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By N. A. Gokcen

THE data of Chaudron,1 Tonosaki,2 and Collins³ on the thermodynamic properties of MOO, disagree widely. These authors, by using essentially similar methods, studied the following reaction: 1/2MoO2(s) + H2(g) == i/2Mo(s) + Ph,o H2O(g),K = —----- [1] Ph2 Tonosaki used a vacuum system consisting of a furnace containing MOO, and a water saturator whose temperature was kept at 25 °C with a thermostat. After repeated evacuation, hydrogen was admitted slowly into the system. The experiments were based upon the fact that at a constant furnace temperature and constant partial pressure of water, the total pressure of gas mixture over MOO? + Mo is constant. Any attempt to vary the pressure by external forces would vary only the amounts of MoO2, and Mo, after which the pressure should return to the equilibrium value in accordance with the equilibrium constant of reaction 1. The actual value of K was determined from the total equilibrium pressure (sum of Ph2o + Ph2) at each temperature. The total pressure of gas was varied within the range of 80.2 to 125.4 mm Hg, within a temperature range of 645" to 823 °C. The results were summarized as log K = 0.9413 — 1444.6/T for reaction 1. The ratio of H2O/H2 was considered to be uniform in spite of the presence of a thermal gradient across the static gas phase. It was shown by Rosenqvist and Cox,' however, that in somewhat similar circumstances the error resulting from thermal diffusion may be large. Collins³ improved Tonosaki's method by attempting to avoid thermal diffusion errors. His equilibrium data were obtained at 700°, 800°, and 900° C, and the result for reaction 1 was expressed as log K = 1.258 — 1822/T. The purpose of this investigation was to study reaction 1, avoiding the thermal diffusion errors, and to obtain equilibrium data in a considerably wide temperature range for the reliable extrapolation of the resulting thermodynamic functions. Experimental Procedure The diagram of the apparatus is shown in Fig. 1. Tank hydrogen was passed through a tube containing platinized asbestos at 425 °C in order to convert a trace of oxygen into water vapor. The flow rate was carefully controlled with a capillary-type flow-meter B and a bubbling column D. The gas was then presaturated sufficiently at P and led into a condenser system immersed in a thermostat, controlled to within ±0.002C. The temperature of the thermostat was measured with a thermometer calibrated against a certified standard. The temperature of P was adjusted to avoid the condensation of unduly large amounts of water as judged from the rate of flow out of the capillary tube M. Argon was purified by passing it through magnesium chips kept at 630°C. After passing through the flowmeter B', it was mixed with moist hydrogen emerging from the thermostat. The resulting gas mixture was then led into the hot zone of the furnace through the heated glass tubing and the silica tube S, thus insuring the same ratio of PII2o/Ph2 from the condenser to the reaction chamber, thus avoiding thermal diffusion. The entire gas system was of all-glass construction, except at the magnesium train. The furnace comprised a glazed alumina tube over which a 15-in. platinum coil was wound. The lower end of the alumina tube was tightly closed with a brass bottom and a silicone rubber gasket, and the upper end with two glass disks, each with a hole of 1/16 in. in diameter, through which the gas mixture escaped into the atmosphere. A back-up coil of kanthal wire facilitated the temperature control of the furnace. A coil of annealed molybdenum strip of 99.99 pct purity, 0.005 in. thick and 0.050 in. wide, and weighing 17 g was hung in the furnace with a 0.010-in. platinum wire attached to one end of a sensitive analytical balance. The temperature of the furnace was measured with a Pt-Pt-10 pct Rh thermocouple checked against a standard. The experimental procedure consisted of heating the gas purification trains, adjusting the gas flow rates, attaining a constant thermostatic temperature, flushing the entire system for 2 hr while heating the furnace to well above the expected equilibrium temperature and then cooling it at a rate of 0.3°C per min during which time the change in the weight of molybdenum was observed on the balance. For a given thermostatic temperature, i.e., a constant Prr,o/Px,, molybdenum oxidized upon cooling slightly below the equilibrium temperature. The procedure was then repeated by heating the furnace and thus reaching a temperature slightly above the equilibrium value. The average of the two temperatures, differing by 2" to 3"C, was considered to be the true equilibrium temperature. In order to determine the stoichiometric composition of the oxide phase present in this investigation,

Jan 1, 1954

WE the undersigned, being all persons of full age and citizens of the United States, and a majority residents of the State of New York, desiring to form a corporation pursuant to the provisions of the Membership Corporations Law for the purpose of incorporating, as provided in Section 6 of Article I. of said law, the existing unincorporated association known as American Institute of Mining Engineers, do hereby make, acknowledge and 6le this Certificate for that purpose, and do Certify as follows: I. That the American Institute of Mining Engineers is an unincorporated association organized and existing with the object of promoting the arts and sciences connected with the economic production of the useful minerals and metals and the welfare of those employed in these industries by means of meetings for social intercourse and the reading and discussion of professional papers, and to circulate by means of publications among its members and associates the information thus obtained. 11. That the persons duly appointed or designated to manage the affairs of said association are designated by the rules thereof Members of its Council; that the under- signed are all members of said Council as the same was constituted on the 29th day of December, 1904.

Jan 1, 1942

By W. F. Carew, F. A. Crossley

The stress for rupture in 500 hr at 1000° F has been reported to be about 13,000 psi higher for Widmanstitten than for equiaxed microstructures for the Ti-7A1-3Mo alloy.1,2 Also, limited data indicated Widmanstatten and equiaxed micro-structures to have approximately the same creep resistance at 1000°F.' Recently, data have been obtained which show Widmanstatten structures to be significantly more creep resistant than equiaxed structures at 1000°F, see Fig. 1. These data represent several ingots and several heat treatments of the solution treat and age type. The thermal history and tensile properties corresponding to the creep data are given in Table I. The creep data were taken from creep-rupture tests. The associ- ated rupture data are given in Table 11. It may be noted that total creep deformations are considerably greater for equiaxed structures. Previous evaluations of stability indicated that all of the structures tested are stable with respect to exposure to stress at elevated temperature.1,2 For a minimum creep rate of 10-4 in. per in. per hr the difference in stress levels is indicated to be about 17,000 psi. Holden,et found an irreversible contraction upon heating three a-ß titanium alloys in the equiaxed condition above their ß-transus temperatures. The alloys were: Ti-6Al-4V, Ti-155 A (5Al-1.3Cr-1.5Fe-1.2Mo), and C-130 AM (44-4Mn). The net contractions in length between the original equiaxed and the final Widmanstatten structures measured at room temperature were: 0.4, 0.4, and 0.7 pct, respectively. The greater creep and rupture resistance of the WidmanstStten structure of the Ti-7Al-3Mo alloy may be associated with the phenomenon

Jan 1, 1959

By I. H. Van Horn

IN the development of the incandescent electric lamp one aim of the investigators has been to establish the fundamentals of lamp design, so that the performance of any new lamp may be accurately predicted. The study of the temperature relations in lamps has done much toward establishing these fundamentals. Fig. 1 gives the typical FIG. 1.-ARRANGEMENT OF PARTS IN MAZDA B AND MAZDA C LAMPS. arrangement of the lamp parts in both the Mazda B (vacuum) and Mazda C (gas-filled) lamps. The outstanding differences are the filament form and the bulb shape. The ideal lamp is one in which the filament operates at a uniform temperature throughout its length. In practice, it is necessary to disturb this uniformity by the introduction of supports and terminals. The amount of cooling at the supports and terminals affects the over-all efficiency of the filament as a light producer. The life of the lamp is ordinarily proportional to the maximum and not to the average filament temperature. The Holborn-Kurlbaum type of optical pyrometer has

Jan 9, 1919

By P. van Meurs, C. van der Poel

From observations of the linear displacement of oil by water from a porous medium as visualized in transparent models, new insight into the mechanism of vircous fingering, as occurring in the case of unfavorable oil-water viscosity ratios, has been obtained. A mathematical description of the observed phenomena leads to formulas. expressing both oil production and pressure drop across the formation as a func-tion of cumulative water injection with the oil-water viscosity ratio as a parameter. Theoretical results are compared with those obtained from scaled laboratory experiments in sand-filled rubes in which water was injected at the one end and production was obtained from the other end. In the second part of the paper the same principles are used to calculate a more realistic production scheme in which production of oil and water is obtained from wells regularly distributed over the field. Results show satisfactory agreement with production data of an actual oil field. Application of the theory is extremely simple as only a limited number of readily available parameters have to be known and results are presented in an analytical form. INTRODUCTION When oil is displaced from a porous medium by water having a viscosity lower than the oil. the oil-water interface is essentially unstable and has the tendency to break up into what are called "fingers" or "streamers". This Viscous fingering has been noted before by Engelberts and Klinkenberg' and has recently been demonstrated in transparent models by van Meurs-. It is highly questionable whether it is at all possible to perform relative permeability measurements In the laboratory reproducing the same distribution of water and oil in the cores as Present in practice. Therefore, we based our descriptive theory on an idealized model which was "tailored" to the fingering mechanism as observed in transparent models under scaled conditions. This approach was the more attractive, as a highly simplified model had already yielded an adequate description of the process. Though we are aware our theoretical approach can be extended in several ways, it is given here in its simplest form to explain the fundamental principles and to avoid complicated formulas. SIMPLEST CASE OF A LINEAR DRIVE Observations and Theoretical Model In Fig. 1 a series of photographs shows how water penetrates into a formation saturated with an oil which is considerably more viscous than the water. The technique for obtaining these photographs, using transparent models, has already been described?. The fingers can be clearly observed and are seen to be mainly parallel to the flow direction. Note that the contours of these fingers are not smooth but show protrusions of water enclosing oil pockets.

By H. I. Kaplan, C. Laird

Pulsating contpresslon experiments have been carried out on coppev single crystals in order to test the adequacy of mechanisms which have been suggested for stage I cvack grouth when tension-compression loading is the mode of- applying stress. Although dejorrrzation continues slowly for many millions 01 cycles under Pulsating compression stresses both within slip bands formed in the first cycle and by the creation of new bands, the morphology of these bands is not typical of fatigue slip bands in general and neither initiation of cracks nor propagation of artificially induced cracks was observed. It is therefore concluded that 1) initiation and growth of cvacks in fatigue is a consequence of repeatedly reviersed, microplastic deformation rather than intermittent, unidirectional deforma-tion, 2) gvowth of- cracks in polycrystalline specimens FAILURE of a plain specimen subjected to cyclic tension-compression stresses is known to take place by two stages of cracking.1,2 The first stage is characterized by propagation along slip planes roughly at 45 deg to the stress axis, in the outermost layers of the specimen, and at a rate of the order of angstroms per cycle. The average macro orientation of a stage I crack is normally at right angles to the stress axis,3 however, except in a few cases, e.g., single crystals in suitable orientation, when both the micro and macro directions of propagation are at 45 deg to the axis. After a stage I crack has penetrated to such a depth that the stress on the remaining section of metal is significantly increased, stage II crack propagation succeeds stage I, at right angles to the stress axis both micro and macroscopically. Direct evidence of the processes occurring during a subjected to pulsating compression stresses is caused by the setting-up of local tensile stresses due to interactions between grains, and 3) both unslipping mechanisms and operation of the "plastic blunting process " within a single slip band are adequate descriptions 01- stage I growth, being fundamentally similar. The slip mode of a material, however, is considered to control whether unslipping or the plastic blunting process takes place. Thus materials of planar slip mode fauor unslipping while those hacing a wavy mode, and capable of cross slip, can fail in stage I by the plastic blunting process. Similarly, and perhaps more importantly, reversed torsion and tension-compression modes of applying stress may be expected to cause the unslipping mechanism and the plastic blunting process, respectively. single stress cycle4,5 shows that stage II crack propagation takes place by a repetitive plastic rounding and closing of the crack tip. This mechanism has come to be called the "plastic blunting process".6 On the other hand, the extreme smallness of the phenomena occurring in stage I growth prevents elucidation of its mechanism by direct observation. Many suggestions have been proposed for this mechanism,'"-" most of them involving some kind of random or systematic unslipping process. It is possible, however, that the mechanism of stage I growth is essentially the same as that of stage 11, but occurring on a much smaller scale.6 In this situation, it has been observed that cracks propagate along slip bands comprising dislocation cell structures and bounded on either side by regions of different dislocation density and distribution." Since there is evidence that such bands are softer than their matrix, * it may be expected that the plastic blunting process occurs within a single slip band during stage I growth, rather than on two gross bands as in stage 11, when the acting stress is greater.= The blunting of the crack tip allowed by coordinated

Jan 1, 1968

An index covering the papers presented at the 19 Open Hearth Conferences from the first meeting at Pittsburgh, April 15 and 16, 1925, through the most recent meeting in Detroit, April 16 and 17, 1936, the reports of the meetings and the discussions that took place has been published, accompanied by a list of the papers presented, grouped by subject matter and a list of books of interest to open hearth men The list below gives the Open Hearth Conference meetings in chronological order and the symbols used to designate them in the index SYMBOL MEETINGS P25 Pittsburgh, April 15 and 16, 1925 C125 Cleveland, Oct 13 to 15, 1925 Ch26 Chicago, April 12 and 13, 1926 P26 Pittsburgh, October, 1926 B27 Buffalo, May 3 to 5, 1927 D27 Detroit, Nov 2 and 3, 1927 Y28 Youngstown, May 3 and 4, 1928 P28 Pittsburgh, Nov 15 and 16, 1928 C129 Cleveland, May 16 and 17, 1929 Ch29 Chicago, Nov 5 and 6, 1929 B30 Buffalo, May 20 and 21, 1930 C130 Cleveland, Nov 20 and 21, 1930 Ch31 Chicago, May 12 and 13, 1931- D31 Detroit, Nov 24 and 25, 1931 P32 Pittsburgh, May 24 and 25, 1932 Ch33 Chicago, June 29 and 30, 1933 P34 Pittsburgh, June 28 and 29, 1934 C135 Cincinnati, April 11 and 12, 1935 D36 Detroit, April 16 and 17, 1936 The list of papers mentioned in the first paragraph above is reprinted on pages 428 to 431 The Index of the Open Hearth Proceedings may be purchased at A I M E headquarters, 29 W 39th St, New York

Jan 1, 1936

By Herbert Everest

THE idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentage expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2: Development, including all necessary yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, drawslate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1 ½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the Beadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high.

Jan 1, 1916

By Herbert A. Everest

The idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. , Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentagc expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2. Development, including all necessary .yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, draw-slate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the deadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high. The item, depreciation of coal reserves and royalties, is usually about

Jan 1, 1916

Discussion of the paper of Dorsey A. Lyon and Robert M. Keeney, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 2117 to 2149. C. D. WOODWARD, Butte, Mont.:-We constructed a small blast furnace, and used first carbon electrodes and an alternating current of about 50 to 75 volts. We were able to smelt the material which was fed to the furnace, but we did not carry on experiments far enough to see just exactly what the power consumption was and compare it with the amount of material treated. We found that the blast seemed to help materially in reducing our charge and cutting down our power consumption. There were several other experiments tried at the time, and I do not call to mind now just what they were. We were trying to do away with the carbon electrode and replace it with an iron one, and it also seems to me we tried to use the ore for an electrode; but I believe that the electric furnace in the reducing end is simply a question of British thermal units per kilowatt-hour compared with British thermal units in a pound of coke. Now, if a pound of coke costs us 0.5 c. and a kilowatt-hour costs us in the neighborhood of 1.5 c., I think there is a big advantage in using coke instead of the electrical current for copper reducing. We tried steam, and a number of other methods, to throw copper down, but it has been some time ago and I am not very familiar with them now. The outlook of the experiments did not seem to be very promising, and we took it more as a thermal problem than an electrical one. PROF. JOSEPH W. RICHARDS, South Bethlehem, Pa.:-Mr. Woodward has spoken of the question as a thermal problem. I would call attention to the fact that electric furnaces usually utilize the heat of the electric current at from four to ten times the efficiency with which the heat generated by fuel is utilized; so it will not do simply to compare by thermal units generated. Your total thermal units may cost much less from coke than from electricity, yet you will find the electrically generated thermal units utilized in a properly designed furnace from three to ten times more efficiently than in the combustion furnace. You must make that allowance is comparing the cost of thermal units generated electrically and thermal units from coke.

Jan 11, 1913

By L. E. Fish, G. S. Wyman

CHUQUICAMATA open-pit mine is capable of producing a total of 105,000 tons daily. When the sulphide plant is operating to capacity the distribution of this quantity will be approximately 30,000 tons sulphide ore, 30,000 tons oxide ore and 45,000 tons of waste. At that time the combined capacity of the oxide and sulphide plants will be about 540 million lb of copper per year. Twenty one benches have been opened to date and all are still active except the three top ones at the northeastern limit of the pit. These have been pushed back to final slope line. The pit is 8850 ft long, 3540 ft wide and 980 ft deep. Benches are not all developed to a standard height as the large development shovels used in cutting new benches have been of different size and range. At this time new benches 521 ft high are being developed with an' 11-yd shovel. Ore haulage through the greater part of the pit operation has been on down grades to the crushing plant. With the opening of bench F-4 in April 1944 the first adverse-grade ore haul was encountered. Of the total tonnage now being mined, 35 pct comes from the four benches that are below the level of the crushing plant yard. Waste had been dumped on the same or nearly the same level as the bench from which it was removed up to the opening of bench F-4. Waste from this and lower levels is hauled to the level of the crusher, yard where it must cross the flow of ore traffic to reach the only available dumping space to the east of the mine. The average ore haul is 3.13 miles one way shovel to crusher, and the average waste haul from that portion handled by railroad is 3.32 miles one way from shovel to waste dump. Waste that is being removed by truck haulage from the upper part of the -mine is hauled an average of 1 mile. Ore loaded by trucks is hauled 3/4 mile to a car loading ramp and then by rail to the crushing plant. Drilling and Blasting There are two types of development for blasting being used: One for opening new benches and the other for bank shooting. The former consists of 9a in. diam churn-drill holes drilled 11.8 ft center to center in rows 10.2 ft apart with the holes staggered in the rows thus giving an equilateral-triangle pattern of 11.8 ft on a side. After the first block, 492 ft long by 71.5 ft wide, is shot in the pattern described above, the cut is advanced by smaller shots consisting of four rows of holes spaced 23 ft center to center with the rows 23 ft apart. These shots are fired in conjunction with side shots in which the holes are drilled 33 ft center to center and 29.5 ft back from the toe, thus squaring up the bank for subsequent shooting. An average bank shot consists of about 60 holes drilled in two rows parallel to the face, 30 holes to

Jan 1, 1952

By David L. Holt, Walter A. Backofen, Anwar-uI Karim

Specimens of a hydrided Mg-0.5 Zr alloy were strained in tension at 500°C and constant rates of 2 x10-3 5 x 10-3, and 2 X 10" min-1. Hydride-denuded zones formed at grain boundaries normal to the tensile-stress direction as a result of magnesium transport during difusional flow. The width of the zones could be measured and the measurement used for calculating the diffusional component of the imposed tensile strain. The strain from diffusional flow was found to increase with imposed strain at a diminishing rate, tending to saturate at approximately 12 pct. Strain rate sensitivity of flow stress was low. The apparent non Newtonian character of the diffusional flow is attributed to a non Newtonian process acting in parallel with it which could be boundary shear. Fracture grows out of voids that form in the denuded zones. DEFORMATION of a grain by diffusion of atoms from boundaries stressed in compression to boundaries stressed in tension is Newtonian viscous,1-3 and evidence has accumulated in recent years that such a process may be responsible for the high strain-rate sensitivity of the flow stress of super-plastic alloys.4"7 One piece of evidence is that experimental stress: strain-rate relationships can be quantitatively explained.5-7 There is also metallo-graphic evidence of diffusional flow in superplas-ticity, but in a limited amount. The formation of striated bands on the surface of superplastically deformed specimens has been attributed to diffusional flow.5"7 The basis of that attribution came from experiments on a coarse-grained, nonsuperplastic and hydrided Mg-½ wt pct Zr alloy which formed hydride-denuded, light etching zones at tension-stressed boundaries when strained in tension at 270?C.6 The origin of these zones had already been traced to the diffusional flow of magnesium atoms to the boundaries.' The particular observations in the more recent work were of striated-band formation on the surface and denuded-zone formation internally, with both the bands and zones having the same width and appearing at tension-stressed boundaries. It was argued that the bands were a surface manifestation of the zones and hence of diffusional flow. Of course in superplastic alloys which do not contain internal metallographic "markers", the surface bands can be the only metallographic indication. In the present work, denuded-zone formation was utilized, as it has been by others,9-11 to extend the observations of diffusional flow and to measure the strain, ed, resulting from it. Grain size had to be large to measure ed with accuracy. The grain size chosen for this study was -30 , and with that a strain of 10 pct from diffusional flow produces a denuded zone only 3 µ in width. The large grain size naturally precludes superplasticity. The observations of diffusional flow were complemented by determining the strain from the other operative deformation modes: slip, e,, and grain boundary shear, egb. An incremental specimen extension is the sum of increments from slip, and grain boundary shear as well as diffusional flow. Division by a common length is required to convert to strain. If this length is taken as the initial specimen length, then imposed engineering strain, e, is given in terms of the component engineering strains by e = ed + es + egb [1] Stress:strain-rate relationships are determined by the way in which this "strain balance" is made up. EXPERIMENTAL Material. Zirconium hydride markers were introduced into the Mg-0.5Zr alloy by annealing in hydrogen at 450°C for 30 min. The hydride concentration was particularly high at zirconium rich stringers, which was fortunate in that the transverse boundaries at which denuded zones form lie perpendicular to the stringers. Grain size after annealing was 30 µ. Photomicrographs of unstrained and strained material are shown in Fig. 1. Procedure. Specimens were strained in tension with an Instron machine at crosshead velocities of either 2 x 10"3, 5 x X or 1 x 10-2 in. min-'. Specimen length and diameter were 1.0 and 0.2 in., respectively, so that initial strain rates in tests at constant crosshead speed were 2 x 10"3, 5 x X and 1 X l0-2 min-1. Tests were made at 500°C which is a compromise temperature at which diffusional flow is still measurable but grain growth is not active enough to interfere with metallographic measurements. The tests were made in a hydrogen atmosphere. Strain Balance. An equation additional to [I] is eg = ed + es [2] where eg is strain measured from grain elongation. Measurement was made of ed, eg, and, of course, e, which enabled all the strains in Eq. [I] to be determined. For this purpose, strained specimens were sectioned longitudinally, polished, and etched. The strain from diffusional flow, ed, was computed by measuring on photomicrographs the width in the tensile direction of denuded zones at either end of a grain XI, X2, adding them, and dividing by twice the initial longitudinal grain dimension L0, Fig. 2. Reported values are the results of measurements on seventy randomly selected grains; 95 pct confidence limits on ed were +1.5 pct strain. To measure eg, the maximum length, L, and the maximum width, W,

Jan 1, 1970

Organization Place Date 1918 American Chemical Society Cleveland, O. Sept. 9-12 American Society of Sanitary En Engineers Chicago, Ill. Sept. 3-5 National Petroleum Association Atlantic City, N.J. -Sept. American Institute of Mining Engineers Colorado Sept. 2-7 National Association of Stationary Engineers.... Cincinnati, O. Sept. 9-13 Association Iron, Steel, Electrical, Engineers Baltimore, Md. Sept. 9-14 British Iron and Steel Institute, Autumn Meeting London, Eng. Sept. 12-13 National Exposition of Chemical Industries New York, N. Y. Sept. 23-28 American Electrochemical Society Princeton, N. J. Sept. 30- Oct. 2

Jan 9, 1918

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By R. M. Waterstrat

The phases occurring in the V-Mn system were studied by means of X-yay diffraction and metallo-paphic techniques, using are-melted alloy specimens annealed in the temperature range 800° to 1150°C and quenched. The bcc solid solution extends at 1250°C all the way from vanadium to 6-manganese. Below 1050°C the a-phase is formed, and the terminal a-manganese phase is stabilized up to about 900°C by vanadium in solid solution. IN the only previous general survey of the V-Mn system Cornelius, Bungardt and Schiedtl reported the existence of three intermediate phases corresponding to the approximate compositions VMn,, VMn, and V5Mn. The phase VMn8 has recently been identified as a o phase2 but the alloy VMn was found to have a bcc structure2 corresponding apparently to the vanadium solid solution rather than to the large cubic unit cell reported by Cornelius et al. 1 Subsequent work by Rostoker and Yamamoto3 has shown that the vanadium-base bcc solid solution extends to at least 15 pct Mn at 900°C. An alloy corresponding to the composition VMn, was examined by Elliott,4 who reported that the as-cast sample as well as samples annealed at 1200o and 1300°C had bcc structures, but that annealing at 1000°, 800") and 600°C produced two phases. One of these phases was apparently the bcc solid solution and the other resembled the o phase structure. Hellawell and Hume-Rothery5 established the phase relationships in manganese-rich alloys above 1000°C, and showed that the o phase in this system is replaced by the 6 Mn (bcc) solid solution at temperatures above 1050°C. These results suggest that a continuous bcc solid solution may exist above 1050°C between vanadium and 6 Mn. The present investigation was undertaken in order to develop more complete information in regard to this system. EXPERIMENTAL METHODS The alloys used in the present work were prepared by arc-melting electrolytic manganese having a minimum purity of 99.9 pct and vanadium lumps with a purity of 99.7 pct. The major impurities present in these metals were carbon, nitrogen, and oxygen and this would account for the small percentage of nonmetallic inclusions observed metal-lographically. The arc-melting was at first performed under a helium atmosphere and it was necessary to keep the melting times as short as possible in order to minimize the loss of manganese by vaporization. It was later found that the evaporation of manganese was considerably reduced when the melting was done under argon atmosphere. The final composition of each alloy was calculated by assuming that the total weight loss during melting was due to evaporation of manganese. Compositions which were calculated in this manner agreed reasonably well with the results of chemical analysis, as shown in Table I. Spectrographic analysis revealed the presence of contamination by tungsten, but in no case was the percentage of tungsten greater then 0.4 at. pct. The specimens were in each case broken in half and the fractured section was examined visually and microscopically for evidence of inhomogeneity. Each specimen was homogenized at temperatures near l100°C, as shown in Table I. After this treatment most specimens consisted of large columnar grains of the bcc vanadium solid solution. The etchant used in most of the metallographic work consisted of 20 pct nitric acid, 20 pct hydro-flouric acid, and 60 pct glycerine. It was found that this etchant would clearly delineate the phases present in these alloys although it does not produce any striking contrast between the phases. For certain manganese-rich alloys, a 1 pct aqueous solution of nitric acid was used. This etchant gave a brown color to the a-manganese phase, whereas the o phase was virtually unattacked and appeared very light as shown in Fig. 1. The etchants used by Cornelius et a1.l were found to produce spurious effects in some of these alloys. In particular, the vanadium-rich alloys etched in hot sulfuric acid often appeared to consist of two phases when both X-ray diffraction and etching with the glycerine-acid mixture indicated the presence of single phase bcc solid solution. A few percent of what appears to be an oxide or nitride phase was found at the grain boundaries and in the interior of the grains, especially in the vanadium-rich alloys. All alloys were annealed in sealed silica tubes containing 1 atm of pure argon and these tubes were then quenched in cold water. Although some manganese loss occurred during annealing, the loss seemed to be confined to the surface of the speci-

Jan 1, 1962

By Samuel M. Shallcross

The principal function of the Bell limestone mine, of the American Lime & Stone Co., at Bellefonte, Pa., is to supply raw material to the company's modern rotary kiln plant at Bellefonte. Because mining costs are naturally higher than many open quarries with more favorably lying strata, it cannot compete, except to a limited extent, in the markets for road material and fluxing stone. In the broad sense, mining does have an advantage over the average open quarry when the end products are designed for and controlled by process and final inspection for the exacting demands of the industrial and chemical markets. Approximately 80 per cent of the output of the Bell mine is sold in the industrial and chemical markets. And as higher standards are set each year for construction and agricultural work, the remainder of the mine output goes to these two classes. Mining produces more fines than open quarries, and these fines are difficult to market; on the other hand, owing to absence of contaminating clay and impure rock, a greater percentage of the output of a mine is marketable. Because of the great production of fines, rotary kilns for lime burning are necessary when the stone must be mined. In 1921 the American Lime & Stone Co. operated 19 quarries and 72 lime shaft kilns spread over an area of 40 miles between Bellefonte and Altoona, Pa. At that time the lime-production capacity of the 72 kilns amounted to approximately 700 tons per day. After 15 years the company operates three rotary kilns and six shaft kilns, with a combined output of approximately 500 tons of lime per day, concentrated in its single Bellefonte plant (Fig. 1). Its only stone-producing plants are now the limestone mine at Bellefonte and a construction-stone open quarry at Union Furnace, near Tyrone, Pennsylvania. Since the middle of 1922, the management of the American Lime R. Stone Co. has been under the direction and ownership of the Warner Company, of Philadelphia. To the Warner Company's experience, gained through successful operation of lime plants in eastern Pennsylvania for many years, and the successful solution of material-handling problems

Jan 1, 1937

By R. L. Fullman, J. W. Cahn

A method is derived for obtaining the distribution of sphere diameters and plate thicknesses (or pearlite spacings) from size distribution functions obtained along randomly oriented lines. The method described offers several advantages over existing methods. IT is often of interest to measure not only the average particle size but also the distribution in sizes of particles dispersed in an opaque material, such as an alloy. Particles of a single size will exhibit a range of apparent sizes on the plane of polish. If the sample contains a distribution of particle sizes, the distribution of observed sizes on the plane of polish is related to the distribution of actual particle sizes by an integral equation, For example, a distribution of sphere sizes, N(r),

Jan 1, 1957

By Gerhard Derge

THE chemical and physical properties of slag systems are of special interest to metallurgists, for nearly all metals are in contact with molten slags during the primary reduction from their ores and then during the subsequent refining and alloying processes that determine their ultimate use, These processes are carried out at high temperatures and involve such complex chemical systems that fundamental scientific knowledge of the properties and behavior of these slags is inadequate. Various hypotheses have been proposed to explain known slag characteristics. The commonest of these regard the molten slag as constituted of compounds that are dissociated in varying degrees into their constituent oxides, as illustrated by Schenck1 and the recent papers of Chipman and co-authors.2-5 Another proposal is that these compounds dissociate by completes or partial ionization, the evidence for which has been reviewed by Martin and Derge.7 Whatever hypothesis is favored, the exact formulation of these compounds is incomplete. It is based on three principal types of information: I, Studies of solidified slag samples by X-ray diffraction, petrography, and similar methods that can be used at room temperature, The correlation of this information with the high-temperature properties of the molten slag is uncertain, 2. Studies of slag-metal equilibria, which have the advantage of providing observations on the molten slag at high temperature, However, even these data present problems of interpretation because they require a complete knowledge of the metal phase, 3, Direct measurements of properties of molten slag at the temperature of greatest metallurgical interest. Viscosity, electrical conductivity, specific heats, and similar properties have been measured, but the data of this type are widely scattered and frequently difficult to interpret, Additional direct measurements of properties of molten slag that can be interpreted in terms of chemical behavior are required, and this paper shows that the standard physical-chemical measurement of reversible electrode potentials, which measure the activity of specific constituents, is a fruitful source of such information, The purpose of this investigation, therefore, was to find an electrode combination that would yield reversible electromotive force measurements when dipped into the molten slag to be studied. Slags of the systems CaO-SiO2 and CaO-A1203-Si02 were chosen for study because graphite containers could be used and the data could be correlated with those for viscosity, conductivity, and other proper-

Jan 1, 1946

By H. B. Bell, P. C. Lidster

A study has been made of the use of ThO2-Y2O3 solid electrolytes to determine activity of chromium in Fe-Cr and Ni-Cr alloys in the temperature range 1300° to 1700°K. This method has been shown to give results which agree with the best vapor pressure measurements. SEVERAL investigations1"5 have been made of chromium activities in Fe-Cr alloys. Most of these have been made by vapor pressure measurements using knudsen cell techniques while one has been made using gas equilibration with H2/H2O gas mixtures. In the present investigation electromotive force measurements have been made on this system using a thoria yttria solid electrolyte. This system was -chosen to establish the technique of measurement at low oxygen potentials, before extending it to Fe-Ni-Cr alloys. The cell used was Pt Cr • Cr2O3 ilThO2 • Y2O3 llalloy • Cr2O3 IPt Since the oxygen pressure of the chromium-chromic oxide mixture lay in the range 10"19 to 1027 atm in the temperature range investigated, thoria yttria electrolytes were used. It was soon found during the investigation that considerable care was required to prepare satisfactory electrolytes and a study was made to find the best method of preparing pellets of TI1O2-Y2O3 solid solution of as near theoretical density as possible. Various compositions in the range 2 to 15 mole pct Y2O3 were studied but the experimental measurements were made with electrolytes containing 4 to 5 pct Y2O3. The method of preparation finally adopted was to use thorium oxalate and yttria as raw materials. The thorium oxalate was decomposed at 900°C for 6 hr to give a reactive powder. This powder was then mixed with the appropriate amount of yttria and ground in an agate mortar for about 4 hr. The powder was moistened with absolute alcohol as a lubricant and pressed in a metal die at a pressure of 7 tons per sq in. The pellets were then heated at 1100°C for 24 hr followed by 24 hr at 1250°C. After this firing cycle the material was reground and repelleted and again fired at 1100° and 1250°C followed by heating for 7 days at 1600°C. It was found in later experiments that a further firing of 4 hr at 2000°C improved the density further and gave pellets of only 2 pct porosity. After firing, the pellets were polished on emery paper and finally with the 2 µ diamond paste. The polishing was followed by a final firing of 24 hr at 1500°C. Investigation showed that this final treatment minimized reaction between the electrolytes and chromium oxide. It was believed that this was because of a lowering of the surface energy of the electrolyte surface. Two types of investigation were carried out to check that the pellets were satisfactory oxygen ion elec-

Jan 1, 1970

By J. G. Early, F. V. Lenel, G. S. Ansell

The isothermal shrinkage rates of copper-powder compacts were determined in the temperature range from 760o to 1060oC. The rates for compacts fabricated from a pure spherical copper powder were compared, on the one hand, with those fabricated from a spherical powder containing a dispersion of aluminum oxide in the copper matrix and, on the other hand, with those fabricated from a pure copper powder having irregularly shaped particles. The irregular copper-powder compacts shrank much faster than the spherical copper-powder compacts and these, in turn, shrank faster than dispersion -strengthened copper -powder compacts. The experiments were run under a small external load. By varying this load, the stress dependence of the isothermal shrinkage rate could be measured for compacts from spherical-copper powders. This stress dependence was found to be approximately proportional to the 4th (dispersion-strengthened alloys) or 4.5th (pure metals) power of the stress. It is concluded that the material-trans port mechanism is due to dislocation climb rather than to Herring-Naharro creep. In this investigation the isothermal shrinkage rates of compacts made of copper powder with an average particle size of 70 g in the temperature range from 780o to 1060°C were studied. The aim of the investigation was to determine the mechanism of material transport responsible for the shrinkage of these compacts. Since the phenomenon observed in this investigation was shrinkage of compacts at high temperatures, it will be necessary to consider all the material-transport mechanisms possible in sintering. It is unlikely that evaporation and condensation, surface diffusion, or even volume diffusion of vacancies, in which surfaces of small radius of curvature are vacancy sources and those of larger radius of curvature vacancy sinks, will materially contribute to a sintering process involving shrinkage, i.e., one in which the centers of adjacent particles move towards each other. The two principal mechanisms which were considered were the so-called Nabarro-Herring mechanism and a plastic deformation mechanism by dislocation movement. Both of these mechanisms were originally conceived to explain creep, i.e., the plastic deformation of metals at elevated temperatures under a constant applied stress. The shrinking of metal-powder compacts which was studied in this investigation can be readily treated as a creep process. This means that the shrinkage rates observed are considered as creep rates under the influence of an applied stress. It is true that the stresses in sintering experiments are generally not constant but decrease continually with time, but, by considering creep rates or shrinkage rates over a relatively short time interval, the small changes in stress during this time interval can generally be neglected. Nabarro-Herring microcreep is a high-temperature deformation mechanism which is caused by a diffusional flow of vacancies from boundaries under a tensile stress towards those under a compres-sive stress. Herring1 has shown that in this type of microcreep the strain rate e is proportional to the applied stress and follows an equation of the form ma, aexp(-Q/RT) [l| where <j is applied stress, Q is the activation energy for self-diffusion, R is the gas constant, and T is the absolute temperature This type of creep can explain the shrinkage of a porous compact. The surface of the pores are subjected to the tensile stress of surface tension, while the grain boundaries are under a compressive stress also due to surface tension. Vacancies diffuse from the pore surface through the lattice to the grain boundaries, where they are eliminated. In order to maintain vacancy equilibrium at the pore surface the vacancies leaving the region of the pore surface are replenished from the volume of the pore, which means that the pores shrink.

Jan 1, 1964