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By R. F. Domagala, M. Hansen, D. J. McPherson

Iodide zirconium was combined with calculated amounts of nitrided zirconium sponge and arc melted to prepare alloys in the 0 to 6 wt pct N region. Annealing treatments were carried out at 21 temperature levels. Metallographic examination of the heat-treated specimens permitted construction of the binary phase diagram from 0 to 6 pct N. Features of the diagram include the peritectic formation of both a and ß solid solutions. The maximum solubility of nitrogen is 0.8 pct in 8 zirconium and 4.8 pct in a zirconium. An X-ray study of nitrided materials was made in the range 6 to 13 wt pct N region because serious nitrogen losses were experienced when attempts were made to arc melt these high nitrogen alloys. PHASE relationships in the Zr-N system have been determined. Due to the inability to retain nitrogen in nitrogen-rich alloys subjected to arc melting, definitive work was possible only in the range 0 to 6 wt pct (0 to 30 atomic pct) N. Sufficient complementary X-ray work was done to permit construction of the binary phase diagram up to 13.3 wt pct (50 atomic pct) N (ZrN). Small ingots were prepared by arc melting master alloys with enough zirconium to produce an alloy of the desired composition. Following pretreatment, alloys were annealed at temperature levels between 600" and 2020°C. Determination of the phase boundaries was then accomplished by metallographic evaluation of specimens quenched from the various temperatures. Incipient melting techniques were used to corroborate solidus curves, and X-ray diffraction was employed to study the nitrogen-rich region of the a + ZrN phase field. Materials Westinghouse Grade 1 iodide zirconium crystal bar served as the base material for this investigation. The as-received bars were lightly sandblasted, pickled in a 20 pct HN0-5 pct HF aqueous solution, rinsed in water and acetone, and dried. They were rolled to about 1/32 in. strip and acid-pickled, followed by water and then acetone rinses. The material was sheared to approximately 1/4 in. squares, cleaned with acetone, and stored for use. Pure zirconium nitride is not commercially available, according to correspondence with possible suppliers. A literature survey indicated that nitrogen may be introduced into zirconium metal by passing nitrogen or ammonia gas over zirconium at an elevated temperature. After considerable experimentation, a train was devised whereby high quality nitrogen was passed through a series of bubblers to remove the last traces of oxygen, through an H,SO, bubbler and a cold trap to remove moisture, and finally over zirconium within a resistance furnace. Hand-picked Bureau of Mines magnesium-reduced zirconium sponge proved more amenable to nitriding than the crystal bar. The nitrogen used was extra pure gas purchased from Linde Air Products Co. The first two bubblers through which the gas was passed contained a solution suggested by L. F. Fieser" as being ideal for removing the last traces of oxygen from nitrogen. The solution contains 20 g KOH, 2 g sodium-anthraquinone ,9-sulphonate, and 15 g NaHSO, dissolved in 100 ml water. It is blood-red and turns brown when contaminated with oxygen. A saturated lead acetate solution, next in the series of bubblers, removed any H2S which might have formed in the 0, removal stage. An H,SO, bubbler and cold trap removed moisture before the nitrogen was passed over the zirconium. The zirconium was contained in a stainless steel screen within a Vycor or quartz tube in a resistance furnace. Unreacted nitrogen passed out of the system through a mercury bubbler. A nitriding run was performed in the following manner: Zirconium was placed in the screen within the furnace tube. The unit was assembled, suitably clamped off, and evacuated to remove the air and moisture within the tube and that trapped by the zirconium sponge. The unit was flushed with nitrogen and evacuated twice. After nitrogen was allowed to pass over the zirconium for about 30 min, the furnace was turned on; several hours were required to reach temperature. The reaction was allowed to continue at temperature for about 4 hr and the screen was then withdrawn from the hot zone of the furnace by means of a Nichrome or Kanthal rod. The screen and its contents were allowed to cool at the end of the tube and were then removed. At 800°C no more than about 5 pct N could be introduced into the sponge zirconium. This material was designated M-1 (master alloy No. 1) and reserved for the preparation of alloys in the dilute nitrogen region. A second set of experiments was conducted at 1000°C; no more than 7.2 pct N could be introduced into the zirconium at this temperature. Enough material (M-2) of this nitrogen content was prepared to produce a second set of alloys to complement the first and to provide alloys in the 0 to 6 wt pct N range. It was hoped that a diffusion anneal plus a re-nitriding might yield a material of higher nitrogen content. The material containing about 7 pct N was therefore given a homogenization treatment at 1000°C for 6 hr under an argon atmosphere. It was

Jan 1, 1957

By Shiou-Chuan Sun

THE fact that coals of different ranks and even of the same rank differ greatly in their amenability to iroth flotation is well known. In recognition of the need for an explanation of this phenomenon, two hypotheses have been suggested. Wilkinsl reported that the floatability of coals increased with an increase of the carbon content or rank. This postulate is handicapped by the fact that bituminous coals that possess moderate carbon contents are actually more floatable than anthracite coals that have high carbon contents, as shown in columns 6 and 9 of Table I. Taggart and his associates' implied that the difference of floatability between bituminous and anthracite coal was caused by the variation of carbon-hydrogen ratio. This is not applicable to the relative floatability of other coals and carbons. For example, column 11 of Table I shows that the carbon-hydrogen ratios of low-floating lignitic coal and non-floating animal charcoal are not only smaller than the moderate-floating anthracite coal, but are also similar to the high-floating bituminous coal. Furthermore, according to this hypothesis, high temperature coke-A (464), Ceylon graphite (1238), and lamp-black (357), all possessing extremely high carbon-hydrogen ratios, should be less floatable than other substances having much lower carbon-hydrogen ratios such as high volatile-B bituminous coal (11.9 to 22), anthracite coal (35.7 to 60.5), lignitic coal (15.6 to 33.6), and charcoal (13 to 26.2). However the former group is actually more floatable than the latter group. In this paper, a surface components hypothesis is Proposed to explain the different floatabilities of coals, carbons, and hydrocarbon minerals. The validity of the hypothesis is experimentally supported by the actual floatability, natural floatability, wettability, and adsorbability for neutral oils of coals, carbons, and hydrocarbon minerals tested. The combustible recovery of the flotation results, as used in this paper. was calculated from Eq. 1: P (100-Ep) 100 RWCP Rc= [1] F (100-E,) C, where R, is the percent combustible recovery; F and P are, respectively, the weight of feed and the weight of concentrate or product; E, and Ep are, respectively, the total percent of ash plus moisture in feed and in concentrate; Ru. is the percent weight recovery: and C, and C, are, respectively, the percent of combustible in feed and in concentrate. Except for ash and moisture content, all chemical components of a coal are assumed combustible. The experimental work included studies on flotation, ultimate and proximate analyses, contact angle tests, extractions of bitumen-A with benzene, adsorptions for liquid hydrocarbons, and wetting tests. Most of the flotation experiments were performed in a laboratory Fagergren machine; others were tested in a small Denver machine. The solid feed for the former was 300 g and for the latter was 30 g. The solid materials used for flotation were crushed to —48 mesh. After the mineral pulp in the flotation cell was agitated for 6 min and the pH was adjusted to 7.5 & 0.2 with sodium hydroxide or hydrochloric acid, a petroleum light oil having a viscosity of 5.73 centipoises at 77 °F was added and conditioned for 2 min. Finally, pine oil was introduced and the froth was collected for exactly 3 min. The weight ratio of petroleum light oil to pine oil was kept constant at 1.5 to 1. Tap water was used for all flotation tests. Contact angles were measured with a captive bubble machine. For each coal sample, three specimens were mounted in transoptic mounts and polished with levigated alumina, first on a sheet glass, then on a cloth-covered metal polishing wheel. The polished specimen was first washed with distilled water and wiped thoroughly on a cleaned linen pad, then transferred into the pyrex cell of the captive bubble machine and conditioned for 6 min., and finally measured for contact angles at three or more points. Except where otherwise stated, the induction time for each measurement was 1 min. The contact angle representing each material was obtained by averaging the measurements of three specimens. The linen pad was first washed with warm distilled water, then boiled 30 min in a 2N sodium hydroxide solution, and finally washed with distilled water until no trace of sodium hydroxide could be detected in the decanted solution. The cleaned linen pad was stored under distilled water. Immediately before using, the pad was rewashed and transferred into a clean pyrex petri dish partly filled with distilled water. The glassware and rubber gloves used were cleaned by soaking in sulphuric acid-potassium dichromate cleaning solution, followed by rinsing with distilled water. The polished specimens were handled only by glass forceps. The ultimate and proximate analyses were made in accordance with the ASTM standard procedures for coal and coke. The extractable bitumen-A was determined by weighing 1 g of —100 mesh sample and placing it in a desiccated and weighed ASTM aluminum-extraction thimble. The thimble was placed in condenser hooks and inserted into an extraction flask containing 100 cu cm of benzene. The flask was heated and the benzene vapor was condensed by water coils. At the end of 24 hr of percolation, the thimble was removed, desiccated, and weighed. Loss in weight of sample was taken as bitumen-A and calculated to dry and ash-free basis.

Jan 1, 1955

By J. A. Berger, G. W. Wiener

Several transition metal nitrides have been prepared and their saturation magnetization determined. On the basis of an atomic model of ferromagnetism involving a consideration of nearest neighbor interactions and the assumption that all atomic moments of the metal point in the same direction, it appears that the nitrogen interacts with d-shell of the transition metal in such a way as to reduce the magnetic moment. THERE is a large class of materials having metallic properties which are formed by a combination of hydrogen, boron, carbon, oxygen, or nitrogen with the transition metals. Several attempts have been made to establish the type of metal-nonmetal bonding in these interstitial alloys because it is believed that many of the physical properties of these materials are determined by the characteristics of this bond. Several of these alloys are ferromagnetic, and thus a powerful method is available for investigating the structures in a direct manner by measuring the saturation magnetization. The latter is a fundamental property of ferromagnetic metals and alloys which depends primarily on the electron distribution surrounding the atom. For the first row of transition metals, this refers specifically to the 3 d-shell. Since bonding involves the electronic configuration between atoms, there is reason to suppose that a relationship exists between ferromagnetism and bond type. In the case of the interstitial structures studied in this work, bonding will refer to the distribution of electrons between the transition metal and the nonmetal. Since these alloys have metallic properties, it is further proposed that any bonding interactions will involve the outer p-shell of the interstitial element and the incomplete d-shell of the transition metal. If this is the case, then the relationship between ferromagnetism and metal-non-metal bonding is established qualitatively. In order to investigate the subject quantitatively, certain transition metal nitrides were chosen because they have simple crystal structures, are ordered alloys, and are ferromagnetic. They also have sufficiently high saturation magnetization to be of technical interest. Currently there are two major theories of ferromagnetism, each of which has been applied to the interpretation of the saturation magnetization in terms of atomic structure. They are usually referred to as the band theory and the atomic theory. The former has found widespread application to the study of pure metals and certain solid-solution allays. However, it has not been applied to the interstitial structures or ordered alloys because it does not interpret the properties directly in terms of the crystal structure. The atomic theory on the other hand is especially suited to the study of interstitial structures because it permits an interpretation of ferromagnetic phenomena in terms of the crystal geometry. As has been pointed out previously, the nitrides have simple ordered crystal structures and, therefore, the choice of the atomic theory for the interpretation of the data is a natural one. One of the prime difficulties with the atomistic theory is that its mathematical justification is much more difficult, and for this reason its general acceptance will depend to a large extent on the value it has in explaining and predicting the results of experiment. Before the presentation of the theoretical basis for understanding the metal-nonmetal bond, it is useful to review the ideas existing prior to this work. Four different interpretations have been given to the metal-nonmetal bond. These are summarized as follows: 1—acceptance of electrons by the nonmetal from the incomplete d-shell of the transition metal, 2—transfer of electrons from the nonmetal to the incomplete shell of the transition metal, 3—no exchange of electrons between the two atoms, and 4— a resonating type of bond involving the p electrons of the interstitial atom giving rise to half bonds. Zener'-4 in a recent series of papers has proposed a new theory of ferromagnetism and has developed an explanation of the observed saturation magnetization of iron nitride (Fe,N) using the concept that nitrogen accepts electrons from the 3d-shell of iron. Jack," on the basis of atom size considerations in iron carbonitrides, has proposed that nitrogen transfers or donates electrons to the inner 3d-shell. He found that the effective size of the carbon atom was less than that of nitrogen and thus suggested that the interstitial atoms give up electrons. Kiessling" has studied the borides of several transition metal atoms and proposed that boron loses one p electron to the transition metal. He postulated that the additional electron added to the metal lattice compensates for the loss in metallic properties which results from the increased metal-metal atom separation. GuillaudT3" has proposed similar arguments from some recent magnetic studies he had made on manganese nitride. However, he did not base his conclusions on a quantitative argument. Pauling," in a recent paper, discussed electron transfer in in-termetallic compounds. He classified nitrogen as a hyperelectronic atom which can increase its valence by giving up electrons. He classified the transition metals as buffer atoms which are capable of either accepting or giving UP an electron. He pointed out that two factors are operating which promote electron transfer because they lead to increased stability. The first is an increase in the number of bonds, and the second is a decrease in the electric charges on the atoms. These ideas when applied to the interstitial nitrides would indicate a viewpoint favoring electron transfer by nitrogen to the transition metal. Hagg7s arguments in favor of no exchange are adequately summarized by Wells." Implicitly, Hagg

Jan 1, 1956

By Mark Malamphy

MOST igneous rocks, and particularly those of the basic type, con-tain relatively high percentages of magnetite and other iron oxides, which give them moderately high magnetic susceptibilities and make them responsible for an important percentage of the magnetic anom-alies observed during the course of geophysical surveys by the mag-netic method. The irregular form and high susceptibility of the igneous rocks make them readily polarizable and the position and strength of the magnetic poles formed are not always explicable by theoretical calculations of anomalies due to certain assumed forms in the normal magnetic field of the area under study. During the last four years our field work has permitted us to make detailed observations of anomalies due to igneous rocks in all the states of south Brazil. The form and position of the igneous rock at some places were fairly accurately known, which made it possible to calculate theoretical anomalies and compare them with the observed data. Some of the data we obtained have been published and additional cases are now awaiting publication. In this paper, we will present a few specific cases where the geology is moderately well known and which we believe to be of considerable value as representing typical magnetic anomalies in the south magnetic hemisphere, where the inclination of the magnetic field is very low. Although the vertical intensity is negative in the Southern Hemis-phere, we have adopted the convention of plotting increase in intensity (negative) as a positive anomaly; i.e., above the x axis. This conven-tion facilitates comparison with anomalies in the north.

Jan 1, 1936

By C. E. Lundin

The Ce-H system was investigated in the temperature range, 573° to 1023°K, and the pressure range, 10-3 to 630 Torr, as a function of 'composition up to 72 at. pct H. Families of isothermal arid isopleth curves were plotted from the pressure-terr~perature-composition relationships. From these curves the solubility relationships were determined for the system. The isopleths are analytically represented by equilibrium dissociation pressure equations. The relative partial molal enthalpzes and entropies of solution of hydvogen in the systerrz were calculated fronz the dissociation pressure equulions and are tabulated. The integral free energies, enthalpies, and entropies of mixing in the Ce-H system were determined from the relative partial quantities and are also tabulated. The standard free energy, enthelpy, and entvopy of reaction of the dihydride phase at kcal per kcal per mole H2, and ?S° = -34. 1 cal per deg mole H2, respectively. The equilibrium dissociation pressure equation in the two-phase region is: UNTIL recently very little was known of the detailed solubility and thermodynamic relationships of the Ce-H system. Two previous investigations1,2 are noteworthy. However, significant discrepancies and omissions exist on analyzing them. The work of Mulford and Holley1 on cerium did not clearly delineate the boundaries of the two-phase region, Cess - CeH2-x. The plateau partial pressures were not thoroughly defined and were considerably displaced in pressure compared to those from the work of Warf and Korst.2 These latter authors concentrated their studies primarily from 823° to 1023°K in the pressure range of 1 to 760 Torr. No data were determined to outline the regions of primary solid solubility and the hydride phase. Also the establishment of the plateau partial pressures was rather limited in scope. In neither work was a treatment conducted of the relative partial molal enthalpies and entropies of solution of hydrogen in the single-phase regions and the integral thermodynamic quantities of mixing throughout the system. Therefore, it was the objective of this research to determine the complete equilibrium solubility relationships and thermodynamic data for the system by pressure-temperature-composition studies. EXPERIMENTAL PROCEDURE The cerium metal for this study was donated by the Reno Metallurgy Research Center of the Bureau of Mines. Total impurity content was 0.13 pct with only 60 ppm O. The metal was checked metallographically and contained only minor amounts of second phase compared to cerium from other sources. Specimen preparation was done in a dry box flushed with argon gas. The surface of a small rectangular piece of cerium (about 0.2 g) was filed with a clean, mill file. Final weighing was done in a tared enclosed vial containing argon gas. The specimen was then loaded quickly into the reaction chamber which was purged several times with high-purity hydrogen gas and then allowed to pump to about 10-6 Torr. The furnace was heated to the reaction temperature and the run started. The equipment used to conduct the hydriding was a Sievert's-type apparatus. Basically it consisted of a source for pure hydrogen, a precision gas-measuring burette, a heated reaction chamber, a McLeod gage, and a mercury manometer. Pure hydrogen was supplied by the thermal decomposition of uranium hydride. The 100-ml precision gas burette was graduated to 0.1-ml divisions and was used to measure the quantity of gas and admit it to the chamber. The reaction chamber was a quartz tube. Prior to each run, the cerium specimen was wrapped in a tungsten foil capsule to prevent reaction of the cerium with the quartz. Control of the temperature was achieved within ±1°K. Pressures in the manometer range were measured to ±0.5 Torr and in the McLeod range (10-3 to 5 Torr) to ±3 pct. The compositions of hydrogen in cerium were calculated in terms of hydrogen to cerium atomic ratio. These compositions were estimated to be ±0.01 H/Ce ratio. The technique used to study the equilibrium pressure-temperature-composition relationships of the Ce-H system was to develop experimentally a family of isothermal curves of composition vs pressure. The range of pressure through which each isotherm was developed was from 10-9 to about 630 Torr in the temperature interval, 573° to 1023°K. RESULTS AND DISCUSSION The hydriding characteristics of cerium are iso-morphous with those of the elements of the light-rare-earth group (lanthanum, cerium, praseodymium, and neodymium) wherein the region from the dihydride to trihydride is continuously single phase.' The structure of this phase is fcc.3 The heavy rare earths form a trihydride,2 which is hcp, separated by a two-phase region from the fcc dihydride phase. The Ce-H system is represented by the family of experimental isotherms in Fig. 1. Due to the small scale required to draw the curves, the experimental points are omitted; however, a total of 240 experimental data points were taken to prepare these curves. The solubility relationships can be deduced therefrom. Three distinct regions of partial pressure and composition can be seen. The region of cerium solid solution is represented by the rapidly rising isotherms in the dilute composition range. In accordance with Gibbs Phase Rule only one solid phase, the cerium solid so-

Jan 1, 1967

By S. Usui, I. Iwasaki

The effect of pH on the adsorption of dodecylamine at the mercury-aqueous solution interface was investigated by differential capacity and electrocapillary measurements. With dodecylammonium acetate, the differential capacity curves showed two desorption peaks in the cathodic branch with their relative intensities varying with the solution pH. With dodecyltrimethylammonium chloride, only one cathodic de-sorption peak was observed in the same pH range. Through thermodynamic analysis of the electrocapillary curves, the adsorption density of undissociated amine was evaluated separately from that of aminium ion. The adsorption densities of the un-dissociated amine and of the total amine increased with increasing pH. The ratio at the interface of undissociated amine to aminium ion was several orders of magnitude greater than the ratio in the solution and increased with increasing pH. The potential at the closest distance of approach of counter ions to the mercury surface was compared with values of zeta potential on quartz previously reported. The most important variable in the flotation separation of minerals is probably the pH of the pulp, and a number of theories have been proposed to explain its effect on the condition of the mineral surfaces, on the dissociation of collectors and of inorganic and organic species (accidentally present or intentionally added) in the pulp, and on the mineral-collector interaction. In the development of a theoretical background for oxide flotation systems, an experimental approach based on electrokinetic measurements has been of much value, although the effect of pH becomes confounded since it governs both the electrochemical conditions of the oxide surface and the dissociation of the collector. For investigation of the adsorption behavior of long-chain collectors on oxide minerals, however, electrokinetic potential measurements are the most widely used technique. Hydrogen and hydroxyl ions are found to be the potential determining species, thereby governing the interfacial electrical conditions. The electrostatic interaction between the charged mineral surfaces and ionized collectors is regarded as the driving force for the adsorption of the collectors. An association of alkylamine collectors adsorbed on quartz surfaces has been postulated from streaming potential measurements, and a term "hemi-micelles" has been proposed.' The possibilities of coad-sorption of undissociated amine along with aminium ion has been inferred from contact angle measurements? and from adsorption studies.~ Electrochemical titration as applied to silver sulfide provides a more quantitative approach to the analysis of the electrical double layer at an ionic solid-solution interfaceqG and the electrochemical evidence for the adsorption of amine at pH 4.7 indicates a specific affinity of dodecylammonium ion towards silver sulfide surfaces, whereas at pH 9.2 the adsorbed species might be free arnine." A combination of differential capacity and electrocapillary measurements on a dropping mercury electrode was reported to be a sensitive method of provid- ing reliable information on the adsorption behavior of dodecylammonium acetate (DAA) at a natural (near neutral) pH.? It was also shown that there were striking similarities in the properties of the double layer and in the adsorption behavior of the amine on mercury and on such ionic solids as quartz, silver sulfide, and silver iodide. The effect of pH on the differential capacity curves at a mercury-sodium fluoride solution interface has been investigated by Austin and Parsonss who reported that between pH 7 and pH 11 there was very little effect. In the present paper, the adsorption behavior of DAA was investigated as a function of pH through differential capacity and electrocapillary measurements and the information gathered was correlated with that available in literature on quartz and silver sulfide. Experimental The apparatus and the method used for determining the differential capacity and the electrocapillary curves were identical to those described previously.' The ionic strength of the supporting electrolyte was fixed at 0.1 M with potassium fluoride, and the pH of the solution with potassium hydroxide. Only the neutral to alkaline range was covered in order to avoid the dissolution of the glass vessel with hydrofluoric acid. Results In Fig. 1 the differential capacity has been plotted against the applied potential at a DAA concentration of 10-' M at three different pH values. The curves are characterized by one capacity peak in the anodic branch, by two capacity peaks in the cathodic branch, and by a marked depression in capacity between the peaks. The depression indicates an adsorption of the arnine in this potential range. One of the cathodic peaks appears at pH 7.3 near -1.4 v and decreases with increasing pH. The other appears at pH 8.9 near —1.2 v and increases with increasing pH. At pH 9.6 only the latter peak is observed. Beyond the cathodic peaks, all the curves tend to converge with the curve in the absence of DAA, implying that two different species are being desorbed in this potential region. The anodic peak near 0.0 v increases markedly with increasing pH. The well-defined anodic peaks at pH 8.9 and 9.6 were accompanied by an appreciable increase in the current flow (in excess of O.luA), and, therefore, is a "pseudo-capacity"'" due to a

Jan 1, 1971

By J. E. Hove

The high temperature physical properties of graphite are reviewed and interpreted in the light of present day knowledge of the mechanisms affecting these properties. The thermal and mechanical behaviors only are discussed and, whenever possible, comparisons are made with other refractory materials. Possible further studies are indicated, including some carbide work. AS long as the term high temperature implied only temperatures up to about 1000°C, the materials problems which arose could usually be handled by fairly conventional metal alloy types, such as the Co-Cr-Ni superalloys, for which there exists a great deal of technology. Perhaps this temperature can still be considered an upper limit for normal applications, but it is certainly true that the number of abnormal applications is increasing rapidly. The advent, in recent years, of ram-jet and rocket missiles and of high power nuclear reactor heat sources has raised a host of questions concerning the basic problem of what material to use in the temperature range up to 2000°C and higher. While there are, of course, many metals, in the second and third transition series, which melt at considerably higher temperatures than this, these metals are, at present, pretty well excluded from practical use by other considerations, such as re-crystallization, chemical activity, or excessive plastic deformation. The behavior of metals, from the standpoint of dislocation theory, is just beginning to be understood and thus there is some hope for the future development of very high temperature metals, but the immediate problems would most logically appear to have solutions involving the nonmetals, such as the refractory ceramics and graphite. For this reason, there is presently a great deal of engineering and experimental research being performed on the latter materials, much of this research being exploratory in the sense of gathering new property data. The situation, so far as graphite is concerned, is somewhat more fortunate than with the other refractory solids in the sense that a great deal is already known about its basic properties. This stems both from the fact that the carbon-carbon bond has been of interest to chemists for a long time (and graphite can be considered as a very large aromatic molecule, if desired) and the fact that its properties, both as a function of temperature and of radiation damage, are of critical importance to nuclear reactor designers. It is still true, of course, that such fundamental questions as why graphite remains solid to such a high temperature and why it has such a high thermal conductivity cannot entirely be answered at the present time. It is, nonetheless, meaningful and instructive to consider such problems in the light of existing knowledge. This is what the present paper will attempt to do. Before going on, it may be appropriate to classify graphite and justify its discussion before readers primarily interested in metals. Graphite is comparatively unique among materials in that there is always a property or group of properties which precludes calling it either a metal, a semiconductor, or a ceramic. It has the high electrical and thermal conductivities of a metal, but the artificial, poly-crystalline types show a negative thermal coefficient of electrical resistivity, generally characteristic of semiconductors. On the other hand, semiconductors, by definition, show an ever increasing resistivity as the temperature is lowered, whereas graphite approaches a finite, and, indeed, a rather low resistivity in the region of 10°K and, furthermore, a good single crystal of graphite has a positive temperature coefficient, as for a metal.' On still another hand, its porosity and brittleness at lower temperatures would put graphite in the ceramic class although, unlike most ceramics, it is readily machinable and has a high resistance to thermal shock. All things considered, it is probably more nearly appropriate to call graphite a metal than anything else. Although certainly outstanding in some ways, graphite has its peculiarities and, especially if the reader is unfamiliar with the data, it is probably valuable to review some representative property variations at high temperature. This review is meant to be mainly illustrative and no attempt has been made to be exhaustive. Review of High Temperature Properties At ordinary pressures, graphite does not melt, but sublimes directly into the gaseous phase at about 3700°C. Although the phase equilibrium diagram is still in some doubt, graphite will melt, at a slightly higher temperature, at pressures in excess of about 100 atm. The chief difficulty of using graphite in an oxidizing atmosphere is that the reaction rate becomes quite high at fairly low temperatures. If a threshold oxidation temperature2 is defined as the temperature at which graphite loses 1 pet of its weight in 24 hr, the value in air is 450°, the value in steam is 700°, and the value in carbon dioxide is 900°C. Efforts are presently being made to raise this threshold temperature either by impregnation with a retardant of some type (sodium tung-state and phosphoric acid, for example) or by a suitable metal or oxide coating. It is probably fair to say that, to date, these attempts have not shown any outstanding success in all respects. Most commercial graphites are fabricated by impregnating a carbon flour (say of coke or lampblack particles) with some type of hydrocarbon

Jan 1, 1959

By E. P. Whittemore

Experience with the ASC multipoint gas lift system was obtained in Colonia zone of the West Montalvo field near Oxnard, Calif. The wells in this pool produce from depths varying from 10,500 to 12,000 ft. Oil gravity is generally 14 to 15' API with a few extremes of 12 and 20" API. Some salt water is produced which causes some very viscous emulsions. Viscosities at 150F (which is the approximate wellhead temperature) vary from 5,000 to 100,000 SSU. Most of the production is by gas lift, although a few wells are produced by rod and hydraulic pump. About half of the gas-lift wells are on continuous flow and the remainder are on intermittent lift using large, ported, pilot-operated valves for single-point transfer of gas from casing to tubing. Gas-liquid ratios vary from about 6 to 10 Mcf/bbl of gross fluid lifted. Wells are produced to a 450-psi trap system. The following remarks will be confined to intermittent lift only, since this is the type of lift which has been achieved with the ASC valve system. The maximum gross fluid which has been produced by single-point intermittent lift is about 350 B/D in 3-in. tubing and 200 B/D in 21/2-in. tubing with gas-liquid ratios of approximately 7 to 9 Mcf/bbl. Some design changes could reduce this ratio. The ASC multipoint system has provided production as high as 480 BOPD in 21/2-in. tubing with gas-liquid ratios just under 4 Mcf/bbl. To be able to apply the multipoint system, it is recommended that a detailed explanation be obtained concerning transition-point pressure and stabilizer setting—what its significance is to the string design, how it may work for or against the operation of the well, how it is related to tubing sensitivity and how it affects the unloading operation. The unloading operation may only be of academic interest in a technical paper, but to the production foreman, unloading and setting the valves in operation is a very real problem and should be understood in detail. One item touched lightly in the paper was the unloading valve. This valve controls the maximum pressure at which the well can be operated. When lifting heavy viscous fluids, it is most important to set this valve for the maximum possible realistic operating pressure at the surface. If the well lifts easily, it is simple to set the ASC valves at a lower operating pressure and the unloading valve will remain closed; but if the well happens to be heavier to lift than anticipated, it may be desirable to operate on the unloading valve itself and use all the energy obtainable at the bottom of the hole. In the Colonia pool very heavy wet-gas gradients are experienced due to the viscosity of the liquid and the dense mist which is left behind a slug of fluid. There are many combination strings of 3- and 21/2-in. tubing. This aggravates the wet-gas gradient problem and provides wet-gas gradients of about 50 to 70 psi/1,000. An advantage which multipoint lift has provided is increased slug efficiency through better maintenance of pressure under the slug and decreased fall back as the slug passes up the tubing. By using multipoint injection, wet-gas gradients have been reduced to about 30 psi/1,000. This has reduced bottom-hole operating pressure and given a production increase. The ASC valve is not a simple device. It's operation is difficult to understand, and it must be understood to be used efficiently in gas-lift design. Operating problems are difficult to diagnose—whether they be caused by the fluid lifted, valve malfunction, lift gas rate, or operating pressure. Calculations and reasoning are required to find out what is causing the problem. Inherent in the ASC valve is the inability to create large pressure differentials across a slug. Large differentials may be required to overcome the inertia of very viscous fluid as it is being accelerated in the bottom of the hole. This is tied back to the design of the unloading valve and is one reason for the importance of setting the unloading valve for the highest possible operating pressure. ~u; to the narrow spread the ASC valves provide, it is impossible to cycle slower than about 24 cycles/day on choke control. If small production of 150 BOPD and less is expected, a surface time-cycle controller will be required if the most economical operation is to be achieved. To achieve a satisfactory operation, the operator must keep a record of any changes made in the operating pressure of the ASC valves. Anything which may cause changes in casing pressure in excess of the stabilizer setting will change the valve operating pressure, and if this is not noted from daily inspection of the well casing-tubing pressure recorder charts, the operator will lose control of the well. Significant results can be achieved using ASC valves; however, considerable knowledge is required to design with them, and attention to detail is required for satisfactory field operation.

Jan 1, 1965

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By A. Spilners, M. Simnad

The rates of formation of the different oxides on molybdenum in pure oxygen at 1 atm pressure have been determined in the temperature range 500° to 770°C. The rate of vaporization of MOO, is linear with time, and the energy of activation for its vaporization is 53,000 cal per mol below 650°C and 89,600 cal per mol at temperatures above 650°C. The ratio Mo03(vapor.lzing)/MoOS3(suriace) increases in a complicated manner with time and temperature. There is a maximum in the total rate of oxidation at 6W°C. At temperatures below 600°C, an activation energy of 48,900 cal per mol for the formation of total MOO, on molybdenum has been evaluated. The suboxide Moo2 does not increase beyond a very small critical thickness. At temperatures above 725°C, catastrophic oxidation of an autocatalytic nature was encountered. Pronounced pitting of the metal was found to occur in the temperature range 550° to 650°C. Marker movement experiments indicate that the oxides on molybdenum grow almost entirely by diffusion of oxygen anions. USEFUL life of molybdenum in air at elevated temperatures is limited by the unprotective nature of its oxide which begins to volatilize at moderate temperatures. Although the oxide/metal volume ratio is greater than one, the protective nature of the oxide film is very limited. Gulbransen and Hickman' have shown, by means of electron diffraction studies, that the oxides formed during the oxidation of molybdenum are MOO, and MOO,. The dioxide is the one present next to the metal surface and the trioxide is formed by the oxidation of the dioxide. Molybdenum dioxide is a brownish-black oxide which can be reduced by hydrogen at about 500°C. Molybdenum trioxide has a colorless transparent rhombic crystal structure when sublimed, but on the metal surface it has a yellowish-white fibrous structure. It is reported to be volatile at temperatures above 500" and melts at 795°C. It is soluble in ammonia, which does not affect the dioxide or the metal. In his extensive and classic investigations of the oxidation of metals, Gulbransen2 has studied the formation of thin oxide films on molybdenum in the temperature range 250" to 523°C. These experiments were carried out in a vacuum microbalance, and the effect of pressure (in the range 10-6 yo 76 mm Hg), surface preparation, concentration of inert gas in the lattice, cycling procedures in temperature, and vacuum effect were studied. The oxidation was found to follow the parabolic law from 250" to 450°C and deviations started to occur at 450 °C. The rates of evaporation of a thick oxide film were also studied at temperatures of 474" to 523°C. In vacua of the order of 10- km Hg and at elevated temperatures, an oxidation process was observed, since the oxide that formed at these low pressures consisted of MOO, which has a protective action to further reaction in vacua at temperatures up to 1000°C. Electron diffraction studies showed that, as the film thickened in the low temperature range, MOO8 became predominant on the surface. Above 400°C MOO, was no longer observed, MOO, being the only oxide detected. The failure to detect MOO, on the surface of the film formed at the higher temperatures does not militate against the formation of this oxide, since according to free energy data MOO3, is stable up to much higher temperatures. At the low pressures employed, this oxide would volatilize off as soon as it was formed. Its vapor pressure is relatively high and is given by the equations" log p(mm iig) = -16,140 T-1 -5.53 log T + 30.69 (25°C—melting point) log p(mm He) = -14,560 T-1 -7.04 log T+1 + 34.07 (melting-boiling point). Lustman4 has reported some results on the scaling of molybdenum in air which indicate a discontinuity at the melting point of MOO, (795°C). Above the melting point of MOO,, oxidation is accompanied by loss of weight, since the oxide formed flows off the surface as soon as it is formed.5,6 Qathenau and Meijering7 point out that the eutectic MOO2-MOO3 melts at 778C, and they ascribe the catastrophic oxidation of alloys of high molybdenum content to the formation of low melting point eutectics of MOO3 with the oxides of the melts present. Fontana and Leslie -explain the same phenomenon in terms of the volatility of MOO,, which leads to the formation of a porous scale. Recent unpublished work by Speiser9 n the oxidation of molybdenum in air at temperatures between 480" and 960°C shows that the rate of weight change of molybdenum is controlled by the relationship between the rates of formation and evaporation of MOO,. They have measured the rates of evaporation of Moo3 in air at different temperatures and estimated an activation energy of 46,900 cal. This compares with the value of 50,800 cal per mol obtained by Gulbransen for the rate of sublimation of MOO, into a vacuum.

Jan 1, 1956

By James R. Stuart

AS the available remaining properties of iron ore reserves on the Mesabi Range are opened up for mining, the various properties located under lake beds are brought nearer an active status. The actual physical problems involved in stripping these properties do not act as a deterrent so much as the legal and political problems that are encountered. When it is proposed to destroy a natural lake that has been used by the public for many years, much local as well as state opposition may be encountered regarding its destruction. Public hearings must be held and some adverse publicity is likely to result. The ownership of the ore under the lake and the rights of the abutting property owners must be settled, and protection from damage caused by a disturbance in surface and subsurface drainage is likely to be demanded by property owners some distance from the proposed mine area. The Embarrass Mine, located near Biwabik, Minn., falls into this classification. A portion of the orebody lies under what was formerly Syracuse Lake, this body of water having been removed in the process of stripping the mine. An additional problem in the case of a meandered body of water is the establishment of a meander line that can be projected downward as mining progresses to form the basis for a satisfactory division between lake bed and upland ore shipments for royalty purposes. Fig. 1 illustrates the complications encountered in maintaining these divisions. A balance point was agreed upon in the center of the lake to make an equable division of lake bed ore to the abutting properties. The entire lake bed has since been adjudged the property of Minnesota. Lake Characteristics Lake bed stripping problems with which this paper is concerned necessarily are limited to a specific type of lake, namely the glacial lakes of the Lake Superior region. One characteristic common to these bodies of water is a deposit of fine black mud or silt on the bottom, frequently underlain by a layer of impervious blue clay. This is also true of the muskeg areas of the region, which present almost identical problems as lakes in stripping. The actual removal of the water and the lake bed material is a routine matter more or less standardized as to equipment, and the period of time required can be estimated easily on the basis of volume and capacity. More important than the foregoing is the execution of preliminary work, and above all, the timing involved. An account could be prepared based entirely on statistical and cost data which would give a very fair picture of the time required and cash outlay needed to effect the removal of a body of water preliminary to stripping the orebody. However, the real interest from the standpoint of the operator and the engineer who carry responsibility for completion of the job lies in the unexpected emergencies and the action of various materials involved in the stripping when the balance has been upset through diversion of water courses and the reduction of the lake level. Runoff and Drainage Lakes are located in natural basins that catch all the rain water and runoff water for a considerable area. Where a lake is involved having an inlet and outlet or a sizeable water course running through it, the drainage area may include a watershed covering many square miles. All available data then must be collected to supply a history extending over as many years for which information can be gathered on the flow of streams, annual rain and snowfall, and most important, the peak flows to be expected. Where the diversion of a stream around the stripping area is a part of the problem, this last factor is of great importance since it controls the cross-section to be selected for the diversion channel and the volume to be removed in its excavation, as well as affecting the hydraulic considerations to be met in the design of the completed channel. Characteristic material in the overburden found at the Embarrass Mine is illustrated in Fig. 2. Well Pumping Pumping from the well holes was started well in advance of the draining of the lake. Fig. 3 shows a gradual lowering of the water table with no noticeable fluctuations during the period in which the lake was being dewatered. Unfortunately, because of tight ground, a maximum flow to the wells was not maintained. This retarded the rate at which the water table was reduced so that in the course of stripping the excavation soon extended below the water table, and the great bulk of the pumping was handled from a system of sumps in the pit itself. Any dewatering program projected by prepumping from wells, a glorified well point system, would have to be started well in advance of the stripping to be of any great advantage. Preliminary drainage of the surface over the mine area is entirely apart from the actual elimination of the lake bed itself. Since the lake is what is called a perched water table because of the impervious character of the lake bottom, the adjoining surface may be dewatered below the surface of the existing lake and the flow will not be affected by the proximity of that body of water. This condition actually has been demonstrated through the establishment of a number of observation holes where a small churn drill was used to put down the holes and a 3-in. pipe was installed for taking water level

Jan 1, 1952

By A. R. Calawa, T. C. Harman, M. Finn, P. Youtz

A study has been made of the growing, annealing, and diffusion parameters in PbSe, Pb1-ySnySe, and Pb1-xSnxTe. Single crystals of these materials have been grown using the Bridgman technique. For all of the above materials the as-grown crystals are p type with high carrier densities. To reduce the carrier concentration and increase the carrier mobility, the samples are annealed either isothermally or by a two-zone method. From isothermal anneals, the liquidus-solidus boundary on the metal-rich side of the stoichiometric composition has been obtained for some alloys of Pb1-xSnxTe and on both the metal- and seleniunz-rich sides for PbSe and alloys of Pbl-ySnySe. In Pbo.935 Sno.065 Se carrier concentrations as low as 5 x1016 Cm-3 and mobilities as high as 44,000 sq cm v-1 sec-1 at 77°K have been obtained. Inter diffusion parameters mere also studied. The ddiffusion experiments mere identical to the isothermal or two-zone annealing experiments except that the samples were removed prior to complete equilibration. The resulting p-n junction depths were determined by sectioning and thermal probing. Inter diffusion coefficients for various temperatures were calculated for both PbSe and Pb0.93Sn0.0,Se. RECENTLY, there has been considerable interest in the PbTe-SnTe and PbSe-SnSe alloys with the rock salt crystal structure. The unusual feature of these systems is the variation of energy gap EG with composition. Several investigations1-3 have shown that EG for the lead chalcogenides decreases as the tin content increases, goes through zero, and then increases again with further increase in tin content. The possibility of obtaining an arbitrary energy gap by selecting the composition is an especially attractive feature of these alloys for applications involving long-wavelength infrared detectors and lasers. In addition, some unusual magneto-optical, galvanomagnetic, and thermomag-netic effects should occur for alloys with low band gaps. If uncompensated low carrier density crystals can be obtained, then a small carrier effective mass, a large dielectric constant, and the resultant high carrier mobility should yield enormous effects at low temperature in a magnetic field. The relative variation of the energy gap with pressure should also be very large for these low gap materials. The primary purpose of this paper is to provide some information concerning the preparation of low carrier concentra- tion, high carrier mobility, and homogeneous single crystals with a predetermined alloy composition. I) DETERMINATION OF ALLOY COMPOSITIONS In all of the work described in this paper, the composition of lead and tin chalcogenides in the alloys was determined by electron microprobe analysis. Separate X-ray spectrometers are used to make simultaneous intensity measurements of the Pb La1 and Sn La1 lines emitted by the sample under excitation by a beam of 25 kev electrons focused to a spot about 2 µm in diam. These intensities are compared to the intensities of the same lines emitted by standards under the same conditions. The standards used are the terminal compounds of each pseudobinary system, i.e., PbTe and SnTe for Pbl-xSnxTe alloys, PbSe and SnSe for Pbl-ySnySe alloys. The composition of the sample is then obtained from theoretical calibration curves which relate the weight fractions of lead and tin in the alloy to the measured ratios of X-ray intensities for the sample and the standards. The lead and tin calibration curves for each alloy system were calculated by using corrections for backscattered electrons,4 ionization,5 and absorption,6 and assuming that the atom fraction of tellurium or selenium in the sample and standards is exactly +. Results obtained by using the microprobe are in good agreement with those obtained by wet chemical analysis. II) CRYSTAL GROWTH FROM THE VAPOR Early work on the vapor growth of PbSe was carried out by Prior.7 He used small chips of Bridgman-grown single crystals as the source material and frequently converted the whole charge of a few grams into one crystal. In the present work, vapor growth occurred using a metal-rich or chalcogenide-rich two-phased alloy powder as the source material. Small, nearly stoichiometric crystals are formed on the walls of the quartz tube. The procedure will now be described in detail. Initially, a 100-g charge containing (metal)o.51(chalco-genide)o 49 proportions or (metal)o.49(chalcogenide)o. 51 proportions of the as-received elements in chunk form are placed in a fused silica ampoule. After the ampoule is loaded, it is evacuated with a diffusion pump and sealed. The sealed ampoule is placed in the center of a vertical resistance furnace. The region containing the ampoule is heated to about 50°C above the liquidus temper-ature for the particular composition used. After about one-half hour at temperature, the elements are reacted and the molten material homogenized. The ampoule is quenched in water. The quenched ingot is crushed to a coarse powder for vapor growth experiments and to a fine powder for the isothermal annealing experiments which are discussed in a later section. Vapor growth experiments were carried out using the powdered, metal-rich or chalcogenide-rich alloys

Jan 1, 1969

By C. B. E. Douglas

The following analysis was stimulated by a previous article on evaluation of the water problem at Eureka, Nev., which describes a method using formulas especially devised to calculate flow potential of extensive aquifers characterized by relatively even porosity and permeability throughout. The present discussion submits that the method was unsuitable for solving the kind of problem occurring at Eureka, where the amount of water available, rather than the flow potential, may have been the vital factor. IN an interesting article on evaluation of the water problem at Eureka, Nev.,1 W. T. Stuart describes how a difficult water problem, or one phase of it, may be evaluated by means of a small scale test. Test data are plotted by a method rendering, under certain conditions, a straight-line graph that can be projected to show how much the water table will be lowered by pumping at any specified rate for a given time. A formula is then used to determine the size of opening, or extent of workings, necessary to provide sufficient inflow to enable pumping to be maintained at that rate. At first glance this might seem the answer to a miner's prayer, but a word of caution is in order. It may not be the whole answer. Moreover, results obtained by the method described are reliable only for conditions approximating those assumed. Even where conditions do not meet this requirement, however, it may be possible to draw helpful inferences from the results, perhaps enough to facilitate another approach to evaluation of a problem. The two formulas Mr. Stuart used, the Theis formula and the one developed from it by Cooper and Jacob, were given field checks a number of years ago in valley alluvials by the Water Supply Div. of the U. S. Geological Survey and found to be reliable when the aquifer is very large in horizontal extent and sufficiently isotropic for the test well and observation wells to be in material of the same average permeability as the saturated part of the aquifer as a whole." Extensive valley alluvials, sands, and gravels can be evaluated in this way, and there are even cases in which the method could apply to porous limestones, such as flat beds of very large areal extent that have been submerged below the water table after extensive weathering. These are sometimes prolific sources of water for towns and industries. It is necessary for them to have been above the water table for some geologically long period of time in a fairly humid climate before submergence because the necessary high porosity and permeability, and large reservoir capacity, are the result of weathering, that is, of solution by the carbonic acid (H,CO3) in rainwater formed by the absorption of CO, from the air by raindrops, and this dissolving action must cease when all the H2CO3 has been consumed by re- action with the carbonate to form the more soluble bicarbonate. Consequently this weathering process is largely restricted to a zone that does not extend much below the water level, and submergence is necessary after the weathering to provide large reservoir capacity and good hydraulic continuity. On the other hand, water courses tend to form along faults and fractures in limestone, and to become enlarged by solution, well below water level when, as often happens, fresh meteoric water is circulated rapidly through them to considerable depth by hydrostatic pressure, as through an inverted syphon. Although the reservoir capacity of such water courses is relatively small they may extend far enough to tap more prolific sources. Cavities, and sometimes caves of considerable size, are found in limestones where the acid formed by the oxidation of sulphides has attacked them. This action can take place as deep below water level as surface water is carried by syphonic or artesian circulation, because the oxygen it carries in solution will not be consumed until it reacts with some reducing agent, such as a sulphide. Moreover, the formation of acid and solution of limestone in this way is not confined to the immediate vicinity of the sulphide. Oxidation of pyrite, for example, results in formation of acid in several successive stages, each taking place as more oxygen becomes available, as by the accession of fresh water into the circulation at some place beyond the sulphides. When the acid thus formed attacks the limestone, CO, is liberated and the ultimate effect of the complete oxidation of one unit of pyrite will be the removal of six times its volume of limestone as the sulphate and bicarbonate, both of which are relatively soluble. The reaction may be continued or renewed along a water course far from the site of the sulphides, where the small electric potential produced by contact with the limestone helped to start the reaction. Mr. Stuart refers2 0 caves in the old mining area in the block of Eldorado limestone southwest of the Ruby Hill fault at Eureka, Nev., and to the cavities encountered in drillholes in the downthrown block on the other side of the fault. Although he interprets these cavities as evidence that this formation was sufficiently isotropic (evenly porous and permeable) to give reliable results by the method he describes, they may, in fact, be entirely local conditions. There is reason to think they were probably formed

Jan 1, 1956

By N. J. Grant, B. C. Giessen, Paul Predecki

ALLOYS of gold, silver, and aluminum with elements of the groups BII, BIII, BIV, and BV were prepared by a rapid quenching technique (splat) and were examined by X-ray diffraction. Five new intermediate phases were found and will be described briefly herein. For the gold and silver systems, the concentration ranges having an electron/atom ratio e/a of 1.4 to 1.5 ("3/2 Hume-Rothery phases") were studied primarily. Master alloys were prepared from high-purity metals (99.9+ pct or better) by melting either in evacuated fused silica capsules or by nonconsum-able-electrode arc melting in an argon atmosphere. Small pieces, 20 to 50 mg, of each alloy were blast-atomized to form a splat, by a technique similar to that described by Duwez and Willens.1 The technique used for this study is described in detail in Ref. 2; it utilizes a resistance-heated graphite crucible with a small hole at the bottom, directed toward a metal substrate or quenching plate. The prepared alloy rests over the fine hole, through which it is expelled by an explosion shock wave in the form of fine droplets (1 to 50 µ) of molten metal onto a copper or silver substrate, which is maintained at about -190°C. The resulting very high cooling rates (see Ref. 2 for quantitative measurements) can prevent the process of nuclea-tion and growth in many instances, resulting in the formation of metastable phases. The splat particles were transferred to a GE-XRD5 diffractometer and maintained at -190°C, where they were examined with CuKa radiation. The samples were then allowed to warm to room temperature or were heated to higher temperatures until the equilibrium structures formed. Of fifteen alloy systems considered, nonequi-librium structures were encountered in six; these are described below and summarized in Table I. In the system Au-Sb a metastable £ phase (A3 type, hcp, a = 2.898 + 0.002A; c = 4.731 * 0.004A; c/a = 1.633) was found in the concentration range Au + 13 to 15 at. pct Sb. This phase is isomorphous with the stable phases in the systems Au-Cd, Au-In, and Au-Sn, all at an average e/a ratio of 1.4 to 1.5. The concentration range of one-phase metastable was deduced from the small amounts of supersaturated gold solid-solution phase present in the splat product. It was found that ? could also be retained by splatting onto a substrate held at room temperature: however, decomposed into the equilibrium phases Au + AuSb2 after heating to 200°C for 1/2 hr, or on holding the powdered splatted alloy at 20°C for several months. Calorimetric measurements will be made in an attempt to decide the question whether ? is metastable at all temperatures or whether it is a stable phase at low temperatures. There is evidence that another phase, possibly also close-packed but with a different stacking sequence, can be obtained by rapid quenching of alloys with a different antimony content. Klement, Willens, and Duwez3 reported the existence of an amorphous phase on quenching Au-Si alloys (25 at. pct Si) to - 196°C. They found that on heating to room temperature another phase of unknown crystal structure was formed. This was confirmed (see Table I); however, the new crystalline phase, designated as ?, could also be formed simply by rapid quenching to room temperature, and even was found to exist already in the as-cast Au + 20 at. pct Si alloy. It was found that ? decomposed into Au + Si on the specimen surface at room temperature. This behavior, and the question whether or not there is an equilibrium-temperature region for ?, have not yet been resolved. It is probable that ? (Au + 20 to 21 at. pct Si) is cubic of the -brass type (D81-3) with a = 9.60, + 0.01A and N = 52 atoms per cell [compare 6 (CU-Sn)4]. Except for two very weak lines, the powder pattern of about thirty lines could be indexed on this basis; however, a determination of the atom positions has not yet been attempted. For Au-Ge the C phase was observed at about 21 at. pct Ge as reported by Luo et at.5 Lattice parameters a = 2.876A, c = 4.73,A, c/a = 1.64 were found. In the Au-Pb system, formation of a ? phase was not observed, but in the lead-rich region at 75 at. pct Pb, broad peaks belonging to an amorphous phase were found. The maximum diffracted intensity occurred at 28 = 32.4 deg which is about 1 deg larger than the position of the (111) line of lead (Cuka). For Ag-Pb, an amorphous phase analogous to the one found in the Au-Pb system was observed; this metastable phase exists probably at about 75 at. pct Pb. Since no lead-rich alloys were tested, all alloys consisted of silver + amorphous phase at -190°C. In A1-Ge alloys, line-rich and complex powder patterns were obtained at about 30 at. pct Ge; they bear similarities to those of aluminum and germanium, but are of lower symmetry; the existence of more than one intermediate phase is possible. The authors are grateful to the Kennecott Copper Corp. for Fellowship support, and ARPA (Contract

Jan 1, 1965

National Minerals Advisory Council A meeting of the National Minerals Advisory Council on August 3rd in Washington, D. C., indicated the vitally important part that the mining industry is to play in the mobilization program. Director James Boyd of the Bureau of Mines told the Council that the Department of the Interior would review the recommendations of all the Council's commodity committees with regard for mobilization planning in the light of the changed international picture. The Council was requested to reactivate its commodity committees and have them gather all available data on supplies, their sources and availability and present and potential production of the minerals and metals represented on each committee. Data on labor, machinery, transportation, automotive and stationary equipment, power, fuel, lumber, water supply are a few of the important items called for in the reports, which are to be presented at a meeting of the Council on September 1 at Salt Lake City. The material in the reports will become the basis for discussing metal and mineral requirements at that time. Discussion at the meeting bared several $64 questions, probably the most important of which are the following: 1. Which of the war-essential metals and minerals and in what quantities can we reasonably expect to get them from abroad under threat of submarines? 2. How are we going to meet the manpower problem posed by (a) migration of labor from mining to manufacturing since the end of World War II and (b) the draft and the calling up of reservists? Opinion was expressed by industry spokesman at the meeting that the function of complying with mobilization requirements be left to those in the industry itself; that is, those having the "know how." This view contended that any administrating governmental agency should be kept as small and streamlined as possible. There was general sentiment against the reactivation of the wartime Premium Price Plan or other bonus plans as a stimulus to production. The thought was emphasized that what was needed was a change in the basic conditions which have fostered the decline in domestic mining activity in the postwar years. One such condition, long overdue for correction, is the tax structure as it applies to mining enterprises. Many quarters both in industry and in government favor tax relief along the lines suggested in the six tax recommendations by the Council to the Secretary of the Interior last December. The Council adopted a resolution expressing a feeling that the following tax recommendations are still feasible and desirable and will accomplish as much toward increasing exploration for new deposits (thereby subsequently increasing production) as will government loans for exploration: (1) Losses from unprofitable ventures should be allowed corporations, partnerships, or individuals as ordinary deduction against current income. (2) Development costs after discovery should be recognized as operating expenses. (3) Allowance for depletion should be made to the stockholder as well as to the corporation. (4) Income should not be taxed without full allowance for losses of loss years. (5) Adequate allowances for percentage depletion should be made. A discussion of the manpower problem led to the Council's acceptance of a resolution advising that "military authorities should proceed with caution in depriving the mining and metallurgical industry of its manpower." The resolution strongly urged that no personnel "directly engaged in exploration, development, production or supervision (of strategic and critical materials) should be drafted for the armed forces, at least until the anticipated demands upon these producers are clarified." Stockpiles The Munitions Board's "Stockpile Report to the Congress" of July 23, 1950 revealed: (1) The total estimated value of the stockpile objective is $4,051,714,510 at the close of fiscal year 1950. (2) The total value of the stockpile on hand, at the close of fiscal 1950 was $1,556,154,352 or 38.4 pct of the total stockpile objective. An additional $494,948,060 was on order, making a total of 50.6 pct on hand plus the amount on order. (3) Materials obtained for the stockpile by the ECA from January to June 1950 amounted to $13,112,085, while development projects by ECA during this period involved the expenditure of $9,322,000, mainly with counterpart funds. Shortly after the start of the Korean conflict it was felt that Congress ould appropriate greatly increased sums for the purchase of materials for the stockpile. This stimulus to the program may increase the dollar earnings of those European nations that are present or potential contractors in our stockpiling program. Such a development would mean that these nations could add to their gold reserves, thereby stabilizing their respective economies and hastening recovery. This seems to be the picture for the next six months anyway. The "bug" appears when it is realized that the increased threat of total world war actually may retard recovery in Europe as nations on the continent may feel inclined to devote more of their resources to defense programs. Industries Essential to Defense The Department of Commerce in response to a request by the Department of Defense issued on August 3, 1950 a "Tentative List of Essential Activities" as a "guide for calling up for active duty members of the civilian components of the Armed Forces." The list includes the following: Primary Metal Industries. Included herein are establishments engaged in the smelting and refining of ferrous and nonferrous metals from ore, pig, or scrap. Metal Mining. This category includes establishments primarily engaged in mining, developing mines or exploring for metallic minerals (ores). This group includes all ore dressing and beneficiating operations. Anthracite Mining, Bituminous Coal and Lignite Mining, Crude Petroleum and Natural Gas Extraction, Mining and Quarrying of Nonmetallic Minerals, Except Fuels. Challenge to the Mining Industry The source of our country's great strength lies in its capacity to produce. In times of stress such things as national morale and manpower are all-important but without a capable industrial machine we would be helpless. That machine must be fed with minerals and metals in order to generate and maintain momentum sufficient to insure success. Consequences of the lack of adequate supplies of essential metals and minerals to increase and sustain our industrial power are not pleasant to contemplate. It is absolutely imperative that we put forth Herculean effort to guarantee ample supplies of such essential materials as copper, lead, zinc, manganese, antimony, mercury, tungsten, tin, chromite, nickel, cobalt, iron ore and rubber. The mining industry faces a challenge more serious than ever existed before in the history of our country. The industry must be equal to the task.

Jan 9, 1950

By Herbert Hoover

NO ONE could fail to be gratified to receive so profound an approbation in his calling from the members of one's own profession. To have re-ceived this distinction from men, many of whom have been professional associates and friends over more than a quarter of a century, warrants a pride. and satisfac-tion that could come from no other source. It has often been said that men are really known only to their brothers in labor and in profession. The long friendship I have received from Mr. Saunders and the distinguished service that he has contributed to the mining industry add to my appre-ciation of this event. Mr. Kelley's remarks bring my mind easily to the beginning of that period of great advance in the engineering, profession of forty years ago. Three great forces contributed to the development of the profession. The first was the era of intense development of minerals, metallurgy; and transportation in our great West. It greatly stimulated invention; it made a demand for training and skill on the part of our engineers. The pressure of national development advanced our American practice beyond the rest of the world. Moreover, the skill of our engineers of that period owes a great debt to American educators. The leaders of our universities were the first of all the educators of the world to recognize that upon them rested the responsibility to provide fundamental train-ing in, the application of science to engineering under the broadening influence and cultivation of university life. They were the first to realize that engineering must be transformed into a profession in the highest sense, not only in the training and character but that the essential quality of a profession is the instillation of ethics. That implies not alone the job but the responsibility of the community for the job. Our uni-versities poured into our development a great stream of men with this background and training. A third distinction that grew in American engineering was the transformation from solely a technical profession to a profession of administrators-the business manager with technical training.

Jan 3, 1928

By C. M. Weld

(Cleveland Meeting, October, 1912.) IN the course of my professional work in China during the fall of 1907, I had an opportunity to visit the iron-ore mines at Ta-yeh in Hu-pei province (long. 114° 50' E., lat. 30° 20' N.). It occurs to me that a rescript and discussion of my notes taken at that time should be of interest, particularly as supplementing the exceedingly interesting recent paper-of Thomas T: Read on The Mineral Production and Resources of China.1 Ta-yeh is situated near the right bank of the Yangtze river. The mines are at an average distance of about 16 miles from the river-port Shi-hui-yao, which in turn is some 70 miles S-E. of Hankow. The ores are shipped to the works of the Han-Yeh-P'ing Iron & Coal Co. at Hanyang, as also very largely to the Japanese government iron- and steel-works at Wakamatsu, Japan. Smaller quantities have been exported to the Pacific coast of the United States. These ore-deposits have been briefly noticed in a number of publications, but, so far as I know, Mr. Read's description, in his chapter on the iron-ores of the Chinese Empire,2 contains the first attempt to define their geological character. In his more recent paper Mr. Read follows very closely the text of his earlier description. To quote briefly from the recent paper, he says : " The iron-ores at Ta-yeh . . . . lie along the contact between a marble and an intrusive body of a dark-gray syenitic rock." He refers to a figure which exhibits in plan, but not in section, the relative positions of the syenite, ore-bodies, and limestone. The contact is seen to trend approximately E-W., with the syenite on the north and the limestone on the south. " The syenite lies to the north of the marble, and a few miles farther north the marble again' appears, iron-ore again being present along the contact." No litho- 1 Bulletin Bulletin No. 63, March, 1912, pp. 293 to 343. 2 The Iron Ore Resources of the World', vol. ii., p. 913 (Stockholm 1910).

Oct 1, 1912

A. 11. P. WYNNE, San Jose de Gracia, Sinaloa, Mex. (communication to the Secretary): In the comparative tests reported by Mr. Tays, the stamp-batteries were provided with various styles and mesh-sizes of screens, while the Bryan mill was run throughout with a slot-punched Russia iron, equivalent to a 30mesh wire screen. This is a somewhat onc-sided method of comparison. Should not the Bryan mill also have had the chance to show what it could do with other screens? I cannot wholly approve Mr. Tays's proposed improvements upon the Bryan mill. In my judgment, the inventor has given us a machine well-adapted to the work for which it was designed, namely, the crushing and amalgamation of soft ores. In my judgment, the principal cause of failure to amalgamate satisfactorily in the Bryan mill is not, as Mr. Tays thinks, the shallowness of the basin, or the lack of copper plate about the inside periphery of the mill, but rather the want of experience and practical knowledge on the part of most amalgamators. Having had experience, not with one Bryan mill only, but with several, and taking into consideration the experience of other capable mill-men, I am positive that the proper way to save gold ill this mill is: first, to use enough quicksilver to keep the amalgam liquid, thus allowing it to accumulate in the annular space between the dies and the rims of the basin; secondly, to replace the old dies, when a little more than half worn-out, with new ones; and, thirdly, to clean up the mill at least every two weeks—or oftener, if rich ore is being treated. It is evident that Mr. Tap endeavored to amalgamate quite " hard," as is done in a stamp-battery, in which case the swash of the pulp would scour off portions of the hard amalgam adhering to the inside plates and dash it through the screens— after which a considerable amount would be lost through floating on, or in, the pulp in its journey over the outside plates, and

Jan 1, 1900