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By R. F. Mehl, C. S. Barrett, A. H. Geisler

THE complicated nature of the property changes that accompany age-hardening has made it necessary to reconsider and to elaborate the simple dispersion theory.1 It has been apparent for some time that more direct information is needed concerning the atomic mechanism of precipitation from solid solution if the theory of age-hardening is to be developed past a purely qualitative stage.‡ Fortunately, in recent years information has been developed which allows us to trace in detail the lattice changes accompanying the process of precipitation from solid solution. This information relates chiefly to the sequence of the several stages of lattice alteration by which the matrix lattice is transformed to the equilibrium precipitate lattice, and permits reasonable inferences to be drawn concerning the states of internal strain that accompany these stages. This information is most complete for the solid solution of copper in aluminum. The equilibrium precipitate is the ?' phase," CuA12." This forms in plates, parallel to the {100} matrix planes. The atom pattern and atom spacing on the plane in ?' lying parallel to this plane are roughly similar to those opposite it in the matrix.' The formation of ?', however, is preceded by the formation of a transition lattice ?', which again lies parallel to the matrix {100} plane, but in this case the matching of atom pattern and spacing across the matrix-precipitate interface is nearly perfect; thus there is lattice coherency across the interface and the interface is a plane of separation in only a very limited sense.4-8 Particles of ?' may be recognized microscopically and form a well-defined Widmanstätten figure.* The ?' lattice is recognized by the occurrence of diffraction lines on powder photograms or of spots on Laue photograms which may be analyzed in the usual ways to determine the lattice type and the lattice spacings. Very recent research 6.7,1 has shown, however, that streaks occur on the Laue photogram at an earlier point in time during the aging process. These streaks have been attributed to a preferential clustering of copper atoms into small platelike elements of only a few atoms in thickness and a few hundred atoms in diameter, probably of the composition of "CuA12," which may be designated as Guinier-Preston zones or aggregates.1 It is difficult if not impossible to determine the lattice type and lattice spacing in these aggregates. It is possible1 that these aggregates are in fact merely ?' in a high state of dispersion; the inclination to argue this is strong, as we shall see later. Never-

Jan 1, 1941

By C. G. Dunn, J. L. Walter

Either (100) (001] or (110) (0011 oriented secondaries, or both, depending on annealing atmosphere and material composition, grew in a matrix of 2-dimensional grains. The growth dependence of secondaries of(hkl) orientation in the iron and 0.6 pct Si-Fe is explained in terms of surface energy driving forces and the dependence of these driving forces on both (hkl) orientation and amount of oxygen at the surace. (100) [001] grains grow when suflicient oxygen is present at the gas-metal interface to make y100 the lowest (hkl) surface energy. When little or no oxygen is present at the gas -metal interface, making Y110 the lowest surface energy, (110) (001 1 grains grow. When Y100 and y110 are approximately equal but lowest among (hkl) surface energies, both (100) (0011 and (110) (0011 grains grow simultaneously. An excess of surface oxygen produces complete grain growth inhibition in high-purity iron; i.e., a 3-dimensional grain structure is retained over long annealing periods. However, a small amount of silicon in the iron reduces the growth inhibiting tendency of oxygen without altering the (hkl) surface energy relationships. Some information on annealing twins is presented. SECONDARY recrystallization in high-purity a iron, cold rolled to strip a few mils thick, was reported recently.' The present paper provides additional information on this subject, also furnishes similar information on both zone-refined iron and a high-purity 0.6 pct Si-Fe alloy, and treats the more important surface phenomena that determine to a large extent whether one or more secondary recrystallization textures will appear during annealing. Previous studies on high-purity 3 pct Si-Fe2,3 have shown the importance of (hkl) and impurity-atom dependent gas-metal surface energies on the formation of secondary recrystallization textures in high-purity material. For example, either (100) or (110) oriented grains may have the lowest surface energy depending on the amount of oxygen at the gas-metal interface, and hence the highest driving force for growth if other growth factors are the same. Another important phenomenon in high-purity material is stabilization of the matrix structure; neither impurities nor a strong matrix texture seem to be necessary when recrystallization and normal grain growth occur, developing a stable two-dimensional type grain structure. The average grain size in such a structure is usually between one and two times the specimen thickness. Consequently, most of the grains have grain boundaries terminating at exposed surfaces, and this results in gas-metal interfaces (or hkl surfaces) as well as

Jan 1, 1962

By S. R. B. Cooke, S. W. Clark

Adsorption of calcium and magnesium by quartz was determined over a wide pH range using flame photometry for solution analysis. A parallelism was noted between calcium adsorption at alkaline pH and flotation of activated quartz by means of various fatty acid collectors. Abstraction of calcium from 100 ppm solutions increased above pH 10.5 until all of the calcium was removed at pH values between 12 and 12.5. Calcium was desorbed from the quartz over the pH range 12.5 to 13.0 and returned to the solution. Precipitation of magnesium hydroxide prevented measurement of magnesium adsorption above a pH of 9.9. A calcium adsorption isotherm determined at pH 9 matched measurements of barium adsorption by other investigators over the range where adsorption corresponded to the equation of a Freundlich adsorption isotherm. In accordance with the Gouy-Chapman theory for adsorption in the diffuse layer, adsorption of the divalent cations under these conditions was four times the adsorption of monovalent sodium. Activation of quartz by multivalent metal cations in flotation using anionic collectors has been studied by many investigators. The occurrence of activator cations in the water of flotation mills makes depression of silaceous gangue very difficult. Alternatively, in some flotation systems it is practical to cause activation of quartz so that it is floated from the desirable mineral values. The present study was initiated to investigate the activation of quartz by measuring adsorption of calcium and magnesium ions as a function of solution pH. Early empirical tests showed that quartz would be activated to flotation using anionic collectors if any of a number of divalent or trivalent metallic cations were present in the pulp.1*2 Subsequent studies by Clemmer,3 Cooke and Digre,4 Cooke,5 Schuhmann and Prakash,6 and Cooke, Iwasaki, and Choi,7-8 showed that flotation depended strongly upon the alkalinity of the flotation pulp, the concentration of the activating cations, and the chain length and structure of the organic collectors used. The concept of the electrical double layer has been applied to the case of the quartz-aqueous solution interface by Gaudin and Fuerstenau,9-10 and by de Bruyn 11 to explain the activation of quartz by multivalent metal cations. Studies of the surface of quartz by Gaudin and Fuerstenau9,10,12 and by Li " using the streaming potential technique contributed much information on the electro kinetic condition of the surface of quartz in the presence of various organic and inorganic ions. Measurement of adsorption of ac-

Jan 1, 1969

W. 0. HOTCHKISS, * Madison, Wis.-It was fortunate, in a way, that this subject of income taxation came up during the war when the government could appeal, on the basis of patriotism, to men of large, business experience to undertake the work of formulating this matter and practically donate their services. At any other time the government would have been unable to secure the necessary number of men of like ability. There is one thing that must not be forgotten. In all cases, except the very largest and most important, the application of this income tax law to a mine is going to be made by clerks whose salaries are fixed by a Congress that appropriated a half billion of dollars for a war purpose and provided that $4000 should be the highest salary paid in the adminis-tration of that great sum. The Congress of the United States unfor-tunately seems almost entirely unaware of the fact that men of ability and experience are worth money and must be paid adequately if proper public service is to be secured. While the formulation of the broad plans for administering the law have been in the hands of. men of unusual ability, we have the future to guard, and, if possible, the mining industry should see to it that the Internal Revenue Department is permitted by Congress to pay salaries that will enable it to secure men of the ability and experience necessary to insure proper consideration of the industry. Another fact to which I would call attention is the difference between the business and the economic, or economists', points of view as to what a mine is. The economists have properly, from their point of view, classed a mine as a wasting asset, but the average metal mine is not a wasting asset if it is a real mine. The economists are correct in a broad sense in that every ton of ore taken out diminishes the orebody by that amount. When it comes to the consideration of an individual mine from the business standpoint, the future of that mine rests on unfore-seeable events-the discovery of new orebodies or the development of old bodies beyond the point of probable expectations. Every going mine must develop new ore each year to replace that extracted, and, .during all but the last one or two years of its activity, actually does so. Instead of being a "wasting asset" such a mine, from the practical stand-point which must control valuation, is an "increasing asset." From

Jan 12, 1919

By H. S. Kalish, F. J. Dunkerley

INTRODUCTION AND PREVIOUS WORK THE determination of the low temperature tensile properties of tin and tin-lead alloys was initiated as part of an extensive research program on the phasial equilibria of tin binary systems below 13.2°C and the property changes associated therewith. A survey of the literature revealed that, although some room temperature tensile data were available for tin and a few tin-lead alloys, there existed no survey report of the low temperature tensile properties of these alloys. Accordingly, the tensile data for tin and a series of pure tin alloys containing from 0.01 to 50 pct lead were measured at eleven temperatures from -196 to +20°C and are presented here in survey form. In addition to the importance of these tensile data as a preliminary to certain fundamental research, this information promises also to be of commercial importance because of the 'extensive use now being made of tin-lead solders in refrigeration and gas liquefaction equipment. Occasional failures of the solders in low temperature service have been attributed by some metallurgists to low temperature mbrittlement, while others hold that failure results from the transformation of the tin-rich phase of the solder from the solid white (beta) to the powdered gray (alpha) form. From the data previously available in the literature, neither of these explanations can be completely corroborated. Cohen and van Lieshout1 have established the beta to alpha tin transformation at 13.2°C and have shown that strain accelerates the transformation. Further, they showed that although I pct lead retarded the transformation, it was not eliminated. It will be pertinent, then, to note whether the beta to alpha transformation of the tin-rich phase will occur in a short-time tensile test. Polanyi and Schmid2 studied the tensile . properties of tin single crystals at -185°C and report that tin is brittle at-this temperature. In 1946, Kostenets3 reported that pure tin was brittle at -196 and -253°C, while lead remained ductile at both these temperatures. He also studied the tensile properties of tin-lead alloys containing from 10 to 75 pct lead at -253, -196 and +17°C. The ultimate tensile strength increased more than two times in going from room temperature to -196°C and about three times that value on further cooling to -253°C. All the alloys were brittle at -6°C except those containing more than 75 pct lead. Even at -253°C, the 75 pct lead alloy retained 36 pct elonga¬tion in a 30 mm-ga. length. Creep is also an important phenomenon

Jan 1, 1948

By Dr. O’Neil Thomas J., Donald W. Gentry

"We can easily represent things as we wish them to be..” -Aesop INTRODUCTION The first eight chapters of this book have primarily addressed the concepts and fundamentals associated with project valuation, particularly as they apply to mining-related projects. Specific emphasis was placed on the engineering and economic parameters and estimates necessary for quantifying line items in pro forma income statements and the subsequent calculation of net annual cash flows. Given these estimates of relative benefits, costs, and annual cash flows for a project, it then becomes necessary to convert these estimates into measures of relative desirability or attractiveness. The criteria and techniques typically utilized to determine project acceptability or desirability are, then, the topics of discussion in this chapter. Before proceeding it may be advantageous to put the problem of project evaluation into proper perspective by reiterating some of the fundamental concepts developed in Chapter 1. The basic premise set forth in this book is that the objective of the firm should be to maximize its value to its owners (stockholders' wealth). This value or wealth is represented by the market price of the firm's common stock; consequently, the primary objective of the firm can be restated as one of maximizing the value (price) of the firm's common stock in the marketplace over the long term. This price, in effect, is an indicator or index of the market's perception of the company's progress and a reflection of the investment, financing, and dividend decisions made by corporate management. These three major decisions are inter- related and jointly impact the market price of the firm's stock. Therefore, if the objective of maximizing the value of the firm to its stockholders is to be achieved, an optimal combination of the three decisions will be necessary. It is important to emphasize that maximizing stockholder wealth is not the same as profit maximization or even the maximization of earnings per share. Corporate managers making decisions based on the objectives of maximizing profits and/or earnings per share must recognize that these objectives do not consider: (I) the timing or duration of expected returns, (2) the business and financial risks of the promised earnings, or (3) the dividend policy of the firm. Obviously all of these factors are of concern to the company's stockholders and will have an effect on the market value or price of the firm's common stock. Maximizing market price per share, or wealth maximization, is therefore the appropriate corporate objective and

Jan 1, 1984

By R. F. Mehl, C. S. Barrett, A. H. Geisler

The complicated nature of the property changes that accompany age-hardening has made it necessary to reconsider and to elaborate the simple dispersion theory.l It has been apparent for some time that more direct information is needed concerning the atomic mechanism of precipitation from solid solution if the theory of age-hardening is to be developed past a purely qualitative stage.‡ Fortunately, in recent years information has been developed which allows us to trace in detail the lattice changes accompanying the process of precipitation from solid solution. This information relates chiefly to the sequence of the several stages of lattice alteration by which the matrix lattice is transformed to the equilibrium precipitate lattice, and permits reasonable inferences to he drawn concerning the states of internal strain that accompany these stages. This information is most complete for the solid solution of copper in aluminum. The equilibrium precipitate is the 0 phase, "CuAl2." This forms in plates, parallel to the {100} matrix planes. The atom pattern and atom spacing on the plane in 0 lying parallel to this plane are roughly similar to those opposite it in the matrix.3 The formation of 0, however, is preceded by the formation of a transition lattice 0': which again lies parallel to the matrix {100} plane, but in this case the matching of atom pattern and spacing across the matrix-precipitate interface is nearly periect; thus there is lattice coherency across the interface and the interface is a plane of separation in only a very limited sense.'8 Particles of 0' may be recognized microscopically and form a well-defined Wid-manstatten figure.* The 0' lattice is recognized by the occurrence of diffraction lines on powder photo-grams or of spots on Laue photograms which may be analyzed in the usual ways to determine the lattice type and the lattice spacings. Very recent research 6, 7, 1 has shown, however, that streaks occur on the Laue photogram at an earlier point in time during the aging process. These streaks have been attributed to a preferential clustering of copper atoms into small platelike elements of only a few atoms in thickness and a lew hundred atoms in diameter, probably of the composition of " CuA12," which may be designated as Guinier-Preston zones or aggregates.1 It is difficult if not impossible to determine the lattice type and lattice spacing in these aggregates. It is possible1 that these aggregates are in fact ?' in a high State of dispersion; the inclination to argue this is strong, as we shall see later. Never-

Jan 1, 1941

By C. E. Mills

The Verde mining district is in Yavapai County, in north-central Arizona. Jerome, the principal town, has a population of 6000 and the two important mines of the district—the United Verde and the United Verde Extension—lie almost within its limits. Perched high on the southwest side of the Verde valley, at elevations of 4800 and 5500 ft. (1463 and 1676 m.) respectively, both mines have long tunnels from their lower levels through which the ore is hauled to the smelters. Clarkdale, the smelter town for the United Verde mines, lies in the Verde valley 4.1 miles (6.5 km.) distant in an air line from Jerome and at an elevation of 3560 ft. The two towns are connected by the standard-gage railroad of the Verde Tunnel and Smelter R. R., a subsidiary of the United Verde Copper Co. This road is 11 miles long and has a compensated grade of 4 per cent. At Clarkdale it connects with the Verde valley branch of the Santa Fe railroad. Clemenceau, the smelter town for the United Verde Extension mine, lies 3.5 miles (5.8 km.) in a southeasterly direction from Clarkdale, at an elevation of 3420 ft. It is connected to the Verde Valley branch of the Santa Fe railroad at Clarkdale by the Arizona Extension R. R., a subsidiary of the United Verde Extension Mining Co. The portal of the haulage tunnel, at an elevation of 3757.5 ft., is also connected with the smelter town by 4 miles of standard-gage track belonging to the same company. In the vicinity of Jerome, the great southern escarpment of the Arizona plateau forms the northeast side of the Verde valley. The general level of this escarpment, or "rim" as it is called, is about 4000 ft. above the Verde River. A corresponding elevation is reached only by the highest peaks of the Black Hills, which bound the valley to the southwest. Two lava-capped mesas, known as "Mingus" and "Woodchute mountains," form the culmination of the Black Hills. That part of the Old Black Hills mining district which lies between their crests and the Verde River is now known as the Jerome, or Verde, mining district. If the smelter towns are not included, a strip about 7 miles long by 3 miles (11 by 5 km.) wide, embraces practically all of the properties belonging to the Jerome district. T-bis strip lies parallel to and along the side of the Verde valley from First View, 2 miles northwest from Jerome, to the

Jan 1, 1925

By J. Stuart Warner

The reaction COSO4(c)?CoO(c) + So3(g)was investigated from 950° to 1170OK by two different methods. The sulfate was decomposed in a previously evacuated space and Pso3 calculated from the measured total pressure on the assumption that Pso2 = 2P02. The sulfate was also decomposed under conditions of controlled oxygen pressure. The difference in the results obtained by these experimental techniques is attributed to the effects of thermal diffusion. The standard free energy change for the reaction is given by ?F° = 66,292 + 17.273T log T — 102.57 T (±300 cal/mole). It was found by DTA and X-ray diffraction studies that CoS04 undergoes a transformation at 617°± 4°C. THE metal sulfates may decompose directly to the oxide or pass through one or more intermediate basic sulfates. Provided the oxidation state of the cation does not change in the course of the reaction, the decompositions may be described in general as follows: One of the major objectives of this work was to determine the validity of Eq. [3] as much of the existing sulfate data has been calculated from decomposition pressures by Eq. [5]. The type of apparatus generally used for these measurements consists of a reaction vessel at high temperature which communicates with a manometer at room temperature. A temperature gradient is thus imposed on the gas phase causing thermal diffusion which brings about a partial separation of the different gaseous species present in an initially uniform mixture. The basic theory and experimental facts associated with this phenomenon are summarized elsewhere1 and it will suffice to state here that the heavier gaseous species diffuses down the temperature gradient. Since thermal diffusion destroys the homogeneity of the gas phase, it is not necessarily true that Pso2 = 2Po2 in any particular volume element of the gas phase even though there are twice as many molecules of SO, as of 0, in the gas phase considered as a whole. Under these conditions, the decomposing sulfate gives off SO, which dissociates to SO, and 0, and all three species are distributed throughout the system by thermal diffusion. However, the sulfate will continue to dissociate until the equilibrium value of PSo3 is established on the hot zone. Thermal diffusion may distribute SO2 and O2 throughout the gas phase in any way so long as Eq. [2b] is satisfied in the hot zone; i.e. Pso2(Po2)1/2=K3pso3. When a steady state is reached, we can characterize Pso2 in that

Jan 1, 1962

By E. W. Davis

On the eastern end of the Mesabi Range, in Northern Minnesota, is a large formation of siliceous rock which contains bands and fine grains of magnetite. The magnetite comprises about 35 per cent. of the rock, the remainder being chiefly quartzite and iron silicates. The rock has a specific gravity of 3.4, a hardness of 7, and is extremely tough. This large deposit was located early in the history of the Minnesota iron-ore industry but has not been utilized because of its low percentage of iron as compared with the other Mesabi ores, and because of the difficulty and expense of any milling treatment that would concentrate the iron. An investigation, begun about 3 years ago, shows that the magnetit,e is finely disseminated throughout the entire formation and that there are bands or lenses of higher- and of lower-grade material, in which the magnetite and silicate are intimately mixed. As a result, the scheme of milling adopted must include a fine-crushing plant. As it is necessary to crush to 200 mesh in order to produce the desired grade of product, the fine crushing is one of the largest items of expense and for this reason has been given detailed study. It is the purpose of this paper to present some of the results of the work on fine crushing, as to both theory and practice. Previous to fine crushing the part of the rock that contains little or no magnetite can be discarded by magnetic concentrators after each reduction in size in the dry-crushing plant. This-makes it possible to fine-crush a minimum of rock and also establishes the feed to the ball-mill as below % in. The fine-crushing problem, then, consists simply in crushing the rock from 1/4 in. to 200 mesh at the minimum expense. Equipment of the Testing Plant In order to determine, among other things, the operating conditions of the ball-mill when working on this rock, a test mill of about 300 T. daily capacity was erected at Duluth, Minn. The fine-crushing plant contains a Hardinge 8-ft. by 22-in. (2.4-m. by 55.8-cm.) conical mill, a 6- by 27-ft. (1.8 by 8.2-m.) Dorr duplex bowl-type classifier, a 41/2-ft. (1.37-m.) standard Akins classifier, and the auxiliary machinery necessary to handle the products. Each machine is driven by an individual motor, each of which is provided with meters for measuring the power required. Over 150 tests have been made in the ball-mill, varying in

Jan 1, 1920

By G. E. Agar, P. L. De Bruyn

Flotation is an important example of a chemical process by which one solid may be separated from other solids in a mixture. In contrast to other chemical processes such as leaching and smelting, separation of solids by flotation is characterized by the establishment of contact among three phases. In froth flotation these three phases are the solid to be floated, an aqueous electrolyte solution, and a gas, usually air. The gaseous phase may be re- placed by an organic liquid phase or this organic liquid phase may be added deliberately as a fourth phase. To gain an understanding of the flotation sys- tem, it is therefore necessary to study the chemical and physical properties of surfaces and to establish the effect of changes in composition of the bulk phases on the nature of the three interfaces, solid-gas, solid-liquid and liquid- gas. With very few exceptions inorganic solids are completely wetted by the aqueous phase so that the first step in flotation is to replace partially the solid-liquid interface by a solid-gas interface. This displacement process is accomplished by the addition of suitable reagents to the aqueous phase. The resultant surface reaction leaves the solid surface with a hydrophobic film. Although the primary objective of these reagent additions is to alter the solid surface, chemical reactions between the added chemical species and those species derived from the dissolution of the solids are unavoidable. For a quantitative description of such reactions and for the determination of the feasibility of a proposed reaction, we rely on available thermodynamic data. The application of thermodynamic principles to flotation systems has contributed much to our understanding of flotation. The changes in thermodynamic properties such as free energy and chemical potential are accessible to experimental determination, and a knowledge of their magnitude allows us to develop and predict a chemical mechanism oy which a solid becomes floatable. In this chapter it will be assumed that the reader has a working knowledge of the general principles of thermodynamics and solution chemistry. A complete discussion of the physical chemistry of flotation must include a treatment of solubility relations, complex ionic equilibria, reaction potential, and the elements of electrochemistry, especially those dealing with reversible electrical cells. The principles related to all of these features of the flotation system will not be discussed in this chapter. The discussion will be concerned with the thermodynamics and electrochemistry of surfaces and the electrical double layer model of the interface.

Jan 1, 1962

By Pei-Yung Huang, C. R. Hayward

In order to understand more fully the actual state of copper lost in copper reverberatory slags, a systematic study on the various reactions between certain metallurgically important copper compounds and synthetic copper reverberatory slags was undertaken by the authors in the Massachusetts Institute of Technology from 1942 to 1945.1 This paper, being only a part of the work completed, is published first because it furnishes the first experimental evidence of the formation of copper ferrite from reactions between cuprous oxide and copper reverberatory slags. Choice of Slag Composition The choice of the composition of the synthetic reverberatory slags to be used was much deliberated. Commercial copper reverberatory slags, neglecting minor constituents, avcrage :2,3 45 to 47 pct FeO; 38 to 42 pct SiO2; 5 to 6 pct CaO; 5 to 6 pct A12O3. Since the ternary system CaO-FeO-SiO2 has been much more thoroughly studied4 and is much simpler than the quaternary system CaO-FeO-SiO2-Al2O3, it was decided to use calcium oxide only instead of calcium oxide and alumina in the synthesis of the slags. Fortunately the composition of the average copper reverberatory slag with calcium oxide substituted for alumina lies in the region of the phase diagram clarified by the investigators. A survey of the phase diagram (Fig I) suggested two compositions: (I) the composition near the invariant point S, which is made up of 46 pct ferrous oxide, 37 pct silica and 17 pct calcium oxide; (2) the composition near the invariant point K with 45.5 pct ferrous oxide, 43 pct silica and 11.5 pct calcium oxide. Composition S has the advantage that the percentages of ferrous oxide and of silica almost exactly coincide with those of the average reverberatory slags, and that of the lowest melting point (1093º ± 3ºC) in the whole ternary system. The calcium oxide content, however, obviously is high. Since we have already substituted calcium oxide for alumina, and since calcium oxide is more basic than alumina, the fulfil- slag is decidedly too basic. Composition K has a slightly higher melting point (1150O + 3OC) and a higher silica content, but this is preferred because the slag is saturated with respect to tri-dymite and thus should not corrode siliceous refractories. Besides, it is less basic than composition S. The filial slag composition, therefore, was chosen to

Jan 1, 1949

By C. H. Brehaut, W. N. Taylor, R. S. Douglas, H. A. Shannon

The name for this new method of mining is derived from a composition of Horizontal, Radial, Diamond, and the drilling is from raises. This method, worked out at Copper Mountain, B.C., is believed to be distinct in many features from other known methods. Geology The ore deposits of copper Mountain are related to the copper Mountain stock, which is an ideal example of the magmatic in in that it grades from a basic gabbro at the border to an almost pure orthoclase-pegmatite core. The "contact" ore bodies lie along the contact of the gabbro and an older complex of volcanies and sediment termed the "Wolfe Creek" formation. These older rocks, collectively referred to as greenstone, had been fractured prior to the intrusion of the stock and were highly metamorphosed during the intrusion. The "outlying" ore bodies are somewhat different in character, though a similar structural pattern prevails. It is believed that a deep-seated magma of which the Copper Mountain stock is a part has given off solutions that caused a partial graniti-zation of some of the volcanies. Mining problems are caused further by the presence of a series of soft barren dikes intersecting a great many of the ore bodies, which are about 30 ft. wide, strike north-south, and dip at a high angle to the east; and are further increased by the presence of four major faults, with a great many minor faults running at acute angles from them. Fault zones average 8 ft. in width, and consist of gouge and breccia. They strike northeast and dip steeply to the northwest. The faults cut through the ore zones and the dikes, and continue on into the gabbro. The general mine structure with two types of Ore bodies, bounded by weak walls and intersected by several series of soft barren dikes and major faults, indicated that a method of mining should be devised that would remove the ore at low cost with the least disturbance of the waste walls and dikes. , Theoretical Estimate of Stoping with the Diamond Drill After consideration of various methods, the most suitable one appeared to be horizontal drilling to break the ore in a manner similar to shrinkage stoping Drilling vertical holes or rings from a bench and breaking to an open slot was eliminated because the weak walls of ore, dike or waste would not permit this type of opening. It was then decided to drill either horizontal radial holes from a raise or parallel holes from a slot cut across the ore body. After discussion, the latter method was believed unsuitable, for the following reasons:

Jan 1, 1946

By C. E. Brown, D. C. Helms, O. W. Shimer

The early history and progress of anthracite stripping, from the first known operation at Summit Hill in 1821 through 1917, was covered in 1917 in a paper by J. B. Warriner,1 then chief engineer, now president of the Lehigh Navigation Coal Co., and stripping methods between 1917 and 1931 were the subject of a paper in 1931 by H. H. Otto,2 of the Hudson Coal Company. . The present paper is confined to methods used today on pitching beds in the western middle and southern anthracite fields. The anthracite industry, not unlike other large industries, had been affected for more than a decade by the adverse economic conditions of the country. High development and operating costs and the competition of other fuels caused some collieries to be closed temporarily, while others were permanently abandoned. In order to obtain a larger tonnage at a lower development cost, a larger stripping program was put into effect by many companies. In 1932, stripping operations accounted for 3½ million tons, or 7 per cent of the total anthracite production of the year, and in the next 10 years this increased until more than double that amount, or over 14 per cent of the total production, was from stripping operations. These figures represent the total production of the entire anthracite region; unfortunately, the figures have not been separated from the different fields. There are some operations in the southern and middle western fields where the tonnage obtained from stripping amounts to more than 30 per cent of the total colliery production. Stripping operations on pitching beds present difficult problems. Owing to the various geological formations, a different situation is encountered in nearly every stripping pit; and because of the proximity of an overlying bed, it is very often necessary to load the entire overburden into trucks and haul it out of the pit to the dump area, although back filling and casting is done whenever possible. Beds in the areas described in this paper pitch from 20" to 90 and some are on the invert in many sharp folds and faults. Planning The more successful strippings have had adequate engineering analysis. Some work is being done without much preliminary planning, but many strippings would have failed completely if close supervision by the engineering department had not been . maintained. Locations of stripping areas are determined as follows: I. By constructing cross sections over the areas and drilling sufficient diamond-drill or well drill holes to study the strata and to determine the location, thickness

Jan 1, 1944

By Earl E. Hunner

At the end of the year 1914, the main North Star incline shaft had reached the 6300-ft. level, and encountered a vein dipping southwest, or exactly opposite to the North Star. Subsequent development failed to find the North Star vein continuing beyond this intersection, so the 6300-ft. level proved to be the bottom level of the North Star vein, and very little ore was found below the 5300-ft. level. The new vein was called No. 1 vein. Development work carried on for the next 10 years discovered a third vein, called No. 2, branching off about 1200-ft. to the east, with a north-south strike and a west dip. No. 1 vein had about 900 ft. of backs above the 6300-ft. level before reaching the boundary line of the property, and No. 2 vein died out about 800 ft. above. During these 10 years, the money spent on development was increased from $54,000 in 1914 to $184,000 in 1924; the ore reserves steadily diminished until it was evident that the mine must close down unless by some means a large tonnage of ore could be developed quickly. Problems and Plans In order to understand what follows, the general layout and equipment of the mine as it was at this time must be briefly described. Fig. 6 indicates the relative positions of the shafts. The North Star (incline) shaft extends from the surface to the 6300-ft. level following the vein, with an average pitch of 26", between limits of 11" to 38. There is at present no hoisting equipment in this shaft between the 2700 and 4000-ft. levels. The Central (vertical) shaft connects with the North Star shaft at the 4000-ft. level, at a vertical depth of 1600 ft. In two compartments of this shaft 4-ton skips made a vertical turn of 15 ft. radius at the 4000-ft. level and ran down the incline to the 6300-ft. level. A cage operated in the third compartment to the 4000-ft. level. Men and supplies were transferred to a truck running on a third track down the incline shaft to the 6300, operated by an electric hoist on the 4000-ft. station. When the shifts changed double-deck cages were substituted for the skips in the vertical shaft, but they stopped at the 4000-ft. station. The truck to 6300 carried 22 men, and it required 8 min. to make a trip down and back. This arrangement was satisfactory as long as a fairly large proportion of the work was

Jan 1, 1930

By C. H. Brehaut, W. N. Taylor, R. S. Douglas, H. A. Shannon

THE name for this new method of mining is derived from a composition of Horizontal, Radial, Diamond, and the drilling is from raises. This method, worked out at Copper Mountain, B.C., is believed to be distinct in many features from other known methods. GEOLOGY The ore deposits of Copper Mountain are related to the Copper Mountain stock, which is an ideal example of the magmatic differentiation in that it grades from a basic gabbro at the border to an almost pure orthoclase-pegmatite core. The "contact" ore bodies lie along the contact of the gabbro and an older complex of volcanics and sediment termed the "Wolfe Creek" formation. These older rocks, collectively referred to as greenstone, had been fractured prior to the intrusion of the stock and were highly metamorphosed during the intrusion. The "outlying" ore bodies are somewhat different in character, though a similar structural pattern prevails. It is believed that a deep-seated magma of which the Copper Mountain stock is a part has given off solutions that caused a partial granitization of some of the volcanics. Mining problems are caused further by the presence of a series of soft barren dikes intersecting a great many of the ore bodies, which are about 30 ft. wide, strike north-south, and dip at a high angle to the east; and are further increased by the presence of four major faults, with a great many minor faults running at acute angles from them. Fault zones average 8 ft. in width, and consist of gouge and breccia. They strike northeast and dip steeply to the northwest. The faults cut through the ore zones and the dikes, and continue on into the gabbro. The general mine structure with two types of ore bodies, bounded by weak walls and intersected by several series of soft barren dikes and major faults, indicated that a method of mining should be devised that would remove the ore at low cost with the least disturbance of the waste walls and dikes. THEORETICAL ESTIMATE OF STOPING WITH THE DIAMOND DRILL After consideration of various methods, the most suitable one appeared to be horizontal drilling to break the ore in a manner similar to shrinkage stoping. Drilling vertical holes or rings from a bench and breaking to an open slot was eliminated because the weak walls of ore, dike or waste would not permit this type of opening. It was then decided to drill either horizontal radial holes from a raise or parallel holes from a slot cut across the ore body. After discussion, the latter method was believed unsuitable, for the following reasons:

Jan 1, 1945

By F. B. Foley

A description of tests made by a committee of the Division of Engineering of the National Research Council to determine whether velocity of impact affects the formation of Neumann bands. FOREWORD Not infrequently, in the case of the failure of a metal structure, such as a bridge, tank, airplane, gun carriage, etc., a doubt arises whether the failure was due, among other causes, to the effect of an overburden or of an explosive maliciously or accidentally exploded in contact with it. As few means of discriminating between the effects of these causes have been known and these means were of a rough qualitative character, it was surprising to find in the record of an inquiry into the failure of a structure, the statement that, by observations on the Neumann bands in the structure, it was possible to determine the speed of breakage in the metal and to distinguish between a fracture caused by an explosion and one caused by an ordinary mechanical strain. Such a criterion, if well proved, was of value to the explosives experts of the Bureau of Mines, the Ordnance Department of the United States Army and the Bureau of Ordnance of the United States Navy, but search showed no record of it in the literature nor did any of the metallurgists and metallographers consulted know of any data on which such a criterion could be based having been assembled and critically studied. However, it was universally held that there was promise in the idea and that there existed unique facilities at the Bureau of Mines Experiment Station at Pittsburgh for the preparation, under definitely known conditions, of material for study by metal-lographers, and that material of value might also be obtained from the proving grounds of the Army and Navy. Therefore, with the approval of the Director of the Bureau of Mines and the Chiefs of Ordnance of the War and Navy Departments, the Chairman of the Division of Engineering of the National Research Council, on April 12, 1919, appointed a Committee on the Investigation of Neumann Bands, which has since been engaged in this study and which, after some slight change in its personnel, consists now of the following: Dr. William Campbell, Advisory Metallurgist, Bureau of Ordnance, Navy Department. F. B. Foley, Metallurgist, Bureau of Mines. S P. Howell, Explosives Engineer, Bureau of Mines. Comdr. O. M. Hustvedt, U. S. N., Bureau of Ordnance, Navy Department. Dr. Charles E. Munroe, Chief Explosives Chemist, Bureau of Mines.

Jan 9, 1922

From the historical account of the coal industry set forth in the preceding pages the reader will have learned that coal is extremely widely spread throughout the United States, and in most places it has been easily found, that it has been remarkably easy to develop, and, where the deposits were available to streams on which it could be transported to markets, it was opened almost as soon as the country was settled. Such were the mines along the James, Susquehanna, Monongahela, Ohio, Kentucky, Cumberland and Big Muddy Rivers. In other localities where such easy means of transport were not convenient, the early production was confined to local use, or to places to which it could be hauled by wagons, but everywhere small mines were established almost as soon as the country was settled, and these increased both in number and size as the available markets grew. These early mines were nearly always opened by local people, and the industry was so far-flung that its growth attracted little or no attention excepting when a labor disturbance or a breakdown in transportation occurred. Had it been concentrated in a few places as most metal industries were, or as the petroleum industry was for many years after it started, it is probable that much better records of its progress would have been kept. When the canal era began, in the eighteen twenties, coal was at first not considered as a valuable source of freight revenue, and much surprise was expressed that the receipts of the Schuylkill Canal were very largely from coal after the first few years, as was the case of the Union Canal though not to such an extent. Even the early railroads to the coal fields did not realize the extent to which the products of mines would figure in their revenues, and it was many years before railroads were built practically solely for prospective coal traffic; indeed this did not happen until some years after the Civil War. Map 12 shows the extent of the canal system in the United States in its relation to the coal fields at the time of its maximum development. It will be seen that only the coal areas in Pennsylvania, Ohio, and to a very small extent in Indiana were able to ship by canal at all, and that only a very small portion of such fields could do so. Looking backward, it is hard to understand the great hopes entertained of canals by their pro-

Jan 1, 1942

By Edward J. Felten

THE oxidation resistance of chromium and Fe-Cr alloys is increased by small additions of yttrium or other rare earth metals.1,2 In addition, the presence of the additives increases the resistance of these alloys to spalling of the external oxide above 1000°C. This ability to prevent spalling appears to be related to the formation of an internal oxide in alloys containing yttrium or the other rare earth metals. The internal oxide is found in the form of a filamentary growth concentrated in the grain boundaries of the unoxidized metal. In alloys containing yttrium this internal oxide was found to be either Y2O3 Or YCrO3.' In an earlier report,' the growth of the external oxide was found to be controlled by diffusion through the oxide layer. Thermal balance studies were used to establish the kinetics of external oxide formation. In this related study the rate of growth of the internal filamentary oxide has also been studied. The results obtained indicate that the growth of the internal oxide is also diffusion controlled. Internal oxidation appears to occur in at least two ways. Using copper alloys containing small amounts of silicon, Rhines 3 found that between 600" and 1000oC a subscale was formed which consisted of particles of silica dispersed in the metallic matrix. The conditions requisite to internal oxidation as suggested by Rhines are: 1) a solubility of oxygen in the alloy, 2) the ability of oxygen to diffuse at an appreciable rate into the alloy, 3) the possibility of the formation by the alloying element of an oxide more stable than the oxide of the host metal, and 4) a relatively low solubility of this oxide in the metal. These conditions are fulfilled for a number of other copper alloys.3-5 Furthermore, the relationship between the depth to which the internal oxide is formed and the oxidation time follows the simple parabolic law.6 The activation energy for copper containing small amounts of silicon, germanium, iron, or manganese as calculated from Rhines data6 is about 48 kcal per mole. In some cases the internal oxide is found segregated at the grain boundaries, especially near the the surface. Observations of this type of oxide formation have been reported by stead,7 Rhines,* and Lustman. Evans" cites these cases as examples of the simultaneous movement of anions and cations during oxidation. Furthermore, Evans states that the presence of a minor constituent with a high

Jan 1, 1962

By E. R. Darby

When bronze is melted in open-flame furnaces a considerable amount of slag is formed during the melting operation. This slag may be incidental to the melting practice or it may be formed intentionally by slag-forming constituents added to the furnace charge to provide a covering for the melted metal When this slag is removed from the furnace, it contains, in addition to the free metal carried with it by the skimming operation, a considerable amount of metal as oxide either combined or mechanically mixed with it. By the usual crushing and washing methods this metallic oxide generally goes to the tailings pile, so diluted by other foundry waste that its recovery is economically impossible. If this furnace slag is not inixed with other foundry waste, practically all of the metal oxide as well as the free metal may be recovered by some form of smelting operation. The first side of the problem to be considered is the amount of slag produced. It may not be true that the majority of foundries have no record of the amount of slag made, but it is certainly true that a very great many either have no record or pay little attention to it. Usually the collection of furnace slag for one normal week's run will be sufficient to determine the advisability of its being handled in a different manner from the crushing and concentrating generally employed. If a large representative sample of the accumulation is crushed and concentrated and the recovery compared with the results of a careful assay, it will be found that a large percentage of the total metal content was lost in the washing process. The percentage of loss will be greater than the metallic oxide content, as some free metal in extremely fine particles always is carried away in grains of slag. From such a comparison, it may be determined whether or not the slag should be melted, sold or treated as before. An example of an actual test of this kind is given below. In a foundry melting 50,000 to 80,000 Ib. of bronze per day, a week's accumulation of furnace slag and ladle skimmings amounted to 16,500 1b.—over three times as much as the foundry superintendent estimated before the accumulation was made. Two tons of this material was taken as representative of the lot. This was ground through a thoroughly cleaned ball-mill and concentrated over a Wilfley table. The