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By E. V. Kleber, V. F. Novy, R. C. Vickery

The constitutional diagram for the gadolinium-cobalt system has been determined. Seven intermetallic compounds have been found at compositions corresponding to the follwing gadolinium-cobalt ratios: 3:1, 1:1, 2:3, 1:2, 1:3, 1:4, 1:5. The GdCor compound has a congruent melting point above 1600°C. whereas the remaining compounds melt incongruently. Eutectics exist at 16 wt pct Co and 84 wt pct Co, with melting points of 620° and 1260°C, respectively. PREVIOUS papers'" have recorded constitutional data for systems of gadolinium with iron and nickel. This paper presents data obtained on alloys of gadolinium with the remaining 3d transition element-cobalt. voge13 found that cerium forms, with cobalt, the intermetallic compounds; Ce3Co, CeCo, CeCo2, CeCo3, CeCo4, and CeCo5 Other workers4-' found the 1:2 compound in this series to exist in a C-15 (face-centered cubic) structure and the 1:5 compound to exist in a hexagonal crystal system. Comparison of earlier data with that now reported will be made subsequently. EXPERIMENTAL Methods of alloy preparation, annealing procedures, and experimental techniques have been previously reported.' The cobalt content of alloy specimens was determined volumetrically. RESULTS Microstructure—The alloy containing 10.5 wt pct Co corresponds closely to the composition of the first intermetallic compound Gd3CO. In Fig. 1 the microstructure is essentially single phase composed of Gd3Co which formed peritectically. The globular areas are remnants of the peritectic reaction which did not quite go to completion. A eutectic occurs between Gd3Co and the next higher compound GdCO, which is followed by a series of peritectic compounds. Fig. 2, an arc-cast alloy containing 20.2 wt pct Co, shows large primary crystals of GdCo etching either white or black (depending on crystallographic orientation), embedded in the eutectic. Similarly, in Fig. 3 at 31.2 wt pct Co, a compound, probably GdCo3, has separated from the melt. Subsequent peritectic reactions were almost completely by-passed as the alloy cooled, the residual liquid freezing as a fine eutectic. In the 47.5 wt pct Co alloy, Fig. 4, the high melting point compound, GdCO4, has solidified as massive, white-etching grains, which reacted with the liquid to form GdCo3 as the second phase. The microstructure of an arc-melted alloy containing 60.5 pct Co is complex, Fig. 5, and presents some difficulty in interpretation. The grey etching matrix is probably primary GdCO4. The dappled phase represents grains of GdCO5 which have separated from the melt on cooling. The discrete spots in this phase are eutectic liquid. Small light etching crystallites are probably GdCO5 formed by the incomplete high-temperature peritectic reaction. GdCO4 is retained in arc-melted specimens containing 70.5 wt pct Co, Fig. 6. White areas of this compound are held in a matrix of GdCO5 the appearance of which is mottled slightly by inclusions of the second eutectic in this system. The development of the spheroidized eutectic at high cobalt contents is depicted in Fig. 7 (80.3 wt pct Co) in which the white massive crystals are GdCO5 compound. Further development of this eutectic is seen in Fig. 8, the primary white phase here being pure cobalt. Thermal Studies—Differential thermal analyses and melting point determinations have yielded the data shown in Fig. 10. The high gadolinium eutectic, indicated by microscopy, was confirmed at about 16 wt pct Co, melting at 620°C . A eutectic also occurs at about 84 wt pct Co melting at about 1260°C. The thermal arrests at 8803 960°, 1060; and 1210°C,

Jan 1, 1962

By Ralph W. M. Lai, D. W. Fuerstenau

This paper presents the results of a study of the principles which control the distribution of ultrafine particles between an oil phase and an aqueous phase. Alkyl sulfonates were used to control the wettability of fine alumina particles and to stabilize the water-isooctane emulsion system in the extraction process. The effects of sulfonate chain length, sulfonate adsorption, pH, and oil-water ratio on the recovery of solids in the oil phase were investigated. Higher extraction was found under conditions which yield higher hydrophobicity and large interfacial area of the oil-water emulsion. Today there is considerable interest in developing methods for the recovery of fine particles from disseminated ores. One such method might be the extraction of fine particulate solids from aqueous suspensions into an oil phase, and one example of a successful industrial process based on oil-water distribution phenomena is the emulsion flotation process used in concentrating manganese ore.' The oil agglomeration process of Farnand, Smith, and Puddington has been suggested for concentrating tin ore. Patents 3-4 for oil flotation processes extend back to that of William Haynes in 1860. The basis of all these processes is directly related to the stabilization of emulsions with solid particles. The first extensive study of oil-water emulsions stabilized by finely divided solids was made by Pickering. Later, Briggs,6 weston, 7 Chessman and King,8 and Leja and Schulman9 experimentally demonstrated how solid particles control water-oil and oil-water systems. Hildebrand10 suggested that fine particles can stabilize an emulsion because the fine particles are partially wettable by both oil and water, that is, that the contact angle must be finite. Takakuwa and Takamori 11 recently applied emulsion inversion to the study of collector adsorption in flotation systems. At the Royal School of Mines, Shergold and Mellgren12 are currently carrying out a comprehensive study of the extraction of fine particles into an organic liquid phase. In the present study of the liquid extraction or oil flotation of ultrafine particles, the recovery of hydro-philic alumina particles of 0.1-µ average diameter was determined after separating an isooctane-water mixture (emulsion) into its two phases. Alkyl sulfonates of different chain lengths were used to change the wettability of the alumina and to effect emulsifi-cation of the oil-water mixture. The effects of the oil-water volume ratio, sulfonate adsorption, and pH on the recovery of the fine particles were investigated. Since the surface charge on an oxide such as alumina is determined by the adsorption or interaction of H+ and OH- with the surface, pH control is extremely important. Furthermore, since the zero point of charge (zpc) of alumina13-14 occurs at pH 9 to 9.4, major adsorption effects with anionic collectors in these systems will occur below pH 9. The results of the foregoing investigation are discussed in terms of the recovery of the fine particles, the oil-water ratio, contact angles, surface tension, the adsorption of sulfonate onto the solid surface, and the emulsifica-tion of oil. MATERIALS AND EXPERIMENTAL METHODS Materials: The particulate solid used in all the experiments was Linde "A" alumina, a high-purity, readily available form of a-alumina. The surface area of the alumina, as evaluated from measurements of stearic acid adsorption from benzene, was 14.5 sq m per gm, based on a value of 20.5 sq Å for the cross-sectional area of stearic acid.13 All aqueous solutions were prepared from triply distilled conductivity water. Other materials included high-purity crystals of sodium alkyl sulfonates, isooctane, and, for pH adjustments, reagent grade hydrochloric acid and sodium hydroxide-Because sulfonic acids are strong acids, they are

Jan 1, 1969

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By Frank T. Weems

The basic operating properties of the DorrClone are discussed and certain metallurgical applications which exploit these properties are presented. An effective method of controlling the consistency of the DorrClone underflow is described. THIS paper presents in a general manner some of the applications of the DorrClone (Dutch States Mines cyclone) in the metallurgical and heavy chemical fields. Technical or mathematical analyses of the design factors and of the mechanism whereby the cyclone effects its separation have been capably treated in a series of previous papers.'-' Engineering and operational aspects of a few installations and laboratory tests have been chosen to emphasize the unique classification characteristics of the DorrClone and to illustrate its utility in certain difficult situations. Its use in coal washing and heavy-media operations are not considered. Some 25 or more industrial installations and applications of the DorrClone in the metallurgical fields are operating at present, ranging from single trial units to complete batteries composed of a multiplicity of units. The data were obtained in part from some of these, and in part from operations of units at the Westport Mill, the laboratories and testing plant of the Dorr Co. at Westport, Conn. The DorrClone was developed, in principle, by the research dept. of the State Mines, Limburg, Holland, during World War 11. Its first use was to dewater a —50 micron sand used as a dense medium in coal washing. The metallurgical wet cyclone was first publicized by Driessen and Yancey and Geer.6 ince then the cyclone has been studied as a means of de-watering, heavy media concentration, and fine size classification. The DorrClone is shown in Figs. 1 and 2. pulp to be treated is pumped through the feed nozzle, A, into the feed cylinder, B. The feed nozzle is of rectangular cross-section, the outermost edge being tangent to the feed cylinder walls. Feed shims, C, may be attached to the innermost walls to adjust the open area of the feed nozzle available to entering pulp. The tangential injection of feed under pressure supports the vortex action of the pulp that has been detained in the cyclone; the centrifugal forces of this vortex throw the coarse particles to the walls of the cone, D, where they collect and pass downward and out through the adjustable apex opening, E. The fine particles contained in most of the water that entered with the feed pulp move to the inner spiral of the vortex and flow out through the vortex finder, F, contained in the orifice plate, G, which forms the top of cyclone chamber. The overflow pulp may then spray freely into a spray chamber, H, surmounting the cyclone or pass directly into a 90' elbow, I, of somewhat larger diameter than the vortex finder. Installation Where more than one cyclone is to be installed, best overall efficiency and ease of control are obtained with equal distribution of both pulp volume and weight of solids among the cyclones, and with the same pressure being maintained at each cyclone feed nozzle. Fig. 3 is a suggested arrangement for feeding two or more cyclones in parallel. Feed pressure control: Under normal operating conditions it is not unusual for some surge to exist in the feed that is to be sent to the cyclone. As will be discussed later, surges in the solids content of the pulp are usually not of great operational consequence, but surges in feed volume are a problem. Several systems have been devised for absorbing surges in feed volume, each involving a surge tank. If the surges are small and fairly rhythmical, a tank of sufficient size might be installed to dampen out

Jan 1, 1952

By B. A. Wilcox, A. H. Clauer

The tensile and creep deformation characteristics of fiber reinforced composites have been studied, primarily at room temperature. The matrix was an alloy of Mg-14 wt pct Li-1 wt pct Al (LA141A alloy) and the reinforcing fiber consisted of continuous wires of a high strength precipitation hardening stainless steel (AFC-77 alloy). The ultimate tensile strength and strength-to-density ratio of composites were much improved over that of the mmix, both at room ternperatwe and up to 200°C. Creep studies at room temperature revealed logarithmic weep behavior for wires and composites and steady state creep for the nonreinforced matrix. It was concluded that creep of the fibers controlled composite creep. The instantaneous logarithmic weep rate of the composite, 6, was related to the average tensile stress on the fibers, where n = 3.4. Mg-Li alloys are the lightest structural metals commercially available, e.g., the alloy LA141A (Mg-14 wt pct Li-1 wt pct Al) is 27 pct less dense than beryllium Although Mg-Li alloys have moderate tensile strengths at normal strain rates and ambient temperature, their creep strength is relatively poor, even at room temperature. Reinforcement by fibers offers a means of increasing the tensile strength as well as overcoming the limitations of low creep strength. With a proper choice of reinforcing filament, composites could be designed with higher strength-to-density and modulus-to-density ratios than the matrix alloy. This communication reports a study on the tensile and creep characteristics of alloy LA141A reinforced with continuous wires of a high-strength precipitation hardening stainless steel (AFC-77 alloy). The matrix has a bcc structure, with possible precipitates of AlLi and MgLi2A1,3 and is ductile at room temperature Stainless steel wire was chosen as the -reinforcing filament since previous work at Battelle4 showed that a good bond between wire and matrix could be produced in this system. For all volume fractions of reinforcing wire, the theoretical composite strength-to-density ratio is greater than that of the matrix and the modulus-to-density ratio is somewhat lower than that of the matrix. EXPERIMENTAL PROCEDURES The 0.004-in. diam AFC-77 wire was obtained from two suppliers: National-Standard Co.* and Crucible 0.13 Si, 0.18 V, 0.011 P, 0.012 S, balance Fe. In the as-received condition, wires from both sources had room-temperature ultimate tensile strengths of 585,000 to 590,000 psi with approximately 2 pct total elongation and 40 pct reduction in area. Composite preparation was accomplished by the vacuum infiltration technique. This involved placing the continuous wires in a mild steel tube (1/8-in. ID), capping one end of the tube with a sheet of the matrix alloy, and evacuating the other end of the tube. When the capped end was placed in a crucible of the molten matrix alloy and held for 31/2 min at 70O°C, the liquid metal was drawn up the tube approximately 7 to 8 in. A cross-sectional view of a typical composite is shown in Fig. 1. This preparation technique promoted excellent bonding between the matrix and wires, and porosity was rarely observed. Test specimens for both creep and tension studies were prepared by machining off the steel tube. Tensile tests on the wires and composite specimens were made in an Instron using a cross-head speed of 0.01 in. per min. The composite specimens had no reduced section and were held in self-tightening grooved serrated grips. Most of the specimens broke in the effective 1-in. gage section. Even though several tests fractured in the grip section, their ultimate strengths were essentially the same as those of specimens which fractured in the gage section. Creep specimens of the composite were gripped in a similar manner, and

Jan 1, 1970

By Y. H. Liu, G. W. Powell

Thin foils made from the gold-rich region of- short (0.15 mm)Au-Ag diffusion couples annealed at 880°C for 1 hr were examined in ax electron microscope and the structures were compared with that of pure gold subjected to identical heat treatments. The couples underwent Plastic deformation (bending) during the diffusional anneal. The foils prepared from the diffusion zone have a higher dislocation density and incidence of subboundaries than pure gold. The plastic bending of the couple can be interpreted in terms of the climb of edge dislocations, which are pavallel to the bend axis, along the direction of diffusion. THE interaction of lattice defects during interdiffu-sion between two metals has been considered by several investigators. In particular, da Silva and Mehl,1 Barnes,' and Bardeen and Herring3 have proposed models based on interactions between point and line defects to account for the Kirkendall effect (marker movement). Each of these models is based on dislocation climb, which can either add or remove lattice planes, induced by an under- or supersaturation of vacancies. A theoretical treatment of the nucleation and growth of voids by the condensation of vacancies has been presented by Seitz.4 Stresses can arise in a binary diffusion couple as a result of the vacancy flux and the gradient in the lattice parameter. rinkmman' has investigated the effects of diffusion-induced stresses on void formation; the results of his analysis suggest that voids can form and grow under the influence of these stresses in the absence of an excess vacancy concentration. The relationships between vacancy supersaturation and diffusion-induced stresses and void formation have been elucidated also by Balluffi and seigle6 who concluded that the stress generated by a vacancy supersaturation aids the growth of voids. Experimental evidence which indirectly shows that stresses are produced by chemical diffusion has been obtained by several researchers.2,7-9 This evidence consists in part of the observation by light microscopy of the formation of subgrains, a process characteristic of the high-temperature deformation of some metals during creep.'0 More striking, however, is the plastic bending of short (0.2 mm) Cu-Ni couples observed by Barnes2 and of boron-doped silicon crystals reported by Queisser.9 Diffusion-induced dislocations in iron and silicon have been investigated by Ainslie et a1.11and Schwutte and Queisser,12 respectively, the former using transmission electron microscopy and the latter an X-ray transmission technique. Matthews and Crawford13 observed the formation of boundaries by the climb of interfacial misfit dislocations in thin, vapor-deposited Au-Pd couples diffused in situ in an electron microscope. In order to obtain additional information on the phenomena associated with defect interaction and diffusion-induced stresses in binary couples, an electron transmission study of foils prepared from the diffusion zone of short Au-Ag couples was undertaken. The initial objective was to investigate the processes occurring at various points of the diffusion zone. Subsequent research demonstrated that foils could be prepared only from the gold-rich region of the diffusion zone using the technique described herein. EXPERIMENTAL PROCEDURE Preparation and Annealing of the Diffusion Couples. The couples were made from annealed, high-purity (99.999 pct). polycrvstalline gold and silver sheets (0.075 mm thick). Shall (0.075 by 5 by 17 mm) elec-tropolished coupons of gold and silver were pressure-bonded by heating the two metals in a graphite device to 770°C and holding at this temperature for approximately 30 min. The quality of the bond was ascertained by metallographic examination at 150 times. Each couple subsequently was annealed at 880°C for 1hr; under these conditions, the extreme portions of the

Jan 1, 1968

By T. F. Mueller, P. A. Witherspoon

For the case of an infinite radial system operating at constant terminal rate, the reservoir engineer often uses the "point source" solution of the diffiusivity equation to study pressure interference effects. At early times and at short distances from the inner boundary these solutions are invalid. The amount of error is often not precisely defined because of mathematical difficulties. Work presented here shows that, if dimensionless time is defined appropriately, previous solutions of the pressure equation can be displayed .as a family of curves on one chart. These solutions include the point source solution (referred to in the field of hydrology as the Theis solution) and other solutions obtained with digital computer methods. With these curves, an exact evaluation of the pressure drop within a reservoir or an aquifer can be made by the engineer. Examples of field problem solutions are presented. In most reservoirs the error involved when the Theis solution is employed is often negligible; whereas, in the calculation of interference effects in an aquifer, a substantial error can occur through such an approach. INTRODUCTION Flow equations are used in petroleum engineering to study the behavior of individual wells and reservoirs. In the case of wells, the pressure response at the wellbore face is the major point of interest; whereas, in the case of reservoirs, the pressure response at the interface of the aquifer boundary is sought. To aid in such studies, the flow equations have been solved in terms of the behavior at these two inner boundaries. Only limited work has been published in regard to the pressure conditions away from these points, i.e., within the reservoir or aquifer. Theis and Mortada are among the few who have reported on this problem. The Theis approach employs the exponential integral and is valid for pressure conditions that occur some distance away from the flow disturbance. It is derived from the concept of a point source, as opposed to a flow across some finite area. The Mortada results, on the other hand, are valid at all points within the reservoir or aquifer. They are presented in terms of dimension less ratios of the radius where the pressure is desired to the radius where the flow rate is measured. Their main use, in the past, has been in aquifer studies. The published results are presented in the form of graphs that are limited to a maximum radius ratio of 64. These graphical results are cumbersome to interpolate at non-integral radius ratios, so that one may be forced to utilize a rather involved analytical expression presented by Mortada. BASIC EQUATIONS The solutions of Mortada and Theis are both based on the diffusivity equation as applied to the case of an infinite radial system subject to a constant terminal rate. The equation is obtained by combining the material balance equation with Darcy's flow equation. The assumptions implicit in the use of this equation are as follows: (1) a single fluid is present that occupies the entire pore volume; (2) the reservoir is horizontal, homogeneous, uniform in thickness, and of infinite radial extent; (3) compressibility and viscosity of the fluid remain constant at all pressures; and (4) fluid density obeys the equation Using the diffusivity equation in situations where the above conditions do not hold will result in errors. These errors (not discussed here) but only the errors which arise in the solution of the equation itself. The diffusivity equation for the homogeneous reservoir conditions cited above can be written in cylindrical coordinates, as To obtain a dimensionless equation, so a single solution can be used for applications of different porosity, permeability and fluid properties, the following transformations are usually made: After these transformations are made, Eq. 3 .can be written in dimensionless form, as MORTADA SOLUTION One solution of Eq. 6 has been given by Mortada.

Jan 1, 1966

By G. H. Bruce, C. E. Williams

The trend toward deeper drilling, together with the attcndant increase in power requirements for circulation of the drilling fluid, has emphasized the need for a critical examination of the factors affecting the removal of bit cuttings from the hole by the drilling fluid. The ability of drilling fluids to lift cuttings is called their carrying capacity. A series of laboratory and field experiments has been conducted to determine the minimrim annular velocity necessary to remove cuttings, and to investigate the effects of properties of drilling fluids on their carrying capacities. Consideration of the results of these experiments led to the following conclusions: 1. Turbulent flow in the well annulus is most desirable from the standpoint of cutting removal. 2. Low viscosity and low gel are advantageous in removing cuttings. 3. Increase in mud weight is effective in increasing carrying capacity. 4. The carrying capacity is higher when the pipe is rotated than when it is not. 5. If turbulent flow can be maintained, an annular velocity slightly higher than the slip velocity of the largest cuttings to be transported should keep the bore hole clean. This implies velocities of 100 to 125 ft per minute rather than the presently used 175 to 225 ft per minute. INTRODUCTION Power Savings by Reduction of Annular Velocities A large portion of the power expended in drilling operations is consumed in circulating the drilling fluid. An important factor in establishing the rate of mud circulation is the minimum velocity in the annulus necessary to remove bit cuttings. Empirically, it has been found that average annular mud velocities of about 200 ft per minute will remove cuttings. It was not definitely known, however, whether annular velocities of about 200 ft per minute were just above the minimum necessary to remove cuttings, or whether such velocities could be materially reduced without sacrifice of the ability of the mud to remove cuttings. It is apparent that if annular velocities could be reduced without impairment of cutting removal, a considerable saving in power requirements would result. Need for Research on Carrying Capacity The ability of a drilling fluid to transport cuttings is called its carrying capacity. Although it has been recognized that the carrying capacity of a mud is affected by mud properties such as viscosity and density1,2,3,4 there have been various views in the industry as to the effects of these mud properties on carrying capacity. The economic importance of the problem of carrying capacity and the scarcity of information on the subject indicated that research on the problem was needed. THE FACTORS AFFECTING CARRYING CAPACITY Qualitative Determination of Factors The mechanism of cutting transport is closely related to that involved in the separation of material by settling processes. A considerable amount of research has been done on settling problems, and discussions of sedimentation theory can be found in standard texts.5,6 Consideration of the information available from these sources, together with consideration of the mechanism of cutting transport, leads to the conclusion that the factors affecting carrying capacity are the dimensions of the system, the physical properties of the cuttings. and the physical properties of the drilling fluid. system Dimensions and Their Effect The dimensions of the fluid circulating system of importance to carrying capacity are the bore hole size, drill pipe size, pump capacity, and pump speed. These dimensions determine the annular velocity of the drilling fluid. Physical Properties and Their Effects The physical properties involved in the interaction between mud and cuttings are the density and shape of the cuttings and the density, viscosity and gel strength of the drilling fluid. The effect of the density factor on carrying capacity is fairly obvious; high density difference between cuttings and fluid results in a low buoyant force and therefore decreases carrying capacity. The effect of cutting shape is less obvious. Although

Jan 1, 1951

By C. E. Williams, G. H. Bruce

The trend toward deeper drilling, together with the attcndant increase in power requirements for circulation of the drilling fluid, has emphasized the need for a critical examination of the factors affecting the removal of bit cuttings from the hole by the drilling fluid. The ability of drilling fluids to lift cuttings is called their carrying capacity. A series of laboratory and field experiments has been conducted to determine the minimrim annular velocity necessary to remove cuttings, and to investigate the effects of properties of drilling fluids on their carrying capacities. Consideration of the results of these experiments led to the following conclusions: 1. Turbulent flow in the well annulus is most desirable from the standpoint of cutting removal. 2. Low viscosity and low gel are advantageous in removing cuttings. 3. Increase in mud weight is effective in increasing carrying capacity. 4. The carrying capacity is higher when the pipe is rotated than when it is not. 5. If turbulent flow can be maintained, an annular velocity slightly higher than the slip velocity of the largest cuttings to be transported should keep the bore hole clean. This implies velocities of 100 to 125 ft per minute rather than the presently used 175 to 225 ft per minute. INTRODUCTION Power Savings by Reduction of Annular Velocities A large portion of the power expended in drilling operations is consumed in circulating the drilling fluid. An important factor in establishing the rate of mud circulation is the minimum velocity in the annulus necessary to remove bit cuttings. Empirically, it has been found that average annular mud velocities of about 200 ft per minute will remove cuttings. It was not definitely known, however, whether annular velocities of about 200 ft per minute were just above the minimum necessary to remove cuttings, or whether such velocities could be materially reduced without sacrifice of the ability of the mud to remove cuttings. It is apparent that if annular velocities could be reduced without impairment of cutting removal, a considerable saving in power requirements would result. Need for Research on Carrying Capacity The ability of a drilling fluid to transport cuttings is called its carrying capacity. Although it has been recognized that the carrying capacity of a mud is affected by mud properties such as viscosity and density1,2,3,4 there have been various views in the industry as to the effects of these mud properties on carrying capacity. The economic importance of the problem of carrying capacity and the scarcity of information on the subject indicated that research on the problem was needed. THE FACTORS AFFECTING CARRYING CAPACITY Qualitative Determination of Factors The mechanism of cutting transport is closely related to that involved in the separation of material by settling processes. A considerable amount of research has been done on settling problems, and discussions of sedimentation theory can be found in standard texts.5,6 Consideration of the information available from these sources, together with consideration of the mechanism of cutting transport, leads to the conclusion that the factors affecting carrying capacity are the dimensions of the system, the physical properties of the cuttings. and the physical properties of the drilling fluid. system Dimensions and Their Effect The dimensions of the fluid circulating system of importance to carrying capacity are the bore hole size, drill pipe size, pump capacity, and pump speed. These dimensions determine the annular velocity of the drilling fluid. Physical Properties and Their Effects The physical properties involved in the interaction between mud and cuttings are the density and shape of the cuttings and the density, viscosity and gel strength of the drilling fluid. The effect of the density factor on carrying capacity is fairly obvious; high density difference between cuttings and fluid results in a low buoyant force and therefore decreases carrying capacity. The effect of cutting shape is less obvious. Although

Jan 1, 1951

By Hiroshi Kato, Shin’ichi Nagashima, Hiroshi Takechi

The preferred orientations detleloped during cold rolling, annealing, and hot rolling of low-carbon steel sheets have been investigated by means of an inverse pole figure technique. And X-ray line profile meas-urerrlents were made for the study of the factor which causes the textural changes between cold-rolled and recrystallized sheets. From ND (sheet normal divection) and RD (rolling direction) inverse Pole figures tlze axis detisity change with increasing cold reduction was measured and the Preferred orientations were analyzed. These results can be explained by crystal rotation due to slip during rolling deformation. The ND axis density change during annealing of cold-rollitig sheets is contrary to the change observed in the case of rolling. It was found that the orientation dependency of stored strain energy resulting from cold rolling apparently corresponds to the textural changes during recrystallization. The texture in hot-rolled sheets was found to be influenced by finishing tenlperatltre. In a sheet kol-rolled below the Ar, temperature a significant differernce in texture between the surface and central layers was obtained, the former having {hkl}(111} and {110 )(uvw) orientations and the latter having {pqr}( 110) orientation. A large number of investigations1 6 have been made on the textures of iron and low-carbon steel sheets, and, since it was shown7 13 that preferred orientations are closely related to deep drawability of steel sheets, studies14" 7 on the influence of production processes upon textures have been made from the standpoint of industrial interest. Generally, the textures in low-carbon steel sheets have several components of preferred orientations, but the overlapping of poles makes it difficult to analyze texture quantitatively from pole figures. On the other hand, an inverse pole figure directly expresses quantitative values of components of preferred orientations as the density of a certain sample axis with respect to the crystallite coordinates. Such a texture representation is desirable for a study of deep drawability. The purpose of the present work is to use both inverse and conventional pole figures, with emphasis on the former, to clar~fy the textural changes ~n low-carbon steel sheets by the production processes of hot rolling, cold rolling, and annealing. And the factor causing the textural changes during annealing is studied using an X-ray diffraction line profile technique. EXPERIMENTAL PROCEDURE Sheet specimens for study of cold rolling and annealing textures were taken from a hot-rolled strip of rimmed steel 2.7 mm thick. The chemical composition is shown in Table I. Although the study of cold-rolling textures requires a suitable material with random orientation, it is difficult to produce a perfect random one. But the used specimens, obtained from the hot-rolled strip with a finishing temperature higher than Ar3, were nearly random, such as shown in Fig. 15. These were cold-rolled to various reductions by a 4-high rolling mill with 150-mm-diam work rolls and 300-mm-diam backup rolls. The influence of hot-rolling conditions such as finishing and coiling temperatures on textures of hot-rolled strips was studied. The chemical compositions of these hot coils are given in Table I1 and the hot-rolling conditions are given in Table 111. inverse Pole ~i~ure -~easurement. Two methods for determining inverse pole figures with high precision have been proposed. The first, by Jetter et a1.,18 consists of obtaining mean densities of divided regions in a unit stereographic triangle by solving simultaneous linear equations to approximate the integral relation between conventional and inverse pole figures. The second, proposed by ~oe'~-'~ and Bunge,22'23 consists of determining an expanded series of spherical surface harmonics which express axis densities from several sets of pole distribution data. Both methods require laborious numerical calculations which make them unsuitable for practical use. However, the axis densities of the locations corresponding to low Miller indices in a unit stereographic triangle can be obtained

Jan 1, 1969

By A. W. D. Hills, M. R. Moore

An apparatus has been developed in which metals can bc solidified under closely controlled unidirectional heat flow conditions. The metals are cooled by a multiple-air jet system, designed so that the heat transfer coefficient is uniformly distributed over the cooled surface. They are contained in a special furnace system of low thermal mass which enables unidirectional heat flow conditions to be rigorously maintained for certain specific cooling rates. This apparatus has been used to study the solidijication of pure lead and pure tin at zero superheat, solid metal thicknesses being determined by thermal analysis, and by dipstick readings. Detailed thermal balances have been drawn up on the experiments in order to determzne the cooling conditions that most rigorously correspond to unidirectional heat flow. Solid metal thicknesses, measured experimentally for these conditions, have beetz shown to agree closely with the theoretical wedictzons of different integml-profile methods. lt is concluded that any errors inherent in the integval-profile method are no greater than the errors inherent in experimental investigations of solidification. A knowledge of the influence of heat transfer rates on the solidification of metals is essential to the understanding and control of industrial casting processes. Since very few formal methods are available for solving the heat conduction equation under solidification conditions several authors have developed approximate techniques for doing so. One of the most promising is the integral-profile technique that has been used by Goodman,1 Hills,2,3 Tien4 and Hills and ~oore' to predict solidification rates under certain specific heat transfer coniditions. In the first of this present series of papers6 Hills has further generalized the method and developed a universal integral-profile technique that can be used to predict solidification rates under a wide variety of unidirectional heat flow conditions. Since the method is approximate it is necessary to investigate the accuracy with which it predicts actual solidification rates. A review of the literature7 shows that solidification rates have seldom been measured under accurately known heat transfer conditions and it was therefore decided to develop experimental techniques that would enable this to be done. Results obtained would then be available as a check on the approximate solutions already developed, and in the development of further solutions that would be applicable under more complicated conditions. The initial work in this experimental program which is described here, is concerned with the solidification of pure metals at zero superheat under unidirectional heat flow conditions. Subsequent experiments have been concerned with the solidification of pure metals at appreciable superheats, under varying unidirectional heat flow conditions and under two dimensional heat flow conditions. In a series of experiments recently initiated, the solidification of binary alloys is also being investigated. The results of these subsequent experiments, together with relevant theory, will be described in later papers. The solidifying metal has been cooled in all the experiments, by a multiple air jet system designed so that the heat transfer coefficient to the cooled surface is constant and is uniformly distributed over the surface. This type of system has been chosen since only one parameter is needed to specify the heat transfer conditions and since these conditions are analogous to those met in certain continuous casting processes.' Air cooling was chosen in preference to water cooling, however, since it offers a wider range of heat transfer conditions, and since it can be investigated by analogous mass transfer techniques. The first part of this paper discusses the existing integral-profile solutions that correspond to the heat transfer conditions used in the experiments, and describes the development and use of the experimental apparatus. In subsequent sections the heat transfer and solidification results are presented, and the various theoretical predictions are compared with them, and with one another.

Jan 1, 1970

By J. A. Stachura

Continuous mining machinery provides the coal industry with one way to compete for a larger share of the total energy market. Various types of machines are discussed and some of the problems with continuous miners, encountered by operators, are reviewed. Equipment manufacturers are working with mine personnel to provide solutions for problems that arise. While coal production over the past 25 or 30 years has been on a horizontal plane, coal's share of the total energy market has declined. To participate more effectively in this total energy market, it is necessary to produce coal more efficiently. It is the obligation of all management, employes, and mining departments to gear the deep mining industry to the rapid progress and changing of today's modern industry. This can be accomplished in the near future with the selection of the proper type of continuous miner best suited to each operator's individual situation. In most mining operations there is tremendous incentive to undertake the continuous mining program. It can reduce the size of the mine greatly by permitting a minimum of working places; it makes pillar recovery work more efficient from the standpoint of overall cost, amount of coal recovered, and safety. The work force can be reduced materially permitting closer and more efficient supervision. It simplifies maintenance because equipment can be more readily standardized. The trend of the coal market favors the use of continuous mining machines. Although there appears to be a general feeling that continuous mining is still a relatively new program and will be slow in replacing conventional mechanical equipment, the fact is that tremendous strides have been made since the first machines were installed in 1948. This program is advancing at approximately the same rate that mobile loading machines replaced hand loading. From 1948 to 1955 there were approximately 450 continuous mining machines in service. In October 1959, a survey revealed that there were more than 700 continuous mining machines in service. Many operators have expressed a desire to undertake this program, but they feel that they could not do so at this time because of one or more of the following reasons: 1) the thickness of their coal seams, 2) seam characteristics, 3) soft bottoms, 4) bad roof conditions, 5) size consist, 6) insufficient flexibility in machines, 7) difficult ventilation problems, and 8) high maintenance costs. With the realization on coal about the same today as it was in 1948, or slightly less and since coal is still failing to participate to a greater degree in the total energy market, it is not surprising that the coal industry is desperately exploring more economical methods for deep mining. The manufacturers are aware that the coal industry is willing to invest in continuous miners if the equipment is built for maximum flexibility, will produce higher tons per man, and assure long life between overhaul programs. CONTINUOUS MINING MACHINES Before discussing details regarding the selection of a continuous miner, let us have a preview of some of the continuous mining machines which are available to the coal industry today. Jeffrey Manufacturing Co.: The machine shown is the Jeffrey 76 A.M. Colmol. This is their most widely used miner, and has been particularly successful in central Pennsylvania and in high-wall mining in western Kentucky. One of the outstanding features of this auger-type miner is its portability. The entire mining range can be changed from its lowest point to the maximum height without stopping the mining operation. Jeffrey 76 B.M. Colmol: This machine is similar to the 76 A.M. model; however, it is built bigger and stronger for a mining range of 50 1/2 to 72 in. This is the model that is now available (Fig. 2). Jeffrey has added two arms to the top row, omitted the odd arm in the bottom row, thus permitting a 50 pct larger throat opening. This eliminates one

Jan 1, 1961

By J. S. Browning

The U.S. Bureau of Mines has been experimenting with flotation processes to separate the spodumene-beryl ores mined at Kings Mountain, N.C. The success to date as well as the present status of the process is discussed. Further experimentation is underway and studies will be made on the milling costs involved in the flotation process. The pegmatites of the Kings Mountain-Lincoln ton, N.C., area constitute the largest known domestic reserve of beryl and spodumene. The reserve is estimated to contain 90 million tons of pegmatic material with 1,280,000 tons of recoverable Li,O(lithia) as spodurnene.' The pegmatites also contain 0.4 to 0.5 pct beryl disseminated throughout the orebodies. The pegmatites may contain a potential reserve of 240,-000 tons beryl, equivalent to approximately 34,000 tons of BeO. Different flotation methods for separating beryl from feldspar or quartz, or both, have been developed by various investigators.2-7 On the other hand, published information is limited on concentration of spodumene-beryl ores. As the response of beryl and spodumene to flotation is essentially the same, a successful separation of the two minerals depends upon use of a selective depressant or a selective collector for one of the minerals. Several years ago, the U.S. Bureau of Mines, recognizing the need for developing a large domestic supply of strategic beryl, undertook studies to develop improved methods of recovering this mineral. Because of the low beryl content of the pegmatites, economic production of the beryl entails its recovery as a byproduct of spodumene flotation. Thus, the development of procedures that would provide for maximum recoveries of the spodumene, as well as the beryl, was a primary objective of the studies. Tests were made on run-of-mine ore and spodumene flotation tailing from the Foote Mineral Co.'s spodumene concentrator at Kings Mountain (Fig. 1). Pet-rographic analyses of the ore and flotation tailing are given in Table I. Detailed analyses and examinations of the ore and tailings revealed that the beryllium was present in the form of a low-alkali beryl. The crystals were small, rarely larger than 14-mesh diam, and were clear and colorless. About 10 pct of the Li20 content of the ore was present in the mica and feldspar components of the pegmatite. In some weathered ores the associated clays also contained small amounts of lithia. This is typical of pegmatites in the Kings Mountain area. LABORATORY TESTS In the past, a vigorous chemical treatment of the spodumene surfaces with acid or caustic soda has been considered essential for satisfactory selective flotation of the spodumene from the other minerals in pegmatites. Many tests were made during this investigation to determine if chemical treatment of the spodumene could be eliminated in favor of a reagent combination that would depress the other minerals while selectively floating the spodumene. A simple reagent combination and process was developed that gave better spodumene recovery and grade of concentrate than the more elaborate acid or caustic treatment methods previously used. The finely ground ore pulps were conditioned with either an ammonium, alkali, or alkaline earth lignin sulfonate and sodium fluoride, and the spodumene was floated and cleaned using oleic acid as the collector. The spodumene tailing was deslimed, conditioned with sulfuric acid and coco amine acetate, and floated to reject mica. The mica tailing was thickened and conditioned with hydrofluoric acid, then washed to remove the acid. The washed pulp was conditioned with sodium hydroxide for pH control, and the beryl was floated and cleaned with oleic acid as the collector. Table II summarizes the results of the laboratory test work. About 77 pct of the spodumene and 75 pct of the beryl were recovered from the pegmatitic material. The spodumene concentrate assayed 6.1 pct Li20 and 0.01 pct BeO, and the beryl product assayed 3.0 pct Li20 and 1.57 pct BeO. Similar results were ob-

Jan 1, 1961

By E. J. Harris

Mathematical solutions to complex mine ventilation problems are possible, but often the airway network is so complex that the mathematical solution becomes tedious and impractical. A fluid network analyzer, designed and built for analyzing mine ventilation problems has been in service at the Bureau of Mines' Pittsburgh station for approximately nine years. Using this analogue as a model, the mine ventilation network is simulated electrically by a combination of series and parallel circruits laid out to conform with actual mine airways. A tungsten filament lamp, referred to as a Fluistor, is used to simulate mine airway resistance. With the analogue ventilation model established, voltages of the proper amplitude to represent ventilating pressures are impressed across the circuit at points where mine fans and airshafts are located. Comparison of mathematical solutions of complex systems against analogue results showed a maximum variance of 3% for pressures and 2% for quantities. An electrical fluid network analyzer, designed and built especially for analyzing mine ventilation distribution problems, has been in service at the Bureau of Mines, Pittsburgh, for approximately nine years. It is a nonlinear, low voltage, fluid network analyzer of the type developed by the late Malcolm S. McIlroy, Professor of Electrical Engineering, Cornell University, who cooperated with G. E. McElroy, of the Bureau of Mines, in adapting the instrument to mine ventilation systems. Several modifications have been made since the original installation, but a considerable part of this paper is drawn from G. E. McElroy's original description of the analogue.' The analysis of water or gas distribution systems2,3 led to the development of this type of network analyzer. Several similar units are now employed by utility companies for this purpose; however, the Bureau of Mines unit is the only one designed specifically for mine airflow problems. Other airflow analogues employing a similar principle have been used at the Central Research Station of the Netherlands State Mines,4 in England,5 and in South Africa.6 Information on these devices indicated they were somewhat inflexible because commercial lamps with a sufficient range of resistance are difficult to obtain. Computers for mine ventilation analysis have been developed in Germany which instead of lamps use a variable resistance to adjust for turbulent flow laws. These units are expensive, but excellent results have been reported in their application. THE ANALYZER Theory of Application: As airflow generally follows the law of turbulent fluid flow, resistance to flow is nonlinear; consequently, the problem has been to find a nonlinear resistance element of suitable range that can be used for electrically simulated airflow. Tungsten filament lamps operated on alternating or direct current approximate the square-law resistance characteristics of mine airflow over a large range below maximum or rated voltage; that is, the voltage drop varies approximately as the square of the current. Consequently, the heart of the network analyzer is a nonlinear resistor known as a Fluistor, which is simply a custom-made low voltage, tungsten filament lamp that is available in a progressive series of relative resistance values ranging from 0.05 to 500 in nominal 5% steps. However, variations in manufacturing large groups result in differences of 1 to 3%, but series arrangements required for high-loss branches can be matched within about 1%. Utilizing a combination of Fluistors and load circuits, the mine ventilation system is duplicated electrically. Intake load circuits are connected from power intake to primary points of the circuit network; segments of unregulated flow along intakes and returns are represented by Fluistors, regulated splits and leakage paths are represented by load circuits and Fluistors of proper capacity; and mine exhausts are connected to ground from the last point of the network to complete the circuit. For the special purpose of representing booster fans or natural draft conditions, boosters or separate source circuits are provided that can be inserted between any two points of a network to increase voltage to the required value. Physical Layout: The analyzer consists of three 42

Jan 1, 1963

By E. I. Alessandrini, A. U. Seybolt

Selective oxidation (preferential oxidation of a single alloy component) can be an important oxidation mode in the case of alloys such as Si-Fe, Al-Fe, and so on, when heat-treating in atmospheres of low oxygen potential. It was the purpose of this work to identify the oxide films formed on iron containing 31/4 pct Si at low oxygen potentials, and to make some observations on the H2O/H2 ratios required to form them. EXPERIMENTAL METHOD HYDROGEN of varying water content offers a very convenient method of establishing a wide range of oxidizing power, or oxygen potential. The method chosen to provide a desired H2O/H2 ratio was to nearly saturate hydrogen at room temperature by bubbling the gas through water at room temperature, and then to wring out the excess water by passing the wet gas through a long glass spiral immersed in a refrigerated liquid at a desired dew point. After oxidation, the samples were examined by electron diffraction in order to identify any film which might be present. PREPARATION AND TREATMENT OF SAMPLES The samples were prepared from a Si-Fe alloy made in the laboratory using electrolytic iron and 96 pct ferro-silico11, vacuum-induction melted, hot-rolled then cold-rolled to 0.015-in. thickness. After the cold-rolling the samples were annealed at 1000°C. Transmission Laue X-ray patterns showed very little evidence of preferred orientation, but it is likely that a weak texture could have escaped notice. The chemical analysis was as follows: 3.34 Si, 0.01 Al, 0.001 0, 0.006 C, 0.002 N. Before each sample was oxidized, it was polished through 4/0 emery paper, cleaned and dried using benzene, carbon tetrachloride, acetone, absolute alcohol, and ether. The method of oxidizing the samples was as follows: Samples about 1/2 by 1 by 0.015 in. were held in the upper, cool part of a vertical vacuum furnace until the desired temperature was established in the furnace, Fig. 1. The sample was then lowered into the uniform hot zone where it was heated quickly to the desired temperature, and subjected to hydrogen of controlled moisture content (dew point) as mentioned previously. Temperatures from 400" to 1000°C were used, and dew points from room temperature to —78°C were explored. In general, the runs lasted about 6 hr, but a few runs were of different duration, as will be indicated later. After the predetermined period of time had elapsed, the specimens were withdrawn by the hollow stainless steel-quartz holder to a cool zone at the top of the furnace. EXPERIMENTAL RESULTS The results are summarized in Table I. Most of the electron-diffraction patterns were obtained by glancing-angle reflection, but in a few cases it was possible to strip the oxide films from the surface by a total immersion in a 3 pct bromine-methanol solution. Some of the thinner stripped films yielded quite satisfactory transmission patterns. However, both diffraction techniques yielded the same results; the patterns while most frequently showing a (low) tridymite, also showed a (low) cristobalite in some cases. A few patterns showed evidence of both phases. Because of the known nonequilibrium behavior of silica, it is not surprising that both forms of silica could coexist on one sample. It was noted that at fairly low dew points such as —40°C, there was a tendency for diffuse diffraction rings, suggesting a very small crystallite size. Also some subsequent experiments to be reported later show that thin silica films can be amorphous.

Jan 1, 1959

By David E. Nicholas

INTRODUCTION In this paper, a numerical process for selecting a mining method, with the emphasis on underground mass mining techniques, such as caving, induced caving, and stoping, is proposed. In the past, selection of a mining method for a new property was based primarily on operating experience at similar type deposits and on methods already in use in the district of the deposit. Then, the chosen method was modified during the early years of mining as ground conditions and ore character were bet- ter understood. Today, however, the large capital investment required to open a new mine or change an existing mining system make it imperative that the mining methods examined during the feasibility studies and the method actually selected have a high probability of attaining the projected production rates. Although experience and engineering judgment still provide major input into the selection of a mining method, subtle differences in the characteristics of each deposit, which may affect the method chosen or the mine design, can usually be perceived only through analysis of measured characteristics. The parameters that must be examined when choosing a mining method include: 1) geometry and grade distribution of the deposit; 2) rock mass strength for the ore zone, the hanging wall, and the footwall; 3) mining costs and capitalization requirements; 4) mining rate; 5) type and availability of labor; 6) environmental concerns; and 7) other site-specific considerations. This paper encompasses a detailed look at the first two parameters since they, plus mining costs, have the greatest impact on the selection of a mining method. The proposed method selection process is for a project where drilling has defined sufficient geologic reserves, but little or no underground development has been done. Since each deposit has its own characteristic geanetry/grade distribution, and rock mechanics properties, mining method selection should be at least a two-stage process. In Stage 1, the deposit is described in terms of geometry, grade distribution, and rock mechanics properties. Using these parameters, the mining methods can be ranked to determine which are most applicable; they can then be considered in general terms of mining and capitalization cost, mining rate, type and availability of personnel, environmental concerns, and other site-specific considerations. In Stage 2, the most likely mining methods are costed out, based on a general mine plan. Mining and capitalization costs are used to determine a cut-off grade fran which a minable reserve can be calculated; economic comparisons can then be made to determine the optimum mining method and economic feasibility. During the mine planning phase of Stage 2, rock mechanics information would be used to provide realistic estimates of underground opening size, amount of support, orientation of openings, and caving characteristics, and open pit slope angles. If ground control or operational problems should be encountered with the methods being considered, modifications could be made. Although planning on paper extends start-up time, it is cheaper to err on paper than to find the error after mining has begun. METHOD SELECTION - STAGE 1 The main purpose of Stage 1 is to select those mining methods which should be considered in greater detail. The simplest way to do this is by defining those characteristics required for each mining method and then determining whether the characteristics of the deposit are suitable. However, no one mining method is so restrictive that it can be used for only one set of characteristics, as indicated by the classification system proposed by Boshkov and Wright (1973). In the mining method selection proposed, geometry, grade distribution, and rock mechanics characteristics are ranked according to their acceptability for ten general mining methods.

Jan 1, 1981

By Milton F. Goodrich

The idea of using water-only cyclones to scalp clean coal from the feed to other devices has recently been gaining in popularity. 1,2,6 An indication of this popularity is that water-only cyclone scalping is already employed in a few commercial plants. Water-only cyclones scalp the feed to a heavy media vessel in the Warwick plant,3 to Deister tables in the Rowland plant,4 and to a heavy media cyclone circuit in the Amherst Coal Company's McGregor plant.5 Despite this trend, we found surprisingly little information in the literature on the performance of water-only cyclone scalping. Therefore, we undertook a study of its technical merits, limiting our investigation to a water-only cyclone scalper in conjunction with heavy-media cyclones. We felt this to be a reasonable combination of equipment, since both processes operate well on the same size coal. First, we developed simple mathematical relationships revealing general characteristics of the system. Next, we compared the clean coal yield and quality obtainable by the system on four different coals with that obtainable by heavymedia cyclone circuit alone and water-only cyclones middlings recirculation circuit alone. This was done by using computer program to simulate the operation of the three systems. Finally, we considered the effect that implementation of the system would have on a typical preparation plant circuit. The System The system chosen for study [(shown schematically in Fig. 1)] comprises two stages: a primary stage (PI) consisting of water-only cyclones, and a secondary stage (P2) consisting of heavy media cyclones. The water-only cyclones separate raw coal at a low specific gravity, producing a relatively small amount of high quality clean coal and a refuse containing much good coal. This refuse is then recleaned in the secondary stage, where the sharper separation of heavy media cyclones, operated at a higher specific gravity, reclaims additional clean coal. Finally, the clean coal from both stages is combined as the system's clean coal, while the refuse from the secondary stage becomes the system's refuse output. Theoretical Analysis Two assumptions lie behind the following theoretical analysis. The first assumption is that partition curves can represent not only the performance of each stage but also the performance of the system as a whole. These curves show partition number as a function of specific gravity, where partition number is either the percent or fraction of a raw coal with a given specific gravity reporting to clean coal. The second assumption is that these curves are independent of the raw coal characteristics. At any specific gravity the system's partition number is solely a function of the partition numbers of each stage. For convenience, we use the following variables in describing this relationship: g = specific gravity, PS(g) = the partition number of the system at specific gravity g, P1(g) = the partition number of the primary stage at specific gravity g, P2(g) = the partition number of the secondary stage at a specific gravity g. For this system, we can show algebraically that the relationship is: [Ps (g) = PI (g) + (1- P1(g)) P2 (9)(1)] With this equation the system's partition curve can be determined if the partition curves of each stage are known. By definition, the partition numbers P1(g) and P2(g) must each be less than, or equal to, one. Placing these constraints on Eq. (1), we arrive at the relations: [Ps(g) % P1(g) and Ps(g) % P2(g)(2)] Thus, at any specific gravity the system's partition number is always greater than that of both stages. [Figure 2] shows typical partition curves. Curve I represents the primary stage, which consists of water-only cyclones operating at a separating gravity of 1.3. (The separating gravity or the

Jan 1, 1979

By N. A. Warner

Dense cylindrical specimens of artificial hematite were reduced in hydrogen over a range 0-f total pressures between 0.1 and 1.0 atm and temperatures between 650" and 950°C. Hydrogen reduction at a total pressure of 1.0 atm in the presence of a chemically inert diluent and reduction in hydrogen/water mixtures were also studied Znter.face movement during reduction was followed using metallographic techniques. Substantiu1 partial-pressure gradients are shown to exist in both the gas film and the reduced iron layer. The results are consistent with a mixed-control mechanism involving an interaction of gaseous-difusion effects with a first-order reversible chemical reaction at the iron/wüstite interface. Over the hydrogen-pressure range of 0.1 to 1.0 atm, the rate is not directly proportional to the hydrogen pressure in the bulk gas stream. Nitrogen dilution of the reducing gas masks the true effect of hydrogen partial pressure by influencing the effective molecular dif-fusivity and gives an apparent first-order relationship. The transport of hydrogen and water vapor across the iron layer is consistent with molecular gaseous diffusion rather than a Knudsen diffusion mechanism. ALTHOUGH the literature on the reduction of iron oxides is very extensive, it is only comparatively recently that a definite quantitative model for the reduction of hematite has been proposed. This development has been largely due to the work of McKewan. An extensive literature review of previous work has recently been compiled by Themelis and Gauvin. Exclusive control by gaseous diffusion in the reduction of hematite has been recently advocated by Kawasaki and coworkers.' Although the results of the latter workers are of considerable practical significance, the extreme porosity (approximately 30 pct) of their specimens precludes their application to a kinetic evaluation of the reduction of dense polycrystalline hematite. The salient features of the kinetic model used by McKewan for the gaseous reduction of dense hematite at temperatures in excess of 560°C (when FeO is stable) are: 1) The reduction is a topochemical reaction taking place through 'the series FeO/FesO4/FeO/ Fe; 2) The gas-solid type of reaction takes place only at the FeO/Fe interface and the internal reduction: FeO proceeds by solid state diffusion; 3) The iron layer formed is quite porous and offers negligible impedance to gaseous diffusion; 4) The rate-controlling step in the reduction is a chemical process occurring at the Fe/FeO interface. In view of the relatively fast reaction rates involved when hematite is reduced in a stream of hydrogen, it seemed possible that the simplifying assumptions of negligible diffusional resistance in both the bulk gas phase and the reduced iron layer could introduce serious errors in the interpretation of experimental reduction data. The present work was initiated to elucidate the effect of gaseous diffusion on the reduction rate and to determine its possible influence in the formulation of the reaction mechanism at the metal/oxide interface. EXPERIMENTAL Procedure and Apparatus. Dense cylindrical compacts of artificial hematite were prepared by sintering pressed compacts of reagent-grade ferric oxide, containing 99.6 pct FezO, in air at 1200°C for 80 hr. The resulting compacts had a final porosity of less than 2 pct and a height and diameter of approximately 1 cm. Variations in sintering atmosphere from air to pure oxygen, in sintering time from 20 to 80 hr, and temperature from 1150" to 1250°C had no effect on the subsequent reduction behavior. The reduction of the samples took place in a vertical tube furnace in which was placed a 1-in. transparent silica tube. The 24-in. heated length

Jan 1, 1964

By Gerald Sherman

MINING asbestos ore by caving at Thetford Mines, Quebec, has been described in the Transactions of the Canadian Institute of Mining and Metallurgy in papers presented by the staffs of The Asbestos Corp. Ltd., in 1934, and Johnson's Co. Ltd., in 1941. The production by caving prior to 1949, in round figures, has been 14,500,000 tons of ore by The Asbestos Corp. and 5,500,000 by Johnson's Co. They are the only caving operations in asbestos at the present time. The Jeffrey mine of the Canadian Johns-Manville Co. at Asbestos, 40 miles southwest of Thetford, is well along in preparations for caving, and The Bell Asbestos Mines Ltd. at Thetford Mines, will follow their examples later. GERALD SHERMAN, Member AIME, is Consulting Engineer, Johnson's Co. and Canadian Johns-Manville Co., New York, N. Y. AIME San Francisco Meeting, Feb. 1949, and New York Meeting, Feb. 1950. TP 2828 A. Discussion (2 copies) may be sent to Transactions AIME before May 31, 1950. Manuscript received Nov. 2, 1949. Nearly all of the asbestos produced by the western hemisphere is mined in those districts of Canada. Since it has been the subject of a number of technical papers, this contribution is offered as a review of progress. Some introductory information seems appropriate, for few members of the American Institute of Mining and Metallurgical Engineers are familiar with the mineral, or how asbestos deposits differ from others in caving characteristics. Asbestos is found in peridotite, a basic intrusive rock irregularly altered to serpentine and erratically fractured, well enough to cave in bulk, but leaving many blocks of solid ore of all sizes scattered through the deposit. Chrysotile is the variety of asbestos produced in Canada. Its fibers are the most valuable commercially, having a good proportion of long fibers, strong, silky, flexible, and suitable for spinning. The fibers occur in all lengths, from the microscopic to 3 in., or more, the shorter are more abundant. Their values depend on length combined with other qualities noted above. They occur as veinlets filling fissures in the peridotite, with crystals compactly arranged in parallel across the cracks. They are sometimes lightly frozen to the walls but the partings are distinct, often marked by particles of magnetite adhering to the fiber ends. Sometimes they break away from the rock in slabs an inch or more in thickness. In wide fissures the fibers are not always continuous, the "broken fibers" being separated by lines of parting with occasional scattered grains of magnetite, which is an injurious mineral to be elirninated when possible. In addition to the chrysotile ''Cross fiber," some "slip fiber" is found in fault fractures in limited quantities of less valuable character. With the exception of the Asbestos Corp, ore, in which the content is said to be more uniform, asbestos is eccentrically distributed through the peridotite. It is the general impression that the higher grades are more often found in hard rock. There are rich pockets and leaner areas which range down through various grades to minor barren sections. Wholesale mining is necessary to make sure of recovering all the high grade for there are no reliable indications of its existence. Ore selection only appears in constructing assay-grade limits for caving. The mill yield of the Thetford ore body in 1934 was 6 to 10 pct of the ore milled, made up of all classes, with occasional runs up to 15 pct or so. Quotations for Canadian asbestos* are as follows: Per Ton No. 1 Crude (% in. and longer) $896 @ $960 No. 2 Crude (Y8 to % in.) 350 @ 545 Spinning (less than % in.) 204 @ 378 Other classes, price diminishing with length. Shorts or "refuse" 24 @ 46 • From Eng. and Min. Jnl., Nov. 1948. Asbestos of even superior quality for electric insulation is found in Arizona limestone. It is white, containing very little combined iron, but occurs only in small quantities.

Jan 1, 1951

By Gerald Sherman

MINING asbestos ore by caving at Thetford Mines, Quebec, has been described in the Transactions of the Canadian Institute of Mining and Metallurgy in papers presented by the staffs of The Asbestos Corp. Ltd., in 1934, and Johnson's Co. Ltd., in 1941. The production by caving prior to 1949, in round figures, has been 14,500,000 tons of ore by The Asbestos Corp. and 5,500,000 by Johnson's Co. They are the only caving operations in asbestos at the present time. The Jeffrey mine of the Canadian Johns-Manville Co. at Asbestos, 40 miles southwest of Thetford, is well along in preparations for caving, and The Bell Asbestos Mines Ltd. at Thetford Mines, will follow their examples later. GERALD SHERMAN, Member AIME, is Consulting Engineer, Johnson's Co. and Canadian Johns-Manville Co., New York, N. Y. AIME San Francisco Meeting, Feb. 1949, and New York Meeting, Feb. 1950. TP 2828 A. Discussion (2 copies) may be sent to Transactions AIME before May 31, 1950. Manuscript received Nov. 2, 1949. Nearly all of the asbestos produced by the western hemisphere is mined in those districts of Canada. Since it has been the subject of a number of technical papers, this contribution is offered as a review of progress. Some introductory information seems appropriate, for few members of the American Institute of Mining and Metallurgical Engineers are familiar with the mineral, or how asbestos deposits differ from others in caving characteristics. Asbestos is found in peridotite, a basic intrusive rock irregularly altered to serpentine and erratically fractured, well enough to cave in bulk, but leaving many blocks of solid ore of all sizes scattered through the deposit. Chrysotile is the variety of asbestos produced in Canada. Its fibers are the most valuable commercially, having a good proportion of long fibers, strong, silky, flexible, and suitable for spinning. The fibers occur in all lengths, from the microscopic to 3 in., or more, the shorter are more abundant. Their values depend on length combined with other qualities noted above. They occur as veinlets filling fissures in the peridotite, with crystals compactly arranged in parallel across the cracks. They are sometimes lightly frozen to the walls but the partings are distinct, often marked by particles of magnetite adhering to the fiber ends. Sometimes they break away from the rock in slabs an inch or more in thickness. In wide fissures the fibers are not always continuous, the "broken fibers" being separated by lines of parting with occasional scattered grains of magnetite, which is an injurious mineral to be elirninated when possible. In addition to the chrysotile ''Cross fiber," some "slip fiber" is found in fault fractures in limited quantities of less valuable character. With the exception of the Asbestos Corp, ore, in which the content is said to be more uniform, asbestos is eccentrically distributed through the peridotite. It is the general impression that the higher grades are more often found in hard rock. There are rich pockets and leaner areas which range down through various grades to minor barren sections. Wholesale mining is necessary to make sure of recovering all the high grade for there are no reliable indications of its existence. Ore selection only appears in constructing assay-grade limits for caving. The mill yield of the Thetford ore body in 1934 was 6 to 10 pct of the ore milled, made up of all classes, with occasional runs up to 15 pct or so. Quotations for Canadian asbestos* are as follows: Per Ton No. 1 Crude (% in. and longer) $896 @ $960 No. 2 Crude (Y8 to % in.) 350 @ 545 Spinning (less than % in.) 204 @ 378 Other classes, price diminishing with length. Shorts or "refuse" 24 @ 46 • From Eng. and Min. Jnl., Nov. 1948. Asbestos of even superior quality for electric insulation is found in Arizona limestone. It is white, containing very little combined iron, but occurs only in small quantities.

Jan 1, 1951