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By M. Garfinkle

LITERATURE data1'3 have suggested that the resolved shear stress necessary to initiate slip in a particular slip system in tungsten single crystals is strongly dependent on the tensile orientation of the crystals. This orientation dependence is contrary to the critical resolved shear stress criterion. Because the Schmid factors for slip on particular {110}( I l l), {112}(Ill), or {123}(111) slip systems in a (100) oriented tungsten crystal are virtually identical to the respective Schmid factors for a (110) oriented crystal, slip should initiate on the same system in both orientations, independent of the values of the critical resolved shear stresses. Thus, it would be expected that, for a critical resolved shear stress model to hold, the proportional-limit stress for these two orientations must be similar. The expectation of similar proportional -limit stresses has not been borne out by reported experimental observations. Rose et nl.' reported that the proportional-limit stress for crystals near the (110) orientation was 92,000 psi, approximately three times that of (100) oriented crystals (31,000 psi). Garlick and probst2 reported greater than a fourfold difference. Most recently, Beardmore and ~ull~ examined tungsten crystals with orientations along the edge of the standard stereographic triangle and showed a gradual increase in proportional-limit stress from 28,000 psi for the (100) orientation to 93,000 psi for the (110) orientation. Again, the difference was more than threefold. As has been emphasized by these investigators, the apparent gross variation of proportional-limit stress with orientation precludes any meaningful analysis of slip mechanics using a simple critical resolved shear stress relation. The purpose of the present communication is to report that, by use of refined strain-measuring techniques, the proportional-limit stresses for the (100) and (110) orientations are quite similar. Obviously, to determine accurately the value of the proportional-limit stress, which is by definition the stress at which nonlinear stress-strain behavior first occurs, the method of measuring strain must be capable of high accuracy. The investigators cited used as the measured elongation the crosshead displacement of the tensile machine. This method, however, introduces extraneous strain contributions from the load train. Because these extraneous contributions are considerable in the initial portions of loading (as has been well-documented in the literature) the proportional-limit stress measurement, dependent on the deviation of a line from linearity, can be significantly affected. During the course of a recent study, room-temperature strain was measured directly on several (100)-and (110)-oriented tensile specimens axially oriented within 5 deg of the desired direction. Each specimen was fitted with two foil strain gages mounted on opposite sides of the 0.08-in.-diam reduced section. The tensile specimens were loaded at a crosshead speed of 0.01 in. per min on an Instron Universal Testing Machine. Strain was monitored through an external bridge circuit in which the specimen strain gages comprised two arms. The bridge output was recorded against the applied load. A strain of 2 x 10"" could be readily detected by this method. Some specimens were subjected to conventional loading while the load for others was cycled between zero and continually increasing values until the first deviation from linearity was observed. The results of these methods were similar, and the stress at which nonlinear behavior occurred was considered the proportional-limit stress. The proportional-limit stresses measured in this study are approximately 5000 and 6500 psi for the ( 100) - and ( 110) -oriented crystals, respectively, a range of values almost a magnitude less than those reported previously. Representative curves are presented in Fig. 1. Because of the nature of the measurements made, the point of departure of a line from linearity is often debatable. While the load-strain curve for the (100) -oriented crystals made a relatively abrupt deviation from linearity, the curve for the (110) -oriented crystal was more gradual, so that the actual point of tangency may be at a lower load than reported. Thus, the difference in values for the proportional-

Jan 1, 1967

By Wilbur Jurden

THE writer's first experience with a nonferrous reduction plant of great magnitude was at the Washoe reduction works of Anaconda some 35 years ago. Here was a plant which had been planned with remarkable skill and foresight considering the time and the state of development of copper-plant practice in the year 1902. The designer utilized topography to fullest extent to provide proper sequence of operations and, what is most remarkable, to leave adequate space for future developments, most of which at that time were unknown. However, the practice then was to locate the various units of the reduction works at the most advantageous points of the existing terrain with little regard for tramming or other auxiliaries and then connect these various units by the essential trackage, conveyor systems, piping, etc., as the need developed. This occasionally led to undesirable track arrangements, sharp curves, and steep grades, especially when it became necessary to extend various portions of the plant. Conveyor systems also became rather complicated, running as they did at various angles, and such items as piping and electrical distribution were often found to be in the wrong place, entirely inadequate in size, or awkwardly arranged for any kind of extension. This condition was not peculiar to Anaconda, for all copper plants at that time were built in the same manner and it was the constant association with these difficulties which, in the year 1925, influenced the layout of the Andes Copper Mining Co. plant. In that plant all trackage was laid out straight and level, all conveyors at right angles to each other with minimum length and number of transfers. All buildings were placed parallel and the main structures were complete for all purposes so that auxiliary buildings and dog houses would not be added later. Piping and electrical work was provided for in the original layout and carefully designed to avoid additions and alterations, and careful study given to every movement of material throughout the entire plant so that it would be accomplished with the least possible effort. Naturally it was hardly expected to attain all these objectives perfectly but our efforts did succeed in creating a plant which was unique and outstanding for its time-1927. It was also most gratifying to find that these design principles contributed to considerable savings starting right in the drafting room, carrying through the construction and ultimately yielding savings in operations and manpower. Not only that, but such a plant gives the observer an impression of symmetry and order, is more attractive to the workmen, and unquestionably eliminates many accident hazards. However, the Andes plant buildings were fitted to the existing terrain instead of having terrain created to fit the buildings-an item which we found advantageous to correct on the next large plant. At Morenci in 1939, all of the desirable features of the Andes plant such as parallel buildings, etc., were incorporated; but we went one step further-power shovels were brought in to make the terrain fit the reduction works. The result at Morenci is well-known and needs no elaboration here, but the success achieved by the design methods used for this and previous plants naturally influenced and guided the layout of the Chuquicamata sulphide plant which is the largest yet conceived. Chuquicamata Plant Design At Chuquicamata several factors not encountered previously complicated the problem to a great extent. The most desirable location for the smelter would allow smelter gases to blow directly into the open-pit mine already producing 60,000 tons of oxide ore per day and employing 1550 men. This, of course, would be a serious condition and, therefore, we were forced to move the smelter to a less desirable location but followed our previous experience at Morenci and made the terrain fit the job. The most difficult problem, however, was the provision for receiving various types of ore both by rail and conveyor. These consisted of: 1-Sulphide-bearing residue from the stockpile from which oxide copper had previously been leached. 2-Sulphide-bearing residue coming direct from the leaching vats. 3-Sulphide ore crushed at the existing crushing plants and hauled to the concentrator in cars. 4-Sulphide ore from the new crushing plant adjacent to the concentrator. 5-Sulphide ore obtained

Jan 1, 1952

By R. M. Genevray, M. B. Bever, E. J. Suoninen

THE martensitic transformation in the ß-phase of the Cu-Zn system has been the subject of several investigations. The transformation is known to be reversible and to be affected by stress. Its temperature range has been determined as a function of composition. In the investigation reported here, the effect of tensile stresses on the transformation was investigated quantitatively. Some information was also obtained on the thermoelastic behavior of the martensite formed in the first stages of the transformation. Most of the experiments were done with alloy E of an earlier investigation;' this alloy analyzed 60.49 pct Cu and 39.51 pet Zn by weight. The methods of shaping and heat treatment were also essentially the same as those previously used. The stress was applied to the specimen immersed in a cooling liquid. The transformation was followed by measuring the electrical resistance with a Kelvin bridge and the elongation with a cathetometer. Fig. 1 shows the M, temperature as a function of stress. Resistance and strain measurements gave essentially identical values. The results suggest a roughly linear relation between M. and in the range investigated, up to 12 kg s mm". At higher stresses, plastic deformation begins to interfere seriously with this relationship. The increase of M, with stress is consistent with published work on the effect of stress on the martensitic transformation. The slope of the curve, 4°C per kg mm ", is of the same order of magnitude as the corresponding value calculated for steel.' Fig. 1 also shows the difference, AM, between the temperature of 50 pct transformation on cooling, as measured by changes in length, and that of 50 pct reverse transformation on heating. This difference, which may be considered a measure of the hysteresis, increases with stress; the decrease at highest stresses is probably associated with plastic deformation. Preliminary work using only resistance measurements was done with an alloy containing 60.15 pct Cu and 39.79 pct Zn by weight. The results indicated higher values o-F M, in agreement with the known variation of M. with composition.' The effect of stress on M, (2°C per kg mm-') was of the same order of magnitude as that shown in Fig. 1 for composition E. An increase in hysteresis with stress was also found. The following experiment was made in order to investigate a partially transformed structure. A specimen of alloy E was cooled to — 85°C under a stress of 4.7 kg mm-'. Under these conditions, the martensitic transformation started but did not go to completion. The stress was then released and the specimen cooled to — 105°C. Fig. 2 shows the measured elongation c. The first change in the slope of the curve indicates the beginning of the transformation under stress. Removing the load at —85°C caused a decrease in length to the value corresponding to the elastic elongation of the parent phase resulting from the applied stress. Hence, the marten-site formed in the first part of the experiment apparently disappears completely and without hysteresis upon the release of the stress. The increase in length on further cooling indicates renewed formation of martensite. These conclusions are consistent with the concept of "thermoelastic" martensite," which has been confirmed by test." Acknowledgments The authors are greatly indebted to Professor M. Cohen for his advice and encouragement. They also thank F. Paxton for assistance. Thanks are due the American Brass Co. which supplied the alloys. References E. Kaminsky and G. V. Kurdjumov: Zhur. Tekhn. Fiziki SSSR (1936) 6, p. 984. A. B. Greninger and V. G. Mooradian: Trans. AIME (1938) 128, p. 337. "J. E. Reynolds, Jr. and M. B. Bever: Trans. AIME (1952) 194, p. 1065; Journal of Metals (October 1952). 'A. L. Titchener and M. B. Bever: Trai~s. AIME (1954) 200, p. 303; Journal of Metals (February 1954). " 3. A. Kulin, M. Cohen, and B. L. Averbach: Trans. AIME (1952) 194. p. 661; Journal of Metals (June 1952). "J. K. Pate1 and M. Cohen: Acta Metallurgica (1953) 1, p. 531. 'C. Crussard: Comptes Rendus (1953) 237, p. 1709. ' G. V. Kurdjumov: Zhur. Tekhn. Fiziki SSSR (1948) 18, p. 999. G. V. Kurdjumov and L. G. Khandros: Dokl Akad. Nouk. SSSR (1949) 66. p. 211.

Jan 1, 1957

By R. E. Reed-Hill

A simple accurate graphical method for orienting fcc crystals using (111) slip traces on a single surface is described. Solutions placing the pole of the surface in a unit stereographic triangle are obtained using a Wulffnet, a (111) standard projection showing the four (111) planes, and a (111) standard projection ruled into unit stereogmphic triangles. The multiplicity of solutions is described in relation to this method. AN accurate, rapid single-surface technique for orienting fcc crystals, using (111) slip or twin traces, has been a matter of continuing interest, as evidenced by the periodic appearance of papers on this subject.'-4 This note presents a new method possessing some advantages over those previously reported. It is an analog to the technique for hcp metals5 where the solution is based upon a stereographic rotation of the specimen surface on a (0002) standard projection that contains the planes of the possible features that might account for the observed traces. In the hexagonal solution, the axis of rotation is the basal-plane trace in the surface. In the fcc case, the axis of rotation is one of the observed ( 111) slip or twin traces. It has been observed that the solution is greatly facilitated if the indicated movements of the (111) traces during the rotation of the surface are plotted directly on a paper copy of a Wulff net as shown schematically in Fig. 1. The trace selected to represent the axis of rotation is assumed to fall at the north-south poles of the Wulff net. The directions of the other traces are then plotted as points at the proper angles around one-half of the basic circle of the net. The loci of points corresponding to the movements of these traces (directions) during a 90-deg rotation of the surface, about the polar axis of the net, are then easily and accurately drawn by following the appropriate small circles on the Wulff net. The solutions to the problem are obtained with the aid of a (111) standard projection of the (111) planes, see Fig. 2. This is normally a Xerox copy of a carefully drawn master made on high-quality tracing paper. The (111) standard projection is superimposed over the Wulff net and rotated about its center until the points of intersection of the loci of surface traces, drawn on the Wulff net, and the great circles representing the (111) traces, all lie along a common meridian of the Wulff net as demonstrated in Fig. 3. This latter great circle represents a possible position of the surface. The pole of this plane is then plotted on the (111) standard projection. Its location in a unit stereographic triangle may be readily obtained by superimposing (over the standard projection) another tracing paper (111) standard projection subdivided into unit stereographic triangles. The result of this last operation is shown in Fig. 4. MULTIPLICITY OF THE SOLUTIONS The maximum possible number of (111) traces on a single grain surface is four. Any one may be selected as the axis about which the surface is rotated. Corresponding to each such choice are three possible surface-pole orientations, related to each other by a simple 120-deg rotation about the pole of the (111) basic circle. This is a direct result of the threefold symmetry of the (111) standard projection. Since each of the four rotations will normally give a different set of three-pole positions, there are, accordingly, twelve possible solutions. Each of these twelve solutions places the surface pole in the same relative position in a stereographic triangle so that all are crystallo-graphically equivalent. Accordingly, only a single solution is required to obtain a basic orientation of the crystal. It should be noted, however, that the orientations obtained by reflecting* any of these twelve orien- *The reflected orientations may be obtained by a 90-deg rotation of the opposite sense to that shown in Fig. 1. This rotation places the correspond-ing trace loci in the lower hemisphere of the Wulff net rather than in the upper hemisphere. tations through the surface are also possible orientations that could account for the surface traces. Thus, a given set of four traces always corresponds to a pair of basic orientations related by reflection through the specimen surface. This ambiguity can only be re-

Jan 1, 1967

By Leslie W. Austin

ROTARY kiln operators will agree that some of the most severe conditions a refractory must stand occur in the hot zone of a kiln burning Portland cement, dead burn dolomite, magnesite, periclase, and similar materials. Frequently the operator is faced with factors beyond his control which drastically shorten the life of refractories. Shutdown due to mechanical failure can be serious if the period is of sufficiently long duration to cause the dropping of coating or the loosening of the lining. A change in slurry can affect the coating and cause ring buildup. A change in type of fuel and its effect upon the flame can cause a shift in location of the hottest zone. Weekend shutdowns or any other interruption can cause the operator trouble and may damage the refractories, since stopping and starting a rotary kiln is certainly more difficult than stopping and starting a motor. Some operators have tried to set an estimate of damage for each shutdown in equivalent days of running time. Conditions affecting the refractory may be roughly grouped in four classes: chemical attack, mechanical stress, thermal shock, and abrasion. Chemical Attack: The drive to obtain maximum production through a kiln demands maximum operating temperatures, temperatures which are limited more by the ringing up or melting of the clinker. This can cause interface temperatures at the junction of coating and refractory which require the use of a basic kiln block to withstand the chemical attack. Chemical changes take place within the refractory itself, especially in chemically bonded or unburned kiln blocks: These changes cause the formation of the ceramic or burned bond. Migrating liquids or fluxes from the kiln charge have an effect within the refractory and lead to mineral or glass formation. The alkalies, sodium and potassium, migrate into the refractory as silicates, chlorides, sulphates or other salts. They may move under capillary action or may be caused to move by volatilization with condensation in the cooler portion. Mechanical Stress: Concentrated stress may be caused by several factors or combinations thereof. 1-The rings of refractories must be kept tight and rigid within the kiln, and this alone demands considerable force to hold the blocks in place. So that the force will not be concentrated, the blocks should fit the circle as perfectly as possible, with the faces in contact overall. 2-As the kiln is heated, thermal expansion takes place at the hot end of the kiln block. Since this disturbs the plane face it too can cause a concentrated stress at the two ends of the block, and shearing stress can be set up within the brick itself because of the difference in expansion between the two ends. 3-If a lining becomes loose and moves in the shell very severe stress can be set up, and as the kiln rotates this load changes and gives the effect of repeated loading. Permanent expansion of the refractory can also cause severe loading. 4-Not least important, flexing of the kiln is frequently the cause of concentrated stresses. Thermal Shock: Thermal shock, the result of heating and cooling too rapidly, occurs on starting and stopping or when a large patch of coating drops, exposing the bricks. Again, its destructive effect is often the result of phase change, liquid to solid or the reverse; dense refractories loaded with glass-forming impurities are particularly susceptible. Thermal shock is a problem with refractories set in the wall or roof of a stationary furnace, and becomes even more serious in a rotary kiln, the tendency to spall being magnified with movement and concentration of stress. Uniform rate of feed and loading insures both better coating and a more uniform stress. Abrasion: If the refractories do not take a coating, abrasion can become a most destructive factor. Movement of the lining in shell or movement of loose blocks causes abrasion, which is also most destructive if the refractories do not take a coating. An analysis of the problem of basic lining for the hot zone reveals, therefore, a number of desirable characteristics: high refractoriness, basic chemical reaction, resistance to spalling, good strength at all stages, ability to take coating, true sizing, volume stability, and abrasion resistance. Increased demand

Jan 1, 1952

By David T. Peterson

Calcium metal was found to deoxidize thorizcm at 1000° to 1200° C. The reaction kinetics were determilled and related to the diffusion coefficients of oxygen in thorium. The solubility of oxygen in thorium, the minimum oxygen concentration, and the diffusion coefficient were determined from 1000° to 1200°C. This firocess results in the lowest oxygen concentrations zohich have been reported for thorium metal. FOR many years it has been known that calcium metal will reduce thorium oxide to thorium metal. This reaction has been the basis for several methods of preparing thorium metal. From the equations giv-by Kubaschewski and vans, ' AF" for the reaction Cao, + Tho,(,) - CaO(,, + Th(,) was calculated and found to be -3.4 kcal at 1000°C, -2.5 kcal at llOO°C, and -2.0 kcal at 1200°c. Thorium is very slightly soluble in liquid calcium, and the solubility of calcium in solid thorium is very low. Consequently these metals would be in essentially their reference states. If thorium containing oxygen were equilibrated with liquid calcium between 1000° and 1200°C, the oxygen content of the thorium would have to be below the solubility limit in thorium. Oxygen is one of the impurities most difficult to remove from thorium and is the most abundant impurity in metal prepared by almost all known methods. Fortunately, oxygen does not have a large influence on the properties of thorium because the solubility in solid thorium is very low. EvenT in thorium containing 100 ppm of 0, particles of thorium oxide can be observed in the microstructure. In view of the incompatibility of thorium oxide and liquid calcium and the low solubility of thorium oxide in thorium, the deoxidation of thorium by this method was investigated. For thorium containing an amount of oxygen well in excess of the solubility limit, the reaction should proceed in the following sequence. The oxygen content of the thorium matrix near the surface would be depleted by the diffusion of oxygen to the surface. At the surface, the oxygen would react with calcium to form calcium oxide. To maintain equilibrium within the thorium, thorium oxide would dissolve to keep the matrix saturated. Consequently, the thorium-oxide particles would disappear first at the surface and then the particle-free rim would grow in thickness. If the rate-controlling step were the diffusion of oxygen through this layer of thorium which was growing in thickness in direct proportion to the amount of oxygen removed, the well known parabolic time law should be observed. If the oxygen concentration at the surface of the thorium and at the inner surface of the deoxidized rim were known, the diffusion coefficient of oxygen in thorium could be calculated from the parabolic rate constant. EXPERIMENTAL PROCEDURE The thorium metal used in this study was prepared by calcium reduction of ThF, by the method described by Wilhelm.' The analysis of this metal is given in Table I. The carbon was determined by combustion, the oxygen by the HC1-insoluble residue method, nitrogen by the Kjeldahl method, and the other elements by emission spectroscopy. A section of this ingot was hot rolled at 600°C to 1/4 and 1/8-in. thick plates. Specimens approximately 7/8 in. square were cut from these plates, and all surfaces of the specimens were cleaned and smoothed by filing with a clean file. Individual specimens were placed in 1-in. diam by 2-in.-long tantalum capsules. Approximately 1 g of clean, high-purity calcium was placed in the capsules and an end closure arc-welded in place. The tantalum capsules were sealed in Inconel crucibles to protect the reactive metals from oxidation. The entire loading procedure was done in a glove box filled with pure argon. The loaded crucibles were placed in a muffle furnace, controlled within 2OC of the desired temperature, for a measured length of time. After the specimen had cooled to room temperature, it was sectioned perpendicular to the large faces and through the mid-point of two of the sides. The sectioned specimen was mounted and polished through Linde A abrasive. The rim which was free of thorium-oxide particles could be clearly observed microscopically as mechanically polished. Twenty measurements of the thickness of the rim were made at equally spaced points far enough from the end of the

Jan 1, 1962

By Carl Altstetter, Roy Adams

Measurements of enthalpy changes and transformation temperatures me reported for the fcc -hcp marten-sitic transformation in pure cobalt single and multi-variant crystals andpolycrystalline specimens. From these measurements the free-energy difference between the two phases was calculated for both heating and cooling and for repeated cycles through the transformation. For heating of single crystals the average value of enthalpy change was 113 cal per mole but on cooling it was 84 cal per mole. This difference is interpreted in terms of lattice imperfections which are introduced during transformation. A calorimeter suitable for measurement of enthalpy changes of athertnal transformations is described. COBALT exhibits an allotropic phase transformation at around 417°C' where an hcp phase (a = 2.507, c = 4.0686, c/a = 1.6228)~ stable at low temperatures transforms to an fcc phase stable up to 1495°C. The transformation has a hysteresis of about 30°C— the exact magnitude of the hysteresis depending on the number of transformation cycles, grain size, impurity content, and prior treatment.'-6 The athermal character of the transformation and surface relief due to transformation shear indicate that the transformation is martensitic. Pole mechanisms for the transformation have been discussed by Sebilleau and Bibrin, Bilby, Seeger,' and Basinski and Christian. These authors have proposed that the cooling transformation is effected by rotation of a Shockley partial dislocation around a pole dislocation with a screw component of 2a/3(111). Such a rotation would change the stacking of close-packed planes from the abcabc fcc stacking to the ababab hcp stacking sequence. This mechanism would predict a lattice correspondence of {111}fcc~~(0002~cp, have been observed. One would expect the heating transformation to occur similarly, involving dislocation movements on the (0002)hcp. However, a high density of pole dislocations and high dislocation ve- locity are required to account for microscopic observations. Furthermore, the high back stress developed on the Shockley partial dislocation as it rotates toward the sessile dislocation associated with it makes it doubtful that this is in fact the transformation mechanism. A mechanism for spontaneous nucleation of partial dislocations1*l2 seems more reasonable. If a single crystal of cobalt is cycled through the transformation but never heated above 600°C, it will remain single."13 However, if the specimen is heated through the transformation and held for a short time at or above 1000°C before cooling, the cooling transformation usually involves all four {lll}fcc habits, resulting in neighboring orientations of hcp differing by 70.5 deg. This condition is termed multivariance and has been observed by Bibring and sebilleau5 and Nelson and ~ltstetter.' The latter authors used single crystals of cobalt grown in an electron beam zone refiner. They proposed that the strongly directional cooling caused the initial cooling transformation to occur on the {lll)fcc most nearly perpendicular to the temperature gradient. Transformation shear on one set of {lll}fcc planes immobilized dislocations on the other {lll}fcc planes involved in the initial transformation. Since there is only one (0002)hcp plane orientation, the product of subsequent heating transformations could have no more than two variants—the original fcc orientation or its twin. If, however, the specimen was annealed at a high temperature, such as 1000°C, there would be relaxation of dislocations into low-energy configurations, and in the absence of a sharp temperature gradient transformation on all four (111 Ifcc planes could occur. If a dislocation mechanism is responsible for the transformation, the perfection of the lattice should affect the reversibility of the transformation. Mutual constraints of neighboring grains in polycrystalline material would inhibit dislocation movement resulting in increasingly sluggish transformation. Indeed, Nelson and Altstetter found the M, to decrease and the A, to increase in going from single crystals to multi-variant specimens to polycrystals. The effect is enhanced by decreasing the grain size of a solid specimen or a powder.6'14 Data on the free-energy difference necessary to initiate the transformation gives information about the driving force of the transformation. This is an important consideration in the understanding of the transformation mechanism. The free-energy change associated with the transformation of the hcp phase to the fcc phase at some temperature T different from the equi-

Jan 1, 1969

By Dirk A. Benham, William A. Warfield

INTRODUCTION Whenever a mine operator is faced with a decision between various mining methods or even equipment selection, he must be assured that his decision will have a positive effect on his economic operating goals. For instance, if a mine operator has a choice of using an LHD mining system or a slusher mining system, then his decision to use either must not only fit within the operational limitations of the mining plan (such as production requirements, physical features of the ore body, mine plant facilities, manpower requirements, etc.), but also assure that the ore is being extracted at the least possible cost. To aid in his decision making process, the operator will commonly compare acquisition costs, operating costs, maintenance costs, and the operational advantages and disadvantages of each system in addition to considering his personal observations and feelings. A decision to select a mining system using only those criteria aforementioned will not necessarily lead the operator down the path to maximizing profits. In addition, the operator should also compare the economic effects of each system upon the overall mining operation. A relatively simple method for determining the economic effects is to calculate and compare the break-even grades for each system under consideration. By implementing such an analysis, the operator can now select a mining system which can operate within the limits of the mining plan as well as extract the ores at the least possible cost, thus yielding maximum profits. This paper will present the geologic and physical conditions of a mine being operated by the Hypothetical Mining Company in the Grants Mineral District. An LHD and Blusher mining system are under consideration and are compared by examining operating costs, equipment availability, and the conventional operational advantages and disadvantages of each. Finally, the economic effects of each system upon the overall mine's operation are compared by utilizing break-even grade analysis. DESCRIPTION AND MINING METHODS The Hypothetical Mining Company has located an ore body in the Grants Mineral District. After extensive surface exploration drilling and cost considerations are evaluated, it is decided to open a mine. The following tasks must now be performed: 1. outline ore body and it's basic characteristics, 2. look at mining methods currently in use in the area, and 3. determine mining techniques and methods for exploitation of the ore body. For the sake of argument, the newly discovered ore body is located towards the western edge of the Grants Mineral Belt. The host rock for uranium mineralization is the Westwater Member of the Morrison Formation. The Westwater Canyon sands are overlain by the Brushy Basin Member of the Morrison Formation and underlain by the Recapture Member, all of Jurassic Age. Westwater Canyon sandstone is a fine to very course grained subarkosic sand with interbedded green and red mudstone. The Brushy Basin Member consists predominantly of greenish-gray mudstones and contains thin lenses of sandstone and a few thin beds of limestone. The Recapture Member consists of alternating thin beds of reddish brown, grayish red, and light gray sandstone and siltstone. The Recapture interfingers with the Cowsprings Sandstone in the area of the mine. The mineralization is based on the groundwater flow within the Morrison Formation. Primary ore distribution is based on three basic parameters: ore migration in a geochemical cell, lenticular mudstone interbedding, and enrichment along fault zones. In areas where two fault systems intersect, ore thicknesses of up to 30 m (100 ft.) are encountered. In those areas where the ore is well defined and thicknesses facilitate full face mining, the mining methods are determined more by ground or operating conditions, equipment availability, and access to working areas than they are in the fringe areas of the ore body where grade control is more important and split-shooting and break-even grade determines the mining technique. If a modified room-and-pillar mining method using systematic development is selected, better control of the ventilation, equipment utilization, and mining plan can be obtained. Figure I shows a typical mining area where the ore is thick,

Jan 1, 1982

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemi-sorption. FLOTATION of quartz is of practical importance as something to be avoided in soap-floating many types of ores. Clean, unactivated quartz is not floated with fatty acids and soaps, such as oleic acid and sodium oleate, in the quantities normally used for flotation. However, data in the literature indicate that almost any multivalent cation will activate quartz if given an opportunity. Thus, a common problem is to prevent activation of quartz by the various inorganic cations inevitably present in flotation pulps. Wark and his coworkers1 have demonstrated the reversibility of the chemical reactions and adsorptions involved in the activation, depression, and collection of the common sulphide minerals. The procedure in much of their work was to bring a mineral surface to equilibrium with solutions of known pH, collector concentration, and activator concentration, and then to test the floatability of the mineral by contact-angle measurement. From the data, graphs were constructed with pH and reagent concentrations as coordinates. These graphs show fields of flotation and fields of nonflotation, separated by narrow transition regions whose locations are shown by so-called contact curves. From the shapes and locations of the contact curves, which roughly separate fields of flotation from fields of nonflotation, a quantitative understanding of the interaction of the reagents with each other and with the minerals often can be deduced. The study of quartz flotation to be described in this paper follows in broad lines the approach of Wark and coworkers. That is, pH, activator concentration, and collector concentration were varied to find equilibrium conditions of flotation and non- flotation, and the results are presented graphically by means of contact curves. However, instead of testing for floatability by measuring the contact angle on a polished surface, a simple vacuum flotation technique was developed and used. Purified oleic acid was the collector and terpineol the frother. Barium activation was studied in some detail, and exploratory studies were made of activation with calcium, aluminum, ferric iron, and stannic tin. Preparation of Materials Quartz: Large lumps of high-grade vein quartz were crushed dry in a cone crusher and rolls. The —20, +28-mesh portion was screened out and used in the subsequent steps. This material was passed through a high-intensity magnetic separator to discard iron, then leached twice with hot concentrated HCl and washed repeatedly with distilled water. The cleaned sand was then wet ground with porcelain balls in a porcelain pebble mill, deslimed repeatedly by settling and decantation to discard —800-mesh material, and again washed with hot HCl followed by distilled water. The resulting stock of quartz was stored under water. Chemical analysis gave 99.8 pct SiO2. Table I gives the size analysis of the quartz used for flotation tests. Calculations from these data, using shape factors given by Gaudin and Hukki9 indicate a specific surface of about 500 cm2 per g. Blank flotation tests in distilled water, and in water with added frother, showed the prepared quartz to be completely nonfloatable and thus indicated the absence of organic contamination. Oleic Acid: The preparation of oleic acid was based on fractional vacuum distillation of methyl oleate2,3 followed by regeneration of oleic acid, and finally fractional crystallization of oleic acid from acetone solutions at low temperatures." The pure oleic acid was stored in a refrigerator. The iodine number of the oleic acid was found to be 90.0 (theoretical 89.93). Oleic acid was used in the form of a dilute water solution of sodium oleate, after preliminary flotation tests showed no effects of form of addition and order of addition of reagents when an adequate conditioning time (that is, 30 min) was provided. Other Reagents: Sodium hydroxide solutions low in carbonate were prepared by first making 1:1

Jan 1, 1951

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemi-sorption. FLOTATION of quartz is of practical importance as something to be avoided in soap-floating many types of ores. Clean, unactivated quartz is not floated with fatty acids and soaps, such as oleic acid and sodium oleate, in the quantities normally used for flotation. However, data in the literature indicate that almost any multivalent cation will activate quartz if given an opportunity. Thus, a common problem is to prevent activation of quartz by the various inorganic cations inevitably present in flotation pulps. Wark and his coworkers1 have demonstrated the reversibility of the chemical reactions and adsorptions involved in the activation, depression, and collection of the common sulphide minerals. The procedure in much of their work was to bring a mineral surface to equilibrium with solutions of known pH, collector concentration, and activator concentration, and then to test the floatability of the mineral by contact-angle measurement. From the data, graphs were constructed with pH and reagent concentrations as coordinates. These graphs show fields of flotation and fields of nonflotation, separated by narrow transition regions whose locations are shown by so-called contact curves. From the shapes and locations of the contact curves, which roughly separate fields of flotation from fields of nonflotation, a quantitative understanding of the interaction of the reagents with each other and with the minerals often can be deduced. The study of quartz flotation to be described in this paper follows in broad lines the approach of Wark and coworkers. That is, pH, activator concentration, and collector concentration were varied to find equilibrium conditions of flotation and non- flotation, and the results are presented graphically by means of contact curves. However, instead of testing for floatability by measuring the contact angle on a polished surface, a simple vacuum flotation technique was developed and used. Purified oleic acid was the collector and terpineol the frother. Barium activation was studied in some detail, and exploratory studies were made of activation with calcium, aluminum, ferric iron, and stannic tin. Preparation of Materials Quartz: Large lumps of high-grade vein quartz were crushed dry in a cone crusher and rolls. The —20, +28-mesh portion was screened out and used in the subsequent steps. This material was passed through a high-intensity magnetic separator to discard iron, then leached twice with hot concentrated HCl and washed repeatedly with distilled water. The cleaned sand was then wet ground with porcelain balls in a porcelain pebble mill, deslimed repeatedly by settling and decantation to discard —800-mesh material, and again washed with hot HCl followed by distilled water. The resulting stock of quartz was stored under water. Chemical analysis gave 99.8 pct SiO2. Table I gives the size analysis of the quartz used for flotation tests. Calculations from these data, using shape factors given by Gaudin and Hukki9 indicate a specific surface of about 500 cm2 per g. Blank flotation tests in distilled water, and in water with added frother, showed the prepared quartz to be completely nonfloatable and thus indicated the absence of organic contamination. Oleic Acid: The preparation of oleic acid was based on fractional vacuum distillation of methyl oleate2,3 followed by regeneration of oleic acid, and finally fractional crystallization of oleic acid from acetone solutions at low temperatures." The pure oleic acid was stored in a refrigerator. The iodine number of the oleic acid was found to be 90.0 (theoretical 89.93). Oleic acid was used in the form of a dilute water solution of sodium oleate, after preliminary flotation tests showed no effects of form of addition and order of addition of reagents when an adequate conditioning time (that is, 30 min) was provided. Other Reagents: Sodium hydroxide solutions low in carbonate were prepared by first making 1:1

Jan 1, 1951

By Joseph Lindmayer, Karl M. Busen

If silicon is in contact with silicon oxide, a heterojunction is formed which induces an inversion layer ("channel"). Influences of impurities and structural parameters on the channel are discussed. There seems to be no relation between surface damage in silicon and the channel. Phosphorz~s doping of the silicon oxide has no pronounced effect on the channel. Channels which were well above the range predicted by the heterojunction model were related to the action of H* at the silicon/silicon oxide interface. Channels which were well below tlze range predicted by the heterojunc-tion model were related to a transition region in the oxide adjacent to the interface. In support of the assumed transition region and the H+ are the ex-periments described on the paper. Furthermore, the concept of metastable OH groups in silzca, known from other publications, has been used to explain certain channel characteristics by the formation and the accommodation of hydrogen in the silicon oxide. THE electrical characteristics of a semiconductor device in many cases are influenced by the conditions of its surface. Because these conditions often are changing with temperature, time, ambient, or other parameters, numerous attempts have been undertaken to stabilize them by suitable processes. Several years ago Atallal studied the silicon/silicon dioxide interface and since then increasing attention has been given to the silicon dioxide as a means to stabilize silicon surfaces. This led to the evolution of the planar silicon devices where oxide layers are used to protect places at which the junctions are intersecting the surface. The formation of the oxide layers on silicon can be carried out by a variety of processes: high-temperature thermal oxidation (exposing the silicon surface to oxidizing agents such as 02, HzO, COz), evaporation of SiOz, or anodic oxidation of silicon. Originally the oxides were considered able to "passivate" the silicon surface and it was understood that they 1) protected the surface against any influence from the ambient and 2) stabilized surface conditions which were necessary for proper device performance. It soon was realized that the second quality was not always obtainable: Atalla and Tannenbaum showed that, in the process of forming a layer on doped silicon by thermal oxida- tion, electrically active impurities in the silicon near the interface are redistributed, thus giving rise eventually to the formation of unwanted charge distributions. The degradation of planar junctions during reoxidation of silicon was observed by Barson et a. Surface studies with silicon planar junction structures indicated that the breakdown voltage depended on the kind of procedure which was applied during or after oxidation.*" Yamin and worthinge report on charge storage and dielectric properties of SiOz films at elevated temperatures. Another influential parameter on the silicon surface is discussed by Lindmayer and usen,' namely the appearance of a surface potential on silicon as a consequence of the heterojunction formed by the silicon/silicon oxide interface. In the present paper the influence of impurities and structural parameters in silicon oxide in connection with the heterojunction is a particular object of our investigations. I) EXPERIMENTAL The material used was 10 ohm-cm p-type silicon slices with either damaged surfaces (mechanically lapped or polished) or damage-free surfaces (chemically polished). The majority of the slices were oxidized for 1 hr and 20 min at 1250°C in a steam + oxygen atmosphere generated by passing O2 through HzO at 90°C (wet oxygen). This led to oxide layers of about 9000 A. On some slices oxide layers were grown in pure oxygen (dry oxygen). By a photoresist process a set of 20 by 200 mil windows was then etched out from the oxide layer. The pattern of the windows is shown in Fig. 1. In a subsequent process n-type pockets with surface concentrations between 2 x 10" and 5 x 10" cm"3 were formed by phosphorus diffusion. This allowed ohmic contacts between the silicon surface and tungsten probes which for electrical measurements were placed into two adjoining windows and

Jan 1, 1965

By L. R. Plitt, E. O. Lilge

Tests carried out with both kaolin and silica slurries show that flocculants of the polyacrylamide type can be used to improve the thickening performance of hydrocyclones. This thickening improvement demonstrates that contrary to previously held theories, flocs can be formed which are capable of resisting the shear forces in a hydrocyclone. For the 1.25-in.-diam cyclone used, optimum thickening occurs when the flocculant solution is injected into the slurry stream at or near the feed inlet. Hydrocyclones offer many distinct advantages over gravity thickeners. These advantages include simplicity, low initial and operating costs, small space requirements, and flexibility of operation. In spite of these advantages hydrocyclones have not been widely used as thickeners. The main reason for their lack of application is that hydrocyclones are unable to efficiently remove semicolloidal particles (less than 5 microns) from the suspending solution. In gravity thickeners, the fine particles can be induced to form particle aggregates, or flocs, which have the settling characteristics of large particles. In the hydrocyclone, however, it was always assumed that the existence of shear would prevent the formation of flocs.1-3 Thus, if no floc structure is retained, the thickening hydrocyclone must be designed so that its separation size is below the size of the smallest particle to be recovered. To obtain a very small separation size requires the formation of very high centrifugal forces, which necessitates the use of small-diameter cyclones. Small hydrocyclones have, in turn, very low throughput capacities which render them impractical for many industrial thickening applications. Thus, if the effects of flocculation cannot be utilized, the usefulness of hydrocyclones as thickeners remains limited to pulps which contain no very fine particles. It is well established that most substances acquire a surface electric charge when brought into contact with an aqueous medium. The repulsive interactions between similarly charged particles act to prevent flocculation. One method of flocculating a dispersion is to neutralize the repulsive surface charges by the addition of an electrolyte. The reduction of the electrostatic surface charge (Zeta potential) then permits the universal van der Waals attractive forces to operate between the atoms of the various particles and form particle aggregates. The lime additions used by thickener operators promotes flocculation by this mechanism. A second method of flocculating a dispersion is by the addition of long-chain macromolecules. In this case flocculation is brought about by a bridging mechanism in which the molecules are adsorbed with part of their length on two or more particles, thus forming a molecular bridge between the particles.4,5 The molecules also form bridges between themselves when a single particle is bonded to several polymer molecules. The type of bonding between the flocculant molecule and the particle may be hydrogen bonding, chemical bonding, or electrolytic attraction.6 The flour and glue which are used as thickener additives are believed to promote flocculation in this manner. In the past decade synthetic long-chain polymers with extremely high flocculating capabilities have been developed. The flocs formed by these new polymeric flocculants are larger and more shear-resistant than those formed by the presence of electrolytes.7 The advent of these flocculants raises the question: Are the flocs formed by these synthetic polymers stable enough to resist the liquid shearing forces in a hydrocyclone? The investigations reported in this paper were carried out in order to provide some answers to this question. EXPERIMENTAL WORK For the bulk of the test work a 5% slurry of commercially available kaolin with an average particle size of 5 microns was used. An initial evaluation of the available flocculants was carried out to determine which flocculant formed the most shear-resistant flocs. A combination of the techniques used by Healy7 and Booth8 was used for this purpose. This initial study revealed that the neutral high-molecular weight polyacrylamides produced the most shear-resistant flocs. One of these flocculants, Separan MGL*, was selected and then used for all the hydrocyclone tests. The flocculant was dissolved in water and added to the hydrocyclone as a dilute solution.

Jan 1, 1968

By James Boyd

Scientists and engineers must concern themselves not only with technical problems, but with the socio-economic difficulties of our scciety. The author states that raw materials are basic to the economic process and that mineral raw materials occur naturally in nearly limitless quantities. He feels that the economic pocess will provide sufficient raw materials for man's needs, and discusses in some detail the world economic environment. As our society grows in complexity, it becomes more urgently incumbent upon scientists and engineers to bring their experiences to bear on the solution, not only of technical problems, but also on socio-economic difficulties. It is to this field that I have directed a large part of my energies. RAW MATERIALS BASIC TO THE ECONOMIC PROCESS There are certain premises, frequently overlooked, which bear emphasis. As economics concerns the production, distribution, and consumption of goods, so our whole society is based on raw materials. Without the basic extractive industries of agriculture and mining, there would be no economic problems with which to deal. This principle has been expressed as "All productivity is based on three factors: 1) natural resources, whose form, place, and condition are changed by the expenditure of 2) human energy (both muscular and mental), with the aid of 3) tools.77* It seems strange that such an obvious statement needs to be made. It is forgotten by all of us, however, at one time or another when we get involved in the complicated theories of economics dealing with the financial and fiscal policies of the various governments, in the financing of enterprises, in marketing, and in labor relations. Unless decisions in all of these fields are made with this, and the rest of the "Ten Pillars of Economic Wisdom," consciously in mind, the resultant policies are doomed to eventual failure, whether these are policies of individuals, companies, or governments. The economic process is basic to society. But economics and the social aspects of humanity are inevitably intertwined. If human needs are not provided for within the basic economic structure, conflicts arise to disrupt the most carefully laid social plans. Similarly, it is the failure of society to cope with basic economic prerequisites that leads to discord within the body politic and thus to eventual economic disaster. We who are engaged in the various aspects of the mineral industry, therefore, have a large part to play in society's future development. MINERAL RAW MATERIALS OCCUR NATURALLY IN NEARLY LIMITLESS QUANTITIES Throughout my career in and out of government, I have formed my decisions on two principles: 1) Fundamentally, all mineral raw materials exist within the earth's crust in quantities greatly exceeding man's needs; 2) Any problems of supply of these materials are primarily economic in nature. With respect to my first assumption, analyses by geochemists enable them to estimate the quantities in which each of the known elements exist in the earth's crust. One of the most cautious and famous suggests that the average copper content of the lithosphere to a depth of ten miles is about 55 ppm. If this is valid, then the lithosphere contains about 1.4 quadrillion (1.4 x 10 ") tons of copper metal — an unimaginable quantity. This is 8,500,000-fold that which has been consumed by man in his existence. To illustrate further, consider that the deepest copper mine in history went to about 9000 ft below the surface; the present copper mines of the world now average less than 1500 ft in depth. We can assume that in most continental land areas of the world it is technically feasible to mine to a depth of at least one mile. The land areas of the world to an average depth of one mile, then, contain 343 billion tons (343 x 10') of copper metal, or 2000 times that which has been extracted to date. I don't believe man will ever expend his energies to extract copper from rocks which contain only a few parts per million, but he knows that each element

Jan 1, 1968

By Jr. Ensign C. O., J. W. Trammell

The two principal methods of room-and-pillar mining practiced at White Pine make it important to predict variations in the thickness and rock types of a stratum called the upper sandstone. In full column mining, the nearly barren upper sandstone occurs between two ore horizons (called upper shale and parting shale, respectively), and is mined as a part of the ore column. In general, full column mining is not practiced where excessive upper sandstone thickness causes full column ore grade to be substantially lower than parting shale ore. In parting shale mining, the upper sandstone forms the mine roof. Difficulty in predicting upper sandstone character arose because the extent and directions of its variations were not apparent in the property's diamond drilling, in which holes are spaced on 1000-ft centers. A study of sedimentary features, undertaken to improve the predictability of trends in the upper sandstone, led to an interpretation of the sedimentary environment in which the sandstone was deposited. Ripple marks, mud cracks, and cross-bedding, as well as other, less well known features, such as channel casts, flute casts, and current crescent casts, were mapped or recorded. These data, coupled with the knowledge of regional facies changes gained from studying drill core, show that the upper sandstone was deposited by a series of streams flowing northeastward over the underlying parting shale. Deposition of the lowermost bed of the upper sandstone, near the ancient shoreline, was locally preceded by erosion of the parting shale, and the greatest thicknesses of sandstone are found in channels scoured out of the parting shale. Awareness of the rather strong linear trends in the upper sandstone makes it possible to project continuous areas of thick upper sandstone through apparently isolated "highs" in the upper sandstone thickness contour map, which is based on drill-hole information. Since local exceptions to general stratigraphic trends exist, a method was also needed for esti- mating, in detail, the thickness and degree of shali-ness of upper sandstone forming the roof in active parting shale mining areas. Because locally the parting shale was thinned by erosion during the deposition of upper sandstone, parting shale thickness is inversely proportional to upper sandstone thickness. Utilizing information taken from short drill holes into the roof, curves were constructed for the correlation of parting shale thickness (measurable in the mine) with the upper sandstone total thickness, as well as the thickness of its basal member. The White Pine mine is situated in Ontonagon County, approximately 6 miles south of Lake Superior, in the Upper Peninsula of Michigan (Fig. 1). The orebody mined by White Pine Copper Co. occurs in the lowermost 20 to 25 ft of the Nonesuch formation, a series of middle- to upper-Keweenawan shales, siltstones, and sandstones. This paper discusses the solution of problems in ore grade control and mine planning arising from the presence of an essentially barren sandstone stratum within the orebody. THE OREBODY AND MINING TYPES A brief description of the orebody and mining methods is necessary to show how the upper sandstone affects mine planning and grade control. In the mine, the cupriferous zone of the Nonesuch shale is divided into three major portions, as shown in Fig. 2. Lying conformably on top of the Copper Harbor formation (or "lower sandstone") is the lower part of the orebody, the parting shale. The parting shale is overlain by the upper sandstone. On top of the upper sandstone, the basal portion (8 ft) of the upper shale is quite similar to the parting shale. Both the upper shale and parting shale are divided into a number of smaller stratigraphic units, each having a characteristic copper content, as shown by the histogram of Fig. 2. Other authors1'3 have dealt in detail with the stratigraphy and mineralogy of the orebody. However, in this article, we are concerned with the major subdivisions only, and it will suffice to note the low copper content of the upper sandstone, shown in Fig. 2. At present, two types of room-and-pillar openings are created in the orebody. In one type, called parting

Jan 1, 1964

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets.1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741.3 Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name 'and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832.4 Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead," the zinc, and the rock or waste materials. The first record of zinc production from -the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons' in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with, considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1952

By F. X. Tartaron

The comparison of actual energy of comminution with theoretical surface energy presents a wide gap. On the other hand, Solid State Theory presents a viewpoint that places actual energy of breakage in understandable relationship with the theoretical strength of materials. Therefore the physical background of comminution theory favors Kick's Law. A11 discussions of comminution theory appear farfetched when comparison is made of the actual energy of crushing with the energy input called for by thermodynamic theory. Quoting Taggart:' "The work required to crush halite was of the general order of 0.10 kg cm per sq cm of new surface produced (based on air percolation) whereas the theoretical energy of this new surface, based on thermodynamic calculation by a number of investigators is 1 x 10'4 kg cm per sq cm or the work expended was 1000 times the new surface energy — an indicated crushing efficiency of 0.10 pct." It is frustrating to probe into comminution theory with the prospect that the energy one is measuring is 99.9 pct waste energy. Do the laws of comminution deal with energy consumed by mineral breakage or with waste heat energy? There seems to be something wrong. Solid State Theory may provide the answer. The following statement is particularly apt to the situation: "The fundamental fact about metals and indeed all materials, is that they are much weaker than they should be. By this, we mean that if the stress required to cause a given plane of atoms to slip past a neighboring plane is calculated from basic atomic and Solid State Theory, the resulting number is some 1000 times too large to be accounted for by errors in approximations or assumptions. We know now that this is due to dislocations, and we even have a fair understanding of the details of the process by which dislocations make a crystal weak. An intriguing possibility which suggests itself, results from the discovery of very minute metallic crystals which are apparently free of dislocations and which have a strength approaching the ideal value for a perfect crystal." Here we have an extraordinary reversal of the situation. The actual energy for breakage is only one-thousandth the theoretical instead of 1000 times as given by Taggart. But the theoretical energy presented by Solid State Theory is the binding energy between atoms (or ions) in a crystal lattice. This is the kind of energy envisaged by Kick's law. The surface energy or the other hand corresponds to Rittinger's law. Comparison of theoretical with actual, on the basis of surface or on the basis of Rittinger's law, seems to present an absurdity; yet Rittinger's law is confirmed by an impressive array of experimental evidence. Comparison of theoretical and actual on the basis of Solid State Theory or, in other words, on the basis of Kick's law seems reasonable; yet there is not a scrap of evidence to verify Kick's law. The situation appears paradoxical, but the developments in the next section of this paper show that the paradox is a hint to a strange turn of events. ENERGY-WEIGHT FRACTION FUNCTION charles3 has presented a final equation showing the relation of energy consumed in comminution to size of product n{-1) a-n + 1 E is energy, C is proportionality constant, a and n are constants and k is the coarsest size in the ground product (size modulus). In the course of study of a large amount of experimental evidence, Charles discovered that in the general case, a-n + 1 approximated zero. For some reason, a and n differed only by unity. But if a—n + 1 = 0 within experimental error, then it is obvious that Eq. 1 states that it requires an infinite amount of energy to grind any sample to any size. This is a case where nature gives an absurd answer to a man-made mathematical equation, hence something is wrong with the equation. An adjustment must be made to fit the equation with experience. TO reveal the nature of the adjustment, the differential equation preceding the second integration in Charles' group of equations is given When Charles integrated the quantity xa-n it was almost impossible for him to realize that he was dealing with a special case of integration and that

Jan 1, 1962

By L. D. Jaffe

From the limited experimental data in the literature, preliminary values were derived for the thermal diffusivity of titanium alloys and for the quenching severity of various mediums used in heat treating them. The thermal diffusivity, under quenching conditions, was found to be approximately 0.006 sq in. per sec. The quenching severity H for brine is approximately 8, for water 4, for oil 0.9, for air 0.08, and for argon 0.02 in. -' The quenching severity of the Jominy-Boegehold end quench water jet against titanium alloys appeared effectively infinite. Using these results, graphs were prepared showing ideal round sizes and Jominy positions having cooled conditions equivalent to those at the center and surface of titanium alloy rounds, plates, and sheets quenched in various media. THE importance of proper quenching of steel parts has long been recognized, and much quantitative work has been done on the subject. Recently, it has become evident that for titanium alloys, too, the rate of cooling from high temperature is important in obtaining high strength and hardness. Maximum hardness and strength may not be attained in the as-quenched condition but rather after a subsequent tempering or aging treatment. Nevertheless, if cooling from temperatures near the ß transus is too slow, for the hardenability of the alloy used, a soft a-ß structure will be produced which will not harden on subsequent aging.'-" The quenching of parts of practical interest can usually be approximated by simple shapes, such as cylinders (rounds) and plates (or sheets) of various thickness and combinations of these. A standard that has been used extensively for comparison purposes is the center of an ideal round: a cylinder quenched in a medium that instantly cools its surface to room temperature." For steel, considerable practical use has been made of the Jominy-Boegehold end quench hardenability specimen,".' and its value for titanium alloys is becoming evident. It therefore appeared worthwhile to attempt comparison of cooling conditions in titanium alloy rounds and plates of various sizes quenched in various media with cooling conditions in ideal round and Jominy quenches. These comparisons are based upon measurements reported in the literature. Quenching Severity of Jominy Water Jet and of Brine As one starting point, there were available the data of Phillips and Tobin8 for positions on a Jominy specimen having the same hardness as the centers of round bars of various size quenched in 5 pct NaOH brine, shown in Table I. The Jominy speci- men is a round bar, normally 1 in. diam and 4 in. long, quenched by a water jet at one end and air cooled on the other surfaces.' After quenching, hardness is measured vs distance from the quenched end on longitudinal flats ground 0.015 to 0.10 in. deep. The data of Table I were based on a subsized specimen of 3/4 in. diam. Since the positions tested are close to the jet quenched end, cooling there may be taken as due solely to the jet, and air cooling can be neglected. These positions are then equivalent to various locations near one surface of an 8 in. plate jet quenched on both sides. Assuming that the thermal diffusivity is constant and that Newton's law of cooling holds, temperature vs time curves at corresponding positions in Jominy specimen and round bar then may be calculated for various assumed quenching severities of the water jet and of the brine. Cooling curves are generally taken to be metallurgically equivalent when they show the same time to reach a temperature halfway between the initial (heat treating) temperature and the final (quenching medium) temperature.5 On this basis, using Russell's tables,9 upplemented by heat flow tables derived in similar fashion by the author for positions near the surface of severely quenched plates, it was found that the experimentally observed equivalence, Table I, was best matched if the quenching severity H of the Jominy water jet was taken as infinite (H = co). This means that the jet quenched surface should be considered as instantly cooled to the temperature of the water

Jan 1, 1956

By R. W. Fountain, John Chipman

The solubility of nitrogen in Fe-B alloys (0.001 to 0.91 pet B) is determined by the Sieverts' technique for temperatures of 950° to 1150°C. The activity coefficient of nitrogen is decreased by boron. The three-phase equilibrium between ? iron, BN, and gas is established and also the four-phase equilibrium between iron, BN, Fe2B, and gas. The above equilibria are calculated for a iron. The relation of these data to hardenability and strain aging of boron-treated steels is discussed. BORON additions are known to enhanbe the hardenability of heat-treatable steels and to assist in the control of strain aging in sheet steel for deep drawing. The increase in hardenability is explained by the theory that adsorption of boron on austenite grain boundaries reduces their free energy and thus retards ferrite and upper bainite nucleation.l,2 Digges and Reinhart3 have shown that the full effectiveness of boron in commercial steels is achieved only when strong nitride formers such as titanium and zirconium are also present. The influence of nitrogen on eliminating the boron contribution to hardenability was also demonstrated by Shyne and Morgan.4 These workers prepared Ni-Mo steels containing either nitrogen or boron or nitrogen plus boron. The nitrogen-plus-boron steels showed the lowest hardenability which was attributed to the presence of stable nucleating particles, presumably nitride. Morgan and Shyne5-7 have shown that boron in the amount of 0.007 pet will completely eliminate strain aging due to nitrogen in low-carbon, open-hearth steels. In addition, by proper control of the boron additions, a rimming steel can be produced. Since the effectiveness of boron on hardenability and eliminating strain aging is influenced by the amount and distribution of the nitrogen in the steel, the present study was. undertaken to determine the influence of boron on the solubility of nitrogen in iron. EXPERIMENTAL PROCEDURE The solubility of nitrogen in Fe-B alloys was measured by the method of Sieverts, which consists of determining the amount of gas dissolved by the metal in a constant volume system. The apparatus employed in this investigation and the experimental details have beendescribed previously.B AMcLeodgage was added to the apparatus to allow measurements at very low pressures. The alloys were melted at reduced pressure in a basic-lined induction furnace using electrolytic iron and ferroboron. Ferroboron was added after the primary deoxidation of the iron with carbon. Since it was difficult to attain a constant low level of oxygen by this procedure, silicon was added after the carbon deoxidation and prior to the ferr obor on addition. The alloys were castas 2-in. sq ingots, heated in argon at 1050loC, and forged to 1/4-in. plate. After forging, 1116 in. was machined from each side of the plate to remove any possible contamination, and it was then cold-rolled to 0.010-in. sheet. The sheet was cut into approximately 1/4-in. squares and pickled in an inhibited H2SO4 solution to ensure a clean surface. In the case of the boron alloys, a hydrogen treatment could not be used for surface cleaning because boron losses resulted. The composition of the alloys is given in Table I. For a solubility determination, a 75-g sample was inserted in a quartz tube and sealed in place in the apparatus. The entire system was evacuated at room temperature and leak tested for 24 hr. If no leaks were observed, the system was heated to the temperature of measurement and again leak tested for 24 hr. If no leaks were detected, the hot volume and solubility determinations were begun. The hot volume was determined at a constant temperature for each run by admitting successive amounts of argon and recording pressure vs volume, which, in all cases, resulted in a straightline relationship. The argon was then removed and the procedure repeated with nitrogen. Successive additions were made until the desired nitrogen content of the metal and equilibrium pressure of the system were obtained. The

Jan 1, 1962

By A. Phillips, D. H. Killpatrick, V. Kerlins

The tensile failure of two dispersion-strengthened Ni-20 Cr alloys was studied and compared to the fracture of a similar alloy with no dispersoid. The fracture characteristics were studied using electron fruc-tography and transmission electron microscopy. In all three cases, the mode of failure was found to be microvoid coalescence. The failure in the dispersion-strengthened alloys was found to have initiated at the particles. The size of the dimples in the fracto-graphs was found to be related to the spacing of the particles, but not to the total elongation before failure. The elongation before failure was found to be related only to the amount of dispersed phase. These results are compared to those predicted by a theoretical model of ductile failure. THE continually increasing strength requirements for creep-resistant materials capable of long life at service temperatures above 1600"F (875"C) have developed considerable interest in dispersion-strengthened nickel-base alloys. The high yield strength of these alloys results from the presence of a dispersion of fine particles of thoria which act to impede the normal motion of dislocations. Many theories have been presented to explain the role of the dispersion in the strengthening of these alloys. A good review of these theories is presented by Ansell.' The ultimate strength of such alloys is not only related to mechanisms raising the yield strength, but also to the amount of work-hardening which occurs. This work-hardening is determined by the work-hardening rate and the amount of plastic strain before failure. Since fracture limits the amount of plastic strain, a study was undertaken to gain a better understanding of the fracture mechanisms in these alloys. Electron fractography and transmission electron microscopy were used to study the fracture characteristics in two dispersion-strengthened alloys and one similar alloy containing no dispersoid. These results are related to the tensile properties of the dispersion-strengthened alloys at room temperature and at 2000°F (1095°C). The results are also related to a theory for ductile fracture. 1) PROCEDURE Standard tensile specimens were made from 0.020-in. (0.051-cm) sheets of three different Ni-Cr alloys. The alloys had nominal compositions of Ni-20Cr, Ni-20Cr-2Th0,, and Ni-2OCr-4Th0,. All alloys were supplied in an annealed condition. The specimens were fractured in tension at room temperature to determine the effect of the dispersed thoria on the fracture appearance. The tensile properties were determined from the average of a minimum of three ten- sile specimens for each condition tested. The fracture surface of all of the alloys was examined by electron microscopy using standard two-stage plastic-carbon replica techniques. Thin foils of the dispersion-strengthened alloys were used to investigate the size and spacings of the dispersion particles, and to investigate sections of the 2 pct thoria alloy taken from areas of the specimen adjacent to the fracture surface. The thin foils were mechanically ground to 0.010 in. (0.025 cm) and then chemically polished to approximately 0.001 in. (0.003 cm) in a solution of 29 g of ferric chloride and 10 ml hydrochloric acid in a liter of water. The polishing solution was maintained at 150"F (65°C) during the thinning operation. The final thinning of the foil was done electrolytically using a solution of 700 ml ethanol, 100 ml 2-butoxy eth-anol, 120 ml distilled water, and 78 ml perchloric acid (70 pct). The potential was maintained at 15 v and the bath temperature at -20°~ (-29°C) during the thinning operation. The final thickness of the foil was approximately l000A as determined from the width of twins boundaries observed in many of the foils. 2) RESULTS AND DISCUSSION The electron fractographs of the fracture surfaces of the alloys are shown in Figs. 1, 2, and 3. The normal interpretation of this type of fractographz is that, as a result of the difference between elastic and plastic properties of the matrix and the particles or other inhomogeneities in the alloy, microvoids are formed

Jan 1, 1969

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

The preferred orientation and earing characteristics of ZS aluminum were studied. An empirical correlation was obtained relating earing behavior and variation of mechanical properties for face-centered cubic metals. Strain-ratio mecsurements for ZS were found to be in good agreement with Hill's theory of plastic anisotropy. DIRECTIONALITY in sheet metals may be revealed by earing in a cupping test, or by the variation of mechanical properties of tensile specimens taken at different angles to the rolling direction. However, no simple correlation appears to exist between earing and variation in mechanical properties in metals and alloys studied.1-5 Aluminum of commercial purity (known as 2S, with a specified minimum of 99.0 pct Al) cast by the direct casting process was chosen in the present investigation since ears at 45° or 0°, 90° can readily be obtained. Preferred orientation is often the principal cause of uneven metal flow, such as earing, and variation in mechanical properties. Consequently, the preferred orientation characteristics of 2s aluminum were initially studied using both X-ray diffraction and micrographic techniques. The variation of mechanical properties with angle to the rolling direction was next determined in an effort to clarify the apparent lack of correlation of earing behavior with mechanical properties for face-centered cubic metals. Strain-ratio*5,8 measurements were finally conducted to determine first, if the maximum values of the strain ratio occurred at the earing positions, and second, if Hill's theory of plastic anisotropy7 is applicable to other face-centered cubic metals such as aluminum. Detailed treatments, working, and heat treating are not given in this paper since this study is not concerned with the methods to produce earing, but to correlate earing with other properties and characteristics of the material. The percentage of earing was measured using the formula specified by the x—Y Aluminum Association, -------- X 100 = pct earing, where x is the height of the ear and y is the height of the valley from the bottom of the cup. Preferred Orientation Characteristics of 2S X-ray Studies: Beck and coworkers8 have determined that the rolling texture of 2s after 95 pct cold reduction may be approximately described by four equivalent ideal orientations near (123) [121]. Also, the recrystallization or annealing texture of 2s aluminum consists of four components of the (123) [l21] type, retained from the rolling texture, and of a (100) [001] or cube texture component." Fig. 1 illustrates the typical appearance of deep drawn earing cups with the corresponding X-ray diffraction patterns. The X-ray patterns were taken with copper radiation using the structure-integrating method'" in which the specimen is scanned during the X-ray exposure. The rolling direction was vertical and the rolling plane was perpendicular to the horizontal X-ray beam. It was found that an increase in the retained rolling texture component resulted in a greater 45" ear height, while an increase in the recrystallization texture component (i.e., cube texture) caused greater 0°, 90° earing. X-ray diffraction studies of 2s cold rolled to final thickness followed by a final anneal have shown that the rolling texture, and also the retained component of the rolling texture after annealing, are increased with greater cold-rolling reduction. The preferred orientation characteristics of 2s sheets given an intermediate anneal between cold rolling followed by a final anneal were found to depend on the position of the intermediate anneal during cold rolling. For instance, three identical samples of 2s homogenized ingot, each 1 in. thick, were given the following cold-rolling and annealing treatments: 1—50 pct reduction, intermediate anneal, 96 pct reduction to 0.020 in. and a final anneal; 2—75 pct reduction, intermediate anneal, 92 pct reduction to 0.020 in. and a final anneal; and 3—90 pct reduction, intermediate anneal, 80 pct reduction to 0.020 in. and a final anneal. The intermediate and final annealing treatments consisted of a 70°F per hr heating rate, held 6 hr at 800°F and air cooled. Typical X-ray patterns after each of these treatments are shown in Fig. 2. It was evident that after treatment 1, the retained rolling texture component predominated; after treatment 2, the retained roll-

Jan 1, 1954