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By L. F. Mondolfo

R301 is a clad aluminum alloy, composed of a core of a duralumin-type alloy clad with a magnesium silicide alloy. It differs from other well-known clad alloys in that the cladding and the core respond to the same heat-treatments. and therefore the cladding contributes appreciably to the mechanical properties of the material. In Table I are reported the compositions of the core and cladding alloy for the material used for the experiments. Table i.—Chemical composition of R~OI Used for the Experiments Element Core Cladding Cu........................ . 0.01 Fe......................... . 0.42 Si......................... 0. 0.71 Mg........................ . 0.99 Mn ........................ . 0.54 In clad aluminum alloys where the core contains copper and the cladding does not, at elevated temperatures the copper tends to diffuse from the core through the cladding. This fact is well known and many investigations have been conducted to determine the features of this diffusion. Most of those investigations cover materials clad with pure aluminum; the present investigation covers a material clad with an alloy. Since the cladding thickness on R~OI varies with the gauge, as shown in Table 2, several gauges were investigated. Material and Procedure he material used was sheet of the following gauges: o.o2~, 0.040, 0.064, 0.102, 0.170 Cut in pieces approximately 3 by 5 in., it was in the cold-rolled state and was soaked for increasing times at the following temperatures: glo°F., 940°F., 970°F. A Lindberg cyclone furnace was used for the experiments. The procedure was as follows: The furnace was started empty with controlling pyrometer set at the determined temperature. After the furnace had reached that temperature, it was run for some 20 to 30 min. empty, in order to allow the temperature inside to become stabilized. Then the first samples were introduced. A load thermocouple, with recording pyrometer, was connected with the sample of the heaviest gauge. About 10 min. was required for the material to reach temperature and the soaking time was counted from that point. After zoo min. soaking, the second group of samples was introduced without stopping the furnace. After 200 min. more, the third group of samples was introduced, and so on, until the last sample had been introduced. Every time the furnace was opened and new samples were added, a drop of temperature was registered by the thermocouple connected with the material

Jan 1, 1946

By Robert B. Sosman

THE chance that a Howe Memorial Lecturer will be able to refer back to a personal contact with the distinguished metallurgist for whom this lectureship is named grows steadily smaller. I did not have the pleasure of his personal acquaintance, but I did receive from him in 1916, quite unexpectedly, a letter commenting on a paper that I had recently published27 on "Types of Prismatic Structure in Igneous Rocks." The paper was written for geologists and petrologists, but it had an application, naturally, to the crystallization of iron and steel in a mold. The letter impressed me at the time, and still does, with the breadth of Professor Howe's interests and his perseverance in keeping up with the progress of any branch of knowledge that had metallurgical connotations. SCOPE OF THE LECTURE During several years preceding the date mentioned I had worked in collaboration with Arthur L. Day and E. T. Allen8 at the Geophysical Laboratory, Carnegie Institution of Washington, on the temperature scale as defined by the nitrogen thermometer at high temperatures. Later (1928-1947) I had the opportunity of helping to apply this experience to the pyrometric problems of the steel industry. Because of the range of its temperatures, the multiplicity of the methods used, and the variety of its purposes, pyrometry in the open-hearth process has always offered a fertile field for both research and development. I am therefore devoting an hour to the subject of temperatures in the basic open-hearth furnace, the furnace that makes most of our steel in the United States. I would direct your attention particularly to the choice of the first two words in the title of the Lecture. One sometimes hears or reads of the "temperature of such-and-such a furnace," as if it were a single, measurable, characteristic quantity. I need hardly emphasize to any one who has ever operated a furnace or a kiln that "temperature of the furnace" is a meaningless phrase. Even if we confine our attention to the space enclosed by the roof, walls, and hearth of an open-hearth furnace, it is doubtful whether any such space has ever been so uniform in temperature that the phrase would have a meaning. In any furnace, there are many different temperatures, capable of being measured at many different places, and actually being measured for several different purposes. These pyrometric purposes, methods, and results, if given in detail, would expand this lecture into a book on the basic open-hearth process. I must therefore confine myself to a statement of the range of temperatures as they have been observed in various parts of the furnace, indicating in a general way why they are measured, and mentioning the methods that are available without attempting to describe these

Jan 1, 1948

By Denis G. Maxwell

INTRODUCTION It is my intention in this paper to deal with gold, uranium, diamonds, platinum, manganese, chrome, vanadium and heavy mineral sands. These are the most important strategic minerals produced by the Republic of South Africa which are not covered in other sessions of this program. In each case I have high- lighted the statistics and peculiar advantages which combine to make South Africa a vital source of these minerals. Before proceeding to give individual attention to these minerals I believe it would be useful to define what I mean by 'strategic'. The Concise Oxford Dictionary defines strategic in the context of materials as 'essential for war'. However it is commonly used in a much broader sense than this (often, in fact, very loosely) and I prefer to define it as 'concerned with the acquisition and maintenance of power, whether economic, political or military.' A VITAL SOURCE In dealing with the individual minerals I have quoted statistics which are contained in Tables 1, 2 and 3. Table 1 clearly shows the absolute size of the South African mineral industry. However, it can also be used to demonstrate the importance of the industry to the South African economy if compared with the GNP in 1980 of about R60 billion. Table 4 illustrates clearly how important South Africa is as a supplier of these minerals to most of the important industrialized countries of the Western World. Gold If anyone had any doubts about the inclusion of gold in a list of strategic minerals I am sure that the above definition of 'strategic' will convince them that it certainly belongs there. Similarly no one is likely to have any doubt about the fact that South Africa is a vital source of supply. Tables 2 and 3 show that in 1980 we had 51% of the world's reserves and accounted for 55% of world production. The figures for the Western World are considerably higher. The only other major producer, of course, is Russia, with small but significant production in the Pacific Rim area coming from Australia, Canada, Latin America, Papua New Guinea, Philippines and the U.S. All South African mine gold production is shipped in bullion form containing about 88% gold and 9% silver to the Rand Refinery which is a modern refinery with large scale units capable of refining half a ton of bullion at a time. The Refinery is equipped to produce standard 'good delivery' gold as well as 9999 gold and 999 silver. The Refinery also produces the 22 karat blanks which are, used by the South African Mint to produce Kruger Rands. It goes without saying that the South African gold mining industry leads the world in all aspects of deep-level, narrow-reef mining technology. The industry's metallurgists, too, have a record of tenacious and continuing efforts to improve extraction to the level of the present finely honed efficient process used on all the modern mines. Uranium In 1980 South Africa had 14% of the uranium reserves of the Western World and accounted for 14% of production. In view of the paucity of data I am not in a position to estimate figures for the total world. All the other major sources of uranium in the Western World are situated around the Pacific Rim, with the U.S. and Canada already being major suppliers and accounting for 38% and 17% of Western World production in 1980. Australian production at the time was small but they have very large reserves and production is already rising rapidly. The U.S., Canada and Australia account respectively for 22%, 19% and 29% of the uranium reserves of the Western World. South Africa has been a major producer continuously for 30 years. Nearly all the uranium produced, amounting to about 115 000 tons up to the end of 1981, was a by-product or co-product of gold extraction. During that time the industry has frequently led the world in technological innovation, and has established a reputation as a reliable producer of a consistent, high-grade product. In the latter respect, it is helped by the fact that production is marketed by one company, Nuclear Fuels Corporation, which also blends, dries and calcines the product from the individual mines and samples and assays it before shipping. Diamonds Diamonds are the rock on which the South African mineral industry is founded. The discovery of diamonds in 1866 gave rise to the first major mineral industry in the country and the profits from diamond mining helped to finance the gold mining industry 20 years later. Although now overshadowed by gold, diamonds are still very important in the overall picture of mineral production and exports, as can be seen in Table 1. There are really three separate diamond markets - gem, natural industrial, and synthetic - and, to be meaningful, statistics should be provided separately. Unfortunately separate figures are not available. The figures in Tables 2 and 3 show that, for gem and natural industrial together, South Africa ranks third in the world in production and second in reserves. South Africa is a major producer of synthetics and probably ranks second in the world after the U.S. Recently, of course, Australia was the scene of a major diamond discovery and will soon become the only

Jan 1, 1982

By William H. Breeding

The remarks that follow are based substantially on experience covering 45 years, 80% of which has been in placer work, rather than on a review of available literature. Most commercial placers have been deposited by the action of water. The richer and more- difficult-to-mine placers are those in the headwater areas where gradients are steepest. The most lucrative placers are generally in inter- mediate areas where volumes are greater, fewer boulders are present, and gradients are from 3% to 1-1/2%. The higher volume, lower grade placers are in the lower reaches of river systems where gradients are lower. Where gold-bearing rivers have discharged into the sea, wave action can concentrate values on beaches, past and present. Most of the rich, readily accessible placers were mined by our forefathers. Current opportunities exist: (1) in remote areas where infrastructure has been absent in the past, or development has been prohibited by adverse ownership - political or commercial; (2) in deposits that could not be mined by equipment available to our forefathers; (3) in deposits unidentified by our forefathers; (4) where the-price-of-product/cost ratio is substantially better than in earlier years; or (5) a combination of those factors. When I entered the placer business in the late 1930s, and subsequently, a prevailing opinion believed that glacial deposits should be avoided as irregular in mineral content and composition, and unrewarding to explore and develop; yet an operator has been mining a fluvio-glacial deposit profitably for the past 17 years. Rich buried placer channels, of ten called paleo-channels were worked in the last century, generally by hand methods, and under conditions that would be unacceptable today. Exploration and mining equipment now available make some of these channels attractive targets. Well-known examples are in California and Australia. The formation of a commercial placer requires a source of valuable minerals. Above primary deposits, there may be eluvial deposits formed by the erosion of gangue minerals and the concentration "in situ" of valuable minerals. Down slope from these deposits are the hillside or colluvial deposits, and below them are the alluvial deposits of redeposited material. Most of the great placer fields of the world are the result of several generations of erosion and deposition. Well-known examples are in California and Colombia. Gold is a very resistant and malleable material, and gold placers may extend for 64 or 80 km (40 or 50 miles) along a river system. Platinum is less malleable, but is very resistant to disintegration. Diamonds are extremely hard, and (especially gem diamonds) may be found over great lengths of a river system. Cassiterite is less resistant to disintegration, and tin placers seldom extend over two miles without resupply from an additional source or sources of mineralizaton. Tungsten minerals are generally more friable, and within a few hundred yards of the source disintegrate to the point that they are uneconomical to recover. Rutile, ilmenite and zircon placers generally result from the weathering of massive deposits, and may be encountered over extensive areas; most are fine grained and durable. What does a geologist or mining engineer look for in placer exploration? The old adage to look for a mine near an existing mine is still valid. You need a source of valuable mineral. Then you require conditions for concentration, which means a satisfactory gradient and/or other conditions that will permit heavy minerals to settle. Nicely riffled gravel, often called a shingling of the bars, is conducive to placer formation. Coarser gravel is logically associated with coarser gold. Excessive clay and/or high stream velocities in narrow channels can carry gold far downstream and distribute it uncommercially over a large area. When material is extremely fine, in situ weathering and concentration become more important. Placers frequently occur distant from lode mines, and one must remember that in a larger watershed the exceptional floods that occur once in a hundred or a thousand years can move great quantities of material long distances. The carrying power of water is said to vary with the fifth or sixth power of its velocity. I am not ready to disagree with Waldemar Lindgren and accept that many commercial placers are substantially enriched by the chemical deposition of gold from solutions; however, I have seen crystalline gold in clayey material quite distant from known sources of primary gold that is dif-

Jan 1, 1985

By G. M. Willis, F. L. Hennessy

The oxygen pressure in equilibrium with silver and Ag2O-B2O3 melts has been measured between 800' and 900°C, to obtain the thermodynamic properties of the liquid. The compound Ag20. 4B20:1 appears to exist in the liquid, which shows marked heat content and entropy effects. A KNOWLEDGE of the thermodynamic properties of binary liquid silicates, borates, and phosphates would be of considerable assistance in the interpretation of the behavior of multi-component metallurgical slags. However, the literature contains comparatively few studies of the thermodynamics of binary slags. The system Ag20-B,O, attracted our attention as it was known to give a single liquid phase,',' in which high contents of silver could be obtained (up to 61 pct Ag according to Foex2). Further, it would be expected that the partial pressure of oxygen over melts in equilibrium with metallic silver could be used to determine the activity of Ag2O in the Ag,O-B,O, system. In many respects, it may be expected that the reaction of a basic oxide with boric oxide would be analogous to its reaction with silica. Liquid immiscibility frequently occurs in both borate and silicate systems. With B2O3 and SiO reaction with a basic oxide presumably involves a breakdown of the three-dimensional network of the acid oxide by reaction with oxygen atoms common to more than one silicon or boron atom. Ag2O-B2O3 was therefore investigated as a model of a slag system in the hope that its thermodynamic properties would assist in understanding those of other systems. Several methods for determining the activity of a component in a slag have been described in the literature. Chang and Derge" used high temperature electromotive force measurements to obtain the activity of SiO2 in CaO-SiO2 and Ca0-Al203-Si02 slags, but the cell reaction in their work is not clear. low has used rate of volatilization and vapor pressure measurements combined with phase diagrams to obtain activities in the systems KO-SiO,, Na,O-SiO, and Li,O-SiO," and PbO-SiO26 Taylor and Chipman7 extrapolated their results for the distribution of FeO between liquid iron and CaO (+Mg0)-FeO-SiOl slags to obtain the activity of FeO in the binary FeO-SiO2 system. In principle, one of the most direct methods for obtaining the activity of a metallic oxide in a phase is by comparison of the equilibrium oxygen pressure for the system metal-pure oxide with that of metal oxide-containing phase. Schenck and othersa have studied the stabilization of Ag2O on combination with other oxides (MO,) in the solid state by measurements of the oxygen pressure in systems of the type Ag-Ag,O-xM0,-MOy-0, (gas). Schuhmann and Ensio" have determined the activity of FeO in iron silicate slags in equilibrium with solid iron, using CO/CO2 mixtures to establish known partial pressure of oxygen. Although the method gives the activity of FeO without ambiguity, the slag is not a binary system, and interpretation of the results in terms of the hypothetical binary system FeO-SiO, is not possible. If a metal is solid at temperatures at which the properties of the slag containing its oxide are to be studied, this method has the considerable experimental advantage that the metal can be used as the container for the slag, and contamination by contact with refractories is avoided. In this work, crucibles for Ag2-B,O, melts were made from silver, and the liquid brought to equilibrium with definite pressures of oxygen gas. The oxygen pressure PO, thus fixes the activity of Ag20 in the liquid silver borate. For the reaction at a given temperature. is substantially constant, is directly proportional to the square root of the equilibrium oxygen pressure. Varying the oxygen pressure changed the silver oxide content of the liquid and it was possible to obtain the activity of Ag2O over a range of composition. Experimental Procedure In principle, the method consisted of bringing melts in silver crucibles or boats to equilibrium at a fixed temperature under a definite pressure of oxygen and analyzing the glass after solidification. Materials: B2O3 glass was prepared from A.R. quality boric acid by fusion in platinum. The silver

Jan 1, 1954

By Kenneth K. Humphreys, J. W. Leonard, T. S. Spicer

INTRODUCTION Reasons for Thermal Drying The continuing increase in the percentage of - '/4 in. ( -6.3 mm) coal produced as a result of the increased use of mechanical mining methods has, over the years, tended to shift considerable portions of the work of producing salable coal to the fine coal cleaning section of the preparation

Jan 1, 1979

By G. V. Woody, J. D. Price

Many producers of by-product coke have spent considerable time and given considerable thought to the use of a substitute for low-volatile coal as an admixture with high-volatile coking coal for charging high-temperature by-product coke ovens. Generally speaking, there are two principal reasons for this, one being occasioned by the lack, in certain localities, of low-volatile coal delivered at a cost sufficiently low to permit its use, and another being the desirability of finding a substitute product with a delivered cost lower than the cost of the low-volatile coal. It was for the first of these reasons that the Colorado Fuel and Iron Corporation was particularly interested in this subject. This paper deals with the experimental work and the results obtained by that company in producing and using low-temperature char. Overcoming Defects in Coke from Colorado Coals Colorado coking coals are all of the high-volatile type and were all laid down during the Cretaceous period, which makes them something like 80 million or so years younger, geologically, than eastern coals. They behave, when coked alone, something like the Pittsburgh-seam coals. They make a very brittle, highly cross-fractured, fingery coke of fair shatter value but quite low in resistance to abrasion. Mixed with a low-volatile coking coal such as Poca- hontas or certain Oklahoma coals, they make a coke that is equal to virtually any of the eastern cokes. The inferior quality of Colorado coke has long been recognized, and a considerable amount of time, effort and money has been spent upon investigations of means for improving its physical properties. While it has been found that some improvement in quality of coke can be secured by the adjustment and control of such variables as pulverization, moisture content, oven temperature and bulk density of charge, and through the addition of certain inert materials such as pulverized coke breeze, pitch, high-volatile noncoking or semicoking coal, the benefits so gained have been comparatively small. There are no low or medium-volatile coking coals available within a reasonable freight-cost distance. Therefore a substitute for low-vol-atile coal was developed through low-tem-perature carbonization—or, more correctly speaking, through the partial devolatiliza-tion—of high-volatile coals. The blending of this lower volatile char with the high-volatile coking coal distinctly improves the quality of coke made from Colorado coal. There is nothing especially new about this idea. In 1908 a Japanese patent was taken out by Kotaro Shimomura, whose claim was: The method of making a nonfingery coke out of bituminous coals, without using natural coals of low-volatile matter; according to which a certain coal is heated at a temperature between about 300°C. and about 600°C. so as to leave about 15 to 25 per cent volatile matter in the coal; this coal is mixed with the

Jan 1, 1944

By J. D. Price, G. V. Woody

Many producers of by-product coke have spent considerable time and given considerable thought to the use of a substitute for low-volatile coal as an admixture with high-volatile coking coal for charging high-temperature by-product coke ovens. Generally speaking, there are two principal reasons for this, one being occasioned by the lack, in certain localities, of low-volatile coal delivered at a cost sufficiently low to permit its use, and another being the desirability of finding a substitute product with a delivered cost lower than the cost of the low-volatile coal. It was for the first of these reasons that the Colorado Fuel and Iron Corporation was particularly interested in this subject. This paper deals with the experimental work and the results obtained by that company in producing and using low-temperature char. Overcoming Defects in Coke from Colorado Coals Colorado coking coals are all of the high-volatile type and were all laid down during the Cretaceous period, which makes them something like 80 million or so years younger, geologically, than eastern coals. They behave, when coked alone, something like the Pittsburgh-seam coals. They make a very brittle, highly cross-fractured, fingery coke of fair shatter value but quite low in resistance to abrasion. Mixed with a low-volatile coking coal such as Poca- hontas or certain Oklahoma coals, they make a coke that is equal to virtually any of the eastern cokes. The inferior quality of Colorado coke has long been recognized, and a considerable amount of time, effort and money has been spent upon investigations of means for improving its physical properties. While it has been found that some improvement in quality of coke can be secured by the adjustment and control of such variables as pulverization, moisture content, oven temperature and bulk density of charge, and through the addition of certain inert materials such as pulverized coke breeze, pitch, high-volatile noncoking or semicoking coal, the benefits so gained have been comparatively small. There are no low or medium-volatile coking coals available within a reasonable freight-cost distance. Therefore a substitute for low-vol-atile coal was developed through low-tem-perature carbonization—or, more correctly speaking, through the partial devolatiliza-tion—of high-volatile coals. The blending of this lower volatile char with the high-volatile coking coal distinctly improves the quality of coke made from Colorado coal. There is nothing especially new about this idea. In 1908 a Japanese patent was taken out by Kotaro Shimomura, whose claim was: The method of making a nonfingery coke out of bituminous coals, without using natural coals of low-volatile matter; according to which a certain coal is heated at a temperature between about 300°C. and about 600°C. so as to leave about 15 to 25 per cent volatile matter in the coal; this coal is mixed with the

Jan 1, 1944

By D. N. Collins, J. R. Stevens

New expressions in the technical assessment of mineral concentration processes have been defined as Concentration Efficiency: Concentration Index: where h is the percentage of head grade; c, percentage of concentrate grade; maximum percentage of concentrate grade obtainable at 100 pct recovery; and R, percentage of recovery. In developing these expressions, the basic requirements for the technical assessment of concentration operations have been defined and the various expressions suggested by others for asses-sing the results of such operations reviewed and criticized. The derived expression for E, depends on a simple combination of recovery (R) and a new term selectivity (s), where S is (c-h)/(c,,, -h), by definition. Limitations of the new expressions are discussed and examples given of applications to concentration operations. It is also shown that the formulae for E, and I, can be applied to scatter diagrams, multi-component systems, size-assay analyses, and classification data. Many methods of quantifying concentrating operations have been proposed. These include direct presentation of weights and assays together with the derived data for other items such as ratio of concentration, recovery, rejection losses, and enrichment ratio. These terms are now accepted mineral dressing nomenclature, and in the assessment of the results of concentration processes, concentrate grade and recovery are together the most widely accepted measures of concentration efficiency. For practical purposes when the concentrate grade is equal to or greater than a predetermined acceptable value, test or operation results may be assessed by the recovery figures alone. Similarly when recovery is within specification, the operation can be assessed by concentrate grade. The ultimate assessment and selection of a concentration operation is—admittedly with a few ex-ceptions—a matter of economics. There is however an intermediate stage in laboratory and pilot plant testing when evaluations must be made without the assistance of an economic yardstick. Numerous techniques are employed to facilitate such evaluations among which may be mentioned graphical plots of grade vs recovery, cumulative weight vs cumulative recovery, scatter diagrams, etc. The desirability of combining metallurgical results and product weights and assays into a single value has been recognized and discussed by many workers1-" with the resultant suggestion and use of such terms as Metallurgical Efficiency ,' Summation Index,' Selectivity Index,'- and Absolute Efficiency B None of these expressions completely satisfies all the requirements of a selectivity or efficiency formula and it is proposed in this report to define the terms, relating to the technical efficiency of a concentrating operation, concentration index and concentration efficiency. Examples are given of their application and a comparison is made of their use in relation to other formulae. All product grades in this paper will be referred to in terms of mineral percentages and a list of symbols is given in Table I. CONSIDERATION OF EXISTING FORMULAE A number of the existing efficiency and index terms used in quantifying concentrating operations are considered below. Metallurgical Efficiency:

Jan 1, 1961

By John W. Cahn

Tile avevage value of the mean curvature of surfaces in a specimen can be precisely delermined by sitrlple measurements performed on random sections or on 1 vojectiotzs of these surfaces. For surjaces enclosing particles, the quanlity is sllown to be velated to the probable tlurrrhev ot particles intevcepted by a random platne and, under certain conditions, to the perimeter of their Projected image and to the increuse in surjace area when the particles grow. Many of the velalionsl~ips are valid for any shape or size of the particles. THE recent growth of the field of stereology or quantitative microscopy is largely the result of the discovery of a limited number of rigorously defined geometrical parameters that can be precisely measured on random sections or projections of a structure without assumption about particle shape.' Such parameters have been the volume fraction and surface area and line length in a unit volume. The present paper introduces the average of the mean curvature of surfaces in a volume, and shows that it can be rigorously measured by counting methods alone. AVERAGE CURVATURE ON A PLANE Consider a curve y = v (x) on the xy plane. The curvature Kp in the plane at any point is given by Kp = (dZy/dx2)/[1 + (dy/dx)2]3/2. The average curvature along the curve may be defined as Kp = (1/L) fKpdl |1] where dl = [(dx)' + {dy)2]1/2 is the length of an element of the curve and .(dl = L is the total length of the curve. The integral on the right-hand side of Eq. [l] will be called the total curvature, It is dimensionless and readily evaluated by substituting d = arctan (dy/dx) to give ah which is the change in direction of the tangents to the ends of the curve, Fig. 1. Hence, Kp = (1/L)(?h) [2] It can also be seen that Eq. [2] is applicable in the limit of sharp corners. Even though the curvature is infinite at a corner, f Kpdl remains integrable and equal to the change of angle. The signs of the quantities Kp and dl are arbitrary and, if the areas adjoining the curve are equivalent, there seems to be no compelling reason to fix the signs. For a closed curve, however, it becomes worthwhile to use the signs to distinguish among different types of areas or phases enclosed. For such a closed curve AO = t2r and As a convention, we shall take the direction of the integration along the curve in such a way that the area or phase under consideration (henceforth called a) is kept to the left side of the curve, Fig. 1. Then Kp is positive where the curve is concave and negative where the curve is convex. For a closed curve Ad - +2n if the curve encloses a, and Ad = —2n if the curve is surrounded by a. PLANAR SECTIONS OF THREE-DIMENSIONAL STRUCTURE OF ISOLATED a PARTICLES Consider next a planar section of a microstructure containing particles of a in a matrix. We place no restriction on shape of connectedness of these particles, and begin by considering the curvature of only the boundary between a particles and matrix' Clumps of contiguous a particles are then treated as a single a particle. The two-dimensional section will show in a unit area na simple closed curves enclosing a particles and nh closed curves of holes, in-

Jan 1, 1968

By Edwin S. Smith

Current research in the phenomenon of liquefaction has indicated some of the in situ material properties which should be considered in the planning of waste deposits. Practical experience in tailings disposal, however, must be combined with the results of theoretical and laboratory studies to develop the most economical and safe techniques against liquefaction failure. One of the many responsibilities of mining engineers and mill superintendents is providing for the safe and economic disposal of wastes (or tailings) that remain after a mineral has been extracted from an ore. The cost of disposal of these tailings exerts a considerable influence on the minimum grade of ore extracted from the mine. It varies to a great extent with each project and is dependent on many diverse factors, such as the local topographical, meteorological and seismological conditions together with the applicable water pollution laws of the area. The process used to extract a mineral from an ore determines the physical characteristics of the tailings and, thus, is one of the more important factors to be considered in the planning of disposal facilities. Today's trends in ore-processing operations complicate the problems of waste disposal: recently developed extraction methods result in the production of finer tailings, the use of lower grade ores produces greater volumes of wastes, and the enhanced land values in the vicinity of industrial plants preclude the former practice of spreading wastes laterally and force the expansion of waste deposits upwards to greater heights. The frequent occurrence of stability problems in waste deposits indicates that many mining organizations may not be employing current soil mechanics concepts and practices. One of the most spectacular, and often most disastrous, types of tailings dike failure is that resulting from liquefaction of the fine particles combined with a breach of the main retaining dike. TAILINGS DISPOSAL Since the many variations in site conditions, ore bodies and extraction processes dictate the general disposal procedures and mill superintendents incorporate special facilities to provide the most economical disposal method consistent with accepted safety standards, it is only natural that there are almost as many different disposal techniques as there are disposal areas. Methods used in the construction of retaining dikes include: (1) transporting granular material from an adjacent borrow pit and (2) mechanically segregating the total effluent and allowing the coarse fraction to form a retaining dike by hydraulic deposition. The location of a plant and the surrounding topography exert the greatest influence on the layout of tailings disposal facilities. Most of the above-ground waste deposits can be classified as side-hill, (Fig. 1,) cross-valley, (Fig. 2,) or stockpile deposits. Side-hill disposal areas are probably the most common type of layout because of the natural advantage of gravity flow, without the disadvantage of river diversion usually needed for tailings ponds formed behind cross-valley dikes. Stockpiles, which are placed on relatively level ground, require four retaining dikes and the effluent generally has to be pumped up into the disposal area. Changing production techniques or developments adjacent to disposal areas have produced variations or partial combinations of these three categories. In a series of articles, Lenhart describes many of the problems encountered in the disposal of tailings and certain innovations used to overcome them. He discusses weaknesses in design and layout and shows that disposal techniques are often learned the

Jan 1, 1970

By Duane E. Eubanks, David B. McDonald

The anticipated growth of western coal production over the next decade is expected to add approximately 274.3 million t/y (302.4 million stpy) to US capacity. The increased output, coming from both new mines and expansion of existing facilities, promises to bring with it a new set of problems for mine planners.

Jan 8, 1976

Secretary's NOTE—on page 176,of this paper, in the fourth line of the first footnote, " 4° " should, be " 1" )'; and on page 180, at the beginning of line 23, ('lined boot" should be " unlined boot." Discussion. Frank Owen, London, Eng. (communication to the Secretary) : In his very interesting and valuable paper, Prof. Snow

Jan 1, 1900

By William Rawles

THIS report is the first of a series planned by The Mineral Inquiry, organized by the American Institute of Mining and Metallurgical Engineers to make factual studies of the world's mineral resources in their political and international relations. The efforts of The Inquiry are directed toward filling certain notable gaps in available information, without duplication of many other investigations touching this field which are being conducted by other organizations, public and private. Much of its work is indirect-suggesting topics for study by other individuals and organizations and aiding in their planning and correlation. The present report is designed to show the nationality of the commercial control of mineral resources, in addition to the geographic distribution of production by nation-in other words, the overlapping of political and commercial control. Detailed accuracy cannot be claimed for the figures presented. For some companies division of ownership by nationality has been very difficult to ascertain, as many companies are reluctant to dis-close the extent of their holdings in other corporations, and the impossi-bility of tracing all stock ownership is obvious. Also there are frequent changes in the distribution of stock control. The percentages given rep¬resent only the participation that has been identified definitely, even though it is known in some cases with fair certainty that larger interests exist, as, for instance, American participation in foreign chromite companies. In bringing together these figures we have had the cordial co-operation of the United States Bureau of Mines, the United States Geological Survey, the Mineral Division of the Bureau of Foreign and Domestic Commerce, the Brookings Institution, and of many specialists connected with the mineral industries covered. The report does not cover the entire list of minerals. Coal, although it is, one of the most important minerals, was omitted because not more than one or two per cent of the world's production comes from mines that are owned by companies foreign to the countries where the coal is produced. Others, such as gold, are omitted because sufficient data have not yet been accumulated. It is hoped that some of these may be covered in subsequent reports. It is not our purpose here to discuss the significance of the figures sub-mitted, but only to point out the general nature of the problems on which they have a bearing.

Jan 1, 1933

By A. W. Dana, F. J. Shortsleeve, A. R. Troiano

The phenomenon of flake formation which may occur during cooling or room temperature aging of large steel sections is caused by a combination of hydrogen and stress. As such, the transformation characteristics of austenite in a particular section play a major role in the occurrence of flakes. Isothermal and continuous cooling studies demonstrated that flake formation is particularly sensitive to nature, distribution, and relative proportions of the microconstituents in the cooled sections. In the absence of transformation stresses, abnormally high hydrogen contents could be tolerated without flaking. Flakes were not found in the absence of transformation stresses. No correlation existed between average hydrogen content and flake formation where cooling stresses were low. Hydrogen in the steel was a necessary but not sufficient condition for flake formation. DESPITE a long history of laboratory and commercial investigations, the mechanics of hairline crack or flake formation in large steel sections remains unresolved. Recent publications1-5 emphasize that flaking is still a serious problem.* immediate cause of flaking is the pressure exerted by entrapped hydrogen gas that collects in small internal cavities or voids during cooling. When the temperature of steel is low enough to preclude extensive plastic flow, the cavity pressures cause local brittle fractures (flakes). 2—The microsegregation and transformation stress theory. According to this theory, flaking results from stresses arising from the volume changes accompanying the formation of martensite and/or low temperature bainite after a large portion of the section has transformed at elevated temperatures. The proponents of the hydrogen theory suggest that stresses from other sources may be additive to the stresses produced by hydrogen and may, in some cases, provide the increment of stress necessary to produce flakes. However, the critical aspect of this theory is that no stress, except that produced by hydrogen pressure, is believed to be sufficient in itself to produce flakes. For practical considerations, this means that the only safe thermal cycle is one that reduces the hydrogen content to some low critical level below which flaking will not occur regardless of the state of transformation. On the other hand, the transformation stress theory does not rule out the embrittling effect of hydrogen. This theory simply proposes that flaking does not occur in the absence of transformation stresses. With the transformation stress theory as a basis, commercial antiflaking cycles would be designed to complete the transformation of austenite prior to cooling to ambient temperatures. Both of these theories appear compatible with industrial experience. Antiflaking cycles, such as cooling at extremely conservative rates or soaking for periods of time in the upper subcritical temperature ranges prior to cooling to ambient temperatures, may allow the following: 1—diffusion of hydrogen out of the section and, as a consequence, reduction of the hydrogen content to some safe minimum value (hydrogen pressure theory) and 2—complete decomposition of the austenite at rela-

Jan 1, 1956

By Richard H. Willey

EFFICIENT handling of large tonnages in open-pit mining demands, primarily, an abundance of well fragmented rock. To accomplish this, a drilling and blasting department composed of a well-knit, versatile group of workers and supervisors must be established. Some costs will remain relatively on the same level over a period of years, while others can be greatly lowered as will be shown. The Utah Copper Division, Kennecott Copper Corp., Bingham Canyon, Utah, has been ever alert for improvements in operating methods and equipment to gain increased efficiency and lower costs. BINGHAM CANYON MINE The greatest dimension of the mine is 8000 ft northeast-southwest with a width of 5000 ft southeast-northwest. From the lowest level, at an elevation of 6040 ft above sea-level, the mountain rises in 27 benches on the west side to W level (elevation 7800 ft) and 21 benches on the east side to Q level (elevation 7460 ft). At the inception of open-pit mining in 1906, the overcapping of leached porphyry had to be removed to reach a commercial grade of sulphide ore. The overcapping or waste now remaining is quartzite with intercalated limestone beds. The limestones are usually soft in nature except in the south-middle portion of the mine where they have been metamorphized along the contact zone. This area usually gives drilling and blasting operations a great deal of trouble. Here occurs the greatest amount of secondary blasting, as the well-established fracture planes cause the altered beds to break into large boulders. Several massive quartzite "ribs" occur in the overburden, and as a result, the uniform breakage is difficult to obtain. The major quartzite areas have been subjected to complex folding, fracturing and faulting because of the intrusion of the underlying stock. The monzonite porphyry, which carries the commercial values, presents no serious drilling problem because of its softness and extensive minute fracturing throughout the mass. A series of major slips or faults pass through the ore body which, although of considerable advantage to the drilling and blasting department, sometimes works to the disadvantage of the track department. Drilling and blasting of the porphyry apparently becomes easier as depth is attained, but highly siliceous streaks throughout the pit still present difficulties. At no place is the rock of such character as to allow excavation without preparations. On certain higher elevations or levels, old fill material can be dug by the shovels, but only at some risk to equipment because of bank slippage. PAST METHODS At the outset of shovel operations in 1906, drilling was done with the Ingersoll-Rand F-24 piston drill mounted on an un-

Jan 1, 1947

By Sidney G. Parker, Jack E. Pinnell

Cubic ZnSe crystals have been grown with HCl, HBr, and I, as chemical transport reagents. The growth of large, well faceted crystals is in the order HCl < HBr < I, with some produced by I, transport being 1 cm on an edge. ZnSe:I crystals showed no strain in cross polarized light and were homogeneous as determined by electron microprobe. These ZnSe:I crystals had a mobility at 25°C of about 300 sq cm per v-sec and contained 103 dislocations per sq cm; further, these crystals were pure cubic and showed both photolumi-nescence and photo conductance. Doping was attempted during crystal growth. ZnS and Zn (S,Se) crystals were also grown by I, transport. ZnSe crystals have been grown by most of the common techniques.1"6 Of these, chemical vapor transport as developed by Kaldis2 showed promise of producing high perfection, cubic ZnSe. With chemical transport, crystals can be grown far below the transition temperature for conversion of cubic to hexagonal structure. A small temperature gradient between source and growth region tends to allow near equilibrium growth, which should lead to good crystalline perfection. Because it is necessary to form volatile compounds, halogens or their compounds are extensively used as transport agents. HCl, HBr, and I, were compared as transport agents for growth of ZnSe crystals. Doping during crystal growth was attempted. ZnS and ZnSxSe1-x crystals were also grown by chemical transport. Electrical and optical properties of the crystals were examined. EXPERIMENTAL Fig. 1 is a sketch of the growth apparatus, which was a modification of Kaldis&apos;. A quartz rod which butted against the seed crystal served as a heat sink to promote growth only on the seed. Initially, crystal growth was attempted with ZnSe prepared by wet chemical methods. Because such a material contained oxides and selenates, HC1 reacted with these oxygen compounds to form H2O and no ZnSe transport occurred. Further, this ZnSe contained about 10 ppm Cu plus other impurities. High purity ZnSe, containing small amounts or no oxygen compounds, was prepared by reaction of H2Se with zinc vapors at 600° to 800°C in an atmosphere of palladium diffused H2.7 The ZnSe thus produced was usually copper free and contained only 1 ppm by weight of magnesium as determined by emission spectrographic analysis. G.E.&apos;s electronic grade ZnS, which had about 10 ppm metallic impurities, was used without further treatment. ZnS-ZnSe charges were prepared by mixing the two compounds together, then sintering 24 hr at 1100°C in an H2S ambient. The gases, HBr and HC1, were commercial grade, of 99.9 pct purity, while the I2 was 99.999 pct pure. The single crystal seeds with (111) faces for growth were mechanically held in place by sealing the outer quartz tube to the growth tube. After the charge was introduced, the system was outgassed at l0-5 torr while heated to 500°C. The desired quantities of transport reagents were then sealed off in the system, which had been cooled to 25°C. The growth tubes were inserted into a furnace with the desired temperature and thermal gradient. The system was allowed to cool for 1 to 12 hr from growth temperatures of about 850" to 25°C. The tubes remained stationary during crystal growth. HCl CHEMICAL TRANSPORT GROWTH Preliminary runs with HC1 had a temperature gradient of 100° to 200°C and no seed or heat sink. Only small crystals were obtained in these experiments as

Jan 1, 1970

By W. P. Haynes

The 1938 production for all of Rumania amounted to 48.984.200 bbl., as compared with 53.239.500 bbl. in 1937. which represents a drop of 4.255.300 bbl., or 8 per cent. Of the various fields. there are only two major increases in production—Baicoi-Tintea with 3.092.400 bbl. And

Jan 1, 1939

By W. P. Haynes

The 1938 production for all of Rumania amounted to 48.984.200 bbl., as compared with 53.239.500 bbl. in 1937. which represents a drop of 4.255.300 bbl., or 8 per cent. Of the various fields. there are only two major increases in production—Baicoi-Tintea with 3.092.400 bbl. And

Jan 1, 1939

By T. J. Koppenaal, M. E. Fine

A yield point effect attributed to short-range ordevi?g (SRO) occurs in Cu base Al. At at 296°K varies with heat treatment, decreasing as the annealing ternperature is raised .from 433Oto 598°K. Davies aid Cahn&apos; observed a corresponding decrease in SRO. (523 °K anneal, measured at 77°K) is approximateZy proportional to the variations of A7 with strain rate and testing temperature are also Consistent with the idea that is associated with SRO. A preliminary investigation of the tensile properties of a Cu-Al single crystals showed the presence of a rather strong yield-point effect (drop in flow stress after initial yielding). The object of this research was to investigate its origin and behavior. a Cu-Al alloys are particularly interesting because diffuse X-ray scattering measurements by Davies and Cahn,&apos; Houska and Averbach,&apos; and Borie established the presence of short-range order. The degree of local order may be changed with heat treatment.&apos; cottrel14 suggested that the presence of local order might result in a yield-point effect, and thus the possibility exists here for experimentally ascertaining the importance of short-range order with respect to yield points in these alloys. Since a 12 pct difference exists between the atomic sizes, elastic or Cottrell locking5 must also be considered. Further, Howie and swanne have shown that the stacking fault energy of copper is reduced by aluminum additions. The width of extended dislocations should thereby increase. hus the conditions appear attractive for Suzuki locking.&apos; Finally, the possibility of stress-induced order at dislocations, schoeck locking,&apos; must also be examined, EXPERIMENTAL PROCEDURE Alloys up to 14 at. pct Al were prepared by induction melting high-purity (99.999 pct) Cu and (99.996 pct) Al in a graphite boat under a dynamic vacuum of 5 X 10o mm of Hg. After homogenizing the ingots at 900°C for at least 24 hr under vacuum, they were rolled with intermediate anneals to strips 1.55 mm thick. Single crystals 10 in. long were grown by lowering strips, contained in a split graphite mold sealed in fused quartz at 5 X lo-&apos; mm of Hg, through a single coil induction heater at a constant rate of 1/2 in. per hr. Tensile specimens 1.25 in. long were cut from the single crystal strips and reduced cross sections about 0.7 in. long and 3.0 to 3.5 mm2 in area were introduced by filing and abrading.8 To remove the worked portion about 10 pct of the cross-sectional area was removed by etching. Back-reflection Laue photographs of a filed and etched specimen were taken before and after annealing at 900°c for 24 hr. Small, well defined Laue spots were obtained with no visual difference in the two photographs. Further, specimens with and without the reduced sections began yielding at about the same stress. Hence, for our purposes, filing and abrading did not affect the structure of the specimens. Each single crystal specimen was annealed at 900° c in vacuum for at least 24 hr and furnace cooled; while cooling through the range of 250o to 200°c, the rate was about 55°C per hr. Orientations were determined by the usual back-reflection Laue

Jan 1, 1962