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By Louis D. Huntoon

After the completion, in 1905, of the Hammond Mining and Metallurgical Laboratory of the Sheffield Scientific School, Yale University, it became necessary to secure and assay a large assortment of ore-samples, and to arrange the results in such a manner as to admit of use with the least amount of work on the part of the instructor. At first, pulp-samples, together with their respective assays, were secured from the mines, smelters, and public aesayers, but this practice did not prove satisfactory. It was not only difficult to secure a suffciently large supply of material, but it was also impossible to secure the proper assortment required for instructing students. It involved considerable work for the donors to select the samples and furnish a record of the assays, and also for the instructors to examine each pulp and determine its character. Some of the samples did not contain a sufficient quantity of pulp for the student to check his work in case his first assay did not correspond to the record. In order to increase the number of samples and to vary the character of the material, pulp-samples were later mixed with barren rock and with sulphides, but this method also was unsatisfactory and has been discontinued. At present, small lots of ore are secured by gift or purchase from engineers, public assayers, and the managers of mines, mills, and smelters. These lots weigh from 10 to 100 lb. each, and are rarely finer than 0.25 in. in size. The larger samples permit of a careful study by the instructor in order to determine the best method of assaying to be followed. In universities in which assaying is taught the record of the samples should fulfill the following requirements:

Jan 1, 1910

By R. A. Morse, P. M. Bridges, L. E. Wilsey, Howard N. Hall, P. L. Terwilliger

Theoretical and experimental investigations of a constant pressure gravity drainage system are reported. Experimental data are presented to show that recovery to gas breakthrough by gravity drainage is inversely proportional to rate. The gravity drainage reference rate, which is numerically equal to the so-called "maximum theoretical rate of gravity drainage" is shown to have no particular significance from a recovery standpoint. Before this rate can be used as a basis of comparison for recoveries, it is necessary that the relative permeability and capillary pressure characteristics and displacing fluid viscosities be identical for the systems compared. A method is presented by which accurate prediction of the performance of a gravity drainage system can be made. Close agreement between experimental and calculated drainage performance shows that steady state relative permeability and static capillary pressure data can be used to describe fluid displacement behavior. The very wide range of liquid recoveries before gas breakthrough which result from production rate variation alone demonstrates the importance of this factor in planning depletion of a gravity drainage reservoir. Calculated results are presented which show that little additional recovery can be expected from a high pressure gravity drainage system between the times of gas breakthrough and attainment of such high gas/liquid ratios as to make further pressure maintenance impractical. INTRODUCTION It bas been recognized for some time that gravity forces play an important part in the recovery of oil from some types of reservoirs. Field experience has shown that under certain conditions, gravity drainage can result in very high oil recov- eries. Qualitative reasoning has led most engineers to the general conclusions that: (1) Where gravity drainage is important, the reservoir pressure should be maintained by gas injection at the crest of the structure to prevent shrinkage of the oil in place and to keep a low viscosity so the oil can drain at the fastest possible rate. (2) Recovery by gravity drainage is rate sensitive. A survey of the literature indicates that while considerable work has been done on the effects of gravity in oil production problems, no satisfactory method of calculating the performance of gravity drainage reservoirs has been reported.1,2,3,4,5 In the absence of any proven method of calculating reservoir performance, the level at which pressure should be maintained and the rate of production for most efficient operation has been open to debate. The limits and general character of the recovery us rate relationships are apparent from the following: As rates of production approach zero, oil will drain by virtue of its own weight fast enough that the saturation distribution is always approximately the same as the original static capillary distribution. That is, at the gas-oil contact the liquid saturation goes from 100 per cent to a value of perhaps 10 to 30 per cent in a very short distance, so that as production continues. a relatively sharp gas-oil contact moves down-structure, leaving a very low residual oil saturation in the originally oil saturated zone. At very high rates of production, the oil that drains by virtue of its own weight is negligible and recovery will approach that of a horizontal gas drive, leaving a much higher residual oil saturation. At intermediate rates of production. recovery should be between these extreme*. Despite the lack of quantitative methods, tentative agreement had been reached on the question of optimum producing rate. It has been generally concluded that rather sharp decrease in recovery would occur at rates of production above the "maximum rate of gravity drainage" and hence this rate should not be exceeded. The "maximum rate of gravity drainage" is defined as the rate of production from a 100 per cent

Jan 1, 1951

By L. E. Wilsey, R. A. Morse, P. L. Terwilliger, Howard N. Hall, P. M. Bridges

Theoretical and experimental investigations of a constant pressure gravity drainage system are reported. Experimental data are presented to show that recovery to gas breakthrough by gravity drainage is inversely proportional to rate. The gravity drainage reference rate, which is numerically equal to the so-called "maximum theoretical rate of gravity drainage" is shown to have no particular significance from a recovery standpoint. Before this rate can be used as a basis of comparison for recoveries, it is necessary that the relative permeability and capillary pressure characteristics and displacing fluid viscosities be identical for the systems compared. A method is presented by which accurate prediction of the performance of a gravity drainage system can be made. Close agreement between experimental and calculated drainage performance shows that steady state relative permeability and static capillary pressure data can be used to describe fluid displacement behavior. The very wide range of liquid recoveries before gas breakthrough which result from production rate variation alone demonstrates the importance of this factor in planning depletion of a gravity drainage reservoir. Calculated results are presented which show that little additional recovery can be expected from a high pressure gravity drainage system between the times of gas breakthrough and attainment of such high gas/liquid ratios as to make further pressure maintenance impractical. INTRODUCTION It bas been recognized for some time that gravity forces play an important part in the recovery of oil from some types of reservoirs. Field experience has shown that under certain conditions, gravity drainage can result in very high oil recov- eries. Qualitative reasoning has led most engineers to the general conclusions that: (1) Where gravity drainage is important, the reservoir pressure should be maintained by gas injection at the crest of the structure to prevent shrinkage of the oil in place and to keep a low viscosity so the oil can drain at the fastest possible rate. (2) Recovery by gravity drainage is rate sensitive. A survey of the literature indicates that while considerable work has been done on the effects of gravity in oil production problems, no satisfactory method of calculating the performance of gravity drainage reservoirs has been reported.1,2,3,4,5 In the absence of any proven method of calculating reservoir performance, the level at which pressure should be maintained and the rate of production for most efficient operation has been open to debate. The limits and general character of the recovery us rate relationships are apparent from the following: As rates of production approach zero, oil will drain by virtue of its own weight fast enough that the saturation distribution is always approximately the same as the original static capillary distribution. That is, at the gas-oil contact the liquid saturation goes from 100 per cent to a value of perhaps 10 to 30 per cent in a very short distance, so that as production continues. a relatively sharp gas-oil contact moves down-structure, leaving a very low residual oil saturation in the originally oil saturated zone. At very high rates of production, the oil that drains by virtue of its own weight is negligible and recovery will approach that of a horizontal gas drive, leaving a much higher residual oil saturation. At intermediate rates of production. recovery should be between these extreme*. Despite the lack of quantitative methods, tentative agreement had been reached on the question of optimum producing rate. It has been generally concluded that rather sharp decrease in recovery would occur at rates of production above the "maximum rate of gravity drainage" and hence this rate should not be exceeded. The "maximum rate of gravity drainage" is defined as the rate of production from a 100 per cent

Jan 1, 1951

By Arne Bergholm

Australian rutile was chlorinated in the presence of CO or carbon. The chlorination velocity in CO was found to be strongly influenced by temperature and proportional to the CO concentration, but independent of the Cl, concentration. In the presence of carbon, the reaction velocity is much higher. The reactivity of the carbon and the distance between the carbon and the rutile surfaces are important variables. The reaction velocity is approximately proportional to the Cl, concentration and independent of the CO concentration of the surrounding atmosphere. Experiments with fluidized-bed chlorination of carbon-rutile mixtures indicate that the motion of the bed has little influence on the reaction velocity. At low temperatures, the chlorination velocity of dense tablets is much greater than that of TiO, coke mixtures suitable for fluidization. The reaction mechanism is discussed. In the industrial production of TiCl, rutile is chlorinated in the presence of carbon. Disregarding intermediate steps, the reaction may be expressed by the following equations: The ultimate object of this study was to find out whether the reaction velocity was higher in fluidized bed operation compared with chlorination of pelle-tized carbon-rutile mixtures. From a literature survey and preliminary experiments it was learned that some basic knowledge about the reaction mechanism was needed for a good experimental design. Therefore the following sets of experiments were carried out: 1) Studies on the reaction velocity in the chlorination of rutile with CO as the only reducing agent. 2) Chlorination of separate rutile-carbon tablets. 3) Chlorination of rutile-carbon tablets at different temperatures with various kinds of carbon, various grain sizes, and various tablet-making techniques. This series of experiments was carried out not only with pure chlorine but also with mixtures of chlorine with argon, CO and CO,. 4) Chlorination of rutile-carbon tablets made in a strictly standardized manner. 5) Chlorination of static-bed rutile-carbon mixtures. 6) Fluidized-bed chlorination of the same mixture. The experiments 4 to 6 formed the final and direct test of the main question: Static bed vs fluidized bedo. Although there are many patents and papers dealing with the general aspects of chlorination, only few experiments from which detailed information can be obtained have been reported in the literature. Pamfilov and coworkersl3 have studied chlorination of TiO, with CO or carbon as the reducing agent. They found that the weight decrease per hour at 600o to 800°C amounted to 11 to 14 pct with CO and 45 to 51 pct with carbon. They suggest that phosgene might be an intermediate in the chlorination of TiO,. Takimoto and Hattori have chlorinated reduced titanium oxide (TiO) and found a very high rate of TiC1, -production. They proved that the composition of the gas from the chlorination of rutile-carbon mixtures contained mainly CO. They reported for instance 74.7 pct CO, and 5.6 pct CO at 800°C at which temperature the Boudouard equilibrium composition is 12 pct CO and 88 pct CO. Seligman and Segerchano6 have studied the chlorination of TiO. They proved that Ti0 and chlorine react rather rapidly at temperatures as low as 300°C. Above 400°C, the reaction was complete. At 500°C the velocity of this reaction was much higher than was chlorination of TiO, + C. McTaggart,7 Nishimura, et al. and Wilskam have chlorinated mixtures of carbon and various kinds of rutile or beneficated ilmenite. Nishimura, et al. also report the results from reduction of TiO, with H2, CO, or carbon. At 900°C only 1 pct Ti O is formed in 2 hrs. At 1100°C 23.1 pct Ti, 0, was formed if elementary carbon was present. No carbide formation occurred below 1400°C. McIntosh and Cofferll have observed that the CO, content of the exit gases from chlorination of rutile and calcined petroleum coke is appreciably higher than found in the Boudouard equilibrium. At 900°C the ratio (CO, /CO + CO) is about 80 pct, whereas the equilibrium value is about 2 pct. W. E. Dunn12 has studied the chlorination rates of several TiO,-bearing minerals with CO + Cl, or COCl . Chlorinations were carried out either in a fluidized reactor or a fixed-bed reactor, both having a 30-mm diameter. The results obtained in both reactors were comparable. It was proved that benefi-ciated ilmenite (i.e., ilmenite from which the iron oxide had been removed by chlorination) was chlorinated 10 times faster thanrutile. Sore1 slag showed an intermediate rate. When phosgene is used, the

Jan 1, 1962

By M. J. Bradley, P. H. Dike

THE art of melting nonferrous metals, in smelting, casting, and in compounding of alloys, is highly dependent on a knowledge of the temperature of the metal. This knowledge may reside in the experience and judgment of the melter, or it may be based on measurement by suitable instruments. It-is our purpose to discuss briefly the instruments available for such measurements, and more at length the present, practice in molten metals, without recommendation. In the limits of time and space allotted, we can do little more than mention the available methods of measurement that are directly applicable in this field. In the 1939 edition of the Metals Handbook, issued by the American Society for Metals, pages 285-310, there is a more extensive discussion of pyrometry than can be attempted here, which should be referred to for details we must omit. For reasons that will appear, thermoelectric pyrometry will constitute the major portion of our discussion of the subject, only limited attention being given to radiation or optical pyrometry. METHODS OF MEASUREMENT THERMOELECTRIC PYROMETRY A thermocouple consists of two dissimilar wires, or a wire inside a tube of a dissimilar material, joined together at one end, which is placed in the region or medium where the temperature is to be measured, while the other ends are connected through suitable lead wires to a voltage-measuring device, which may be a millivoltmeter or a potentiometer. An electromotive force is generated in the circuit, which is proportional to the difference in temperature between the hot junction and the junctions of the lead wires to the measuring instrument. If this latter temperature is maintained constant, the e.m.f. measured varies only with the temperature of the hot junction. Every thermoelectric pyrometer is calibrated for a particular cold-junction temperature such as 0°F., 32°F. or 75°F., but usually it is more practical to compensate instrumentally over the probable range of cold-junction temperatures than to maintain the junction at the particular temperature for which the instrument was calibrated. Thermocouple Materials Only a very few pairs of metals are in frequent use for pyrometric measurements. The most generally used are iron-constantan for temperatures up to 1500°F. in oxidizing and up to 1800°F. in reducing atmospheres and chromel-alumel, which may be used in oxidizing atmospheres at temperatures up to 2400°F. for intermittent service. The accuracy of the latter deteriorates rapidly at temperatures above 2200°F. Platinum vs. platinum-rhodium thermocouples may be used up to a temperature of 3000°F , but they are easily contaminated by slow absorption of

Jan 1, 1946

By E. S. Machlin

The wzodel of Blandin and Diplunt;, generalized to include a phase factor, is applied to the carbon-carbon interaction in iron. Darken&apos;s "energetic" model is generalized to include not only first neighbor interactions but further neighbor interactions as well. On the bases of these generalized models relations are derived for the activity of carbon in both austenite and ferrite in terms of the carbon-carbon Pair interaction energies. A single function then yields the pair interaction energies consistent with the experimental activities of carbon in both ferrite and austenite. Thus, a simple explanation is given for the observation that the nearest-neighbor interaction between carbon is repulsive in austenite and attractive in ferrite. Certain consequences of this approach are explored. OnE object of the present paper is to attempt to take into account the consequences of electrostatic contributions to the carbon-carbon pair interaction energy for carbon as a solute in iron. Friedel&apos; has shown that oscillations in electrostatic potential are to be expected about a solute atom in a metallic solution. Blandin and 6lant6&apos; have shown that such oscillations yield an interaction energy between pairs of solute atoms that obeys the relation: W{ = A cos(2ftFri + 4>)/(kFri)3 [l] where kF = Fermi wave vector, ri = distance between solute atoms comprising the pair7 <p = phase factor dependent only on electronic nature of solute and solvent, A = coefficient dependent only on electronic nature of solute and solvent. Machlin3 found that Eq. [I] accurately described the pair interaction energy derived from short-range order measurements based on field ion microscope observations of dilute alloys of platinum. He also found that the value of the phase factor $ derived from residual resistivity measurements agreed well with that obtained from the analysis of the short-range order data. Harrison and paskin4 were able to predict the long-range ordering energy of 0 brass using Relation [I] and residual resistivity values to predict the value of the phase factor $. Machlin5 has repeated their analysis and applied it to the prediction of the long-range ordering energy in AgZn and AgCd with excellent agreement between prediction and experiment. Both A and $ are independent of the crystal structure. The Fermi wave vector depends uniquely upon the conduction electron concentration per unit volume in the spherical approximation of the Fermi surface. Thus, Eq. [I] is expected to apply to both fer- rite and austenite with only one set of values of A and $. Mossbauer studies6 yield the result that iron has one 4s electron. We shall make an assumption found to hold previously for platinum3&apos;7 and nikel, which is that only the s electrons are involved in shielding the perturbing potential of carbon. With this assumption, kF = 1.35 A-l. Although A and $ may be obtained from certain mdels&apos;&apos;&apos; we shall take A and $ to be empirical constants in the spirit of Kohn and osko.&apos; Thus, Eq. [I] involves two adjustible parameters. Consequently, two independent relations in A and $I are required in order to evaluate them for carbon as a solute in iron. We may use a recent analysis of Aaronson, Domain, and poundg who showed that Darken&apos;s energetic model,1° as well as others, can be used to describe the activity-temperature data for carbon in iron in both the aus-tenitic and ferritic phases. Darken&apos;s model takes into account only first neighbor pair interactions. For our needs, all neighbor pairs need to be taken into account. It is convenient to generalize Darken&apos;s model. The result for the partition function for austenite is: over the temperature range 800" to 1200°C and where the uncertainty corresponds to one standard deviation. Eq. [4] effectively yields only one relation. Another relation is required to obtain unique values for A and $. One property of Eq. [4] is that it is independent of crystal structure. Hence, data for a iron can be used to obtain another relation. To arrive at this relation we must generalize Eqs. [2] and [3] so that they may be applied to the bcc a iron. The result is that:

Jan 1, 1969

By CHARLES SIMENSTAD

COAL preparation, or coal washing, is not a new subject to the Pacific Northwest. Most of the coals mined in this state smaller than lump, and nearly all such sizes mined on the Pacific slope of the Cascade mountains, require washing before marketing. A large proportion is being put through some washing process, but coal washing in this state, has not reached the efficiency it is practicable to obtain. In fields where the coal, as mined, contains less impurities, where working conditions make it practicable to eliminate some of these impurities under- ground, or where a large proportion of the run-of-mine coal is lump, coal preparation does not present the problem it does in our local field. Here the coal is associated with considerable impurities and the slack, or steam sizes, aggregate about 80 per cent. of the output. Therefore, approximately 80 per cent. of the coal mined must be washed to become a valuable fuel. Washing then becomes a major operation, as hand picking of the 20 per cent. of lump is a standard and comparatively simple operation. Some coal beds here are so interstratified with bone, of a more or less massive nature, that hand picking of lump is not economically efficient. In such cases the coal must be crushed fine enough for it to be freed from the bone, so that the two may be separated. Coal cleaning in this field, therefore, must receive skilled attention to purify this fuel and raise its stand- ard. Without greater efficiency in cleaning, the importation of coal from outside the state will be even greater than in the past.

Jan 1, 1926

By Rudolf Gahl

Page Tests Conducted in Small Test Mill................... 577 Demonstration Tests Conducted by Minerals Separation Co....... 577 Flotation Tcsts Conducted by Inspiration Co............. 577 Sampling of Orebody........................ 578 Other Results Obtained in Small Test Mill.............. 580 Tests Conducted in Large Test Mill................... 580 Flow Sheet............................. 580 Discussion of the Value of Preliminary Table Treatment......582 Callow Flotation Installation..................... 585 Flinn-Towne ~nstallation...................... 587 Cole-Bergman Installation...................... 588 Methods Followed in Competitive Tests................ 590 Advantage of Pneumatic Flotation Machines.............. 591 Pneumatic Machines of Minerals Separation Co............. 591 Float Skimming Device......................; 594 Experimental Inspiration Flotation Machine.............. 596 Arrangement of Callow Cells in Series ................ 598 Variations of Flow Sheet....................... 599 Coal Tar as Flotation Agent..................... 600 Experience with Primary Slime.................... 602 Influence of Iron on Flotation..................... 605 Influence of Chemicals on Primary Slime............... 607 Fineness of Crushing Desired.................... 607 Operation of Large Concentrator..................... 608 Flow Sheet of Callow Sections.................... 608 Flow Sheet of Inspiration Sections.................• 609 Inspiration Flotation Machine of Steel Construction.......... 610 Minerals Separation Section..................... 613 Filter Plant............................ 613 Results Obtained in Commercial Concentrator............... 616 Tonnage............................... 616 Air and Power Consumption...................... 616 Floor Space............................ 617 Water Consumption......................... 617 Recovery............................. 618 Oil Consumption.......................... 618 Operating Costs for Flotation..................... 620 Discussion of Results Obtained.................... 620 Prospects for Futurc Development of the Flotation Process......... 621 Porous-Bottom Experiments..................... 622 Raising the Grade of Concentrates.................. 624 Recovering Carbonates by Flotation................. 624 Recovery of Silicates by Flotation.................. 627 General Theory......................... 628 Conclusion............................... 630

Jan 1, 1917

By H. F. Yancey, O. L. Furse

Although usually coal is screened with punched plate screens having round-hole perforations, there has been an increasing use of wire-cloth screens with square apertures, especially at pneumatic coal-cleaning plants and at other plants that use vibrating screens. It is therefore desirable to establish a relationship between round-hole and square-hole screens for use not only in experimental investigations of coal-preparation problems, but also for use in commercial operations. This need is apparent whenever it is necessary to substitute wire cloth for punched plate and still retain the familiar trade-size product. Although similar work probably has been done by industrial organizations concerned with the screen sizing of coal, the only recorded study of the subject other than the one described herein of which the authors are aware is that of Roesler,1 who screen-sized two samples of limestone and two of zinc concentrates crushed to pass 1/2 in. with a set of square and round-hole screens and found the ratio to be 1:1.23. Procedure In the course of a study of various methods for determining the friability of coal,2 it was necessary to screen-size the products of a series of friability tests of three different coals with both round-hole and square-hole screens. The coals used were from the Roslyn No. 3 and the Bayne mines in Washington and the Rolapp mine in Utah. Twenty 25-lb. samples of each of these three coals at 3 to 2 1/2 in. round-hole size were dropped from a height of 10 ft. to a concrete floor, and then screened with

Jan 1, 1934

By G. Das, S. V. Radcliffe

The substructural features developed in tungsten as a function of annealing temperature (up to 2200°C) and type of material [undoped and doped powder metallurgy (PM) tungsten and electron beam melted tungsten] have been investigated by transmission electron microscopy. For doped PM tungsten wires, characteristic "par-ticulate" substructural features developed rapidly with increase in annealing temperature above 700°C. The features consisted of parallel rows of elongated or circular shapes (500 to 1000A diam) lying along the direction of the wire axis and were identified as internal voids by diffraction contrast experiments. In recrystallized doped PM rod, larger voids were observed and were identified by precision dark field analysis to be cubic in shape and bounded by (100) planes. In marked contrast with both the doped PM materials, recrystallized undoped PM rod exhibited only very occasional and randomly arranged voids. Furthermore, no voids were observed in either material after electron beam melting. The high concentration of voids in the doped PM materials is attributed primarily to vaporization of doping additions or their pvoducts situated at the original grain boundaries , whereas the few voids in undoped material are considered to be traces of microporosity which were not eliminated during sintering. A tentative mechanism is suggested for the dezlelopment of the voids in relation to the processing sequences (sintering and working) and to the subsequent annealing. In recent years, a characteristic substructural feature consisting of rows of small elongated or circular regions of light contrast lying along the direction of working has been seen in thin foil electron microscopy studies of annealed sheet, wire, and rod tungsten. These features were present in the published micrographs of sheet by Weissmann et al.1 and of wire by Meieran and Thomas2, although the authors did not draw attention to them. Wronski and Fourdeux3 observed similar features in sintered rod tungsten (it was not specified whether or not the material was doped*) and interpreted them on the basis of their ap- particles, based on extraction replica evidence from the fracture surface of the initial hot-rolled slab material from which the sheet was prepared. No diffraction contrast experiments on the features were reported in any of these studies. The present investigation was undertaken with the primary objectives of: a) identifying the nature of these substructural features in tungsten by electron diffraction contrast experiments, since the contrast for voids can be expected to differ from that for crystalline or glassy particles, and b) elucidating the origin of the features and their development. For the latter purpose, doped and undoped powder metallurgy tungsten was obtained as rod and wire to represent different stages of reduction during final processing. These materials were examined both in the as-processed condition and after annealing to successively higher temperatures. In addition, the same doped and undoped materials were examined after vacuum melting in rod form. I) MATERIALS AND PROCEDURE Doped powder metallurgy (PM) tungsten wire (commercial purity 99.9 pct W) was obtained in the as-drawn and surface ground condition (0.030 in. diam "ground seal rod"). Doped and undoped tungsten rod (0.075 in. diam) representing an earlier stage of final processing was obtained from the same commercial source (Refractory Metals Division, General Electric Co.). Lengths of both the doped and undoped rod materials were single-pass melted in an electron-beam zone refiner to examine the effect of vacuum melting on the substructure. Annealing was carried out in a tungsten crucible in a tantalum strip resistance furnace under a vacuum of l0-15 mm Hg. Longitudinal sections of the wire and rod materials were examined by light and electron microscopy. The preparation of thin foils suitable for electron transmission from 0.030 in. diam tungsten wire and the rod specimens was carried out by means of a high-precision microjet technique developed to provide lack of jet stability and precise control of the area thinned. The method is described in detail elsewhere.&apos; The foils were examined in a JEM 6A electron microscope using a goniometer stage (±20 deg tilt, 360 deg rotation) and operated at 100 kV. To minimize contamination problems a 200 µ condenser aperture was used in conjunction with a useful beam current of 50 µA. II) RESULTS AND DISCUSSION A) Diffraction Contrast Analysis. In order to determine the optimum conditions for the development of the substructural feature, a series of isochronal 30 min annealing experiments were carried out on specimens of the doped PM tungsten wire. The transmission electron microscopy analysis showed that the as-drawn wire, Fig. 1(a), consists of &apos;fibers&apos; whose long axis is closely parallel to the wire axis of (110). The fiber width averages some 0.5 µ. Dense disloca-

Jan 1, 1969

By Fred C. Bond

THEM is acute need for a method that will measure the size distribution of finely divided materials, particularly when the particle sizes are smaller than the openings of the finest screen cloth regularly available for testing. Several methods and types of apparatus are in common use, but none are entirely satisfactory for all purposes. The commonest methods are microscopic grain counts, elutriation by liquids or air, and sedimentation. Valuable results have been secured by air elutriating devices, typified by the Haultain Infra-Sizer and the Pearson Air Analyzer. These are especially useful in obtaining sized fractions that can be examined and analyzed separately. They are particularly advantageous because the sizes of the samples produced are adequate for a considerable amount of subsequent manipulation and testing. The sedimentation methods, such as the Wagner and Klein turbidimeters, and the sedimentation balance, do not deliver sized fractions, but they do provide a means for reasonably rapid comparative determinations of fineness and surface area, which are important in many industries. The sedimentation balance consists essentially of a balance pan suspended in a liquid that contains falling particles. These are weighed as they settle onto the pan, and the time rate of settling is determined. The size distribution is computed from the settling rate. The balance appears to be inherently more accurate than any of the other methods, and it has served a useful purpose in the study of numerous problems, but there have been limitations which have restricted its scope of application. Because of the essential simplicity of the method, it was felt that intensive investigation might lead to modifications that would make it a valuable tool for the operators in industries that require precise information regarding the results secured from grinding equipment. A large number of size-distribution determinations below 200 mesh were required in connection with a particular research problem in fine grinding. The sedimentation balance was used in this investigation, and a technique was gradually developed that ultimately gave good results. A serious difficulty appeared early in the work. It was discovered that less material always settled on the sedimentation pan and more on the bottom of the containing jar than could be accounted for by calculation, and also that the material that settled to the bottom of the jar outside the edges of the pan was considerably finer than that on the pan. This segregation was so great that the distribution obtained for the finer sizes was obviously very much in error; material ground to about the degree of comminution of cement might have only 60 per cent of the calculated weight actually settling on the pan, and the error increased with the fine-

Jan 1, 1939

By Frank H. Macpherson

FEW people realize the tremendously important part that Saline and Pulaski Counties in central Arkansas have played in the winning of the war. The present favorable war situation might have been very different if we had not had the bauxite produced during the past two years from these two small areas, a total of about 15 square miles. As a general statement, it can be said that the Arkansas bauxite deposits are, individually, relatively small, irregular in thickness, somewhat scattered and, therefore, compared with coal or iron-ore or open-pit copper deposits, relatively ill adapted to a rapid and drastic expansion in output. A few years ago it would have seemed an impossibility for these two counties to produce over six million long dry tons of bauxite in one year. However, with the cooperation of the operators, that is what happened in 1943, when the state of Arkansas achieved the seemingly impossible with a production of 6,020,684 1 long tons (dried equivalent) of bauxite ore during the calendar year, compared with an average annual production of 323,725 2 tons in the period 1931 to 1941. It became apparent early in the year 1941 that the United States soon might be drawn into the war, and, therefore, that an immediate study of the domestic metal situation was necessary. Aluminum is by far the most important metal used in the manufacture of aircraft and many other implements of war and it was evident that there would have to be a tremendous increase in the domestic production of bauxite, which is practically the only present source of this metal. As more than 95 per cent of the known bauxite reserves of the United States was in Arkansas, careful preliminary studies were made in that state by representatives of the United States Bureau of Mines, the United States Geological Survey and the War Production Board in the summer of 1941. Reserves in other states also Were studied. The result of these studies was a recommendation by the War Production Board that mining in the bauxite-producing areas of Arkansas be geared for maximum production, in order to supply sufficient ore for this country&apos;s war requirements and, if necessary, to supply the war needs of Canada as well. The annual goal of eight million long tons (dry Weight) set for Arkansas by the War Production Board Was exceeded in 1943, when in August the output reached a rate equivalent to more than 8,300,000 long tons, dry basis. Because of the comparatively small tonnage of bauxite reserves, Congress appropriated funds for an extensive drilling program to be carried out by the Bureau of Mines in cooperation with the Geological Survey. Drilling began in May 1942 but owing to delay in delivery of drills, and trucks upon which the drills were mounted, the program did not get into peak operation until early in 1943. By the end of July 1944, the Bureau of Mines records indicated an addition to the Arkansas reserves of 10,000,000 long tons (wet basis) in place

Jan 1, 1945

THE Copperton mill in reality was a sort of proving ground. It was. designed to serve three purposes: (1) to verify the accuracy of the mine sampling by actually treating substantial tonnages of ore, (2) to demonstrate on a reasonably large scale the percentage of the copper in the ore that could be recovered, and (3) to permit the testing of various kinds of machines and devices for crushing and concentrating the ore, so as to guide the engineers in designing the proposed 6000-ton milling plant at Garfield. As a matter of fact, it contained a lot of second-hand equipment sent by MacNeill and Penrose from the works of the United States Reduction &Refining Co. in Colorado; and it was by no means a model installation. Probably it would have been even less efficient, except for the skillful work of George O. Bradley and-Frank-G, Janney, both of whom had worked with Jackling at Mercur. Bradley was a mechanical engineer employed to design the plant. Later he directed the design of the concentrator at Garfield and remained as consulting mechanical engineer to the company until 1915. Janney, who afterwards became manager of mills, a position that he retained until his death in 1916, was a skilled mechanic at the Golden Gate mill at Mercur when Jackling built the plant in 1897. His flair was making machinery do what it was intended to do. He was not an ore-dressing engineer in a technical sense; but Jackling him- self was a specialist in this respect as well as in designing and construction and Janney&apos;s mechanical aptitude made him a particularly valuable man in the Utah organization. John McDonald, another of Jackling&apos;s Mercur associates, was the first mine superintendent. With practically all of the available funds earmarked for construction of the Copperton

Jan 1, 1933

By Anthony William Deller

MAJOR battles in the present war have been fought in American research laboratories. Without the outstanding contributions made by our scientists, engineers, and technologists in mining and metallurgy, many of our achievements in the present war would have been impossible. All recognize the fact that modern research has found the answer to problems which seemed insuperable. For the duration, many of these inventions must be maintained as war secrets. But aside from these, the record of accomplishment in peace and war, especially in the past few decades, known to all, demonstrates what scientific and industrial research can accomplish. It may be regarded as axiomatic that the development of modern industry is based principally upon research devoted to making inventions and discoveries. We have come a long way from the sixteenth century, when Agricola published his classical "De Re Metallica:" With our vastly increased knowledge we are now commercially producing dozens of metals entirely unknown at that time. Further, with our modern methods of research and development, a large number of invaluable alloy compositions have been produced. These include numerous age-hardenable alloys, cupronickels, "Everbrite", "Everdur", "Duralumin", "Elinvar", "Hastelloy", "Inconel", "Invar", "Monel", "Ni-chrome", "Ni-Hard", "Ni-Resist", "Permaloy", "Silcrome", stainless steels and irons and other alloy steels, "Stellite", and zinc die-casting alloys.

Jan 1, 1945

By S. C. Lindsay, D. J. I. Evans

With each year that passes hydrometallurgical processes are being more widely used to recover base metals from ores and concentrates. Generally these processes involve liquid-solid separation of metal-bearing liquors from barren residues. This may be done by countercurrent decantation, thickening and filtration, or filtration or centrifuging only. Choice of method is governed by chemical and physical characteristics of the pulps; cost of labor, power, materials, and waste disposal; and even by the amount of capital available when building plans are under way. Since all these controlling factors change with time and location and with technological advances, the separation method must be re-evaluated when new plants or expansion programs are contemplated. At the Sherritt Gordon refinery in Fort Saskatchewan, Alberta, an ammonia pressure leaching process recovers nickel, copper, and cobalt from the nickel concentrate produced at Lynn Lake, Manitoba.&apos; Pregnant liquors are separated from the barren leach residues in a thickener-filter liquid-solids circuit. The decision to use this type of separation was based on a summation of economic factors during the pilot plant and design stages. Although countercurrent decantation was not incorporated in the refinery flowsheet, the problems solved in thickening the leach residues and the experience gained in operating the commercial plant are pertinent to countercurrent decantation, particularly where it is applied to slow-settling leach slurries or slurries having a relatively high vapor pressure of a valuable, obnoxious, or poisonous vapor. The nickel sulfide concentrate is leached under pressure with air and ammonia to dissolve the nickel, copper, and cobalt and most of the sulfur, leaving a residue of iron and siliceous material to be discarded.2,3 During a subsequent distillation stage to recirculate part of the ammonia used for leaching, copper is recovered from the solution as a high-grade sulfide. After further purification to remove traces of copper and undesirable sulfur compounds, the nickel is precipitated by hydrogen under pressure at an elevated temperature. The cobalt, together with unreduced nickel, is precipitated as a mixed metal sulfide by hydrogen sulfide for eventual cobalt recovery. The end solution is then evaporated to produce a crystalline ammonium sulfate byproduct. As shown in Fig. 1, the nickel concentrate is leached in a continuous, two-stage, countercurrent circuit. In the first stage, or adjustment leach, the fresh concentrate is partially leached (60 pct) in the liquor produced in the second stage, or final leach. This system is necessary to produce a solution suitable for treatment in the copper sulfide precipitation step. Both stages are carried out at 170" to 180°F under a total pressure, with air, of 100 to 120 psig in large, horizontal, water-cooled autoclaves. The autoclaves are divided by overflow weirs into four compartments to give a certain amount of staging during leaching. In each compartment a relatively slow-speed (100 to 150 rpm) agitator maintains the nickel concentrate in suspension. Adjustment leaching, carried out in two autoclaves in parallel, is followed by a thickener-filter liquid-solids separation. Final leaching is done in two parallel trains of three autoclaves using aqueous ammonia to extract metal values from the adjustment leach residues. This too is followed by a thickener-filter circuit. Here the residues are repulped once with aqueous ammonia and twice with process water and refiltered after each repulping stage, to insure a maximum recovery of soluble metal, before being discarded to the tailings pond. The adjustment leach is followed by a seven-disk Dorr-Oliver American filter, with an agitator in the filter boot to keep the residues in suspension. Agitators have also been installed beneath the four ten-disk filters following the final leach thickener. Slurry Characteristics: A typical analysis of the Lynn Lake nickel concentrate, produced by differential flotation,&apos; is given in Table I. The concentrate comprises 35 to 49 pct pentlandite, (Fe, Ni)S; 3 to 6 pct chalcopyrite, CuFeS,; 1 to 4 pct pyrite, FeS2; and 24 to 42 pct pyrrhotite, Fe1-2S (where x varies between 0 and 0.2) together with various siliceous minerals such as actinolite, hornblende, and talc. During leaching, the sulfur in the pentlandite, chalcopyrite, and part of the pyrrhotite is successively oxidized through the thiosulfate and poly-thionate radicals to sulfate and sulfamate." Simultaneously the nickel, copper, and cobalt are converted to soluble ammine complexes and the iron in the pentlandite, chalcopyrite, and part of the pyrrhotite is oxidized to a spongy hydrated iron oxide. The hydrated iron oxide is readily broken down by the agitation needed to sustain the concentrate in suspension and to maintain a satisfactory mass transfer of the oxygen from the air to the slurry. Table I shows typical chemical, screen, and

Jan 1, 1961

By W. F. Campbell

One of the deepest coal operations today is the Springhill No. 2 mine of Cumberland Railway & Coal Co., subsidiary of Dominion Coal Co. Ltd. Mining is now conducted at a slope distance of 14,000 ft, with 4400 vertical ft of cover. The record of Springhill No. 2 can be said to contain the history of bumps in the Province of Nova Scotia. The Springhill coal field forms part of the Cumberland field of Carboniferous age. There are seven mineable fields in the area, numbered in the order they were discovered. Mining has been carried on in all but the No. 4 and 5 seams, but present operations are confined to the No. 2. Fig. 1 shows a vertical section through the seams. Opened in 1873, No. 2 mine was first worked from parallel slopes driven from the outcrop of No. 2 seam down to the 7700 level. As the mine went deeper, a two-place auxiliary slope was driven from the 6900 level to the present workings. A transfer level at the 7800 connects the main haulage with the auxiliary haulage slope. The No. 2 mine plan is shown in Figs. 2 and 3. The seam is bituminous, averaging 8.5 to 9 ft thick. There is a well defined parting 14 to 16 in. from the roof, and this roof coal is harder than the rest of the seam. Average pitch at the outcrop is 30°; at the 6500 level, 20°; at the 7900 level, 16°; and at the 13,800 level, 10°. Immediate roof and floor strata consist of beds of variable thicknesses of shales, grading to arenaceous shales to shaly sandstones to sandstones. A characteristic of the strata is the appearance and disappearance of sandstone bands, of considerable thickness, over distances of several hundred feet. MINE HISTORY Entrance to No. 2 mine is obtained by three parallel slopes, separated by 100-ft pillars except in the upper portion of the mine where the pillars are smaller. Main haulage levels were originally driven 600 ft apart for room and pillar extraction. A parallel drainage and intake airway was driven below each haulage level and a parallel return airway (counter level) above. Pillars 80 to 100 ft wide separated these two servicing entries from the center haulage level. Inclines up to 700 ft apart were driven off the levels up pitch to the upper levels, and at 40-ft centers level rooms 12 ft wide were driven off each side of the inclines, separated by crosscuts every 50 to 100 ft. The rooms were driven up to 350 ft or until they were holed into the room from the adjoining incline—a plan followed until the level reached its boundary. When the entire area between two levels had been divided into pillars, the pillars were extracted from the boundary back to the slope pillars. As depth of cover increased, mining conditions required larger pillars, and at the 3300 level, where NECESSARY PRECAUTIONS FOR MINING AT DEPTH In Room and Pillar Extraction: 1) Pillars must not be too small. 2) Pillar size must be uniform. 3) Extraction lines must be kept as straight and as uniform as possible. If they are irregular they should be brought back into line slowly. 4) Peninsulas of coal surrounded by gob must be avoided. 5) Pillars should not be disturbed after they have been formed. If it is necessary to increase the width of an opening, stripping or ribbing should be done gradually over a long distance. If increased width of openings is necessary at certain locations and can be anticipated, the openings should be widened during development. 6) Under no conditions should small pillars be left in the gob that would interfere with caving. In Long wall Operations: 1) Pillars formed during development for retreating longwalls should be uniform in size and shape. 2) Pillars should not be split after development, but if pillar splitting does become necessary it should not be carried out within the zone influenced by the working faces. 3) Width of levels should not be increased after the levels have been developed. If it does become necessary to widen a level it must be done gradually over a long distance by stripping or ribbing the coal ribs. 4) Extraction lines should remain uniform. If several walls are worked together—one behind the other—a uniform distance must be maintained between them and the wall faces must be kept as straight as possible. If the walls do get out of line, they must be brought back into line gradually to allow for a gradual change in the stress distribution over the area. 5) Stumps of coal, roof supports, etc. must not be allowed to remain in the gob to interfere with caving. 6) Midwalls should be built as rigidly as possible and properly maintained.

Jan 1, 1959

By I. R. Kramer

The stress associated with the surface layer was deter-minedfor iron and molybdcnum. These measurements show that the surface layer plays a very important role in the plastic deformation of bcc metals. From the relationship between strain arui the surface layer stress and the back stress imposed by the internal obstacles, it is Proposed that the term of of the Hall-Petch equation is related only THE large number of studies on plastic flow of bcc metals have failed to delineate clearly the nature of the obstacles influencing dislocation movement. Since there have been numerous articles which have reviewed the mechanisms proposed to explain various aspects of plastic flow of bcc metals (for example, conrad1 and Keh and weissmann2), no attempt will be made here to discuss their conclusions. However, it should be to the surface stress while the term Kfd is the stress required to move dislocations through internal obstacles. Further, it was possible to derive the well-known empirical equation C = , and it is shown that the terprz a is comprised of two terms: a term Cs associated with the surface layer stress and Ci related to internal dislocation obstacles. pointed out that in all of these publications it has been tacitly assumed that plastic flow occurred uniformly throughout the cross section of the specimen (except, of course, for the Luders band region), It has generally been further assumed that the nature and distribution of the obstacles opposing the motion of dislocations are of the same type throughout. The role of the surface layer in bcc metals has not, as yet, been considered even though its role in fcc metals has been shown clearly to be of primary importance.3-5 Investigations on fcc metals have shown that at the onset of plastic deformation a surface layer acts to

Jan 1, 1968

By Alden B. Greninger

Many recent contributions in the field of theoretical metallography have been concerned with crystallographic definitions or descriptions of various phenomena. The lattice orientation of the crystal being studied must be known before such definitions can be made. Practically all the metallic crystals used do not have crystal faces—other than the unknown "plane of polish"; hence resort must be had to other means of determining orientations. Some use has been made of etch-figures, slip bands, optical properties, etc., for this purpose; but for accurate results, and because of the fewer restrictions to their use, the various X-ray methods have been the most satisfactory.&apos; The methods that have been used may be divided into two general classes: (1) oscillation methods, and (2) transmission Laue methods.² The disadvantages of these methods are numerous, chief among them being the time required to determine a complete orientation. No method proposed so far is wholly satisfactory for determining the orientation of a small macrograin within a polycrystalline aggregate.

Jan 1, 1935

By Alden B. Greninger

Many recent contributions in the field of theoretical metallography have been concerned with crystallographic definitions or descriptions of various phenomena. The lattice orientation of the crystal being studied must be known before such definitions can be made. Practically all the metallic crystals used do not have crystal faces—other than the unknown "plane of polish"; hence resort must be had to other means of determining orientations. Some use has been made of etch-figures, slip bands, optical properties, etc., for this purpose; but for accurate results, and because of the fewer restrictions to their use, the various X-ray methods have been the most satisfactory.&apos; The methods that have been used may be divided into two general classes: (1) oscillation methods, and (2) transmission Laue methods.² The disadvantages of these methods are numerous, chief among them being the time required to determine a complete orientation. No method proposed so far is wholly satisfactory for determining the orientation of a small macrograin within a polycrystalline aggregate.

Jan 1, 1935

By Joseph Singewald

INTRODUCTION FOR a number of years Russia, India and Brazil have outranked .all other countries as producers of manganese ores. During the 5 years immediately preceding the European war, the average annual production of Russia was 844,000, of India 694,000 and of Brazil 200,000 long tons. Since the outbreak of the war there has been a considerable falling off in the Russian and Indian production, particularly in the former, whereas the production of Brazilian ores has been greatly increased, amounting in 1914 to over 250,000 tons and in 1915 to nearly 350,000, with conditions favorable for a still larger production in 1916. Manganese is one of the few industrially important metals that are not produced in the United States in quantity commensurate with our needs, so that we have been compelled to import annually about 300,000 tons, having a value of over $2,000,000, and these ores have been obtained from the three countries mentioned above. Their rank as contributors to our imports of manganese ores was India, Russia, Brazil; and in 1913, Brazil contributed only one-fifth. During the past 2 years Brazil has furnished rapidly increasing quantities to this country, and, with the falling off of imports from India and Russia, has become our principal foreign source. In 1914, Brazil furnished two-fifths of the imports and more than India or Russia, and in 1915 over nine-tenths. The manganese ores of Brazil are consequently of more than usual interest to us at this time.

Jan 10, 1916