the same throughout, will be mentioned later. Consider a brass-mill mold, either round or rectangular in cross section, standing several feet high. It may require a minute or two for pouring. Before it is full, the metal in the lower end has partly solidified and contracted somewhat away from the mold. The initial shell of solidification follows not far behind the top of the column, even if the central portion is liquid for a long way down. There is, then, a space between metal and mold that is filled with zinc vapor. The vapor has no exit, because the liquid at the top of the rising column is not only in contact with the mold but keeps UP a fresh supply of zinc vapor, which has, perhaps, already begun to condense in the lower portion. The cycle is short. Within a minute or two after the mold is full the bar has cooled enough to condense all the zinc. Some of it may condense on the mold; it might be expected that most of it would condense there, but the mold at that time is well coated with a velvety carbon residue from the oil dressing, overlain by fluffy zinc oxide that commenced covering the inside of the mold as soon as pouring began. The clean skin of the freshly cast brass offers a preferred surface for condensation. Reabsorption is quick and inevitable. A simple experiment indicates that a surprisingly large quantity of zinc vapor is trapped in the space between mold and

Jan 1, 1941

By M. B. Holt

At this time, when electricity in its various applications is attract- ing so much interest, and with such good reason, it has been suggested to me that, as the Aspen Mining and Smelting Company, of Aspen, Pitkin county, Colorado, was the first mining company in America to employ electrical power for hoisting, a record of what has been accomplished by this company would be of interest. Early in the year 1888 the question of power for the development of the ore-horizon lying below the level of the tunnel through which the property is worked, became one of importance. The conditions under which it was proposed to use power were as follows (See Fig. 1): The company's mines lie upon the northwestern slope of a mountain, the ascent of which is at an angle of 28' from the horizon. The ore, consisting of argentiferous galena and other silver-bearing mineral, is found at or near the plane of contact of a superincumbent mass of blue limestone, with an underlying bed of dolomite. This contact plane dips north 30° west, at an angle of 60' with the horizon. A tunnel 1000 feet in length, driven south into the mountain near

Jan 1, 1892

By K. H. Coats

This paper presents the development and solution of a mathematical model for aquifer water movement about bottom-water-drive reservoirs. Pressure gradients in the vertical direction due to router flow are taken into account. A vertical permeability equal to a fraction of the horizontal permeability is also included in the model. The solution is given in the form of a dimensionless pressure-drop quantity tabulated as a function of dimensionless time. This quantity can be used in given equations to compute reservoir pressure from a known water-influx rate, to predict water-influx rate (or cumulative amount) from a reservoir-pressure schedule or to predict gas reservoir pressure and pore-volume performance from a given gas-in-place schedule. The model is applied in example problems to gas-storage reservoirs, and the difference between reservoir performances predicted by the thick sand model of this paper and the horizontal, radial-flow model is shown to be appreciable. INTRODUCTION The calculation of aquifer water movement into or out of oil and gas reservoirs situated on aquifers is important in pressure maintenance studies, material-balance and well-flooding calculations. In gas storage operations, a knowledge of the water movemenr is especially important in predicting pressure and pore-volume behavior. Throughout this paper the term "pore volume" denotes volume occupied by the reservoir fluid, while the term "flow model" refers to the idealized or mathematical representation of water flow in the reservoir-aquifer system. The prediction of water movement requires selection of a flow model for the reservoir-aquifer system. A physically reasonable flow model treated in detail to date is the radial-flow model considered by van Everdingen and Hurst.1 In many cases the reservoir is situated on top of the aquifer with a continuous horizontal interface between reservoir fluid and aquifer water and with a significant depth of aquifer underlying the reservoir. In these cases, bottom-water drive will occur, and a three-dimensional model accounting for the pressure gradient and water flow in the vertical direction should be employed. This paper treats such a model in detail — from the description of the model through formulation of the governing partial differential equation to solution of the equation and preparation of tables giving dimensionless pressure drop as a function of dimensionless time. The model rigorously accounts for the practical case of a vertical permeability equal to some fraction of the horizontal permeability. The pressure-drop values can be used in given equations to predict reservoir pressure from a known water-influx rate or to predict water-influx rate (or cumulative amount) when the reservoir pressure is known. The inclusion of gravity in this analysis is actually trivial since gravity has virtually no efFect on the flow of a homogeneous, slightly compressible fluid in a fixed-boundary system subject to the boundary conditions imposed in this study. Thus, if the acceleration of gravity is set equal to zero in the following equations, the final result is unchanged. The pressure distribution is altered by inclusion of gravity in the analysis, but only by the time-constant hydrostatic head. The equations developed are applied in an example case study to predict the pressure and pore-volume behavior of a gas storage reservoir. The prediction of reservoir performance based on the bottom-water-drive model is shown to differ significantly from that based on van Everdingen and Hurst's horizontal-flow model. DESCRIPTION OF FLOW MODEL The edge-water-drive flow model treated by van Everdingen and Hurst1 is shown in Fig. la. The aquifer thickness b is small in relation to reservoir radius rb. water invades or recedes from the field at the latter's edges, and only horizontal radial flow is considered as shown in Fig. 1b. The bottom-water-drive reservoir-aquifer system treated herein is sketched in Fig. 2a and 2b. Here the aquifer

By L. Guttman, C. S. Barrett, J. S. Bowles

THE transformation from the face-centered cubic (Al) to the face-centered tetragonal (A6) structure in certain alloys of the indium-thallium system reported in the preceding paper1 exhibits many interesting crystallographic and metallo-graphic features, most of which presumably are similar to those of other cubic to tetragonal transformation such as, for example, the one which occurs in the chromium-manganese system,2 the one which, judging by microstructures and X-ray pat-terns,3 probably occurs in the copper-manganese system, the ones accompanying ordering in FePt,4 Copt,&apos; and AuCu,5 and the transformation near 115°C in barium titanate.6,7,8 If, as mentioned in ref. 1, the transformation is second order, the phases in equilibrium do not differ in composition, and it is both possible and likely that the mechanism is a diffusionless one. The present paper reports the results of a detailed study of the indium-thallium transformation and presents a theory for the atomic movements of the transformation which accounts for the observations. Experimental Crystallography of Transformation Markings The transformation from cubic to tetragonal in the indium-thallium alloys produces microstruc-tures of the type illustrated in fig. 1. These structures are observable either as relief effects produced on a smooth surface by the transformation, or after polishing and etching. These microstruc-tures are markedly similar to those found in copper-manganese alloys: iron-platinum alloys," and barium titanate,7,8 where it has been demonstrated that the lamellae are parallel to the (101) planes of the original cubic crystal. That the observed lamellae in indium-thallium alloys are also parallel to the (101) planes was proved by the following analyses performed on single grains in polycrystal-line samples containing 20.75 atomic pct T1. A study of the microstructures of a number of specimens revealed the fact that in many specimens the transformation markings continued unchanged in direction across the (111) interfaces of annealing twins. This at once demonstrates that the lamellae are parallel to a plane that is common to both parent and twin orientations, i.e., a plane in the zone <112>. This conclusion is compatible with the lamellae being parallel to (101) planes. To make a more complete determination of the habit plane of the lamellae, two grains were selected in which annealing twins ({Ill) twins) had been present in the high-temperature modification (cubic). The traces of the annealing twins, and of the transformation markings in both parent and twin crystals, were plotted together in stereographic projection, and a parent-twin pair of orientations was found that accounted for all lamellae as (101) planes. The projection of one of these twinned grains is reproduced in fig. 2. In view of these results the probability of there being another solution for the habit plane remote from (101) is very small, but to explore the possibility that the habit plane deviates slightly from (101), a more precise analysis was undertaken. In this, the orientation of a grain in the cubic modification was determined from a Laue back-reflection photograph taken with the sample at 90°C. The traces of the lamellae that formed in this grain on cooling were plotted in the stereographic projection

Jan 1, 1951

By B. M. Larsen, L. S. Darken

SOME years ago we collated all laboratory data then available to us on the distribution at equilibrium of manganese and of sulphur between metal and simple slags, and used the results in setting up an equation, which proved to be in accord with determinations of the distribution ratio of manganese and of sulphur between liquid steel and open-hearth slag, over a wide range of operating conditions. We now present an outline of our calculations and inferences on these two closely related questions, which are of especial importance because of the advisability of conserving manganese wherever possible. In general, our results show that in the open-hearth furnace equilibrium between metal and slag, with respect to both manganese and sulphur, is substantially attained during the finishing period of the heat, or in about a half hour to an hour after any additions were made. This implies that the end result will be identical whether manganese is added as alloy or metal to the liquid steel or as oxide to the slag. The proportion of total manganese retained by the steel is favored: (I) by use of the minimum amount of lime, consistent with other operating limitations, required to give a lime-silica ratio of about 2.4, which usually corresponds to a minimum practical slag volume; (2) by a high concentration in the slag of MnO relative to FeO; (3) by high operating temperature. Absorption of sulphur by the slag from the liquid metal is favored: (I) by large slag volume; (2 by a high concentration of active (or "free") lime and of manganese oxide dis) solved in the slag; (3) by low concentration of silica, iron oxide, and phosphate dissolved in the slag. DISTRIBUTION OF MANGANESE BETWEEN SLAG AND METAL AT EQUILIBRIUM On the basis that the distribution of manganese between liquid steel and slag is controlled by the reaction [FeO + Mn = MnO + Fe [A]slag metal slag metal] the distribution at equilibrium, at a given temperature, is expressed by the equilibrium constant for this reaction, which is defined as [K" aMoO aFeI1 dFec) aMn] where a is the activity of the molecular species denoted by the subscript. The activity of each of the several components, although it is not known precisely, is approximately equal to the concentration of each component expressed as its mol fraction; that is, as the ratio of the number of mols of the given component to the total number of mols of all the molecular species present in the metal or slag, as the case may be. Moreover, since the activity, or concentration, of iron in the liquid metal is in fact substantially constant, the term [afe] may be combined with K" to give a new constant K&apos;. Making these changes,

Jan 1, 1942

By Zay Jeffries

IT has been known for a number of years that certain aluminum alloys could be hardened by quenching from a temperature of about 500° C. Immediately after quenching the total increase in hardness is not marked, but the hardness increases gradually so that after about 4 days, it has reached a maximum. The actual increase in hardness probably continues for weeks or even months, but in order to detect the differences after 4 days, it is necessary to make tests at rather long intervals. So far as is known to the authors, Alfred Wilm was the first to discover these properties in aluminum alloys.1 The aluminum-copper alloys are amenable to this treatment and the addition of magnesium makes the heat-treatment effect more pronounced. Until recently heat treatment of aluminum alloys has been confined to the worked alloys in the form of sheet or extruded products. The alloy that has been most used in this connection is known commercially as duralumin and has about the following composition: copper not exceeding, 6 per cent.; magnesium not exceeding, 2 per cent.; manganese not exceeding, 1 per cent.; commercial aluminum,2 balance. A typically good alloy in this range consists of: copper, 4 per cent.; magnesium, 0.5 per cent.; manganese, 1 per cent.; commercial aluminum, balance. This alloy can be rolled into sheets or extruded; and in this condition it can be heated in the ordinary furnace atmosphere to a temperature near 500° C. and quenched in water. After about 4 days it will have the following physical properties: tensile strength 55,000 to 60,000 lb. per sq. in. (3866 to 4218 kg. per sq. cm.); elongation in 2 in., 15 to 20 per cent. No harmful effects of the furnace gases or even the air, if the heating is affected in an electric furnace in an air atmosphere,

Jan 9, 1919

THE paper on "Specific Efficiency of the Blast Furnace" by Richard Franchot, which was pub-lished in the September issue of MINING AND METALLURGY, was presented for discussion at the Pitts-burgh meeting. The session opened with a discussion among S. P. Kinney, Mr. Franchot, and J. N. Reese as to the methods and results in sampling for cyanides in the blast furnace. This is covered in detail in writ-ten discussion below. T. T. Read objected to the use of the term specific efficiency as likely to be confusing. The blast furnace has to perform the function of reducing Fe, 03 to Fe as well as that of fusing the metal and, since the reduc-tion of Fe, 0? involves a CO equilibrium, it seems mis-leading to calculate efficiency on the basis of burning the , all carbon to CO? since it is not possible. Mr. Franchot replied that he used the term in the sense defined by Richards in his "Metallurgical Calculations;" but a subsequent search failed to reveal the term in Richards, who says (page 245, 1st edition): The practical conclusion is that a blast furnace generates in itself not much over 40 per cent of the calorific power of the fuel used, and rejects nearly 60 per cent; by using part of the waste gases to heat the blast, however, some of this rejected heat, to an amount representing net 5 to 10 per cent of the calorific power of the fuel used, is returned to and injected bodily into the furnace, thus rendering available for the purpose of running the furnace about 50 per cent of the calorific power of the fuel as a maximum. The efficiency with which the furnace applies this 50 per cent usefully to the objects of the reducing and smelting is another question for investigation.

Jan 2, 1927

By Folke H. Kihlstedt

Geophysical prospecting for ore has been more or less at a standstill during the present crisis owing to the lack of interest in base-metal exploration. A notable exception is the increased use of electrical and magnetic methods in prospecting for gold, the former in Canada, United States, Sumatra, Africa and Western Australia, and the latter on the Rand. This application to gold prospecting marks a forward step in ore prospecting because it is based upon mapping the geology instead of a direct location of the ore in question, as the amount of gold in an average gold ore is far too small to have any influence upon the physical properties of the rock in which it occurs. The development has thus gone in the same direction as geophysical prospecting for oil; that is, to find geological conditions that are genetically connected with the gold ore as exemplified by known occurrences. Naturally the success of this indirect prospecting depends upon two things: (1) the soundness of the geological conclusions upon which is based the hope of discovering new deposits, because a great many geological "set-ups" are barren, and (2) the ability of geophysical methods to help the geologist to collect the pertinent geological data within reasonable cost, after which favorable structures may be drilled or trenched. The first point is one to be considered by the mining geologist. In the vicinity of producing mines or promising prospects, the chances of finding ore in quartz veins and shear zones are great, and a survey giving the location of these features will materially guide both surface and underground exploration. Even in outside exploration the chances are sometimes favorable. As an example may be taken the Northeastern Ontario-Western Quebec gold zone. Large areas are here covered with glacial and post-glacial deposits and a simple study of available maps reveals the fact that less than 30 per cent of the area can ever be even preliminarily prospected by cominonly used prospecting methods, of which the pick and shovel are the main tools. Considering the number of deposits found, it is evident that many deposits remain hidden below the overburden. The knowledge of gold geology gives a starting point

Jan 1, 1934

By J. H. Brophy, A. L. Prill, H. W. Hayden

An error lzas been found in the conventional way of determining the time exponent of linear shrinkage for the solution-precipitation stage of liquid-phase sintering. Large amounts of shrinkage due to initial rearrangenzent result in erroneously low exponents and tni si dentification of the rate-controlling process. Two alternative, mathematically correct, methods of calculation are applied to existing data on Fe-Cu, WC-Co, and Tic-Ni-Mo. This results in reidentifi-cation of the rate-controlling process and somewhat better agreement with related data. In each case the solution-precipitation process is solution-controlled and not diffusion-controlled. THE liquid-phase sintering process has been the object of a considerable amount of research for many years. Linear shrinkage vs time data at various sintering temperatures have been used recently in an attempt to identify the prevailing mass-transport processes in a number of metallic and nonmetallic systems. There is an analytical error inherent in the methods previously used for handling data during the "solution-precipitation" stage of shrinkage which must be recognized before the correct rate-controlling processes can be determined. It is the purpose of this paper to show how the error has affected previous conclusions and to suggest corrections. A number of mechanisms have been proposed to explain experimental observations during sintering in the presence of a liquid phase. The earliest of these is due to Price, Smithells, and williamsl who attributed the observed growth of the solid phase to the simultaneous solution of smaller particles and growth of larger particles. From the work of Lenel et al.2,3 and Gurland and Norton4 has come the suggestion that the liquid-phase sintering process may be divided for analysis into three steps: 1) rearrangement upon formation of the liquid phase; 2) solution and reprecipitation of the solid phase in the liquid; and 3) "coalescence" or a final stage of sintering, characteristically slower than the solution-precipitation process. The sequence of steps is shown schematically in Fig. 1. Kingery5 presented a thermodynamic and kinetic analysis of the first two "stages" of the liquid-phase sintering process in which the functional dependence of lineal shrink- age on time at temperature was used to identify the rate-controlling mass-transport process. Table I summarizes the time dependences and processes considered by Kingery.5 This method of data presentation has been employed by Kingery and coworkers for Fe-CU,6 WC-Co, and a number of nonmetallic systems," and bv Whalen and Humenik for Tic-Ni-Mo and Fe-cu.8 In each of these studies, the rate-controlling step during the "solution-precipitation" stage was identified from the slope of the curve in that stage on a logarithmic plot such as Fig. 1. However, in all of the previous studies, at a certain composition, relatively large amounts of shrinkage due to rearrangement preceded the solution-precipitation stage. Under these conditions, the larger the amount of rearrangement shrinkage the

Jan 1, 1965