Search Documents

By Brahm Prakash, R. Schuhmann

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

Jan 1, 1951

By J. B. Wagstaff

The raceway in front of the tuyere of the blast furnace has been studied quantitatively and a correlation obtained for the penetrastudiedtion of the blast. Some evidence is presented for the height and width of the raceway which suggests that all the raceways of a ffurnace overlap. The size of the coke in this zone has been measfurnaceoverlap.ured photographically during normal operation and results are given for the various areas. IN an earlier paper,' it was shown that a raceway exists opposite each tuyere of a blast furnace. This raceway is formed by the jet effect of the air emerging from the tuyere and consists essentially of a turbulence in which coke particles are recir-culated at high speed. Its presence was deduced originally from observations on movies taken with a high-speed camera through the tuyeres of various furnaces and was confirmed by experiments made on a model. In the model described,' this raceway was shown as operating in a vertical plane only, although there was a suggestion in the motion-picture film exhibited at that time that the raceway was three dimensional, unless artificially restricted. There was also some doubt then about the factors influencing its size. This paper describes the next steps in the investigation. Since the size of the raceway is obviously of importance in the operation of the furnace, it seemed worth while to study the subject more carefully. It is probably in this region that about half the coke in the furnace is consumed, so that the movement of the stock column may well be controlled by raceway behavior. Furthermore, there is some evidence to suggest that the coke is packed densely in the center of the furnace to form the "dead man" and more loosely above the raceway. It is therefore probable that the bulk of the gases passing up the stack flow from the top surface of this raceway. Clearly then, a knowledge of this critical zone is of interest to the blast furnace operator, and the first half of this report is devoted to a quantitative discussion of the subject. A further topic of interest among operators is the degree of breakdown of coke in the furnace, with which is inseparably linked the importance of a strong coke. Indeed, the whole question of the optimum size and type of coke may be as dependent on the condition of the coke in the bottom of the blast furnace as at the top. Attempts have been made from time to time to obtain samples of coke from the tuyeres and other furnace openings but they all suffer from the fact that the coke is filled to a varying degree with metal and slag and is probably broken up by the very act of taking the sample. It has proved difficult to make any reliable studies of coke size by these methods. However, it did seem possible to use the highspeed movies mentioned earlier1 to estimate the size of the coke. These movies provide an accurate record of individual coke particles so that, in theory at least, it should be possible to measure the size of the particles one by one and to obtain, for the first time, information on the coke being blown around the raceway under actual operating conditions while the furnace is performing normally. Such a study has been made and is discussed in the second half of this paper. The results obtained enabled the blast furnace data to be correlated with the model results given in the first half. Raceway Size In order to make a quantitative study of the size of the raceway it was necessary to devise some apparatus of laboratory scale, which could be handled quickly and easily. This focused attention on models, which in turn means that the laws of similarity governing this particular process must be ascertained. Method of Procedure: Since the work was to be carried out on a smaller scale than the blast furnace, the linear dimensions of the model became unimportant provided that the scale was known; it is only important to insure that the container does not affect the raceway being observed. The studies therefore were carried out in a glass-sided box, 11 in. high x 7 in. wide x 3 in. deep, using air jets ranging

Jan 1, 1954

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 R. E. Steiner, A. F. Weinberg, R. J. Van Thyne

Uranium solubility in molten bismuth was determined in the temperature range 350" to 600°C, varying from 0.09 wt pct to near 2 wl pct, respectively. Zirconium and magnesium, simulated fission products, and sodium were found to increase or decrease uranium solubility in bismuth in the temperature range 370° to 420°C, depending upon the addition or temperature. Data were obtained by both chemical analysis of samples obtained by filtration from saturated melts and by measurement of the variation of electrical resistance with temperature. IN a study for the United States Atomic Energy Commission, a Liquid Metal Fuel Reactor Experiment (LMFRE) was proposed in which a solution of uranium dissolved in molten bismuth served as both the fuel solution and the primary coolant. It was proven necessary to add other elements to this solution for certain desirable properties: 1) zirconium acts to prevent reaction of the fuel solution with the graphite core (which would result in the formation of uranium carbide) and also to inhibit mass transfer of iron and chromium from the walls of the reactor tubing to cold spots resulting in blockage; and 2) magnesium acts as a deoxidant and prevents loss of uranium. In addition to these intentional additives, other elements entering the solution are: 1) fission products which are a by-productof reactor operation, 2) sodium (the secondary coolant) through possible small leaks in the heat exchange system, and 3) iron and chromium from the container walls. Each of these elements may affect the solubility of uranium in the fuel solution. Since the concept of this reactor is based upon a liquid solution, the relatively low solubility of uranium in bismuth will dictate the minimum temperature at which the reactor heat exchanger can be operated. The solubility of uranium in liquid bismuth and the effect of some additives on this solubility have pre- viously been investigated1-&apos; (more recent results6" became avaliable after initiation of this investigation). However, some discrepancies between the various results have been noted, and also the cumulative effect of the several additions or impurities was not known. This investigation was commenced to obtain definitive results. MATERIALS Analyses of the materials used are presented in Table I. Although ostensibly of extreme purity (99.999 wt pct), the bismuth was found to contain a significant amount of impurities, probably oxides. Except for determination of the binary system, Bi-U, prior to use all bismuth was hydrogen-refined and

Jan 1, 1962

By O. Mellgren, A. M. Gaudin, P. L. De Bruyn

The adsorption density of ethyl xanthate on pyrite was determined as a function of xanthate concentration. Surface preparation of the mineral appears to have asafunctionsome effect on the subsequent adsorption process, A monolayer of xanthate on the surface is exceeded only in presence of oxygen. The effect of OH- , HS- (and x and CN- S=)and on the amount of xanthate adsorbed was investigated. Competition between OH- and X- (xanthate) ions for specific adsorption sites is indicated over a wide pH range. IN the flotation of sulfide ores, xanthates are most commonly used to prepare the surface of the mineral to be floated so that attachment to air takes place. The quantity of agent required to make the mineral hydrophobic is usually very small, of the order of 0.1 to 0.25 lb per ton of mineral. Details of the mechanism of pyrite collection are for the most part unsettled. Adsorption of collector has long been believed to involve an ion exchange mechanism as demonstrated for galena&apos; and for chalcocite.2 In the work on chal-cocite it was also demonstrated that a film of xanthate radicals unleachable in solvents that dissolve alkali xanthates, copper xanthate, or dixanthogen was formed at the surface of the mineral. The unleachable product increased with increasing addition of xanthate up to a maximum corresponding to an oriented monolayer of xanthate radicals. Pyrite is extremely floatable with xanthate if its surface is fresh.9 ut the floatability decreases rapidly as oxide coatings increase in abundance. Pyrite shows zero contact angle when in contact with ethyl xanthate solution at pH higher than about 10.5;4 at neutrality, a contact angle of 60" is obtained at a reagent concentration of 25 mg per liter. Alkali sulfides and cyanides are pyrite depressants. In this study of pyrite collection the writers have sought to relate measured xanthate adsorption to the method used in preparing pyrite, to the presence or absence of oxygen, to concentration of hydroxyl, hydrosulfide, sulfide, and cyanide ions. The principal experimental tool has been radioanalysis," " using xanthatcx marked with sulfur 35. Experimental Materials Pyrite: Unlike most sulfides, pyrite is a poly-sulfide. The structure given by Bragg7 resembles that of sodium chloride, the iron atoms corresponding to the position of sodium and pairs of sulfur atoms corresponding to the position of chlorine. The edge of the unit cell in pyrite is 5.40 A and in halite 5.63 A. The S-S distance in pyrite is 2.10 A; the Fe-S distance, 3.50 A: and the Fe-Fe distance, 3.82 A. Natural pyrite from Park City, Utah, was used in this investigation. Pyrite 1 was obtained by hand picking pure crystals. Pyrite 2 and Pyrite 3 were obtained from finer textured crystalline material containing inclusions of silicates. The same cleaning technique was utilized for the preparation of Pyrite 2 and Pyrite 3, whereas a different cleaning technique was used for Pyrite 1. Pyrite 1 was prepared as follows: The crystals were ground in a porcelain ball mill and the 200/400 mesh fraction was separated by wet screening with distilled water, followed by washing for 1 hr with deoxygenated distilled water acidified with sulfuric acid to pH 1.5. The acid was removed by rinsing with deoxygenated distilled water on a filter until a pH of 6.0 was reached in the effluent. This filtration was carried out under nitrogen. The sample was then dried in a desiccator under nitrogen. The period of time for which this pyrite sample was in contact with water containing oxygen was about 4 hr. The specific surface as determined by the BET gas adsorption method was 582 cm2 per g. Final material assayed 53.12 pct sulfur and 46.5 pct iron (theoretical, for FeS,: S, 53.45 pct; Fe, 46.55 pct). After crushing, Pyrite 2 and Pyrite 3 were washed with 1 M HCl. rinsed, and fed to a laboratory shakinq table to remove the small amount of silicates. The concentrate obtained was ground in a laboratory steel ball mill. The 200/400 mesh fraction was separated by classification in a Richards hindered settling tube. This fraction was then given a final wash with 0.1 M HCl and deoxygenated water was filtered through the sample. The final effluent showed a conductivity equivalent to that of a solution having a salt concentration of 0.3 ppm. Aqueous hydrogen sulfide solution was then added to the sampln (about 100 ml saturated H,S solution to about 1000 g pyrite under a few hundred milliliters of water) which was stored wet under nitrogen. The sample stored in this manner showed no indication of formation of iron oxides, whereas iron oxides appeared

Jan 1, 1957

By A. Simkovich, R. W. Lindsay

The possibility of using sodium sulfide slags to remove copper from ferrous alloys has been investigated by Jordan1 and by Langenberg.2, 3 In these studies, such slags were determined to be capable of removing copper and sulfur from the melt. The present work represents additional effort to clarify the effects of temperature on copper removal. The experiments were performed in a 17-lb induction furnace. Graphite crucibles contained the melts and kept the baths saturated with carbon. Temperatures were measured with a calibrated optical pyrometer and were controlled by manipulation of power input to the furnace. Estimated accuracy of temperatures in this investigation is ± 10°C (18°F) for measurements prior to slag additions, and + 20°C (36°F) after slag formation. The procedure consisted of melting 800 g of electrolytic iron. During this step, powdered graphite covered the exposed iron surface. After a predetermined temperature was reached, copper shot was added. A sample of the molten alloy for chemical analysis was then aspirated into a silica sheath. Next, a slag-forming mixture of sodium sulfite and graphite was added instantaneously to the melt. The sodium sulfite amounted to one-tenth the charge weight of iron; sufficient graphite was added to combine with oxygen in the sodium sulfite, assuming formation of carbon monoxide and reduction of the sulfite to sulfide. Subsequent to the slag addition, the molten alloy was sampled periodically, with the exception of heat A in which no intervening samples were taken between the slag addition and the end of the run. The iron was poured into a graphite mold, and the ingots sectioned and drilled for samples. Results of selected heats are presented in Table I. Analyses of samples drawn from the iron prior to slag addition are listed under zero time. Two samples from heat D were reported with copper contents greater than the initial concentration in the bath. Owing to the gradual but complete disappearance of slag during this heat, it is believed copper momentarily became more concentrated in the upper portion of the bath while reverting from the slag. This is the region from which samples were drawn. It should be noted that analysis of the ingot was equal to the copper content at the time of slag addition. The terminal temperatures of heats D and E, and the initial sulfur content of heat A are also to be noted. Because of the large temperature drop which occurred when slag was formed in heat D, power input to the furnace was increased in heat E after the slag addition, causing a higher terminal temperature. In heat A, the initial sulfur concentration was relatively high as compared to heats B through E owing to contamination by some slag remaining in the crucible from a previous heat. It is evident from Table I that copper was removed at the onset of slag formation. Roughly 30 pct of the copper was taken into the slag, with the exception of heat D, which had approximately 50 pct removed. For a comparatively short time of slag-metal contact, it appears that no gain is to be made in copper removal through use of high or low temperatures. If the slag initially formed remains in contact with the iron for an extended period, temperature has a marked effect upon copper removal, as can be seen by studying results for the two extremes in temperature. At about 1425°C, the copper level remained relatively constant after the initial removal by the slag. However, in the region of 1670°C, a definite reversion of copper occurred. Reversion was incomplete in heat D, and complete in heat E. The final temperatures of heats D and E differed by about 75°C. This temperature difference is thought to be the reason for only partial copper reversion in heat D. It is believed the effects of temperature noted above are related to the evolution of a white fume, which appeared in every run except heat A. (In the case of heat A, the fume was practically indiscernible.) After each slag addition, a yellow flame formed for about 5 sec. When the flame subsided, a white fume appeared. Upon contact with surrounding cooler surfaces, this fume deposited as a white solid. In the experiments made at 1425°C, evolution of fume continued unchanged to the end of the runs. However, heats D and E exhibited a different behavior. A very noticeable decrease in fume evolution from heat D was observed. Furthermore, this heat had much less slag remaining than did runs A through C when the experiments were terminated. No slag remained at the end of heat E; evolution of fume from this heat ceased prior to pouring. Spec-trographic analysis of the white deposit indicated sodium to be the major metallic element, with the maximum concentration of iron and copper as 0.1 and 0.01 pct, respectively. It is supposed the white fume observed in these experiments is principally sodium oxide (Na2O), formed by oxidation of sodium in the slag and subsequent sublimation. (Sodium oxide is a white to gray substance in the solid state; at 1275oC, it sublimes.4) According to this mechanism, elevated temperatures would accelerate removal of sodium from the slag, sulfur pickup by the

Jan 1, 1961

By C. H. Bowen

NDUSTRY needs a generally applicable means of defining average grain size and grain size distribution. Students of sediments hade explored this field, employing methods that might also prove useful in engineering problems. Before attempting to solve specific problems it is well to review the derivation of commonly used grade scales and the reasons for their selection. This aspect of the problem seems largely to have been lost, and a review of basic factors may suggest causes for failures in using size analysis data. Three facts are implicit in selection of a grade scale: 1) Most particulate mixtures are continuous distributions of sizes, and any grade scale that may be employed is an arbitrary means of visualizing that distribution. 2) For purely descriptive purposes, any grade scale, regardless of the rationality of the class intervals, will be satisfactory if it is accepted by a sufficient number of workers. 3) For analytical purposes, class intervals must be small enough to define the continuous distribution accurately. Further, where statistical studies are involved, a fixed relationship should exist between classes or grades. An argument in favor of geometrically related size grades lies in the fact that most particulate mixtures contain such a wide range of sizes that use of an arithmetic diameter scale is practically impossible. Udden, who recognized this fact in 1898,&apos; proposed one of the first grade scales based on a regular geometrical interval. Udden used 1 mm as his basic diameter and a ratio of 2 (or 1/2) between classes. In 1922 Wentworth2 re-examined Udden&apos;s grade scale, retaining the same class interval and basic diameter, but extending the scale in both directions and renaming the classes. In 1930 (Ref. 3, p. 82) the American Society of Testing Materials proposed what is now known as the U.S. Standard fine sieve series, also based on the 1 mm diam, with a v2 ratio between sieves. This, then, is a one fourth Udden-Wentworth series in the sizes below the 4 mesh sieve. The U.S. Standard coarse sieve departs from the 1 mm base and uses inches; hence it is not a direct continuation of the fine series. The U.S. Standard series would thus seem to possess all the attributes of a good grade scale, which it is. It has a large number of classes (sieves). in fact too many for practical use in its entirety. This v2 subdivision of the Wentworth grades has led to the common use of two v2 sieve series, the half-Wentworth and the engineers&apos; series. Geologists and sedimentologists favor the half Wentworth, or 18, 25, 35, 45 sieves, etc., whereas the engineers, preferring round numbers, utilize the other half of the U.S. Standard grade scale in the 16, 20, 30, 40 sieves, etc. The fixed geometrical ratio between classes is an advantage in statistical analysis, but the unequal classes cause some complications in calculations. This is especially true when moment measures are used. It was to simplify these calculations that Krumbein in 1934 devised the phi scale. Phi is defined as being equal to —log2 of the diameter in millimeters. Selection of logarithms to the base 2 relate the phi scale directly to the Wentworth grade scale in such a manner that the whole or fractional diameter values 2, 1, 1/2, 1/4 mm, etc., become rational whole numbers, —1, 0,1, 2, etc. Since this is an arithmetic rather than geometric series, calculations are facilitated. When the logarithm is multiplied by —1 the phi values below 1 mm become positive, those coarser than 1 mm negative. Because of its relationship to the Wentworth grade scale (and in turn to the U.S. Standard fine sieve series) it is not necessary to use the transformation equation to calculate the phi value for each individual sieve; this can be done graphically as shown in Fig. 1. It should be noted that this graph may be extended in either direction to include the range of sizes most commonly used by the individual worker. Application to Statistical Analysis Any attempt at systematically relating size analysis data to properties involves a statistical study whether it is recognized as such or not. Since this is true it would seem more logical to use measures and devices related to the general body of statistical theory. Several methods are available for studying particulate mixtures. One of the most commonly employed, and also the most often misused, is the histogram or block diagram. If its limitations are recognized and provided for, the histogram is a very useful tool. According to conventional practice, the bars of equal width are plotted and the values noted in terms of diameters, when in point of fact, log diameter is implied by such notation. Further, the histogram is sensitive to choice of grade scale and size of class interval, either of which may color the result. Grade scales whose classes are not related by fixed intervals are particularly difficult. Another basic weakness of the histogram is that it pictures a continuous distribution as a series of discrete grades.

Jan 1, 1957

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

By T. P. Chen, F. W. Bowdish

Studies made with a captive bubble apparatus on the sulfidization and collection by amyl xanthate of true chrysocolla specimens have defined the ranges of pH value and sulfide concentration which permit contact between the bubble and the mineral surface. Titanium compounds were the most effective of the materials found to activate the sulfidization of chrysocolla. With titanium activation, the contact angles and the ranges of pH value and sulfide wncentration giving bubble contact were all increased. Chrysocolla ores were concentrated by flotation. Chrysocolla ores occur at many localities in grade and quantity sufficient to make mining and millin feasible, but no satisfactory method of concentratio has been found. Although chrysocolla may be leached with acid, only those ores without acid-consuming gangue may be leached economically. Because of its potential importance, a study of the conditions nece sary for flotation of chrysocolla has been carried ou The literature contains a few references to flotation of chrysocolla. Two methods were developed by the U. S. Bureau of Mines.1,2 The first consisted of a fatty acid soap and a high xanthate as collectors of chrysocolla from a synthetic ore, while the second involved the use of hydrogen sulfide and xanthate. Ludt and DeWitt3 demonstrated the difference in adsorptive powers of chrysocolla and quartz for bas triphenyl methane dyes and suggested the use of butyl, hexyl or octyl-substituted malachite green as collector. Jackel4 emphasized the effects of combin tions of reagents such as Aerofloat 31, pine oil, and Reagents 404 and 425 with sodium sulfide and zinc hydrosulfite as conditioning agents. Although he reported recoveries of 89% from a synthetic ore and 98% from a natural ore containing azurite, malachite, chalcopyrite and chrysocolla, careful application of Jackel&apos;s method to chrysocolla from Tyrone, N.M., failed to give a high recovery. MATERIALS AND TECHNIQUE Samples from Inspiration, Ariz., and Tyrone and Magdalena, N. M., were used for experimentation and verified as true chrysocolla by leaching tests, specific gravity tests and X-ray diffraction. Chrysocolla does not dissolve at pH 4, although malachite and azurite do. Chrysocolla is about half as dense as the copper carbonates. X-ray diffraction analyses by the powder camera method confirmed the samples as true chrysocolla. A captive bubble apparatus, which cast an enlarged image of the air bubble and the mineral surface upon a screen, was used to check on the character of the surfaces. The specimens were prepared by grinding a flat surface on a glass plate using fine abrasive; then they were washed and kept in distilled water until they were to be treated with reagents. Before each reagent treatment, the specimen was carefully checked for cleanliness in the captive bubble apparatus. It was assumed that the surface was clean if, after fine grinding and washing of the specimen, the bubble would not stick. Specimens were handled with glass forceps, and precautions were taken to avoid contamination of the mineral surfaces. Contact angle measurements were carefully made several times on each treated specimen to obtain reliable average values. EFFECT OF pH VALUE AND SODIUM SULFIDE CONCENTRATION In each experiment, a specimen with a freshly ground surface was immersed for 10 min in a solution of sodium sulfide, washed and immersed for 15 min in a solution containing 30 mg per 1 of potassium amyl xanthate. The specimen was then washed again in distilled water and tested for contact angle in the captive bubble apparatus while submerged in distilled water. In this series of experiments, the pH of the sulfidizing solution was varied from 3 to 7, and the concentration of sodium sulfide, containing 60% Na2S, was varied from 50 to 650 mg per 1. Many combinations of pH value and sulfide concentration resulted in no contact between the bubble and the surface, but over a limited range of conditions, contact angles varying from 24ºto 52ºwere obtained. The data in Fig. 1 show sulfidization conditions that lead to bubble contact and those that do not. The region of contact is surprisingly small, which may indicate why flotation of chrysocolla involving sulfidization has proven so difficult in practice. Several features of the system are illustrated in Fig. 1. In the region between pH values of 4 and 6 with sodium sulfide concentrations below about 350

Jan 1, 1963

By W. B. Dancy, A. Adams

Laboratory test work on heavy liquid separation of sylvite from halite is reported. Numerous tests were run on sylvite ore sized in the ranges of 4x20 mesh, 10x65 mesh, 8x100 mesh, -8 mesh and -10 mesh with heavy liquids in the range of 2.05 to 2.15 sp gr. From the test results, it was concluded that, with the type of ore under study and a size in the range of -8 mesh, a recovery as high as 90% could be achieved with a product grade of 70% KCl. However, a final product at an acceptable recovery cannot be made with one pass, and the float must either be further processed with heavy liquids or dried and sent to a conventional froth flotation circuit. Potash ores occurring in this country consist essentially of sylvite and halite plus minor amounts of magnesium sulfate salts and montmoril-lonite-type clays. Recovery of potash minerals from evaporite ores in the North American potash fields is accomplished almost exclusively by use of amine flotation. European practice involves froth flotation as well as solution-crystallization processes. Laboratory and pilot plant test work has been reported in Europe and the U. S. on the application of heavy media separation to potash ore beneficiation. Work was probably discontinued because of lack of ore with the required very coarse liberation characteristics (1/8 to 1/2 in. liberation size). Sylvite, with a gravity of 1.99, and halite, with a gravity of 2.17, appear to be ideal for separation by heavy liquids, which are now available in gravities from 1.59 to 2.95. This paper reviews preliminary results obtained from laboratory test work on heavy liquid separation of sylvite from halite. TEST WORK The heavy liquids used in the tests under discussion were chlorobromethane, with a specific gravity of 1.923, and dibromethane, with a gravity of 2.490. These liquids, completely miscible, were combined in the proportions needed to give a mixture having the desired specific gravity. Feed for the laboratory tests was mine-run ore screened to the desired mesh sizes. In conducting the tests, the sample was fed at a constant rate into a stream of heavy liquid and the mixture directed into a small separatory vessel. The float overflowed into a collecting pan while the sink collected in the bottom of the separatory vessel and was removed at the end of the test. Approximately 500 g of feed constituted a charge. Pulp density of the feed was kept low to prevent particle to particle interference in separation. With feed in the range of 8x100 mesh, a pulp density of under 10% solids by weight was found advisable. With coarser feed the pulp density could be carried as high as 15% solids. Time of separation was very rapid. In the case of 4x20-mesh material, separation was effected in 15 to 30 sec; with -10-mesh feed, separation required about 1 to 2 min. SPECIAL EQUIPMENT Since heavy liquids are toxic to varying degrees, all separatory work was carried out in a standard laboratory fume hood. It was noted that complete removal of fumes was not being effected; therefore the hood construction was modified, resulting in a completely satisfactory arrangement for heavy liquid test work. In the interest of safety, details of this fume hood are reported here. Unlike most fumes, heavy liquid fumes tend to settle and flow like water, rather than to rise like a gas. Working on this assumption, a standard water drain was installed in the hood. Across the front of the hood a 1-in. barrier was constructed. In the rear of the hood a false back was installed, with an adjustable sliding door on both the bottom and top of this panel. As shown in Fig. 1, the exhaust fan pulled a vacuum behind the barrier, sucking the heavy fumes from the bottom of the hood. Another addition was the drying box, shown to the right of the hood. This is simply a box covered on top with hardware cloth and connected by a 6-in. inlet to the hood. Sample trays made of fine mesh wire filter screens were found ideal for drying samples. With this arrangement, air flowed completely through the sample and all fumes were drawn into the hood. In use, it was found effective to cover with a

Jan 1, 1963

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958

By P. W. Bennett, E. M. Elkin

A new technique for the recovery of palladium and platinum and sludge from go12 electrolyte eliminates many of the drawbacks of the zinc-dust cementation process. In the electrolytic refining of gold by the Wohwill method, the electrolyte contains about 100 gpl gold as AuCl, and 100 gpl free hydrochloric acid. Any palladium and platinum contained in impure gold dissolve and accumulate in the electrolyte. When their concentration approaches that of gold, it becomes desirable to discard the electrolyte and to recover the gold and the palladium and platinum. As practiced at Canadian Copper Refiners until recently, discarded electrolyte was treated with sulfur dioxide to precipitate the gold. 2AuCl3 + 6H2O + 3SO2 —2Au + 6HC1 + 3H2SO4 The gold precipitate was filtered off, washed with water, and melted with scrap anodes. The wash water was added to the gold-free filtrate and treated with zinc dust to cement the palladium and platinum. PdCl2 + Zn — Pd + ZnCl2 PtCl4 + 2Zn—Pt + 2ZnCl2 The precipitate was separated and sold as Pd-Pt concentrate, and the barren filtrate and wash water were discarded. The receipts of palladium and platinum at this company were too small to warrant their refining. The process suffered from a number of disadvantages. Handling the gold-free solution was disagreeable, as it was laden with sulfur dioxide even after blowing with air overnight. Subsequent treatment with zinc dust was also troublesome. A large excess of zinc dust was required. Spray, acid mist, and vigorous evolution of hydrogen accompanied the reaction. The Pd-Pt precipitate was grossly contaminated with occluded unreacted zinc and with yellow granules of sulfur, thought to be formed by the reduction of sulfur dioxide and its derivatives. Repeated boiling of the precipitate with hydrochloric acid and washing with water decreased somewhat, but did not eliminate, the zinc. After drying overnight, the precipitate hardened to a clinkerlike mass that was difficult to grind. Grinding, handling, and packing created heavy dusting. Even weighing in an analytical balance generated dust, indicating that the powder was easily charged with static electricity. The total palladium and platinum content of the powder varied from 56 to 72 pct, with an average of 65 pct. The powder dissolved in aqua regia with difficulty, leaving a small amount of insoluble residue. However, the main drawback of the process was the strongly hygroscopic Pd-Pt product. The powder showed a marked though erratic increase in weight and a corresponding decrease in assay on exposure to air, even during the course of laboratory analysis. Such changes were the source of disagreement between Canadian Copper Refiners and the buyers of concentrate. The cause of the change in weight is believed to be hydrogen, liberated by the action of zinc dust on hydrochloric acid. Some of this hydrogen was adsorbed so strongly by the precipitate that it could not be driven off during the drying. On exposure to air it oxidized catalytically, forming water. Experiments soon showed that the entire process of treatment of discarded electrolyte then in use would have to be changed. A small quantity of apparatus was moved into the gold room, and experimentation was held to a minimum compatible with the development of a new process. PALLADIUM AND PLATINUM PRECIPITATION It was thought that zinc could be replaced with a reagent other than a hydrogen-evolving metal. Semiquantitative experiments showed that platinum and palladium could be precipitated completely from the gold-free solutions with such well-known reagents as sodium formate, formic acid, formaldehyde, and paraformaldehyde. Sodium formate was the preferred reagent, since the other three are volatile and somewhat toxic. The gold-free solutions used were those saturated with sulfur dioxide, partly freed of sulfur dioxide by aerating overnight, completely freed by boiling, as well as strongly and weakly acidic, neutral, and alkaline solutions. The reactions are complex and can be represented by these over-all equations: PdC1, + HCOONa — Pd + NaCl + HC1 + Co2 PtCl4 + 2HCOONa — Pt + 2NaCl + 2HC1 + 2CCX, Strongly acidic solutions required as much as five times the theoretical amount of sodium formate as did solutions neutralized to pH 4.5. Increasing the pH decreased the consumption of

Jan 1, 1965

In the past few years the hydraulic aspects of rotary drilling have received considerable attention. It is generally recognized that accurate prediction of circulating pressures is desirable, particularly in those areas where a delicate pressure balance is necessary to prevent both blowouts and lost circulation. For calculation purposes the standard hydraulic equations for the pipeline flow of Newtonian fluids have been altered to facilitate similar calculations for non-Newtonian or plastic fluids.&apos;.&apos; This latter classification includes most colloidal types of drilling muds. Two properties: plastic viscosity, un, and Bingharn yield value, Y,, are commonly used to define the flow characteristics of a plastic fluid. These are commonly obtained with multispeed viscosimeters such as the Fann V-G meter.3 Normally these measurements are obtained at some surface temperature, and are not corrected to circulating temperatures for calculation purposes. This may result in considerable error, particularly in the laminar flow calculations which comnlonly apply to the drill pipe-borehole annulus. The purpose of this study was to investigate the effect of temperature on the flow properties of some water-base muds. A laboratory model Fann V-G meter was used for measurement of viscosities, yield values and gel strengths. An aluminum water jacket with O-ring seals was machined to fit around the mud cup of the Fann meter. All mud samples were prepared by thorough mixing in a Hobart mixer. Viscosity and gel strength measurements were made with the Fann meter using standard procedures. Flow properties of muds were studied at 80, 120, 160 and 180°F. The compositions of the test samples are given in Table 1. The typical effect of temperature on the flow properties of these drilling muds is illustrated by Fig. 1. It was found that plastic viscosity and apparent viscosity decreased with an increase of temperature. However, the curves were not linear, and did not appear to follow any definite trends or patterns. Yield point data showed much more scattering.&apos; Since a usable method for predicting drilling mud flow behavior with temperature was desired, an attempt was made to find a simple relationship between these variables. The basic approach used by Havenaars for predicting the same effect on turbulent viscosity resulted in Fig. 2, which shows that

By Howard Scott

Manganese being perhaps the least expensive of the metallic alloying elements that can be advantageously added to iron in considerable quantities, the basic characteristics of its alloys with iron are properly receiving much attention at present. In the course of a dilatometric investigation of this system several years ago, the writer observed certain transformation phenomena not found in other iron-base alloys. The interpretation of those effects was not evident from the information then available, but recent X-ray studies clear up at least one anomaly and permit a more satisfactory description of the results than would have been possible otherwise. Aside from the new phenomena observed, the effect of manganese on the Arg transformation is of timely interest because of its bearing on the critical compositions of austenitic alloys, the applications of which are now being greatly extended. It is known that the properties of these alloys are unstable unless the As transformation occurs at a tem-pcrature considerably below atmospheric and that manganese is very effective in depressing the temperature of this transformation. The writer aims to establish a simple general relation which expresses the location of Ar3, quantitatively in terms of the content of the common alloying elements used in steel. A step in that direction is taken here. Early dilatometric investigations1 indicated that only one transformation other than the eutectoid A1 occurs in the high-iron alloys of the iron-manganese system. Dejean represents the Arg point as falling to 0" C. with a manganese content of 11 per cent and carbon content of between 0.30 and 0.40 per cent. Esser and Oberhoffer," sing alloys low in carbon (under 0.03 per cent), arrived by extrapolation at 14 per cent for the manganese content at which As is depressed to 0" C. Not until both low carbon contents and alloys having a manganese content

Jan 1, 1931

By P. A. Laxen, H. R. Spedden, A. M. Gaudin

WHEN a mineral particle is fractured, bonds between the atoms are broken. The unsatisfied forces that appear at the newly formed surface1 are considered to be responsible for the adsorption of ions at the mineral surface. A knowledge of the mechanism and extent of ion sorption from solution onto a mineral surface is of interest in the development of the theory of flotation.2,3 Study of the adsorption of sodium from an aqueous solution on quartz offers a simple approach to this complicated problem. The availability of a radioisotope as a tracer element meant that accurate data could be obtained.4,5 Three main factors which appeared likely to affect the adsorption of sodium are: 1-concentration of sodium in the solution, 2-concentration of other cations in the solution, and 3-anions present in the solution. Hydrogen and hydroxyl ions are always present in an aqueous solution. By controlling the pH, the concentration of these two ions was kept constant. The variation in the amount of sodium adsorbed with variation in sodium concentration was then determined under conditions standardized in regard to hydrogen ion. The effect of concentration of hydrogen ions and of other cations was also measured. A few experiments were made to get a preliminary idea on the effect of anions. The active isotope of sodium was available as sodium nitrate. Standard sodium nitrate solutions were used throughout these experiments except when the effects of other anions were studied. It was found that sodium adsorption increased with sodium-ion concentration, but less rapidly than in proportion to it. Increasing hydrogen-ion concentration, or conversely decreasing hydroxylion, brings about a comparatively slight decrease in sodium-ion adsorption. Increasing the concentration of cations other than hydrogen or sodium decreases somewhat the adsorption of sodium ion. It would appear as if the kind of anion is a secondary factor in guiding the amount of sodium ion that is adsorbed. Materials and Methods Quartz The quartz was prepared as in previous work in the Robert H. Richards Mineral Engineering Laboratory4 except for the refinement of using de-ionized distilled water for the final washing of the sized quartz, prior to drying5 To minimize the laborious preparation of quartz, experiments were made to determine whether the sodium-covered quartz could be washed free of sodium and re-used. The experiments were successful as indicated by lack of Na&apos; activity on the repurified material and by its characteristic sodium adsorption. Table I gives the spectrographic analyses of the quartz used. The quartz ranged from 16 to 40 microns in size, averaging about 23 microns (microscope measurement), and had a surface of 1850 sq cm per g (lot I), 2210 (lot II) and 2000 (lot III) as determined by the Bloecher method.6 Radioactive Sodium Method of Beta Counting for Adsorbed Sodium: Na22, the radioisotope of sodium, possesses convenient properties.7 It has a half-life of 3 years, thus requiring no allowance for decay during an experiment. On decay it emits a 0.575 mev ß radiation and a 1.30 mev ? radiation. The decay scheme is illustrated in the following equation: [Y Nam S. - &apos;Net 3 years] The ß radiation is sufficiently strong to penetrate an end-window type of Geiger-Mueller counting tube. This, in turn, makes it possible to use external counting, a great advantage in technique. Furthermore, it permits the assaying of solids arranged in infinite thickness, while assaying evaporated liquors on standardized planchets. The equipment used was standard and similar to that employed by Chang8 The original active material was 1 ml of solution containing 1 millicurie of Na22 as nitrate. This active solution was diluted to 1000 ml. Five milliliters of this diluted active solution was found to give a quartz sample a sufficiently high activity for accurate evaluation of the sodium partition in the adsorption measurements. Also, 1 ml of final solution gave a sufficiently high count for precision on the liquor analyses. The sodium concentration of the diluted active solution was 1.2 mg per liter, so that 6 mg of sodium for 60 ml of test solution and 12 g of quartz was the minimum amount used. The active solution was stored in a Saftepak bottle. Procedure for Adsorption Tests: The method consisted of agitating 12 g of quartz with 60 ml of solution of known sodium concentration for enough time to establish equilibrium between the solution and the quartz surface. The quartz was separated as completely as possible from the solution by filtering and centrifuging. The activity on the quartz and in the equilibrium solution was measured and the partition of the sodium was calculated from the resulting data. The detailed procedure for the adsorption test is set forth in a thesis by Laxen5 In brief, it included the following steps: 1-Ascertainment of linearity between concentration of Na22 and activity measured. 2-Evaluation of factor to translate activity on solid of infinite thickness in terms of activity on an evaporated active film of minute thickness, on the various shelves of the counter shield. 3-Taking precautions to avoid evaporation of water during centrifuging.

Jan 1, 1952

By C. N. J. Wagner, E. N. Aqua

Dilute Fe-Si alloys (0 to 5 wt pet Si) were cold-worked by making filings at room temperatuve. The analysis of the broadening of the X-vay powdev pattern peaks yields anisotropic particle sizes normal to the (110), (100), and (112) planes which are a consequence of deformation faults (probability a) and twin faults (probability p). The combined fault probability (1.5a +ß) increases from 0.01 in pure iron to 0.02 in Fe-5 pct Si. The root mean square stvains vary with crystallographic ovientation due to elastic anisotropy of the alloys, and increase from about 0.0035 in pure iron to 0.005 in Fe-5 pct Si measured in the [100] direction. Assuming that deformation and twin faults occur with equal probability, the X-ray data would imply that the stacking-fault energy does not change upon alloying. THE crystallographic aspects of the plastic deformation of dilute Fe-Si alloys has been studied many times since the observation of glide band structures by Barrett et a1.l These glide bands were much straighter than the wavy bands observed by Gough2 in unalloyed iron nearly 40 years ago. This difference in the morphology of the slip band structure has been verified many times, most recently by replication electron microscopy3 and dislocation etch-pitting methods.4,5 These observed differences in deformation behavior have been considered to result from the degree of cross slip that occurs in pure iron and in the Si-Fe alloys. The lowered tendency for cross slip would result from a lowering of the stacking-fault energy of iron by the silicon alloying addition. Experimental evidence for this lowered stacking-fault energy in the Si-Fe alloys, obtained with transmission electron microscopy, is the observation of dissociated dislocations in Fe-6 pct si6 and Fe-3 pct Si alloys,7 and the observation of sharply defined cellular substructures in Fe-3 pct si6,7 compared to the more diffuse substructure in zone-refined iron.&apos; The activation energy for recovery of cold work is greater in the Si-Fe alloys than it is in pure iron, indicating that cross slip has become a more difficult process.9 The purpose of this investigation is to determine the change in stacking-fault probability from the broadening of powder pattern peaks in dilute silicon-in-iron solid solutions, when cold-worked at room temperature. The knowledge of the stacking-fault probability and the dislocation density, both of which can be deduced from line-broadening analysis of two orders of reflections from (110), (200), and (112) planes, will permit us to calculate the change in stacking-fault energy upon alloying silicon into iron. Previous X-ray investigation of cold-rolled Fe-5 wt pct Si (100)[001 ] single crystals did not reveal any line broadening, whereas cold-rolled Fe-3.1 wt pct Si (001) [110] and (001)[100] single crystals10 showed line bro2dening as a consequence of small particle sizes (-200A) and root mean square strains (-0.25 pct), which are comparable to those found in cold-worked iron filings.11 EXPERIMENTAL PROCEDURE A series of five Fe-Si alloys containing nominally 1 to 5 wt pct Si and pure iron were cold-worked by making filings at room temperature. The actual silicon content deviated at the most by +0.1 pct which is of no significance in the present study. The carbon contents of these materials were less than 0.02 pct. The X-ray diffraction patterns were measured with molybdenum, cobalt, and iron radiation using a GE and a Philips diffractometer. The (110)-(220), (200)-(400), and (112)-(224) pairs of reflections were

Jan 1, 1968

By J. E. Worthington, W. H. Spence, I. T. Kiff

Gold was discovered at the Haile mine in Lancaster County, South Carolina, in 1827 or 1828, and since that time the mine has been worked intermittently by both open-pit and underground methods until its forced closure in 1942 by World War II. Production figures are incomplete, especially for the early years, but the total gold produced is estimated to have been greater than 200,000 oz. Thus, the Haile mine has been the most productive gold mine in the eastern United States. The upper, residually enriched ores were relatively rich, but the bulk of the production has come from the mining of lower grade ores. General Geology The Haile mine is located in late Precambrian or early Paleozoic rocks of the Carolina slate belt at the edge of the Atlantic Coastal Plain [(Fig. 1)]. The metamorphic grade is lower greenschist facies and the rocks have been folded into a sequence of northeast-trending isoclinal folds. The gold is associated with siliceous, pyritic, and kaolinized felsic pyroclastic and tuffaceous rocks in an interbedded volcanic and volcanoclastic sequence of felsic to mafic tuffaceous rocks and argillaceous sediments [(Fig. 2)]. The ore bodies occur in two northeast trending zones approximately 500 m apart; each zone is 30-70 m wide and 600 m or more in length, with possible extensions to the east beneath the Coastal Plain sediments. Mineralogy. Gold in the Haile mine is always associated with siliceous and/or pyritic ores. The gold occurs in at least three states: As native gold as originally deposited; as residual gold derived from the breakdown of pyrite; and as gold included in pyrite. Major associated minerals in addition to quartz and pyrite are kaolinite, sericite, and iron oxides. Minor molybdenite, arsenopyrite, pyrrhotite, copper sulfides, sphalerite, rutile, and topaz are also present. Petrology. The gold-bearing ore zones vary from highly siliceous rocks to pyritic massive sulfide lenses. This variation is most easily seen today along strike from the Haile pit to the Red Hill pit. Ore grade material still exposed in the wall of the Haile pit consists of a highly siliceous and very thinly bedded rock containing minor pyrite. Along strike, the character of the mineralization changes to pyritic massive sulfide lenses occurring interbedded with siliceous horizons at the Red Hill pit. The siliceous rocks vary from the thinly-bedded material as just described from the Haile pit to silicified fragmental-appearing rocks to totally recrystallized cherty rocks lacking any recognizable primary features. Scattered, apparently at random, throughout the very thinly-bedded and very fine-grained ore face of the Haile pit are seemingly anomalous silica-rich clasts or concretions up to 5 cm in diameter which will be discussed later in this paper. Alteration. One of the most striking features of the Haile deposit is the alteration mineral assemblage which is intimately associated with the siliceous and pyritic ores. This altered material has been intersected in drill core at depths greatly exceeding the modern weathering profile and is, therefore, of hydrothermal origin rather than from supergene processes. This "sericite," actually a fine-grained mixture of sericite, kaolinite, and quartz, can be shown to stratigraphically underlie the gold- quartz-pyrite zone, and is well exposed in the open pit just southeast of the Haile and Bumalo pits. Relict textures indicate that this highly altered material was originally a felsic ash flow. Other similar alteration zones have been found in outcrop and drill core underlying the remaining ore bodies. Thus each of the mineralized zones consists of two parts: A siliceous and/or pyritic gold-bearing ore zone which is stratigraphically underlain by a zone of high alumina minerals, in this case sericite and kaolinite along with variable amounts of quartz. A green chrome mica, presumably fuchsite, is present in trace amounts in the high alumina zone. Genesis An adequate model to explain the origin and distribution of the gold deposits in the Carolina slate belt is presently lacking. Worthington and Kiff1 suggested a volcanogenic origin for certain gold deposits in the North Carolina slate belt from the waning exhalations of felsic volcanic piles. They also pointed out that such an origin has similarities to many epithermal precious metal deposits located in more recent volcanic piles in the western United States. A further key to the understanding of the genesis of the gold mineralization at the Haile mine is the close association of the mineralization in siliceous and sulfidic horizons to the genetically related and stratigraphically underlying high alumina alteration. Such high-alumina alteration is common around felsic volcanic centers in the Carolina slate belt and the mineralogy as seen today consists of some combination of kaolinite, sericite, pyrophyllite, kyanite, andalusite or sillimanite depending on the local prevailing grade of metamorphism. Accompanying the high-alumina alteration are large quantities of pyrite and iron-oxide minerals as well as characteristic minor accessory minerals often including base metal sulfides, fluorine-bearing minerals (topaz, fluorite, apatite), titanium-bearing minerals (ilmenite, rutile),

Jan 1, 1981

By J. Alfred Berger, O. M. Katz

The Zv-Hf-H ternary system was studied between 500° and 900°C at pressures less than 1 atm of hydrogen gas between 1 and 60 at. pct H. A new and unique microgravimentric apparatus was used. Cizanges of slope on pressure-hydrogen composition isothernis delineated phase boundaries. These boundaries separatecl the three regions, a, 0, and y—so designated to correspond to the regions of the Zr-H binary system—from the multiphased areas between them. A eutectoidal decomposition was found with the ß region phase or phases decornposing into a lamellar product on quenching to rool ter,zperatuve. Reproducible decomposition-pressure hysteresis occilrved lnainly at lower hydrogen cornpositions and at lower temperatures across multiplzase vegions between a and 0 and a and y. Tire effects of hqfniur7z on the hydriding charactevistics of zirconiurrz weYe as follows: 1) stabilization of the a and y vegions while destabilizing the 0 region; 2) a/?preciable elevation of the decomposition pressrkres in the multiphase region between the a and /3 field; 3) ~nouenzent of the eutectoid reaction to high te~nperatures; 4) reduction in the total qiiantity of hydrogen absorbed under one atmospheve of Hz p7-essure; and 5) introduction of a split deconzposilion at the eiitectoiclal poinl in pa?? of the ternavy. Assuru~ptions based on an ir-2terstitial vandonl-solulion rtioclel 0.f hydrogen in metals slzowed that the bindit~g energy between solute sites prednnzinatecl at low /i?!dvogen concentrations. However, at high hydrogen contents the entropy was the predorninatlt factor in determining the stability of the Zr-Hf-H al1o.s. This was interpreted to mean a scarcity of filrtlzer itltevslilinl solute sites caused by hydrogen-hydvogen intet-actions in the metal lattice. INTEREST in the reaction of hydrogen with metals has increased in the past few years for the following reasons: 1) the formation or use of high hydrogen potential environments in nuclear reactors; 2) the reaction of hydrogen with alloys in nuclear reactors with the accompanying deleterious effects on the mechanical and corrosion properties; 3) the theoretical implications of thermodynamic data on the theory and rules of alloy formation in the metal-hydrogen systems; 4) the use of hydrogen-containing fuels in rocket engines; 5) the need for a process of making fine metal powders of high-melting reactive metals; and 6) the beneficial impregnation of superconducting alloys with hydrogen. In nuclear pressurized-water reactors, the problem exists of limiting the hydrogen pickup of zirconium alloys which are utilized as fuel cladding, heat shields, and support members. In general, zirconium alloys have good mechanical and corrosion-resistant properties in high-temperature water. However, hydrogen is absorbed from the corrosion reaction between metal and water, greatly accelerating the formation of the corrosion product ZrOz as well as mechanically embrittling the underlying metal. In addition, recent observations1 at zirconium to hafnium welds showed that secondary elements in zirconium can have an appreciable, and somewhat unexpected, effect on hydrogen absorption. This paper lists the thermochemical data in the range 500" to 900°C for the equilibrium reaction of four high-purity Zr-Hf alloys with hydrogen. Phase boundaries and thermodynamic functions are determined while the structural data will be presented in a future paper. In general, the Zr-Hf-H system approximates the well-known, eutectoidal, Zr-H diagram2,3 with modifications introduced through the behavior of hafnium.4,5 The Hf-H system,&apos; published while this work was in progress, provided a consistent trend with the Zr-Hf-H data. PREPARATION OF Zr-Hf ALLOYS Table I presents a complete flow chart of the preparation procedure. The zirconium and hafnium crystal bars were completely immersed in high-purity kerosene and slowly cut into thin wafers. Wafers were then cold-sheared into approximately 1-g pieces, thoroughly cleaned, weighed, and inserted into the furnace. The alloys, B-2, B-4, B-6, and B-8, were then nonconsumable arc-melted under 500 mm of purified argon. Additional purification of the argon was accomplished by melting a large titanium button each time an alloy was re-melted or a different alloy melted. Each alloy button, which weighed 25 g, was remelted four times in an approach to complete homogeneity. Material losses were less than 0.02 wt pct. Alloy buttons were alternately cold-rolled and vacuum-annealed into 10- and 20-mil sheets. Table I1 gives the composition of the four alloys used. Very little elemental segregation existed be-

Jan 1, 1965

By Robert E. Green, Edmund G. Henneke

Expressions are derived for the orientation factors associated with dislocation damping due to dislocation motion in the basal slip plane. The results of the calculation of these factors for zinc crystals aye pvesepzted and the relevance of these results to ultvasonic-attenuation measuvements is discussed. In this paper expressions are derived for the orientation factors associated with damping of sound waves due to dislocation motion in the basal slip plane of hexagonal crystals. For zinc crystals, the orientation factors for dislocation motion in one slip direction only, as well as for simultaneous motion in all three basal slip directions, have been calculated at 1-deg intervals. The results of these calculations are presented in the form of contour diagrams in a sector of a standard polar coordinate graph. The calculations have been carried out for both the traveling- and the standing-wave conditions. The relevance of these results to ultra sonic-attenuation measurements is discussed. Associated with the calculation of the orientation factors, the modes of elastic-wave calculation were also evaluated, according to the procedure previously employed by Hinton and Green.&apos; The results of these calculations are not presented here since similar results have been reported earlier by Mus-grave.2 However, it should be mentioned that, although the qualitative features of Musgrave&apos;s results are preserved, the quantitative values are somewhat altered since different values for the elastic constants of zinc were used. ORIENTATION FACTORS It is convenient to refer to three different sets of coordinate axes, Fig. 1, in the following calculations. First, there are the Miller-Bravais axes which are used in describing the Miller-Bravais indices of planes and directions in a hexagonal crystal. These axes are denoted in the conventional manner by a,, a,, a,, and c. The second set of axes used is a rectangular Cartesian coordinate system located as shown in Fig. 1. Note that the x axis coincides with the close-packed direction al [2110], the z axis coincides with the c axis [0001], and the y axis is orthogonal to these two lying in the [0110 ] direction. Finally, a form of Eulerian angles is used to locate the position of the specimen axis. The angle between the specimen axis and the c axis is denoted by 0 and the angle between the projection of the specimen axis on the basal plane and the a, direction is denoted by ø. The majority of observations on plastic deformation of metal crystals indicates that such deformation is caused by the motion of dislocations on a limited number of crystalline planes and in a close-packed direction. In crystals with hexagonal symmetry the most prominent slip plane is the basal plane (0001) and the slip directiops- are one of the close-packed directions, [2110], [ 12101, or [ 1120]. To properly set the stage for the following work, several points must be clarified. First, we want to consider the contribution to the attenuation of elastic waves by dislocations only. While the Granato-Lucke, model for dislocation damping is presently the one most quoted, it is not necessary to refer the present work to it. The only assumption that need be made is that the motion of dislocations absorbs energy from the elastic wave and therefore attenuates it. It is immaterial here how this energy absorption occurs. However, because dislocations have been observed to move most freely on the (0001) plane for hexagonal crystals undergoing plastic deformation, it will be assumed that it is the dislocation motion on this plane which attributes to the observed attenuation. The basic orientation factor, Ok, is essentially the same as that used by Green and Hinton for fee crystals,4 defined by the equation

Jan 1, 1968

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