Search Documents

By V. B. Kurfman

The temperature, strain rate, and orientation deDendence of defbrnzation of single-crystal MgAg has been examined. The crystals exhibit a tendency to single glide and little or no hardening at 25°C for many orientations. A much higher hardening rate is observed when multiple glide occurs, such as can be initiated by surface defects. The tendency for easy glide becomes less dependent on surface preparation and orientation as T — 100°C and bars so tested often fail after one-dimensional necking-. At T > 200°C (transition temperature for single-crystal notch sensitivity and poly crystalline ductility) single glide diminishes and two-dirnensionul necking begins. The crystals do not strictly obey a critical resolved shear stress law, but show the influence of {loo) cracks in determining the slip mode. The results are correlated with the difficulty of sciperdzslocation intersection and semibrittle behavior of this compound in single-crystal and poly crystalline form. Comparisons are made with the slip selection mode observed in tungsten, with the reported observations of easy glide in bee metals. and with the mechanical behavior of poly crystalline MgAg. PREVIOUS work on tensile deformation of polycrys-talline MgAgl and bending deformation of single-crystal MgAg2 has shown that the compound is semi-brittle (i.e., notch and grain boundary brittle). If this semibrittleness is supposed to result from the difficulty of multiple glide (associated with the problems of superdislocation intersection) one might expect single crystals deformed in tension to show pronounced single glide and strong orientation dependence of hardening rate. These experiments were done to examine this supposition and to study the tensile deformation of a highly ordered system which may be considered bcc if the difference between the two kinds of atoms is ignored (actual structure: CsC1). EXPERIMENTAL Single-crystal ingots were grown by directional freezing as previously described.' These ingots were sliced into a by a by 2 in, rectangular bars by electric discharge machining, then round tensile bars were conventionally machined to 1/8-in.-diam by 1-in.-long reduced section. The bars were typically tested without an anneal because of the problem of magnesium vapor loss and they were typically tested as mechanically polished. The analyses are within the same limits as those reported earlier; i.e., the average composition for each specimen is within 0.5 at. pct of stoichiometry, while the total range from end to end in a given specimen varies from 0.7 to 1.4 at, pct. There has been no indication in the results of any variation in slip or fracture mode attributable to the composition fluctuations. The slip systems were determined by two-surface analysis of the bars after testing to failure at room temperature. Single glide was so dominant that there was little difficulty in identification of the dominant slip system even though the tensile elongation to failure often approached 7 to 8 pct in room-tempera- ture tests. Elevated-temperature testing was done in a silicone oil bath and low-temperature testing was done in liquid Np or a dry-ice bath. All stress measurements are reported as engineering stress unless otherwise specified, and crosshead travel is used as the strain measurement. RESULTS The tendency toward single glide is best seen in the pictures, Figs. 1, 2, and 3, which depict deformation at fracture as a function of test temperature. While it is possible to find regions of secondary slip by careful microscopy, such regions are very small. The development of a ribbon-shaped configuration from an initially round section bar pulled at 100°C is typical, occurred by single glide, and illustrates the degree to which such glide continues. At temperatures =100°C the bars typically show elongation of 20 to 50 pct by predominently single glide. Despite the large elongation, fracture even at 150°C occurs in a brittle mode, Fig. 2, in the sense that it is an abrupt failure which shows no discernible necking in the second dimension of the bar's cross section (i.e., there is no appreciable action of any slip modes which would decrease the broad dimension of the cross section). Near 200°C the fracture mode changes slightly. Although most of the sample extension is by single glide, after the bar develops the characteristic ribbon shape it begins to neck in the second (i.e., broad) cross-sectional dimension. The bar becomes very thin in the "necked down" region, Fig. 3, and the reduction in area approaches 100 pct. Often there oc-

Jan 1, 1967

By W. S. Sanner

Seven samples were analyzed from the major reserves of the Bottom Red Ash bed, Northern anthracite field. The data show variations in thickness and geo-logic structure and in the amount of coal that can be recovered. Recovery of market-grade (less than 17 percent ash) coal is high for all samples, ranging from 76 tp 95 percent in the 3-1/4-by 9/16-inch sizes. The occurrence of coal containing less than 5 percent ash varies from 19.5 to as much as 57.5 percent. Crushing the coarse sizes to minus 9/16 inch increased recovery of market-grade coal in six of seven bed samples. Crushing released sufficient impurities to cause substantial increases in yield of coal containing less than 5 percent ash for all samples. Analyses of the float coal indicate that the quality of coal is comparable for all samples.

Jan 1, 1967

By A. Vaillant

Classifiers that are used to separate granular materials into a coarse and a fine stream, are considered. Fundamental Equations are presented which apply to all, such Classifiers and all feed size distributions. The practical, real Classifier is compared quantitatively with the perfect. Classifier. A Formula is introduced for the Real Classifier Selectivity, that defines what portion of each particle size is sent to the coarse outlet duct. This selectivity function is characterized by three Parameters, one of which (b) is a numerical measure of the sharpness of classification. A new dimensionless variable (J) appears in Equations and Graphs for recirculating load, product size distribution and specific surface in the classified product. This variable combines the effects of sharpness of separation, Classifier setting and fineness of a Rosin-Rammler Type feed to the Classifier. Classifier efficiency is defined, based on Specific Surface.

Jan 1, 1967

By E. W. Bauman

The important role that mineral aggregates assume today in the construction and chemical industries is best attested to by the astronomical tonnage figures for the annual production of aggregates in the United States as reported by the U. S. Bureau of Mines in their "Minerals Yearbook." For 1965, latest publication available, the tabulation shows a total of 1,716,000,000 tons made up of 30,000,000 slag, 908,000,000 sand and gravel, and 778,000,000 crushed stone. A much more favorable statistical position for slag can be established by listing the world-wide tonnage. According to Dr. A, W. Lee, Director, British Slag Federation, in a paper presented before the 8th Annual conference of the British Institute of Works and Highway Superintendents, he estimates that the probable world consumption of slag is in the order of 100,000,000 tons showing--Great Britain with 10,000,000 tons, Germany, 15,000,000 tons, France, 1$000,000 tons and the United States with 30,000,000 tons. Thus it is seen that the United States accounts for almost one-third of the world's consumption of slag and the 3 west European powers another third.

Jan 1, 1967

By F. W. Jessen, G. F. Sears

Cavities in massive salt for the purpose of storage of liquid hydrocarbons have assumed a prominent position in recent years. This paper describes a program to facilitate leaching operations for the formation of specifically shaped storage cavities. Various forms and sizes of cavities may be possible through use of the techniques developed. INTRODUCTION The creation of large under ground storage facilities for natural gas and liquified petroleum gases has been practiced for many years. Use of cavities obtained through solution of salt for this purpose is a fairly new and novel approach, but has gained increasing importance due mainly to the economics of this type of mining operation opposed to hard rock mining or surface and pit storage.1 The idea of controlling the shape of any cavity dissolved from massive salt has not been a prime consideration of companies engaged in the formation of storage space. This has been due mainly to an insufficient knowledge of the mechanics of the leaching process and a dearth of published information dealing with both the desirability and ease of control possible for this type of operation. Two distinct advantages are readily ascertainable. From a stability standpoint (i.e., the ability of the completed cavity to withstand stress imposed by overburden pressure and lateral tectonic stresses) a controlled cavity may be generated which will yield the most favorable attitude to these external forces and remain in operable use for longer periods of time.2 A second advantage is that for a given volume, a sphere (which is one controlIed shape possible) represents the minimum surface area exposed. This may become more important in the future when dealing with refrigeration and product losses in underground cavities. The basic solution mining process, without regard for controlling the final shape, is quite simple in that the equipment and materials required to dissolve a cavity in a salt dome or layered salt section are a source of fresh water, a circulating pump, several strings of tubing and a means of disposal of the return brine. To add control measures involves the use of an inert blanket material above the position at which solution proceeds. Initially the annulus of the largest wash pipe is filled with this blanket down to the top of the proposed cavity. The bottom of the proposed cavity is determined by the depth of the original drilled hole or a blanking plug. The blanket material is added incrementally in stages and displaces the water in the enlarging cavity downward. Where the blanket material has displaced the water, no further solution takes place. The rate at which to add this controlling blanket material so as to be able to form specifically shaped cavities of any particular size is of considerable importance. THE PROBLEM It is desirable to have as much information as possible as to the behavior of the mining operation since visual observation obviously is impossible. It would be advantageous, for example, to have a step-by-step program showing how much salt is to be removed, at what rate and the time required. This type of program would facilitate procurement of surface equipment and provide for proper management of manpower requirements, as well as yield a host of intangible benefits. As mentioned earlier, only rule-of-thumb estimates were available until recently and made this type of pre-planned operation haphazardous at best. Recent work3-' has done much to clear up the ambiguities and inaccuracies attendant to solution mining of salt and has helped to place the operation on a more scientific basis. This study attempts to correlate much of the information thus far developed into a complete mining program; to extend the idea of controlled solution mining to include not only spheres, but solid conic sections of the ellipsoidal variety; and to refine the computation of the rate of removal of salt during the various stages of mining. Dommers3 and IIemson4 made extensive laboratory

Jan 1, 1967

By G. L. Barthauer

It is a distinct privilege to have the opportunity of addressing this group today. Although I, myself, am not a mining engineer, I have always had profound admiration and respect for those of you who each day live and work according to the tenets of Murphy's Law. Murphy, as some of you may know, was an Irish philosopher whose sole contribution to his field was the observation, "In the mining industry, if anything can possibly go wrong, it will!" You have heard this afternoon a variety of papers, touching on or describing in some detail imminent problems which presently face the coal industry. Many of these problems show no signs of diminishing. Actually, there is good reason to believe that they will intensify as governing bodies pass additional enabling legislation and regulatory agencies adopt more stringent "Rules and Regulations". In many cases, these groups are seemingly oblivious or indifferent to either cost factors or technical problems involved in achieving the desired goals.

Jan 1, 1967

By A. O. P. Casbarian

An analytical method is developed to determine the variation of the ultimate lateral resistance of a plate in a cohesionless soil with depth. This analysis is based on a modification of Rankine's classical earth pressure theory and the theory of plasticity as applied to soils. The ultimate resistance is defined as the product of the effective stress at the midpoint of the plate, the area of the plate and a dimensionless variable termed the ultimate resistance factor. This variable has been plotted us the depth ratio: i.e., the ratio of of the depth of embedment us the height of the plate. The resistance of a plate may then be calculated using the values af the ultimate resistance factor from the chart provided or the equation may be programmed for use in an analysis of anchor systems in cohesionless soils. It is emphasized that the analysis is semitheoretical. The theory has been compared with experimental results reported in the literature and results indicate general agreement. Actual field tests are necessary to further verify this theory. INTRODUCTION With the exploration for oil offshore in waters all over the world, it is of importance to determine the behavior of the various soils in relation to their ultimate resistance to deformation. Examples of such problems are the holding capacity of anchors and the lateral resistance of piles. Very little information is available in the published literature on the design and performance of anchors, in either cohesive or cohesionless soils. The analysis developed in this report is the first step in obtaining a solution for the determination of anchor holding capacity in a cohesionless material. DISCUSSION The theory used in the analysis is based on the ultimate strength of the soil and is the maximum resistance developed by the plate against further movement. In such a state the elastic deformations are disregarded in comparison with the plastic deformations. Hence, the plate can be considered as completely rigid. The theory of plasticity determines the three unknown stresses at any point by means of two equilibrium conditions for a small earth element in combination with the failure condition. However, the exact solutions can only be carried out in a few simple cases such as, for example, when the rupture or failure lines are straight (Rankine theory) or with spiral and straight rupture lines (Prandtl theory).l Kiitter 2 derived a single equation expressing the variation of the stress in any given rupture line. To utilize this equation, it is necessary to know the stress in the rupture line at a certain point. Unfortunately, this is difficult to obtain unless the rupture line intersects the free surface at a certain angle or when the earth is cohesion-less and unloaded. A method to overcome this is to consider only the boundary conditions at both ends of a rupture line without investigating the equilibrium of the earth above the rupture line. This assumes that the rupture lines meet the surface at statically correct angles, so that boundary stresses may be determined. As Kiitter's equation furnished a relation between these stresses, the unit earth pressure may be calculated at the point where the rupture line meets the wall. Another way in which Kotter's equation may be applied is in investigating the equilibrium of a soil mass above a rupture line. This method assumes that the failure or rupture line is known and that the boundary stresses in the rupture line at the ground surface can be determined. In this case it is possible to determine the earth pressure from the equations of equilibrium.

Jan 1, 1967

By W. P. Haynes

To evaluate the potential use of battery-powered vehicles, the Bureau of Mines studied the performance of a battery-powered truck propelled by a 9.6 kw dc motor in considerable detail under various driving conditions. Some of the factors investigated were: Gear ratios, payload (1,468 lb maximum), road grade, stops made per mile, and charging time. With the exception of some difficulty with brush wear, operation and maintenance of the truck was generally trouble-free. On range tests carrying an 868-lb payload a distance of about 50 miles at an average speed of about 19 mph in normal driving, the Bureau's battery-powered truck averaged 1.6 to 1.9 miles per ac kwhr input. Stop-and-go tests indicated that prospects for the use of battery-powered trucks in house delivery service within city limits are good; performance of the Bureau truck was sufficiently high for that type of service. The estimated power cost of 2.65 cents per mile while making 10 stops per mile was attractively low from the standpoint of fuel costs.

Jan 1, 1967

By J. C. Trantham, J. W. Marx

From 1955 to 1958 the Phillips Petroleum Co. conducted a series of small scale counterflow combustion field tests in a tar sand about 60-ft deep and 6 to 12-ft thick near Bellamy, Mo. A total of seven different well patterns, including conventional five-spots and seven-spots plus a 15-well line drive pattern and a 10-well radial pattern, were employed. Electric heaters, gas burners and wellbore fuel packs were tested as counterflow ignition devices. Direct drive ignition followed by reversal to counterflow operation was not possible in this case because of formation plugging by the semi-solid tar during the direct drive phase. Sustained counterflow burning was achieved with wellbore fuel packs under conirolled conditions which permitted close correlation of fire front velocity, air velocity, air-oil ratios and yields for a wide variation in process parameters. The tar in place was about 10' APl with a reservoir viscosity greater than 500,000 cp. The composite oil prodlcct was 26' API with a viscosity in the range of 5 to 15 cp. Vertical sweep efficiency in the line drive pattern was essentially 100 per cent, with no evidence of gravity segregation. Original tar in place ranged between 800 and 1,000 bbllacre-ft, specific rock permeability averaged about 800 md before burning and effective air pemeability with residual oil and water in place (at the time of ignition) averaged 250 md. The volume of oil recovered during stabilized line drive counterflow burning represented about 67 per cent of the volume of tar originally in place in the burned out zones. A minimum was found when air-oil ratios were plotted against formation air velocities, and a limiting air velocity existed below which the counterflow fire front reversed and burned back over its own trajectory, using the residual coke left behind. INTRODUCTION Early in 1955 the Phillips Petroleum Co. began preparations to field test counterflow underground combustion as a method for producing oil from high viscosity hydrocarbon deposits, such as bituminous (tar) sands. At that time all published information on the use of underground burning io increase oil recovery was based on the direct drive process, in which the fire front moved through the formatim in the same direction as the injected air stream. Such direct burning is not applicable to many heavy oil sands or to most tar sands because the heat-thinned native hydrocarbon congeals in the cold rock ahead of the fire front forming a gas permeability block which smothers the combustion. Laboratory experiments first demonstrated that this permeability blocking would not occur if the fire front and the air stream moved in opposite directions. In this counterflow (reverse) burning process the heated heavy oil or tar is driven back through the combustion zone. Here thermal cracking produces relatively light hydrocarbon products which pass through the heated rock behind the fire front (largely in the vapor phase) and are incapable of causing a gas permeability block.' Exploratory wring revealed that consolidated bituminous sandstones with a combination of tar saturation, permeability, thickness and depth favorable for experimental field testing existed at several points in western Missouri. By the fall of 1955, a test site for determining technical feasibility of the counterflow process under field conditions had been selected near Bellamy, Vernon County, Mo. — about fifty miles north of Joplin. The pay zone was too thin and too shallow to constitute a commercial prospect, but these same factors permitted controlled experimental operation with a small compressor installation and with a large number of monitor wells. These were needed for adequate observation of the underground phenomena at reasonable cost. RESERVOIR CHARACTERISTICS The reservoir chosen for the experimental field tests was a 124 thick section of tar sand extending from 49 to 61 ft subsurface, with shale and siltstone laminations sealing the top and bottom. This section was part of a larger 30-ft thick tar sand deposit which extended both above and below the test zone. Core analyses showed that the 12-ft sand was effectively subdivided into two approximately equal sections, with the lower zone being consistently more permeable than the upper zone. The most significant laboratory core data are summarized in Table 1. The line drive test which is the principal subject of this report utilized only the lower 6-ft thick zone, with the upper zone regarded as part of the 55-ft overburden. WELL PATTERNS The over-all field test program involved a total of seven different well patterns, including conventional five-spots and seven-spots plus a 15-well line drive pattern and a 10-well radial pattern. In some of the tests the full 12-ft zone was employed; in other cases only the more perme-

Jan 1, 1967

By J. E. House, J. L. Drobnick, R. R. Swanson, D. W. Agers

The paper describes an improved process for the recovery, purification, and concentration of copper values from acidic dump leach liquors. The process employs a new liquid ion exchange material (LIX® 64 reagent) capable of separating copper values from iron and other impurities common to dump leach liquors in the typical pH range of 1.4-3.0. Properties and performance of the reagent are discussed. Design scale-up information for both pilot plants and commercial installations, including preliminary economics, are presented. Economic considerations include the production of electrolytic copper from the purified copper aqueous concentrate. INTRODUCTION Since the commercial acceptance of the liquid ion exchange process in the mineral processing industry, it has been predicted that a wide selection of chemical reagents to efficiently and economically separate many metals in solution under a wide range of conditions would be commercially available. Several years ago, General Mills began a basic research and development program to determine if it were possible to develop a reagent to extract economically metallic ion from existing process streams. Progress toward reaching this objective was reported in late 1963, when development of quantities of LIX -63 reagent, a highly selective reagent for the extraction of copper, was introduced. The properties of this reagent are described in a previous paper.' The development work with this reagent indicated that perhaps the largest potential application for a copper extractant, the treatment of acidic dump leach liquors, would not be economically possible with LIX-63 because of limitations of the reagent to extract copper below a pH of 3.0. The relatively small pH adjustment required with LIX-63 would result in a significant processing cost and the undesirable precipitation of ferric hydroxide from the liquor. This pointed to the need for another copper extractant which would extract copper from dump leach liquors without pH adjustment. General Mills was successful in synthesizing a molecule which does not have the pH limitation of the LIX-63, yet possesses the selectivity-solubility properties required of a metallic extractant. The resulting LIX-64 reagent broadens the pH range from which copper can be effectively extracted and separated from other metallic impurities. Of special significance is the fact that the reagent, LIX-64 was developed by combining the technical requirements of a specific problem in hydro-metallurgy with theoretical organic and physical chemical principles.

Jan 1, 1967

By Murray Weintraub

The Bureau of Mines investigated the aerodynamic factors that control the functioning of the blowcase devices used in pneumatic processes for dispersing solids. Alumina powder of 8-micron mass-mean diameter was placed in 0.75- inch-diameter by 10-inch-long cylinders} and a blowcase technique was employed. It was determined that dimensions and gas rates for discharging the powder from the cylinder by a stream of gas may be calculated by assuming a fluidization mechanism and extrapolating residues to a desired finite quantity or to zero at the terminal velocity. The state of agglomeration of a powder at discharge may be determined from measurements of the velocity required for zero residue.

Jan 1, 1967

By Virgil A. Bilyeu

The Pan American Mine, in southeastern Nevada, is about 26 miles west southwest of Pioche, Nevada, and about 266 miles southwest of Salt Lake City, Utah. The mine is located in the Comet Mining District, and since the ore occurrence is generally the same in the Comet and well known adjacent Pioche Mining Districts, the history of the two shall be considered the same. The Pioche District began about 1869 with the production of high grade gold and silver from fissure veins in the Cambrian Prospect Mountain quartzite. Peak production was reached in 1872 and then declined with sporadic production until 1924, when Combined Metals Reduction Co. shipped its first, complex sulphide ores mined from the CM bed at Pioche. Extensive ore bodies in this CM bed attracted interest in the same bed in the Comet District. Small pockets of high grade silver-lead were mined from fissures cutting the CM and other limestone beds above. Exploration in the CM beds at this time proved a large area of mineralization too low grade to be considered ore.

Jan 1, 1967

By T. R. Ingraham

The thermal decomposition of Nickel (II) sulfate was examined by determining the total pressure of SO3, SO2, and O2 developed over a sample when it was heated in an evacuated system fitted with a Pyrex-bellows mercury manometer. The equilibrium pressures were combined with previously published data on the oxide and sulfide systems to establish a predominance volume diagranz for the Ni-S-O system over the normal range of roasting temperature and gas composition used in the treatment of nickel minerals. The protective coating of NiSO4 which forms on the oxide during roasting, and which causes the sulfation reaction to almost cease, may be disengaged readily by temperature cycling. When nickel sulfide is roasted in air, either nickel sulfate or nickel oxide may be formed by altering the experimental conditions. Because of the differing stabilities of various metal sulfates and oxides, possibilities exist for separating some of the constituents of mineral mixtures by the technique of selective sulfation and oxidation roasting. The experimental information required to make reasonable predictions of the possibilities of this technique is incomplete and, in some cases, in conflict. As part of a general program of providing some of the required information, a study EXPERIMENTAL Materials, Apparatus, and Procedure. Fisher certified reagent-grade nickel sulfate hexahydrate was used in the source material for both nickel sulfate and nickel oxide. The manufacturer's analysis indicated: 0.001 pct Fe, 0.005 pct Cu, 0.003 pct Co, 0.001 pct C1, other 0.15 pct. The material was dehydrated, before use, by heating at 420°C for 5 hr. The oxide was prepared by roasting the sulfate in an open tray in a muffle furnace for 10 hr at950°C. Preliminary experiments which were made by thermogravimetric analysis and X-ray diffraction analysis established that only one crystal form of nickel oxide and one of nickel sulfate were involved in the equilibrium shown in Eq. [I]. The apparatus used in the experiments was described previously.1,3 The main feature is a corrosion-resistant, flexible Pyrex bellows which is used to protect the mercury in a u-tube manometer. Before a run was begun, the apparatus was washed thoroughly with hot hydrochloric acid and water and dried by heating under vacuum. The mixed sulfate -oxide sample was contained in a platinum bat rest-

Jan 1, 1967

By William R. Hardwick

Present information indicates that a copper deposit can be safely fractured with a nuclear explosive and the copper successfully recovered by the in-situ leaching method. The process is not yet at the stage where it can be presented t ) the mining industry as a technique proven in all its aspects, but it is a method with high success potential. Unknown factors must be evaluated by a full-scale test before the economics or the extent of the use of the method by the mining industry can be predicted. Preliminary calculations indicate that the cost of fracturing a copper deposit nuclear explosives may range from l.8 to 55.2 cents per ton or in some case: less than a cent per pound of recoverable copper. Copper may be recovered at less cost than by conventional methods. Deposits with 4 pounds of recoverable copper may be economic and production from a deposit may begin in less than half the time required for conventional methods.

Jan 1, 1967

By James W. Estep

The sulphur shortage that has occurred in the past two years has directed attention more and more toward sour natural gas reserves, Wellhead values of sour gas vary widely depending upon the acid gas composition, location, market conditions, and necessary processing charges. Preparation of a saleable residue gas is considerably more involved and expensive than that normally encountered in treating a sweet gas. Recent revisions in sales gas specifications have increased capital costs and processing charges. The presence of hydrogen sulphide and carbon dioxide in a natural gas contributes to shrinkage factors that may approach 50%. Fuel required by the gas sweetening process and sulphur slant tail gas incineration also contributes to a further reduction in the volume of residue gas available for sale. The additional processing facilities involving gas sweetening, mercaptan removal, expanded utilities, sulphur recovery and sulphur handling, all demand additional operating and maintenance staff. A favorable sulphur market under certain conditions may justify the processing of sour gas whereby sulphur is either a by-product or a co-product, Five typical cases involving different compositions of sour gas are presented for evaluation. Capital costs, operating problems, and operating costs are presented in detail for full economic evaluation

Jan 1, 1967

By P. G. Hallof

The exploration staff of Campbell-Chibougamau Mines Limited has used past experience to delineate several areas of geologic interest under Lake. Chibougamau. Du• ring the winter of 1964-65, a reconnaissance induced polarization survey was carried out on the lake. First, several lines were surveyed over the known mu1.erahza-tion at the Henderson mine; the grid was then extended into the areas of interest. Several anomalous zones were located and drilled; one, of these proved to be of economic interest.

Jan 1, 1967

By Jr. Driscoll

The thermoelectric properties of yttrium monoarsenide, yttrium monoanti-monide) and ternary materials from the Y-As-Se and Y-Sb-Te systems were investigated by the Bureau of Mines from room temperature to 1,000° C. The maximum calculated figure of merit for yttrium monoarsenide and yttrium monoantimonide was 2 x 10-5 ° C-1. YAs and YSb were semimetallic conductors with low Seebeck coefficients and electrical resistivities. The values of the Seebeck coefficient and electrical resistivity of the Y-As-Se and Y-Sb-Te materials were between those of the yttrium selenides and tellurides and of YAs and YSb, respectively. The maximum power factor calculated for the Y-As-Se materials was 2 X 10-5 watt/cm ° C2 which yielded an estimated figure of merit of 6 X 10-4 0 C-1. Most of the Y-As-Se compounds had high Seebeck coefficients and moderately low electrical resistivities. The Y-Sb-Te materials were degenerate semiconductors with power factors of 6 X 10-8 watt/cm ° C2.

Jan 1, 1967

By D. A. Elgillani, M. C. Fuerstenau

With either sulfonate or oleate as collector, quartz responds to flotation with moderate additions of calcium only at moderately high pH, where some portion of the activator has hydrolyzed to caOH+ . Calculations of the concentrations of various ionic and precipitated species of calcium and collectors suggest that the products of [(CaOH+) (RSO3)] and [(CaOH+)(01-)] determine whether flotation is obtained under specific conditions. Ion products on the order of 10-12 were calculated for both the sulfonate and oleate systems. The activating effect of calcium ion in nonmetallic flotation systems is of considerable interest because of the normal presence of calcium in natural water. As a result, this phenomenon has received quite some attention in the past. Kraeber and Boppel1 showed that quartz could be activated by calcium above pH 10 with sulfonate as collector. The feasibility of selectively separating quartz from hematite with calcium activation at relatively high pH was demonstrated by Clemmer, Clemmons, Rampacek, Williams, and stacy.2 Cooke and Digre3 showed with a bubble pick-up method that the minimum quantity of calcium ion required as activator for complete pick-up of particles occurs at pH 11.5 for an addition of 20 mg per liter sodium oleate. They also showed that larger additions of calcium (10-fold increase per unit decrease of pH) must be added for complete bubble pick-up as the pH is reduced. Schuhmann and Prakash,4 using a vacuum flotation technique, found that quartz could be floated with moderate additions of calcium chloride and oleic acid at neutral pH, providing the metal ion was present in stoichiometric excess over the quantity needed to form the normal soap with oleic acid. They also reported that calcium will function as an activator only in basic media. More recently, Eigeles and volova5 have shown that essentially complete flotation of quartz is obtained with 6 x 10-4 mole per liter calcium chloride and 1.7 x 10-5 mole per liter sodium oleate at pH 11.6. while no flotation is obtained at about pH 10.9 and below. The importance of adsorption of activator and collector at the air-liquid interface is also demonstrated in these systems. The important role that metal ion hydrolysis assumes in quartz activation systems was also demonstrated recently.6-8 A detailed investigation of metal activation in sulfonate flotation of quartz was undertaken in one system7 and yielded a number of interesting and important observations. Quantification of the data of this system7 to the extent desired was not possible, though, because certain species could neither be ignored nor accounted for accurately. These difficulties can be circumvented when calcium is involved as activator. This detailed analysis was undertaken to obtain a more quantitative explanation of calcium and metal ion activation in quartz flotation. EXPERIMENTAL MATERIALS AND METHODS Sodium alkyl aryl sulfonate9 mol wt 450, and pure potassium oleate were used as collectors. All other reagents used were reagent grade in quality, i.e., n-amyl alcohol as frother, KOH for pH adjustment, and calcium chloride. Conductivity water, made by passing distilled water through an ion exchange column, was used in the investigation. Quartz was prepared by leaching the sized sample (48 x 150 mesh) with HC1 until no iron could be detected in the leach liquor. The experimental equipment and procedure were the same as that described previously.6,10 EXPERIMENTAL RESULTS As the presence of precipitates was noted in all of the systems to which ca++ and collector were added, experiments were undertaken to determine the solubility products of calcium sulfonate and calcium oleate using a nephelometer. With this technique, collector is titrated into a known solution, which in this case was 5 x 10-5 mole per liter CaCl2 at pH 5.5. Upon precipitation of the calcium-col lector salt, e.g., calcium oleate, light is scattered and detected

Jan 1, 1967

By G. R. Love, M. L. Picklesimer

Aboue 950°C the Nb-Zr system consists of a completely miscible bcc solid solution, commonly called the phase. Between 950 and 600°C, and between 20 and 85 pct Nb, the phase deconlposes, after sunciently long times, into two bcc solid solutions. The pct Zr alloys are conveniently descibecl with T-T-T (time-temperature-transformation) curves having a nose at about 2 hr at 700°C. The reaction rate varies only slowly with zirconium content and negligibly with oxygen contanzination; it is speeded up by a factor of 10 to 15 by 90 pct cold ulork and slowed dou by n factor oj 10 to 30 by a two-hundrecljold increase in grain size. Nb-r alloys with compositions between 40 and 85 pct Nb have been the basis for the majority of commercially important superconducting materials. In part because of their commercial promise, more is known about these alloys than about most other high-field superconducting materials. At the same time, there is considerable disputed or incomplete metallurgical information. For example, although Rogers and tkins&apos; indicate a monotectoid reaction at approximately 600°C and a two-phase 01 + 0, field extending between 20 and 85 pct Nb and to a maximum of 95OGC, erhout&apos; has reported that this entire region would be a single homogeneous B were it not for oxygen contamination. Again, although it has been shown that relatively short-time heat treatments in the vicinity of 700CZ significantly improve the ability of short wire samples to carry high currents in high magnetic fields at 4.2K, these observations have never been fully correlated with the structural change or changes occurring during the anneal. We intend to investigate in detail the effect of metallurgical variables, including heat treatment, on the superconducting properties of hard superconductors. To verify that our experimental techniques are valid and to establish a relative standard against which other materials may be measured, we feel it advisable to know the behavior of the Nb-Zr alloys under a variety of processing conditions. As an initial step toward this goal, we have determined in detail the kinetics of the transformations in Nb-Zr alloys. EXPERIMENT A number of problems had to be solved before beginning any fruitful work on the reaction kinetics in this system. While solving some of these problems, either by chance or by design, small amounts of information were obtained about alloys containing 40, 50, 60, 65, 67, 70, and 75 pct Nb, bal. Zr. In addition, a large range of grain sizes and a range of temperatures considerably greater than the range indicated by Rogers and Atkins phase diagram were examined. We will, however, report in detail only the results obtained for the Nb + 33 pct Zr and Nb + 25 pct Zr alloys at three grain sizes, two levels of oxygen contamination, and the temperature range 550 to 950°C. These data are most complete, but the other data are sufficiently complete to indicate the kind and magnitude of the variation of the transformation kinetics outside this range. The first and most difficult problem encountered in this inquiry was one of sample homogeneity. When Nb-Zr alloys are arc- or electron-beam-melted on a cooled copper hearth, solidification is sufficiently slow that there is appreciable coring in the cast structure and a large variation of grain size across the button thickness. Both these factors significantly affect the apparent reaction rate in the system. A two-step solution to the problem was attempted; an arc-melting and drop-casting technique has been developed by conald that greatly reduces the as-cast grain size and virtually eliminates coring segregation. Ingots made in this way exhibited no detectable (3 pct maximum) zirconium segregation. Before it was evident just how good this technique was, we attempted to supplement it with rather long-time, high-temperature annealing of the cast ingots. This annealing was carried out in evacuated and sealed (seal-off pressures < 1.0 x 106 torr) quartz capsules lined with tantalum foil at 1400 to 1450 C for 8 to 72 hr. There were two principal effects of this treatment: the grain size increased to a fairly uniform 150 p, and the surface and all grain boundaries near the surface acquired a film of a second phase, tentatively identified as an oxide (possibly additionally contaminated with silicon). There was no evidence that this 1400 C treatment had affected the zirconium segregation. High-temperature annealing was subsequently used only for grain-size control, but anneals of longer than 4 hr at temperatures greater than 1000°C were performed in dynamic vacuums (pressure no greater than 1.0 x lo torr). Any contamination resulting from these treatments was well below the limits of detection of our techniques. All samples, as cast, were cold-swaged to at least 85 pct reduction in area. The samples called cold-worked were tested as swaged. The minimum re-crystallization anneal for these alloys was about 12 hr at 1050 C; this produced an equiaxed grain diameter of about 4 to 8 P. Annealing for 4 hr at 1450°C produced a grain size of about 80 to 150 p; and annealing for 4 hr at 1650aC, close to the melting point of many of these alloys, produced a grain size of 0.5 to 1.0 mm. At all temperatures, the larger grain size was

Jan 1, 1967

By Arnold M. Lansche

[Selenium and tellurium are rare elements in the earth&apos;s crust. Their scarcity and primary dependence of supply on copper and lead output are in consonance with the relatively small quantities consumed at 1964 level. Production depending on output of copper and lead creates a supply situation- which would appear to militate against development of large-scale usage of these elements. However, supplies have adequately met. demand in the past decade and can be increased substantially if warranted by new uses for either selenium or tellurium Ores of copper and lead re the principal sources of selenium and tellurium. Byproducts from the smelting and refining of copper and lead concentrates are incorporated in refinery anode slimes from which selenium and tellurium metal are recovered.]

Jan 1, 1967