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By Robert H. Richards

If one holds up a prism, or wedge of glass, with narrow angle, say l° to 2O, and compares the transmitted image with the image seen above or below the prism, the former will be found to be thrown to one side by an amount varying with the angle of the wedge. Speaking of the two rays as the direct ray and the bent ray, we may say, that when the bisecting plane of the prism is at right angles to the line of sight, the angle between the direct ray and the bent ray will be constant for any given prism. If now we place a prism or wedge of glass in such a position that it half-covers the objective of a telescope, we shall obtain on looking through it two images of every object seen—one image by the direct ray, which comes through the unaovered half of the objective, the other which comes through the prism, and is the image by the bent ray. The angle of divergence of these two raps will be constant and unalterable, whether the telescope is directed to a near object, with its eye-piece at increased distance from its objective, or upon a ' distant object, with eye-piece nearer to the objective. That is to say, if the " throw " or apparent dislocation of the image is one foot in one hundred feet it will be two feet in two hundred, ten feet in a thousand, and so on. The usual form of stadia-telescope has at the focal point of the objective two spider-lines, placed at a definite distance apart and intended, let us say, to represent a throw of one foot in one hundred feet. But as the distance between lines remains the same, while the distance from the objective to the webs diflers with every variation in the distance of observed objects, it follows that the angle between the lines of sight which these two spider-lines define cannot be con stant, but must vary with every increase or decrease of distance be tween the instrument and the object viewed. If in Fig. 1 we represent three positions of the spider-lines by W1, W2, W3, and three objects corresponding to those positions by 01, O2, 0, at near, medium and distant positions, we see at

Jan 1, 1892

By Mark J. Klein, Robert A. Haggins

The pesence of a small amount of oxygen was found to cause significant grain growth restraint in 99.99 pct Ag. This behavior does not seem to be due to the presence of an oxide dispersion, but instead, to be related to the existence of minute amounts of oxygen m solid solution. In the amounts necessary to cause the observed effects it was found that oxygen migrates with extreme rapidity, probably along extended structural imperfections, such as dislocations and grain boundaries. A number of years ago, hastton' observed that the presence of small amounts of oxygen caused a retardation of grain growth in silver. He attributed this behavior to the formation of a fine dispersion of oxide particles by the internal oxidation of trace impurities. More recently, wood2 also found that oxygen caused grain growth restraint in copper. Experiments using both copper and a dilute solution of aluminum in copper led to the conclusion that the effects observed in copper were due to oxygen in solid solution, rather than an oxide dispersion. During the course of a study of the influence of several solutes on the recrystallization of Silver, the presence of minute amounts of oxygen was found to exert a very significant influence on grain growth. The experiments reported here were undertaken to determine whether the grain growth restraint observed in silver due to the presence of oxygen is related to oxygen in solution or to an oxide dispersion. Because of the difficulty in measurement and control of the very small oxygen concentrations causing the observed effects no attempt was made to treat the data in a quantitative manner. EXPERIMENTAL PROCEDURE AND RESULTS Silver of 99.99 pct purity was melted and cast into slabs in a vacuum of better than 0.1 p. These slabs were reduced 50 pct by cold rolling to a thickness of approximately 2.5 mm. Samples cut from this material were annealed in vacuum for a series of different times at 803°C. Another group of samples was similarly annealed in air. This temperature is well above the recrystallization range for silver of this purity (150" to 200°C). The mean grain diameter of these samples was measured, both in the center and near the edge, using the line-intel-cept method, and is shown as a function of annealing time in Fig. 1. The grain size in the centers of the samples annealed in air was found to be larger than that of the surface regions, while the grain growth behavior of the edge and center regions annealed in a vacuum was approximately the same. he edge measurements were taken at a distance of 0.2 mm from the surface.) A few samples were annealed in nitrogen. Their behavior was identical to that of the vacuum-annealed samples. Samples annealed in oxygen had approximately the same grain growth behavior as those annealed in air, leading to the conclusion that this behaviouer is due to oxygen. The microstructure of a sample annealed in air at 803" for. 30 minutes is shown in Fig. 2. The smaller grain size near the surface is characteristic of the iir-annealed specimens. Possible causes of this grain growth restraint in the air-annealed specimen are the formation of an im-

Jan 1, 1962

By C. W. Price, G. A. Wheeler

BERYLLIUM has always been a difficult material to prepare for metallographic examination. Severe surface deformation occurs during mechanical grinding and polishing, and a suitable bright-field etchant has not been developed. Certain electrolytic techniques produce satisfactory polishes, but these techniques have not yielded entirely consistent results. Furthermore, they have not provided satisfactory bright-field etches. The purpose of this paper is to describe an electrolytic polishing technique that can also provide an excellent bright-field etch. In addition to the bright-field etch, other advantages of this technique are excellent reproducibility, rapidity, and ease of polishing and etching. Experimental Technique. Specimen Preparation. Beryllium specimens may be either mounted or unmounted for the use of this technique, and the size of the specimens is not important. Specimens with surface areas up to 4 sq cm have been electrolytically polished and etched successfully with this technique. The most important requirement is to provide a good electrical contact to the specimens, preferably on the back of the specimens. When specimens are mounted, a hole is drilled through the back of the mount so that the specimen can be contacted with a probe. Conventional metallographic grinding techniques are used to prepare the specimens for electrolytic polishing. The specimens are ground on silicon carbide papers. Grinding through the usual series of papers down to the 600-grit paper is recommended to minimize the electropolishing time, although successful electrolytic polishes and etches have been obtained by grinding only through the coarser papers. A continuous flow of water is used during grinding to provide lubrication and to minimize contamination. Mechanical polishing is not required because high rates of material removal are achieved with the electrolytic polishing. Polishing. Most of the results reported in this paper were obtained with a "Disa-Electropol", manufactured by Struers Scientific Instruments of Copenhagen, Denmark. Satisfactory results have also been obtained with the conventional beaker technique of electrolytic polishing, however, the conditions applicable to both the Disa unit and the beaker technique will be described. Stainless-steel cathodes are used in both techniques. In the beaker technique, a flat cathode, with an area at least twice that of the specimen, is used, and the specimen is held parallel to the cathode. The distance between the specimen and the cathode affects the quality of the electropolish. This distance is set at about 7 mm in the Disa unit; the same distance produces the best results in the beaker technique. Specimen-cathode distances greater than about 20 mm produce considerable relief, while distances less than 7 mm cause high current densities that are difficult to control. An electrolyte of 2 pct HC1, 2 pct HNO3, and 2 pct HCIOl in ethylene glycol gives the most satisfactory results for both electrolytic polishing and etching. This electrolyte was developed by altering the composition of a similar electrolyte used for electrolytic lapping1 and adding HClOl to improve the etching characteristics of the electrolyte. Variations of from 0.25 pct to as much as 5 pct of each of the acids do not significantly affect the results, but the greatest amount of control and reproducibility can be obtained with the composition of 2 pct of each acid. Generally, lesser amounts of the acids produce lower polishing and etching rates, while greater amounts of acids cause high current densities that are difficult to control.

Jan 1, 1968

By A. L. Labbe, J. J. Donoso

Hard-won experience in the operation of smelting plants has pointed the way to the most efficient design and economical operation of a baghouse system for recovery of metallurgical fumes. This paper treats of the principles governing bag filtration and the factors most directly affecting bag life. Methods for temperature conditioning of smoke previous to filtering are described and the evolution of bag shaking mechanisms is brought up to date. THE history of the bag filter process, as applied to the recovery of metallurgical fumes, goes back to 1876, when the Lone Elms Works of Joplin, Missouri, erected a two-section baghouse for filtering fumes from Scotch Hearth lead smelting furnaces. By modern standards this installation was crude, but the basic operating principles remain essentially the same today. Great progress, however, has been made in the mechanization of the process by the use of devices for mechanically and automatically cleaning bags of adhering fumes, in the use of instruments for recording and controlling the variables affecting baghouse operation, and in the conditioning of the smoke previous to filtering. The principles governing bag filtration of fumes were investigated in 1929 by A. J. Seitz of American Smelting and Refining Co., and he arrived at the following conclusions: 1. Pressure on Bags at Constant Volume: The rate at which the pressure builds up in a bag with constant volume is directly proportional to the time, with the same kind of smoke and dust burden (fig. 1). 2. Pressure on Bags for Change in Volume: With equivalent smoke and dust burden, the bag pressure increases with the square of the increase in volume. The rapid increase in pressure is due to (1) increase in volume, (2) increase in amount of dust on the bag, and (3) more intense packing of the dust at higher velocity (fig. 2). 3. Effect of Intensive Filtering on Pressure: Since the pressure builds up directly as the square of the volume, with the same smoke and dust burden, it is necessary to shake the bags in that increased ratio, and vigorously, to keep the pores of the bag open and pressure down. This increases the wear on the bags and the dust losses. 4. Effect of Shaking on Dust Losses: The losses are in direct proportion to the frequency of shaking because the loss takes place immediately after shaking, before the bag seals itself. Tests show that with hand shaking the last 20 pct of the removable dust requires seven times more shaking per unit weight of dust than the first 80 pct. 5. Effect of Fineness of Dust on Losses: The fineness of the dust particle increases the loss in two ways: (1) When the bag is clean a greater percentage of it will pass through before the bag seals itself; (2) fine material penetrates farther into the bag pores, building up pressure faster and necessitating more shaking, thereby increasing the loss. The physical principle of operation of the bag filter is simply the separation of solids from a gaseous carrier through a porous medium, and is analogous to the filtration of solids from liquid carriers. Basically, the mechanics of the process are as follows: Fume-laden gases drawn from the operating fur-

Jan 1, 1951

By J. R. Lewis, J. H. Hamilton, V. D. Hanson, K. W. Heinlein, J. M.194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1'/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By J. H. Hamilton, J. R. Lewis, K. W. Heinlein, V. D. Hanson, J. M. 194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1'/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By T. P. Papazoglou, N. A. D. Parlee, W. C. Phelps

PRACTICAL designs for good "home made" molybdenum furnaces are hard to find in the literature. The one described briefly below left something to be desired but was good enough to operate as a rather long (30 in.) two-zone furnace for periods of over 2 months at temperature (1400° to 1600°C) without shutdown. The principal problem in our application was to establish a temperature profile with two distinct constant-temperature zones and no dip between the zones. This and other problems mentioned below would not arise, however, if both zones were operated at the same temperature and if a shorter furnace were acceptable. Since molybdenum windings oxidize very rapidly when heated in air, impervious refractory tubes were used for both the furnace core and the outer jacket tube shown in the accompanying figures. A protective atmosphere of forming gas (10 pct H2, 90 pct N2) was continuously passed through the gas-tight annular volume between the core and outer jacket. The water-cooled copper ends shown in Figs. 1 to 3 sealed the ends of the tubes. Annealed molybdenum wire (40 mil) was wound on the recrys-tallized alumina core (35.5 mm OD by 28.5 mm ID by 30 in.) at a spacing of 0.36 in. The spacing was reduced to 0.29 in. close to the ends to partially compensate for end heat losses. Fig. 2 shows two molybdenum-ground lead taps off the center of the winding to produce the two independect temperature zones. The ground lead was insulated with 1/4-in. alumina tubing and connected to the grounded copper end by passing the wires through the gas port tube and crimping each wire into notches cut at the mouth. To prevent sagging of the core at high temperatures, refractory "troughs" were cut from alumina and mullite tubing to cradle the core. The support "troughs" which were 6 in. long and located about 5 in. from each end of the annular volume also served as radiation shields. In view of the different expansion coefficients and operating temperatures of the core and outer jacket (3.5 in. OD by 3 in. ID by 30 in.), the ends of the tubes were smoothed and squared to produce good sealing surfaces, and the core was ground 1/32 in. shorter than the outer tube. The power terminals were passed through their respective conductor ports which were internally lined with alumina tubing and sealed with Armstrong adhesive A-6 leaving about 1 in. of protruding lead wire for external connections. (Adhesive A-6 is a two-component system consisting of an epoxy-type resin with a mineral fibrous filler and an amine catalyst.) Figs. 1 and 3 show the entire assembly of concentric refractory tubes, copper ends, and sealing o rings, held together with two brass flanges and 3/8-in. bolts with heavy springs for dimensional flexibility. The furnace shell (19 by 19 by 30 in.) consisted of a steel frame with sheets of transite forming a box. The copper ends protruded through two 3.75-in. circular holes cut at the center of each transite square end. Thermal insulation next to the transite exterior was effected by K-23 B&W bricks alld in a central volume of approximately 9 by 9 by 30 in. a row of K-30 bricks was lined up to support the hot assembly. The rest of the volume around the outer

Jan 1, 1965

By T. F. Baily

DURING the period of the War, in both this country and Canada, a number of attempts were made to make pig iron from steel scrap in the FIG. 1.-EXPERIMENTAL FURNACE. 5000 KW. CAPACITY: 150 TONS 2 PER CENT. SILICON PIG IRON PER DAY. electric furnace, and a considerable tonnage of white pig iron was produced, the silicon and manganese additions being made with ferroalloys; but it is believed that the experiments conducted at Canton, Ohio, in

Jan 1, 1930

By T. L. Berger, M. C. lnman, M. F. Rose

AFTER shock loading, mechanical twinning has been observed by several authors1-4 for a variety of fcc metals and alloys. It has been shown2,3 that low stacking fault energy materials deform primarily by the formation of mechanical twins and stacking faults during dynamic loading. Further, Inman, Murr, and Rose2 have shown that a correlation exists between the average hardness and the volume percent of mechanically twinned material in stainless steel. It has been observed in nickel' that mechanical twins are produced during shock deformation if the peak pressure associated with the shock wave is greater than 350kbar. It is the purpose of this work to discuss shock-hardening mechanisms in nickel and to determine to what extent the initial condition influences the final response. The "flying plate technique" was used to generate planar shock waves with peak pressures as high as 710 kbar. The details of the technique5" have been discussed elsewhere and will not be repeated here. The material used in this study was polycrystalline nickel of 99.98 pct purity and had an initial grain size of 0.18 mm. Prior to any deformation, all samples were given a stress-relief anneal of 6 hr at 600°C in a vacuum furnace. After annealing, the initial condition of the samples was changed by cold rolling individual samples from 0 pct to a maximum of 80 pct reduction in thickness. This provided samples with controlled initial histories for shock deformation. Immediately after shock loading, the specimens were slowed and quenched by suitable arrangement of buffer material and water bath. This technique minimized specimen damage and the effects of shock heating on the samples. Since deformation twins in high stacking fault energy fcc metals are only produced at high stress levels, the initial defect configuration prior to mechanical twinning should greatly influence the number of twins formed. In fact, Zukas and Fowler7 have found that prestraining in iron and steel greatly suppresses the formation of mechanical twins during the shock deformation process. Following the methods of Warren and warekois8 and Cohen and Wagner,9 X-ray diffraction profile analysis was performed to determine the incidence of stacking faults and mechanical twins in the deformed nickel. Analysis was made on samples in the annealed, cold-rolled, shock-loaded, and cold-rolled—shock-loaded states. The estimated accuracy of the peak position measurements is *0.01 deg in 2?. Therefore the accuracy in the measurement of the relative peak shifts between the (400) and (222) diffraction line profiles is on the order of *0.04 deg in 2?. None of our measurements indicated peak shifts greater than 0.04 in 28. This would correspond to a fault density of -1 pct and sets the limit on the sensitivity of our instrument. Similarly, analysis of the shift in the center of gravity of the diffraction line profiles yields the minimum measurable value of the twin fault probabilitya to be -0.002. Fig. 1 shows mechanical twinning in nickel shock-loaded to 400 kbar. In this photograph, fault intersections are clearly visible. The thickness of the individual fault lamellae was determined from the image widths of their lines of intersections by the

Jan 1, 1968

By A. G. H. Andersen

A FEW years ago, while the writers worked on the constitution of ternary and quaternary alloys consisting mainly of iron, chromium and silicon1,2 some information on microstructures and hardness, not incor-porated in the published work, was obtained. It has not been possible to do any further work on the properties of the alloys. The information possessed is hereby presented for the interest it. may have in connection with the constitution of the alloys.

Jan 1, 1937

By W. D. Forgeng, R. E. Grace

The rate of copper corrosion in liquid lead was studied as a function of flow rate and temperature in a closed isothermal loop system between 350" and 500°C. Kinetic data were interpreted with the Nernst-Brunner equation and the Chilton-Colbum mass transfer analogy to heat transfer. Over the range of temperatures and flow rates employed the solution rate constant varied from 2 x 10-3 to 20 x 10-3 cm per sec. The rate constant was found to vary with the 0.76 power of the Reynolds number in the range 7500 < Re < 40,000. The results of this investigation indicate that the Cu-Pb corrosion process is controlled largely by diffusion across a boundary layer. particularly at the higher temperatures. The effective boundary layer for diffusion is found to be approximately three to four times smaller than the momentum boundary layer. CORROSION by liquid metals has long been a problem during melting, static containment, and casting. Recently, considerable emphasis has been placed on the understanding of liquid metal corrosion mechanisms in an effort to prevent mass transfer and eventual failure in liquid metal heat exchangers and in liquid-metal-fueled nuclear reactors. The present investigation was conducted to determine the relationships which exist between corrosion kinetics and hydrodynamic conditions in an isothermal loop containing a simple liquid metal-solid metal corrosion system. The Cu-Pb system was chosen for its simplicity and ease in handling. Previous isothermal Cu-Pb corrosion studies1-3 have been conducted in both static and dynamic systems but not under conditions which allow easy "scale-up" to engineering systems, such as liquid metal heat exchangers. On the other hand, most liquid metal loops which have been used to evaluate mass transfer rates4 have been driven with temperature gradients under conditions of natural convection and unknown flow rates. The primary contributions of this research are twofold: 1) to evaluate the design of a simple isothermal loop corrosion system and 2) to investigate liquid metal corrosion behavior under well-defined hydrodynamic conditions. MATERIALS AND EXPERIMENTAL WORK The apparatus used to measure the Cu-Pb corrosion rates is shown schematically in Fig. 1. The loop system consisted of 1) a reservoir made of Type 316 stainless steel containing a centrifugal pump and 2) stainless-steel tubing containing a removable copper test section. The copper test section was OFHC hard-drawn tubing, 11.0 0.05 cm in length and 0.775 ±0.005 cm ID. A Type 316 stainless-steel venturi meter* was inserted in place of the copper test sec- tion in order to determine the lead velocity as a function of pump speed at each temperature. The Cu-Pb system was chosen for this study because the solubility of copper in liquid lead is large enough to measure accurately but small enough that the diameter of the copper tubing was not changed significantly by corrosion. There are no intermediate phases formed in the system, and the solid solubility of lead in copper is vanishingly small so that back-diffusion cannot occur. Molten lead (corroding grade; 99.94+ pct Pb, 0.0015 max pct Cu) was charged from an auxiliary melting furnace into the system under an argon cover to minimize oxidation. The lead was then maintained at constant temperature by a stirred molten salt bath in which the loop was immersed to a depth just above the lead level in the reservoir, Fig. 1. The total volume of lead in the system was 320 2 cu cm.

Jan 1, 1969

By Robert J. Linney

Reserve Mining Co. produces 5 million tons of iron ore pellets per year. The finished product runs about 65.50 pct Fe, with 7.75 to 8.00 pct SiO2. Less than 12 pct arrives at the blast furnaces smaller than 3/8 in. This plant was originally designed to produce 3.75 tons annually-not 5 million. To handle the increased tonnage a fifth fine crushing line was added, and some very minor changes were made.

Jan 9, 1959

By B. Ramaswami, Y. Nakada, R. Schaefer

EARLY investigations1 of the tensile properties of fcc single crystals did not reveal the easy glide region of crystals with axial orientations inside the standard stereographic triangle. With the production of high-purity metal, a region of easy glide became known1 Consequently, modern work of fcc single crystals ordinarily uses high-purity metal. In order to determine whether further purification of aluminum beyond 99.992 pct purity would result in significant alteration of the tensile properties of single-crystal specimens, metal was zone-refined and crystals grown from the purified material were compared to crystals grown from the original material. Some similar studies were made of gold and silver crystals. Aluminum. The material to be zone-refined was cast into bars 1/4 by 1/2 by 8 in. A zone, approximately 1-1/2 in. in length, was passed over each bar eighteen times at a rate of 6 in. per hr. Melting and stirring were carried out by an induction heater under a vacuum of approximately 10"5 mm Hg, the metal being held in a graphite container. he last 3 in. of each processed bar were then cut off and discarded. No quantitative determination of purity was made, but the effectiveness of the zone refining was tested by a method first suggested by Walton, Tiller, Rutter, and winegard3 depending on the relation between cell structure and impurity content. Rutter and Chalmers4 have shown that a cellular substructure is associated with the presence of impurities during solidificatibn, and have shown how solute rejection at the freezing interface can lead to such an association. Thus the observation of cells gives evidence of impurities being present. To look for cells, the last quarter of a zone-refined bar was cut off, remelted, and grown rapidly into a single crystal. It was then heavily electropolished and thermally etched.~ Cells were revealed by the presence of many parallel rows of etch pits. The same process was carried out with material from the first half of the bar, but no cells were detected. These observations give evidence that a significant amount of impurity material was moved to the "dirty" end of the zone-refined bars. Further indication of purification was obtained from recrystallization experiments. It was found that, after cold rolling to a 97 pct reduction in thickness, the material from the pure end of the bar re-crystallized at room temperature, but the material from the "dirty" end did not. Other investigators6 have previously reported that zone-refined aluminum recrystallized at room temperature. The purified material was cast into 1/4 by 1/4 in. bars, and seeded crystals were grown in a (100) orientation and in a single slip orientation where the Schmid factor is m =2. The crystals were then cut to the desired length, annealed for 24 hr at 610°C, furnace-cooled, and etched in Tucker&apos;s etch. The crystals were tested in a floor-model Instron machine, using a gage length of 4 in. and a strain rate of 1/2 pct per min. The resulting curves are shown in Fig. 1. It is seen that the differences between the curves for zone-refined and original-purity materials are small. These differences are within the scatter observed for crystals of the original purity. In particular, the zone refining does not significantly affect the onset of the different "stages" of the curves.

Jan 1, 1964

[TWEEDY, A. M., and VOGEL, FELIX A.: The Briquetting of Flue Dust in the United States by the Schumacher Process, XLVII, [viii], 338. TWEEDY, GEORGE A., and BEALS, ROGER L.: Cyanide-Plant and Practice at the Minas del Tajo, Rosan&apos;o, Sinaha, Mez., XLI, [xl], 324. Twin Falls plant, Peninsular Power Co., LI, 298. Twins, deformation lies in manganese steel, LI, 882. 2,000-Ton Leaching Plant at Anaconda (LAIST and ALDRICH) (With Discussion), LV, [xxiii], 866.]

Jan 1, 1918

By J. W. Pugh

PREFERRED orientations in cold-rolled body-centered-cubic metals appear to have remarkable similarity. The as-cold-rolled textures in iron,1,2 silicon-iron,3 mild steel,4 molybdenum, vanadium,&apos; chromium,&apos; tantalum,&apos; tungsten,&apos; and a-zirconium-niobium alloys10 have been determined. These textures have, in some cases, been represented by congrUOUS pole figures and have all been described in one or both of two ways. They have frequently been reported as predominantly (100) <011> orientation

Jan 1, 1959

By Neil H. Bostick

A specialized electronic data system processes data taken from normal gologic maps and stratigraphic tables to yield predictions of uranium resource location for an area such as much of the United States. Strata underlying specific quadrangle cells of the earth are evaluated for uranium favorability on the basis of selected criteria of lithology, structure, and stratigraphy but not specific past production in the area. Stored data are computer processed to yield volume of likely host rock beneath each cell to a given depth and at a given favorobility level. Results are plotted by machine directly as maps. A specialized and simplified procedure for predicting approximate uranium resources has been devised at Stanford Research Institute (SRI). The prediction is based on general geology beneath specific area cells of the earth in order to provide complete volume coverage; the strata underlying these areas are evaluated for uranium mineral favorability on the basis of a few selected geologic criteria in order to allow rapid data input. In order that the predictions be generally applicable, they are based on geological factors only, not past ore production of a given cell. Since the bulk of U.S. resources are in elastic host rocks, only such strata are considered in this work. This paper describes the structure and some details of the resource system and how it was tested in an area of the western United States.

Jan 1, 1971

By J. J. Scott

During 1967 wage settlements and strikes ran at a high rate and it has been estimated that union gains of 4 ½ % to 5% per year have been attained in many instances. When these events are coupled with the need to operate on lower grade and deeper deposits the need for technological advances in underground mining becomes obvious. Emphasis has been on ways to reduce the number of manshifts required per ton of product. Most significant advances have been made in rapid advance of primary mine openings, development of mobile, rubber-tired, front-end loaders and hauling units, high-speed drilling with rotary and rotary percussive units and a serious look at nuclear devices to break large volumes of ore cheaply. Boring Developments in bit manufacture and metallurgy have allowed continuous boring-type machines to be used on harder formations than was deemed feasible in the past. Hecla Mining Co. has been using raise boring equipment for drilling 9 7/8 in. diam pilot holes in raises which are later reamed in an upward direction to 48 in. diam. The machine used is a J. S. Robbins model 42R raise borer which has been used successfully to drill two relatively long 48-in. diam holes. The first was 520 ft in length at a dip of 76º, and the second was a 785-ft vertical hole.

Jan 2, 1968

By Charles Means

A THOROUGHLY modern coal-handling system has been installed at the Pittsburgh Terminal R. R. & Coal Co.&apos;s new No. 8 shaft, or Coverdale mine, about 11 mi. (17.7 km.) from Pittsburgh on a spur f the Montour R. R. The Pittsburgh seam, which is the one worked, is reached at a depth of 342 ft. (104 m.). Power is delivered to the mine at 22,000 volts over a transmission line of the bow-arrow type of construction. A private telephone line connecting the various mines and the main office in Pittsburgh is carried on the same poles. The high-tension line is brought to an outdoor, step-down, transformer substation, where three 833-kv.-a., single-phase, 60-cycle, oil-insulated, self-cooled transformers reduce the potential to 2300 volts. The transformers are protected by electrolytic, aluminum-cell lightning arresters, horn gaps, and choke coils. High and low-tension busses, together with fuses, choke coils and air-break horn-gap switches with their remote-control operating mechanism are mounted on a substantial structural-steel tower 25 ft. (7.6 m.) in height. MINE SUBSTATION The transformer substation is adjacent to a pressed-brick building that houses direct-current substation equipment together with switchboards, etc., for the control and distribution of power. This building also contains the hoist for the men and supplies.

Jan 9, 1921

By H. B. Muller

Single stage autogenous grinding of run-of-mine ore was selected by Broken Hill South Ltd., parent Company of Cobar Mines Pty. Ltd. for the concentrator to treat 60,000 tons per four weekly period of copper and complex copper-lead-zinc sulphide ores from the C.S.A. Mine at Cobar, Australia. Ore is mined by mechanised horizontal cut-and-fill stoping and is prepared for grinding by crushing under- ground in a 60 in. x 48 in. jaw-crusher set at 5 in. closed side setting. The ore is reduced to 80% minus 76 microns in two parallel sections each of one 22 ft. dim. x 7 ft. long autogenous grinding mill, powered by a 1,250 H.P. motor, operating in closed circuit with a single stage 20 in. dim. cyclone. Broken Hill South&apos;s original concept of the concentrator required the maximum use of reliable instrumentation, centralised control, and the minimum of manning. To achieve this objective the control room, housing the process recorders, alarms, closed circuit television monitors and all the control equipment, overlooks the grinding and flotation sections, and the whole plant is serviced by a public address system. The grinding units and associated equipment are located in the open and are serviced with the aid of a mobile crane and a mobile hydraulically-operated liner lifting device.

Jan 1, 1970

By J. Howard Boyce

INTRODUCTION The Tsumeb Mine is situated in the northern part of South West Africa, approximately 335 air miles north-east of the port of Walvis Bay on the Atlantic Coast, and 240 air miles north of Windhoek, the capital of South west Africa. The mine came into operation in the early 1900&apos;s and has operated continuously since then except during the two world warm and the depression years of the thirties, Tsumeb Corporation purchased the property in 1947, installed a new plat, and ham built a modern tom. In addition to the concentrator which began operations in 1948, a copper smelter was started in 1962 and a lead smelter in 1963, Tonnage of ore treated in the concentrator has varied through the years, reaching a maximum of 70,000 tons per month with a current mined tonnage of approximately 50,000 tons per month. Copper concentrates, lead concentrates and zinc concentrates are produced, and for a number of years the production of a germanium enriched concentrate was an important feature of the operation, Both copper and lead concentrates are smelted in Tsumeb, while the zinc concentrates are shipped overseas. Ore change, and the installation of the smelters in Tsumeb have necessitated modifications in the concentrator operation. This paper will describe current plant practice in some detail and will touch briefly on the techniques used from 1954 to

Jan 1, 1970