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By Franklin Allison

THE influence of cobalt and nickel on the properties of steel might be readily expected to be very similar. The two elements occupy close and somewhat unusual positions in the periodic table, their chemical behavior and compounds are so completely analogous and their various physical properties, including the ferro-magnetic characteristic, show such a remarkable correspondence, that there is no doubt of a strong resemblance existing between these two metals. However, the variety of, properties that are obtained from steel is due, largely to the fact that the element iron possesses allotropic forms. Alloying elements that affect these allotropic forms, or alter the nature of the allotropic changes, exert a profound influence on the resulting steel. In the case of nickel and cobalt, although these elements are very similar chemically and physically, it is found that they differ distinctly in their individual effects on the critical points of iron. Both nickel and cobalt form a series of homogeneous solid solutions with iron at temperatures above 1800° F. The difference in their behavior does not occur until changes begin to take place in the solid state below this temperature. An inspection of the respective thermal equilibrium diagrams at the lower temperatures makes the opposite effects of nickel and cobalt clearly evident.

Jan 3, 1927

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By Stanislaus Skowronski

THE melting of cathode copper, usually containing 99.98+ per cent. Cu, would appear to be a simple matter. Owing to the well known affinity of copper for sulphur, however, so much sulphur is absorbed by the copper during the operation that a long and expensive refining process is required to remove it, and the final product is inferior in purity and lower in electrical conductivity than the original cathode copper. The furnace operation, as now conducted, includes the following stages: 1. Charging the cathodes into the furnace. 2. Melting the cathodes. 3. Oxidizing period-saturating the copper with cuprous sub-oxide and bringing the metal to the condition known as "set copper;" this is now done by blowing air into the molten bath of metal. 4. Reduction period, otherwise known as "poling," whereby the oxygen previously added is removed by covering the molten bath completely with coke or charcoal and introducing large poles of green wood; this reduces the cuprous sub-oxide, bringing the copper to the "tough pitch" stage ready for casting. 5. Casting Period. Many experiments have been made to obviate the long periods of oxidation and subsequent reduction, with no success. While the copper first, cast will be good, as indicated by the set surface of the bars, it is impossible to hold it to the "tough pitch" condition, and long before the whole charge has been poured the copper will become "overpoled," entailing numerous difficulties. Sulphur has always been blamed by the furnace men as the reason why this refining process is necessary, yet the literature on the subject does not sufficiently emphasize how small a proportion of sulphur will ruin copper for practical purposes.

Jan 3, 1918

By G. W. Doepel, W. P. Sibley

A three-dimensional, multilayer, mathematical model has been developed for predicting performance by mis-c.ible displacement. Areal and vertical coverage, total gas-oil ratio, propane-oil ratio and oil-recovery factor for the multilayer system are computed as functions of hydrocarbon pore volumes of solvent injected. The technique has been programed for use on the IBM-7090 digital computer and was used in calculating perfortnance for continuous miscible displacement of a five-spot, 50-layer .system. Results reported are based on the use of published laboratory data for a single, uniform five-spot layer. However, the technique is sufficiently generalized to calculate performance of any system of layers of any pattern configuration and mobility ratio, provided data are available to describe perfortilance of a single layer. The technique has also been adapted for predicting oil recovery resulting from various gas-driven slug sizes. The results presented for both continuous and slug-type injection show, that permeability variation has 0 significant effect on oil recovery but that solvent composition, within the range studied, has a lesser effect on oil recovery. INTRODUCTION Experience from many of the miscible-phase displacement projects currently in operation indicates that the injected solvent flows preferentially through the more permeable zones in the reservoir. This preferential solvent flow in the high-permeability streaks results in a reduction of vertical coverage (C,) and an attendant decrease in oil recovery. The multilayer technique presented in this paper uses areal coverage (C.,) data for a single layer from the literature and from laboratory measurements of displacement efficiency (herein referred to as microscopic coverage, C,,) to calculate areal coverage (C,) and vertical coverage (C,) for the multilayer system. Oil recovery for the multilayer system is obtained as the product of (Ca) (Cm) (Cv). It is assumed that C, in a single layer is 100 per cent; therefore, (C,) (Cm) is a measure of oil recovery from that layer. Data presented in the literature for a single layer generally deal only with areal coverage and/or microscopic coverage —but seldom specifically with oil recovery. Nevertheless, this technique is flexible in that either Ca and C,, of an individual layer or a direct measurement of oil recovery for the layer may be used. In developing this technique, the problem arose concerning a method of determining the relative amounts of solvent that will enter each layer during any increment of solvent injected. To solve the problem, a dimensionless equation was derived for expressing solvent injected into each layer as a fraction of total injection. A finite-difference method was applied in integrating the equation to determine oil recovery for the system of layers as a function of pore volumes injected. The technique subsequently was programed for solution on a digital computer. The method was applied to predict performance of a deep California reservoir under continuous miscible displacement and has been extended for calculating oil recovery as a function of slug size for gas-driven slug injection. The technique may be applied to any petroleum reservoir that can be adequately represented as a layered system, by considering permeability variation,1-3 permeability range and average permeability. Each layer is then assigned its appropriate permeability, porosity and hydrocarbon saturation. RESERVOIR VOLUME CONTACTED BY SOLVENT It has been recognized for some time that high microscopic coverage (C,) can be obtained by injecting a miscible agent, such as propane, into an oil reservoir. These high values of C,,, have been demonstrated in the laboratory and are normally between 80 and 100 per cent, depending on solvent and reservoir oil composition as well as on reservoir temperature and pressure. In miscible-phase displacement, because there is no interface between solvent and oil, effective permeability to solvent is the same as that to oil. (This assumes that oil saturation has not changed and that water is immobile.) Consequently, mobility ratio (M) simply reduces to that of viscosity ratio . Perrine4 and Habermann5 have shown that large

By G. H. Anderson

H. R. HANLEY*—I have been asked to discuss briefly the development of rotating cathodes for the electrolytic deposition of cadmium. The earliest recorded use of rotating cathodes was by Hoepfner at Frufurt, Germany about sixty years ago. He elec-trolized zinc chloride solution using diaphragms to separate electrodes. In the early experimental work of the Bully Hill Copper Mining and Smelting Co., Shasta County, Calif., rotating aluminum cathodes 4 ft in diam were used in the electrolysis of an acid zinc sulphate solution. Finished cathodes weighing up to 400 lb were produced. Because of mechanical difficulties, this type of cathode was abandoned for zinc, but was later used for cadmium because of the relative smoothness of deposit in comparison with stationary plates with comparable current densities. Cadmium sponge which forms on the cathode at moderate current densities (without special treatment) is entirely eliminated by a slow rotation. The rate of rotation of the cathode has an effect on the mechanical nature of the deposit. A high rate of rotation concentrates the adhering electrolyte on the shaft; a moderate rate appears to concentrate on the cathode a short distance out from the shaft tending to corrode the deposit in the form of a ring. At a very slow rotation (2 to 3 rpm) the adhering electrolyte gravitates nearly vertically, thus avoiding the cutting ring referred to above. The true explanation for the smoother deposits obtained on rotating cathodes may not be given definitely as the numerous factors involved are not thoroughly understood. Smooth deposits are obtained when the orderly growth of the metal crystals in the cathode lattice are disorganized. Thus the crystals form and grow for a very short interval when they are arrested and a new crystal forms. The continued growth of the original crystals provides large crystals and a rough deposit. Also if the acidity of the electrolyte is low, hydrogen gas bubbles adhere to the deposit. As the cathode is rotated the gas surface is brought into the atmosphere where they burst; thus the deposit is made on a surface relatively gas-free. An aluminum hub distance piece was riveted to each aluminum disk 4 ft in diam, slipped on a 4 1/2 in. steel shaft and pressed tight to prevent acid electrolyte seeping through to the shaft. The 9-cathode assembly was supported on insulated bearings. Electrical contact to the shaft was made through what was equivalent to a copper pulley. Sufficiently high conductivity brushes were placed on the face of the pulley to lead the current to the cathode bus bar. The assembly was driven by a link belt contacting a sprocket insulated from the shaft. The lead anodes were semicircular and supported on porcelain insulators placed on the bottom of the cell. Two anodes were provided for each cathode to permit an 8-in. space between them without increasing the ohmic resistance. This ample spacing permitted easy stripping of deposit with the assembly in place. Cathode cadmium was melted under 650 W cylinder oil. After casting, the primary slabs were remelted under molten caustic soda and cast into pencils 1 1/32 in. in diam. Rotating cathodes for deposition of cadmium are used at Risdon, Tasmania, and at Magdeburg, Germany. W. G. WOOLF*—This paper is very-interesting to me because in our work at the Electrolytic Zinc Plant of the Sullivan Mining Co. we had an exactly similar problem—that is, a method of producing cadmium from our purification residue, the recovery of the contained copper as a copper precipitate which could be sent to a copper smelter and the production of merchantable cadmium. It is interesting to me, not knowing of the work of the Risdon people, how closely we approximate them in their main metallurgy, diverging at several interesting steps which I would like to discuss for just a moment. For example, at Risdon they oxidize their purification residue. In our practice we take the current residue as it is produced in the purification department of the zinc plant and process it in the cadmium plant. The only oxidation that it obtains is the oxidation in the presses, the dumping of the presses and the collection and transportation of the residue to the cadmium plant. We find that the leaching of that residue does not necessarily require the oxidation step that the Risdon people evidently find necessary. The discussion of oxidation comes in again in the matter of the treatment of the precipitated cadmium sponge with zinc dust which again at Risdon is oxidized but which we do not attempt to oxidize except as it oxidizes itself in the storage. There is a partial oxidation which cannot be avoided, as Mr. David-sou pointed out, but we make no attempt to attain a complete oxidation and we dissolve the cadmium sponge in the sul-

Jan 1, 1950

Jan 1, 1881

By F. R. Morral

HOT-DIPPED heat-treated zinc coatings on sheet steel were examined with X-rays. The phases identified were compared with those listed in the modern literature of the zinc-iron system. The zeta (FeZn13) phase is shown to correspond to the phase found by Schueler (1925) in heat-treated zinc coatings. An investigation was undertaken to determine the phases present in hot-dipped FIG. I.-SYSTEM IRON-ZINC ACCORDING TO J. SCHRAMM.11 heat-treated zinc coatings on steel sheets and wire. This kind of coating on wire was studied some years ago by Schueler1 by microscopic and chemical methods. He reported a new phase, FeZn10. Later, Clark and Stillwell,2 using X-ray crystal analysis, identified a new phase in ordinary galva-nized wire, which was assumed to be the FeZn10 phase.

Jan 1, 1940

By W. McA. Johnson

The object of this paper is to describe the several necessary characteristics of the zinc-retorting charge and to show how by certain improved methods, the large excess of coal, over that theoretically required, can be reduced, thereby effecting an increase in the smelting capacity of a retort plant as well as other economies. In the southwestern zinc works a charge of roasted zinc ore with a side mixture of carbonates and silicates, analyzing on the average 50 per cent. Zn, is mixed with 55 or more per cent. of its weight of a mixture of coal and coke analyzing 60 to 70 per cent. fixed carbon. The ore is of varying fineness and the coal and coke are usually crushed in a semi-moist condition through a 3/8-in. screen. With ores of a slaggy nature, this coal proportion is increased, sometimes to as much as 100 per cent. of the weight of the ore. With carbonate ores, or where anthracite screenings with higher carbon content are cheap enough, the percentage of reducing agent is cut to 45 per cent. or even somewhat lower. The average charge per retort for a monthly run of a plant on ore, as distinguished from retorts on blue-powder, ladle-skimmings and other between-products, runs from 60 to 66 lb., with a mean of 63 lb. for the Southwest with standard retorts of 8 1/2-in. inside diameter and 50 in. long inside, using roasted sulphide ores. Conditions will change these figures, but 40 lb. of ore per cubic foot of retort space is, I believe,'an average in the better-operated plants. Since the rise in the price of spelter, the tendency has been to overcharge. To remove the slag that is formed when retorts are overcharged, extra men are put on to "gum-chisel" the retorts to prevent "set" furnaces and "butchered retorts." Taking the above conditions as premises, we find that from four to six times the amount of coal necessary for the reduction is used, as is evidenced by the fact that the residues analype as high as 35 per cent, fixed carbon. There are good reasons why this excess of fixed carbon must be used and perhaps these can be best explained by a description of the metallurgical requisites of a zinc charge. We may enumerate these as follows: 1. High reducing power, especially at the end of the shift, to reduce any carbon dioxide to monoxide and bring about a good condensation

Jan 1, 1918

Discussion of the paper of H! M. CHANCE, presented at the Pittsburgh meeting, October, 1914, and printed in Bulletin No. 91, July, 1914, pp. 1461 to 1466. R. V. Norris, Wilkes-Barre, Pa.-In his resume of coal-taxation methods Mr. Chance has apparently confined his investigation to bituminous coal, so a summary of the anthracite situation as to coal assessment should be of interest. The anthracite region of Pennsylvania in its 484 square miles of measures lies in the counties of Wayne, Lackawanna, Luzerne, Carbon, Schuylkill, Northumberland, and Dauphin, and in each of these different methods of assessment have been adopted. The assessments for taxation in Pennsylvania are made under Acts of the Legislatures of 1841 and 1842, under which the assessors are re¬quired to "assess rate and value every subject of taxation according to the actual value thereof, and at such rates and prices as the same would bring at a bona fide sale after due notice." Under this general law coal land has been assessed as such, but up to about 25 years ago the coal was assessed at rates so low as to be unimportant, though the rates in each township and municipality, even in the same county, showed no agreement or ascertainable basis. In the late '80s, the need for greater revenue drew attention to the possibilities of collecting much larger taxes from the coal lands and, spurred by public demand, the assessing authorities in some counties undertook a reassessment of coal. The method pursued was to demand from each owner a statement of the area and average remaining thickness of coal on his lands. On this return, made in some cases under oath, an assessment based on foot-acreage was levied, with a resulting great increase in valuation, but the valuation per foot-acre was low and no serious objection was made. This assessment with occasional increases of rate remained in force until 1907, when, in the triennial assessment of that year, a complete revision of the valuation was again attempted, engineers were employed by the counties to check the accuracy of the returns from the owners, and, based on the foot-acreage thus obtained, an enormous increase, to $65 per foot-acre, was made in the valuation. This assessment was resisted in the courts by appeals resulting in very extensive litigation, and a final determination of the validity of the 1907 and all subsequent assessments has not yet been obtained. In the first appeals tried the rate calculated on a foot-acre valuation was reduced by the Luzerne County court to about $45, while the. Lackawanna County courts reduced a larger rate to $67. The county authorities were, how-

Jan 4, 1915

By Charles H. White

Methods in colorimetry are based on the assumption that the intensity of the' color of a definite volume .of solution is directly proportional to the quantity of the color-producing substance present. In the preparation and examination of colored solutions there are three ways in which the depth of color may be varied. These variables are : (1) the quantity of' the coloring matter; (2) the quantity of the solvent, or the dilution ; and (3) the thickness or depth of that portion of the solution examined. With these three variables as a basis, three general methods in colorimetry have been devised, all of which have been used in the determination of carbon in steel. In the first method, the quantity of the solution is kept the same for both the unknown steel and the standard, and the sections compared are of equal thickness, while the quantity of the standard is varied until the color obtained is of the same intensity as that produced by a definite quantity of the unknown steel. In laboratories where this method is used there is prepared and kept for permanent use a series of solutions with varying amounts of the standard, representing the percentages of carbon from the lowest to the highest demanded in that laboratory. These standard solutions are kept in bottles or test-tubes of equal diameter, and are arranged in series from the faintest to the most intense color, so that it is a simple matter to find the place of the unknown steel in the series, when the solution is of a definite volume and is contained in a tube of the same diameter as the standard tubes. In the second method, equal quantities of the standard and of the unknown steel are dissolved, the standard is diluted to a definite volume, and the solution of the unknown steel is di-

Jan 1, 1907

By T. D. Yensen, R. K. McGeary

While we have not been primarily interested in the determination of hydrogen in the alloys that we have been dealing with, we arc very glad to cooperate in this symposium on sampling and analysis for hydrogen, as we can never tell when we may find it desirable to perform such analyses. Vacuum-fusion Method Our primary interest has been oxygen and carbon, in the analysis of which we think we have become very proficient. Our oxygen-analyzing apparatus is of the vacuum-fusion type (Fig. I). This employs Dr. Derge's type of graphite cruciblc and slit sleeve, with graphite powder between, for heat insulation (Fig. 2) supported in a clear quartz bulb connected to the Pyrex glass system by means of a carefully made ground joint. The latter is kept cool by means of an air blast from a circular copper tube surrounding the joint. The crucible is heated by high-frequency induction and the 24 one-gram samples successively dropped in from the horizontal tube at the top by means of a magnet. The gases (H², CO, N2) are passed through the CuO catalyst held at 300°C. (converting CO to CO²I, H² to H²O) and the CO² and H2O frozen out in the main liquid air trap, while N2 is evacuated. At the end of the operation, shown by the termocouple vacuum gauge indicating a pressure of 10-4 mm. Hg, the analyzing system is isolated by means of the mercury cutoffs, and the CO² and H2oare determined by transferring first the CO2 and then the HzO to the small trap at the right. This is done by placing liquid air on the small trap, then removing the liquid air from the large trap for one minute, which is sufficient to evaporate the Co2 and transfer it to the small trap. The CO2 is then measured by closing the mercury cutoff at the extreme right, removing the liquid air from the small trap, warming it up to room temperature, adjusting the mercury level to one of the three marks on the glass tube (corresponding to three volumes calibrated so that I mm. Hg of CO² corresponds to 0.001, 0.002 and 0.010 mg., respectively, of oxygen) and finally reading the pressure. Liquid air is now replaced on the large trap, the mercury cutoff at the right opened, transferring the COz back to the large trap. The cutoff to the pumps is then opened, the liquid air removed for one minute, thus evacuating the CO², the pumps again cut off and the H2 O transferred to the small trap as before, except that it is now necessary to warm the large trap with a Bunsen burner in order to evaporate all the H2O. This requires about three minutes, as the inner tube must be heated to at least 100°C. In order to prevent H²O from being adsorbed on the glass walls, the glass tubing beyond the catalyst (i.e., to the right in Figs. I and 3) is kept heated to about 150°C. by means of Nichrome ribbon as shown in Fig. 3. For a I-gram sample the above amounts of oxygen (0.001, 0.002 and 0.010 mg.) 35

Jan 1, 1945

By Colin Baker, G. C. Smith

Investigation of the initiation and propagation of ductile failure in OFHC copper was undertaken to determine the role of nonmetallic inclusions. The effect of inclusion initiated voids on the formation of the internal cavity and the final shear separation was studied by metallographic eranzination of strained test pieces. A strain anneal technique was used to enlarge the voids under uniaxial stress conditions to elinzinate triaxial stress effects. Measurements of void size us stress and strain were made to show the point at which void im'tiation begins and becomes an important factor in the deformation process. The work of separation of copper-cuprous oxide was determined to attempt to correlate the breakdown of the matrix inclusion interface with void initiation and propagation. The zloid shape and position relative to the tensile axis suggested an interface breakdown mechanisnz of initiation. Evidence is presented that shows a basic similarity between the central cavity propagation and the 45-deg shear portions of the failure. DUCTILE fracture has been studied by a number of workers1-lo and attention drawn to the importance of hard second phase particles in the initiation of the failure. Holes formed at the matrix-particle interface can elongate by plastic deformation and then subsequently expand sideways to link up and produce a major crack. This is usually observed first in the center of the macroscopically necked region of a test-piece where the hydrostatic stresses are at a maximum. As the crack spreads sideways towards the free surface of the specimen, well defined shear zones develop from the crack tip and the final separation is along a direction at approximately 45 deg to the stress axis. This shear failure may also be associated with voids formed adjacent to second phase particles. In this way a cup and cone type fracture is produced. The stage at which separation takes place between particles and the surrounding matrix has not been clearly identified. In addition, although researchers have dealt with anisotropy of tensile behavior" as a result of material fabrication variables, not much is known about the microstructural features of aniso-tropic behavior. In the present work evidence on these points is presented in relation to the behavior of copper containing second phase particles of cuprous oxide. I. MATERIALS AND PROCEDURES EMPLOYED The material used was +-in. diam or 2-in. sq cold-drawn OFHC copper bar which contained 0.6 pct by volume of cuprous oxide inclusions. These ranged in COLIN BAKER, Junior Member AIME, formerly at -mnF of Metallurgy, University of Cambridge, Cambridge, England, is presently Research Scientist Reynolds Metals Co., Richrnand, Va. G. C. SMITH, Member AIME, is Senior Lecturer, Department of Metallurgy, University of Cambridge. Manuscript submitted June 20, 1967. IMD size from approximately 1 to 6 p in length and 1 to 4 p in width. The shape was generally slightly ovoid. Tensile tests were made on specimens having a gage length of 2.5 cm and diameter of 0.643 cm. Metallographic examination was carried out by sectioning deformed and fractured specimens; in addition fracture surfaces were examined optically and with a scanning electron microscope. Some measurements of the work of separation between copper and cuprous oxide were made, using a sessile drop technique which was a modification of that used by Kingery and umenick." The best metallographic results were obtained by using a vibratory polisher, which minimized smearing of the surface. 11. RESULTS A) Initial Experiments. Specimens from the +-in. diam rod were annealed for 2 hr at 650°C in uacuo, at which temperature complete recrystallization occurred without any change in the form of the inclusion. They were then fractured at temperatures from -190" to 600°C. Cup and cone fractures were obtained at all temperatures from -196" to 400°C. With increase in temperature there was, however, a continuous increase in the extent of the central transverse area and a corresponding decrease in the shear portion of the fracture. Above 400°C, the fractures became intergranular. Sections of specimens tested below 400°C revealed extensive small voids which were always associated with inclusions. However, the voids only reached dimensions greater than the inclusion size in the region of the macroscopic neck, where they were many times longer. Lateral expansion was found only near the fracture surface of the test pieces. As observed by Puttick, the voids were either (a) triangular holes initiated in the direction of the tensile axis and elon-

Jan 1, 1969

By R. H. Jamison

A relatively new haulage system is described. Employed by the Delmant Fuel Co.. the "Full Dimension" system provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. This system is said to provide a higher percentage of recovery as well as additional safety and production. Delmont Fuel Co. is employing a comparatively new system of transportation known as a Full Dimension system. Cne of these systems has been in operation for a year at the company's 10-B Mine as a part of a conventional section. A second was installed at the No. 10 mine in late 1960 to handle the production of a Colmol in a pillar section. SYSTEM COMPONENTS A Full Dimension system is a haulage system that provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. The equipment required for this system consists of a series of interconnected chain conveyors that are mobile and articulated. They will retract or extend a sufficient distance for the development of a five-entry system; or, in the Colmol pillar section, it provides reach of 210 ft in all directions from the section belt. The components of this system are: l)One 160-ft chain line placed in tandem with the belt conveyor. It has a self-propelled drive, is 20 in. wide and 9 in. deep. Moving this conveyor requires the assistance of a loading machine or cutting machine. 2) One 40-ft piggyback that discharges along the entire length of the 160 ft chain conveyor. 3) A mobile bridge carrier, which is a self-propelled conveyor with four wheel steer and four wheel drive, twenty-eight feet long, it delivers coal to the receiving end of the piggyback. Axles steer individually making possible almost lateral movement. 4) Another 40-ft piggyback, duplicate of item 2 that delivers coal along the entire length of item 3 (mobile bridge carrier). 5) A second mobile bridge carrier, similar to the first, which deliver coal to the piggyback (item 4). 6) A third 40-ft piggyback, duplicate of items 2 and 4. This pig is attached to the loading machine and delivers its coal along the length of the second mobile bridge conveyor. Since the original preparation of this paper, the Delmot Fuel Co. has been able to eliminate the 160-ft chain conveyor. This was accomplished by connecting the outby piggyback directly to a loading machine with an extended boom. The loading machine loads directly onto the belt. This change has resulted in a substantial reduction in moving time and greatly increased flexability. A single trailing cable powers the entire string of equipment. It is attached to the side of the equipment in such a way as to keep it off the ground and afford maximum protection. The tramming rate of this equipment is 90 fpm. The conveyor capacity in a conventional section at Delmont's mines is 7.5 tpm and in the Colmol section is 5.5 tpm. This regulation is a simple function of conveyor speed. To visualize operation of this equipment, it would be well for me to touch briefly on local conditions in the Upper Freeport seam in which we mine. (Also, see the photographs of some of the equipment in use.) DELMONT'S TOPOGRAPHY The Delmont Fuel Co. operates two mines in this seam in Westmoreland County, Pa. The No. 10 mine, which was opened in about 1912, is now almost worked out. Depending on economics in the industry, it has a life of two to four years on a declining production basis. A year ago a new drift mine was opened which is called No. 10-B. It is about two miles from the cleaning plant and is connected thereto by an overland belt conveyor. The new mine is being developed at a rate calculated to take up the slack as the old mine plays out. The Upper Freeport seam averages 4.2 ft in thickness in the area of the Delmont mines. It carries 4 in. of boney coal at the top of the seam and a middle man of from 2 to 4 in. We mine just above a 1-in. slate parting which has 4 to 6 in. of highly laminated coal beneath it. This material normally makes a very firm bottom. The roof varies from dark shale to sand rock and 36-in. bolts are placed on 4-ft centers for roof support. All working places are driven 20 ft wide on development and 25 ft wide on retreat. Selection of mobile chain conveyor equipment when it became available, was a very natural move for Delmont Fuel to make, because chain conveyors and piggybacks had been in use at the company's mines for about 12 years. Grades in the new mine

Jan 1, 1961

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine (— 325-mesh) sand assaying 95.5 per cent SiO2, and a very he (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotation Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine fractioh was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1947

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine (— 325-mesh) sand assaying 95.5 per cent SiO2, and a very he (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotation Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine fractioh was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1947

By J. Walter Snavely

PRESENT-DAY processing in many industries, calcining, sintering, briquetting, beneficiation and nodulizing, increasingly calls for the handling of large volumes of hot bulk materials. Various types of conveyors have been employed. This discussion will cover the factors governing their selection. For temperature ranges up to 400°F, or approximately 200 °C, a wide range of conveyors is available. Special constructions of rubber conveyor belts, steel conveyor belts, vibrating and shaker conveyors, apron conveyors, and drag chain conveyors, all are used successfully. As temperatures go well above 400 2F, however, choice of conveyors is narrowly limited. This paper will consider the problem of handling bulk materials only where the temperatures exceed 400°F. The arbitrary selection of 400 °F as a dividing point undoubtedly can be challenged, as special conveyor belting constructions are available which are suitable for temperatures in excess of 400°F. However, when the relatively short life of such belts and the cost of their replacement, with the attendant down time, are balanced against the reliability and long service life of the properly designed steel constructed units to be discussed, there is little question in any operator's mind that the special belts are more expensive to use. Because the conveyors under study are for the handling of bulk materials, inevitably including a high proportion of fines, obviously wire mesh belts cannot be included for consideration. Even though this type of conveyor is widely used at high temperatures, i.e., for carrying glassware through a lehr, it is unsuited for the conveying of bulk materials, and therefore will be excluded from further discussion in this paper. Preliminary to the study of the conveyor itself is the determination as to whether the material is to be cooled while it is being handled, or whether the processing requires retention of all heat and the maintenance of a given temperature range. In the majority of cases cooling is incidental to or part of the handling process, when the handling, for example, follows completion of sintering, roasting, calcining, refining, or some other process. To meet such operating conditions successfully, the conveying medium used must have: 1—a construction capable of withstanding maximum initial temperatures of the material being handled. 2—a construction providing efficient heat transfer for cooling. 3—a construction providing dependable operation and long life with minimum service requirements, and 4—a construction providing controlled and efficient conveying. Under the usual conditions of cooling during the handling, the construction selected to withstand the initial maximum temperatures does not necessarily involve using alloys, as excellent results can be achieved with normal carbon steels and cast irons, when they are properly applied and proportioned. The earliest and simplest type of conveyor for handling very hot materials is the cast steel drag chain conveyor, still widely used for handling hot cement clinker, as illustrated by Figs. I and 2. Because of the rugged and generous proportions of the chain link design, low carbon steels are entirely suitable for the links. The pins, however, must be alloy steel. The simple, rugged construction of this type of conveyor makes it readily capable of withstanding high initial temperatures, even though the chain is operating buried in the material. The drag-chain type of conveyor has advantages and limitations. Although the efficiency of the heat transfer is relatively poor, the life of the conveyor is reasonably long, and because of its crude simplicity it does not require much servicing. However, as a conveyor, it is limited in capacity, and largely limited to horizontal runs. Furthermore, because of the crude design, heavy weight, and the chain operating at the temperature of the material, greatly reducing permissible operating chain pulls, this type of conveyor is limited to relatively short centers. Another type of conveyor that has been used for very hot materials is the cast pan conveyor. Because of its very generous proportions the cast pan, which is made of either cast iron or malleable iron, can withstand initial maximum temperatures. It also provides efficient heat transfer for cooling. Further, it is on efficient conveyor construction, which can be used for inclines. Because the chain employs rolling friction instead of sliding friction, and is not in the maximum temperature zone, much longer centers are possible. It is this type of conveyor that is frequently used in the casting of various metal pigs, pig iron, and aluminum; it is obvious, therefore, that very high initial temperatures are being handled. With this kind of conveyor the return run is frequently sprayed with water to accelerate heat transfer. The build-up of residual heat in the very heavy cast pans is thus overcome. The outboard roller steel pan conveyor is an improved pan conveyor' which provides high rates of heat transfer and substitutes formed steel pans for the heavy cast pans. It is a very efficient conveying medium. The details of this particular construction are clearly shown in Fig. 3. An early application of this type of conveyor is shown in Fig. 4. In this case the conveyor units are handling roasted phosphate rock at average temperatures of 1000" to 1500°F, and frequent maximum temperatures as high as 1900°F. Several widths are used. The capacity of the unit at a speed of 50 fpm is approximately 30 tph per inch of width at peak loadings, average capacity being about 1/3 of peak loading. The assembled conveyor is shown in Fig. 5, with views of both the top and the underside to show all the construction details. In particular, the following general design principles were carried out in this construction:

Jan 1, 1954

By Robert B. Fulton

It is a pleasure to talk to you about the world economics of industrial minerals on this occasion of AIME's 100th anniversary. In order to fit this topic into the time and tenor of such a session as this where you are hearing subjects equally broad in scope one after the other, I will deal with the subject through the medium of eight minerals selected to illustrate import- ant economic factors. I must admit to some bias in the selection, since I picked personally familiar ones to elucidate facts of life in the industrial minerals business. It is a very interesting field, different in many respects from metals and certainly far different from petroleum. The minerals chosen are celestite, diamond, fluorspar, salt, sulfur, ilmenite, rutile and zircon. Celestite The market for celestite is mainly in the U. S. where it is consumed in the manufacture of strontium compounds for red-colored flares and other pyrotechnics, for the production of electrolytic zinc, for glass manufacture, and for the manufacture of ceramic ferrites. Smaller markets exist in Europe and Japan. Production comes from England, Mexico and Spain, and a new mining operation is being started in Nova Scotia. There is no significant production in the U. S. nor in the other principal consuming countries, except England; so it is mainly an international commodity. Celestite illustrates many of the factors important in the economics of industrial minerals. To have a marketable product, grade and freedom from deleterious elements are important for materials used by the chemical industry. In the celestite used to make strontium nitrate, calcium is the most objectionable of the commonly associated cations. Selling price is adversely affected because nitrate manufacturing costs are greatly increased if calcium is pre sent, hence less can be paid for the raw material. Meeting specifications can add significantly to the vendor's cost; hence a balance must be struck between the producer's capability and the consumer's requirements, permitting a reasonable profit for both. Sizing of industrial mineral products is often important to the consum-

Jan 1, 1971

By Phillip J. Maddex, Ole Skaarup

The quite diverse backgrounds of the authors and their mutual interest in reducing transportation costs has produced some interesting innovations. A few words about their background will tell the reason for this joint paper. . . One is a shipping executive with a background of shipowning, ship management, ship and cargo brokerage and chartering, and most facets of the steamship industry, both American and International. The other is a chemical engineer, thoroughly brought up in the American chemical and mineral industries, with several top managerial positions behind him. They began cooperating when U.S. Borax was contemplating a change in its export shipping program and studying the freight cost for shipping borax from The mines in the Californian desert to Europe, and eastern U. S. markets. In the 20 Mule Team days, borax was exported in 50 pound bags. In 1960 it was still shipped in 50 pound hags. Three main products were shipped: two were anhydrous, and all three were sensitive to impurities. The products were bagged at the plant in the Mohave Desert, and shipped about 150 miles by rail to the port in Los Angeles. From Los Angeles, shipment was made by Conference steamship lines at Conference rates. Anyone who has shipped bagged material to Europe by steamer knows that the tab is anywhere from $20 to $30 or more per ton.

Jan 1, 1970

By A. L. Benham, W. E. Dowden, W. J. Kunzman

The mechanism involved in creating a miscible fluid displacement of typical reservoir fluids through porous media by light hydrocarbon mixtures usually involves a number of mixing and separation stages in the contact area. This mechanism has been illustrated on ternary diagrams for miscible displacement of reservoir fluid by rich hydrocarbon gases. A relatively simple method has been devised for calculating the approximate conditions for a miscible displacement of reservoir fluid by rich, light hydrocarbon gases or by LPG mixtures. Miscibility is favored by increased pressures, decreased temperatures, light reservoir fluids and light hydrocarbon-displacing fluids rich in the heavier components. The conditions for miscible displacements of five typical reservoir fluids by various light hydrocarbon mixtures were determined for a range of temperatures from 70" to 260°F and pressures from 1,000 to 3,000 psia. A correlation of these conditions was obtained using the variables of temperature, pressure, C,+ molecular weight of the reservoir fluid, C,+ molecular weight of the displacing fluid, and mol per cent methane in the displacing fluid. This correlation has been applied to data appearing in the literature with very good success. A partial check of the correlation has been made using a step-wise procedure with a windowed PVT cell. INTRODUCTION The term, "miscible fluid displacement", may be defined as any oil-recovery displacement process where there is an absence of a phase boundary or interface between the displaced and displacing fluids. A water flood would not be a miscible fluid displacement, although cycling in a condensate reservoir would be. Another example is displacement of oil by gasoline. Not so obvious, however, are the conditions under which fluids (such as propane, mixtures of propane and methane, or other similar systems) will give a miscible fluid displacement of oil. There are four standard miscible-fluid-displacement processes being used for the recovery of oil from an "nderground reservoir—(1) high-pressure dry-gas miscible displacement, (2) rich-gas miscible displacement, (3) miscible slug using a slug of LPG followed by a lean dry gas under miscible conditions, and (4) gas driven-slug process using a slug of LPG followed by a lean dry gas under immiscible conditions. The first problem faced by the engineer designing a miscible displacement process is to determine the conditions for miscibility. A number of papers have included discussions of the process of obtaining miscibility in these processes.1-8 However, methods for calculating the conditions for miscibility have not been published. This paper includes a method for calculating the conditions for a miscible displacement of reservoir fluid by a rich gas (LPG diluted with methane or natural gas). In this paper, a rich-gas miscible displacement process is defined as an oil-recovery process in the reservoir. The cost of a rich-gas miscible displacement will be decreased by diluting the LPG with methane or natural gas as much as possible. Therefore, it is important to be able to predict the maximum dilution for miscibility. MECHANISMS The mechanisms for obtaining miscibility between a displacing mixture of light hydrocarbons and displaced reservoir fluid often involve a number of displacing and mixing stages. As stated previously, the cost of a rich-gas miscible displacement is decreased by diluting available LPG with natural gas. The problem is to predict the maximum dilution allowed while maintaining miscibility. To get a picture of the displacing and mixing stages, it is helpful to consider these complex hydrocarbon mixtures as pseu do-ternary systems. Graphical representation of a complex system as a pseu do-ternary system can be thermodynamically rigorous if enough variables are used to describe the character of each pseudo-component in one of the phases. However, even if the characterization of each pseudo-component is not complete, pseudo-ternary representation does serve to illustrate the mechanism for obtaining miscibility and can be used to make some general conclusions concerning the conditions for obtaining miscibility. A phase diagram for the ternary system, methane-