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By Thomas C. Wilder

ALTHOUGH there is an abundance of thermodynamic information available in the literature today, no experimental data on the standard molar properties of formation of zinc oxide have ever been reported for the temperature range 419.5" to 908°C. This is the temperature range for which zinc metal is a liquid. so that the formation reaction is: Data have been reported for Reaction [I] where zinc is a solid or a gas. Maier. Parks, and nderson' measured the electromotive force of the cell: in order to obtain the free energy of the reaction: over a limited temperature range. With the known properties of H2O, those of ZnO were calculated. These workers found that t;j of ZnO is -75,930 ± 150 cal per gfw and Mtf is-83,000 300 cal per gfw at 25°C. akolkin made similar measurements to obtain ilF7 and AH; of -75,870 and -82,540 cal per dw respectively at the same temperature. Truesdale and arirlg and Kitchener and 1gnatowicz4 measured the equilibrium of the reactions: respectively, in order to determine Ff of ZnO from the known values of CO, CO2, and H/X The thermodynamic properties of ZnO, presumably taken from the above investigations, are tabulated by Kubaschewski and vans' and Wicks and Block" The latter reference also gives values for the range where zinc is a liquid. These numbers are probably acquired from the known extremities of the range plus Cp data of the components. In the present work the electromotive force of the cell: was measured in order to calculate the properties of ZnO from the well known values of NiO for the temperature range where zinc is a liquid. In Cell [B] the cathode reaction is: and at the anode: The overall reaction is: for which the free energy change is: where is the electromotive force in volts of cell [B] after correction for the Pt-Ta thermocouple. The standard molar free energy of formation of NiO is very well known as a function of temperature. Kiukkola and agner's data are used in most tabulations. The error is probably less than I00 cal per gfw, The latter authors were the first to use solid mixed oxide electrolytes in galvanic cells. Their work with NiO and other oxides was performed with cells similar to Cell [B] using a variety of reference electrodes of fixed oxygen potential. There may be a small composition range for which either NiO, ZnO or both exist. If nonstoichiometry does exist for these oxides, then the properties described herein are those for the metal saturated compositions. The simple cell design was the same as described elsewheres except that the molten metal electrode was about 125 g of pure zinc including a small (0.25 to 0.50 g) amount of ZnO. The mixture of pure nickel and NiO powders making up the reference electrode was contained in a tube of the ZrOl + CaO electrolyte which was immersed in the molten zinc. A platinum contact was made to the reference electrode; tantalum to the zinc. A refractory cement plug was necessary at the open end of the electrolyte tube in order to prevent zinc vapor from contaminating the reference electrode. Experimental details such as the vessel containing the cell, furnace, temperature control and measurement, and electromotive force determinations of the cell have all been outlined previously. The temperature was varied randomly in the range between the freezing and boiling points of zinc. After each temperature change, time was allowed for the drift at the new temperature to be less than +;"c whereupon a series of three cell readings was taken at least 10

Jan 1, 1970

By H. L. Overton, J. B. Lipson

The concept of sodium single-ion equivalent activity as developed by Gondouin, Tixier and Simard,' was used to determine the filtrate resistivity-activity relationships for 150 laboratory and 49 field drilling muds. With the exception of the native and phosphate high-resistivity muds and the calcium-treated muds, the filtrate resistivity-activity relationships were found to be within 20 per cent of that for pure sodium-chloride solutions. Gondouin, et al,* previously suggested a method of treating the high-resistivity muds in quantitative electrical log analysis. A new treatment is presented which uses the P-alkalinity of the filtrate as a correlating parameter in handling the calcium-treated muds in quantitative electrical log analysis. INTRODUCTION Gondouin, Tixier, and Simard in 1957 introduced new interpretive procedures for analyzing the self-potential curve of electric logging. They re-emphasized the previous work of the Schlumberger brothers, Wyllie and others, and re-defined the activity concept in terms of a number which they termed the effective resistivity. Their procedures are particuarly effective when the connate-water resistivities are less than 0.08 ohm-m or greater than 0.3 ohm-m. For gyp muds where the calcium content is known, a procedure was presented which allows calculation of the connate-water resistivity with improved accuracy. These procedures assume that one of the fluids in the system is a pure sodium-chloride solution, but they are not limited by the assumptions. For the fresh and saline connate waters, for instance, it is assumed that the drilling mud filtrate is a pure sodium-chloride solution. For the gyp mud interpretive procedure, it is assumed that the connate water is a pure sodium-chloride solution. Brown,' on the other hand, has suggested that it would be extremely fortuitous if drilling mud filtrates could all be treated electrochemically as sodium-chloride solutions, and he cites some examples where discrepancies exist. A study has been made of the deviations of sodium single-ion equivalent activities, an of drilling mud filtrates from that of pure sodium-chloride solutions. Drilling fluids in common use on the Texas and Louisiana Gulf Coast and laboratory-controlled muds were investigated. Using a cell patterned after Gondouin, et al, (Fig. I), the equivalent activities were measured. It should be remembered here that the activity measured is merely an index which uses numbers that represent the sodium single-ion equivalent activity a,., of the mud filtrate. The membrane potential developed across the cell was determined with a null-type potentiometer. The resistivities of the solutions were measured by the four-electrode method. The resistivity system was calibrated by using saturated KC1 solution as a standard. Chemical properties of the drilling fluids studied were measured with standard API-approved equipment for the particular property in question. For the activity measurements, two shale discs were chosen to serve as cationic membranes. These two discs were both 1 1/4 in. in diameter and, respectively, 1/4 and 1/8 in. in thickness. The membrane potential, Em, was determined using the shale discs as cationic selective membranes separating two standard sodium-chloride solutions of known mean activity. Then by comparing the constant Km in the following equation, with the theoretical value of — 59 mv at 75OF, the perfection of the membrane was determined. The variation of Km for the two membranes used in this experimental work, as well as one of the many unsuitable membranes, is shown in Fig. 2. The measurement of the activity of mud filtrate involves the balancing of its combined individual activities against a standard solution of known mean activity. At the point where there exists zero potential across the membrane cell, the activity effects may be assumed equal for all practical purposes. The mechanics of making a null reading of this sort requires having a large number of different mean activity solutions. A more practical method involves

By Charles Stein

The effects of several different prior treatments an the creep behavior of 99.9995 pct aluminum at 260°C and 1000 psi canstant stress are compared with annealed specimens. Quenching from 538oC, irradiation with 2 mev electrons, and tension-compression room temperature fatigue damage were used to change the substructure of the specimens prior to their creep testing. The quenched and the irradiated specimens showed a larger primary and transient stage contribution to the creep curve than those specimens which were only annealed prim to creep testing. The specimen receiving prior fatigue damage at room temperature showed no first stage or transient creep when tested under identical conditions as the above specimens and had an average creep rate seven orders of magnitude lower than that of either the annealed, the quenched or the irradiated creep specimens. In a previous electron transmission microscopy investigationl of the substructure developed during the creep of pure aluminum at 0.57 Tm, it was noted that a large concentration of vacancy loops and dislocation loops were present in and near the subboundaries while the interior of the subgrains contained few of these defects and had a low dislocation density, see Fig. 1. Yim and Grant2 and Hazlett3 have shown that the presence in nickel of a substructure developed by prior cold work effectively reduces the primary and transient contributions to the creep curve. Hultgren,4 McLean,5 Gervais, Norton, and Grant,6 Chang and Grant,7 and others have shown the same effect with the higher stacking fault energy material, aluminum. However, the specific defect responsible for the change in the creep behavior developed by the prior cold work was not established. Specifically, what effectively interfered with dislocation motion in these materials at elevated temperatures? Was it jogs on dislocations produced by their interaction with other dislocations or with vacancy or dipole loops, see Fig. 1, or did the subgrain walls determine the mean free path for glissile dislocations? This would be a realistic possibility only if subgrain boundaries are effective barriers to dislocation motion. The ability of a subgrain boundary to act effectively against glissile dislocations depends on the number and the arrangement of dislocations in the subboundary,8 which in turn is a function of the creep strain.' This paper compares the creep rate of specimens possessing a large vacancy concentration to that of annealed specimens and with specimens having a stable subboundary wall containing Lomer locks produced by prior fatigue damage. EXPERIMENTAL PROCEDURE Vacancy Loops. Aluminum creep specimens, 99.9995 pct pure, were machined from zone refined rod to a 2-in. gage length and a cross-sectional area of 0.025 in. These specimens were annealed at 538oC, 1000°F, for 10 min and quenched into ice water. They were then reheated to the creep test temperature of 260°C, 500°F, and held at this temperature for 18 min to al-low for vacancy condensation and loop formation. The specimens were subsequently creep tested at 1000 psi constant stress at 260°C. The temperature along the gage length was monitored by three Pt-Pt, 13 pct Rh thermocouples, two embedded in the upper and lower shoulders of the specimen and one attached to the mid-point of the gage length- The temperature gradient was held to ± l°C or less throughout the Creep test. Elongations were measured by two LVDT's 'connected in an averaging 'On-figuration, with quartz extension arms placed against the specimen at the extremes of the 2-in. gage length. i K 3J?--' * *H - >^> - ' ^L r 77 . *y J OJL, J Fig. 1—Transmission electron micrograph of pure aluminum strained 8.2 pct at 260°C and 1000 psi constant stress. Note the presence of numerous loops near the light dislocation tangle

Jan 1, 1970

By K. E. Brown, F. W. Jessen

By utilizing an 8,000-ft experimental field well equipped with 10 gas-lift valves and 10 Maihak pressure recorders, gas-lift tests were conducted with port sizes ranging from 5/16 through I in. The well was equipped to provide accurate means of measuring surface pressures, temperatures, quantity of injection gas and fluid production. The tests were conducted in 2%-in. OD tubing, and the well was making 95 per cent water. A complete evaluation of gas-lift-valve port sizes shows the relationship of per cent recovery, gas-liquid ratios, minimum pressure created at the operating valve and horsepower requirements for each port. The length of time necessary for the fluid in the tubing to reach equilib.rium conditions after each cycle was recorded. Fallback of fluid at depths of 477, 969, 1,685, 2,493 and 4,290 ft was noted. For each port size, pressure loads of 2.50, 300, 350, 400 and 450 psi were lifted with a valve operating at approximately 550 psi at 6,000 ft. Gas-liquid ratios for each load were varied from excess gas to a gas volume per cycle whereby the load failed to reach the surface. Numerous curves are presented in evaluating the accumulated dara. The results show a 1-in. port to be the most efficient under all conditions. The production of intermittent liquid slugs against different-sized surface chokes was evaluated. These tests were conducted from a 7/16-in. ported valve at 4,072 ft. Tests indicate that, when possible, a %-in. in diameter choke or larger should be used at the surface. In the past few years most of the advancement in gas-lift operations has been made in continuous-flow operations. Yet, it is estimated that at least 70 per cent of the wells on gas lift in the United States are of the intermittent type. Since the term "slug flow" is sometimes used in both intermittent- and continuous-flow operations, it would be well to distinguish between the two types of flow. Continuous-flow gas lift is defined as a method whereby the fluids are produced at a continuous rate at the surface. This generally requires a continuous injection of gas through a surface choke; however, various other control devices sometimes are installed to eliminate freezing, to shut-off gas during natural flow periods, etc. The actual flow of fluids in the tubing may be of the slug type (one of many flow patterns known to exist in continuous flow). Intermittent flow is defined as a method of gas lift whereby the liquid is produced in separate piston-type slugs. Perhaps this type of flow could best be thought of as a ballistic-type flow where the liquid leaves bottom as a piston, propelled by a slug of expanding gas. Gas generally is injected through some type of control at the surface at predetermined intervals. However, the valve may have characteristics whereby gas can be injected through a small choke and still result in a ballistic-type flow. The purpose of the experimental work was to evaluate the most efficient port size to be used on the operating valve for the ballistic type of lift and, in addition, to establish the importance of utilizing a surface choke large enough to allow slugs to be produced without detrimental effects. This work is part of a compre- hensive study of both intermittent-and continuous-flow gas lift, representing a joint project conducted by the Ohio Oil Co., the Sun Oil Co., Otis Engineering Corp, and The U. of Texas. The problem of evaluating port sizes has been given little previous attention. Some work undoubtedly has been done which has not been published to date. Some tests were conducted when the wireline, mechanically-opened valve (Nixon) first came on the market. This valve was capable of utilizing full tubing area as its port size. It is known that this was a very efficient valve, but to the authors' knowledge the results of tests have never been published. EXPERIMENTAL EQUIPMENT These tests were conducted on an actual field well, the Ohio-Sun Unit Well No. 2-E, in the North Markham-North Bay City field, Matagorda County, Tex. The well incorporated 23/8-in. OD tubing and produced 95 per cent water. Since the running of equipment was to be quite elaborate and expensive, a well was selected in which both intermittent- and continuous-flow tests could be conducted. This particular well was capable of producing in excess of 1,000 B/D of liquid (95 per cent salt water), yet with a 3/64-in. in diameter bottom-hole choke, production was controlled to 82 B/D. Most of the intermittent tests were conducted at this low rate. Figs. 1 and 2 show all the surface and down-hole equipment. As can be seen, every attempt was made to insure that ample equipment was available for reliable testing procedures. Fig. 1 shows the surface testing equipment. The input gas was controlled first by a regulator, then

Discussion of the paper of WALTER 0. SNELLING, presented at the New York meeting, February, 1915, and printed in Bulletin No. 100, April, 1915, pp. 695 to 704. A. F. LUCAS, Washington, D. C.-Are the results of the experiments in practical operation? WALTER 0. SNELLING.-There have been no large-scale operations up to the present. The necessary conditions of temperature are easily arranged for, but the required condition of pressure, 800 lb., is not so simple a matter. The company has been working on this problem, and at the refinery in Oklahoma work of this kind is now being carried on. The way in which an ordinary still will leak at 50 lb., and more at 100 lb., leaves some doubt in the mind of many as to what can be clone at a continuous working pressure of 800 lb. On the other hand, the fact that they are making ammonia abroad by a method which employs higher pressures and temperatures than anything involved in this method, leads me to believe that the difficulties, which will undoubtedly arise when this is attempted on a commercial scale, can be overcome. WILLIAM N. BEST, New York, N. Y.-Is there any change in the calorific value of these oils?

Jan 5, 1915

By Luther Wagoner

For a number of years I have, at odd times, tried to perfect a method of assay sufficiently delicate to find and estimate minute quantities of gold and silver. The object in view was to examine rocks remote from veins or mineral areas, in order to test the probability of the lateral-secretion theory. Having succeeded in measuring approximately the amount of gold and silver contained in one cubic centimeter of sea-water, I present a detailed account of the method employed, with some of the results obtained. General Method of Assay. The use of purified lead without other flux, blowpipe methods, extraction by cyanide, and measurement of the beads by the microscope. Purged Lead. The method of testing is by cupellation. On a properly prepared cupel, it is easy to cupel and find a bead of 0.002 mm. diameter. Such a bead of silver weighs about 1 part of a milligramme. A bead 0.02 mm. in diameter weighs about 1/24,000 mg., and a close approximation can be made to its weight. It is desirable that the lead used in the assay should be low in silver so as to avoid large correction, but as the parted

Jan 1, 1902

By Wittenau, E.

OPERATION of a pilot mill of 100 tons' daily capacity during 1930 and 1931 proved that the copper minerals of the Colorado and Clay sections of the Morenci ore body could be successfully concentrated by the flotation process. Economic conditions caused postponement of further testing work until late in 1938 when the former No. 6 Concentrator at Morenci was converted into an experimental mill of 1500 tons' daily capacity which was known as the Test Concentrator. The Test Concentrator began continuous operation on April 3. 1939, using such crushing and grinding equipment of the No. 6 Concentrator as was appropriate. Modern classifiers and certain flotation machines were installed and later other types of flotation machines, a large grate-type grinding mill. mechanical screens, a short-head cone crusher, and a disk filter were added to the test equipment. Operation was uninterrupted until Dec. 31, 1941.

Jan 1, 1942

By J. W. Halley

THE effects of tin on steel have become increasingly important because of the necessity of using poorly detinned scrap, tin cans, and terne plate, in the open hearth. Since a tin can contains about 1.5 per cent tin, it would be possible to have up to 0.75 per cent tin in steel made with a 50 per cent scrap charge. In order to use tin-containing scrap, it is necessary to know the effect of tin on various grades of steel. The following investigation is by no means comprehensive but covers a number of steels in which tin content is important. PUBLISHED WORK A number of investigations of the effects of tin on rolling quality and physical properties have been published. It has been found that tin increases strength and hardness and reduces ductility and notched impact resistance. The decrease in ductility and notched impact resistance becomes more marked as the carbon decreases. McKimm1 found no change in Olsen ductility, Rockwell hardness, yield strength, tensile strength or elongation in cold-reduced strip containing up to 0.124 per cent tin. A tin content of 0.15 per cent gave difficulties in rolling and increased the hardness but 0.21 per cent tin showed no change in properties from material containing very little tin. No explanation of this contradiction is offered. Andrew and Peile2 investigated steels containing 0.10 to 0.25 per cent carbon and up to 0.63 per cent tin. They found an increase in tensile strength 0f from 100 t0 200 lb. per sq. in. for each 0.01 per cent tin on normalized samples. The elongation and reduction of area were decreased and the notched impact resistance was reduced markedly. Quenched and drawn samples did not show embrittling if water-quenched from the draw temperature of 1148°F. (620°C.). They could find no evidence of tin under the microscope. Whitely and Braithwaite3 investigated the effect of tin on rail steel containing 0.55 to 0.60 per cent carbon. They found that 0.08 per cent tin caused a marked decrease in elongation, notched impact resistance and the degree of bend. The tensile strength was increased in some tests and not in others. Bolsover and Barraclough4 investigated the effect of tin up to 0.50 per cent on a 0.35 per cent carbon steel and on several alloy steels with from 0.30 to 0.40 per cent carbon. They found a comparatively uniform decrease in notched impact resistance with increasing tin on quenched and drawn samples. If the samples were quenched from the drawing temperature or reheated to the drawing temperature and quenched, the embrittling was much less severe. Steels containing molybdenum suffered much less embrittlement than other steels. EXPERIMENTAL WORK Effect of Tin on Low-carbon Rimmed Steel Increasing quantities of tin were added to four ingots of a heat of the following ladle analysis: carbon, 0.08 per cent; manganese, 0.39; phosphorus, 0.009; sulphur, 0.025.

Jan 1, 1942

By C. M. Kleeman

All mines have a finite life and the time comes when the ore is depleted, or it has become uneconomical, or for some other reason which may be peculiar to that particular mine, it must be shut down. The operation at that point of time still has some assets such as mining and milling equipment, workshops, and in the case of a gold mine, an unknown quantity of gold which has been trapped in the treatment plant. Recovering this gold can be done at small cost and can be extremely profitable. Recovery of these hidden assets requires a good knowledge of the treatment plant, i. e. , as say values of heads, tails, and of the products produced; and if mercury for amalgamation was used, where? A careful study of available plant records can be extremely valuable in forming a working plan for adequate recovery. In the following paragraphs the various unit operations and equipment where gold can collect are described and suggestions made for its recovery. Subsequently, a case history will be given for dismantling a gold plant which had a complex flowsheet and which had been in operation for many years. Looking for trapped gold in a crushing section would invariably not be worthwhile since crushing is normally a dry process and there would be few areas where free gold could collect. In the milling section, however, many areas can literally become small gold mines in themselves. Material in the bottoms of grinding mill scoop boxes can be high-grade, and even the scale on the sides of the boxes should be scraped off and collected. Mill liners should be removed and all scale, rust, and residue collected. Residue in the bottoms of rake or spiral classifiers, if they have been used, should be removed as should the scale and rust on the sides of pump sumps, launders, etc. The following procedures should be used to extract gold from these collected materials. All steel (ball scats) should be removed, screened out, and reserved. Fine material should be placed in an amalgam barrel, water added and ground with the amalgamation tailings being sampled and assayed. If high-grade, tails should be reserved; if not, discarded. The coarse material that was screened out should be placed in a heap, regularly sprayed with water and allowed to rust. After several months this material should be screened again and the fines amalgamated.

Jan 1, 1981

By D. G. Wilder

D.G. Wilder I found the suggestion that the amount of displacement of a fault can be numerically related to the thickness of gouge or breccia to be both intuitively satisfying and intriguing. I have long agreed that there is some type of relationship between the amount of gouge and the amount of displacement of faults. I congratulate the author for developing a numerical relationship between them. However, I am concerned that the limits for applying this relationship be fully understood. An underlying assumption in this approach is that there is either a uniform thickness of gouge or breccia along a given fault or the thickness does not vary widely. Since it is not always possible to confirm this, the displacements derived by this method should be viewed with caution unless significant fault extent can be observed. At the Nevada Test Site, in drifts constructed in granite for test emplacement of spent nuclear reactor fuel, we found a fault with 0.3 to 0.4 m (12 to 16 in.) of clay gouge. Within a few meters of this location, the fault had no clay gouge, but rather consisted of a highly fractured zone with significantly altered rock and some slickensides. Based on Fig. 1, the 0.3 to 0.4 m (12 to 16 in.) thickness of gouge would indicate a displacement in excess of 30 m (98 ft). However, no gouge thickness would indicate essentially no displacement. Based on a quartz vein that terminated on the fault, and is not identified nearby, an estimated displacement of more than a few meters was made. This estimate is consistent with that obtained using the regression line proposed in the paper if the 0.3 to 0.4 m (12 to 16 in.) thickness for the gouge is used. However, using the regression curves with zero thickness would not yield results consistent with what was observed in the field. Therefore, it is important to recognize that the suggested procedure would properly yield a range of probable displacements. ? *Work performed under the auspices of the US Department of Energy by the Lawrence Livermore National Laboratory under Contract W-7405-Eng-48. Reply by E.C. Robertson It is certainly true that the t (thickness) of gg-bx (gouge and breccia) on a fault does vary along the fault. My observations have been that near the termination of a fault, the displacement d is small and the t is also small, whereas the maximum d and t will usually be found in the central part of the fault. The information on gg-bx and t of the fault found in granite in the NTS tunnel by Mr. Wilder could be interpreted somewhat differently than he does. He speaks of the fault changing within a few meters from 0.3 to 0.4 m (12 to 16 in.) of clay gg to "a highly fractured zone with significantly altered rock and some slickensides," but no gg. The highly fractured rock may be taken to be bx, rock not so finely ground as gg but still crushed by the fault movement, equivalent to the gg in my usage, and probably occupying about the same t. Mr. Wilder's estimates for the fault in the NTS tunnel for t of 0.3 to 0.4 m (12 to 16 in.) and for d of a quartz vein, in excess of "a few meters," would place the point on the low side of the central trend line in my Fig. 1, at the lower limit. There is, of course, a problem with determining d using displacement of only one planar surface. It would be greater or lesser depending on the rake of the movement. Finally, estimating the d of a fault from its t should be made with awareness of our present uncertainties, as pointed out by Mr. Wilder. Although the central trend line in my Fig. 1 has a ratio of d/t of 100, I have put the limiting ratios at 10 and 1000. Understanding of the values of the ratio will be improved only with collection of more data, for which the discussion of Mr. Wilder is much appreciated. ? G.C. Waterman E.C. Robertson's paper provides significant information to a geologist attempting to deduce fault offset by noting the products of structural dislocation. However, considerable mapping in underground and open-pit mines, and examination of structures produced in different geological settings, have convinced me that gouge and breccia thickness are controlled by geological conditions and fault movement. The following paragraphs suggest geological variables that control them. 1. Depth of Loading A near-surface fault resulting from tensional stress has more breccia/gouge than is produced by a similar stress at considerable depth. A deep-loaded compressional stress may produce a linear zone of schist, or structural dislocation may occur along an earlier formed belt of schist. Such "shear zones" are common in Canadian mines in precambrian rocks. In neither case can offset be directly deduced by an analysis of the minimal gouge/breccia in the shistose rocks. At greater depth, stress may be partially to wholly relieved by flowage. I vividly recall first noting the regional "Midas Thrust" in the Lark mine, Bingham Mining District, UT (where we called the structure the North Fault). My recorded notes, as I remember them, showed a narrow gouge streak separated the "Jordan" and "Commercial" limestone units from impure, muddy limestone beds of uncertain stratigraphic position. The visible structure did not indicate the great importance of this premineral fault

Jan 1, 1985

By Hans-Jürgen Schaller, Horst A. Brodowsky

Activity coefficients of boron in palladium were determined at concentrations up to PdB0.23 by reducing B2O3 between 870" and 1050°C in a controlled H2-H2stream and measuring the resulting weight gain. The deviations from ideal behavior closely resemble those of the system Pd-H and are interpreted in terms of three principles: 1) The solute atoms occupy octahedral interstitial positions. 2) They donate their valence electrons to the 4 d and 5s bands of palladium, raising its Fermi energy. 3) The lattice strain energy is lower for two nearesl -neighbor interstitial particles than for two farther separate ones. SOLID solutions of hydrogen in palladium are a useful subject for studying thermodynamic aspects of the formation of alloys and of nonstoichiometric systems.1-3 The activity of hydrogen is readily measurable to a high degree of accuracy,4'5 even at low temperatures where the deviations from ideal behavior are more pronounced, and its simple structure facilitates an interpretation of these deviations in terms of a detailed model. Two effects are discussed to account for the non-ideal properties:3 An "electronic" effect, connected with the rise of the Fermi energy, as electrons of the interstitial hydrogen atoms enter the electron gas of the metal, and an "elastic" effect, due to an interaction of the regions of strain around each interstitial atom. The electronic effect is based on the idea that the lowest energy levels of the dissolved hydrogen atoms are higher than the Fermi energy, so that the electron will not occupy a localized state but enter into the electron band of the metal.6 The elastic effect is based on the observation that dissolved hydrogen distorts and expands the palladium lattice. The hypothesis is put forward that the elastic strain energy is lower for two adjacent dilatational centers than for two separate ones; i.e., they attract each other. The resulting pair interaction can be used to calculate an elastic contribution to the thermodynamic excess functions by means of one of the statistical methods. This model permitted a detailed description of the solution properties of hydrogen in palladium3 and in palladium alloys.798 An extension of the approach to describe the excess functions of substitutional palladium alloys is possible.9 In order to further test and refine the model, an investigation of other interstitial alloys was started. Palladium dissolves considerable amounts of boron in homogeneous solid solution.10 The palladium lattice expands linearly up to nB = 0.23 (nB = B/Pd atomic ratio), the highest concentration studied." The expan- sion, extrapolated for 1 mole of interstitial per mole of palladium, is 17 pct of the lattice constant of pure palladium vs 5.7 pct in the case of hydrogen.12 The fact that the lattice expands rather than contracts is a strong indication that interstitial positions are occupied. According to neutron diffraction experiments, hydrogen occupies the octahedral sites of the fcc lattice.13 Unfortunately, this direct evidence is not available for the Pb-B system, mainly because of the high-reaction cross section of boron with thermal neutrons. However, by way of analogy and on the grounds of the rather close similarities between the two systems to be reported here, it seems safe to attribute octahedral positions to the dissolved boron, too. At higher boron contents, compounds of stoichiomet-ric compositions are reported such as Pd3B, which has the structure of cementite,14 so that a close structural relationship seems to exist with the system r Fe-C. In their study of hydrogen absorption in Pb-B alloys, Sieverts and Briining noted that alloys with an atomic ratio of about nB = 0.16 are no longer homogeneous15 This observation was confirmed in an extensive X-ray investigation.11,16 The phase boundaries of two miscibility gaps were established. One two-phase region was stable below a transition temperature of about 315°C and extended from nB = 0.015 to 0.178. The other one extended from nB = 0.021 to 0.114 slightly above the transition temperature and had an apex at nB = 0.065 and 410°C. All phases involved have the fcc structure of pure palladium with lattice expansions proportional to their boron contents. The occurrence of miscibility gaps, i.e., the coexistence of dilute and concentrated phases, points to an energy of attraction between the dissolved particles, in the Pb-B system as well as in the Pd-H system. The filling up of the electron bands seems to be analogous, too, in the two systems, as indicated by the hydrogen absorption capacit15,17,18 and by the suscepti bility of Pd-B alloys.l8 In both types of experiments, boron acts as an electron donor. A chemical method was used to measure the activity of boron in palladium. Boron trioxide was reduced in a moist hydrogen stream: B2O3 + 3H2 = 2B + 3H3O [l] At known activities or partial pressures of boron trioxide, hydrogen, and water, the activity of boron could be calculated from the law of mass action. The equilibrium concentration of boron corresponding to this activity was determined as the weight gain of the sample. EXPERIMENTAL The samples consisted of small pieces of foil of 0.1 mm thickness and about 100 mg weight. The palladium was supplied by DEGUSSA, Germany, and stated to be

Jan 1, 1970

Jan 1, 1925

By Raymond B. Ladoo

Abrasives include the substances, natural or artificial, that are used to grind, polish, abrade, scour, clean or otherwise remove solid material, usually by rubbing action but also by impact (sandblasting, for example). They do not include abrasive tools-for instance, lathe tools and files-or polishing agents such as waxes, which act by filling pores. Detergents and cleaners whose action is chemical rather than physical are omitted although some chemical-action polishes and cleaners may also contain solid abrasives; for example, many automobile and metal polishes. General Considerations The most important physical properties of materials that qualify them for use as abrasives are hardness, toughness (or brittleness), grain shape and size, character of fracture or cleavage, purity or uniformity. For making bonded grain abrasives such as grinding wheels, additional important factors are stability under high heat and bonding characteristics of grain surfaces. The economic factors of cost and availability are always important. No one single property is paramount for any use. For some uses extreme hardness and toughness are needed, as in diamonds for drill bits; for others, the factors of greatest importance are hardness and ability to break down slowly under use, to develop fresh cutting edges when grains become worn-for example, in garnet for sandpaper, neither highly cleavable or friable grains nor extremely tough grains are wanted. For still other uses, great hardness is objectionable; for example, abrasives for dentifrices and for glass-cleaning soaps. For the most efficient use in the more critical applications, the different types of abrasives are rarely completely interchangeable; thus, while crushed quartz and garnet are both used in sandpaper, the papers are not at all interchangeable in their use applications. In the last analysis, the choice of a high-grade abrasive depends upon the quality and quantity of work done by the abrasive per unit of cost. Initial cost of an artificial abrasive. may be much greater than that of a natural abrasive but the artificial abrasive may do so much better work than the natural one, and do it so much faster, that the ultimate cost will be less. It is on this basis that artificial abrasives have largely replaced natural abrasives for grinding metal. Classification Abrasives may be divided into two general classes, natural and manufactured. The former includes all rocks and minerals used for abrasive purposes without chemical or physical change other than crushing, shaping or bonding into suitable forms. Manufactured or artificial abrasives are made either by heat or chemical action from metals or mineral raw materials. On page 2 is a list of the most important abrasives, classified as to inherent types and as to forms in which they are used industrially. For most types of use there are manufactured products that can be substituted for the natural products, usually at higher initial

Jan 1, 1960

By Oscar A. Glaeser

THE mine and plant of the United Verde Copper Co., located in Yav-api County, at Jerome, Arizona, have been described in various technical publications and, therefore, a brief outline of its essential features will suffice. The orebodies are of the schist replacement type, the main sulfide mass being a large lens-shaped body approximately 7 acres in cross-sectional area. In general, the mineralization is found along the iron-schist contact, with the orebodies extending about 1000 ft. along this contact and varying from a few feet to 250 ft. in width. The mine is located on a mountain slope and has openings into it on three horizons, as follows: 1. Raises and ore passes open vertically into the shovel pit. These extend down to the 1000-ft. level, where they terminate as loading bins in the Hopewell adit. The main fan intake air shaft, which also extends vertically down to the 1000-ft. level, has its surface opening on the hillside at the extreme north end of the shovel pit, and the main mine air return exhausts through a raise at the extreme south end of the shovel pit. 2. The 500-ft. level adit. 3. The Hopewell adit on the 1000-ft. level. The mine has a vertical downward extent of 2000 ft. below the 1000-ft. level. Underground operations are carried on through two main shafts, both of which are wholly underground. Shaft 5 is the ore-hoisting shaft. It has three compartments and extends from the 800 to the 3150-ft. level. Shaft 6 is the service shaft, and is used for lowering and hoisting men and material. It has a large cage compartment and three smaller compartments for counterweights, ladderway, and pipe and power lines. This shaft extends from the 500 to the 2550-ft. level, and is now being extended to the 3000-ft. level. An air shaft for the downward movement of fresh air from the fan has been provided, and extends to the 2550-ft. level at this time. Work has been started to extend this air shaft to the 3000-ft. level. A return airway for vitiated mine air, in the form of raises and short connecting drifts, extends downward to the 2100-ft. level. The underground mining operations extend from the 300 to the 3000-ft. level. There are about 45 stopes in active operation on the 17 levels

Jan 1, 1929

By Joseph M. Dawson

Although there was a very great increase in drilling activity along the fault line of south Texas during 1937 as compared with the previous three or four years, and although seven new fields were discovered, it cannot be said that any major development of importance took place. There were slightly over 550 completions during 1937, of which nearly 350 were oil wells as compared with some 275 the preceding year, of which 169 were oil wells. This increase in drilling activity was occasioned almost entirely by the greater interest in and development of Austin chalk production, principally in Caldwell, Guadalupe and Bastrop Counties. Seven fields were found, of which a brief description is given in Table 3. None has as yet assumed anything of major importance, and it is doubtful whether any will be classified as major discoveries through further development. Two were producing from the Edwards limestone, one from serpentine and the other four from the Austin chalk; the latter may be classified as extremely erratic in nature. The major discouraging feature of the year was the rise and subsequent fall of interest and importance of Austin chalk production. Many new operators came into this district to drill and produce Austin chalk oil toward the first of the year, but in the latter months the production decline from this formation was so rapid and the production so erratic that drilling operations fell off materially and interest largely abated. This was particularly true of the Branyon field in Caldwell County, and the same statement is applicable to the Pearsall field, Frio County. The latter field, covering some 5000 acres and owned and controlled entirely by the Amerada and Rycade, had early completions from the Austin chalk as high as 22,000 bbl. per day. The decline was extremely rapid and the maintained production from 17 wells is only around 700 bbl. per day after a total recovery of less than one-half million barrels. Present estimates of recoverable oil from Chalk vary from 2000 to 5000 bbl. per acre. A new process for the completion of producing wells from the Edwards limestone, particularly applicable to old producing areas, was developed, in that the penetration of the Edwards lime in drilled wells is now only 1 or 2 ft. as compared to between 15 and 75 ft. in wells drilled several years ago. Many new wells have been drilled in old Edwards lime producing

Jan 1, 1938

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By J. E. Hilliard, J. W. Cahn

calculation has been made of the standard deviations to be expected in the measurement of volume fractions by areal analysis, lineal analysis and four point-counting Procedures. The effect of experimetztal errors is not included. Our conclusion is that a point count using a two-dimensional grid will be the most efficient method of volume fraction analvsis providing the grid spacing IS coarse enough. The predicted standard deviation for such an analysis is shown to he in good agreement with that determined experimentally. A metallurgist wishing to estimate a structural property by means of quantitative metallography is often faced with a choice between several different procedures. In such a case, he will naturally wish to know: a) Which procedure is the most efficient in the sense of requiring the least effort for a given precision? b) For a given procedure, what are the conditions for maximum efficiency? c) Under these conditions, how many measurements are required to attain a given precision? It is our aim in this paper to provide at least partial answers to these queries as they relate to the estimation of volume fractions from measurements on a random two-dimensional section of an opaque specimen. The commonly used techniques1,2 for volume-fraction analyses are based on one or more of the following principles: i) For an areal or Delesse3 analysis: That the areal fraction of a three-dimensional feature intercepted by a random plane provides an unbiased estimate of the volume fraction of that feature. ii) For a lineal or Rosiwal4 analysis: That the fractional intercept on a line passing at random through a two-or three-dimensional feature provides an unbiased estimate of, respectively, the areal or volume fraction of that feature. iii) For a point-count analysis: That the fractional number of randomly or regularly dispersed points falling within the boundaries of a two-dimensional feature on a plane, or within the boundaries of a three-dimensional feature in a volume, provides an unbiased estimate of, respectively, the areal or volume fraction of that feature. The absence of bias referred to in these principles does not, of course, imply that the results of an analysis will be free of error, but only that the expected result is equal to the true one. Application of the foregoing principles provide the following six possible experimental procedures: a) An Areal Analysis—This involves the measurement with a planimeter (or other means) of the area of a constitutent intercepted by the plane of polish b) A Two-Dimensional Random Point Count—Apos-sible experimental procedure for this analysis is to superimpose a sheet of transparent graph paper on a micrograph, and then use a table of random numbers to select coordinates for the points. c) A Two-Dimensiotuzl Systematic Point Count-Similar to b) except that the points are distributed in a prescribed manner, usually at the corners of a lattice superimposed on a micrograph or the screen of a projection microscope. d)A Lineal Analysis—A measurement of the fractional line length intercepted. It is usually performed by traversing the specimen under the cross hairs of a microscope and recording the distance travelled in each constituent. e) 4 One-Dimensional Random-Point Count— This could be performed by traversing the specimen with stops at random intervals to identify the constituent then present under the cross hairs. f) A One-Dimensionul Systematic Point Count— Similar to e) except that the traverse is stopped at prescribed (usually equal) intervals. It will be noted that the point-counting procedures b) and c) can be regarded as methods of estimating the areal fraction. Similarly, e) and f) are indirect methods of making a lineal analysis. All of the six procedures described above involve at least one stage of sampling. Thus, quite apart from any experimental errors, there will always be a statistical uncertainty associated with the final results. It is with the calculation of this uncertainty that we will be concerned. To avoid unnecessary repetition, the statistical relationships that will be used in the calculations are collected together in the appendix. For a fuller discussion reference should be made to one of the many textbooks5,6 on statistics. For symbols relating to the specimen structure we will follow as far as possible the terminology used by Smith and Guttman.7

Jan 1, 1962

By Roswell P. Angier, Herbert S. Kalish, Henry H. Hausner

POWDER metallurgical methods as applied to zirconium are of great interest because they permit not only the fabrication of parts directly to shape with a minimum loss of material but also the utilization of scrap, and the metal may be processed at temperatures well below the melting point. There are two difficulties involved in the melting of zirconium: 1—High vacuum is necessary for melting, and 2—zirconium reduces nearly all oxides so that it is difficult to find a suitable crucible material. The purest material, zirconium crystal bar, is expensive, and the methods of subsequent fabrication are limited. Cold rolling or extrusion of zirconium crystal bar is suitable for making certain shaped parts, but where melting is involved, the purity of the metal is invariably decreased, and a method of casting to a definite shape does not seem to be feasible. Where it is necessary to hot roll, zirconium must be covered with a steel or copper jacket to protect it from the air. This is necessarily a costly operation. Where annealing of zirconium is necessary it must be conducted in a high vacuum. In an investigation of powder metallurgical methods for zirconium, the zirconium powder used should be characterized by the following properties: 1— Highest degree of purity, 2—uniformity in particle shape and particle size distribution, 3—low compression ratio, and 4—satisfactory strength of the compacted powder (green strength). Zirconium powders have been made by two methods: By the reduction of zirconium tetrachloride (ZrC1,) with sodium,2' *•B "nd by the reduction of zirconium oxide with calcium and calcium chloride." The first method represents one of the earliest attempts to make pure zirconium metal, and the sintered product from this powder is extremely brittle. The second method of making zirconium powder is far more successful, but the powder made in this way has an extremely low apparent density and, therefore, a high compression ratio which makes it difficult to use it for powder metallurgical purposes. For these reasons the above-mentioned zirconium powders were not used in this investigation. To obtain sintered zirconium compacts of highest 'purity and good ductility, it was first necessary to produce a suitable powder. Pure zirconium has been made chiefly by two methods. The iodide method, from which the product is "crystal bar" zirconium, was developed first.'." By this purification process very soft, pure, and ductile zirconium metal can be produced. The second method, developed by the U. S. Bureau of Mines, is the reduction of zirconium chloride with magnesium." The sponge zirconium thus produced is also soft, pure, and ductile.

Jan 1, 1952

By J. Campbell

Various models are discussed for the evaluation of the negative pressures which may occur in solidifying materials which exhibit various deformation modes: elastic-plastic, Bingham, viscous, or creep flow. The inadequacy of the previously proposed elastic-plastic solution for solidifying metals is revealed by comparison with the more reliable creep results which are given graphically for aluminum, copper, nickel, and iron. The maximum tensions experienced in the liquid phase of solidifying spheres ranging in size from large castings to submicron powders are in the range from —10' to —105 atm for these metals. THERE has been much recent interest in the negative pressures associated with the volume change on solidification and in the possibility of the occurrence of cavitation. Considering the freezing of a highly supercooled liquid, an attempt to evaluate the stresses in the liquid ahead of the rapidly moving solidification front has been made by Horvay1 on a microscale and by Glicksman2 on a macroscale. In a casting of a wide freezing range alloy, the pressure differential due to viscous flow of residual liquid through the pasty zone has been discussed by Piwonka and Flemings,3 In a previous publication4 the author has attempted to estimate the negative pressure occurring in the residual liquid of a spherical casting, employing an elastic-plastic model to describe the collapse of the solidified shell under the internal tension. An earlier model assuming a rigid shell was shown to be inaccurate by many orders of magnitude. The elastic-plastic model is critically reviewed here, and other models are developed which are thought to be more closely related to metals and other materials near their melting points. The spherical geometry (Fig. 1) is chosen because the highest shrinkage pressures would be developed, although the analyses are readily adaptable to cylindrical geometry. A parallel sided casting experiences little internal tension because of the relatively easy dishing inward of the sides. (This commonly observed phenomenon has previously been attributed solely to atmospheric pressure.) Furthermore, small regions of confined liquid in a large solidified volume of a casting approximate reasonably well to spherical geometry. ELASTIC-PLASTIC MODEL The author has shown4 that as solidification proceeds the internal hydrostatic tension builds up until the elastic limit of the shell is exceeded. At this point the internal pressure is closely -2Y/3. Subsequently a plastic zone spreads from the inner surface toward the outer surface of the shell. When the whole casting is deforming plastically a rather more generalized analysis taking account of the externally applied pressure PA + 2y/b gives the internal pressure as: P = Pa + 2y/a + 2ys/b - 2 Y In(b/a) [1] The 2y/a and 2ys/b terms result from the tendency of the liquid-solid and solid-vapor interfaces to shrink, reducing their energy, and thereby helping to collapse the solid phase and compress the liquid phase. The 2y/b term would be important only for powders. The last term arises because of the plastic restraint of the solid, resisting collapse and so effectively expanding the residual liquid. From Eq. [I] it is easily shown that there is a minimum in the pressure at the radius amia= y/Y [2] which is of the order of 103K for the metals aluminum, copper, and iron, and corresponds to the minimum pressure Pmin = 2 Y[l-ln(bY/y [3] The results of a fully worked out elastic-plastic solution are given in a previous reporL4 The main criticism which may be leveled at this analysis when applied to metals at their melting points is the strong dependence of the yield stress on the strain rate. The strain rate varies with both solidification conditions (e.g., whether chill-cast or slowly cooled) and during solidification, as is indicated in the following section. Thus an appropriate choice of Y is very arbitrary. Before proceeding to a discussion of models which are strain-rate-dependent, it is necessary to evaluate the strain rate as a function of the rate of solidification. SOLIDIFICATION RATE Various empirical relations have been deduced5 for the rate of thickening of the solid shell by pour- out tests on partially solidified spheres. These, however, are unsatisfactory for our purposes since they become very inaccurate when the liquid core is very small. A theoretical approach is therefore necessary, and some solutions are set out below. Making the assumptions of constant surface temperature of the casting during freezing, no superheat, and a material freezing at a single temperature, Adams8 deduces the approximate solution: which becomes when b » a: Employing a semiempirical approach vallet6 finds:

Jan 1, 1969