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NAME AND OBJECT. SEC. 1. This Institute is incorporated under the Membership Corporations law of the State of New York ; its corporate name is American Institute of Mining Engineers; 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 Members. All members shall be equally entitled to the privileges of membership, except¬ing that Honorary Members, Junior Members, 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. The following classes of persons shall be eligible for membership in the Institute, namely : as Members, all professional mining . engineers, geologists, metallurgists, or chemists, and all persons actively engaged in mining and metallurgical engineering, geology, or chemistry ; as Associates, all persons desirous of being connected with the Institute who in the opinion of the Board of Directors are suitable.

Jan 1, 1923

By D. T. Camp, J. A. Clum

WITH experience, the indexing of back-reflection Laue patterns for cubic crystals is usually a straightforward matter. However, for noncubic systems where in general tables of interplanar angles are not available or when inexperienced personnel are involved, properly indexing a crystal is normally a tedious, time-consuming affair. The authors have written a digital-computer program which calculates and prints out a table of interplanar angles for an arbitrary crystal structure and then systematically searches through this table attempting to obtain a consistent fit to back-reflection Laue data supplied by the user. Of course it is possible simply to calculate a desired set of interplanar angles should this be the only requirement. The program was originally written in ALGOL for use on a Control Data G-21 computer and a preliminary version is described elsewhere.' A version in FORTRAN TV which has been run on an IBM 7040 and one in FORTRAN II run on an IBM 1620 are also available. While it is not the purpose of this communication to describe the program in detail, an indication of how the program operates may help one to decide whether the program would be useful under particular circumstances. The hypothetical example chosen involves the indexing of the zinc crystal which has a hexagonal structure and a c/a ratio of 1.856. Input data for this example consists of only nine cards: The first card controls the list of families of planes to be considered. If the Miller indices of a prospective family are (h k 1), the first digit places a limit, in this case 3, on the absolute values of any individual index. The second digit limits the sum (h +k + l + 1), in this case to 5. Therefore the families (013) and (112) are considered but the family (122) is not. All systems, including the hexagonal, are treated using three-digit Miller indices. The second card contains the crystallographic parameters required to define the system. The first digit is 0 if the system is rectangular, that is all angles are 90 deg. If the system is nonrectangular, as in this example, the digit 1 so indicates. The next three values are the lengths of the sides of the reference cell, a, b, c. The final three entries on the card are the angles, a, 8, y between the sides. The digit 0 on the third card indicates that more data follows, the digit 6 shows that a set of six planes is to be indexed, and the value 2.0 indicates that agreement between a tabulated angle and an input angle is considered adequate if their absolute difference is less than 2 deg. The fourth card lists the measured input angles. a12 = 38 deg, a13 = 24 deg, and so forth. More input angles are listed on cards five to eight with the latter giving the input angle 058 = 40 deg. The final card in this series indicates by the digit 1 that no more data is to be processed. The program operates in the following manner: 1) The input data from cards 1 and 2 is read, thereby indicating the crystal system to be examined and the crystallographic parameters of the particular crystal to be indexed. 2) A table of interplanar angles is generated. 3) The input angles contained on cards 4 to 8 are read. 4) The measured input data is systematically compared with the generated table of standard angles. All combinations of poles whose interplanar angles compare with the requisite, specified accuracy are printed out. If no adequate fit is found an indication of this is printed. A typical line from the generated table of interplanar angles is: (001) - (111) 74.92 64.99 It can be shown that in the Miller-Bravais system the first value corresponds to the angles between (0001)-(1121) and the second to that between (0001)-(1011). This particular set of data gave the following result for the planes 1 to 6: (011) (110) (121) (111) (112) (010) The ALGOL version also has the option of simply providing the Greninger-chart coordinates ? and 6

Jan 1, 1967

By Arnulf Muan, W. C. Hahn

Activities of "FeO" in "FeO"-MgO solid solutions have been determined in the temperature interval 1100" to 1300"C by equilibrating oxide samples with pure metallic iron in atmospheres of known oxygen partial pressures. The end members "FeO" and MgO form a complete solid solution series. The system shoztls a considerable positive deviation from Raoult's law. A large amount of data has been accumulated during the past 50 years on stability relations existing among various crystalline and liquid phases in oxide systems of importance in metallurgical processes. By comparison, very little is known about the thermodynamic properties within the various homogeneous phases present in the systems, particularly the crystalline phases. The present paper deals with determination of activities as a function of composition in solid solutions of wüstite ("FeO") and periclase (MgO). Such solid solutions ("Magnesiowustites") have simple sodium chloride structure and are known to be continuous between the end members at elevated temperatures (> 1000°C) and strongly reducing conditions such as prevail in contact with metallic iron. Wiistite has a defect structure in which the ratio of oxygen to iron varies and never has as low a value as 1:1 corresponding to the simplified formula FeO. Within the range of stability of this phase, the composition changes between a maximum oxygen content of approximately 25.6 wt pct (corresponding to the formula FeO.83O) and a minimum oxygen content of approximately 23.2 wt pct (corresponding to the formula Fe0-96O). The activity data to be reported in the present paper are based on using as standard state the wüstite which is in equilibrium with metallic iron at a particular temperature, i.e., "FeO" of minimum oxygen content at each temperature. For sake of simplicity, the formula FeO will be used to represent the wustite phase in the remainder of this paper. Consider the reaction 2 FeO solid solution = 2 Femeta1 + Ozgas in which FeO dissolved in another oxide dissociates to metal and gas. The activity of FeO in the solid solution, relative to "pure FeO" in equilibrium with metallic iron as a standard state, is then expressed as Here Po, and po2 * are the partial pressures of oxygen of the gas phase in equilibrium with metallic iron and the oxide solid solution or "pure FeO", respectively. The equation holds only if the metal phase is essentially pure Fe. The oxide combinations used in the present investigation fulfill this requirement. Magnesia is so much more stable than FeO that for all practical purposes aFe = 1 in the metal phase present in all mixtures used in this investigation. I. EXPERIMENTAL METHOD 1. General Procedure. The experimental method used was analogous to that developed by Foster and Welch1 and used recently by Hahn and Muan.2 It consisted of four steps: a) Determination of the equilibrium between FeO and Fe. b) Determination of d-spacing change with composition in oxide solid solutions (FeO-WO). c) Equilibration of various members of the solid solution series with a metal phase and a gas phase of known oxygen partial pressure. d) Determination of compositions of solid solutions resulting from step (c) by means of data obtained in step (b). 2. Starting Materials. General Aniline and Film Corp. carbonyl iron powder, type HP, was used as the source of metallic iron. A representative analysis was as follows: Fe, 99.6 pct; C, 0.01 to 0.04 pct; 0, 0.1 to 0.3 pct; N, 0.00 to 0.05 pct. Other elements were present as traces only. Wiistite was prepared from "Baker Analyzed" Reagent grade Fe2O3 of reported analysis as follows: Assay (Fe2O3), 98.8 pct; insoluble in HC1, 0.02 pct; not precipitated by NH,OH, 0.02 pct; Zn, 0.003 pct. Production of wüstite was achieved by a two-step

Jan 1, 1962

By H. L. Riley, B. W. Gandrud

AS far as the present writers know, this system of flotation has not been used elsewhere in this country, but in the last couple of years it has been introduced, with minor variations, at one plant in England and one in Wales.' The system has been described and discussed in a number of publications.2-5 The following is quoted from an abstract of the latest of these,5 a paper presented at an International Conference on Industrial Combustion in 1952. On the basis of experience to date with the commercial plants, it is believed that the kerosene-flotation process incorporates all the necessary elements to make it greatly superior to anything else now available for treating of fines in wet processes of coal preparation. Additional study and investigation are still needed, however, to determine if certain phases of the process can be improved to such an extent as to make it generally satisfactory and acceptable to the industry. Further improvements will be needed with respect to the capacities of the flotation cells and the reagent consumption. The situation referred to above explains why an investigation is being made of the possibilities of achieving better cell capacities. Results obtained from this investigation, which is still in progress, are believed significant with regard to both cell capacity in general and the relation of cell design to cell capacity in particular. Commercial equipment now being used in a laboratory-type investigation should have performance characteristics similar to those of the larger machines. Equipment and Procedures: All flotation tests have been made in a standard Denver sub-A 24x24-in. unit cell of 12-cu ft volume. Cell modifications to make it more suitable for the tests were an adjustable front-wall section for varying cell depth and a perforated scraper-drag assembly for removal of the float product. There is also an apron dry-coal feeder, a gravity-feed water supply, reagent feeders, and a centrifugal pump that feeds the mixture of coal, water, and reagents into the flotation cell. A wattmeter connected into the drive-motor circuit records the power requirements of the impeller throughout each run. Dry coal, water, and reagents are all fed through a pan-type intake to the feed pump. A Sturtevant blower was set up to furnish air for supercharging. A centrifugal pump with a garbage-can intake provides for disposal of refuse flow to an outside settling tank. Figs. 1 and 2 show the flotation cell; Fig. 2 also illustrates the blower for supercharging. For purposes of this investigation, the percentage by weight of the feed coal recovered in the float product under a standard set of conditions has been considered as the criterion of cell capacity. The authors realize that such a criterion may be somewhat unorthodox, as the term cell capacity is usually understood to refer to feed input and ordinarily takes into account the ash analyses of the float product and refuse. However, the word capacity is flexible enough so that Webster gives one definition as maximum output, a definition which seems to justify, at least partly, acceptance of the above criterion. It has been the authors' experience in the Birmingham district that the ash-reduction efficiency of the coal-flotation process is generally satisfactory and that the only real problem is to increase the rate of float recovery so that the feed rate to any given bank of cells can be increased without undue loss of coal in the refuse. Originally it was planned to operate the flotation cell to simulate continuous operation during sampling periods. It was assumed that operating for reasonable time with feed coal, water, and reagents turned on would stabilize conditions so that the weight of float coal discharged during a fixed time interval would be an accurate measure of the rate at which the coal was being floated. It developed, however, that this supposition was erroneous. The float coal, caught for fixed time intervals and weighed, gave widely varying results in duplicate runs. Efforts to correct this trouble failed, and it was decided to try to operate on a batch-test basis, whereby all the float coal produced during a run on a known weight of feed coal would be caught in tubs, dewatered, and weighed. This method gives consistent and reproducible results, with total float product weight rarely varying by more than 3 or 4 pct on duplicate runs. The standard test procedure is as follows: A 132-lb sample of dry feed coal is weighed and placed in the feed hopper. The feeder is adjusted for a rate of 800 lb per hr. Feed water and reagents are turned on, and the feed and refuse pumps are started. One minute later the impeller is started. Six minutes are allowed for the cell to fill up with the water-reagent mixture. The feed of dry coal is started at the end of this 6-min period. One minute later the float-coal removal drag is started. The float coal is caught in one tub for the first 6 min after the flow of feed coal starts. Tubs are then changed, and the float coal is caught in a second tub until the feed coal runs out, when the tubs are again interchanged to catch the float coal for the remainder of the run in the first tub. The cell is kept running for 3 min with the water and reagents on after the feed stops to allow residual float coal to be removed. At the end of a test the wet float coal in both tubs is weighed and the total weight recorded. The product in the second tub is used for moisture determination and screen-size analyses. When the

Jan 1, 1956

By John D. Hanks

INTRODUCTION Chevron's Uranium Mill is located near Panna Maria, Texas; 70 miles southeast of San Antonio. Designed by Kaiser Engineering, the Mill will process a nominal 2500 dry T.P.D. of uranium bearing ore containing 15% uncombined moisture. Earl Torgerson, San Mateo, California, is the consulting metallurgist on process design. Feed to the plant consists of a mixture of high, medium and low grade sandy day ore; the average grade of the ore will be 0.7% throughout the life of the project. The ore is delivered to the mill via truck and stored by type in individual piles on a flat storage area. The ore is fed by conveyor to a semi-autogenous grinding mill. The SAG mill discharge slurry is pumped either to a storage tank or to the first of five mechanically agitated leach tanks where both H2SO and NaC103 are added. Following leaching, the slurry is mixed with thickener No. 2 overflow before being pumped to a six-thickener, countercurrent decantation circuit where the solution containing uranium is separated from the leach residue. The residue is washed essentially free of solubilized U308 values at the sixth thickener and discharged to an adjacent tailings pond. The first thickener overflow, containing approximately 0.4 grams U308 per liter, is filtered for clarification and sent to the liquid ion exchange (solvent extraction) section. The pregnant aqueous solution is mixed with organic solvent containing amine on which the complex uranyl sulfate ions are absorbed. The immiscible aqueous and organic solutions are mixed and separated in each of the four stages of solvent extraction. The final pregnant organic is directed to a stripping section. A strip solution containing (NH4)2S04 (ammonium sulfate) and NH4C1 (ammonium chloride) is contacted and separated from the organic in each of the four mixer-settler units. The strip solution is mixed with NH3 and the uranium precipitates along with trace amounts of sulfate, chloride, and ammonia. After washing in a thickener, the uranium precipitate or yellowcake is centrifuged and dried in a multiple hearth roaster. Overflow from the yellowcake thickener is recycled back to the stripping section of the solvent extraction circuit. The dried yellowcake concentrate contains more than 98% U308; diluents include H20, ammonia, chloride, etc.. Impurity concentratons in the product are sufficiently low, after drying, to permit direct shipment to refinement installations. ORE RECEIVING Ore from one or more mines will be stored on a pad adjacent to the uranium processing mill. Ore stored in the area may total 200,000 tons or more, which is equivalent to over a two month treatment reserve for the mill. In addition to providing surge capacity, the proposed storage facility permits natural oxidation of the ore and affords an area which, in turn, improve plant U308 recovery and reduces consumption of oxidizing reagents. The uranium bearing ore is segregated at the storage area into several distinct types. One group can be identified by its sandy, clay-like matrix and will be distinguished by its U308 concentration of high, medium, or low. Other ores contain substantial quantities of carbonaceous shale, typical of ore in the area. Feed is recovered from any one or combination of piles, in accordance with operating requirements and recovery factors. The run-of-mine ore, at 15% moisture, is transferred to a stationary, 24-inch by 24-inch grizzly. Undersize material falls into a 280-ton surge hopper located below grade, while oversize material is removed for preliminary size reduction. A sump-pump is located at the reclaim hopper to recover excess mositure and ore spillage for processing. Plant feed is continuously drawn from beneath the hopper at an average rate of 120 tons per hour by means of an apron feeder. The hopper is equipped with a low level safety system to protect the apron feeder assemble. The ore is transferred to a belt conveyor, elevated and discharged into a semi-autogeneous grinding (SAG) mill. Water addition at the discharge point is automatically controlled by a feed rate, moisture analyzer system located on the belt conveyor. GRINDING Minus 24-inch ore, at a rate of 120 mosit tons per hour and cyclone underflow at a rate of 137 moist tons per hour are combined with a controlled 186 gallons per minute fresh, warm plant process water. The mixture, at an average 68% solids, is fed to the SAG mill. The viscous mass is subjected to grinding at a temperature moderately above ambient resulting from use of the 140ºF well water. The 16.5'x5.0' Marcy Mill is driven by a 500 H.P. A.C. Motor. The mill will rotate at 73.4% of critical speed or 14.06 R.P.M. Eight percent of the mill's volume will be occupied by steel grinding balls; 22% by pulp. The mill is designed to rotate in either direction, in order to obtain maximum lifter and liner life. The SAG mill is equipped with a trommel having ½ -inch slots for removal of oversize material at its discharge and the large or oversize material is collected and either recirculated periodically or discarded. The undersize slurry is diluted to 64% solids with the addition of hot well water or mine water to the mill discharge sump. The flow is automatically controlled utilizing data from a gamma density meter. The slurry from the mill discharge sump is pumped to hydrocyclones to classify the particles. Underflow from the cyclones, at 80% + 28 mesh, is recycled to

Jan 1, 1979

By Peter L. Raffo, Walter R. Witzke

RHENIUM has a melting point of 5750°F, the second highest value among the metals.1 Its refractory nature should thus make it a useful material at high temperatures. The only available data on the high-temperature properties of bulk rhenium are from the work of Flagella and Tarr.2 These authors have shown that, over a wide range of temperatures and stresses, no appreciable linearity exists in a log-log plot of the minimum creep rate or rupture life vs stress. The purpose of the present letter is to describe some exploratory experiments designed to study the creep rate-stress-temperature relation and other facets of the creep of rhenium. The tests were carried out in vacuum in the temperature range 2200" to 4000°F (0.43 to 0.72 Tm,) at stresses ranging from 4 to 20 ksi. The powder metallurgy rhenium was purchased as 0.04-in.-thick sheet. The sheet is nominally 99.97 pct pure with the major impurities being iron (70 ppm) and molybdenum (25 ppm). The sheet was machined into specimens having a 1-in. gage length and a 0.25-in. gage width. The specimens were preannealed in hydrogen at 3000°F for 1 hr. This treatment produced

Jan 1, 1970

By Vernon F. Swanson

An equation has been developed which will permit the direct determination of free settling velocity of any sized particle encountered in mineral benefici-ation. The equation is based on Newton's Law with the friction factor defined as a function of the laminar boundary layer. Development of the equation is described first from an empirical trial-and-error approach. Examples are given of actual settling data from the literature versus calculated data to prove the validity of the new equation. Newton's law of resistance of a fluid to flow around a sphere is expressed in Eq. 1. Free settling velocities can be determined by setting Eq. 1 equal to the gravitational force on a particle and solving for V. Experience has shown that when the solids are greater than 1 mm in diam, or when Reynold's number for the system is greater than 800-1000, the friction factor term, fD, essentially becomes a constant, and Eq. 2-b becomes Newton's equation for free-settling and is applicable under the conditions stated. Another special condition exists for particles smaller than 0.05 mm or when Reynold's number for the system is less than three. In this case fD=24/NRor24µ/pL,dV,. Eq. 2-b can then be rewritten as

Jan 1, 1968

By W. L. Freyberger

In a study of the kinetics of flotation of coal, employing a single-cell continuous flotation system, Jowett and Safvi1 pointed out that it was experimentally impossible to float all of the coal which could be floated in a batch tests. On the other hand, deBruyn and Modi, in a discussion of their paper, suggested that under certain conditions the recovery of floatable material from a similar one-cell continuous system would have been 100%.2 Both of these pairs of authors make the assumption that a state of perfect mixing exist in the flotation cell. This assumption, usually made by investigators studying flotation kinetics in a continuous system, has been confirmed experimentally on several occasions. The work is summarized by Arbiter and Harris.3 It is the purpose of this note to demonstrate that, for conditions of perfect mixing, one of two conditions exist: 1) the recovery of floatable material is 100%, and the flotation is infinitely fast, or 2) the recovery of floatable material must be less than 100%. This theorem can be proved simply in the following manner: 1) If recovery, R, of the floatable material is 100%, the concentration in the tailing must be zero. This follows from the definition of recovery as

Jan 1, 1964

NAME AND OBJECT. SEc. 1. This Institute is incorporated under the Membership Corporations law of the State of New York ; its corporate name is American Institute of Mining Engineers; 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 Members. All members shall be equally entitled to the privileges of membership, except¬ing that Honorary Members, Junior Members, 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. The following classes of persons shall be eligible for membership in the Institute, namely : as Members, all professional mining . engineers, geologists, metallurgists, or chemists, and all persons actively engaged in mining and metal¬lurgical engineering, geology, or chemistry ; as Associates, all persons desirous of being connected with the Institute who in the opinion of the Board of Directors are suitable.

Jan 1, 1923

By Robert K. Steunenberg, Irving Johnson, James B. Knighton

The activity coefficient of plutonium in liquid magnesium, over the temperature range 650° to 800°C, was obtained from measurements of the distribution of plutoninm between a 50 mole pct MgC12-30 mole pct NaCl-20 mole pct KC1 molten-salt mixture and liquid Zn-Mg alloys. For dilute solutions (0.08 at. pct Pu) the activity coefficient of plutonium was found to vary from 10.1 at 650°C to 12.2 at 800°C. The activity coefficients of plutonium in dilute liquid solutions of plutonium in uranium, silver, lanthanum. cerium, and calcium were estimated to be. The distribution data indicate a value of about 0.1 at 800°C for the activity coefficient of PuCl3 dissolved in the above ternary salt mixture. LIQUID magnesium and several liquid alloys of magnesium with metals such as zinc and cadmium have been shown to be useful solvents in pyrochemical processes for the recovery of uranium and plutonium from discharged nuclear fuels,' and for the separation of transuranium elements.' The present study was undertaken to determine the activity coefficient of plutonium in liquid Pu-Mg alloys in support of process-development work. The activity coefficient of plutonium in liquid magnesium was determined from experimental data on the distribution of plutonium between a liquid ternary MgC12-NaC1-KC1 salt mixture and various liquid Zn-Mg alloys. The distribution data were used to calculate the ratio of the activity coefficients of plutonium in liquid zinc and in liquid magnesium. The activity coefficient of plutonium in liquid magnesium was then computed from the known activity coefficient of plutonium in liquid zinc. It was not necessary to know the thermodynamic properties of the molten-salt system explicitly. The major features of the Pu-Mg system have been reported by Schonfeld.3 At the temperatures of interest in the present study, i.e., above about 600°C, the phase diagram indicates the existence of a wide liquid-miscibility gap, with the plutonium-rich liquid containing about 8 at. pct Mg and the magnesium-rich liquid containing about 10 at. pct Pu at the intersection with the solidus regions. Additional data on the compositions of the two equilibrium liquid phases obtained in this laboratory4 have defined the miscibility gap up to the consolute temperature (at about 1040°C). EXPERIMENTAL PROCEDURE AND RESULTS Materials. The 50 mole pct MgC12-30 mole pct NaC1-20 mole pct KC1 salt mixture was prepared by melting the required proportions of reagent-grade NaCl and KC1 with anhydrous MgC12. The molten salt was then purified by contacting it with liquid Cd-30 wt pct Mg alloy (at 450°C) to reduce oxidizing impurities, followed by filtration through a stainless-steel frit (pore size, 65 µ) to remove solid MgO formed during the reduction. The purity specifications of the zinc, magnesium, and plutonium were 99.999, 99.8, and 99.85 pct, respectively. Apparatus. The liquid salt and metal were contained in a tantalum crucible inside a graphite secondary vessel. The crucible assembly was located inside a resistance-heated stainless-steel furnace tube. The furnace tube was closed by means of a stainless-steel cover, which was attached by bolts, with a neoprene O-ring serving as a gas-tight seal. The top of the furnace tube was water-cooled to protect the O-ring. The furnace-tube cover was provided with a tantalum thermowell, a tantalum stirrer, and a port through which sampling tubes could be inserted and materials could be added to the melt without admitting air to the furnace tube. Vacuum and an argon atmosphere were available through a side-arm on the furnace tube. The furnace temperature was regulated by a proportional controller that was actuated by a chromel-alumel thermocouple between the furnace tube and the heating elements of the furnace. The melt temperature was measured by means of a chromel-alumel thermocouple in the tantalum thermowell. The accuracy of temperature measurement was ±3°C. The salt and metal phases were intermixed by a motor-driven tantalum paddle positioned at the liquid interface. The tantalum crucible was provided with four baffles to increase the turbulence. The sampling tubes consisted of 1/4-in.-OD tantalum tubing that terminated in a tantalum frit (Kawecki Chemical Co.; average pore size, 30 µ). Procedure. The zinc, magnesium, plutonium, and salt were charged to the tantalum crucible; then the system was evacuated and filled with argon. The melt was brought to the desired temperature, and agitated for 1 to 2 hr. After allowing the salt and metal to separate, both phases were sampled. Filtered samples were obtained by immersing the end of the sampling tube in the liquid and increasing the argon pressure sufficiently to force the liquid salt or metal through the frit into the tantalum tube. The sample was then partially withdrawn into the cooler portion of the furnace tube and permitted to solidify before being removed. The temperature sequence for sampling at each magnesium concentration was 800°, 700°, 600°, 650°, and 750°C. The composition of the liquid-metal phase was varied by incremental additions of magnesium in a series of experiments at low magnesium

Jan 1, 1967

By C. L. Sollenbeger, R. B. Greenwalt

PERHAPS the most widely used dispersants or gangue depressants in nonmetallic flotation are sodium silicates, which vary in silica-to-soda ratio from 1 to 3.75. Typical manufactured silicates in order of decreasing solubility and increasing amounts of silica are Metso, silica-to-soda ratio of 1.00; D, 2.00; RU, 2.40; K, 2.90; N, 3.22; and S-35, 3.75.* References in flotation literature1,2 to the use of sodium silicates are often weak because they fail to mention the type of silicate used. Metso and silicate N have occasionally been mentioned, but when the type of silicate is not mentioned, it is usually assumed to be N, the cheapest of the soluble silicates and the one recommended by sodium silicate manufacturers as a flotation agent. In the All is-Chalmers Research Laboratories a systematic study was made of the effect of different alkali-silica ratios on the concentration by flotation of two scheelite ores. One of these was a high grade ore from the Sang Dong mine in Korea. The effect of such factors as pH; addition agents; and conditioning time, temperature, and pulp density on the flotation efficiency of this ore have been described previously. The other ore was a low grade ore from Getchell Mines Inc., Nevada. The mineralogy and techniques of concentrating this ore have been described by Kunze. Hereafter these ores will be referred to as the Korean and Nevada ores. Experiments were made with both to determine the effect of three factors—-type of silicate, concentration of silicate, and pH of the pulp—on recovery and grade of tungsten in a rougher concentrate. Average WO, content of the Korean ore was 1.50 pct and of the Nevada ore 0.27 pct. The predominant tungsten mineral in both ores was scheelite, which was accompanied by a small amount of powellite. The powellite and scheelite were finely disseminated through both ores and required a —200 mesh grind for liberation. Major gangue minerals in the Korean ore, in decreasing order of abundance, were amphi-boles, quartz, biotite, garnet, fluorite, and calcite. Bulk sulfides composed about 3 pct of the total weight. Gangue in the Nevada ore, in descending order of abundance, was garnet, alpha quartz, calcite, phlogopite, wollastonite, and amphiboles. Sulfide minerals were 3 to 4 pct of total weight. Batch flotation experiments were made with 500-g samples of ore, each sample wet-ground to 90 pct passing 200 mesh. The finely ground ore was floated in a Fagergren batch cell at 25 pct solids. The natural pH of the Nevada ore was 8.9 and of the Korean ore, 8.5. The D, RU, K, N, and S-35 sodium silicates were obtained in colloidal dispersions with varying amounts of water. The most alkaline, Metso, was in dry powdered form. For convenience in addition, 5 pct solutions by weight were prepared from each of the silicates, on the basis of dry sodium silicate dissolved in the correct amount of distilled water. Chemical analyses of the various silicates are given in Table I, together with the pH of the 5 pct solutions. A preliminary bulk sulfide float was made with secondary butyl xanthate as the collector and pine oil as the frother. The WO] analysis of the sulfide concentrate was nearly 1 pct for the Korean ore and about 0.1 pct for the Nevada ore. The tungsten contained in the sulfide concentrate constituted about 3 pct of the total tungsten in each ore. No effort was made to recover these tungsten values. The scheelite was floated with oleic acid. Adjustments in pH were made with sulfuric acid or sodium carbonate. A 1 pct solution of 85 pct Aerosol OT was sprayed on the froth and sides of the cell during the scheelite float to aid in dispersing the minerals and to decrease the entrapment of gangue particles. Six tests were planned for each of the six types of silicate in which concentrations of 1, 2, and 4 1b of silicate per ton of dry ore were investigated at both 6.5 and 10 pH. All tests were made at room temperature. The performance of each silicate was judged from the grade and recovery of WO, in the scheelite rougher concentrate. Tungsten recovery was calculated on the basis of the scheelite remaining in the ore after the preliminary sulfide float. Testing of each silicate at three levels of concentration and two levels of pH required 36 tests with each scheelite ore. Variance analyses were performed on the concentrate grades and recoveries to determine whether or not the type of sodium silicate, the concentration of sodium silicate, or the pH significantly affected recovery or grade. Results Concentrate Grade: A variance analysis of the concentrate grades for the Korean ore showed that concentration of the silicate and pH of the ore pulp were major factors in producing a high grade concentrate. Also, the silica- to-so da ratio was important as an interaction with pH. The concentrate grade vs silica-to-soda ratio is plotted in Fig. 1. The curves show that the concentrate grade improved with an increase in concentration of sodium silicate and also

Jan 1, 1959

By T. R. A. Davey

IN 1947 the author became interested in the fundamental aspects of the desilverizing of lead by zinc, conducted some experimental work, and searched the technical literature for all available fundamental data. Since then a revival of interest in the subject in Europe resulted in the appearance of quite a number of papers. It became evident that it would be more profitable to collect together and examine thoroughly the results of various workers, than to attempt to duplicate the experimental determinations. There are many inconsistencies in the various publications, and it is opportune to review at this time the present status of knowledge on the Ag-Pb-Zn system. There is also a need for a clear description, in fundamental terms, of the various desilverizing procedures. This paper is presented in four sections: 1—There is an historical review of the origins of the Parkes process, of the results of many attempts to find a satisfactory fundamental explanation for the phenomena, and of the modifications proposed to date. 2—A diagram of the Ag-Pb-Zn system is presented. This is believed to be free of obvious inconsistencies or theoretical impossibilities, although thermodynamic analysis subsequently may reveal errors. 3—The fundamental bases of the various desilverizing procedures, which have been used up to the present day, are described; and a new method is suggested for desilverizing a continuous flow of softened bullion in which the bullion is stirred at a low temperature in two stages producing desilverized lead at least as low in silver as that from the Williams continuous process and a crust which, on liquation, yields a very high-silver Ag-Zn alloy. 4—A suggestion is made for the revival of de-golding practice, following a recently published account which does not seem to have attracted the attention it deserves. The terms "eutectic trough" and "peritectic fold" as used in this paper are synonymous with "line of binary eutectic crystallization" and "line of binary peritectic crystallization" as used by Masing.' The German literature on ternary and higher systems is rather extensive and a fairly general system of nomenclature has arisen, whereas in English usage the corresponding terms are not as well established. For this reason the meanings of terms used in this paper, together with the equivalent German terms, are given as follows: 1—Eutectic trough—eutektische rinne: line at which a liquid precipitates two solids S1 and S2 simultaneously. If the composition of a liquid which is cooling reaches this line, it then follows the course of this line until a eutectic point is reached, or until all the liquid is exhausted. The tangent to the eutec-tic trough cuts the line joining S1S2. 2—Peritectic fold—peritektische rinne: line at which a solid S1 and a liquid L transform into another solid S2. If the composition of a liquid which is precipitating S1 reaches the line, on further cooling only S2 is precipitated. The liquid composition moves from one phase region (L + S1) into the other (L + S2), and does not follow the course of the boundary. The tangent to the peritectic fold cuts the line S1S2 produced nearer S,. 3—Liquid miscibility gap, or conjugate solution region—mischungslucke: the region within which two liquid phases coexist in equilibrium over a certain range of temperature. A system whose composition is represented by a point in this region comprises one liquid at high temperature; then as the temperature is progressively reduced, two liquids, one liquid and one solid, one liquid and two solids, and finally three solids. 4—Liquid miscibility gap boundary—begrenzung der flussigen mischungsliicke: the line along which the surface of the miscibility gap dome, considered as a solid model, intersects the surrounding liquidus surfaces. 5—Tie lines—konoden: lines joining points representing the compositions of two liquids, a liquid and a solid, or two solids, in equilibrium. In binary systems the only tie lines customarily drawn are those through invariant points, e.g., through the eutectics of the Pb-Zn and Ag-Pb systems, or the various peritectics of the Ag-Zn system, as in Figs. 1 to 3. In ternary systems it is desirable to draw sufficient tie lines to indicate the slopes of all possible tie lines. 6—Ternary eutectic point—ternares eutektikum: point at which liquid transforms isothermally to three solids, S1, S2, and S Such a point can lie only within the triangle 7—Invariant peritectic (transformation) point— nonvariante peritektische umsetzungspunkt: (a) — On the miscibility gap boundary, the point at which two liquids and two solids react isothermally so that L, + S, + L, + S2. (b)—On the eutectic trough, the point at which a liquid and three solids react iso-thermally so that L + S, + S2 + S3. Such a point must lie on that side of the line joining S,S which is further from S,. (c)—A further possibility, not found in this ternary system, is that the point is at the intersection of two peritectic folds when the reaction concerned is L + S, + S, + S Historical Introduction Karsten discovered in 1842 that silver and gold may be separated from lead by the addition of zinc.2 Ten years later Parkes used this fact to develop the well known desilverizing process which bears his

Jan 1, 1955

By G. S. Cole, H. Biloni, G. F. Bolling

An experimental investigation of sovlze low .melting point alloys sJtows that a substvucture of isolated depressions can be the first manvestation of constitutional supercooling on solid-liquid interjaces veuealed by decanting. Electron-tni cvop vobe and wletallo gvaplic esanzinations, in tlze bulk belzind the interjace, oj the segregation associated with these isolated areas substantiate tlzei&apos;v depressed nature, since a solute of ko < 1 is enriched, and a solute of ko > 1 depleted. In contrast, the pox structuve, a set of projections often veported in the literature, leaves no trace oj. segvegation. These obserl;atims, accovlrpanied by a brief review of recent literature, point to inconsistencies between experirrental obsevvation and the idea that the fornzation of a projection is a causal step in the development of a cellular substructure. An argument is presented to show instead how it is plausible for substantial depvessiom to form in the pvesence of constitutional supercooling at dislocations threading the solid-liquid interjace. THE development of constitutional supercooling during growth from the melt leads to the formation of the cellular solidification substructure. This well-founded association between structure and instability has been basic in understanding cellular substructure and micro segregation; however, the initial formation of structure seems unclear. Rutter and Chalmers,&apos; in definitive experiments and theory, noted that in the presence of constitutional a planar interface might break down: "resulting in the formation of a small projection on an initially plane or uniformly curved interface." That is, the breakdown from a planar to a cellular interface was implied to be initiated via a projection into the unstable liquid. Later, Walton et (11. found that a structure of isolated projections, termed "pox", appeared at solid-liquid interfaces decanted under growth conditions near the onset of constitutional supercooling; the pox were taken as the indication of the instability promoted by the supercooling. Tiller and Rutter4 in their extensive work studied the shape transitions at decanted interfaces which were generally observed to proceed as— pox, "irregular cells", elongated cells, regular (hexagonal) cells, and so forth. The pox varied in size from lo-&apos; to 1CT4 cm, and tended to disappear as cells increased in number and regularity, but as noted,4 the first real array of cells did not seem to be a development from the pox. In fact these authors implied a lack of connection because they stated that the pox are denser on "irregular cells", and as cell boundaries increase in number (i.e., the cells become smaller) there is less need for the pox which do dis- appear. Thereafter, most authors dealing with either experiment or theory have accepted the reality of pox and have used them as a criterion for the onset of constitutional supercooling. In contrast, Spittle, Hunt, and smiths have now suggested that pox are irrelevant artifacts comprised of such things as entrapped oxide. This proposal invokes the observations of weinberg6 and chadwick7 each of whom have shown that the act of decanting leaves a residual liquid on a decanted interface; the remnant solid layer of the order 10 p may thus contain particles that might have been transported from the external surfaces, or elsewhere, during decanting. With the incentive of this suggestion,= some further experiments and a reexamination of the literature have been conducted, in order to question the validity of pox as evidence of an instability and to examine the initial development of the cellular substructure. 1) EXPERIMENTS Single crystals of zone-refined tin (-99.9999 pct) were grown from the melt in a controlled fashion with various, small concentration additions of lead and antimony, for which ko < 1 and > 1, respectively. The crystals were decanted at conditions near the onset of constitutional supercooling and were thus appropriate for observation of slight perturbations. It was possible to observe two types of small departure from smooth or "planar" interfaces in both cases of lead or antimony additions. Some were projections and others, if in regular array of any type, were depressions. The crystals were etched with suitable reagents progressively dissolving the decanted interface surface; projections left no record, but depressions were continuously associated with spotlike areas contrasting with the rest of the interface. Traverses were made with the beam of an electron microprobe across the regions of contrast; with lead addition the persistent spots were lead-rich, and with antimony addition the persistent spots were antimony-poor. This is consistent only with a dominant role for depressions, because if the projections had left spots but were incorrectly catalogued, a reversed observation should have been made; that is, the Pb(ko < 1) should have been depleted and the Sb(ko > 1) enriched. In the work of Cole and inegard, and elewhere, regular arrays of structure associated with the initial stage of instability have been shown, in photographs and represented as pox or projections. We believe this to be erroneous, by inference, since whenever a regular array was observed, in the present examination, it consisted of depressions, regardless of the nature of the solute, ko 1. Fig. 1 is reproduced8 as an ideal example of the possible optical illusion involved; the observer can satisfy himself from the distribution of illuminated areas that the markings are depressions. Fig. 2 from the present investigation is an interference photograph of an interface similar to that in Fig.

Jan 1, 1967

By J. F. Myers, F. M. Lewis

DISCUSSION L. E. Djingheuzian (Canadian Dept. of Mines and Technical Surveys, Ottawa)—In their Summary the authors say: "Reconciling the grinding efficiency with good metallurgy is still a problem." In the discussion of the first paper8 in his reply to W. I. Garms, Mr. Myers states: "Our grinding process with smooth I-in. balls has reduced by nearly one half the metallic losses in the fine micron sizes of the tailing. This is simply because less of the fine micron sizes are produced. Since the + 65 mesh size is the same as formerly, a higher percentage of the intermediate sizes are developed. These sizes have the highest floatability, require the least reagents, and use less floating time. "These factors contribute so heavily to the overall economies that dropping our power grinding gain from 28 pct back to 19 pct is a small detail. However, we feel that this is only a momentary situation and that eventually the best features of the grinding and flotation processes can be brought together, which is as it should be." Italics are mine. The above statements, to me, appear to be the answer to the opening statement in the Summary. Denoting the costs at different power grinding gains as: Power Grinding Power Grinding Gain, 28 Pct Gala, 19 Pct Cost of grinding G G1 Cost of flotation F F1 Value of metallic losses T T1 where G1 > G2 F3 < F, and T1 < T, we have: G1+Fl+T1<G +F+T. Since the authors accept the idea that "grinding in flotation plants becomes part of the &apos;conditioning&apos; of the feed to flotation",4 i.e., that in flotation the ball mill is primarily a conditioning machine, it can be postulated that Tennessee Copper grinding at cost G1 is more efficient than grinding at lower cost G. This can be directly inferred from the Conclusion of the paper. Mr. Myers also emphasizes this at the end of his reply to Mr. Garms: "that grinding is for the purpose of preparing flotation feed and not grinding per se." This, to me, in the final analysis means that when the efficiency of grinding is weighted against the conditioning factor, the former becomes a function of efficient conditioning, hence, within the system in which proper conditioning is the dominant factor, the best grinding efficiency is provided by grinding which will contribute towards the optimum conditioning. This brings us again to the statement: "that if every grinding unit were considered as a conditioner for each following step, efficient grinding plants would become much easier to design."&apos; In other words, grinding equipment should be balanced against the flotation equipment and against chemical reactions taking place in the system. F. C. Bond (Allis-Chalmers Mfg. Co., Milwaukee)— The authors&apos; discussion of the probable ball motion in a slow speed high dilution mill is very interesting. When the 1-in. balls have worn down to about one fourth of their original weight they apparently first develop a flat surface; as wear progresses this flat face becomes concave, and other concave faces appear. It seems more probable that the first flat face may form at the softest part of the ball surface, and that each succeeding contact tends to force this flat face into sliding contact with a larger round ball; than that the flat faced ball tends to pair off with a particular round ball and to travel with it continuously. When the small worn ball has a flat face and is in sliding contact with a large round ball, the surrounding large balls will assume a more or less definite pattern, and slide against the worn ball, thus producing secondary concave faces. The primary concave face seems to be larger and better developed than the secondary faces. The ball charge can be divided into "concaves" which show at least one concave surface, "intermediates" which have developed flats or incipient concaves, and "rounds." Ball slippage is always present in a tumbling mill, and the mutual ball movement is necessarily a combination of rolling and sliding. The sliding motion is apparently concentrated upon the smaller worn balls which nest between the surrounding larger round balls. When each worn ball starts its upward path in the mill its primary flat or concave surface fits against a larger round ball, and the round ball slides upon it. The action may be something like that of the ball separator in a ball bearing, except that the worn sliding balls are always under considerable pressure. The material is ground under the combined influence of breakage 1—by impacts between falling balls and between falling and supported balls, 2—by being nipped between rolling balls, and 3—by being rubbed between the sliding balls. The rubbing action will be increased in the presence of worn balls with concave surfaces. The rubbing action probably produces a considerable portion of the finely ground slimes in the product. The worn balls commonly approach tetrahedrons in shape, and are very different from concavex, each of which has two equal opposed concave surfaces. Concavex were designed only to grind upon themselves, and not for use in combination with grinding balls. Their action in a grinding charge is very different from

Jan 1, 1952

By H. P. Leighly, F. C. Perkins

THERE have been several statements in the literature about the difficulty of producing single crystals of high-purity (99.99pct) by the strain-anneal method. Consequently, investigators tend to employ low-purity aluminum for their single-crystal experiments, or else resort to the Bridgman or other techniques which depend on solidification for the production of single crystals. The following paragraphs describe a solid-state method for the manufacture of single crystals of high-purity aluminum which should provide crystals with greater perfection than those formed by solidification. The starting material for this method of producing single crystals was secured from the Aluminum Corp. of America and has a purity of 99.99 pct, the balance being trace amounts of impurities. Specimens approximately 1 by 4 in. are sheared from sheet having a nominal 0.050 in. thickness. The specimens are given a preliminary anneal at 640°C for approximately 3 hr in order to remove fabrication strains and to produce an average grain size of about 1/4-in. diam. The specimens are then etched in Tucker&apos;s etchant (45 pct HC1, 15 pct HNO3, 15 pct HF and 25 pct H2O) to remove the oxide film. Critical strain is applied by wrapping each specimen about a 1 3/4 in. round and subsequently straightening it against a flat surface. The high-purity aluminum sheet is sufficiently soft that this operation can be accomplished by ordinary finger pressure. The specimens are immediately annealed again at 640°C for 3 hr and reetched for examination. Ordinarily, considerable growth of certain of the grains will have occurred, and occasionally a single crystal will be produced on the first attempt. The procedure of alternately straining, annealing and etching is repeated until the majority of a batch of specimens contains usable crystal sizes. Typical examples are illustrated in Fig. 1. The greatest changes in crystal sizes are produced in the initial treatments. As the average crystal sizes get coarser in the later treatments, the sever- ity of the strain must be increased in order to produce grain boundary- migration. This increase in severity is effected by decreasing the diameter of the round used for straining (to 1 1/2 in., for example) and/or wrapping the specimens about the round twice, with opposite faces in contact with the round, before flattening. Usually the strain treatments described are not severe enough to produce nucleation in coarse grain high-purity aluminum. The growth of grains occurs by strain-induced grain-boundary migration. It has been observed that the grain boundaries move most readily during the first hour or so of each annealing treatment and that the rate of movement decreases with extended holding times at temperature. Prolonged annealing treatments are therefore not usually beneficial. Similarly, the rate of growth of each crystal appears to depend upon the orientation of the crystal with relation to those of its neighbors. Frequently island grains are formed after the initial heat treatment as the result of slow grain-boundary migration. These sometime become stationary during later heat treatments. Twin orientation interfaces are frequently developed during annealing. These imperfections can usually be removed by increasing the severity of strain to produce actual nucleation of new grains of more favorable orientation at the imperfection interfaces. The largest single crystals produced in our laboratory by the above method measured 4 by 1 by 0.050 in. Examination of Laue back-reflection patterns from a limited area of the specimens, gave no evidence of polygonization. Probably there is some indication of polygonization in the original grain area provided a more sensitive technique is used for detection. Experiments to produce wider specimens were less successful, possibly because wider sheets increase the complexity of the strains induced by deformation and promote widespread nucleation. Grain boundary migration occurs preferentially in a direction parallel to the longitudinal axis of the specimen. The choice of specimen geometry with respect to the rolling direction of the sheet appears to be immaterial in regard to the production of single crystals.

Jan 1, 1961

By I. R. Kramer

In his technical note entitled "Specimen Temperature During Electropolishing of Aluminum Crystals" Dr. Nakada 1 reported that the temperature of his aluminum specimen increased 65°C when it was polished electrolytically; he therefore concluded that the temperature of the specimens used in our work (Nakada&apos;s Refs. 4 and 5) similarly increased by this amount. From this he further concluded that the change in plastic-flow characteristics associated with the removal of the surface by electropolishing during deformation were due solely to a temperature effect. We believe that these conclusions are not justified. It is, of course, realized that the amount of the temperature increase is associated with the total heat transfer of the entire system, including the efficiency of the stirring action, heat flow through the grip on the specimen, and so forth. Of great importance is the type of electrolyte used for polishing. For example, in Dr. Nakada&apos;s experiment a methanol-perchloric acid solution was used. This solution has a relatively low conductivity and for the current densities of 0.3 amp per sq cm the electrical input appears to be 30 to 40 w. In our research a methanol-nitric acid solution was employed and for the size of specimen employed by Dr. Nakada the power never exceeded 7 w. There are major differences in his test apparatus which also appear to affect the temperature rise. In our case a steel specimen holder was used in order to provide a good heat sink and conduction path. The cathode was a coiled aluminum tube through which a coolant was passed to provide additional cooling capacity near the specimen. The stirrer was placed in a position to obtain maximum cooling efficiency. On the other hand Dr. Nakada used a teflon grip to support the specimen at one end only and a flat copper cathode, thus providing a much less efficient heat sink as compared with the metal grips used in our studies. The relationship between current density and the temperature increase for two sizes of aluminum crystals tested in our tensile apparatus is shown in Fig. 1. The temperature measurements were determined from a thermocouple spot-welded to the surface of the specimen. The thermocouple was protected from the acid by means of a thin film of stop-off lacquer which covered just the outer surface of the thermocouple junction. Whereas Dr. Nakada reported a time delay of 8 min to obtain thermal equilibrium, our temperature change occurred in less than 15 sec. This emphasizes the importance of the heat transfer in the entire system. However, readings were taken over a period in excess of 5 min to assure thermal equilibrium. The data given in Fig. 1 were also confirmed by determining the thermal expansion of the specimen in a sensitive creep apparatus. As may be seen, for the smaller specimen the increase in temperature is 8°C for a current density of 1.5 amp per sq in. The majority of our data on the effects of surface removal on the plastic behavior of aluminum were obtained at current-density values at or below this value. Above 1.5 amp per sq in. the temperature rise is uncertain; for under certain conditions, especially in the vicinity of 2 amp per sq in., the anode can become polarized and the current flow is erratic. This condition is easily recognized because wide fluctuations in the creep and stress-strain data make it virtually impossible to obtain reliable results, a condition which was always avoided in our work. The temperature increase for the larger specimen is less than that of the smaller specimens below a current density of 1.5 amp per sq in. At a current density of 2 amp per sq in. the temperature increased 13°C. The above temperature data are in accord with observations made during creep experiments. Using a transducer capable of detecting strain in the order of 10 -6, temperature changes in excess of 8°C were not detected when the current density was increased to as high a value as 1.5 amp per sq in. Finally we wish to point out that our experimental data cannot be explained solely in terms of a heat effect. For example in Fig. 1 of Ref. 2, the change in slope of Stage II as a function of the rate of re-

Jan 1, 1965

By Harry R. Hall

The proposition to make pig-iron from magnetic concentrates and cobbed ore with charcoal-fuel weighing from 12 to 20 lb. per bushel is, on the face of it, not inviting; but the work that has been done under these adverse conditions may be worthy of record as furnishing a comparison with coke blast-furnace practice under similar conditions.&apos; The record of the furnace-practice is briefly summarized as follows :—the charcoal was made of hemlock, pine, spruce, balsam, maple, beech and birch woods. All the concentrates, containing from 62 to 65 per cent. of iron, would pass 1/8in. screen, of which about 34 per cent. would pass an 80-mesh screen. The cobbed ore, varying in composition from 32 to 47 per cent. of metallic iron and from 23 to 32 per cent. of silica, would pass through a 1-in. ring, and of this ore about 25 per cent. would pass a 1/8-in. screen. The furnace was blown-in October 20, 1903, with charcoal which was supposed to weigh, when dry, 18 lb. to the bushel. Some charcoal similar to this was stored in the stock-house for about a year, and was then found to weigh 12 lb. per bushel.

Jan 1, 1906

By R. V. Whetsel, V. R. Garfias

It is estimated that during 1936 world consumption of petroleum, its products and related fuels reached an all-time peak of 1,757,778,000 bbl., an increase of 143,000,000 bbl. over 1935. As indicated in Table 1, world output of petroleum and related fuels exceeded demand by 92,510,000 bbl. during the year, as compared to an excess output of 88,000,000 bbl. during 1935. Production within the United States exceeded domestic demand by approximately 51,000,-000 bbl. in 1936, as compared to an excess output of 53,000,000 bbl. in the previous year. Outside the United States, production exceeded demand by 41,510,000 bbl. in 1936 and by 35,000,000 bbl. in 1935, indicating that foreign output is making no progress toward curtailment in line with demand. Table 1 also combines excess production in the United States with the favorable balance of exports over imports, showing that stocks in the United States were reduced by 23,400,000 bbl. in 1936, as compared to 22,447,000 bbl. in the previous year. In the rest of the world, excess output over demand combined with the excess of imports over exports from the United States indicates an unfavorable balance of approximately

Jan 1, 1937

(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, indicated by a *. A complete list was published in the Year Book, issued as a supplement of the Bulletin for March, 1918.) ALLEN, Roy H., Capt., Air Service, Bureau of Aircraft Production, Washington, D. C. *AMIDON, CLAUDE E., Chemical Warfare Section&apos;, Astoria, L. I. *ARCHBALD, HUGH, 311th Infantry, 78th Division, American E. F. ASKIN, THOMAS B. H., Radio Station, Cavite, P. I. *AYER, FRANK A., Lieut., Commanding Officer, -270th Aero Squadron, American E. F. BARBA, W. P., Lieut. Colonel, Ordnance, U. S. A., Washington, D. C. *BARRETT, LESLIE P., 1st Lieut., 5th Field Artillery, American E. F. BLICKENSDERFER, F. C., 1st Lieut., N. A., Ordnance Dept., Edgewood Arsenal, Edgewood, Md. *BROOKS, FLOYD R.., Lieut., Canadian Engineers, Seaford, England. BROWN, GEORGE M., Co. G, 1st Replacement Regiment, Washington Barracks, D. C. *BRUNEL, LOUIS J., Lieut., 5th Field Artillery, Brigade Headquarters, American E. F. *BRUNTON, D. W., Cons. Engr., Naval Consulting Board, Washington, D. C.

Jan 11, 1918

In a letter to Prof. L. P. Breckenridge, chairman, urging the continua-tion of the Fuel Conservation Committee of the Engineering Council, Edwin Ludlow calls attention to the millions of tons of culm in the an-thracite region that can be utilized in properly equipped plants and the electricity generated transmitted to. New York, Philadelphia, and other nearby cities. He says: "Experiments have shown the economic value of No. 4 Buckwheat, which passes through a 1/16-in. mesh and over a 1/32-in. mesh. Culm passing through a 1/32-in- mesh is pulverized until it passes through a 300 mesh and is then burned under boilers. Two plants of this kind are now in successful operation in the anthracite region. Another rather interesting experiment was using 30 per cent. of this pulverized coal with 70 per cent. of oil in the generation of steam in the place of pure oil. This utilization of the fine anthracite, which has always been a waste product and thrown away, is not only fuel conservation in its truest sense but in the case of a bituminous strike, when no strike is contemplated in the anthracite field, will furnish fuel to manufacturing plants that otherwise would be obliged to shut down."

Jan 12, 1919