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By James T. Wood

Stratigraphy, structure and closure of the Paloma field are discussed, with some details regarding the present state of development. The Paloma anticline, a large dome modified by faults, is the largest single closed dome in the southern San Joaquin valley without surface outcrop or topographic expression. It owes its discovery to geophysics. Of nine wells drilled to the Stevens sand, seven are productive. The ultimate productive area may approximate 3000 acres when the limits of the field are determined. Introduction The Paloma field derives its name from the Paloma school, ½ mile north of what is now the center of the field. The field is 17 miles southwest of Bakers-field, in the south central part of the San Joaquin Valley (Fig. I). The present developed area and its expected extensions are in townships 31 and 32S., R.26E. It is the most southerly of several fields that produce from the Stevens sand zone. The first well to produce condensate was a joint test drilled by the Western Gulf Oil Co. and The Texas Co. near the center of sec. 3, T.32S., R.26E., well No. 54-3. This well was completed on Aug. 31, 1939, at a depth of 10,178 ft. The initial production was 2280 bbl. a day of condensate and 14,750,000 cu. ft. of gas (½-hr. gauge) through a ?-in. bean. Prior to the drilling of this well, the Ohio Oil Co. discovered and developed a small gas field, which embraces a part of three townships about their common corner but lies chiefly in secs. 31 and 32, T.31S., R.26E. The first well was completed on July I, 1934, at a plugged depth of 5239 ft. after having been drilled to 7957 ft, production was found in a 59-ft. zone IIO ft. below the Fourth Mya bed, within the Sari Joaquin clays. Until recently this field was known as the Buena Vista gas field. It is now carried by the Division of Oil and Gas as the Paloma gas field. Production within the Paloma gas field is derived from six wells, producing from several different zones in the interval from 4177 to 5548 feet. Geology Stratigraphy The sediments penetrated by wells in the Paloma held range in age from Recent to Upper Miocene. The entire section, which exceeds 10,000 ft. in thickness, conforms generally to the standard section of the south central San Joaquin Valley. The producing zone, sometimes called the Paloma sand, is the approximate equivalent of the Stevens sand zone of the Ten Section, Canal, Coles Levee and other fields. The stratigraphic section is shown in Table I. Paloma Producing Zone* (Stevens Sand Zone, Upper Miocene) .-The Paloma producing zone (Fig. 2), frequently referred to as the Paloma sand, is encountered at depths ranging from 9950 to 10,300 ft. In general, it is a massive, firm to hard, not too well sorted, fine to medium fine sand. It is interbedded with thin, hard, calcareous sand shells of similar texture and cherty

Jan 1, 1942

By R. R. French

Investigations were conducted by the lndiana Geological Survey for some dry methods of bene-ficiating low-grade gypsum ore. Seventy-two batch and continuous flow tests were performed with a roller mill, rod mill, pebble mill, electronic color sorter, electrostatic separator, and an air separator. Approximately 650 size analyses and 550 chemical analyses were performed during the investigation. Batch samples were tested by the Survey, and most of the continuous flow tests were handled by com-mercial laboratories. INTRODUCTION Although many companies producing gypsum have been interested in dry beneficiation, very little information has been published in the past. One of the most comprehensive investigations was reported by the Canadian Department of Mines.5 Other reports of interest have been published by the Canadian Department of Mines,' the South Australian Department of Mines,2 and the U.S. Bureau of Mines.3 Some beneficiating practices, such as hand picking, selective mining, crushing and screening, milling, and air classifying, were already in use in the Canadian industry prior to MacPherson's investigations.5 MacPherson examined the effect of tabling, air separation, screening, flotation, electrostatic separation, calcining, washing, and various types of milling, in his efforts to eliminate dolomite, limestone, and small amounts of silica and clay from the gypsum ore. RAW MATERIAL Gypsum and anhydrite occur within the lower part of the St. Louis Limestone (Mississippian) in southwestern Indiana. Gypsum in single beds 10 or more ft thick occur at Shoals, Martin County; near Freedom, Owen County; near Bloomfield, Greene County; and in the Devonian strata of LaPorte County. Only the Shoals deposit has been exploited commercially. The western edge of the Shoals deposit is contaminated by a continuous bed of shale, about 1 1/2 to 2 ft thick, near the top of the evaporite and by thin continuous strata or irregular masses of carbonate rock. The low-grade material used in these investigations was obtained from the waste pile of the National Gypsum Co.'s plant at Shoals. The waste ore averaged about 67.4% gypsum and was contaminated by various amounts of shale, dolomite, and limestone. All the material had been previously crushed and screened to minus 1 1/4 to plus 3/8 in. X-ray analyses of powdered, sedimentated, gly-colated, and heat-treated samples of the shale showed that it was composed of slightly structurally disordered illite, Fe-rich chlorite, and very finegrained disseminated silt. The clays and silt were partly cemented with carbonate material. Light-gray argillaceous limestone and gray or brown porous dolomite made up most of the carbonate rock contamination. Crushing Characteristics: The rate unweathered gypsum, carbonate rock, and shale reduce in size was determined by crushing handpicked samples in a rod mill and screening at minus 100 mesh (.0059 in.). Fig. 1 graphically represents the data. These data indicate that carbonate rock should be relatively easy to separate from gypsum, but that shale should be difficult to separate from gypsum by differential crushing. Controlled Samples: Eight rod mill tests of hand-picked samples of mixed gypsum and shale and mixed gypsum and carbonate rock were made. Sample weight and crushing weight were standardized, but the scalping screen size was varied from 3 to 10 mesh (.265 to .0787 in.) in order to obtain the optimum purity and recovery. The data obtained from the tests (see Table I) illustrate that the relatively hard carbonate rock allows appreciable beneficiation (13.6 to 18%), but the low strength gypsum and shale allows only minimal separation (3.1 to 7.5%).* Semi control led Samples: During the investigation considerable slaking of the shale was noted when the ore was subjected to alternate wetting and drying. As demonstrated in laboratory tests, as much as 76% shale can be removed by two cycles of wetting and drying in a 48-hour period with subsequent screening at .265 in. Five ore samples, stockpiled and weathered, were screened, rod-mill crushed and rescreened at .157 or .187 in. with excellent results (Table 11).

Jan 1, 1967

By John M. Sharp

John M. Sharp, Jr. (Assistant Professor of Geology, University of Missouri, Columbia, Mo.)-Geothermal gradient and surface temperature data anomalies are potentially useful tools in mineral exploration, These data, however, may reflect a number of geologic and hydrologic processes as well as the thermophysical properties of the geologic materials. The author presents a relatively simple model which may be of value in mineral exploration. The lack of consideration of other processes, however, limits its usefulness. For instance, because of erosion or sediment deposition over geologic time the land surface must be considered a moving boundary in such a model. Evaluation of this effect coupled with annual surface temperature and thermal conductivity variations can result in temperature profiles such as shown in [Figs. 4 to 7]. These effects and others, such as ground-water movement," must be considered and filtered from the data before proper interpretation is possible. Another assumption which should have been discussed is, why were the observed mean surface temperatures not used for extrapolation of the gradients in [Figs. 4, 6, and 7?] Conductive heat transport responds to the actual boundary (land surface) temperature, not to theoretical temperatures in the absence of perturbations. Benfield presents a theoretical model for calculating perturbations in the geothermal gradient for erosion and sedimentation. The temperature (T) at any time (t) and any depth below the surface (x) is given by [ ] where T is the mean surface temperature, G, the geothermal gradient, v, the rate of erosion (or sedimentation),[ ], the long-term rate of change of T0, and k is the thermal diffusivity of the rock. Using the author's values and a rate of erosion of 1.54 cm per yr for a period of 10,000 yr, a thermal profile as shown in [Fig. 10] will be generated. A thermal diffusivity of 0.0164 sq cm per sec was used, corresponding to a thermal conductivity of 0.0082 cal per cm per sec per C. In [Fig. 11], a sediment deposition rate of 3.0 cm per yr for a period of 1000 yr was used to generate the subsurface temperature profile. A surface temperature of 24.3°C was used, as specified in [Fig. 9, 7], was set equal to zero for both cases. Superposition on the generated profiles in [Figs. 10 and 11] of the annual surface temperature variation perturbation (which may be calculated in a number of ways) will yield profiles similar to those in [Figs. 4 and 7]. Similar profiles can also be produced by regional ground-water flow. Models like the author's can be of value in mineral exploration. Other processes, however, which affect surface temperatures and gradients must first be considered, calculated, and filtered from the data if anything but the crudest qualitative interpretation is to be made. Robert C. Edmiston (Author's reply)-Professor Sharp has raised three objections to my interpretation of the subsurface temperature data I collected at Silver Bell. These are: 1) No correction was made for erosion or sedimentation,

Jan 1, 1977

By L. Adler

Analysis of a mine roof can be based on fixed-end beam behavior. The author here analyzes the effects of zero restraint at deflecting beam supports. Formulae are given for determining permissible support deflections. Analysis of a mine roof in a bedded deposit usually follows the assumption that the behavior of the roof is analogous to a beam with both ends fixed. Yet it is recognized that the end restraint itself varies with the thickness of cover, so that at shallow depth the restraint is minimum, and at great depth, maximum.' Consequently, an analysis on the basis of fixed ends or full restraint represents only one extreme in the possible range of conditions. Since the actual amount of restraint is difficult to ascertain, a more rational approach is to consider the two extreme cases of end restraint. Such situations regarding uncertainty of the restraining effects of supports are not uncommon in structural analysis, and are handled in a like manner.' Since previous studies9)4 have dealt with full restraint, this paper will analyze the effects of zero restraint at deflecting beam supports. In such a case a mine roof with several intermediate supports becomes analogous to a uniformly loaded continuous beam on simple supports, as shown in Fig. 1A. Such a structure is statically indeterminate to a degree equal to the number of intermediate supports, n. Therefore, with the three equilibrium equations, n additional equations are needed. A number of methods exist to analyze this problem, but The Theorem of Three Moments is Simplest. The continuous beam is divided at each support into a sequence of simply supported single span beams, with the unknown end moments, as appears in Fig. 1B. Continuity requires that the common ends of adjacent beams have the same slope (Fig. 1C). eiB= B(i+l)~= ei(i+l)and MiB= Mti+I)A= Mi(i+i). [1I A Glossary of all terms is given in Table I. Furthermore, if we first consider that the supports do not deflect, the slope at a support is the sum of the slopes due to each of the individual loads shown in Fig. 2, yielding:

Jan 1, 1961

By J. D. Allen

In the coal preparation plant at U.S. Steel Corp.'s Concord mine in Alabama, the product moisture was reduced from about 20% to 9% through the development of a workable flowsheet for dying the clean coal. Some of the high points of the conversion of the Concord washer from vibrating screen dewatering to centrifuge drying are presented in this article. The coal preparation plant at the United States Steel Corp.'s Concord mine near Bessemer, Ala., began operations in June, 1950. The basic cleaning circuit consisted of a Chance cone for cleaning the 2-in. x 3/8-in. fraction and concentrating tables for cleaning the -3/8-in. coal fraction. Dewatering of the final product was accomplished with 35-mesh vibrating screens. However, because Concord coal is very friable with 85% passing 3 mesh and 25% passing 28 mesh (Fig. I), a considerable quantity of fines passed through the dewatering screens. To recover these fines, the dewatering screen effluent was sent to two 70-ft-diam hyd rose para tors from which the underflow returned to the dewatering screens and the overflow went to a 175-ft-diam thickener for water clarification, as illustrated in Fig. 2. As might be expected, there was considerable recirculation of the -35-mesh coal in this system. This degraded the coal and led to higher moistures. It also resulted in a fines bleed-off of plant water and thickener underflow to a fine-coal settling pond. In 1961, vacuum filters were installed to supplement the screens and to recover the fine coal lost in the bleed-off. The above system resulted in a product moisture of about 20%. Further reduction to approximately 11% was accomplished by drainage from railroad cars during transit to the coke plant. Control of the moisture at unloading was poor because of variations in operating levels and railroad car availability, but since the Concord coal constituted a relatively small percentage of the over-all coal mix at the coke plant, its high moisture and fluctuations did not seriously hamper coking operations. Early in 1964, however, the proportion of Concord coal was increased with a resulting increase in the moisture of the total coal mix, and this caused difficulties in the coke plant. The increased participation of Concord coal also brought about a shortage of railroad cars which raised the moisture content to a point where the coal became almost impossible to handle. In an effort to alleviate this problem, U.S. Steel initiated a research program to develop a flowsheet for reducing the moisture to approximately 10% and to determine whether the resulting product would be satisfactory for the coke plant. Under this program, an experimental flowsheet for evaluating both a vibratory centrifuge system and a solid bowl centrifuge system was drawn up as shown in Fig. 3. A horizontal vibratory centrifuge was installed to reduce the moisture in the dewatered coarse coal, and a four-cell, flotation machine was installed to clean the -65-mesh portion of the 3/8-in. x 0 table raw coal feed. Several shortcomings in the above circuits were apparent. For one thing, large quantities of repulping water were required. Furthermore, since about half of the flotation feed was of a size consist that could have been cleaned on the tables, these circuits would result in an unnecessary increase in reagent costs and facilities. Maintenance of the hydro separators and pumps for pumping the underflow to the filters,

Jan 1, 1969

THE following employment items are made available to AIME on a non-profit basis by the Engineering Societies Personnel Service, Inc., operating in cooperation with the Four Founder Societies. Local offices of the Personnel Service are at 8 W. 40th St., New York 18; 100 Farnswarth Ave., Detroit; 57 Past St., San-Francisco; 84 E. Randalph St., Chicago 1. Applicants should address all mail to the proper key numbers in care of the New York Office and include 6c in stamps far forwarding and returning application. The applicant agrees, if placed in a position by means of the Service, to pay the placement fee listed by the Service. AIME members may secure a weekly bulletin of positions available far $3.50 a quarter, $12 a year. MEN AVAILABLE Graduate Mining Engineer, 38; married, two children, Dutch nationality. Experience bauxite open-pit dragline, stripping and mining, nickel silicateous ore exploration, tin lode mining and dredging, ore dressing research, administrative. work, all in responsible positions. Read, write and/or speak seven languages. Desires responsible position. Presently located Netherlands. M-655. Mining Geologist, 35, married, 3 children, Canadian citizen. Eight years' experience Canadian gold mines as geologist, chief engineer, shift boss, 3 years in western U. S. as surveyor, junior geologist, etc. Desires position of responsibility. Prefer western United States. M-656-478-E-3-San Francisco. Mining Engineer, 30 years' experience open-pit and underground coal and iron mines. Held positions engineer to manager. Desire position of responsibility in foreign field. M-659.

Jan 1, 1952

By Nathaniel Arbiter

T. M. Morris (School of Mines and Metallurgy, Rolla, Mo.)—Rate studies promise to help quantify flotation operations. The author's exposition of rate studies is therefore laudable. However his exposition of a second order rate equation rather than a first order rate equation is not convincing. It is important to use the correct equation, otherwise erroneous conclusions may be drawn. The data assembled by the author and plotted according to a second order rate equation can be plotted according to a first order rate equation. As a matter of fact much of the data used reveals that a second order rate equation is not applicable. According to the method used by the author for plotting data, the reciprocal of the slope of the line is the maximum possible recovery. Physically this cannot exceed 100 pct, yet several graphs yield maximum recoveries greater than 100 pct. For example: l—Curve 2, Fig. 1 yields a value of 120 pct; 2—Curve 2, Fig. 2 yields a value of 115 pct; 3—Fig. 7 yields a value of 110 pct. The real test of the graphical method is, as the author points out, whether or not the earliest points agree with a straight line. The second order rate plot does not fulfill this requirement. Fig. 10 shows the data used by the author to obtain Curve 2 in Fig. 1, plotted according to both a second and first order rate equation. Fig. 11 shows the two rate plots made from data published by Ludt and DeWitt.19 In Fig. 10 the first order rate plot gives a very good fit and obeys the boundary condition of passing through the origin. For the second order rate plot a straight line can be drawn through the last three points but the earliest points are not in agreement with this line. The reciprocal of the slope of the straight line gives a value of maximum possible recovery of 112 pct. The maximum possible recovery according to the first order rate plot is 78 pct. In Fig. 11 the first order rate plot again fits all of the points very well and passes through the origin. The second order rate plot shows wide deviation of the earliest point, indicating that the second order rate equation is not applicable. The reason for the fit of data to the second order rate equation during longer flotation times is that R, the cumulative percent recovery, changes slowly toward the end of the flotation operation, so that when t/R is plotted against time, one is, in effect plotting t against t and one would certainly expect a straight line. The author attempts to show that the second order rate equation is correct by citing evidence which shows that recovery is a function of initial concentration. This is a debatable point especially since, in the author's original presentation of this paper at the St. Louis

Jan 1, 1952

By C. B. Smith, A. G. Guy

The power and convenience of tensor analysis are employed in deriving the equations that describe three-dimensional diffusion in an n-component system plus vacancies. A "Kirkendall coordinate system, '' defined in principle by Kirkendall markers, is used in defining the intrinsic diffusion coefficients that are of primary interest. The defining equations are then transformed to local cartesian coordinates and to fixed cartesian coordinates, the latter being convenient for experimental measurements. Finally, it is shown how the intrinsic diffusion coefficients can be calculated from the experimentally determined quantities. It is generally recognized that any treatment of the essential diffusion phenomena in metallic systems—the jumping of atoms from one lattice position to another—must be based on movement of the atoms past Kirkendall interfaces.' A logical extension of this idea is the proposal that, fundamentally, diffusion must be described mathematically in terms of "Kirkendall coordinates." The basic equations thus obtained may then be transformed into other coordinate systems, when such transformation leads to simplification of either the experimental or mathematical treatment of diffusion data. Tensor analysis is clearly the appropriate mathematical tool for this purpose, since it provides methods both for determining general equations that are valid for all coordinate systems and for conveniently transforming the equations from one coordinate system to another system related to the first in some known fashion. For the present purpose three coordinate systems are used, and it is convenient to begin with a description of the relations among these three systems. FICK LAW FOR KIRKENDALL COORDINATES The reference coordinate system, and the one that is useful for most experimental measurements, is a cartesian coordinate system, xi (i = 1, 2, 3), fixed with respect to some point in the specimen, such as an end of the alloy bar undergoing diffusion. A second set of coordinates, the Kirkendall coordinate system, is one that in principle can be described by the location of inert markers (Kirkendall interfaces). A convenient mathematical relation between Kirkendall coordinates, x'i, and the reference cartesian coordinates, ii, is obtained by specifying that the Kirkendall coordinate system coincide with the cartesian system at time, t = 0. After diffusion has occurred for some time, t = t, the Kirkendall coordinate system will differ from the reference system in the manner shown schematically for the one-dimensional case in Fig. 1. The third coordinate system is Darken's1 local cartesian-coordinate system, xi, which is identical with ii in orientation and units, but which moves with the point under consideration in the Kirkendall coordinate system. Even though most diffusion problems are essentially one-dimensional, there is an advantage in first carrying out the analysis of the three-dimensional problem to gain a clearer insight into the assumptions that are made in the usual one-dimen-

Jan 1, 1962

By J. H. Frye, J. W. Caum, R. M. Treco

IT has been suggested that: "Insofar as the hardening due to a solute depends upon the increase of lattice parameter produced by it, it is reasonable to suppose that this hardening might be related to the resistance which the pure solvent would offer to increase in its parameter."' It was further suggested that this hypothesis might be examined by comparing the hardnesses of analogous silver and copper alloys, since the stress necessary to produce a given lattice expansion is quite different in the two metals. Suitable data2,3,4 are now available on these alloys, and we shall proceed to an examination of the hypothesis quoted above. Hardness, Lattice Parameter and Ionic Overlap Increase in ultimate Meyer hardness has been plotted against increase in lattice parameter in Figs. I and 2. It is believed that most of the values are accurate within the limits indicated by the rectangles. These data are also presented in Table I. Ultimate Meyer hardness, P u, is defined by the equation: where L = load, D = diametei of an impression having the same diameter as the testing ball, which in this case is 4 mm. Increase in lattice parameter is defined as the difference between the lattice parameter of the solid solution and that of the pure solvent. The curves in Figs. I and 2 are based on equal solute concentration and an equal degree of plastic deformation pro duced by the testing ball. Hardnesses are corrected to a constant grain size for copper and silver alloys, respectively, and these two grain sizes are approximately equal. Furthermore, all solutes are from the same period of the B subgroup of the periodic table as their solvent. The question arises as to whether Or not lattice parameter is the only factor controlling the hardness of these alloys. There

Jan 1, 1943

By J. H. Frye, R. M. Treco, J. W. Caum

IT has been suggested that: "Insofar as the hardening due to a solute depends upon the increase of lattice parameter produced by it, it is reasonable to suppose that this hardening might be related to the resistance which the pure solvent would offer to increase in its parameter."' It was further suggested that this hypothesis might be examined by comparing the hardnesses of analogous silver and copper alloys, since the stress necessary to produce a given lattice expansion is quite different in the two metals. Suitable data2,3,4 are now available on these alloys, and we shall proceed to an examination of the hypothesis quoted above. Hardness, Lattice Parameter and Ionic Overlap Increase in ultimate Meyer hardness has been plotted against increase in lattice parameter in Figs. I and 2. It is believed that most of the values are accurate within the limits indicated by the rectangles. These data are also presented in Table I. Ultimate Meyer hardness, P u, is defined by the equation: where L = load, D = diametei of an impression having the same diameter as the testing ball, which in this case is 4 mm. Increase in lattice parameter is defined as the difference between the lattice parameter of the solid solution and that of the pure solvent. The curves in Figs. I and 2 are based on equal solute concentration and an equal degree of plastic deformation pro duced by the testing ball. Hardnesses are corrected to a constant grain size for copper and silver alloys, respectively, and these two grain sizes are approximately equal. Furthermore, all solutes are from the same period of the B subgroup of the periodic table as their solvent. The question arises as to whether Or not lattice parameter is the only factor controlling the hardness of these alloys. There

Jan 1, 1943

By Robert Hunt

The American Institute of Mining and Metallurgical Engineers was the first American technical organization to consider steel-rail specifications and sections. If I am not mistaken, the first contribution on those subjects to the Proceedings of the Institute was presented by Prof. Thomas Egleston, in 1874, and was entitled "Investigations on Iron and Steel Rails made in Europe in the Year 1873.1" The next contribution was Dr. C. B. Dudley's report of his investigation for the Pennsylvania Railroad of the wear of steel rails in their tracks, in 1878, which contribution led to a very memorable discussion that extended through several meetings.2 In 1880, Mr. C. P. Sandberg3 presented an article on rail specifications and inspection and rail sections. This was followed, in 1881, by A. L. Holley,4 in his exhaustive report on the then used American rail sections. The next contribution, I believe, was made by myself' at the Buffalo meeting, in 1888, on steel-rail specifications, which I followed,6 in 1889, by a paper on steel-rail sections. Since that time there have been numerous contributions to the Transactions of the Institute in relation to the manufacture of steel rails. The American Society of Civil Engineers officially took up the investigation of the relation of rail sections to the wear of car wheels in 1888. In the prosecution of this work, the Society appointed a special committee, whose labors extended through several years and resulted in an exhaustive report to the Society in 1893.7 Later, another committee was appointed to consider specifications for the manufacture of steel rails; the labors of this committee extended over a period of several years, a final report8 being made in 1910. It was my fortune to be the secretary of both of those committees.

Jan 9, 1919

By J. T. Norton, H. Blumenthal, S. J. Sindeband

The interstitial phases formed by the transition elements with carbon, nitrogen and boron constitute a unique class of substances which are of considerable technical interest because of their well developed metallic properties and their high hardness and melting points. They are also of interest from the point of view of structure. It was pointed out by Hagg1 some years ago that the principal factor determining these structures was the relative sizes of metal and metalloid atoms. He predicted that if the ratio of the radii of metalloid to metal atoms was less than 0.59, the metal atoms would be arranged in a close-packed fashion of relatively simple type. This prediction has been amply confirmed. It appears, however, that this specification applies most generally to those structures in which the metalloid atoms occupy isolated positions in the lattice and there is no tendency for strong binding between the metalloid atoms themselves. The series of borides of the type MeB2 formed from the transition elements of the fourth (Ti, Zr) and fifth (V, Cb, Ta) groups of the periodic table provide an opportunity to investigate this situation further. If one takes as a basis the radius of 0.87 A for the boron radius and the Goldschmidt values for the metal atoms in 12 coordination, then the radius ratio varies from 0.54 for zirconium to 0.64 for vanadium, which latter is considerably greater than Hagg's limit of 0.59. It was of interest, therefore, to see if these five elements would form borides of the MeB2 type and have simple isomorphous structures. When this investigation was initiated, relatively little information had been published on the structure of this group of borides. McKenna2 reported the preparation of ZrB2 by a carbon reduction process. Its structure was hexagonal and the approximate lattice constants were given as a = 3.15 A, c = 3.53 A. Ehrlich3 reported the structure of TiB2 as being of the C 32 type with a = 3.02, c = 3.21 and c/a = 1.06 A. Recently, Kiessling4 published the results on ZrB2, showing it to be one in which the metal atoms are arranged in a simple hexagonal lattice with the boron atoms in flat sheets, midway between the layers of metal atoms. The structure is C 32 type with a = 3.169, c = 3.530 and an axial ratio =1.11 A. He also states that CbB2 and TaB2 are isomorphous with ZrB2. Preparation of Borides The borides of titanium, zirconium, columbium and tantalum were made by the electrolysis of fused salt baths after the method described by Andrieux.5 The bath, totalling about 6500 g in each case had the following general composition: Metal Oxide + B2O3 + CaO + CaF2. The actual proportions of the ingredients were determined for each individual case. Although Andrieux reports the preparation of vanadium boride by the same electrolytic method, several attempts proved unsuccessful. However, it was possible to prepare VB2 by the carbon reduction method used by McKenna to produce ZrB2. The mixture consisted of V2O5 + 6B2O3 + 11C and the reduction was carried out in an induction heated graphite crucible. This method proved satisfactory for ZrB2 and TiB2 and undoubtedly the others could be made in the same way. The borides, which were all in the form of fine gray metallic crystals from the electrolytic bath or gray powders from the reduction process, were carefully separated from other constituents and analyzed chemically for metal and boron content. The densities of the boride powders were measured using a standard 10 cc pyenometer at 18°C and ethyl benzene proved to be the most satisfactory liquid for this purpose. X Ray Technique The boride crystals, as prepared, were too small for single crystal examination so that powder techniques were employed. Patterns of the finely ground borides were prepared, using the Norelco recording X ray spectrometer. The lines on these patterns were indexed with the aid of Hull-Davy charts and the approximate dimensions

Jan 1, 1950

By E. A. Elevatorski

Wollastonite, named after William H. Wollaston, an English chemist, is a calcium metasilicate, CaSiO3. It has a short history as an industrial mineral. The earliest production of wollastonite is reported to be from a deposit near Code Siding, located north of Randsburg, Calif. At this locality small tonnages of wollastonite were quarried during 1933-34 and 1938-41, and processed into mineral wool. This operation was largely experimental and virtually no United States production was again reported until the 1950s when a large deposit near Willsboro, N.Y., was developed by the Cabot Corp. A processing plant was placed on-stream in 1953, with nearly continuous production to date, and it is currently operated by Interpace Corp. Since 1958, wollastonite deposits in the Little and Big Maria Mountains of Riverside County, and in the Panamint Range of Inyo County, both in California, have operated intermittently for production of both ornamental and commercial wollastonite. During 1973, the United States was the major producing country, furnishing about 75% of the world's output. Recent production also has come from Finland, Mexico, and Kenya. Small amounts have been shipped intermittently from India, USSR, New Zealand, Republic of the Sudan, Republic of South Africa, and South-West Africa. The principal use of wollastonite is in the manufacture of ceramics, especially wall and floor tiles. Other uses are for paints, fillers, adhesives, reinforcing agents, plastics, fluxes, and glazes. Mineralogy Pure wollastonite, CaSiO3, has the composi¬tion of 48.3 % CaO and 51.7 % SiO,. However, it is seldom found in the pure state due to the ease with which it takes into solution the metasilicates of manganese, magnesium, iron, and strontium. Predominantly, wollastonite occurs as a contact metamorphic deposit forming between limestones and igneous rocks. Commonly associated minerals are garnet, diopside, epidote, calcite, and quartz. It has a specific gravity of 2.8 to 3.0, and hardness of 4.5 to 5 on Mobs' scale. When pure, it has a brilliant white color, but with impurities it may be grayish or brownish. Luster is vitreous to pearly. Melting point of wollastonite is about 1540° C. Wollastonite occurs in coarse-bladed masses, rarely showing good crystal form. It is usually acicular or fibrous, even in the smallest of particles. The most unique property of crushed and ground wollastonite is its cleavage. Fragments of crushed wollastonite tend to be needle-shaped, imparting a high strength, and this property is the basis for many of its uses. The fiber lengths are commonly in the ratio of 7 or 8 to 1, length to diameter. Some crystals of wollastonite fluoresce under short-wave or longwave ultraviolet light, or both; colors ranging from yellow-orange to pink-orange. Specimens may also show phosphorescence. The brightness of wollastonite is a property of considerable importance to the paint industry. Material, 99% pure, with a size of -325 mesh, has a General Electric reflectance rating of 92 to 96%. Chemically, wollastonite is inert and this property makes it useful as a filler and reinforcing agent. There are two polymorphs of calcium silicates: wollastonite, a low temperature form, and pseudowollastonite, a high temperature form. Inversion of wollastonite to pseudowollastonite occurs at about 1120° C, resulting in an increase in the coefficient of expansion and a color change. Pure white wollastonite, on inversion, may change to a cream tint, or various shades of red or brown. The color change is thought to be due to the presence of iron and strontium.

Jan 1, 1975

By John G. Simmons

A tutorial account of the tunnel effect between metal electrodes separated by a thin insulating film is presented. Energy diagrams of metal-insulator -metal sandwiches are briefly discussed, and the isothermal and thermal J-V characteristics for symmetric and asymmetric tunnel junctions are derived from a generalized expression. Experimental data from tunnel junctions are presented, and it is shown how, correlation between theory and experiment permits a determination of some of the oxide properties. Finally, the barrier heights in AT-Al2O3-A1 junctions determined by various inr3estigators are discussed. The object of this paper is to give a tutorial account of the electric-tunnel effect between metal electrodes separated by a thin insulating film and to illustrate its utility in the study of metal-oxide interfaces. This paper will emphasize the theoretical aspects of the electric-tunnel effect: however, a brief account of the more recent experimental work will be presented. A more comprehensive experimental study of the tunnel effect will be presented in the following paper by Drs. Pollack and Morris, see p. 497. We will open with a discussion of the energy diagrams for metal-insulator-metal junctions and what we mean by the electric-tunnel effect. Following this is an account of the theoretical studies of Sommerfeld and Bethe, and Holm, and the recently reported calculations on symmetric and asymmetric junctions using the generalized formula. The theoretical presentation is concluded with calculations on the temperature dependence of the tunnel J-V characteristic. In the experimental section, it is shown how barrier heights and thicknesses are determined by comparing the experimental data with the theory. Finally, the results of various investigators on the determination of the barrier heights in Al-Al2O,-A1 systems are discussed. I) POTENTIAL BARRIERS BETWEEN CLOSELY SPACED ELECTRODES Since it is necessary to supply energy to a metal in order to observe electron emission from its surface, it is evident that the potential energy of an electron within a metal is lower than that of an electron at rest outside the metal. It follows, then, that the electron must undergo a change in potential as it crosses the metal-vacuum interface. In the Sommerfeld1 free-electron theory of metals it is assumed that the free electrons, the electrons responsible for the electrical characteristics of metals, move in a constant potential environment in the metal, and are contained within the metal by a potential barrier at the metal-vacuum interface. It will be evident that the idea of a constant potential environment is a very idealized model, since the potential energy of an electron is influenced by all of the atoms and the other electrons within the metal; however, its use is justified by the many excellent results it gives.2 Fig. l (a) is the energy diagram of the model envisaged by Sommerfeld; the metal is represented by a potential-energy well of depth V,; the vacuum level represents the energy of an electron at rest outside the metal. At 0°K the electrons, which exist in pairs in discrete energy levels, fill up the well to an energy ?, the Fermi level. The distance, Vo - ?, from the Fermi level to the vacuum level is known as the work function of the metal, and represents the minimum energy required to free electrons from the interior of the metal at 0°K.

Jan 1, 1965

By J. H. Frye, R. M. Treco, J. W. Caum

IT has been suggested that: "Insofar as the hardening due to a solute depends upon the increase of lattice parameter produced by it, it is reasonable to suppose that this hardening might be related to the resistance which the pure solvent would offer to increase in its parameter."1 It was further suggested that this hypothesis might be examined by comparing the hardnesses of analogous silver and copper alloys, since the stress necessary to produce a given lattice expansion is quite different in the two metals. Suitable data 2,3,4 are now available on these alloys, and we shall proceed to an examination of the hypothesis quoted above. HARDNESS, LATTICE PARAMETER AND IONIC OVERLAP Increase in ultimate Meyer hardness has been plotted against increase in lattice parameter in Figs. I and 2, It is believed that most of the values are accurate within the limits indicated by the rectangles, These data are also presented in Table I. Ultimate Meyer hardness, Pu, is defined by the equation: [P" - D2] where L = load, D = diameter of an impression having the same diameter as the testing ball, which in this case is 4 mm. Increase in lattice parameter is defined as the difference between the lattice parameter of the solid solution and that of the pure solvent. The curves in Figs. I and 2 are based on equal solute concentration and an equal degree of plastic deformation pro¬duced by the testing ball. Hardnesses are corrected to a constant grain size for copper and silver alloys, respectively, and these two grain sizes are approximately equal. Furthermore, all solutes are from the same period of the B subgroup of the periodic table as their solvent. [ ] The question arises as to whether or not lattice parameter is the only factor controlling the hardness of these alloys. There

Jan 1, 1943

De la pirotechnia. Libri .X. dove ampiamente si tratta non solo di ogni sorte & diuersita di miniere, ma anchora panto si ricerca intorno á la prattica di quelle cose di quel the si appartiene á l'arte de la fusion ouer gitto de metalli come d'ogni altra cosa simile I questa. Composti per it S. Vanoccio Biringuccio]

Jan 1, 1942

By Clarence F. Zeuch

Mr. Shaw's work showed the dodecyl mercaptan as a very good collector compared with other collectors on Cu-Mo ores, Mo ores, and for Au and Au-Ag ores plus Pt ores. However, as one who has had extensive experience on Cu-Mo ores, the work is not complete in making only a rougher concentrate with a collector; the work must be carried on in making a Cu-Mo separation. Did Mr. Shaw do any of this work? Before trying the dodecyl mercaptan as a Cu-Mo collector for the Cu-Mo ore it would be well to have any new work conducted knowing how the reagent acts on the Mo. Will it allow a good separation? The use of the collector at Palabora was also mentioned. From some earlier literature I thought that the reagent has been used as an oxide Cu collector after the sulfide mineral was floated out. I thought the pH at this point was about 1.5-3.0. I would like to have this point clarified.

Jan 1, 1982

[Appalachian Petroleum Section: James M. Kimbrough, chairman; John H. Adair, first vice-chairman; Walter L. Diddle, second vice-chairman; Byron M. Finch, secretary-treasurer. Balcones Section: Robert V. West, chairman; Frank W. Jessen, first vice-chairman; Zane V. Morgan, second vice-chairman; Brantley G. Myers, secretary-treasurer. Meets 4th Wed., Petroleum Club (San Antonio) or Terrace Dining Room (Austin) Billings Petroleum Section: Charles B. Evans, chairman; L. M. Lyon, first vice-chairman; Gordon E. Lansford, second vice-chairman; Fred. A. Haddenhorst, secretary-treasurer. Meets 3rd Tues., Billings Petroleum Club. California Coastal Section: Robert L. Richardson, chairman; David K. Hayward, first vice-chairman; W. Jackson Bransford, secretary; William N. Harwell, treasurer. Meets 3rd Thurs., Pierport Inn, Ventura. Caracas Petroleum Section: Edwin A. Moo re, chairman;. Jose Cirigliano, vice-chairman; Frank B. Davis, secretary; Robert N. Powers, treasurer. Meets 2nd Tues., Caracas Theater Club.]

Jan 1, 1961

By Frank A. Herald

Through the generous cooperation of the authors of the papers in this chapter, a great wealth of fundamental data is being made available. The same plan of securing and submitting data will be carried forward next year, with the hope of making a nearer approach to complete presentation of the data contemplated in the plan. In order to conserve space, we are printing below such footnotes as have general application to Table 1 in the papers in this chapter: Footnotes to Column Headings Table I In areas where both oil and gas are produced, unless gas is marketed outside the field, such areas are included in column headed "Oil." Manufacture of casinghead gasoline and carbon black is interpreted as outside marketing of gas. Production per acre is determined by dividing into the number of barrels of oil the sum of the number of acres assigned to "Oil" plus such number of acres of the total assigned to "Oil and gas" as represents the portion thereof occupied by oil. Wells producing both oil and gas are classified as "Poducing oil only" unless gas from them is marketed off the lease. W. water; G. gas; A, air. Bottom-hole pressures are preceded by "e." All other figures represent pressures at casinghead with well closed. P, paraffin; A, asphalt; M, mixed. 0 Cam, Cambrian; Ord, Ordovician; Sil, Silurian; Dev, Devonian; Mis, Mississippian; MisL, Lower Mississippian; MisU, Upper Mississippian: Pen. Pennsylvanian; Per. Permian; Tri, Triassic; Jur. Jurassic; CreL, Lower Cretaceous; CreU. Upper Cretaceous; Eoc, Eocene; Olig, Oligocene; Mio. Miocene; Pli. Pliocene. 6, sandstone: SH, sandstone, shaly; Ss, soft sand; H, shale; L, limestone; LS, limestone. sandy; C, chalk; A, anhydrite; D, dolomite; Da. arkosic dolomite; GW, granite wash; P, serpentine. Figures are entered only for fields where the reservoir rock is of pore type. Figures represent, ratio of pore space to total volume of net reservoir rock expressed in per cent. "Par" indicates that the reservoir rock is of pore type but said ratio is not known by the author. "Cav" indicates that the reservoir rock is of cavernous type; "Fis," fissure type. A, anticline; AF, anticline with faulting as important feature; Af, anticline with faulting as minor feature; AM, accumulation due to both anticlinal and monoclinal structure; H, strata are horizontal or near horizontal; MF, monocline-fault; MU, monocline-unconformity; ML, monocline-lens; MC, monocline with accumulation due to change in character of stratum; MI, monocline with accumulation against igneous barrier; MUP, monocline with accumulation due to sealing at outcrop by asphalt; D, dome; Ds, salt dome; T, terrace; TF, terrace with faulting as important feature; N, nose; 9, syncline. Information will be found in text as indicated by symbols: A, name of author, other than above. who has compiled the data on the particular field; C, chemical treatment of wells; G, gas-oil ratios; P, proration; U, unit operation; R, references; W, water; 0. other information.

Jan 1, 1934

By R. K. Iyengar, G. C. Duderstadt

An investigation was carried out to determine the effect of Ca-Si additions on the dissolved oxygen content of liquid steel. An apparent equilibrium was reached after holding the melt for some time when the total oxygen content of the melt was identical with the dissolved oxygen. Results of the investigation show that for deoxidation with silicon and manganese the apparent equilibrium is reached after 8 to 12 min and the oxygen content of the melt is in good agreement with values reported in the literature for similar steel compositions. Ca-Si additions decrease the dissolved oxygen content appreciably below that obtained with Si-Mn deoxidation. It is postulated that some calcium dissolved in liquid steel at the time of its vaporization combines with dissolved oxygen to form CaO which then fluxes the manganese silicate, thereby lowering the activity of the deoxidation products. Slow flotation of calcium silicate inclusions is attributed to their slower growth rate. An addition of aluminum (to yield <0.005 pct Alsol) prior to introduction of Ca-Si improves the kinetics of removal of resulting deoxidation products. WITH the introduction of continuous casting of billets and blooms, application of calcium-containing alloys for deoxidation purposes has gained new interest. More so than in conventional ingot casting, the degree of control of the dissolved and total oxygen contents in steel for continuous casting can determine the success or failure of the operation because these affect both surface (pinholes) and internal quality of the billets. Of the available deoxidizers, only silicon has so far found wide application. However, in the typical range of application (0.20/0.30 pct Si), it is too weak a deoxidizer to suppress pinhole frequency to the level desirable for all but the least demanding appli-cations (approximately 10/50 pinholes per sq ft, de-pending on degree of subsequent reduction). On the other hand, Vincent&apos; has shown that it requires ap-proximately 0.007/0.008 pct A1 soluble in the steel to suppress pinhole formation to <10 pinholes per sq ft. In view of the difficulty of consistently maintaining this aluminum level and the associated problem of tundish nozzle constriction2 when aluminum is added to the ladle, steelmakers have searched for other de-oxidation alloys to circumvent the problem. Calcium-containing alloys offer such a possibility. When employed as a partial substitute for silicon, CaSi (30 pct Ca/60 pct Si) additions are reported to yield adequate pinhole contro1.3 Its use simultaneously preserves the "fluidity" of the steel and ensures good castability.4 Oxygen contents as low as 0.007 to 0.014 pct have been obtained with CaSi additions varying between 3 to 8 lb per ton.5-7 Due to the high degree of vaporization of calcium at the temperature of molten steel and the different methods used to introduce it into the liquid, the reported results show considerable scatter and lack of consistency. According to several investiga- deoxidation tors, the flotation characteristics of CaSi products are not as favorable as those found for alumi-num but are similar to those established for silicates with low alumina content.678 Thus, cleanliness ratings were improved when CaSi was replaced by a com-bination of Si + A1 followed by CaMnSi additions.5,9 In view of the lack of a comprehensive description of the deoxidation potential of CaSi alloys, a study was undertaken of the effect of CaSi additions on oxygen content in comparison to Si + Mn and Si + Mn + Al. The alloy chosen for this study contained 63 pct Si, 32 pct Ca, <3 pct Al, and 2.5 pct Fe. EXPERIMENTAL PROCEDURE The experiments were divided in three groups as shown in Table I. In each group, the CaSi addition was increased successively while maintaining the melt composition constant. The experiments in the third group were designed to determine the effect of prede-oxidation with small amounts of aluminum. Ingot iron, containing 0.02 to 0.03 pct C and 0.03 to 0.05 pct Mn, was melted in a 100-lb magnesia crucible. Pig iron was added to the melt under air in an induction furnace to attain the desired carbon content of 0.05 to 0.15 pct and thereby control the initial oxygen content of the melt between 0.04 and 0.015 pct. After removal of the slag, the furnace was sealed and argon was introduced through an opening in the graphite cover, Fig. 1. When the desired temperature was reached, electrolytic manganese was added to the melt and pin samples were taken (7 mm silica tubing) to establish initial oxygen content. To obtain maximum efficiency, CaSi and metallic silicon were introduced in a sealed steel pipe into the bath. This procedure assured minimum loss of calcium through vaporization as the deoxidants were always released at the bottom of the melt. Samples were taken at regular intervals and quenched in water. Temperature of the bath was measured with a Pt/Pt-10 pct Rh thermocouple. Portions of the pin sample were used for oxygen analysis; five determinations per sample were made to obtain average oxygen content. Oxygen analyses were made by the inert carrier gas fusion method with frequent cross checks with the vacuum fusion method.

Jan 1, 1970