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By Laurence Delisle, Aaron Finger

THE alloys formed by the addition of nickel to iron by convelltional metallurgical procedures show physical properties that differ widely from those of the individual metals. The effect of alloying on the mechanical properties is most pronounced in alloys containing between 15 and 25 per cent nickel. The tensile strength of an alloy containing, for example, 20 per cent nickel and 80 per cent iron (with 0.06 per cent carbon), in an annealed condition, is above 150,000 lb. per sq. in.? owing to a martensitic type of transformation, as compared with about 50,000 lb. per sq. in. for pure iron in the same condition. The object of this work was to determine whether the same beneficial alloying effect could be obtained in alloys of iron and nickel prepared by powder metallurgy technique. PROCEDURE Bars and pellets for tension and compression tests, respectively, were made. Two groups of specimens were prepared: (I) by cold-pressing the powders and sintering; (2) by cold-pressing the powders, sintering, repressing and resintering. Raw Materials Nickel powder supplied by Metals Disintegrating co., marked MD-151, was used. A typical analysis of this powder was as follows: sulphur, 0.040 per cent; copper, 0.18; iron, 0.23; manganese, 0.002; magnesium, 0.001 5 ; lead, 0.035; tin, 0.025 ; silicon, 0.08; chromium, 0.008. Electrolytic iron powder supplied by Plastic Metals Inc. was used. A representative chemical analysis for impurities other than gases, supplied by the vendor, gave the following results: carbon, 0.005 per cent; manganese, 0.002; silicon, 0.003; phosphorus, 0.001; sulphur, 0.004; nickel, 0.008. The fractions from these powders passing a 3 2 5-mesh sieve were reduced in hydrogen at 600°C. for one hour, to clean the surfaces of the particles. The reduced powders, slightly caked, were ground in a mortar and screened again through a 325- mesh sieve. Mixing The proper proportions of the powders were mixed overnight on rolls, in glass jars fitted with iron-wire baffles to break aggregates of particles that might form. Pressing All specimens were compacted on hydraulic presses. The pellets were compacted in a cylindrical die, ½ in. diameter. The ratio of thickness to the diameter of the pellets was maintained as nearly as possible at 0.9, as specifed for standard short compression pieces. The tensile bars were made by compacting 30 grams of powder. Their shape is shown in Fig. I. The compacting pressures for the pellets and the tensile bars were:

Jan 1, 1946

By R. H. Brown, F. Keller

Because of the extensive use of Alclad 24s alloy sheet in aircraft construction, there is much interest in the metallurgical changes caused by heat-treatment of this product.1, 2 One of these changes is the result of the diffusion of soluble elements from the core into the coating during treatment. This paper deals with the thermal treatments that may influence diffusion and with the practical effects of diffusion on the serviceability of the product. As is well known, Alclad 24s-T sheet has a core of high-strength alloy combined with surface layers of relatively pure aluminum. The microstructure of a cross section of this product is shown by Fig. I. Not only does the relatively high-purity coating have a high resistance to corrosion, and hence serves as a mechanical barrier between the core metal and the corroding environment, but in addition this coating exerts electrochemical protection over areas where the core may be exposed, such as cut edges. In addition, the coating affords electrochemical protection to projecting heads of A17s-T, 17s-T and 24s-T rivets, or attached members made from these alloys. Under some conditions of heat-treatment, however, diffusion of copper into the coating will alter the electrochemical relationship between the coating and the core to the extent that only partial protection will be obtained under some corrosive conditions. The reason for the electrochemical protection lies in the fact that the coating layers of pure aluminum are anodic to the alloy core. The addition of copper to aluminum produces alloys in which the potential is lower than that of the pure metal. Moreover, as the amount of copper in solid solution in 24s alloy is increased, the potential of the alloy becomes more cathodic. When heat-treated at 920°F. and quenched in cold water, 24s-T is about 0.16 volt cathodic to pure aluminum. However, copper will diffuse from the core into the coating, and if it reaches the surface the potential will be altered in a cathodic direction. Thus the potential difference between the core and coating may be less than 0.16 volt, depending uponIthe thickness of the coating, and the time and temperature of heat-treatment. Of the methods available for determining the extent of diffusion of copper from the core into the coating of Alclad 24s-T, two are generally used; namely, the metallo-graphic and the solution potential methods.3 Metallographic Methods Metallographic methods will reveal the progress of diffusion of copper into the coating. For instance, the diffusion zone in the coating intermediate between the coating surface and the core is shown in Fig. I. This photomicrograph is of a polished and etched cross section of 0.040-in. Alclad 24s-T sheet. Diffusion proceeds into the coating more rapidly along the grain boundaries than through the body of the grains. Consequently, the diffusion zone is not of uniform width

Jan 1, 1944

By F. Keller, R. H. Brown

Because of the extensive use of Alclad 24s alloy sheet in aircraft construction, there is much interest in the metallurgical changes caused by heat-treatment of this product.1, 2 One of these changes is the result of the diffusion of soluble elements from the core into the coating during treatment. This paper deals with the thermal treatments that may influence diffusion and with the practical effects of diffusion on the serviceability of the product. As is well known, Alclad 24s-T sheet has a core of high-strength alloy combined with surface layers of relatively pure aluminum. The microstructure of a cross section of this product is shown by Fig. I. Not only does the relatively high-purity coating have a high resistance to corrosion, and hence serves as a mechanical barrier between the core metal and the corroding environment, but in addition this coating exerts electrochemical protection over areas where the core may be exposed, such as cut edges. In addition, the coating affords electrochemical protection to projecting heads of A17s-T, 17s-T and 24s-T rivets, or attached members made from these alloys. Under some conditions of heat-treatment, however, diffusion of copper into the coating will alter the electrochemical relationship between the coating and the core to the extent that only partial protection will be obtained under some corrosive conditions. The reason for the electrochemical protection lies in the fact that the coating layers of pure aluminum are anodic to the alloy core. The addition of copper to aluminum produces alloys in which the potential is lower than that of the pure metal. Moreover, as the amount of copper in solid solution in 24s alloy is increased, the potential of the alloy becomes more cathodic. When heat-treated at 920°F. and quenched in cold water, 24s-T is about 0.16 volt cathodic to pure aluminum. However, copper will diffuse from the core into the coating, and if it reaches the surface the potential will be altered in a cathodic direction. Thus the potential difference between the core and coating may be less than 0.16 volt, depending uponIthe thickness of the coating, and the time and temperature of heat-treatment. Of the methods available for determining the extent of diffusion of copper from the core into the coating of Alclad 24s-T, two are generally used; namely, the metallo-graphic and the solution potential methods.3 Metallographic Methods Metallographic methods will reveal the progress of diffusion of copper into the coating. For instance, the diffusion zone in the coating intermediate between the coating surface and the core is shown in Fig. I. This photomicrograph is of a polished and etched cross section of 0.040-in. Alclad 24s-T sheet. Diffusion proceeds into the coating more rapidly along the grain boundaries than through the body of the grains. Consequently, the diffusion zone is not of uniform width

Jan 1, 1944

By Nathaniel Arbiter

THE separation of minerals by flotation can be regarded as a rate process, with the extraction of any one mineral determined by its flotation rate, and the grade of concentrate by the relative rates for all the minerals. So regarded, the significant variables for the process are those that control the rates. These variables are of two types, the first describing the ore and its physical and chemical treatment prior to flotation and the second characterizing the separation process in the cells. This paper will examine the variation in rates for a group of separations, will show that a simple rate law appears to govern, and will consider the relation of the control variables to the rates. The use of rate constants for evaluation of performance and efficiency will be discussed. Flotation involves the selective levitation of mineral and its transfer from cell to launder. The flotation rate is the rate of this transfer. It may be defined by the slope of a recovery-time curve for any cell in a bank, or at any time in batch operation. The objective in flotation rate study is an equation expressing the rate in terms of some measurable property of the pulp. This can be either the concentration of floatable mineral in weight per unit volume1,2 or a relative concentration, which will be a function of the recovery." A rate equation for an actual flotation pulp will contain at least two constants, both to be determined from the data. One of these, the initial concentration or proportion of floatable mineral, is not necessarily equal to the feed assay because of nonfloatable oversize or locked particles." The other, a rate constant, is a measure of proportionality between the rate and the pulp property on which the rate depends. The value of the rate constant will be determined by the values of all variables which control the process and will be changed by significant changes in any of them. It is, therefore, a direct measure of performance. Where recovery or grade change continuously with flotation time, the rate constant will be independent of time and will characterize the entire course of the separation. Development of Rate Equations Rate equations can be developed either by analysis of the mechanism of the process or by direct fitting of equations to recovery-time data. Sutherland's attempt by the first method' suggests that the effect of particle size variation on the rate complicates the derivation of a simple equation applicable to an ore pulp. A further problem with an ore is the concentrate grade requirement, which usually involves a variable rate of froth removal. Thus the final rate for any cell may depend on the froth character and froth height, as well as on the pulp composition. This does not imply that each cell cannot reach a steady state2 in which the rate will depend ultimately on pulp composition. The second method is the fitting of rate equations consistent with the necessary boundary conditions* to experimental recovery-time curves. On the assumption that under constant operating conditions the flotation rate is proportional to the actual or relative concentration of floatable mineral in the pulp, a generalized rate equation may be expressed as follows: Rate = Kcn [I] where K is the rate constant, c is some measure of the quantity of floatable mineral in the pulp at time t, and n is a positive number. In previous rate studies, the value of n has been taken as 1, either by direct assumption," or as a result of the hypothesis that bubble-particle collision is rate determining.' A first order equation results, which after integration in terms of cumulative recovery R, leads to Loge A/A-R = Kt [2] The quantity A is the maximum possible recovery with prolonged time under the conditions used. No conclusive proof for the validity of this equation in flotation has been advanced. The evidence cited in its support consists entirely in the demonstration that it appears to apply to a limited number of recovery-time curves."' It will be shown subsequently that this procedure is not sufficient to establish the order of a flotation rate equation. The possibility that the equation may be of higher order therefore requires examination. If, in particular, the exponent in eq 1 is assumed to be 2, then after integration there results R = A2Kt/1 + AKt [3] with K again a rate constant and A the maximum proportion of recoverable mineral. Eq 3 may be

Jan 1, 1952

By R. T. O’Brien, J. M. W. Mackenzie

A microelectrophoresis technique has been used to measure the zeta potential of quartz over a range of pH and Ni (11) and Co (Ilj concentrations. Results have been discussed in terms of adsorption of Ni (11) and Co (11) cations and hydroxide precipitates. Possible relationships between the experimental results and the formation of lateritic nickel deposits have been discussed. The adsorption of metallic cations to the surface of silicate minerals is of significance in many industrial and natural processes. In an earlier publication,' the adsorption of Fe (111) to the quartz surface, a phenomenon of importance in flotation, was studied qualitatively using microelectrophoresis techniques. This paper reports the extension of this work to an investigation of the adsorption from aqueous solution of Co (11) and Ni (11) to the quartz surface. The adsorption of these metals was of interest for two reasons; firstly, the importance of such an adsorption process in the formation of nickel deposits by the lateritization of nickeliferous rocks, and secondly, the comparison of electrophoresis data for these metals, whose hydroxides precipitate in the pH range 7-8, with similar data for Fe (111) where the hydroxide first precipitates at pH 2.8. The development of laterites is essentially a near-surface geological process involving vertical and, in part, horizontal redistribution of elements derived from rock-forming minerals. This redistribution of elements leads to the formation of layers within the lateritic profile which are enriched in one or more of these elements and, depending on the nature of the original rocks, continued lateritization may give rise to iron ores, bauxites or nickeliferous laterites. The mechanisms discussed in this paper deal with the last mentioned type of deposit although the considered factors are applicable to laterites in general. Nickeliferous laterites are derived from basic and ultrabasic rocks containing ferro-magnesian silicates. These rocks may be partly or wholly converted to serpentinites. Besides their high Mg, Si and Fe contents, they also contain small amounts of Mn, Al, Cr, Ni, Co, Cu and Zn. The nickel content is usually near to 0.2-0.3%, but for economic use this must be increased during lateritization to over 1% and preferably over 2%. The zone of nickel enrichment is usually at some depth within the lateritic profile. Unpublished work by Mr. O'Brien on nickel-bearing laterites in eastern Australia indicated four visually different types of material which were separated from crushed samples under a stereo-microscope. Although chemical analyses indicated that all four types of fragments contained 2-4% nickel and smaller amounts of cobalt, X-ray diffraction and optical studies revealed only quartz, hematite and chromite as well as, in one of the types, a serpentine mineral. No distinct, nickeliferous minerals were detected. Furthermore, it was possible to extract some 60% of the nickel content of the laterites with cold 0.5N hydrochloric acid. It was therefore postulated that much of the contained nickel was present as amorphous, adsorbed material. In considering this fact, as well as the problem of the actual method of transfer and fixation of elements within the lateritic profile, it was further postulated that the transfer of these elements probably took place partly as solutions and partly as colloids and that the fixation of these was caused by adsorption and flocculation. The existence of nickel in solution in water draining from serpentinites has been confirmed by Smurov.2 Although we have no evidence for the transportation of nickel in the colloidal state, it is well accepted that colloids are important in the formation of laterites. Previous work by Mr. Mackenzie' has already indicated that iron may be adsorbed on quartz under conditions which could be expected in a lateritic profile, but no information was available for nickel or cobalt. Although adsorption possibly also takes place on minerals other than quartz, the latter is significant because silica is an important constituent of the silicate parent rocks. During lateritization, it is released to form free quartz and it is likely that it is initially extremely fine-grained. This would provide abundant opportunities for adsorption and it was de-

Jan 1, 1970

By A. C. Fraker, H. H. Stadelmaier

In the systems Mo-Fe-C, Mo-Co-C, and Mo-Ni-C the phase boumdaries of the cubic Ij carbides were determined at 1000°C and the variation of the lattice parameter with composition was measured. In the system Mo-Fe-C a single ternary carbide, Mo3Fe3C, was observed. In Mo-Co-C and Mo-Ni-C two new n phases of the type M12C were found: Mo6C06C and Mo6Ni6C. At 1000°C they coexist with the M6C phases MO3-4,CO3-2C and Mo3-4Ni3-2C, respectively. Mo3C03C and Mo4C02C are not separate phases but parts of a single M6C phase region, and the same is true of Mo3Ni3C and Mo4Ni2C. THE n carbide structure type has been known since it was determined by westgren' in 1933, yet reliable information on phase compositions of this type is ap-parently scarce. The following selection of suggested formulas illustrates the extent of the confusion con-cerning compositions: W2-3Fe4-3C,1W6Fe6C,2 W2_4Fe4-2C,3 (W,Fe)6C,4 Tj.-W4Co2C and nz-W6Co6C,5 n1-W3Co3C and n2-w4Co2C,5 W3C03C and w6Co3C2,7 MO3Fe3C,8n2-M04Fe2C,9 (MO,Fe)6C,4 MO3CO3C,8 n2-MO4CO2C.9 Many of these compositions obviously overlap, some disagree in carbon content. The use of the designations 77, and n2 is not consistent. There is uncertainty whether two carbides M6C coexist or whether there is only one M6C with a wide range of substitution of one metal for the other. The existence of M12C seems questionable altogether if one ignores the results of a neutron diffraction study of W6Fe6C supporting that particular metal to carbon ratio.'' More reliable information is available for the iso-typic oxides studied by Nevitt,11 Nevitt, Downey, and orris," and Meussner and carpenter.l3 That n car-bides and oxides are structurally identical, particularly with respect to the location of the metalloid atom, has been demonstrated clearly by Stadelmaier and Meussner14 and also by Parthe, Jeitschko, and sadagopan.15 Therefore the present paper, which attempts to establish the phase boundaries of n defini-tively for three carbon systems, has implications for the whole family of filled Ti2Ni types. EXPERIMENTAL PROCEDURE The samples were prepared in two stages. Pellets weighing 3 g were compacted from powders of the elements and presintered in a closed tube of fused silica that had been evacuated to lO-4 mm Hg. This was fol- lowed by arc melting the sintered samples under argon. Remelting the button once was sufficient to produce homogeneous specimens. The metal powders had the following minimum purities in wt pct: molybdenum 99.5, iron 99.0, cobalt 99.7, nickel 99.9, graphite 99.94. Carbon determinations by the gas volumetric method were obtained from thirty-one of the one hundred samples in the annealed condition. They were limited to compositions that were critical to placing a boundary line in the isothermal sections. The carbon losses were comparatively low and averaged less than 1.5 at. pct. For annealing, the samples were sealed in tubes of fused silica evacuated to 10-4 mm Hg. The annealing schedule was 20 hr at 1250°C followed by 150 hr at 1000" ±10°C. For the water quench the silica glass tubes were broken mechanically to speed the quench. The specimens were examined in the as-cast and annealed condition by standard metallographic techniques. Some of the as-cast and all annealed samples were studied by X-ray diffraction using a 114.6-mm powder camera with filtered Co radiation. The lattice parameters of the n phases were determined by graphic extrapolation to 8 = 90 deg on a plot against sin2 8. It is estimated that the maximum error in the lattice parameters is ±0.002A.. The phase fields were determined by the disappearing phase method, relying on X-rays for phase identification and on microscopy to detect small amounts of a phase. RESULTS Pertinent information on the binary side diagrams has been summarized by Hansen and Anderko16 and Elliott17 and will not be discussed further. Photomicrographs documenting the equilibria at 1000°C have been selected to avoid redundancy due to the similarity of the three systems. Thus there will be only one picture showing an equilibrium with Mo2C, and so forth. Molybdenum-Iron-Carbon. An isothermal section of most of the Mo-Fe-C system at 1000°C is shown in Fig. 1. No attempt has been made to distinguish between a and y iron or to determine their phase boundaries. A conspicuous absence is the phase Fe2 1Mo2C6 with the Cr23C6 structure, supporting the suggestion by Krainer4 that it decomposes by annealing at high temperatures. The only ternary phase is found in a narrow range around Mo3Fe3C (MO43Fe43C14). A microstructure with a two-phase mixture of Mo + n is seen in Fig. 2. The variation of the lattice parameter of n with the estimated composition is shown in in Table IA. The phase diagram does not agree with the results of Kuo9 who reported n2-Mo4Fe2C. Our lattice parameter Varies from 11.095A at the iron-rich end to 11.140Å at the molybdenum-richoend of the phase field whereas Kuo9reported 11.26Å for Mo4Fe2C. We are in agreement with the older work

Jan 1, 1970

BARBOUR, PERCY E., Deputy Supt., (Captain), New York State Troopers; Captain, 22nd Regiment, N. Y. N. G. BARLING, 1i. B., 1st Lieutenant, Engineer Officers' Reserve Corps. CHAPMAN, R. H., Major, Engineer Officers' Reserve Corps. DUNN, THEODORE S., 1st Lieutenant, Engineer Officers' Reserve Corps. DWIGHT; ARTHUR S., Major, Engineer Officers' Reserve Corps. FORBES, CARROLL R., Captain, Engineer Officers' Reserve Corps. GOODSPEED,. G. E., JR., 1st Lieutenant, Engineer Officers' Reserve Corps.

Jan 7, 1917

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Renato G. Bautista, Gary A. Beetner, Anthony L. Frey

A multistage mixer-settler extractor described in earlier hydrodynamic studies was simulated on a computer. A mass transfer mechanism was derived based on a pseudo distribution constant and a pseudo first-order rate constant. The constants were adjusted to give the proper output concentrations and proper form of the stage-wise concentrations. The results obtained by computer simulation of start-up conditions show very good agreement with experimental data. A second simulation was constructed which demonstrated that the model can be greatly simplified and still yield good results.

Jan 1, 1974

ARTICLE I NAME AND OBJECT SEC. I. This Institute is incorporated under the Membership Corporations law of the State of New York; its corporate name is American Institute of Mining and Metallurgical Engineers, Inc.; and its objects are such as are stated in its Certificate of Incorporation. ARTICLE II MEMBERS SEC. 1. The membership of the Institute shall comprise four classes, namely: 1. Members; 2. Honorary Members; 3. Associates; 4. Junior Associates. All members shall be equally entitled to the privileges of membership, excepting that Honorary Members, Junior Associates, and Members and Associates whose residences shall be outside of the United States, Mexico, and Canada shall not be entitled to vote. Members and Associates residing within the United States of America, Mexico, and Canada, .and not in arrears for dues, shall be entitled to vote in person at the meetings of the Institute, or, as hereinafter provided for, by letter ballot. SEC. 2. Members shall comprise all those persons who on the third Monday of February, 1918, were members of the Institute, and in addition thereto, all those thereafter elected or transferred into the class of Members. A person to be eligible for election or transfer into the class of Members must be at least 27 years of age and must have had at least six years' employment in the practice of engineering, mining, geology, metallurgy or chemistry, during at least three years of which he must have held positions of responsibility in one or more of these fields. Graduation from the scientific course of a college, approved by the Committee on Membership, shall be considered equivalent to two years' employment, as required in the previous sentence. Employment as a teacher of engineering, mining, geology, metallurgy or chemistry, if in direct charge, may be considered a position of responsibility as specified in the second paragraph. Persons employed in research or any scientific literary work or in teaching in the scientific departments of colleges, approved by the Committee on Membership, who at the same time are engaged in consulting or in the active practice of mining, geology, or metallurgy, shall be entitled to consider the time so spent in active practice as equivalent to an equal length of time of employment in positions of responsibility, provided the work done or the positions held seem to the Committee on Membership to warrant the equivalency. The requirement of three years' employment in positions of responsibility maybe waived by the Committee on Membership in the case of persons who have done notable original work in mining, geology, or metallurgy, or have won distinction by

Jan 1, 1925

By R. E. Reed-Hill

Recently Levine applied the Conrad-Wiedersich17 equation where H is the activation enthalpy, v the activation volume, T* the effective shear flow stress, T the absolute temperature, and ? the strain rate, to an analysis of titanium single-crystal data. The activation energy as a function of testing temperaturs obtained from specimens oriented for single (1010) prismatic slip is shown in Fig. 12. This diagram indicates a different linear dependence of H on T in the temperature interval from 220° to 300°K than from 0° to 220°K. Levine interpreted the break in this curve as indicating a change in the thermally activated mechanism controlling the flow stress. The purpose of this communication is to point out that the data in Fig. 1 is not consistent with the concept of a flow stress controlled by a simple thermally activated mechanism between 220° and 300°K. The Conrad-Wiedersich relation assumed that the strain rate can be expressed by:

Jan 1, 1968

By N. E. Rogen, N. J. Grant

AGE-hardening in nickel-chromium-titanium-aluminum alloys in the composition range characterized by the Nimonic alloys, is dependent upon the precipitation of the Ni3(AI,Ti) (y') phase.1 This phase is found to precipitate at about 700°C.1 Ni,Al is a face-centered-cubic ordered structure with a lattice parameter of 3.560 A. The atomic configuration is such that the aluminum atoms occupy the cube corners and nickel occupies the cube faces with titanium substituting for the aluminum preferentially. A maximum of three out of every five aluminum atoms can be replaced by titanium atoms in the -y 'phase.' When the solubility of the y'-phase is exceeded for titanium the T(Ni,Ti) phase is formed. The T-phase is an ordered close-packed hexagonal intermetallic structure which shows little solubility for the other elements in these alloys. The lattice parameter for T-phase is a - 2.5454 kr and c = 8.2900 kx units."

Jan 1, 1959

By M. E. Wadsworth, T. L. MacKay

Sintered UO, samples were leached in sulfuric acid solutions of various concentrations. A pressurized system was used so that it was possible to investigate the kinetics of the reaction to 270°C with oxygen overpressures as high as 900 psi. The rate was observed to be a function of the concentration of hydrogen ions and directly proportional to the partial pressure of oxygen. Evidences are presented which indicate that a UO2 surface site reacts with a molecule of water to form a hydroxyl complex which in turn can dissociate with the characteristics of a weak acid. A rate determining step has been proposed which involves the reaction between an oxygen molecule and the hydroxyl complex on the surface of UO,. ThE 2 principal methods for uranium dissolution are carbonate and acid leaching. The sulfuric acid leach is the more popular and is used for the treatment of the majority of the ores of Africa, Canada, and the United States. Low recoveries in basic leach circuits led investigators at the University of British columbia1 to study leaching of uranium ores in pressurized vessels. Early in the study of dissolution of uranium it was found that only uranium in the hexava-lent state could be leached in acid or basic circuits. Therefore, the use of oxygen over-pressure in an autoclave offered an interesting approach to solving the low recovery problems of carbonate leaching. UO, was used in this study because it could be obtained in high purity and also because it is representative of the most refractory of the primary uranium minerals. A pressurized system was used to provide a means whereby important temperature and pressure parameters could be varied for the evaluation of the kinetic processes. The mechanism for the dissolution of UO2 should be similar for any of the uraninite type minerals. At the present all of the kinetic studies that have been conducted have been in carbonate media. Peters and Halpern2 carried out a kinetic study of the leaching of pitchblende. Their specimens were pitchblende ores selected on the basis of high grade and homogeneity. Pearson and wadsworth3 conducted a kinetic study of the dissolution of UO2 in carbonate solutions with results very similar to those obtained by Peters and Halpern. EXPERIMENTAL The UO2 as received* was found by spectroscopic *Mallinckrodt reagent, supplied by the Atomic Energy Commission. analysis to have a purity of 99.94 pct 0.03. It was ground in a mechanical agate mortar and screened through a 400-mesh sieve. Thin flat disks of UO2 were prepared by pressing the sized powder in a specially constructed die at a total pressure of approximately 25 tons per sq in. The pressed disks were approximately 0.25 cm thick, 1.6 cm diam, and 3.5 g in weight. As pressed, the samples were approximately 65 pct of theoretical density. It was essential that a binder such as poly-vinyl alcohol be added before pressing to prevent formation of cracks when fired. It was found by trial and error that pressed disks sintered close to 1870°C in a hydrogen atmosphere, reach densities between 91 and 91.5 pct of the theoretical density. At this density the samples had no measurable porosity based upon a 2-hr emersion in boiling water. An X-ray examination of these sintered UO2 disks showed them to be identical with the original UO2 received. Reaction of UO2 with the alundum furnace core was prevented by placing molybdenum between the samples and the core according to the method of Corwin and Eyerly.4 Leaching studies were carried out in a specially designed autoclave, the details of which have been presented e1sewhere.6 The main features of the equipment as applied to a kinetic study are: 1) solution samples may be removed during the course of a single run; and 2) temperature and agitation are carefully controlled. Two-ml samples removed from the autoclave during the course of a run were analyzed for U3O8 content with a Beckman model DK-2 spectrophotometer by the method of

Jan 1, 1959

By R. Malmstrom, H. Heden, K. Likin

Sublevel caving techniques at Luossavaara-Kiirunavaara AB (LKAB) in Sweden have undergone significant modification over the years as management strives to maintain mine viability amidst rising costs and erratic iron ore prices. LKAB, which is 96% state-owned, is Sweden's principal iron ore producer and operates three iron mines north of the Arctic Circle. By far, the largest of LKAB operations is the Kiirunavaara (Kiruna) underground iron mine which extracts an estimated 30 million t/y of ore, Other iron mines include the Malmberget underground nine, rated at 12 million t/y of ore, and the Svappavaara open-pit mine, rated at about 4.5 million t/y. Mined ores from Kiirunavaara and Svappavaara are shipped by rail to the northwestern harbor of Narvik, while Malmberget ores are sent to Lulea, located to the southeast. LKAB iron ores are marketed by Malmexport AB, which is owned jointly by LKAB and Granges, Sweden's base metal producer.

Jan 10, 1979

NEW YORK LOCAL SECTION Executive Committee Louis D. HUNTOON, Chairman ARTHUR S. DWIGHT, Vice-Chairman THOMAS T. READ, Secretary, Woolworth Bldg., New York, N. Y. E. MALTBY SHIPP, Treasurer GEORGE F. KUNZ W. DE L. BENEDICT Meeting, Jan. 16, .1914 A meeting of the New York Local Section of the American Institute of Mining Engineers was held at the Engineering Societies Building, 29 West 39th Street, New York City, Jan 16 1914 the Chairman, -L. D. Huntoon, presiding. Forty members weld present. In the absence of the Secretary, C. A. Bohn was appointed Secretary pro teen. N. H. Darton, Geologist of the U. S. Geological Survey, delivered an address on. The Occurrence of Gas in Coal, illustrated by lantern slides of charts, curves, and tables showing various features of gas occurrences. J. W. Richards, J. T. Beard, W. L. Saunders, L. D. Huntoon, and Mr. Darton participated in the discussion which followed the address. THOMAS T. READ; Secretary.

Jan 3, 1914

By H. L. Gilbert, W. W. Stephens

Production of anhydrous zirconium tetrachloride by direct chlor- ination of a zirconium oxide carbon mixture in a silica-brick-lined chlorinator is described. Theory and thermodynamics of reactions are discussed. A pilot-model chlorinator and full-scale production equipment are described and operating data are included. ANHYDROUS zirconium tetrachloride required as a starting material in the Kroll process for production of ductile zirconium has been produced in this country principally by chlorination of the carbide or "carbonitride" made by reduction of zircon sand concentrates with carbon in the arc furnace. This process and the equipment used have been fully described.'.' It is well known that zircon or badde-leyite ores may be chlorinated directly with chlorine in the presence of carbon, and this one-step approach would seem at first glance to be preferable to the two steps involved in the carbide-chloride operation. As previously discussed1 considerations leading to adoption of the longer process were: 1—Chlorination of the carbide proceeds rapidly at temperatures below 500°C, whereas temperatures above 900°C were considered necessary for chlorination of zircon-carbon mixtures. 2—The highly exothermic nature of the carbide-chlorine reaction makes it self-sustaining, while heat must be supplied continuously in direct chlorination of the ore. 3—Silicon was thought to be chlorinated along with zirconium in the direct chlorination of zircon, leading to high chlorine consumption. In production of carbide in the arc furnace, silicon is driven off as silicon monoxide and does not enter the chlorinator. 4—For efficient direct chlorination, the ore must be finely ground and intimately mixed with carbon, and the mixture briquetted. 5—Direct chlorination requires a much larger chlorinator for a given production capacity than chlorination of carbide. Recently it has become necessary to produce large quantities of zirconium metal from a chemically purified zirconium oxide. Since the cost of the latter is high, the relatively high losses encountered in production of carbide in the arc furnace could not be tolerated, and a graphite resistor furnace5 was developed for production of carbide, which was then chlorinated in equipment previously used for chlorinating arc-furnace carbide. This method of operation was quite satisfactory, and losses in the carbid-ing step were minimized. However, operating costs were relatively high, and the process did not lend itself particularly well to large-scale operation because of the multiplicity of small units required and the hand labor needed to load, unload, and maintain the furnaces. To chlorinate the carbide, a vertical-shaft chlorinator was used in which the charge was heated with a central split graphite-rod resistor.' Operation of this chlorinator was somewhat less satisfactory with the resistor furnace carbide than it had been with the arc-furnace carbide due, principally, to the differences in physical properties of the carbide. The arc-furnace carbide is obtained as a fused metallic-appearing mass which can be crushed to -1/4 in. with production of a minimum of fines, whereas the resistor-furnace product is lightly sintered and produces a large proportion of fines in crushing and handling. These fines tend to pack in the chlorinator and promote channeling. which results in poor chlorine efficiency and low capacity. Data based on production of 22,000 lb of chloride from resistor-furnace carbide in this equipment are shown in Table I. Necessity for increasing production of chloride from about 2,000 to 15,000 lb per week led to further investigation of possible methods for direct chlorination of the oxide or oxide-carbon mixtures. Direct chlorination of the pure oxide presents much less difficulty than chlorination of zircon sand or oxide ores. The oxide is easily ground to —200 mesh, and no silica is present to cause excessive consumption of chlorine or contamination of the product. Theoretical Considerations The principal chemical reactions involved in chlorination of mixtures of zirconium oxide and carbon may be written as follows: ½ ZrO2 (c) + C(c) + Cl2(g) = ½ ZrCl,(g) + CO(g) [I] 1/2 ZrO, (c) + CO(g) + Cl2(g) = ½ ZrCl,(g) + CO2(g) 121 ½ ZrO2 (c) + ½ C(C) + Cl2(g) = 1/2 ZrCl4(g) + ½ CO2(g) [3]

Jan 1, 1953

(The following list contains the names of those members of the Institute of whose connection with military service we have only recently become acquainted; it also includes the names of a few who have recently been promoted or transferred. A complete list was published in the Year Book, issued as a supplement of the Bulletin for March, 1918.) ANDREWS, CLARENCE W., 2d Lieutenant, 12th Batt. Infantry Replacement Regiment. ANDREWS, JOSEPH C., Ordnance Dept., U. S. A. ARLUCK, A. A., Headquarters, 1st Corps of Artillery, A. P. O.. No. 759, A. E. F.

Jan 9, 1918

By Barton K. Cross, Ronald J. Sparks

The question of' providing training for miners in the United States has been answered by Public Law 95-164, which stipulates not only that the mine operator will provide training, but generally how much and what material will be covered. At an operation employing 500 miners, the eight-hour refresher training requirement will necessitate an annual investment of over two man-years in labor, a figure that does not include the time spent in training new hires and job transfers. Such an investment can and should be cost effective, while catering to miners' needs.

Jan 6, 1979

By H. L. Hartman

The past 15 years have seen rapid advances in the metallurgy of materials for drill machinery and bits, but rock drilling itself continues to be largely an art. Jet piercing, roller bit rotary drilling, and rotary percussion are all promising new techniques, but the percussion rock drill, still used for 90 pct of the blastholes in U. S. hard-rock mining, has undergone no major modification since pneumatic machines were first used successfully in the 1860's. This singular lack of progress stems directly from ignorance of the nature of impact failure of rock and the fundamentals of rock penetration in general. For nearly three-quarters of a century, mining engineers did not know what went on at the bottom of a drillhole. More and more, in recent years, investigators in the field have come to realize that to understand percussion drilling they must study the basic action of percussion bits penetrating rock—not the per- formances of commercial rock drills, with their many attendant variables that cannot be isolated or controlled. Using a vertical drop tester, the writer investigated the sequence of rock failure under the impact. of a simple, chisel-shaped bit when changes were made, singly, in significant variables. Results of this and other research are reported here. BASIC CONCEPTS The formation of a crater in rock under the impact blow of a bit can be observed experimentally. Directly beneath the edge of the bit the rock is crushed into fine powder; toward the sides of the crater it chips out into relatively large fragments. An explanation of these phenomena has been offered by Drilling Research Inc. on the basis of work sponsored by this organization at Battelle Memorial Institute (Ref. 1,1954, Summary, pp. 9-18). High-speed photography and force waveforms obtained from strain gages mounted on the bit near the cutting edges recorded the sequence of events in crater formation with a die-shaped or blunt, wedgc-

Jan 1, 1960