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By Nicholas J. Grant, Yoichi Ishida, Arthur W. Mullendore

An inert particle -marker technique was developed to provide a direct measurement of grain boundary sliding during creep in tile interior of aluminum specimens. Groin boundary sliding in the interior as measured by this technique was found to be nearly the same or slightly lower than that measured on the surface. These data disagree with those obtained by the grain-counting technique developed by W. A. Rachinger. In tests where both techniques were used, the grain-counting technique gave large and variable values of grain boundary sliding. It is shown that the grain-counting technique gines erroneozcs results because of a preferred direction of grain growth during creep. A question which has long plagued those who study grain boundary creep is whether or not surface measurements yield a true representation of the deformation in the interior of a specimen. The first attempt to measure grain boundary sliding in the interior of a creep specimen was made by Rachinger.1 His technique is based on the measurement of average grain diameters, both parallel and perpendicular to the tension axis, followed by a calculation of the grain elongation from these measurements. The difference between total elongation, Et, and the calculated grain elongation, Eg, is considered to be the result of grain boundary sliding, Egb. Surprisingly large values of Egb/ Et (90 pct) came out of creep tests of pure aluminum at temperatures above 250°C (480°F), utilizing the Rachinger technique. This value is very much larger than the values of 8 to 15 pct measured by numerous investigators on the surface of a specimen. Values obtained on the surface by Rachinger, however, were consistent with those measured by the displacement of reference scratches. Chaudhuri and Grant2 repeated the Rachinger method for A1-10 pct Zn, and found the same large values of Egb/Et inside the specimen. However, from purely geometrical considerations, grain boundaries cannot slide without grain deformation; therefore, they questioned the validity of Rachinger's grain-counting method. Several authors2,3 have suggested that grain boundary migration during high-temperature deforma- tion may tend to restore the equiaxed grain shape in order to minimize interfacial energy. Rachinger4 tested this hypothesis and found that it did not hold. Grains elongated by rapid extension retained their elongated shape during a subsequent slow creep test. It appears that some other factor was responsible for the abnormally high values of .Egb/El. McLean and Gifkins5 made a survey of the effect of grain size on the ratio Egb/ Et, and proposed that the high values were the consequence of a small grain size. They failed to explain the low value of Egb/Et which Rachinger observed on the surface of the small grain size specimens, and suggested that the surface effect on the value of Egb/Et does not always occur. Many of the disagreements arising from Rachinger' s work stem from the fact that there was only one method of estimating grain boundary sliding in the interior of a specimen, and that method was of an indirect nature. If a marker line of some sort could be introduced inside of the specimen, one could make measurements just as clearly as on the surface. In the present investigation, the technique utilizes a layer of finely dispersed oxide particles inside the specimen introduced by hot press-bonding of two pieces of aluminum. Rachinger attempted a similar technique but with the oxide distribution he obtained there was so much interference with grain boundary motion that quantitative measurements were not attempted. MATERIALS AND EXPERIMENTAL PROCEDURE A) Materials Preparation. Two pure aluminum rods (impurity content in percent: Si, 0.002; Cu, 0.004; Fe, 0.002; Mg, 0.000: Zn, 0.000; V. 0.001) 1 in. diameter by 2 in. high were jacketed end to end in a 4-in.-diam cylinder of commercial-purity aluminum, 4 in, high, which had a l-in.-diam hole. The l-in.-diam mating surfaces of the two pure aluminum rods had been electropolished initially. This interface provided the oxide-film internal -marker plane after the hot press-bonding process, see Fig. 2. The composite was annealed for 1 hr at 900 F and hot upset more than 50 pct reduction in height in one step. The resulting slab was then rolled to 1/2 in. thickness; the pure aluminum test material was cut out and annealed at 900°F for 1 hr and cold cross-rolled to 1/8 in. thickness. The sheet was machined into specimens with a 1/8-in.-square cross section and a 1/2-in. gage length. The following heat treatment was given the specimens: Specimen 5A series: annealed at 1000°F for 5 min for grain-size control and then cooled to 700°F and held for 15 hr for stabilization.

Jan 1, 1965

By T. M. Allen, W. H. White

THE Greenwood-Grand Forks area of southern central British Columbia, known as the Boundary District, has a long history of mining exploration and production. At the turn of the century this was the premier copper mining camp in the British Empire, its total production amounting to some 20 million tons. Most of this ore came from the great Granby mines at Phoenix, but the Motherlode mine at Deadwood camp, 6 miles to the west, and several mines in Summit camp, 5 miles north of Phoenix, made important contributions. The large deposits were exhausted in 1918 and the district since has seen only desultory exploration and salvage operations. The orebodies are mineralized skarn zones in limestone members of a thick series of Upper Paleozoic sedimentary and volcanic strata. Chalcopyrite is the primary ore-mineral. Copper carbonates and silicates occur sparingly in outcrops, but the oxidized zone generally is very shallow. Much of the surface is mantled by glacial drift which in most places ranges in thickness from 2 to 15 ft. In some of the hanging valleys, however, the glacial drift may be as much as 100 ft thick and may assume drumlin-like forms. In 1951 an ambitious program aimed at the discovery of new orebodies and important extensions of abandoned deposits was launched by Attwood Copper Mines, Ltd. In this district so thoroughly searched by an earlier generation of prospectors, any orebody which had remained undiscovered must have little or no surface indication. Consequently, in addition to the basic detailed geological work, the program of exploration included magnetometer and self-potential surveys. Geological bets and geophysical anomalies were tested further, prior to diamond drilling, by a study of copper distribution in tree twigs and/or in the soil. The soil sampling and analytical methods used and some of the results seem of sufficient importance to warrant this paper. The authors had done some plant sampling in this and other districts, using the dithizone neutral-color-end-point method (Warren and Delavault, 1948, 1949; White, 1950),1-3 but they were unfamiliar with its soil application. Finally, after much experimenting in the field, they adopted the methods described here. These methods are not entirely original or defensible on theoretical grounds, but under field conditions of rapid sampling and analysis the results are reliable enough to be of use. Fig. 1, which shows the results of duplicate analyses of duplicate soil samples taken at 50-ft intervals across an anomalous zone, indicates the relative dependability both of the sampling and analytical methods. Sampling and Analytical Equipment A 2-ft piece of 1-in. solid drill steel, one end sharpened to a broad, conical point. The steel is marked at 1 ft from the point. A 2-ft piece of ½-in. black iron pipe, one end filed to a bevelled cutting edge. The pipe is marked at 1 ft 3 in. from the cutting end. A 3-lb hammer. A plastic or rubberized sheet about 18 in. square. Moisture-proof assay pulp envelopes. A 10-mesh seive made from window screen with the paint burnt off. A small assay spatula. A pan balance sensitive to 10 mg. Two ignition trays about 4 in. square, made of sheet iron turned up along the edges. A Coleman two-burner gasoline stove. An asbestos board about 5x8 in., used as a hot plate on the gasoline stove. A circular aluminum rack to hold 8 test tubes while refluxing (design of Almond and Morris). Pyrex Glassware Large refluxing test tubes, 25x200 mm, marked at 40 ml volume. Breakers, 20 ml. Pipettes, 1, 5, and 10-ml capacity. Graduate, 50 ml. Shaking cylinders, 100 ml, glass stoppers. Burette, 25 or 50-ml capacity, with holder. Chemical Supplies 1 N sulphuric acid. Hydroxylamine hydrochloride, solid crystals. Fisher Alkacid test paper. Copper standard solution. Dithizone standard solution 60 mg per liter. Water reasonably free of metals. Soil Sampling Method: The problem of how to take a soil sample is extremely crucial. The method outlined below, adopted after a number of tests, has the advantages of uniform pattern, uniform depth, and uniform size of sample. The area to be tested was marked off by chain and compass lines 100 ft apart, normal to the strike of possible ore deposits. Numbered stakes were set at 50-ft intervals along these lines and a soil sample was taken at each stake in the following manner. The drill steel was driven into the ground normal to the slope of the surface to the marked depth of 1 ft, moved slightly from side to side, then carefully withdrawn. The iron pipe was inserted to the bottom of this hole, tapped down to the marked depth of 1 ft 3 in. and withdrawn; the 3-in. soil plug in the

Jan 1, 1955

By E. V. Potter

For a nurnber of years, it has been known that manganese made by electro-deposition under certain conditions is ductile while under other conditions it is very brittle. The ductile metal is gamma manganese normally stable only between 1100 and 1138°C1; the brittle metal is alpha manganese, stable up to 727OC. The ductile metal is not stable, but gradually changes to the brittle form; the time required to complete the transfornlation is about 20 days at room temperature. Other observations have indicated that the transformation is completed in 10 to 15 min. at about 125°C, while at — 10°C, no appreciable change occurs in 9 months. The properties of gainma and alpha Illanganese in the pure state are ordinarilj difficult to determine because the gamma structure cannot be retained by normal quenching procedures and alpha manganese is so brittle, it is difficult to obtain specimens free from flaws. In a recent investigation2 some properties of gamma and alpha manganese were determined by studying the ductile electrolytic metal and determining the changes in its properties as it transformed to the brittle alpha form. These investigations provided an excellent opportunity for following the progress of the transition and studying its mechanism. The results of a series of such investigations are reported in this paper. Procedure Various properties of manganese were determined starting with the metal in the original ductile gamma form and following the subsequent changes in its properties as the metal transformed to the brittle alpha form. These observations were made at various temperatures, the data providing information regartling the mechanism of the transformation as well as the effect of temperature 011 the transition rate. Structure and resistivity values gave the most significant results, so this paper is concerned primarily with them. The structure was studied microscopically as well as by X ray diffraction. The resistivity was determined on strips of the metal by measuring the potential drop across a given length of the specimen. Current was passed through the specimen by wires soldered to its ends, and the potential connections were made by wires looped around the specimen near its center. The current was determined by the potential drop across a standard resistor connected in series with the specimen, the potential drop being measured on a potentiometer. In the temperature range from room temperature to 100°C an ordinary drying oven was used to heat the specimen. This was entirely satisfactory except at 100°C, where the time required to heat the specimen was long compared to the transition time, making the initial section of the resistivity curve unsatisfactory. To overcome this limitation, at 100°C and higher a thermostatically controlled oil bath was used to heat the specimens. The block on which the specimen was mountetl was plunged into the hot oil at the start of each test. The heating time was thereby reduced from 5 min. to about 6 sec, and dependable resistivity values could be obtained through 160°C. At this point the whole transition, including the warm-up time for the specimen, required only about 20 sec and it was not considered worth while trying to extend the temperature range further. Aside from the heating problem, the problem of making a sufficient number of accurate resistivity determinations became more and more difficult as the temperature was raised. Using the manually operated potentiometer, 100°C was about as far as it was possible to go. At this temperature and above, a self-balancing photoelectric recording potentiometer was used. Its response was quite rapid, and it proved to be entirely satisfactory all the way through 160°C, where the tests were stopped because of the specimen heating problem rather than any limitation of the potentiometer recorder. The metal used in these tests was prepared at the Salt Lake City laboratory of the Bureau of Mines. The method of preparation is discussed in a paper by Schlain and Prater.3 The sheets were about 2 3/8 by 5 3/16 in. and varied from 10 to 16 mils in thickness. They could be cut readily into pieces suitable for the various tests. X ray and microstructure determinations were made on pieces about 1/8 to 1/4 in. wide and about 1 in. long, while resistivity measurements were made on strips as long as possible and about 55 in. wide. The thickness of each sheet was not uniform over all its surface. This had no bearing on the X ray and microstructure determinations, but sections as nearly uniform and free from flaws as possible were chosen for the resistivity determinations. The gamma manganese was electro-deposited at 30°C, the time of deposition ranging from 5 to 12 hr for each sheet. Whenever possible, the tests were started directly after the metal was stripped from the cathode; otherwise the sheet was placed immediately

Jan 1, 1950

By M. V. Nevitt

In the systems Ti-Mn-O, Ti-Fe-O, Ti-Co-O, and Ti-Ni-O the bounda.r-ies of the Ti2Ni-type phases were determined at one or more temperatures and the variation of the lattice parameter with oxygen content was determined. Densities were calculated from the lattice parameters and compared with measured density values. The: results indicate that the occurrence of the phase in these systesms can be correlated qualitatively with valency electron concentration, and that the role of oxygen is that of an electron acceptor. The lower limit of oxygen solubility appears to be determined by the valencies of Mn, Fe, Co, and Ni, while the maximum oxygen concentration coincides with the filling of the 16 (c) positions of the O 7h - Fd 3m space group. THE suggestion has been made by several investigators'" that the phases having the cubic E9,-type structure, and known as 17-carbide-type, double-carbide-type and Ti,Ni-type, are members of a family of electron compounds. This concept has been given additional support by recent work8 in which new isostructural phases involving second and third long period combinations were found, and which provided further evidence of the regularity of occurrence of the phase in terms of periodic table relationships. In this laboratory attention has been focused on the isomorphs containing titanium, zirconium, or hafnium, and the role that oxygen plays in their occurrence. In some binary systems Ti,Nitype* phases occur having the formula A,B where A is the titanium group element. Based on previous workq and the present investigation, oxygen is known to be soluble in two of these binary phases, Ti,Co and Ti2Ni. It is probable that oxygen is also soluble in the other phases of this kind. In other binary systems the Ti,Ni-type phase does not occur, but does occur in the corresponding ternary systems with oxygen .3-5 The experiments described here were performed to determine whether the occurrence and composition of certain of the Ti,Ni-type phases could be related to an electronic effect and whether oxygen's stabilizing role is exerted through an influence on the electron: atom ratio. The ternary systems Ti-Mn-O, Ti-Fe-O, n-Co-O, and Ti-Ni-O were selected for study for two reasons: First, several schemes have been proposed for first long period elements which, although not in quantitative agreement, show a generally consistent trend for the variation of valency with atomic number. Although for a transition metal the term valency is difficult to define and is generally not a constant number which can be applied to all alloys, it is usually assumed to be an index of the number of electrons per atom involved in metallic cohesion. Second, the determination of the Ti2Ni-type phase boundaries was facilitated by the fact that the phase relations in several of these ternary systems have been investigated by other workers."' EXPERIMENTAL PROCEDURE___________________ The alloys were prepared by arc melting crystal-bar titanium, reagent grade TiO, and electrolytic manganese, iron, cobalt, and nickel. Each button was remelted at least three times. The metals had a minimum purity of 99.9 pct except the nickel whose purity was 99.4 pct, the major impurity in this instance being cobalt. The preparation of the manganese alloys was attended by the customary difficulties associated with the vaporization of manganese. The technique used in this case was to add approximately 10 pct extra manganese to the original charge and to continue remelting the button until the final weight was in agreement with its intended weight. At least three alloys in each system were analyzed chemically and the results, even for the manganese alloys, were in good agreement with the intended compositions. A few additional alloys in the Ti-Mn-O system were prepared by the sintering of mixed powders in evacuated quartz tubes followed in some cases by arc melting. For annealing, the alloys were wrapped in molybdenum foil and placed in fused silica tubes containing zirconium chips. The fused silica tubes were evacuated at room temperature to a pressure of 1 x l0-6 mm of Hg and sealed. These capsules were then annealed for 72 hr at an external pressure of 5 x 10-5 mm of Hg in a vacuum furnace whose temperature could be controlled to + 1°C. The success of this procedure in avoiding significant oxygen or nitrogen pickup was indicated by the bright, ductile condition of the molybdenum foil and by the complete absence of a microscopic reaction layer on the specimens. This method did not permit rapid quenching of the specimens but in no case did metal-lographic examination indicate that a solid-state transformation had occurred on cooling. Metallo-

Jan 1, 1961

By P. G. Shewmon, F. N. Rhines

THE determination of the self-diffusion coefficient of magnesium has been made possible recently by discovery1-1 of a radioactive isotope, Mg28 having a half-life of 21.3 hr,1 and subject to manufacture in useful quantity. In the present research this material was condensed from the vapor phase upon a surface of high purity magnesium. The progress of diffusion of the tracer atoms into polycrystalline magnesium was followed by machining layers and measuring the change in the intensity of radiation as a function of the distance of each layer from the surface. The self-diffusion coefficient was found to be 2.1 X 10-8 sq cm per sec at 627°C, 3.6 X 10-9 sq cm per sec at 551°C, and 4.4 X 10-10sq cm per sec at 468°C; the activation energy is about 32,000 cal per mol. Experimental Procedure Since there was no other published measurement of a diffusion velocity in any magnesium-base material, is was necessary to employ a number of new experimental techniques. The short half-life of Mg28 made it necessary to complete the entire experimental procedure within three or four days. This meant that the work had to be done where a cyclotron was readily accessible and that all operations, prior to the diffusion heat treatment, had to be so designed as to minimize their time requirements. Unusual problems were imposed also by the chemical reactivity of magnesium, its high vapor pressure, and the fact that no satisfactory method for elec-trodepositing magnesium on magnesium is presently available. Finally, the machining and handling of the easily air-borne radioactive-magnesium chips involved certain health hazards, resulting in the need for further experimental restrictions. Preparation of Mg28 The Mg28 was produced in the Carnegie Institute of Technology syncrocyclotron by the neutron spallation of chlorine.5 his involved bombarding a 2 gram crystal of high purity NaCl with a beam of 350 mev protons for a period of 2 hr, after which the crystal was dissolved in warm water and the Mg28 was concentrated and purified by chemical means (see Appendix). About 50 microcuries of Mg28 thus were obtained in the form of magnesium oxinate (8 hydroxyquin-olatc?), which was ignited in air to produce MgO. This in turn was reduced to magnesium metal vapor, by the method of Russell, Taylor, and Cooper," in the vacuum apparatus shown schematically in Fig. 1. Here the essential part is a tantalum ribbon, slightly dished to receive the MgO. The ribbon, pre- viously outgassed at high temperature, is heated to about 1700°C by passing an electric current through it, whereupon tantalum oxide is formed, magnesium vapor is released almost instantaneously, and condensed partly upon the diffusion sample. Diffusion-Sample Preparation: Hot-extruded magnesium rod, 21/32 in. round was used in making the diffusion specimens. The magnesium analyzed as follows: 0.004 pct Al, 0.027 pct Fe, 0.040 pct Mn, 0.0004 pct Cu, 0.0002 pct Ni, and less than 0.01 pct Ca, 0.0004 pct Pb, 0.0011 pct Si, 0.001 pct Sn, and 0.001 pct Zn. A brief study of the crystal texture of this material revealed a sharp fiber texture with the (001) plane roughly parallel to the extrusion axis. Cylindrical samples 1/2 in. long by 5/8 in. were machined from this rod, the end faces dressed on 3/0 emery, and lightly etched with 20 pct HC1 in water. These samples then were annealed for at least twice the intended time of diffusion, at the intended diffusion temperature, in order to stabilize the grain structure at about 1 mm average diameter. The annealing treatments were conducted in argon in the same apparatus and in the same manner as the subsequent diffusion treatments, which will be described presently. Thus, a strain-free plane surface was produced, but there remained a layer of MgO which had largely to be removed before the layer of Mg28 was deposited. Most of this layer was taken off by two light passes over 3/0 emery paper. The balance of the oxide and a thin layer of metal were then removed by etching 5 to 10 min in 4 pct nital (4 pct HNO3 and 96 pct ethyl alcohol) made with absolute alcohol. There followed immediately three quick rinses in: 1-49 1/2 pct methanol, 49 1/2 pct acetone, and 1 pct formic acid, 2-50 pct methanol and 50 pct acetone, and 3-pure benzene. This procedure is essentially that of Sturkey.7 The resulting surface, which was of almost elec-tropolished brightness, remained plane and was free of cold work. It could be kept clean by storing under benzene, or in a desiccator; short exposure

Jan 1, 1955

By K. P. Gupta, A. K. Pal, L. Chandrasekaran, U. P. Singh

The Mn-Sn binary system, investigated at the high-manganese end and between 500° and 1000° C, shows four phases at temperatures below 727"C, namely the u Mn, the p Mn, the Mn3 Sn, and the Mn, Sn phases, while at higher temperatures only the last three phases remain stable. The solubility of tin in a Mn is very small and the maximum solubility of tin in P Mn phase appears to be abollt 10 at. pct Sn. The solubility range of the Mn,Sn and the Mn, Sn phases is 23.0 to 26.0 and 37.0 to 40.0 at. pct Sn, respectively. The lattice parameter of the 13 Mn phase increases with increasing tin content. The Mn, Sn thase is hexagonal with and appears to be basically of the Ni3 Sn type structure except that for quite a few X-ray diffraction lines the calculated and observed relative intensities do not agree well. The Mn-Sn binary system has been studied by several investigators.''~ Their results indicate that three intermediate phases, namely, the Mn3Sn, Mn,Sn, and MnSn, phases, exist at high concentrations of tin. However, so far the proper phase equilibrium has not been established at the manganese-rich end and very little data is available for the composition range between pure manganese and Mn3Sn. Moreover, earlier investigators differ in their opinion about the exact composition range at which the Mn3Sn and Mn,Sn phases appear and some doubt has been cast by some investigators3 regarding the structure of Mn3Sn phase which has been reported to be isotypic with Ni3Sn (Mgscd type) structure. In this investigation an attempt has been made to establish the proper phase equilibrium between pure manganese and Mn + 50 at. pct Sn composition in the temperature range of 500" to 1000°C. PROCEDURE The raw materials used were from three different sources. For exploratory work five alloys containing 5, 10, 15, 20, and 25 at. pct Sn were prepared using 99.9 pct pure manganese and tin supplied by E. Merck & Co., Germany. The rest of the alloys and one more 25 at. pct Sn alloy were prepared using 99.9 pct Mn supplied by Gallard Schlesinger Chemical Mfg. Corp., U.S.A., and 99.999 pct Sn supplied by Semi Elements Inc., U.S.A. Weighed amounts of manganese and tin were melted in recrystallized alumina crucibles in an inert gas (argon) high-frequency induction melting furnace. By careful control of temperature and time of melting the losses were reduced to below 0.2 pct in all cases. Since the losses were very small, no attempt was made here to analyze the samples chemically. Alloys were wrapped in molybdenum foil and sealed in small evacuated fused silica capsules. The alloys were annealed at different temperatures, controlled within + l°C, for sufficiently long periods to attain proper phase equilibrium, and subsequently quenched in cold tap water. The annealing periods used at different temperatures were 15 days at 500°C, 7 days at 600°C, 5 days at 700" and 750°C, 3 days at 800°, 850°, and 900°C, 2 days at 968"C, and two alloys annealed at 968°C were reannealed for 10 hr at 1000° C. From each annealed specimen, a part was utilized for metallographic study while another piece was used for X-ray diffraction study. 1.0 pct HNO, solution and oxalic acid solutions of concentrations 0.05 to 1.0 pct were used for etching Mn-Sn alloys above and below the MnsSn composition, respectively. Since all alloys were brittle, X-ray specimens were prepared using the as-crushed -325 mesh alloy powders. Only one 25 at. pct Sn alloy powder was reannealed in an evacuated silica capsule at 800°C for 5 min and water-quenched. X-ray diffraction patterns . for the Mn3Sn phase with the as-crushed and the reannealed powders did not show appreciable change. Un-filtered iron radiation at 25 kv, 15 ma was used with either Norelco 114.6-mm-diam Debye Scherrer Camera (for phase identification) or Norelco 12-cm-diam symmetrical focusing camera (for lattice parameter determination of the 6 Mn phase). The estimated accuracy of lattice parameter determination for the focusing camera was * 0.001A. RESULTS AND DISCUSSIONS The results of metallographic and X-ray diffraction study made with different alloys are shown in Fig. 1. The variations in lattice parameter with composition for the p Mn and the Mn3Sn phases are given in Tables I and 11, respectively, and the lattice parameter as a function of composition for the p Mn phase is shown in Fig. 2. The lattice parameter of the p Mn phase increases with increasing tin content while for the Mn,Sn phase the data obtained from two two-phase alloys and one single-phase alloy indicate increase in a,, and decrease in c, parameters with increasing tin content. The results, Fig. 1, indicate that the solubility of

Jan 1, 1969

By R. D. Nelson, S. D. Dahlgren

The conditions of temperature, strain rate, and total strain favoring deformation by grain boundary sliding, slip, or deformation with concurrent recrystallization were evaluated for alpha plutonium. Grain boundary sliding and slip were studied by examining structures formed on polished surfaces with 4.5 pct compressive deformation. The conditions required for recrystallization were determined from compressive deformation us time curves. At 110" C and at a strain rate of 10&apos;per min, deformation was almost exclusively by grain boundary sliding whereas slip was the predominant deformation mode at -10°C and lo-&apos; per min. Defarmation at intermediate tenzperatures and strain rates produced structures showing mixtures of grain boundary sliding and slip. Recrystallization occurred concurrently with deformation only after a critical strain was reached. About 14, 6, and 3 pct strain was required at 105", llO°, and lZO°C, respectively, before recrystallization started, irrespective of the strain rate in the range of lo-&apos;1 to 10&apos;4 per nzin. L HE temperature and strain-rate dependency of the modes of plastic deformation were evaluated for high-purity as-cast alpha plutonium. The techniques used to study slip and grain boundary sliding were similar to those employed previously for alpha plutonium by Bronisz and Gorum,&apos; and spriet.&apos; Recrystallization was investigated using the methods reported by Nelson.3 Bronisz and Gorum&apos; found that slip occurred on more than one slip system at room temperature, and suggested that grain boundary sliding also contributed to the deformation. spriet2 found that deformation was predominantly by slip at room temperature, and observed one to three orientations of slip traces in individual grains. In addition, he deformed polished samples at 10O° C, but concluded that the deformation character of alpha plutonium at 100° C and at room temperature were not essentially different. Several investigators reported that twinning was only occasionally observed.1&apos;2&apos;4 Nelson3 recently found in creep tests that high-purity alpha plutonium would recrystallize concurrently with compressive deformation at temperatures between 25" and 115°C. EXPERIMENTAL PROCEDURE Electrorefined plutonium having less than 300 ppm total impurities was received from Los Alamos Scientific Laboratory in the form of +-in. diarn cast rods. Major impurities were americium, <I00 ppm, and tungsten, (60 ppm. Less than 25 ppm each of other impurities were present. The rods were cut into right half-cylinders 0.35 in. long and 0.25 in. in diam, and the flat faces along the cylinder axes were metal-lographically polished. To avoid loss of the metallo-graphic polish by oxidation, the polished samples were deformed in a silicone fluid heat treating medium contained in a chamber evacuated to a pressure of lo-&apos; torr. Deadweight loading was used to compress the samples between 4.5 and 20 pct at strain rates between 10"4 and lo- &apos; min-&apos;. Strain rates were determined by dividing the total deformation by the elapsed time of the test, even though the strain rate was not constant during deformation. Curves of strain vs time for samples deformed with a constant compressive stress3 show that the strain rate changes by less than a factor of four during the testing of a given sample, which is small compared to the three orders of magnitude change in strain rate investigated. Two samples were loaded into an ordinary metallographic mounting mold and impacted by striking the mold piston with one blow of a 2-lb hammer. Deformation temperatures for the two impact samples were 25" and 90°C. The samples, following deformation, were washed in carbon tetrachloride and their polished and deformed surfaces were immediately inspected metallographi-cally. Subsequently, the deformation structures were replicated with acetate replicating tape for later electron microscope examination, and the samples were then repolished and etched to reveal the microstruc-tures of the deformed metal. Second-stage carbon replicas were prepared from decontaminated acetate replicas using the procedure developed by Miller, Bierlein, and astel.&apos; Additional samples, which were 0.35 in. long by 0.25 in. in diam, were used to obtain constant load deformation vs time curves. These samples were deformed 10 to 20 pct with loads of 15 and 20,000 psi at 120°C; 30, 45, and 55,000 psi at 110°C; and 20 and 30,000 psi at 105°C. Samples were polished and etched to determine whether or not recrystallization or twinning had occurred during deformation. RESULTS AND DISCUSSION The experiments showed that alpha plutonium deforms by grain boundary sliding, slip, and deformation with concurrent recrystallization. Deformation by twinning was insignificant. Metallographic examination of polished and deformed surfaces revealed grain boundary sliding and slip. The start of recrystallization was ascertained from constant load deformation vs time curves. Deformed surface structures of samples compressed 4.5 pct under four different conditions of temperature and strain rate, Fig. 1, show how the deformation mode changes from predominantly grain boundary sliding to predominantly slip with increasing strain rate and decreasing temperature. Deformation at 110°C and 10" 4 min-&apos;, Fig. l(a), was predominantly by grain boundary sliding. Only a few slip traces can be seen in Fig. l(a).

Jan 1, 1969

By T. B. Massalski, H. W. King

An hcp phase may be induced by cold working the ß&apos; phase of the Ag-Zn system. This phase reverts to ß&apos; on subsequent aging. No phase change occurs on cold working the o phase, but ß&apos; is formed when the deformed alloy is subsequently aged at room temperature. It is concluded that for alloys near 50 at pct Zn the ordered bcc ß&apos; phase is the equilibrium structure at room temperature. WhEN the disordered bcc ß phase of the Ag-Zn system is cooled to temperatures below 258o to 274oC, it transforms to a complex hexagonal phase <o.1,2 The nature of the o ß=o transformation has been the subject of some discussion,2&apos;3 and the structure of o has been described in detail.&apos; The latter phase appears to be stable on aging at room temperature but decomposes following cold work. When alloys containing approximately 50 at. pct Zn are rapidly quenched from the 0 phase field, the ß ? o transformation may be suppressed; but the ß phase undergoes an ordering reaction (ß ? ß&apos;). The ß&apos; structure may also be obtained as a result of cold working and aging at room temperature.4 Kitchingman, Hall, and Buckley4 have suggested that the decomposition of (o following cold work proceeds in two stages, (o ? ß followed by ß ? ß&apos;, but did not confirm this by experiment. When the ordered &apos; phases in the systems Cu-Zn5 and Ag-Cd6 are cold worked, they become unstable and transform to a close-packed hexagonal phase (( ) indicating that when order is destroyed in a ß&apos; structure the close-packed hexagonal phase may in many cases be more stable. It thus became of interest to study more closely the effect of cold work and annealing on the stability of both the ß&apos; and o phases in a Ag-50 at. pct Zn alloy. Predetermined weights of spectroscopically-pure Ag and Zn, supplied by Johnson and Matthey, were melted and cast under 1/2 atm of He in transparent vycor tubing. The ingot was homogenized for 1 week at 630°C and quenched into iced brine. Since mechanical polishing was found to induce a phase change, sections were first polished at room temperature, sealed in tubes under 1/2 atm of He, reannealed for several days at 630o or 200°C and then quenched into iced brine. Sections of the alloy thus prepared were found to be homogeneous when examined under the microscope. The sample quenched from 630°C (ß -phase region) was pink in color, whereas the sample quenched from 200°C (o-phase region) was silver. The latter sample showed the characteristic hexagonal anisotropy when examined under polarized light. Filings of the alloy were examined at room temperature, after various heat treatments, using an RCA-Siemens Crystalloflex IV diffractometer with filtered CuKa radiation. The X-ray reflections from flat powder specimens quenched from 630o and 200°C and sieved through 230 mesh were recorded graphically at a scanning speed of 1/2 deg per min. The resultant patterns are shown in Figs. 1(a) and 1(b) and may be identified as those of the 8&apos; and <02 structures respectively. The lattice parameter of the ß&apos; phase was determined as 3.1566Å.* This value compares very well withthatto be expected for a 50 at. pct Zn alloy from the data of Owen and Edmunds? and indicates that no loss of Zn occurred during casting. In order to study the effect of cold work upon the ß&apos; and o phases, filings made at room temperature and sieved through 230 mesh were mounted immediately in the diffractometer-i.e., without a strain-relief anneal. Changes in structure on subsequent aging were followed by scanning repeatedly over the regions of the low index reflections of the ß&apos; and o structures-i.e. , 28 from 35 to 44 deg. Immediately after filing the 8&apos; specimen, additional diffraction peaks were observed in the low-index region of the pattern, as shown in Fig. 1(c). These additional peaks do not coincide with those of the o structure, Fig. l(b), but may be indexed as the (10.0), (00.2), and (10.1) reflections of an hcp phase (<) with nearly ideal axial ratio. However, this hexagonal phase appears to be very unstable since within a very short time at room temperature it reverts back to the ordered ß&apos; phase, the reversion being complete within seven hours. The 5 ? ß&apos; reversion reaction is, therefore, very similar to those already reported in Cu-Zn5 and Ag-Cd6 7&apos;alloys. The action of filing caused the deformed surface of the originally pink ingot to become silver in color, indi-cating that the ( phase possesses similar reflecting properties to the o phase. Hence, the subsequent

Jan 1, 1962

By Vard H. Johnson

IN 1943 the U. S. Bureau of Mines undertook drilling in an effort to develop new reserves of coking coal in an area near Paonia, Colo., as a part of an attempt to alleviate the shortage of known coking coal of good quality in the western United States. Geologic mapping of the area was undertaken by the U. S. Geological Survey with the purpose of first furnishing guidance in location of drillholes and later aiding in interpreting the results of the drilling. The drilling program was under the general supervision of A. L. Toenges of the U. S. Bureau of Mines. J. J. Dowd and R. G. Travis were in charge of-the work in the field. Geologic mapping was started by D. A. Andrews of the Geological Survey in the summer of 1943 and was continued from the spring of 1944 to 1949 by the writer. The first few holes drilled failed to locate coking coal, but in the summer of 1944 coking coal was discovered by drilling 6 miles east of Somerset, Colo., the site of present mining. In the succeeding years, 1945 to 1948, 100 to 150 million tons of coal suitable for coking were blocked out by drilling. The ensuing discussion of the geologic controls on the distribution of coking coal in the area is based on the geologic mapping as well as the drilling done in the Paonia area, more complete descriptions of which have appeared or are in process of publication.1-5 In order that the possible geologic controls affecting the present distribution of coking coal may be considered, it is necessary to discuss briefly the indicators. of coking quality coals. Coking Coal Coal that cokes has the property of softening to form a pastelike mass at high temperatures under reducing conditions in the coke oven. This softening is accompanied by the release of the volatile constituents as bubbles of gas. After release of the contained gases and upon cooling, a hard gray coherent but spongelike mass remains that is referred to as coke. This substance varies greatly in physical properties and, to be suitable for industrial use, must be sufficiently dense and strong to withstand the crushing pressure of heavy furnace loads. Western coals have a generally high volatile content and therefore form a satisfactory coke only when they attain a rather high fluidity during the process of heating and distillation in-the coke oven. When this high degree of fluidity is developed, the volatile constituents escape and leave a finely porous coke. On the other hand, when the degree of fluidity is low the product is an excessively porous and therefore physically weak mass that is called char.6 Small quantities of oxygen present in coal are believed to decrease the fluidity of the material during the coking process and to favor the development of char rather than coke. In consequence, coal chemists have for some time considered the possibility of developing an index to coking. qualities by inspection of chemical analyses of coals.7 A formula has now been developed that does permit a rough preliminary estimate of the cokability of coal on the basis of the analysis on an ash and moisture-free basis. Coals may be eliminated as possible coking fuels if the oxygen content is greater than 11 pct. Similarly the ratio of hydrogen to oxygen must be greater than 0.5 and the ratio of fixed carbon to volatile constituents must be greater than 1.3. If the coal, on the basis of these limiting factors, appears to have possible coking qualities, the following formula permits determination of the coking index: Coking index =[ a+b+c+d 5] a equals 22/oxygen content on ash and moisture- free basis, . b equals two times the hydrogen content divided by oxygen content on moisture and ash-free basis, c equals fixed carbon/1.3 x volatile matter, and d equals the heating value on moist, ash-free basis/13,600. Coking indices higher than 1.0 suggest that the coal will coke, and indices above 1.1 indicate good coking tendencies. Although generally usable, this formula is not completely satisfactory because the percentage of oxygen shown in ultimate analyses is derived only by difference; i.e., by subtracting the sum of the percentages of the constituents determined analytically from 100 pct.8,9 Although the coking index indicates the coking tendencies of coal, it is necessary to make physical tests of coke before its industrial value can be determined. The U. S. Bureau of Mines has developed a standard procedure for determining the approximate strength of coke that would be formed from a given coal. In this test one part of ground coal, mixed with 15 parts of carborundum, is baked to form a standard briquette. The weight, in kilograms, necessary to crush the briquette is termed the agglutinating index. This test determines the relative fluidity attained in the coking process by measuring the cementing strength of the coal in the briquette. A

Jan 1, 1952

By Joseph Fry, Ralph H. Espach

In the spring of 1943, when it was evident that the Tensleep bandstone in the Elk Basin Field, Wyoming and Montana, held a large reserve of petroleum, Bureau of Mines engineers obtained samples of oil from the bottom of nine wells and analyzed them for such physical characteristics as the volumes. of gas in solution. saturation pressures or bubble points, shrinkage in volume caused by the release of gas from solution, expansion of the oil with decrease in pressure, and other related properties. The composition of the gas in solution in the oil was studied. The pressures and temperatures existing in the reservoir and the productivity characteristics of the oil wells were determined. The data obtained indicate that the oil in the Tensleep Reservoir of the Elk Basin Field has unusually varying physiral characteristics, such as a saturation pressure of 1,250 psia and 490 cu ft of gas in solultion in a barrel of oil at the crest of the structure and a saturation pressure of 530 psia and 134 cu ft of gas in solution in a barrel of oil low on the flanks. The hydrogen sulfide content of the gas in solution in the oil varies from 18 per cent for oil on the crest to 5 per cent for oil low on the flanks of the structure. Of even greater significance is the fact that these and other variable characteristics of the reservoir oil are related to the position of the oil in the structure. Many geologists and petroleum engineers have considered that all the oil in a petroleum reservoir has rather uniform physical characteristics and that equilibrium conditions prevailed in all underground accumulations of oil and gas; that such is not always so is borne out by the results of the study by the writers. INTRODUCTION The Rocky Mountain region is one in which may be found striking examples of the unusual in oil and gas accumulations, as is evident from the following: The high helium content (7.6 per cent) of the gas in the Ouray-Leadville limestone sequence in the Rattlesnake Field, New Mexico, and gases of similar helium content in other fields; 50" to 55&apos; API gravity distillate in solution in carbon dioxide gas and recoverable through retrograde condensation, in the North McCallum Field, Colorado; the occurrence of gas, oil, or both in closely related structures contrary to the usual concepts of gravimetric segregation: the accumulation of gas and/or oil in structures closely related to other structures that apparently are more favorable but do not contain oil or gas accumulations; the high hydrogen sulfide content (as high as 42 per cent) of the gas associated with oil in some fields in the Big Horn Basin, Wyoming; and the wide range of fluid chararteristics found in the Elk Basin reservoir. Elk Basin, an interesting old oil field that has been producing oil from the Frontier formation since 1915, is situated in a highly eroded basin resulting from the erosion of the crest of an anticline and some of the underlying softer shales. The field came back into national prominence during 1943 when it became known that it was the largest single reserve of new oil discovered in the United States that year. The Tensleep sandstone was found to contain oil in November. 1942, when a well drilled to a depth of 4,538 ft (44 ft into the Tensleep sandstone) flowed oil at the rate of 2,500 B/D. By the end of 1949, 137 oil-producing wells and five dry holes had been drilled, and approximately 32 million bbl of oil had been produced. Approximately 6,000 acres may be considered productive of oil in the Tensleep Reservoir, and estimates of the oil that will be produced average 200 million bbl. The Tensleep Reservoir has further interest because it ha-greater closure than any oil field in the Rocky Mountain region; the closure of the Elk Basin anticline is variously estimated at 5.000 to 10,000 ft. of which the top 2.00 ft of the structure contained oil. SUBSURFACE OIL SAMPLING Fig. 1 is a structural map of the Elk Basin Tensleep Reservoir, on which the nine wells used in this study and the numbers correvponding to the well designations hereafter referred to are shown. Wells 1. 2, 3, 4, and 8 were tested and sampled during October and November. 1943. and Wells 5, 6. 7, and 9 during June and July, 1944. An electromagnetic type sampler developed by the Bureau of Mines and described by Grandone and Cook&apos; was used in obtaining the subsurface oil samples. As the wells were tubed nearly to bottom, the sampler was run as far as possible in the tubing hut never below the top perforations. The following procedure was used in testing and sampling the wells: A well was shut in for at least three days, after

Jan 1, 1951

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually average about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment. Surface deposits contain sulphur in the same range of concentrations as the salt dome deposits, i.e., from 10 to 50 pct sulphur, associated with various gangue materials such as silica, limestone, and gypsum. The pirincipal distinction, then, does not lie in the percentage of sulphur contained in the ore, but in the geological nature of the deposit. A recent study&apos; of the world sulphur supply situation estimated 1950 sulphur production in the free world countries at 5.6 million long tons, of which the United States produced 5.2 million tons, or 93 pct of the total. While America&apos;s domestic needs alone would have been covered by national production, about 1.4 million tons were exported during the same year. Despite all the steps which are being taken to restrict use of elemental sulphur and to force the fullest possible development of alternate sulphur sources here and abroad, the deficit in elemental sulphur production will probably increase with time. As a result of intensive prospecting for oil throughout the Gulf Coast area discovery of significant new salt domes is held unlikely. With the growing scarcity of sulphur and what appears to be an inevitable rise in price, recovery from deposits not amenable to Frasch-process mining assumes greater economic importance. Untapped Reserves The most important deposits in this category are located in Sicily, where elemental sulphur occurs in Miocene limestone and gypsum formation. Sulphur content of these ores ranges from 12 to 50 pct with an estimated average of 26 pct. Although quantitative estimate of these reserves is not available it is held that they exceed 50 million tons of sulphur. Similar deposits occur also on the mainland which contribute about one-third of Italy&apos;s total current annual production of 230,000 tons, but these are known to be nearing exhaustion. Significant surface deposits of volcanic origin are located in South America, Japan and western United States, silica being characteristic gangue con-stituent. The largest of these deposits are in South America. More than 100 extend over a zone 3000 miles long, paralleling the west coast of South America. &apos;Total sulphur content of these deposits has been estimated to be as high as 100 million tons. The main islands of Japan also possess at least 40 known volcanic sulphur deposits with probable reserves of 25 to 50 million tons.&apos; Prospected reserves in western United States might amount to 2 million long tons, principal deposits being located in the northwestern part of Wyoming, southern Utah, and eastern California. Volcanic deposits of lesser importance are found around the Mediterranean, in Turkey and Greece, and in Africa, Egypt, Abyssinia, and Somaliland. Beneficiation Methods Different methods of beneficiation have been used in these various locations. In Italy the Calcarone kiln and Gill regenerative furnaces are used exclusively. Both utilize heat liberated by burning part of the sulphur in the ore to liquify or vaporize the remaining sulphur, which is recovered by solidification or condensation. The Calcarone kiln is of conical shape, generally 35 ft in diam at base and 18 ft high. A kiln of 25,000 cu ft capacity burns for about two months and yields about 200 tons of sulphur. The Gill furnace consists of a series of chambers with domed roofs. Sulphur is burned and melted in one chamber at a time and the hot combustion gases are used to preheat the ore charge in the subsequent cell. These furnaces operate on a cycle of 4 to 8 days. The recovery yield of both systems is about 65 pct. Sulphur losses amount to 25 pct through the combustion to sulphur dioxide; about 10 pct is retained in discarded calcines. Ores containing less than 20 pct are not considered suitable as furnace feed. These methods are not only wasteful because of the low recovery obtained, but represent a serious atmospheric pollution problem. Sulphur produced ranges from 96 to 99 pct purity and thus does not match Texas or Louisiana sulphur. Owing to the present shortage, sulphur in the Middle East sells

Jan 1, 1954

By Joseph Fry, Ralph H. Espach

In the spring of 1943, when it was evident that the Tensleep bandstone in the Elk Basin Field, Wyoming and Montana, held a large reserve of petroleum, Bureau of Mines engineers obtained samples of oil from the bottom of nine wells and analyzed them for such physical characteristics as the volumes. of gas in solution. saturation pressures or bubble points, shrinkage in volume caused by the release of gas from solution, expansion of the oil with decrease in pressure, and other related properties. The composition of the gas in solution in the oil was studied. The pressures and temperatures existing in the reservoir and the productivity characteristics of the oil wells were determined. The data obtained indicate that the oil in the Tensleep Reservoir of the Elk Basin Field has unusually varying physiral characteristics, such as a saturation pressure of 1,250 psia and 490 cu ft of gas in solultion in a barrel of oil at the crest of the structure and a saturation pressure of 530 psia and 134 cu ft of gas in solution in a barrel of oil low on the flanks. The hydrogen sulfide content of the gas in solution in the oil varies from 18 per cent for oil on the crest to 5 per cent for oil low on the flanks of the structure. Of even greater significance is the fact that these and other variable characteristics of the reservoir oil are related to the position of the oil in the structure. Many geologists and petroleum engineers have considered that all the oil in a petroleum reservoir has rather uniform physical characteristics and that equilibrium conditions prevailed in all underground accumulations of oil and gas; that such is not always so is borne out by the results of the study by the writers. INTRODUCTION The Rocky Mountain region is one in which may be found striking examples of the unusual in oil and gas accumulations, as is evident from the following: The high helium content (7.6 per cent) of the gas in the Ouray-Leadville limestone sequence in the Rattlesnake Field, New Mexico, and gases of similar helium content in other fields; 50" to 55&apos; API gravity distillate in solution in carbon dioxide gas and recoverable through retrograde condensation, in the North McCallum Field, Colorado; the occurrence of gas, oil, or both in closely related structures contrary to the usual concepts of gravimetric segregation: the accumulation of gas and/or oil in structures closely related to other structures that apparently are more favorable but do not contain oil or gas accumulations; the high hydrogen sulfide content (as high as 42 per cent) of the gas associated with oil in some fields in the Big Horn Basin, Wyoming; and the wide range of fluid chararteristics found in the Elk Basin reservoir. Elk Basin, an interesting old oil field that has been producing oil from the Frontier formation since 1915, is situated in a highly eroded basin resulting from the erosion of the crest of an anticline and some of the underlying softer shales. The field came back into national prominence during 1943 when it became known that it was the largest single reserve of new oil discovered in the United States that year. The Tensleep sandstone was found to contain oil in November. 1942, when a well drilled to a depth of 4,538 ft (44 ft into the Tensleep sandstone) flowed oil at the rate of 2,500 B/D. By the end of 1949, 137 oil-producing wells and five dry holes had been drilled, and approximately 32 million bbl of oil had been produced. Approximately 6,000 acres may be considered productive of oil in the Tensleep Reservoir, and estimates of the oil that will be produced average 200 million bbl. The Tensleep Reservoir has further interest because it ha-greater closure than any oil field in the Rocky Mountain region; the closure of the Elk Basin anticline is variously estimated at 5.000 to 10,000 ft. of which the top 2.00 ft of the structure contained oil. SUBSURFACE OIL SAMPLING Fig. 1 is a structural map of the Elk Basin Tensleep Reservoir, on which the nine wells used in this study and the numbers correvponding to the well designations hereafter referred to are shown. Wells 1. 2, 3, 4, and 8 were tested and sampled during October and November. 1943. and Wells 5, 6. 7, and 9 during June and July, 1944. An electromagnetic type sampler developed by the Bureau of Mines and described by Grandone and Cook&apos; was used in obtaining the subsurface oil samples. As the wells were tubed nearly to bottom, the sampler was run as far as possible in the tubing hut never below the top perforations. The following procedure was used in testing and sampling the wells: A well was shut in for at least three days, after

Jan 1, 1951

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By D. C. Hilty

It has long been known that in melting high-chromium steels, some of the carbon might be oxidized out of the melt without excessive simultaneous oxidation of chromium, and that higher temperatures favor retention of chromium. The advent of oxygen injection as a tool for rapid decarburization of a steel bath permits significantly higher bath temperatures, and it was quickly recognized that the use of oxygen injection facilitated the oxidation of carbon to low levels in the presence of relatively high residual chromium contents. Up to the present time, however, specific data pertaining to the chro-mium-carbon-temperature relations in chromium steel refining have not been available. Individual steelmakers have evolved practices more or less empirically, but there has been very little real basis for predicting how effective any given practice can be in permitting maximum oxidation of carbon with minimum loss of chromium. The current investigation, therefore, was undertaken in an effort to establish the fundamental carbon-chromium relationship in molten iron under oxidizing conditions. As reported below, the equilibrium constant and the influence of temperature on that constant have been derived for the iron-chromium-carbon-oxygen reaction in the range of chromium steel compositions with what appears to be a fair degree of precision. The practical application of the result will be obvious. Experimental Procedure The laboratory investigation was carried out on chromium steel heats melted in a magnesia crucible in a 100-lb capacity induction furnace at the Union Carbide and Carbon Re- search Laboratories. The charges for the heats consisted of Armco iron, low-carbon chromium metal, and high-carbon chromium metal, the relative proportions of which were calculated so that the various heats would contain from approximately 0.06 pct carbon and 8 pct chromium to 0.40 pct carbon and 30 pct chromium at melt-down. When the charges were melted, the bath temperatures were raised to the desired level, and the heats were then decarburized by successive injections of oxygen at the slag-metal interface through a ½-in. diam silica tube at a pressure of 30 psi. The duration of the oxygen injections was from 30 sec to 2 min. at intervals of approximately 5 to 30 min. It did not appear that length or frequency of the injection periods had any significant effect on the results; cansequently, no effort was made to hold them constant and they were controlled only as was expedient to the general working of the heats. Between successive injections, the heats were sampled by means of a copper suction-tube sampler that yields a sound, rapidly-solidified sample representative of the composition of the molten metal at the temperature of sampling. This sampling device is a modification of the one described by Taylor and Chipman.1 An attempt was made to vary bath temperatures between samples, but it quickly became evident that, unless the variations were small or unless the new temperature was maintained for a minimum of 15 min. during which an injection of oxygen was made in order to accelerate the reactions, a very wide departure from equilibrium resulted. For most of the runs, therefore, temperature was maintained relatively constant at approximately 1750 or 1820°C. A few reliable observations at other temperatures, however, were obtained. Temperature Measurement The high temperatures involved in this investigation were measured by the radiation method, utilizing a Ray-O-Tube focused on the closed end of a refractory tube immersed in the metal bath. The immersion tubes employed were high-purity alumina tubes specially prepared by the Tona-wanda Laboratory of The Linde Air Products Co. These tubes were quite sturdy under reasonable mechanical stress at high temperature. They were unusually resistant to thermal shock, and chemical attack on them by the melts was slow. With care, it was found possible to keep these tubes continuously immersed in a heat for as long as 5 hr at temperatures up to 1850°C, before failure by fluxing occurred. The Ray-O-Tube—alumina tube assemblage was similar to those supplied commercially for lower temperature applications. In operation, the alumina tube was slowly immersed in the molten metal to a depth of approximately 5 in., and the device was then clamped solidly to a supporting jig where it remained for the duration of the run. A photograph of the equipment, in operation with Ray-O-Tube in place and oxygen injection in progress, is shown in Fig 1. When in position in a heat, the instrument was calibrated by means of an immersion thermocouple and an optical pyrometer. For calibration through the range of temperatures from 1500 to 1650°C, a platinum -platinum + 10 pct rhodium thermocouple in a silica tube was immersed alongside the alumina tube. Output of the Ray-O-Tube in millivolts and the

Jan 1, 1950

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By Paul R. Bluekamp

The Mather Mine property is composed of a 5.2 sq km (2 sq mile) area within the Cities of Ishpeming and Negaunee which are located in the Upper Peninsula of Michigan. Production in this mine started in 1943; it ran continuously until its closing in 1979, having produced slightly over 55 million tons of iron ore. This mine was a joint venture between several steel companies and The Cleveland- Cliffs Iron Company (CCI), with CCI being the fee holder and operator. For the first 17 years of operation, the ore was shipped in its natural state directly from the mine to the steel mill. By 1960, iron ore pellets were on the market and they proved so superior to the natural soft Mather ore that the latter became difficult to sell except for the coarser portions. It was decided to develop a pelletizing process for the Mather ore, and this was accomplished by 1965 when the first Mather pellets were produced. From this date, all but the coarse fraction of the Mather ore was shipped as pellets. The geological setting is that of a large east-west trending synclinorium which plunges to the west. The lowest member of this trough-like structure is, for the most part, a quartzite-like graywacke, the upper 20 m of which grades into a softer, fine grained slate. Lying conformably on this graywacke footwall member is an iron- formation member which is over 1,300 m in thickness. This iron-formation is cow posed of thin alternate bands of iron oxides and chert and is intruded by a number of diorite sills - some up to 122 rn (400 ft) in thickness. The north limb of this synclinorium dips at approximately 45&apos; and bottoms out at about 1,000 m from surface in the central part of the mine. There are two sets of faults, many of which are intruded by diorite dikes, which trend east-west and southeast-northwest. Displacements are varied, reaching a maximum displacement of 243 m (800 ft). The ore is found lying directly on the slatey footwall and its position is largely controlled by the faults and dikes, with the bulk of the ore being on the upper side of these structures. The ore is composed of soft earthy hematite and martite with vertical thicknesses up to 122 m (400 ft) although the average thickness would be closer to 46 m (150 ft). The ore averaged 60% iron and 7% silica on a dried analysis. The Mather Mine is located on the north limb of the syncline and was worked from two shafts, the deepest of which. was 1,09 7 m. These two shafts are about 2 km apart and serviced a total of 8 working levels between them during the life of the mine. The level spacing was about 61 m (200 ft). The main haulageways were driven parallel to the ore/footwall contact in the hard cow petent graywacke wherever possible. On the lower levels, deeper into the footwall, naked development was common. This material graded into roof bolting ground towards the upper stratigraphic portion. As drifting progressed further into the upper stratigraphic portion of the footwall, progressively stronger steel sets had to be used. From the main haulageway, cross-cuts were turned into the orebody on 61 m (200 ft) centers and extended as far as needed to recover the ore available to that particular cross-cut. The main haulageways and cross-cuts were driven 3 m (10 ft) high, 3 m (10 ft) wide at the top and 4.6 m (15 ft) wide at the bottom. This configuration would accomodate a set composed of a 2.7 m (9 ft) cap on top of two 2.7 m (9 ft) legs angled out at 18&apos;. A sill plate was used under each leg to prevent its sinking into the ore below. The type of steel used was dependent on the expected weight to be experienced. Sets were placed on 1.63 m (5 ft 4 in.) centers; however, in extremely heavy areas, it was sometimes necessary to install sets on 0.8 m (2 ft 8 in.) centers. On the top two levels (5th and 6th), the ore was considerably hard-er and was mined by sub-level stoping and long hole drilling. Very light steel sets were used in the main haulageways and cross-cuts and timber was used in the production drifts. Some con- crete production drifts were installed on 6th level, but proved to be uneconomical. However, as mining reached greater depths, the ore became softer and more massive, reaching its maximum vertical heights on 11th and 12th levels. On levels 7 through 10, yielding steel sets were used extensively in the slusher drifts. While they were satisfactory on 7th and 8th levels, their success diminished with depth and they were

Jan 1, 1981

By G. M. Schwartz, D. M. Davidson

THE Duluth gabbro outcrops containing sulphides of copper, nickel, and iron are located on both sides of State Highway No. 1 an airline distance of 8.5 miles southeast of Ely in northeastern Minnesota. The region of known sulphide occurrences includes parts of sections 5, T. 61 N., R. 11 W., and parts of sections 25, 26, 32, 33, and 34, T. 62 N., R. 11 W. These sections, given in Fig. 1, are all in Lake County, Minnesota. Part of the area, which lies entirely within the Superior National Forest, is shown on the topographic map of the Ely quadrangle. The original discovery was made in 1948 when a small pit was opened in weathered gabbro rubble for use on a forest access road. A shear zone had caused unusual decomposition in this glaciated area, and the resulting copper stain was noted by Fred S. Childers, Sr., an Ely prospector, who began searching the outcrops along the base of the intrusive. He was joined in further exploration by Roger V. Whiteside of Duluth. In the summer of 1951 a small diamond drill was moved into the area and a hole 188 ft deep was drilled, passing through 11 ft of glacial drift into sulphide-bearing gabbro. This paper is a preliminary report on the geology of the newly discovered ore. The Duluth gabbro is one of the largest known basic intrusives and may be defined as a lopolith.&apos; It extends northeastward from the city of Duluth as a great crescent-shaped mass that intersects the shore of Lake Superior again near Hovland, 130 miles to the northeast, see Fig. 2. The distance around the outside of the crescent is nearly 170 miles. The form of the intrusive is simple at Duluth where it ends abruptly north of the St. Louis River; at the east end, however, the gabbro splits into two elongated, sill-like masses separated mainly by lava flows and characterized by minor irregularities. The outcrop reaches a maximum width in the central part where it is about 30 miles across, and a maximum thickness of about 50,000 ft. It may be significant that the sulphides occur at the base of the thickest part. The lopolith has segregated into rock types ranging from peridotite to granite. The most abundant types are olivine gabbro, gabbro, troctolite, anortho-site, and granite. Of lesser importance quantitatively are peridotite, norite, pyroxenite, magnetite gabbro, and titaniferous magnetite. Grout estimates that two-thirds of the gabbro at Duluth is olivine gabbro. Variations in the percentages of plagio-clase, augite, olivine, and magnetite-ilmenite constitute the only essential differences found among the basic rock types. The predominant mineral is plagioclase, mainly labradorite. Plagioclase and olivine seem to have crystallized early, and the olivine rich rocks, usually troctolite, are found in the lower part. Segregations of titaniferous magnetite are abundant near the base of the gabbro along the eastern part and also occur far above the base. These have recently been described in detail by Grout.&apos; Near the top, segregation has produced a gradation to granite, or "red rock," as it is known locally. This consists of quartz, red feldspar, and hornblende. The red rock forms a zone with a maximum width of nearly 5 miles but is quantitatively unimportant from Duluth northward for 35 miles. In Cook county, where the gabbro splits, each of the two sill-like masses has a red rock top somewhat thicker in proportion to the gabbro below than in the main central mass. The intrusive ranges from coarse to medium in grain size and from granitoid to diabasic in texture. Throughout much of the Duluth gabbro in Minnesota banding and foliation are well developed, as Grout has emphasized.V he bands are mainly a result of variation in the percentage of minerals, as in troctolite with alternating bands high in olivine and in plagioclase. A few bands may consist largely of one mineral, as is true of some segregations of magnetite. Many of the banded rocks show a clearly developed parallelism of platy plagioclase crystals, and both banding and foliation are believed to conform to the floor of the lopolith. Throughout its extent in Minnesota the Duluth gabbro dips east and south toward Lake Superior. It is generally believed to extend beneath Lake Superior and is found as a smaller mass exposed along the north side of the Gogebic district in Wisconsin and Michigan. The dip at and near the base ranges along most of its length from 20 to 40°, but at places the internal banding dips even more steeply. The dip of the upper part is much less, and if it is assumed that the flows along the north shore of Lake Superior are a dependable indication, it does not exceed 15". The formations shown in Table I which are intruded by the gabbro range from Keewatin to Middle Keweenawan in age. They present a significant picture. At the top, the gabbro and its accompanying

Jan 1, 1953

By Rollyn P. Jacobson

THE town of Pacific, in Jefferson County, Mo., is 127 miles west of St. Louis. Since the area lies entirely on the flood plain of a cutoff meander of the Meramac River, it was considered a likely environment for accumulation of commercial quantities of sand and gravel. Excellent transportation facilities are afforded by two major railways to St. Louis, and ample water supply for washing and separation is assured by the proximity of the river. As a large washing and separation plant was planned, the property was evaluated in detail to justify the high initial expenditure. An intensive testing program using both geophysical and drilling methods was designed and carried out. The prospect was surveyed topographically and a 200-ft grid staked on which electrical resistivity depth profiles were observed at 130 points. The Wenner 4-electrode configuration and earth resistivity apparatus" were used. In all but a few cases, the electrode spacing, A, was increased in increments of 11/2 ft to a spread of 30 ft and in increments of 3 ft thereafter. Initial drilling was done with a rig designated as the California Earth Boring Machine, which uses a bucket-shaped bit and produces a hole 3 ft in diam. Because of excessive water conditions and lack of consolidation in the gravel there was considerable loss of hole with this type of equipment. A standard churn drill was employed, therefore, to penetrate to bedrock. Eighteen bucket-drill holes and eight churn-drill holes were drilled at widely scattered locations on the grill. The depth to bedrock and the configuration will not be discussed, as this parameter is not the primary concern. Thickness of overburden overlying the gravel beds or lenses became the important economic criterion of the prospect.** The wide variety and gradational character of the geologic conditions prevailing in this area are illustrated by sample sections on Fig. 2. Depth profiles at stations E-3 and J-7 are very similar in shape and numerical range, but as shown by drilling, they are measures of very different geologic sequences. At 5-7 the gravel is overlain by 15 ft of overburden, but at E-3 bedrock is overlain by about 5 ft of soil and mantle. Stations L-8 and H-18 are representative of areas where gravel lies within 10 ft of surface. In most profiles of this type it was very difficult to locate the resistivity breaks denoting the overburden-gravel interface. In a number of cases, as shown by stations M-4 and H-18, the anomaly produced by the water table or the moisture line often obscured the anomaly due to gravel or was mistaken for it. In any case, the precise determination of depth to gravel was prevented by the gradual transition from sand to sandy gravel to gravel. In spite of these difficulties, errors involved in the interpretation were not greatly out of order. However, results indicated that the prospect was very nearly marginal from an economic point of view, and to justify expenditures for plant facilities a more precise evaluation was undertaken. The most favorable sections of the property were tested with hand augers. The original grid was followed. In all, 46 hand auger holes were drilled to gravel or refusal and the results made available to the writer for further analysis and interpretation. When data for this survey was studied, it immediately became apparent that a very definite correlation existed between the numerical value of the apparent resistivity at some constant depth and the thickness of the overburden. Such a correlation is seldom regarded in interpretation in more than a very qualitative way, except in the various theoretical methods developed by Hummel, Tagg (Ref. 1, pp. 136-139), Roman (Ref. 2, pp. 6-12), Rosenzweig (Ref. 3, pp. 408-417), and Wilcox (Ref. 4, pp. 36-46). Various statistical procedures were used to place this relationship on a quantitative basis. The large amount of drilling information available made such an approach feasible. The thickness of overburden was plotted against the apparent resistivity at a constant depth less than the depth of bedrock for the 65 stations where drilling information was available. A curve of best fit was drawn through these points and the equation of the curve determined. For this relationship the curve was found to be of the form p = b D where p is the apparent resistivity, D the thickness of overburden, and b a constant. The equation is of the power type and plots as a straight line on log-log paper. The statistical validity of this equation was analyzed by computation of a parameter called Pearson&apos;s correlation coefficient for several different depths of measurements, see Ref. 5, pp. 196-241. In all but those measurements taken at relatively shallow depths, the correlation as given by this general equation was found to have a high order of validity on the basis of statistical theory.

Jan 1, 1956

By O. N. Carlson, S. A. Bradford

Discontinuous yielding in tensile tests was observed in V-O alloys in the temperature ranges of 150° to 175°C and also 350° to 400°C. The magnitude and intensity of the serrations were found to vary considerably with oxygen content. Maxima were observed in tensile and yield strengths and in the strain-hardening coefficient at the higher temperature only. The strain rate sensitivity was observed to be negative between 150° and 400°C. THIS investigation was undertaken to study the effect of oxygen on the tensile properties of iodide vanadium in the temperature range of 25o to 450°C. Brown1 observed an increase in strength between room temperature and 400°C in vanadium metal, and found that oxygen and nitrogen had a rather pronounced effect on the strength and ductility. A maximum in the tensile strength was observed by Rostoker et al.2 near 300oC and by Pugh3 around 450°C for calcium-reduced vanadium. Pugh also found a maximum in the yield strength and in the strain-hardening exponent, and minima in the elongation and strain rate sensitivity at the same temperature. Eustice and Carlson4 reported the appearance of serrations in the stress-strain curves between 140° and 180°C in iodide vanadium containing 600 ppm O. These anomalies in the mechanical properties indicate that strain aging occurs in vanadium, but the impurity or impurities responsible for the above-mentioned effects have not been identified. The phenomenon of strain aging is usually characterized by the return of the yield point after interruption of a strength test. In the temperature range where strain aging occurs, the yield and tensile strengths attain maximum values, elongation and strain rate sensitivity exhibit minima, and discontinuous yielding is generally observed in the stress-strain curve. Cottrell5, 6 has postulated that strain aging is due to the migration of solute atoms to dislocation sites to produce locking after the dislocations have broken free from their impurity atmospheres during the initial yielding. At the strain-aging temperature the process is a dynamic one in which the solute impurity atoms diffuse to the vicinity of the moving disloca- tion producing "locking" which gives rise to maxima in the tensile strength and serrations in the elongation curves. Cottrel17 has noted that discontinuous yielding in iron occurs when the diffusion coefficient of nitrogen, D, and the strain rate, i, are related by D = 10-9 €. EXPERTMENTAL PROCEDURE The vanadium metal employed in this study was prepared by the iodide refining process as described by Carlson and owen.8 A representative analysis of the vanadium used in this investigation was: 150 ppm O, <5 ppm N, <1 ppm H, 150 ppm C, 150 ppm Fe, 70 ppm Cr, <50 ppm Si, 30ppm Cu, 20 ppm Ni, <20 ppm Ca, <20 ppm Mg and <20 ppm Ti. Alloys containing from 200 to 1800 ppm O, all of which lie in the solid solution range of the V-O system, were prepared by arc melting vanadium together with portions of a high-oxygen master alloy. The master alloy was prepared by tamping pure V2O5 into holes drilled in a vanadium ingot and arc melting this five or six times in an inert gas atmosphere, inverting the button between each melting step. The oxygen content of the master alloy was then determined by vacuum fusion analysis. Vanadium containing less than 150 ppm O was prepared in the following manner. A bar of iodide vanadium was deoxidized by sealing it in a tantalum crucible with a few grams of high-purity calcium. This was held at 1100°C for 4 days to allow time for the oxygen to diffuse to the surface and to react with the calcium vapors. The calcium oxide product was later dissolved from the surface of the bar with dilute acetic acid. In this way vanadium containing from 20 to 50 ppm O was prepared. Sample Preparation. The are-melted ingots were cold swaged into 3/16-in. diam rods and these were machined into cylindrical tensile specimens with a reduced section of 1.00-in. length and 0.120-in. diam. The test specimens were annealed for 4 hr at 900°C in a dynamic vacuum of mm of Hg to remove hydrogen from the metal. This recrystal-lization treatment produced a uniformly fine-grained structure with a mean grain size of approximately 0.06-mm diam. The oxygen contents reported in this paper were determined by a vacuum fusion analysis of the tensile specimens after testing. Analyses for other interstitial or metallic impurities showed no significant changes from that of the original material. Tension Tests. Tension tests were performed on a screw-driven tensile machine at a constant cross-head speed of 0.01 in. per min. Tests at elevated temperatures were carried out by heating the

Jan 1, 1962

By F. W. Glaser

Metal-carbon-boron powder mixtures were hot pressed and the resulting specimens were studied by X-ray diffraction. It was found that regardless of the starting combination of the metal, carbon, or boron powders, a metal boride phase was always the major component in these samples. In the absence of carbon the boride phase formed on hot pressing depended only on the amount of boron present. Two new phases of the system Ti-B were found. They are Ti2B and Ti2B5. The existence of a controversial face-centered cubic phase of formula TiB was confirmed. Electrical resistivities were measured for various boride phases. It was found that the diborides are generally better conductors than the monoborides of the same metal. THE carbides and borides of the transition elements have very high melting points, in the range 2500° to 4000°C, and are therefore of interest as high temperature materials. The literature on the stability or chemical reactivity of these carbides and borides is very scarce. Various investigators&apos;-" have demonstrated a relative instability of certain carbide phases in the presence of boron or boron-containing substances. In a recent publication, Glaserl demonstrated the stability of zirconium-boride (ZrB,) in the presence of carbon at temperatures in excess of approximately 2900°C, while during a preliminary investigation of boride phases, Steinitz&apos; concluded that the diborides are stable in the presence of carbon while the monoborides of the fourth and fifth group are not, forming diborides plus carbides instead. Nelson, Willmore, and Womeldorph" have elaborated on the reaction B,C + 2TiC = 2TiB, + 3C, which was known to occur because of a relative instability of B,C and the great tendency towards TiB, formation at relatively low temperatures (approximately 1200°C). A similar study, involving as starting materials TiO, and B,C and resulting in TiB,, was recently described by Honak4, who observed the beginning of an exothermic reaction of a Ti0,-B,C powder mixture, which, when preheated in a hydrogen atmosphere to approximately 950°C, was carried to about 1600 °C by the heat of reaction. To shed more light on reactions of this type (Metal-C-B), the final product apparently always resulting in a boride phase at the expense of a carbide phase," a systematic investigation was started * Boride phases of various metals, as reported to date, are listed in Table I. and the following is an account of some of the results that were obtained. Materials, Preparation of Samples, Testing Methods The raw materials employed for this work consisted of various carbide, boride, and metal powders. as well as of boron and graphite powders. In cases where commercial grades of carbides were considered unsuitable because of low purity or excessive amounts of graphitic carbon, such carbide powders were prepared by this laboratory. The procedure for the preparation of carbide powders (zirconium carbide, titanium carbide, tantalum carbide, and niobium carbide) consisted of mixing graphite and the respective metal hydride powders in stoichio-metric proportions and subsequent heating of such mixtures in a hydrogen atmosphere in carbon crucibles. The heating was by high frequency to temperatures ranging between 1700" and 2100°C. The resulting carbide was then comminuted and screened to the desired particle size. ZrB, and TiB, powders were produced by the electrolysis of fused salt baths, according to the method described by Andriex.. The borides of niobium, vanadium, tantalum, molybdenum, chro-ium, and iron were obtained by mixing the respective metal and boron powders in the desired proportions. Such metal-boron mixtures were heated in a high frequency furnace to form boride powders. For each metal-carbon-boron group (Tables I1 through XI) a metal, its hydride, carbide or boride were mixed with carbon, boron or boron carbide powders. The additions of carbon, boron or boron carbide powders to any of these metals or metal compounds were calculated to satisfy a particular carbide or boride phase that according to the literature (Table I) had definitely been established by X-ray diffraction work. Samples of powder mixtures were hot pressed in graphite molds that were heated by direct conduction. The specimen dimensions were approximately 2.5X1X1 cm. Hot pressing temperatures were measured optically and maintained for approximately 30 sec under a constant pressure of about 1.3 ton per sq in. Wherever possible, an attempt to obtain maximum specimen density was made by temperature variation. Electrical resistivity testing was done by measuring potentiometrically the voltage drop over a length of 1.5 cm for a current of 10 amp, at room temperature. To obtain electrical resistivities for specific carbide or boride phases, values were plotted as a function of the respective sample densities

Jan 1, 1953