Search Documents

By Robert J. White

Strong's equation for calculating bottom-hole pressure reduction due to gas cutting of drilling mud is corrected, resulting in a simpler equation which is easier to use. Use of the equation is illustrated by examples. It is shown that reduction of weight of a 10 lb/gal mud by 50 per cent at the surface due to gas cutting produces only n 3 per cent decrease in hydrostatic head at the bottom of a 5,000-ft well. It is also shown that at the surface 5 volumes of air per volume of mud are required to reduce the apparent mud weight from 8.5 to 6.5 lh/gal in a 3,000-ft well. Illustrative graphs are included to show the effect of various degrees of gas cutting on apparent mud weight at depths of 2,000, .5,000 and 10,000 ft and to show variation of apparent mud weight, with depth for a 100 lb/cu ft (13.37 lb/gal) mud gas cut to 90 lb/cu ft (12.03 lb/gal) at the surface. INTRODUCTION In 1938, Strong' published the following equation for calculating reduction in bottom-hole pressure due to gas cutting of drilling mud h = 1/D [p + pn/100 In p + p(1-n/100)/p(1- n-100)] where: h = depth in feet P = pressure in atmospheres at depth h due to mud column only D = hydrostatic pressure in atmospheres of a col-umn of uncut mud 1 ft high p = back pressure at well-head (in atmospheres) n/100 = fraction by volume of gas in mud at well-head at back pres-sure p This equation was recently found to contain an error due to an incor-rect expression for variation of vol-ume fraction of gas with pressure. Values calculated by the original equation are close to the correct values for small percentages of gas for which it was derived but become increasingly erroneous as the gas content increases and are completely unusable at the very high gas con-tents employed in aerated mud. DERIVATION The nomenclature used is the same as in Strong's paper. The pres-sure dP exerted by a lamina of thickness dh at a depth h ft is ex-pressed by Eq. 1 where x = volume fraction of gas at depth h dP = D (l - x)dh . . . (1) When we express x in terms of the wellhead per cent gas n, the well-head pressure p and the pressure P at depth h, the result is Eq. 2. x = (np/p+p)/(100-n) + (np/P + p) When this is substituted in Eq. 1 and rearranged to separate the var-iables the result is dh = 1/D [dp + np/100-n dp/(P + p]. Integration between appropriate lim-its results in or if the wellhead is open so that p = 1 atm and we rearrange Eq. 4, the result is loss in head = hD — P = n/100-n In (P + 1) ...(5) Eq. 5 can be solved easily by successive approximation using hD as a first approximation for P on the right-hand side and calculating a first value for loss in head. This can then be used to get a new value of P using the left-hand side of the equation. Values converge very rap-idly even for quite large values of per cent by volume gas n. EXAMPLES The effect on bottom-hole pressure of 50 per cent gas cutting of the mud in a 5,000-ft well using 10 Ib/ gal mud will be calculated, we have: h = 5,000 ft 10 X 7.48 144 X 14.7 = 0.0353 atm/ft hD = 176.5 atm n = 50 Substituting these values in Eq. 5 and using hD as a first approxima-tion of P on the right, we have, approximate loss in head = 50/50 In (176.5 + 1) = 2.30 log 177.5 = 5.17 atm. We calculate a new value of P as follows:

Jan 1, 1958

By M. C. Flemings, A. A. Tzavaras

Analytical and experimental procedures are descrihed for quantitative determination of volume-fraction microporosity in metals by mimorarlio-graphg. The analysis assumes spherical pores but allows for the observed fact that pores or portions of pores smaller than a given thickness do not project an image on the film. Volume-fraction microporosity was measuured in cast low-alloy steel and shown to vary from 0.6 x 10-4 to 10 x 10 depending on mode of solidification and on whether the material was air- or vacuum-cast. The aim of work described herein was to develop a method of quantitative measurement of the volume fraction of microporosity in cast metals. Microporosity, present in nearly all castings and ingots, is always interdendritic in nature and may be large enough to be seen on a polished surface with the naked eye. More generally, it is smaller in size, and its detection more difficult. Three techniques might be considered for measurement of microporosity: 1) lineal analysis of a polished metallographic surface, 2) density measurement, or 3) quantitative microradiography. As will be shown later, volume fractions of porosity in directionally solidified castings may be the order of 10-4 to 10-3; such small amounts of porosity are difficult to identify and measure on a polished surface. Density measurement is also not a suitable technique (for alloys) when the volume fraction of porosity is low. This is because the density of the base metal is often not known within the necessary accuracy, due to microsegregation or partial phase transformations. In general, it is not possible to completely eliminate microsegregation without also altering the porosity. Because of the limitations of the foregoing techniques, quantitative microradiography was employed to determine amount of porosity. In recent work, microradiography has been found to be useful for qualitative measurement of porosity, including determination of size and shape of individual pores;&apos;)&apos; however, quantitative determination of amount of porosity has not been attempted. I) CALCULATION OF VOLUME-FRACTION POROSITY The following analysis employs procedures similar to those of Cahn and Nutting3 and Hilliard.4 Major simplifications are that pores are assumed to be spherical and to be present in sufficiently small quantities that no overlapping occurs. A basic difference from the previous treatments is that the limited "sensitivity" of microradiography is considered. Assume first 1) cavities are spherical and of uniform diameter, 2) cavities are randomly distributed, 3) volume-fraction porosity and specimen thickness to pore radius ratio are sufficiently small that overlapping of the projected images of cavities is negligible, 4) sensitivity is limited (i.e., holes or portions of holes of less than a given thickness do not project an image on the film). The sensitivity, is defined as: where 5 = minimum thickness of void from which an image will be projected on the film and t = specimen thickness. Now, from Fig. 1, it is possible to distinguish between two different types of holes. These are 1) holes whose centers lie within the region t/2 > X > 6/2, where X is distance from one surface of the specimen (positive into the specimen), and 2) holes whose centers lie within the region (6 — r) < X < 6/2, Holes whose centers lie outside X = (6 -r) project no image on the film. These two types of pores are considered separately below. Type 1 Pores. When 6 is greater than 0, the image of a spherical cavity of radius r appears on the film as a circle of radius a which is smaller than r. From geometry, Fig. 1: a2=r2-52/4 [2] where a = radius of image of pore on film and r = radius of pore. Since the pores are randomly distributed and spherical V =4/3pr2Nv [3] where V = volume fraction of voids and Nv = number of pores per unit volume and the number of images of Type 1 pores per unit area of film, N&apos;A, is

Jan 1, 1965

By H. Kurushima, S. Tsunoda

An investigation of the hydrometallurgy of Cu-Zn concentrates by The Dowa Mining Co. of Japan is described and a detailed flowsheet of the Kosaka hydro-metallurgical plant is given. Sufficient operating experience has shown that the Dorrco FluoSolids system has proven successful for recovery of both zinc and copper by roasting in a fluidized bed. At present, 2400 metric tons of concentrates per month are being treated and recoveries of 93 pct Cu and 65 pct Zn are being effected. A SURVEY of the Japanese chemical industry in 1949 revealed, among other things, that sulphur-bearing blast furnace gas from the Kosaka smelter of The Dowa Mining Co. Ltd., Kosaka, Japan, was being exhausted to the atmosphere as waste. In addition, it was costing that company an average of ten million yen ($28,000) annually for damages to the surrounding property caused by the sulphur dioxide in the gas. It was recommended by this survey that the gas be utilized for the production of sulphuric acid by the contact process. Mr. Kurushima visited the United States in May 1950 to investigate the engineering, construction, and operation of a contact acid plant. Any acid produced at Kosaka would necessarily have to be 98 pct H,SO, to prevent corrosion in railroad tank cars transporting it to the point of use. However, the production of 98 pct acid requires constant volume and SO, content of the gas. As the converter gas at Kosaka fluctuates in both volume and quality, it was decided to evaluate the smelting process with relation to other methods of recovery. This evaluation revealed a relatively new type of roaster, the Dorrco FluoSolids reactor, which was reported to be capable of roasting the copper sulphide concentrate in an excess of air with the production of an SO1 gas strong enough for acid production and a water-soluble copper sulphate calcine. The copper sulphate then could be leached from the roasted ore and electrolytic copper recovered from the leach solution in much the same manner as practiced by the Chile Exploration CO. for treating naturally occuring copper sulphate ores. Use of a reactor of this type would eliminate the need for a blast furnace and converter and, in addition, would effect savings in labor and fuel. The Kosaka ore is a microscopically fine mixture of copper, zinc, and lead sulphides. The lead and a small amount of the barite present can be separated by flotation, but any separation of copper and zinc proved extremely difficult. In blast furnace operation at Kosaka, zinc had been slagged off until a stockpile of about 12,000,000 tons containing 5 pct Zn had been accumulated. Mr. Kurushima felt that if any of the zinc in the concentrate was sulphat-ized and leached with the copper, it also could be recovered electrolytically. The Dorr Co., sponsors of fluidization technique in the metallurgical field, suggested that sulphate roasting be tested out on a laboratory scale and that, if successful, a commercial Dorrco FluoSolids system be constructed. They indicated that the results from such laboratory tests could be transposed to a commercial scale without pilot planting the roasting operation. It was therefore decided to send samples of flotation concentrate to the West-port Mill, laboratories of The Dorr Co., for testing. The objective of Mr. Kurushima&apos;s visit to the United States thus had changed from investigation of sulphuric acid plants to a fundamental examination of the hydrometallurgy of copper and zinc. Testing Program In February 1951, Mr. Tsunoda went to the United States to witness the application of the FluoSolids system to sulphate roasting at The Dorr

Jan 1, 1956

By Goldye Leeds, Marian Balicki

A SURVEY of literature reveals the scarcity of experimental evidence relating to work hardening and reannealing of silver.&apos;-&apos;" With the exception of the well-known X-ray investigation of silver and other metals by W. A. Wood? the effect of progressive cold working has not been studied. The determination of thermal stability of silver in the work-hardened state by means of isochronal or isothermal annealing has been confined apparently to very high reductions only. In the survey made, no study was encountered which endeavored to produce a complete picture of the changes in various properties due to cold working and subsequent reannealing. To fill these gaps and to assemble data that could be used for the determination of activation energies and of frequency coefficients of primary recrystallization, a systematic experimental approach, which was developed by one of the authors, has been applied to pure silver. Material and Processing The silver used was Handy and Harman fine silver, with a spectroscopic analysis at the testing stage of Cu, 0.0001 to 0.001 pct; Te, 0.0001; Pb, Si, Ca, Mg, and Mn, traces. The material was supplied in the form of a rod 0.406 in. diam and was converted into test samples according to the procedure condensed in Table I. It should be mentioned that air annealing was adopted to simulate industrial practice, though it was realized that annealing in a neutral gas or in a vacuum would be preferred. The usual precautions have been taken to eliminate any structural changes and property changes resulting from the storage of the samples prior to annealing or testing. Thus the wires were kept at the fully annealed stage, and the final wire drawing, which produced the desired degree of work hardening, was done just prior to the final annealing or testing. During the short waiting period, the samples were kept in a thermosflask filled with ice. Even with these precautions some strain-age-hardening or self-annealing could occur. To test such a possibility, wires of different reductions were prepared in the manner adopted and portions of them were immediately tested. The remaining portions were similiarly tested after different periods of storage, both at room and melting-ice temperature. Within experimental error, all samples obtained from a wire of a given percentage of reduction yielded identical values. This result qualified the material for further tests and interpretations, because it showed that either strain-age-hardening and self-annealing were absent, or that these effects fully balanced each other. The final annealing was carried out in an electric resistance furnace automatically controlled to within a 10°F. It was realized that the temperature fluctuations were greater than desired for this type of work. TO diminish the errors due to fluctuations of temperature, the hot junction of the thermocouple was placed as close to the samples as possible; the time of soaking was reduced to 15 min; and the number of samples annealed at different temperatures was increased. It will be seen from Table I that some tests were performed on the same material at rising temperature. These will be discussed elsewhere. Room temperature testing was limited to the determination of some of the mechanical properties that are known to be additive in a degree sufficient to be used for indirect study of the kinetics of the reannealing process, Tensile tests, carried out on a manually operated J. Camillon machine, accurate to within permitted the determination of the ultimate stress (U.S.1 and of the 2 in. elongation, (A,). The proof stress (d,e) has been determined from a series of bend tests carried out on a Tinius Olsen Stiffness Tester by using the offset method. This involved accurate determination for each sample of load-angular deflection curves, finding a

Jan 1, 1952

By V. Koump, T. F. Perzak, R. G. Olsson

The vate of dissolutioz of carbon in molten Fe-C alloys urns studied by votating- cylindrical g-vaphite sairlples in a stationary crucible containing the melt. The rate of dissolltion was determdined from the change in the bnth co)npositiorz. Alccss-tvansfev coeffi&apos;cients, calcilated with the assurlption that the inte)yacial liquid ulas saturated with carbon, were found to be in agreement with an existing Correlation fov mass transfer jroni votating cylinders. It is corzcllrded tlat, itz the wnge of- Reynolds&apos; numbers itzvestigated (790 to 18,000), the vate of dissolutio of carborz in Fe-C alloys is controlled by tile rate of carbon difirsion from the interface. The dissolution of carbon in molten Fe-C alloys has been the subject of a number of investigations. Most of these investigations have been inconclusive because of poorly defined experimental variables. Two of the more sophisticated studies are summarized below. Dahlke and Knacke&apos; studied the phenomenon by, in part, immersing graphite rods into 0.5 pct carbon-steel for different lengths of time. Since the experiments were conducted in air or in CO atmospheres, the melt was strongly agitated by gas evolution. It was concluded that the change in bath composition as a function of immersion time was in agreement with the assumption that the dissolution process was transport-controlled. The hydrodynamic conditions in these experiments, however, were not well-defined and gave no indication of the nature of the dissolution process at higher fluid velocities. In a recent study Shurygin and ruk investigated the dissolution of carbon in molten Fe-C, Fe-Si, Fe-P, and Fe-Ni alloys. Small graphite disks were rotated in a crucible containing molten alloys under such conditions that only the lower face of the disk was exposed to the melt. The results of this investigation indicate that, over the range of velocities (100 to 1000 rpm) and bulk liquid concentrations (0.7 to 4.5 wt pct C) considered, the dissolution of carbon in the iron alloys is limited by the diffusion of carbon from the interface to the bulk liquid (a transport mechanism). For the particular geometry used by these authors, exact analytical solutions of hydrodynamic and diffusion problems are aailable. Hence the experimental results obtained by Shurygin and Kryuk permit the evaluation of diffusion coefficients of carbon in the four systems investigated. In the present investigation a series of experiments were conducted in which the rate of dissolution of carbon in Fe-C alloys was measured under conditions which permitted wide variation of the fluid velocity and the carbon concentration of the bulk liquid. Graphite cylinders were rotated at different speeds in a stationary concentric crucible containing various molten Fe-C alloys. THEORETICAL DEVELOPMENT In general, the resistance to mass transfer from a solid interface to a turbulent liquid is assumed to reside in a laminar boundary layer adjacent to the interface. If it is also assumed that the diffusion coefficient is independent of concentration, the steady-state rate of dissolution for the case of arbitrary liquid and solid densities can be obtained from the integration of the equations of continuity. The derivation is similar to that used by Lommel and chalmers4 for the case of pl = ps. The solution, expressed in terms where N = rate of diffusion of the solute from the interface. Exact analytical solutions of the hydrodynamic equation for the boundary layer thickness are available for only a limited number of geometries. Consequently, a mass-transfer coefficient, k, is usually

Jan 1, 1967

By R. S. Hooper

IN attempting to describe recent mining developments at the Bunker Hill mine, it may be well to recall first the old days of hand mining when holes were laboriously drilled by hand to a maximum depth of three feet, and were shot one or two at a time, in order that the miner could take advantage of every slip, crack or bedding plane, and so save as much drilling as possible. Broken muck, of course, was loaded by hand shoveling into wheelbarrows or small cars, and heading advances or stope ton¬nages were governed by the number of men that could find room to work in any given place. The first big improvement over these old methods was the introduction of the reciprocating type of air-powered rock drill. This innovation, crude as it was, enabled the miners to set up a machine and drill a complete round of holes in one shift; but broken muck was still loaded by hand, and the size and rate of advance in headings were limited by time required to remove the broken rock. The next improvement of consequence was the appearance of the hammer-type rock drill, which greatly increased drilling speed over the reciprocating type, and made possible the completion of a full cycle in one shift, by the use of crossbars in headings, and permitted drilling and mucking to be carried on at the same time. Rock drills continued to improve as time passed, and finally the smaller and more efficient mucking machines appeared on the scene, and it became possible to muck out a round of broken rock and have a clean setup for one or more drill columns and machines. A complete cycle of mucking, drilling, and blasting was completed in one shift with an average advance of five feet. These conditions prevailed for several years, but as the mine grew deeper, the distances to transport the workers grew longer, and this fact together with the shortening of the working day, soon presented the problem of attempting to complete a cycle of operations in a shift of only six hours at the working faces. Anticipating this condition, the operating heads at the Bunker Hill had started an intensive study of more and better mechanization, and over a period of comparatively few years have adopted and placed in service methods and machines that make it possible for a three-man crew to consistently advance headings seven by nine feet in cross section, at the rate of eight feet, and more, per shift, doing all the mucking, switching, drilling and loading. These improvements consist of the following major items: I. The development of a light, flexible, easily transported jumbo equipped with pneumatic columns (Fig I), that has decreased the "setting up" and "tearing down" time an average of 1 1/2 hr per shift, or 25 pct of the available working time. These jumbos were developed at the Bunker Hill mine by members of the staff and consist of two pneumatic columns attached to a suitable carriage. They fold

Jan 1, 1947

By T. Reed Scollon

Coal reserves of the Rocky Mountain area are vast and are more than adequate to meet significantly increasing demands in the immediate future. Practically all of the increase in coal demand for the next decade or so will be in the electric power utility sector. All ranks of coal are available in the area, and from a technological viewpoint, all are suitable for electric power generation. Location of economically recoverable coal deposits, location of generating plants with respect to power demand areas and sources of cooling water, and availability and price of other fuels are the major factors which, individually or in combination, will determine the ultimate selection of Western coal deposits to be developed for electric power generation. The low sulfur content of most western coals and the non-urban location of many power plants should tend to minimize problems of air pollution. If projections for nuclear power prove correct, increases in demand for Rocky Mountain area coals beyond about 1980 will depend primarily on the extent of commercial development of synthetic fuel processes. Current research activities on the conversion of coal to liquid and gaseous fuels are promising. Most western coals range in rank from lignite through high-volatile bituminous coals and are particularly amenable to conversion. Vast reserves of Rocky Mountain coals have long lain dormant primarily because of their remoteness from major market areas. Recently, however, innovations in coal transportation and electric power transmission methods, combined with unprecedented demands for electric power in the West, have resulted in large-scale development of some of these coals. In the past few years, announcements of coal mine openings and transportation plans in the Southwest have been filled with superlatives: The Navajo mine in northwest New Mexico is expected to be the largest mine in the country in the near future. The largest single coal agreement ever made provides for coal shipment from northeast Arizona to Nevada via the largest and longest coal-slurry pipeline ever used. An 1100-mile haul of coking coal from northern New Mexico to California by unit train began last year. It is the purpose of this report to evaluate the coal resources of the Rocky Mountain area, and to project their potential utilization. In projecting energy demand levels on a national basis, the Bureau of Mines has developed some rather complex techniques to cope with uncertainties. Using these same techniques, an estimate has been made of consumption of Mountain area coal in 1980. No attempt was made to estimate coal consumption beyond 1980, but new market potentials are discussed briefly. The present and potential coal-producing fields of the Rocky Mountain area are in six states comprising the Mountain Region - Arizona, Colorado, Montana, New Mexico, Utah, and Wyoming. Since approximately 19% of the coal produced in these fields is shipped to the Pacific Region (Washington, Oregon, Cali-

Jan 1, 1969

By A. H. Daane, R. L. Myklebust

The yttrium-manganese system has been investigated by thermal, metallographic, and X-ray diffraction methods. There are three intermetallic compounds present: YMn2 which melts congruently, YMn4, which undergoes syntectic decomposition, and YMn,, which undergoes peritectic decomposition. The compound YMn4 is ferromagnetic at room temperature with a Curie temperature of 214°C. There are eutectics at 25.2, 60.9, and 82.0 wt pct Mn which melt at 878°, 1100°, and 1075°C, respectively. Crys-tallographic data are given for YMn4, and YMn12 . The terminal solid solubilities are low. In a general program of study of yttrium metal in this laboratory some alloy systems of this metal with elements of the first transition period have been examined. This work was originally instigated by experiences in cladding yttrium with jackets of some protective metals such as Inconel for high-temperature service in air.&apos; In some cases a low-melting phase was observed to form between the Inconel and the yttrium resulting in failure of the samples. In characterizing this reaction, a survey was made of the systems of yttrium with chromium, manganese, iron and nickel,, and it was found that a eutectic was formed between these metals and yttrium on the yttrium-rich side of the system. This present study of the Y-Mn system was carried out to examine in more detail the alloying nature of yttrium, and to correlate trends that have been observed in previous related studies. It was observed by Voge13 that in the systems of lanthanum, cerium, and praseodymium with each of the metals in the first transition series, the tendency to form compounds diminished in the order nickel, cobalt, and iron while no compounds were formed with manganese, chromium, or titanium. Beaudry and Daane4 observed a similar behavior in the systems of yttrium with some members of the first transition series except that the tendency for compound formation was greater. Both the Y-Ti5 and Y-Cra systems consist of simple eutectics, while in the Y-Fe,7 Y-CO, and Y-Ni4 systems, there are four, eight, and nine compounds, respectively. In addition to the above trends, the similar atomic radii and electronegativities of yttrium and thorium invite a comparison between their alloying behaviors with a common element such as manganese. In crys- tallographic studies, Florio et al.&apos; have identified three intermediate phases in this system which are ThMn2, Th6Mn23, and ThMn,,. Gschneidner and Waber10 have examined published information on alloy systems of the rare-earth metals and have correlated this information with current alloying theory. From their study, they predicted that the Y-Mn system would contain one intermetallic compound. On the basis of this prediction, the trend in the alloying behavior of yttrium with the elements of the first transition series and the alloying behavior of thorium with manganese, one might expect from one to three intermetallic compounds to form in the Y-Mn system. A consideration of Hume-Rothery&apos;s rules of alloying based on size-factor, electronegativity, and valency suggested a small terminal solubility and possible compound formation. The present study was undertaken to confirm these predictions of low terminal solid solubility and compound formation and to establish the general alloying behavior of yttrium with manganese. EXPERIMENTAL Materials. The manganese used in this investigation was obtained from the Foote Mineral Co. as electrolytic plates of 99.9 pct stated purity; the yttrium metal was prepared in this laboratory. Table I gives the analyses of these materials. For the solubility studies at the yttrium-rich end of the alloy system, distilled yttrium, whose major impurity was 200 ppm Ti, was used. Preparation of Alloys. The alloys were formed by comelting the two metals in an are-melting furnace under an atmosphere of argon. The buttons thus formed were inverted and remelted three to five times to promote homogeneity. Due to the high vapor pressure of manganese, it was assumed that the weight lost during are-melting was all manganese. This assumption was based on the very good agreement observed between calculated compositions and chemical analyses of several alloys. The compositions of the dilute alloys used for solid solu-

Jan 1, 1962

By E. N. Smith, J. C. Redmond

The increasing need for materials capable of withstanding higher operating temperatures for various applications such as gas turbine blading and other parts, rocket nozzles, and many industrial applications, has brought consideration of cemented carbide compositions. The well known usefulness of cemented carbides as tool materials is attributable to their ability to retain their strength and hardness at much higher temperatures than even complex alloys. However, it has been found that the temperatures encountered in cutting operations do not approach by several hundred degrees1 those involved in the applications mentioned above where the interest is in materials possessing strength and resistance to oxidation at temperatures of 1800°F and above. At these latter temperatures, the tool type compositions which are made up essentially of tungsten carbide are found to oxidize very rapidly and to produce oxidation products of a character which offer no protection to the remaining body. As a further consideration, the density of the tungsten carbide type compositions is high, from about 8.0 to 15.0. The refractory metal carbides as a class are the highest melting materials known as shown by Table 1 which summarizes the available data from the literature for the carbides of the elements which are sufficiently available for consideration for these uses. The density is also included in the table, since as mentioned above it is an important consideration in many of the applications for which the materials would be considered. It has been established that in the tool compositions the mechanism of sintering with cobalt is such as to result in a continuous carbide skeleton and that the properties of the sintered composition are thus essen- tially those of the carbide.2 On the hypothesis that this mechanism holds to a greater or less degree in cementing most of the refractory metal carbides with an auxiliary metal, it appears from Table 1 that titanium carbide compositions would offer possibilities for a high temperature material. Titanium carbide has extensive use for supplementing the properties of tungsten carbide in tool compositions. Although the literature contains several references to compositions containing only titanium carbide with an auxiliary metal,3,4,5,6 it may be inferred from the meager data that such compositions were deficient in strength and were considered to have poor oxidation resistance.7 Kieffer, for instance, reports the transverse rupture strength of a hot pressed TiC composition at 100,000 psi as compared to up to 350,000 psi for WC compositions. The work described herein was undertaken to determine the properties of compositions consisting of titanium carbide and an auxiliary metal and to improve the oxidation resistance of such compositions. It appeared possible that the inclusion of one or more other carbides with titanium carbide might improve the oxidation resistance and also that this might be more desirable than other means from the point of view of maintaining the highest possible softening point. Consideration of the available carbides in Table 1 suggests tantalum and columbium carbides because of their high melting points and general refractoriness. The work on improving oxidation resistance was concentrated on the addition of tantalum carbide or mixtures of tantalum and columbium carbide. The auxiliary metals used included cobalt, nickel and iron. It was also desired to learn the general physical properties of these compositions. Experimental Procedure The compositions used in this study were made by the usual powder metallurgy procedure applicable to cemented tungsten carbide compositions. The powdered carbide or carbides and auxiliary metal were milled together out of contact with air. In some cases cemented tungsten carbide balls and in other instances steel balls were used to eliminate any effect of tungsten carbide contamination. A temporary binder, paraffin, was then included in the mix and slugs or ingots were pressed with care to obtain as uniform pressing as possible. The ingots were presintered and the various shapes of test specimens were formed by machining, making the proper allowance for shrinkage during sintering. Thereafter the shapes were sintered in vacuum at temperatures of from 2800 to 3500°F. Final grinding to size was carried out by diamond wheels under coolant. The titanium carbide used contained a minimum of 19.50 pet total carbon and a total of 0.50 pet metallic impurities as indicated by chemical and spectrographic analysis. It was found by X ray diffraction examination with

Jan 1, 1950

By R. Bakish, C. S. Barnett

IN experiments on tantalum strained in tension, Bechtold did not observe deformation-twinning even at a temperature as low as that of liquid air.&apos; This is an unexpected behavior for a metal of body-centered-cubic structure. Therefore, it is desirable to determine whether the lack of twinning is an inherent characteristic of this particular metal, or whether the reluctance to twin can be overcome under suitable conditions. The following experiments were performed under conditions favorable to twinning and cleavage in their competition with slip, namely, with impact loading on large-grained sheet at low temperatures. The crystallography of cleavage in tantalum sheet was also investigated carefully because of the remarkable ductility of tantalum at low temperatures and because recent tests on oxygen-embrittled tantalum had disclosed unusual cleavage behavior for a metal of this structure: cleavage primarily on (110) planes and only occasionally on {100}.2 The tantalum, supplied by Fansteel, was of 99.9 pet purity, with the chief impurities being iron (0.03 pet) and carbon (0.03 pet). Coarse grains were produced by the strain-anneal method, the anneal being 4 hr at 2000°C in a vacuum of 10-3 mm Hg, followed by furnace cooling. Impact deformation was applied in various ways: a) by hammer blows on an anvil, the specimen and anvil being immersed in liquid nitrogen ( —196°C), b) by driving a precooled center-punch into the sheet, and c) by bending around a crystallographically determined axis with a hammer blow. The impact-deformation produced bands and slip lines on the surface, Fig. 1. The bands have every appearance of being deformation twins, a) They are tilted with respect to the untwinned surface. b) They have the appearance of Neumann bands in iron, thin lens shape, some having serrated edges. c) The orientation in each is different from that of the surrounding matrix and uniform within a given band, see slip lines within them in Fig. 1. d) Upon repolishing and etching, the difference in orientation and the resemblance to Neumann bands is again evident, Fig. 2. (In this micrograph the slight bend in the bands as they cross grain boundaries was noted frequently and results from the high degree of preferred orientation in the sheet.) e) The bands persist after annealing treatments that would have removed them had they been a low-temperature martensitic phase, e.g., after an anneal of 1/2 hr at 900°C. When back-reflection Laue data were plotted for single-surface analysis of several grains, every trace of the bands in each grain could be explained by (112) planes. In addition, two of the bands were located in space by traces on two surfaces, and these also were found to lie on (112) planes within the error of the stereographic plot. Since (1 12) is the anticipated composition plane for deformation twins in body-centered-cubic metals it is felt that the observations listed above are conclusive evidence for the identification of the bands as twins. Similar tests with impact at —77° and at 25°C produced no twinning. It is concluded that the critical shear stress for slip at the rates of straining used was always low enough to prevent the applied stress from reaching the value required for twinning at these temperatures. It was noted that the serrations along the edges of the twin bands exhibited parallel crystal facets. Several similarly oriented notches are visible in Fig. 1, and a remarkable set is shown in Fig. 3. It appears from Fig. 1 and similar areas that notches are frequently associated with the intersection of a slip line with a twin, although not all such intersections are at notches. In some instances a slip line appears to enter the twin, cross it, reflect from the opposite side and recross it, producing a pattern resembling a notch but without producing the notch itself, see Fig. 1. It is suggested that the stress concentration at the end of a slip line is responsible for these details. The strain energy associated with the pile-up of dislocations at a twin boundary may be lessened by untwin-ning or failing to twin a portion of the twin band, or

Jan 1, 1959

By W. C. Thurber, C. J. McHargue

THE deformation textures of metals have been extensively studied because of both the practical implications in metal fabrication and the fundamental insight into the behavior of the metal during deformation. Most studies have been concerned with pure metals or solid-solution alloys and relatively little attention has been directed toward more complex alloy systems. Barrett&apos; and Brick2 have summarized work on multiphase systems in terms of the relative ease of deformation of the phases. Thus, if the phases are deformed about equally, each develops its usual texture for the particular fabrication process. This behavior has been reported for Ag-Cu and Cd-Zn alloys3 and 62-38 brass.4 A dispersion of hard particles in a softer matrix interferes with the normal reorientation of the matrix lattice during deformation, thus distorting the texture or in some cases preventing its development. Increasing the amount of carbon in steel has been observed to increase the randomness of the ferrite5 and almost random textures have been reported for A1-12 pct Si eutectic alloy wires.6 It has also been noted that if the second phase is platelike, deformation causes the platelets to line up in a common direction. The present work concerns a quantitative study of the deformation textures in aluminum-uranium alloys containing 5 or 13 wt pct U. Alloys in this composition range are characterized by a pure aluminum matrix in which the intermetallic compound UA14, is dispersed. From a fundamental standpoint, this system affords the opportunity for studying the effects of a dispersed phase on texture development in an essentially pure metal matrix. Further, these alloys are commonly used as fuels for nuclear research and testing reactors, and a knowledge of the preferred orientation may be useful in evaluating physical and mechanical properties as well as response to fabrication. The deformation textures in sheets of aluminum (99.996 pct), Al-5 pct U alloy, and Al-13 pct U alloy were determined for reductions in thickness of 90 pct by cold rolling. The texture in an alloy rod of the latter composition which had been extruded at 455°C was also determined. It can be readily ascertained from the aluminum-uranium phase diagram that the 5 pct U alloy is hypoeutectic and contains 7.3 wt pct (3.4 vol. pct) UAl4. The 13 pct U alloy has the eutectic composition and contains 18.9 wt pct (9.4 vol. pct) UAl,; however, because of nonequilibrium solidification conditions, the typical eutectic microstructure was not developed. The alloys were prepared by open-air melting in graphite crucibles. The slab castings were cropped and cold reduced 90 pct in thickness on a two-high mill. The slabs were reversed after each pass. Although the uranium-bearing alloys exhibited edge cracking, they could be cold rolled to the desired thickness. Small coupons 1.5 by 0.75 by 0.1 in. thick were sheared from the center of the rolled plate with a major axis parallel to the rolling direction. To obtain a specimen of sufficient thickness for machining of the X-ray diffraction specimen, three coupons were laminated into a single three-ply sandwich with a suitable adhesive. Adhesives which could be cured at room temperatures were selected in order to prevent recrystallization during bonding. Bondmaster M-648 (Rubber and Asbestos Corp.) was used for the pure aluminum laminate and Armstrong A-6 (Armstrong Products Co.) was used for the alloys. The alloy for extrusion was cast in a cylindrical graphite mold! cropped and machined into a 3-in. diam by 4-in. long billet. This billet was preheated to 455°C, upset to 3.125 in. diam, and extruded on a 700-ton hydraulic press at a speed of 6 fpm through a flat-face, 1-in. diam die. The rod was quenched with a water spray as it emerged from the die. Pole-distribution data were obtained using the diffractometric method described by Jetter and Borie.7 A spherical X-ray diffraction specimen of

Jan 1, 1961

By D. S. Bloom

RECENTLY a report has been published on an investigation of the MO-RU system by E. Raub.&apos; In this report it is stated that below approximately 1200°C the system consists of two terminal solid-solution phases and the intervening two-phase field; at 1200°C and above, an intermetallic compound labeled Mo,Ru, makes an appearance. Along with some other little known systems in which it was considered possible for a u phase to form, the Mo-Ru system had been under investigation in this laboratory. Since molybdenum and ruthenium have very high melting points and also since ruthenium is expensive, the investigation was confined to X-ray diffraction by sintered powder compacts. The compositions which were studied were (in atomic percentages): 75 Mo-25 Ru, 67 Mo-33 Ru, 50 Mo-50 Ru, and 25 Mo-75 Ru. All compacts were made from the pure powders, and no chemical analyses were attempted. Each compact weighed about 5 g. All annealing was done in evacuated Vycor capsules, and even though the capsules completely collapsed at 1250°C, they still admitted no air. The results of this investigation in general corroborate the work of Raub. Up to 1150°C the two terminal solutions and the intervening two-phase field were found, and at 1200°C and above an intermetallic compound was found. However, the composition limits of the compound were estimated to lie close to 70 atomic pct Mo rather than 62.5 atomic pct Mo as Raub had reported. At least the X-ray diffraction results indicated that the 75 atomic pct Mo specimen consisted of the molybdenum solid solution plus the compound, while the 67 atomic pct Mo specimen consisted of the ruthenium solid solution plus the compound and the equiatomic composition contained only very little of the compound. The most interesting observation, however, was that the diffraction pattern of the compound was similar to that of the well known u phases, except that unusually large lattice constants were indicated. The d-values, or interplanar spacings, of the diffraction lines (as produced by filtered copper KCX radiation) of the compound are given in Table I, as determined from specimens annealed at 1250°C for 9 hr and then quenched in water. Also shown are the lines of the Mn-Mo u phase as given by Decker, Waterstrat, and Kasper,&apos; the lines of the Fe-Mo a as given by Goldschmidt," and the lines of the Mo-Ru phase multiplied by constants in order to make comparison with the Mn-Mo and Fe-Mo patterns easier. These results indicate that the compound in the Mo-Ru system is apparently isomorphous with the other o- compounds. The rather large lattice constants of the Mo-Ru a phase are acceptable in view of the comparatively large size of the constituent atoms. The data in Table I for the Mo-Ru compound check quite well with those given by Raub, not shown here, though there are some discrepancies between the two sets of data in the estimated intensities of a few lines. Although very little is known at present about this u phase, there are some things which may be pointed out, mainly in comparison with other better known u phases. Even though molybdenum and ruthenium are in the same columns in the Table of Elements as chromium and iron, respectively, the Mo-Ru u phase contains much more molybdenum than ruthenium, while the Cr-Fe u phase centers around or very nearly around the equiatomic composition; furthermore, the Mo-Ru u is stable only above approximately 1200°C, but the Cr-Fe u is stable only below about 825 °C. With respect to its being unstable below 1200°C, the Mo-Ru u is similar to the other binary molybdenum u phases in the Co-Mo and Fe-Mo systems. Similar to the latter two systems, the Mo-Ru u apparently exists over a rather narrow composition range, whereas the Cr-Fe u at lower temperatures is stable over a much wider composition range. The Mo-Ru u thus is closer to the other molybdenum u phases than to the Cr-Fe cr phase. The most striking feature of the Mo-Ru u phase, however, is the absence of any of the transition elements of the First Long Period.

Jan 1, 1956

By J. N. Ong, J. P. Ratledge, J. H. Boyce

SINCE German operators opened the Tsumeb mine in the early 1900&apos;s, continuous operation has been interrupted only by enforced shutdowns during two world wars and the depression of the 1930&apos;s. Original metallurgical practice was blast furnace smelting. This was applicable only to low zinc ore, leaving large tonnages of high zinc ore to be stockpiled on the surface. A gravity concentrator using tables, buddles, and jigs operated for several years, but differential flotation of the ore did not get beyond the experimental stage. When the property was put up for sale by the South African Custodian of Enemy Property there was an estimated 434,000 tons of ore on surface dumps. The dump material, assaying 4.48 pct Cu, 15.8 pct Pb, and 10.9 pct Zn, had become highly oxidized during a stay of many years on the surface. But these dump ores had to be used in evaluating possible metallurgy because the shaft became flooded after the 1940 shutdown and underground ore was unavailable. Many companies examined Tsumeb but quickly lost enthusiasm because of the formidable metallurgical problems presented by the dump ore. However, on the basis of extensive test work, the present company decided that differential flotation would be successful and purchased the property after competitive bidding. Geology and Mineral Association is Complex The Tsumeb mine lies near the northern boundary of South West Africa at an altitude of 4200 ft. The village of Tsumeb is 335 air miles northeast of Wal-vis Bay, territorial port of entry. The orebody is a hydrothermal replacement deposit associated with an irregular intrusion of aplite cutting folded pre-Cambrian dolomite beds. In plan the deposit is a lens of 250x600 ft maximum dimensions. In section the orebody dips 55" south from the surface to 2000 ft, below which there is a reversal in dip to 75" north. Mineralization is localized in the aplite and in the adjoining brecciated dolomite. In the upper levels the ore is largely confined to a single lens, but below the 1890 level massive sulphide ore forms north and south lenses that partially enclose a pipelike body of medium to low grade disseminated ore, part of which is of too poor a grade to mine. Drilling below the 24th, or bottom level, has disclosed that a high percentage of the mineralization from the 3000 to 3300-ft horizons is comprised of copper and lead oxides. The deep oxidation may be local, since it appears due to circulating waters in a major fault cutting the orebody. Tsumeb ranks as one of the highest grade base metal deposits in the world. It is geologically and mineralogically famous for 41 copper, lead, zinc, vanadium, and germanium ore minerals, but its complex mineral assemblage is not attractive to metallurgists. In addition, there is a gradual change in oxidation intensity from the upper levels to the lowest level now operating, which is the 24th. This is well illustrated by the copper minerals. In the upper levels malachite and azurite are predominant, in the intermediate levels chalcocite, and in the lower levels tennantite. Throughout the mine there are sufficient heavy metal soluble salts to activate zinc, and large amounts of depressants are required even for so-called clean sulphide ores. When the same sulphide ore is exposed to prolonged oxidation aboveground, as was the case with the dump ores, reagent requirements become fantastic. Plant Construction First staff members arrived at Tsumeb in March 1947, and by May essential services had been restored and mine unwatering began. In 2 1/2 months company employees built a temporary jig and sorting plant for treating existing dumps to provide a saleable product during high metal prices. Meanwhile, concentrator design and construction was contracted to a South African firm, with subcon-

Jan 1, 1956

By H. D. Kessler, R. J. Van Thyne

Phase diagrams of the titanium-rich portion of the ternary systems from 0 to 10 wt pct Al and 0 to 1 wt pct 0, N, and C were determined. Micrographic analysis of annealed high purity arc melted alloys was the principal method of investigation and was supplemented by X-ray diffraction. TITANIUM-ALUMINUM alloys exhibit excellent properties, particularly for elevated temperature use. For this reason, the Materials Laboratory, Wright Air Development Center, sponsored an investigation of the effect of the interstitially soluble contaminants, oxygen! nitrogen, and carbon, on the Ti-A1 system. Using high purity arc melted alloys and micrographic analysis of annealed samples as the chief method of investigation, titanium-rich partial phase diagrams were determined for the ternary systems. Experimental Procedure Materials: The titanium used in the preparation of the alloys was iodide crystal bar (99.9+ pct pure) produced by the New Jersey Zinc Co. and Foote Mineral Co. The aluminum was obtained from the Aluminum Co. of America in the form of sheet. The given analysis was: Si, 0.0006 pct; Fe, 0.0005; Cu, 0.0022; Mg, 0.0003; Ca, <0.0006; Na, <0.0005; and Al, 99.99. High purity titanium dioxide purchased from the National Lead Co. was used in the preparation of the oxygen-bearing alloys. The spectrographic analysis of this material was as follows: SiO,, 0.07 pct; Fe,O,,, 0.002; A1,0,, <0.001; Sb,O,, <0.002; SnO,,, <0.001; Mg, <::0.001; Cb, <0.01; Cu, 0.0004; Pb, 0.002; Mn, <0.00005; W, <0.01; V, <0.002; Cr, <0.002; Ni, <0.001; and Mo, <0.002. Special spectroscopic graphite rod purchased from the National Carbon Co. was used in the preparation of the Ti-A1-C alloys. Alloy Preparation: Alloy ingots weighing 10 grams were melted in a nonconsumable tungsten electrode arc melting furnace using water-cooled copper hearths and a helium atmosphere. The arc was struck on a tungsten stud in the copper block to minimize contamination from the electrode. Details of the techniques have been reported.&apos;, &apos; Each ingot was inverted and remelted a minimum of four times to insure homogeneity, without opening the furnace. Remelting small control ingots of iodide titanium under similar conditions resulted in no measurable hardness increase, indicating contamination-free melting conditions were used. The alloy charges and resultant ingots were weighed to the nearest milligram. Only small weight losses were obtained upon melting, indicating that the actual compositions were very close to the nominal compositions. Check analyses substantiated this. Oxygen and nitrogen were added during ingot preparation as master alloys. An oxygen master alloy containing 25 wt pct 0 was prepared by melting iodide titanium and pressed titanium dioxide (40 pct 0) powder. The Ti-25 pct 0 alloy is more easily handled during weighing and charging than the pressed TiO, powder. Chemical analysis indicated that titanium-nitride received from an outside source contained large amounts of impurities (2.7 pct Ca). Therefore, a master alloy containing approximately 12 pct N was prepared by melting nitrided sponge titanium. As the sponge melted with difficulty, the alloy was diluted by addition of titanium to lower the melting point. The chemical analysis of the final alloy was 6.69 pct N. Both master alloys are friable and were used as —8 +16 mesh lumps resulting from crushing the ingots. NIelting the Ti-A1-0 and Ti-A1-N compositions five times produced homogeneous ingots. The preparation of homogeneous carbon-bearing alloys caused considerable difficulty. Alloys containing less than 1 pct C were prepared using spectrographic graphite rod, 1/8 in. diameter by 1/8 in. long. However, the graphite did not dissolve even with long melting times and many remelts. The ?/s in. pieces of graphite were used rather than powder because the low density powder is troublesome in arc melting.

Jan 1, 1955

By G. A. Vissac

The only fine coal washer with proved automatic controls, the feldspar jig is capable of good efficiencies even at low separating gravities, handles a variety of products, and treats 150 tph and over. IN continental Europe the feldspar jig is used almost exclusively for cleaning the fine sizes of coal; it operates with an artificial bed, made up generally of pieces of feldspar or of any other hard and heavy rock with good cleavages. A primitive form of this machine requiring close supervision was used in this country years ago, but output was small. The modern machines are generally air-pulsated Baum-type jigs, with capacities up to 150 tph, and treat sizes from % in. to 0, usually previously dedusted or deslimed. These jigs are equipped with very accurate and sensitive automatic controls insuring satisfactory operation even at low separating gravity and with feeds irregular in quality and quantity. This paper will deal more particularly with coal cleaning separation, but the technical studies apply equally to all minerals. Theory of the Feldspar Jig To understand fully the operation of the feldspar jig and to appreciate its potential value, it is necessary to measure and analyze the motion of small particles of various specific gravities, falling in various media, under various conditions. As no practical formulas were available for that purpose, the writer has developed new theories and established new sets of formulas, which cannot be demonstrated here. They are applicable to all minerals and to many problems of coal or mineral concentration. These formulas do not pretend to absolute accuracy, but results, tested by comparison with experimentally recorded data, are sufficient for practical purposes, as they give definite indications of trends, relative values, and order of magnitude of the many factors involved. For pulsating motions no earlier formulas are available. In the case of continuous currents or motions, such formulas as Rittinger&apos;s, Allen&apos;s, or Stokes&apos;, over a hundred years old, apply only to special shapes and sizes of particles, involve factors and coefficients difficult to measure or appreciate, and deal only with free motions. They are of little use in solving practical problems. Units used with the formulas presented here are in the CGS system. Lateral friction and suction have been taken into account according- to methods currently used in aerodynamics, and only hindered motion is dealt with. As the feldspar jig is primarily intended to create a fluid bed condition, the mixture of fluids and solids obtained is homogeneous, and the apparent specific gravity of the separating medium is calculated accordingly. For instance, if 20 pct of the volume is coal and shale, with an average of 1.60 sp gr, in a liquid of 1.10 sp gr, the actual specific gravity of the separating fluid is 0.80 x 1.1 + 0.20 X 1.6 = 1.20. The fundamental equations are as follows: The general motion of a particle falling in a medium of specific gravity equal to u, hindered motion. is dv/dt = g/Ls [ L (s-u) - u/g v2 [1] g = 981 cm per sec, acceleration of gravity, and L = thickness of the particle of specific gravity (s) and velocity (v). The initial acceleration dv/dt = g (1-u/s ) [2] is a function of the specific gravity of the particle, only for a given medium, and independent of its dimensions—an ideal separating condition. The terminal velocity Vt = gxL- (s-u)/u, [3] for a given value of u, is a function of s and L. Two particles, L, s, and L2 s2, are said to be equivalent if they have the same terminal velocity, and accordingly L1/L2=s2-u/s1-u. [4] Two such particles cannot be separated by continuous currents, in any concentrating device. Because the initial accelerated motion, which provides for efficient separation, is followed by a constant velocity, which has a limiting effect, it is essen-

Jan 1, 1956

By A. L. Jamieson, H. Rosenthal

The action of mercury on stressed brass to produce cracks was known before Moore, Beckinsale and Mallinson1 showed that actual season cracking did not occur spontaneously but could be induced by ammonia. These investigators studied other substances, including diphenylamine, without finding that anything other than ammonia could cause season cracking Grimston2 reported season cracking in cartridge cases stored in wooden boxes wetted with dilute sulphuric acid pickling solution. Season cracking associated with sulphur dioxide, water vapor and air has been reported by Johnston,3 although he reported that trimethylamine and pyridine did not cause cracking. Jevons4 has ascribed season cracking of brass in certain instances to trimethylamine, "aldehyde amine," "ketone amine" and pyridine, but no experimental work was done evidently to prove this definitely. Pyridine was investigated by Morris,5 who found it to have considerable cracking power. In the present investigation, the object was a qualitative evaluation of a number of representative amines with respect to their ability to cause season cracking. Thus it was desired only that an appreciable vapor pressure of each amine be obtained in order to make the test as severe as possible, without attempting to test equivalent concentrations for comparative purposes. MEthods Specimens TWO types of specimen were used: 1. Unannealed 70-30 brass CUPs formed from 0.040-in. thick sheet. Height of cup Was 13/8 in. and diameter was 17/8 inches. 2. First draw pieces of caliber .50 cartridge cases in the unannealed condition (70-30 brass)- The cups were representative of a thin wall and thin base specimen, whereas the draw pieces have a thick base (1/4 in.) and a thick wall (approximately 3/32 in.). Both types of specimens contained high residual stress and cracked in less than one minute in a solution of I per cent mercurous nitrate and I per cent nitric acid provided that this was receded by a 30-sec. pickle in 40 per cent nitric acid. Specimens were prepared as follows: (1) degreased in trichlorethylene, (2) rinsed in H2O, (3) pickled 5 min. in 10 per cent by volume H2SO4 (1.84 SP. gr.), (4) rinsed in H2O, and (5) dried. Amirtes Tested In selecting the amines, representatives of the three series of amines, and one series of substituted amines were chosen. The aliphatic series was represented by the methyl and ethyl primary, secondary and tertiary amines. These are the simplest aliphatic amines. For the aromatic series, the phenyl primary, secondary and tertiary amines were selected as being the simplest. The heterocyclic series was represented by pyridine, which is a simple heterocyclic base (tertiary amine).

Jan 1, 1944

By D. W. Fuerstenau, D. A. Sullevan

In the experimental work for this comparison of wet and dry grinding, it was found that the size distributions for wet grinding operations are characterized by a constant value of the distribution modulus, (a; whereas, for dry grinding the distribution modulus decreases slightly with increased size reduction. Wet grindability seems independent of the weight of material in the mill while dry grindability depends on weight. Experimental data and eqmtions for each process are presented for noncleavable isotropic material, for cleavable or material in which there is some form of internal stress, and for energy consumption. The equation ED/EW=BK-? (where ED and Ew are, respectively, energy for wet and dry grinding, k is the size modulus, and B and ? are constants) was found to give the ratio of energy. Dry grinding is generally considered to require more energy than wet grinding. Taggart&apos; states that power consumption in wet grinding ranges from 60 to 90 pct of that required for dry grinding. Bond&apos; states that dry grinding requires one third more power than wet grinding. It is interesting that Rose and Sullivan&apos; note this same difference in power required for wet and dry grinding, but they state that on purely mechanical grounds it is difficult to see any great difference in fundamental principles between wet and dry milling, since dry milling may be regarded as wet milling with a fluid of low viscosity (i.e., air vs water). Yet it must be the viscosity of the fluid that is responsible for the differences between wet and dry grinding. Probably the most comprehensive investigation comparing wet and dry grinding has been that made by Coghill and Devaney. They compared the grind-ability of both chert and dolomite in a 19x36-in. batch ball mill under both wet and dry conditions. In these tests the weight of ore in the mill was varied from 35 to 200 lb and each grinding test was carried out for such period of time that the amount of -200-mesh material in the product was essentially the same for each test with a given material. Cog-hill and DeVaney&apos;s data for the grindability of dolomite, which are reproduced in Fig. 1, show that a comparison between wet and dry grinding is com- plex. The comminution of a material by a wet grinding method is characterized by a constant grindability over a wide range of mill loadings. Since grindability is the weight of material comminuted per unit of energy expended, it is equivalent to (w)/ (p)(t), where W is the weight of material being ground, P is the power, and t is the grinding time. In the case of wet grinding, the grindability of dolomite in the Coghill and DeVaney experiments was 0.546 * 0.009 ton per hp-hr when the load of dolomite in the mill ranged from 50 to 200 lb. On the other hand, for dry grinding no such simple relationship was found to exist between power, time, and weight of material in the mill. Dry grinding efficiency appears to depend greatly upon the amount of material within the mill. When the mill contained

Jan 1, 1961

By G. F. Bolling

STUDIES of grain boundary migration in zone-refined metals have all shown that the rate of migration is greatly reduced by small added solute concentrations. However, it is apparent that a difference exists between boundary migration during normal grain growth and single boundaries migrating in a bicrystal to consume a substructure. To effect the same reduction in velocity in the two cases, much more solute is required for grain growth than for the single boundary experiments. One case is available for direct comparison; both Bolling and winegardl and Aust and utter&apos; added silver and gold to zone-refined lead to study grain growth and single boundary migration, respectively. For comparable reductions in migration rates, about 500 times more solute was required to retard grain growth than to retard the single boundaries. A reason for this difference is suggested here. The rate of grain boundary migration is dependent on solute concentration and must therefore also depend on the solute distribution; i.e., regions of higher solute concentration encountered by a moving boundary must produce greater retardation and thus could determine any observed rate. A dislocation substructure can be the source of a nonuniform solute distribution since it can attract an excess concentration of certain solutes. In fact, it is probable that the solutes which impede grain boundary migration most would segregate most severely to a substructure for the same reasons. Thus a dislocation substructure present in a crystal being consumed could locally magnify the concentration of solute confronting an advancing grain boundary. In the single boundary experiments a low-angle substructure, within single crystals obtained by growth from the melt, was used to provide the driving force to move a grain boundary; in grain growth, no substructure of this magnitude was present. The increased solute concentration at subboundaries should be given approximately by C, = G e c,/kT, where t, is a binding energy and CO the bulk concentration. To account for the difference between the two experiments in the Pb-Ag and Pb-Au cases, C, must be the concentration impeding the single boundary migration, and a value of t, = 0.25 ev is necessary. This is reasonable, even though calculation on a purely elastic basis gives t, = 0.12 ev. because electronic effects must enter for silver and gold in lead. The compound AuPbz forms3 and the metastable compound AgrPb has been reported to nucleate at dislocations prior to the formation of the stable, silver-rich phase.4 Other observations support the hypothesis that a magnified solute concentration impedes the single boundary migration. For example, some crystals were grown by Aust and Rutter at concentrations of ~ 0.1 wt pct Sn and 2 x X at. pct Ag or Au which exhibited a cellular substructure, and in these crystals no boundary migration was observed. It is therefore evident that the higher concentrations at cell boundaries drastically inhibited migration. Inclusions would not have been responsible for this inhibition since according to recent work on cellular segregation,5 no second phase should have occurred in the segregated regions at the cell boundaries for the conditions of growth used, at least in the Pb-Sn system. In the purest lead, only the "special" boundaries observed by Aust and Rutter gave rise to the same activation energy as that obtained in grain growth. It is reasonable to suppose that the structure of special boundaries does not favor segregation at low concentrations and thus solute, or an inhomogeneity in its distribution, would have no effect. Random boundaries, on the other hand, are affected by solute and the substructure would enhance residual concentrations in the zone-refined lead, leading to a higher activation energy. It is clear, even without a detailed theory, that the apparent activation energies and exact solute dependence in the two experiments must be different as long as the non-uniform solute distribution produced by the substructure is important. Recrystallization experiments should also be susceptible to the same kind of local segregation at subboundaries or disloca tion cell walls; a suggestion similar to this has been made by Leslie et al.&apos; Following the arguments presented here, the effects of a given solute concentration would be like those observed by Aust and Rutter if segregation occurred, and like those of grain growth otherwise. This work was partially supported by the Air Force Office of Scientific Research; Contract AF-49(638)-1029.

Jan 1, 1962

By R. E. Jenkins

The Bell Canyon member of the Delaware Mountain group has yielded quite a large number of fields in which completion and production problems have been numerous and complex. Reserves are difficult to estimate due to the problem of evaluating the formation water saturation, and the feasibility of water flooding is questionable. Extensive laboratory investigations were undertaken to determine if normally measured rock and contained fluids properties could account for many of the peculiar characteristics of the formation. A statistical study or routine core-analysis data was made. Capillary-pressure data were obtained by several techniques. Fresh-water and brine permeabilities were measured and relative permeabilities of oil, water and gas were determined. Some qualitative wetta-bility tests were performed, and the petrography of a few thin sections was studied. The results of this work are presented, with typical or average data shown for each test. The permeability -to- water characteristic of the Delaware was found to be abnormally poor. Correlations of laboratory and field performance data indicate high water saturations in much of the formation. Pore geometry is highly uniform. INTRODUCTION The Delaware formation has undergone extensive exploration and development in the past few years. This has resulted in the discovery of quite a number of fields—producing oil for the most part. A large percentage of the productive wells have produced water along with the oil. Often there is no apparent pattern for the percentage of water cut experienced. In several instances vertical displacments of over 200 ft have been established between communicating wells in which the formation appears to be about the same, yet the structurally high wells yield a much higher water cut than the lower wells. Drill-stem tests have contributed little to well completion because they nearly always show small amounts of water (usually mud filtrate), regardless of what is produced later in the life of the well. Other formation evaluation methods have had no more than limited success in indicating ultimate productivity. Consequently, predicting the type of fluid productivity of Delaware wells has been very difficult. Hydraulic fracture treatments are almost universal, and unstimulated Delaware wells have not exhibited the productivity that normally would be expected of wells with the specific permeabilities encountered. Some operators have indicated that their Delaware wells need frequent re-stimulation. The uncertainty of reservoir water saturations reported for the Delaware has made difficult the task of evaluating oil in place in the reservoirs and estimating primary recovery. The unreliability of this data also will cause feasibility studies for water flooding or other secondary-recovery methods to be problematical. Several opinions have been voiced to explain why the Delaware performs as it does. Chronically poor well completions, dynamic water and tilted water tables, capillary inequi-libria, peculiar relative permeability characteristics and oil-wetness are among the suggested reasons. The dual purpose of this paper is (1) to describe the principal rock characteristics, and (2) to present an evaluation of those characteristics so that approaches can be made to explain the irrational behavior of wells completed in the Delaware. STUDY OF ROUTINE CORE-ANALYSIS DATA A statistical study of routine core-analysis data has been made to relate fundamental rock properties and characteristics of the formation to well productivity. For the most part, well production data were obtained from operators and augmented by production data from commercial information services. Only data which were available in detail and which were considered to be accurate were used. The core data used were obtained on samples from the perforated or open interval only and, thus, may not always represent the total productive interval in a well. All the core-analysis data included in this study are from the Bell Canyon member of the Delaware and were obtained by "plug-type" techniques, as opposed to whole-core or full-diameter techniques. Porosity values were obtained by the summation-of-fluids technique, permeability values are air permeabilities corrected for slippage to an equivalent liquid permeability, and the oil and water contents were determined by means of a high-temperature downdraft retort. All the procedures used are discussed at length in Ref. I. The core-analysis data were sorted

By H. Macpherson

Durhm has a well earned reputation for supplying some of the finest coking caals in the world. The caals, in general, vary in rank from 301 to 501/2. Durham has traditionally produced foundry coke for the major proportion of the foundries in Great Britah. Durham coals can be described as clean, soft coals, which yielded, with the old hand methods of coal getting, low ash small coal because the extraneous shale was normally harder than the coal and occurred in larger pieces. Under these conditions, with a plentiful supply of cheap labor, it used to be sufficient, at most collieries, to hand clean the large sizes of coal which could then be remixed with the untreated small coal to produce carbonization coals, ready for the coke ovens, at under 5 pct ash. With the introduction of mechanical methods of coal mining, the coal gradually required more cleaning. The first method adopted to resolve this problem was the introduction of pneumatic dry cleaners for the small coal. Although such machines had little effect on the fine coal (say below 1/8 in.), they could clean the intermediate sizes of coal. This, coupled with hand cleaning the larger sizes above 1-1/2or 2 in., resulted in a combined run-of-mine mixture below 6 pct ash, capable of maintaining the quality and reputation of the cokes produced. In more recent years, the intensification of mechanization and power loading, coupled with gradual exhaustion of the cleaner seams, has created the need for a more complete method of coal cleaning. This particularly applies in the fine sizes (say below 1/50 in.) which normally vary, under present day conditions, between 15 and 35 pct ash and are much too dirty to be included in the raw state in a carbonization mixture. This pronounced change has been accelerated because legislation controlling the dust conditions of mine airs for the prevention of pneumoconiosis has resulted in tk~e extensive use of water underground and a consequent increase in moisture content of the run-of-milie output. The presence of damp fine coal decreast the efficiency of prescreening and dry cleaners, so that this type of preparation for low ash coking coals is decreasing, although it is still used satisfactorily for industrial coals in the medium ash range. Table I shows the gradual increase in mechanization, the reduction in manpower, the increased use of explosives per ton of coal extraction, and the increase in the proportion of coal cleaned by mechanical means in Great Britain. Although similar figures are not available for Durham Div. until after the date of nationalization of the coal industry, it is probable that the increase in mechanical cleaning, particularly by dry cleaners, was more marked in the Durham collieries than elsewhere in the country. As dry clealiers were replaced by coal washeries in the Gritish ccal industry, no special attempts were made to recover the slurry, with the result that large outflows of dirty water were allowed, deposits of slurry came in to lagoons and neighboring streams, and a proportion of fine material was lost from the coking coal. This position, coupled with the higher moisture of the washed coking coal, resulted in adverse effects on coke oven throughputs and coke quality. It is now realized that the natural coal fines are an essential ingredient of coking coals in obtaining the correct coke structure in metallurgical cokes. This, together with economic pressure, led to the introduction of flocculation and filtration plants for the recovery of slurries, and later, when the ash contelits of the filter cakes were too high, to the introduction of froth flotation equipment. After this position had been reached, the tailings from the froth flotation pIants were, in many cases, still allowed to constitute an undesirable effluent. Recent legislation on river pollution has changed this picture; it is now necessary to provide a circuit which is completely closed so far as solids are concerned. The gradual increase in the coal cleaned by wet methods and froth flotation in Durham Div, is shown in Table 11. It is now an accepted feature of new Washeries that. froth flotation should be an integral part of the washing process from the initial installation of the plant.

Jan 1, 1962