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By R. P. Rothwell

IN January last I published in the Engineering and Mining Journal a table giving, with a considerable degree of accuracy, the production of anthracite coal for the year 1874. At that time it was impossible to give any correct estimate of the production of bituminous coal, for the returns I had been able to secure from such distant and scattered districts were not always forwarded punctually, and the work of -obtaining anything like a true estimate of the coal mined in districts from which I was unable to obtain official returns was both laborious and tedious. As will be seen from the detailed reports furnished herewith, I have succeeded in obtaining tolerably accurate returns of 20,093,704 tons; and upon careful revision by numerous well-informed local authorities, my first estimate of the amount of coal not officially returned to me has grown very considerably; indeed, as I get more information on the subject I find that our bituminous coal production greatly exceeds the amount we have been accustomed to hear it placed at, aggregating as it does, 26,465,704 tons. Mr. Andrew Roy, State Inspector of Mines for Ohio, has reported the production of Ohio at 4,168,445 tons, while his first estimates were but 3,820,000 tons; and similarly, in other cases, the more full the returns direct from the mines the more incorrect do I find all previous reports. I am able now to furnish, not an accurate statement, indeed, but by far the most full and reliable one that has yet been published of our coal production. The amount of labor involved in obtaining this admittedly imperfect report has been simply enormous. In the following tabular statement it will be perceived that Pennsylvania produces nearly one-half of all the bituminous coal mined * Since this paper was read fuller reports from some of the States whose production had been "estimated" have modified the figures given. The most serious error is in California, which produces probably 200,000 tons less than these figures. Oregon and Washington territories also yield smaller amounts than here reported; but these reductions are probably fully compensated by fuller reports from other States that were here underestimated. The correct aggregate amount will not vary much from that here given. Early in 1876 I will be able to give a fuller and more accurate report for the year 1875.

Jan 1, 1875

Sailing westward from the Society Islands in the fall of 1774, England's noted explorer Captain James Cook discovered New Caledonia-that long, linear island that has played such an important and varied role in the history of nickel. Upon his return to England the following year, Cook's announcement of this uncharted island generated governmental interest, but its distance from the Island Empire, the troubles in the colonies, and the difficulties with France combined to keep English interest indirect and in the background of the then-current events.

Jan 10, 1968

By Victor deWolfe

INTRODUCTION The Henderson Mine, located near Empire, Colorado, utilizes a continuous panel caving system to extract ore as one of the world's major producers of molybdenum. Any mine using a caving-by-gravity technique of mining must rely on closely controlled draw of the caved ore. This control is essential to insure proper caving action, to avoid damaging load concentrations of weight and to minimize the dilution of ore with waste material. Henderson is no exception. Draw control is a major factor in all production planning, from long- range plans to short-range and day-to-day ore scheduling. Draw control is reviewed constantly and administered daily in an effort to optimize production efficiency, ore recovery, and cave management. MINING METHOD The cave at Henderson is massive, moving slowly through large panels that are 244 m (800 ft.) wide by 610 m (2,000 ft.) long. Generally two cave areas are drawn at one time. The areas under active draw vary in size but can be as large as 244 m (800 ft.) by 244 m (800 ft. ) containing 400 draw points. Each draw point contains 45,360 mt (50,000 st) on the average and takes about two and one half years to exhaust. A complete panel is worked for seven to ten years. No pillar exists between panels, but rather a buffer zone of broken ore, or "static face," is left in each panel to be drawn with the adjacent, yet-to-be-caved panel in efforts of minimizing dilution of a working area from an exhausted one. (Figure 1) Production drifts are driven on 24.4 m (80 ft.) centers through the ore body. Between the production drifts are funnel-shaped draw bells on 12.2 m (40 ft.) x 24.4 m (80 ft.) centers to receive ore from the cave. Each bell is accessed by two draw points, one from the production drift on either side, thus forming a 12.2 m (40 ft.) x 12.2 m (40 ft.) draw pattern. Extraction of the ore is by rubber-tired, 3.8 m3 (5 yd3) load-haul-dump equipment. The LHDs then tram the ore a maximum of 49 m (160 ft.) to ore passes. Cave initiation and bell development are done from the undercut drifts which are parallel to and 17 m (55 ft.) directly above the production drifts. Longhole rings are drilled and blasted from the undercut drifts to define the bells and establish the undercut for caving. (Figure 2) DRAW CONTROL Since the cave line at Henderson is constantly advancing, it is necessary to be continually initiating new cave at one end while exhausting it at the opposite end. There must exist, therefore, an angle on the ore-waste contact in the broken rock from initiation to exhaustion. The basic concept of draw control is to keep this angle as smooth and even as possible, particularly at the time of exhaustion. If this is achieved, draw points are exhausted more or less in a line, avoiding pockets of remaining ore surrounded by exhausted areas. These pockets would cause spotty ore extraction at the time of exhaustion, increasing the amount of dilution occurring while introducing the potential for significant weight problems in the production area. To arrive at the desired angle on the ore- waste contact, maximum tonnage percentages are assigned to each row of draw points increasing at 10% or 15% increments (depending on cave size and velocity of draw) working away from the cave line. The available tonnage indicated by these percentages is the maximum allowable tonnage to be extracted from each draw point until the available tonnage percent- age is increased. As the cave moves, these percentages increase for each draw point regularly. However, in general the tonnage drawn from each draw point is kept at about 50% of this allowable maximum in order to maintain adequate available tonnage in the cave to sustain production for seven months if cave initiation were to cease. This available tonnage cushion is a safeguard built into the draw control program at Henderson to accommodate fluctuations in the rate of cave advance. When draw points move past the row of 100% tonnage availability, they are drawn past the desired 50% at the same increments per row until exhausted. (Figure 3) To achieve proper draw control, the number of LHD buckets to be taken from each draw point is assigned daily. The actual buckets taken, which may at times deviate from the

Jan 1, 1981

By R. C. Buehl, J. P. Morris

F. D. Richardson—The authors are to be congratulated on this further contribution to our knowledge of the thermodynamics of the interaction between sulphur and carbon and silicon in liquid iron. As the authors state, the influence of carbon and silicon on the activity coefficient of sulphur in liquid iron is clearly of great importance in the blast furnace, since it must cause a three to fourfold improvement in the partition of sulphur between slag and metal. The influence of increasing temperature in further increasing the activity coefficient of the sulphur in the metal in the blast furnace by increasing the carbon content is also of interest. This effect, however, is probably only part of the reason for the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing temperature. Other contributing factors are the lowering of the oxygen potential in the presence of carbon by increasing temperature and the probable increase in the activity coefficient of the lime in the slag for the same reason. The former of these effects, which works via the (CaO) + [S] = (CaS) + [O] equilibrium, might possibly account for a 70 pct improvement in the sulphur partition and the latter might give a further 50 pct improvement. C. Sherman—I would like to compliment the authors on their very careful research. If I may, I would like to show results of calculations on the carbon-sulphur-iron system similar to the ones that were shown in our paper for the silicon-sulphur-iron system. For Fe-S-C ternary system k=PHgs/PH2 x 1/(f1°) (f2°) (%S) where fs = sulphur activity coefficient fs' = fs for Fe-S system of equal pct S f3° = f2/f2 for Fe-S-C ternary system This same analysis has been used on other systems, but the results shown in fie.- 7 are for carbon and silicon. L. S. Darken—I would like to make two brief comments in addition to complimenting the authors on an apparently very precise and accurate investigation. The first is that the present work is in agreement with a calculation by Larsen and myself." Our calculation (much less precise than the present work) was based on: (1) Unpublished work on the sulphur content of molten iron (1.5 pct at 1500°C) in equilibrium with graphite and an iron sulphide slag; (2) the distribution coefficient of sulphur between slag and carbon-free liquid iron. We expressed the result in a form equivalent to log 7. = 0.18 [%C] which gives an activity coefficient (?s.) of sulphur only slightly higher than the authors find and certainly within the precision of the earlier work. My second comment concerns the correlation of the thermodynamic findings with atomistics. A rough pic- ture of the atomic arrangement in the liquid solution is rather easily conceived for this particular liquid solution containing iron, carbon, and sulphur. Carbon has a very much stronger affinity for iron than for sulphur. Hence we may conclude that a sulphur atom will but seldom be adjacent to a carbon atom—since this would be a position of high energy. From the metallic radii of iron and carbon we know that six iron atoms pack neatly around one carbon atom. Thus each carbon atom in retaining this shell of iron atoms (which latter may not be replaced by sulphur on account of the high energy requirement) decreases the available positions for each sulphur atom by six. Hence each atomic percent of carbon decreases the equilibrium sulphur content by 6 pct (of itself). Or, at low concentration each atomic percent of carbon increases the activity coefficient of sulphur by 6 pct. This is in good agreement with the observed increase (6 or 7 pct at low carbon content). It is indeed gratifying to find a case where, by such simple reasoning, quantitative agreement is found between precise data and the modern picture of the atomistics of the metallic state. J. P. Morris (authors' reply)—We would like to point out that there is an error in the equation on p. 322 of the paper. The third equation should read: ½S2 (gas) + H2 (gas) = H2S (gas) The authors wish to thank everyone for the interest they have shown in the paper. In regard to the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing the temperature, Dr. Richardson is correct in stating that the cause can be attributed only in part to the increase in activity coefficient of sulphur resulting from the rise in carbon plus silicon content of the metal with rise in temperature. However, this factor is probably an important one. The results of one experiment, performed since this report was written, indicate that at a constant temperature the addition of silicon to a melt saturated with carbon causes an increase in the activity coefficient of sulphur even though the carbon solubility is lowered. In this test, 2.5 pct silicon was added to a melt saturated with carbon and maintained at 1400°C. Although the carbon content dropped from 4.85 to 4.1 pct, the activity coefficient of sulphur was increased by about 20 pct.

Jan 1, 1951

By J. C. Scully, G. Sanderson

The electrolytic formation of y hydride in thin foils of Ti-5Al-Z.5Sn alloy and four Ti-A1 binary alloys has been studied by electron microscopy and electron diffraction. Two orientation relationships have been found: THE adsorption of hydrogen by titanium alloys from aqueous solutions has been widely reported. This can occur, for example, in both alkaline' and acidic2 pickling baths, in chemical polishing374 and electro-polishing Solutions,3,4 during cathodic polarization5" and under certain corrosion conditions,7,8 Etching reagents may also produce visible precipitates, sometimes after an aging period.9 During an investigation into the stress corrosion susceptibility of a titanium alloys by thin foil electron microscopy many observations have been made of hydride precipitation which has been deliberately promoted by techniques that are described below. The characteristics of hydride precipitation are reported and the observed crystallographic orientation relationships are discussed with reference to distortion in the titanium lattice created by the presence of the hydride. EXPERIMENTAL TECHNIQUE One grade of commercial titanium and five alloys were examined. Their composition and hardness values are shown in Table I. All the materials were in the form of thin sheet, 0.05 to 0.065 cm thick, which had been annealed after cold rolling. Foils suitable for transmission electron microscopy were made in two ways: a) by chemical polishing; and b) by elec-tropolishing. The solution used for chemical polishing was a mixture of 44 pct HF and 73 pct HNO3 in the volumetric ratio 2:3. During this polishing specimens were manually moved up and down in the solution. The reaction was veFy rapid, specimens thinning to foil thickness (2000A) in 5 min. Since chemical polishing solutions tend not to cause hydride contamination if they are kept cool3 no attempt was made to prevent the solution from becoming hot during the thinning operation. Electropolishing was performed at -30° C with a mixture3 composed of 295 cu cm methyl alcohol, 175 cu cm n-butyl alcohol, and 30 cu cm perchloric acid with an applied voltage of 40 v after chemically thinning specimens to approximately 0.017 cm thickness in a cool solution of 44 pct HF, 40 vol H2O2, and water in the volumetric ratio 1:6:3. EXPERIMENTAL RESULTS 1) Ti-5A1-2.5Sn Alloy. The Ti-5A1-2.5Sn alloy was severely contaminated by hydrogen during the chemical polishing. Foils obtained in this way contained large amounts of the y hydride phase. A typical micrograph is shown in Fig. l. The laths were identified as the hydride phase by selected-area diffraction of a large number of foils examined in a tilting stage that could be turned through 40 deg. Using the lattice spac-ings of the alloy obtained by X-ray diffraction as a standard the a value of the hydride was determined fromo the electron diffraction patterns to be 4.38 to 4.40A. The accuracy of this figure is only about ±2 pct. Further evidence of the identification was obtained by vacuum annealing chemically polished foils of the alloy at 500°C. This resulted in the dissociation of the hydride with the production of dislocations and possibly vacancy loops as shown in Fig. 2. The hydride

Jan 1, 1968

By C. L. Clark

EVERY user of steel should ask himself whether or not he is taking full advantage of the discoveries of the steel metallurgists during the last few years, or is merely buying grades that looked to be the best five or ten years ago for various specific purposes and have since proved perfectly satisfactory in his plant. If he has not kept abreast of the times, then it is probable that something better has been developed, or something just as good that costs less. Though this article will have to do with steel for high-temperature applications primarily, marked advances have been made in steels for other purposes as well.

Jan 1, 1940

Aamot, Olav Crone, (M'29) Chem. Engr., Elektrokemisk, Raadhusgt. 23, Oslo, Norway. Abadilla, Quirico A., (M'3S) Min. Engr., Dir., Bu¬ reau of Mines. Manila, P. I. Abbott, Agatin Townsend, (J'40) Jr. Engr., Phelps Dodge Corp., Morenci Branch, Morenci. Ariz. Abbott, Clarence E., (M'04) Civil Engr., V.P., Ten¬nessee Coal Iron & R.R.. Co., 1542 Brown -Marx Bldg.. Birmingham, Ala. Abel, Raymond L., (M'28) Chem. Engr., Command¬ing, Officer, Pittsburgh Chem. Warfare Procurement Dist., American Bank Bldg., Pittsburgh. Pa. Abel, Walter D., (M'32) Alin. Engr.. Div. of Cor¬porations, State of Calif., 403 California State Bldg., Los Angeles, Calif. Abele, Louis T., (M'31) 423 3rd Ave., Decatur, Ga. Abell, Leo M., (M'31) Mill Supt. & Met., Emperor Gold Min. Co., Led., Vatnkoula, Fiji Islands. Abernathy, G. E., (M'38) Geol., Kansas State Geol. Survey, Pittsburg, Bans. Abernathy, Jack H., (M'40) Peer. Engr., c/o Superior Oil Corp., National Bank of Tulsa Bldg., Tulsa, Okla. Aborn, Robert H., (M'10) Asst. Met., Research Lab., U. S. Steel Corp., Lincoln Highway, Kearny, N. J. Abraham, A. W., (M'39) Peer. Engr., Richfield Oil Corp., 555 S. Flower St., Los Angeles, Calif. Abramowitz, William L., (J'40) Chem., Industrial Inventions, Inc., 103 Foster Se., Peabody, Mass.

Jan 1, 1942

By J. B. Dyess, H. A. Miley

Tarnish films on some of the common metals (particularly on copper and silver) have been of much scientific and commercial concern for a long time, but before the development of the electrical method1 of measuring film thicknesses there was no satisfactory quantitative approach by which to study the separate parts of these complex films. This method has been adapted to the measurement of duplex (cuprous-cupric) oxide films on copper.2 In the present work it has been used to measure tarnish films on copper when they were mixtures of cuprous oxide and cuprous sulphide, and also when they contained cuprous oxide, cupric oxide and cuprous sulphide, each component part being measured at a different potential value. Oxide-coated copper specimens, prepared by heating copper in air at chosen temperatures for given periods, were exposed to moist hydrogen sulphide in air to forml the various oxide-sulphide films studied. The comparative resistances afforded by (I) different thicknesses of oxide films, and also (2) equal thicknesses of oxide and sulphide films, against severe attacks of hydrogen sulphide were measured. The electrical method requires that the time be observed for the cathodic reduction of each part of the film substance, and these respective time values multiplied by the constant current give the corresponding quantities of electricity expended. Then each millicoulomb per square centimeter represents a film thickness of M X 108____ 2DX 96,494 X 103 where M is weight of oxide (or sulphide) giving one gram-atom of oxygen (or sulphur) for cathodic reduction, D is density of the oxide (or sulphide) and t is the thickness in Angstrom units per millicoulomb per square centimeter. The value of t for cuprous sulphide (cubical) is 14.27, and is obtained by substituting 5.78 for D and 159.20 for M in equation a. Previous values hf t for cuprous and cupric oxides (12.36 and 6.39, respectively) were used in the present work.

Jan 1, 1939

By J. B. Dyess, H. A. Miley

Tarnish films on some of the common metals (particularly on copper and silver) have been of much scientific and commercial concern for a long time, but before the development of the electrical method1 of measuring film thicknesses there was no satisfactory quantitative approach by which to study the separate parts of these complex films. This method has been adapted to the measurement of duplex (cuprous-cupric) oxide films on copper.2 In the present work it has been used to measure tarnish films on copper when they were mixtures of cuprous oxide and cuprous sulphide, and also when they contained cuprous oxide, cupric oxide and cuprous sulphide, each component part being measured at a different potential value. Oxide-coated copper specimens, prepared by heating copper in air at chosen temperatures for given periods, were exposed to moist hydrogen sulphide in air to forml the various oxide-sulphide films studied. The comparative resistances afforded by (I) different thicknesses of oxide films, and also (2) equal thicknesses of oxide and sulphide films, against severe attacks of hydrogen sulphide were measured. The electrical method requires that the time be observed for the cathodic reduction of each part of the film substance, and these respective time values multiplied by the constant current give the corresponding quantities of electricity expended. Then each millicoulomb per square centimeter represents a film thickness of M X 108____ 2DX 96,494 X 103 where M is weight of oxide (or sulphide) giving one gram-atom of oxygen (or sulphur) for cathodic reduction, D is density of the oxide (or sulphide) and t is the thickness in Angstrom units per millicoulomb per square centimeter. The value of t for cuprous sulphide (cubical) is 14.27, and is obtained by substituting 5.78 for D and 159.20 for M in equation a. Previous values hf t for cuprous and cupric oxides (12.36 and 6.39, respectively) were used in the present work.

Jan 1, 1939

By Robert P. Koenig

FOR the first time other than on occasion of war the people of the United States are experiencing full-scale participation in world affairs. Public concern has seldom been so involved with conditions and events beyond our own continental limits. Our interest is not wholly of recent origin. Over the years the activity of the United States as a world power and an important trading nation brought about the widening of our political interests and the progressive interrelation of our economy with that of the rest of the world. Some may see in our present phase of world-wide activity the natural consequence of our continued growth. Such observers must also be aware that within recent years our importance in world affairs has been considerably increased. For within the space of an extremely short period events have occurred that have literally jet-propelled the United States into the role of a leading participant in the conduct of world affairs. Little of the choice of this role was left to us. The decline of power in western Europe, the ascendancy between the Elbe and the Pacific shore of an obviously antagonistic force, our earnest endeavor to secure international control of atomic energy, the necessity of providing relief and rehabilitation to war-ravaged areas, and the police work that must be done in occupied territories are problems which have risen in great measure without our help-but they cannot be solved without it.

Jan 1, 1948

By J. L. Huitt

Some of the problems associated with the acid treatment of oil wells are related to the control of the composition of formation treating acid. One test usually made at the well site is the determination of the acid concentration. Heretofore, the concentration of the acid delivered to the well site usually has been interpreted from a hydrometer measurement of the specific gravity of the acid solution. Experience has shown that this indirect measurement of acid concentration can give erroneous results. More reliable results are, of course, obtainable by the titration method; however, this method is not easily adapted to well-site testing. To overcome these objections, a device was developed for use in the field which permits determination of the concentration of treating acid both quickly and accurately at the time of the acid treatment. Its purpose was three-fold, namely, to determine: (1) the concentration of the fresh treating acid, (2) the concentration of the spent acid returning from the well, and (3) the solubility of formation samples in treating acid. The device was designed to operate on the principle of the stoichiometric relation between the calcium carbonate—hydrochloric acid reaction and the carbon dioxide gas resulting from the reaction. By confining the carbon dioxide gas within a reaction cell of the type shown in Fig. 1, the pressure inside the cell is increased. Since the pressure is an indication of the quantity of carbon dioxide gas produced, it may be used to estimate the concentration of the acid, provided an excess of calcium carbonate is used. From information on the volume of the reaction cell, the volume of acid used, and the pressure and temperature of the gas after the completion of the reaction, the concentration of the treating acid may be computed. The reaction cell is essentially a pressure vessel equipped with a pressure gage, temperature indicator. pressure release valve, and a special sample holder. Plastic and metal construction makes the cell both rugged and acid-resistant. The sample holder is so designed that a sample of calcium carbonate (or sample of carbonate rock) can be dropped into the acid

Jan 1, 1956

By M. E. Wadsworth, J. S. Cho

The kinetics of silicate adsorption from aqueous colloidal (SO2 . nH2O) silicate sols on thoria (ThO 2) surfaces were determined at 100, 150 and 200°C in an autoclave system. The amount of chemisorbed silica on the thoria surface was determined by means of infrared spectroscopy. First order kinetics were observed in all cases with the anomalous result that the first order rate constant was found to vary with the initial amount of silica added to the system. These results have been correlated quantitatively by means of a strain model of adsorption involving two types of surface silicate bonding, both of which may be readily observed in the infrared spectrum. Initial adsorption occurs rapidly to produce one type of surface silicate followed by a slow transition to a second type of surface bonding, the latter transition representing the slow step in the adsorption process. Initial interest in silicate adsorption on thoria arose from the dispersive effect silica has on finely divided thoria in aqueous suspensions. Spaepen et al studied the equilibrium adsorption of silica on thoria and noted increased adsorption up to approximately 215°C with a notable decrease in adsorption at temperatures above 215°C. Similar results were observed by Cho and Wadsworth2 who identified two types of surface silica by infrared spectroscopy. At temperatures of 100°C and above, rapid initial adsorption occurred producing a surface species with maximum infrared absorbance at 9.4 After extended periods of time, the 9.4 species converted to a more strongly bonded surface silicate with maximum infrared absorbance at 10.3. The latter was very strongly bonded and difficult to remove by desorption in hot distilled water, while the 9.4 species could be removed more readily by similar treatment. The 9.4 and 10.3-p species were distinguishable from free unadsorbed silica which had maximum infrared absorbance at 9.1 The kinetics of the formation of the strongly chemisorbed silica (10.3) on thoria as determined by infrared examination are presented in this study. EXPERIMENTAL PROCEDURE As a source of silicate, Ludox SM (16.36% SiO2) or Ludox HS (30.25% SO2), DU Pont colloidal silica in aqueous suspensions, was added directly to a thoria slurry made up of one liter of distilled water and one gram of thoria. The thoria used was designated D-16-650 and was received from the Chemical Technology Div. of Union Carbide Nuclear Co. It was prepared by firing thorium oxalate at 650°C. The D-16 thoria had a surface area of 32.4 ± 0.6 sq m per g as determined by nitrogen adsorption.* This thoria 'Determined by Unlon Carbide Nuclear Co., ORNL. had a total adsorption capacity for silicate of approximately 19.4 mg of silica per gram of thoria.1 This amounts to approximately one SiO2 molecule for every surface thoria site. In all cases in this study.

Jan 1, 1963

By Orville R. Lyons

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

Jan 1, 1953

By Orville R. Lyons

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

Jan 1, 1953

By CYRIL W., William Campbell, Henry M. Howe

THIS paper gives the results of an investigation of the behavior of the argentiferous cobalt-nickel arsenides of Temiskaming, Ontario, in roasting, made in the metallurgical laboratories of the School of Mines, of Columbia University in the City of New York. The ore was kindly given by the owners of the La Rose and Trethewey properties at Cobalt, Ontario, and Mr. E. J. Hall, Tutor in Assaying in Columbia University, has helped us much. I. OBJECT OF THE INVESTIGATION. The object of the investigation was to learn:- 1. The temperature at which the arsenic is most rapidly expelled; 2. The thoroughness with which it is expelled by prolonged roasting at this temperature; 3. The effect of adding charcoal (a) near the end of the roast and (b) at the beginning of the roast. II. NATURE OF THE ORE. The important ores mined in the Temiskaming deposits are native silver-with small amounts of dyscrasite (Ag6 Sb), argentite (Ag2S), pyrargyrite (Ag3SbS3)-smaltite (CoAs2), chloanthite (NiAs2) and niccolite (NiAs). Mispickel (FeAsS) and cobaltite (CoAsS) also occur in smaller quantities. The average composition of the ore shipped from this district for the first six months of 1905 -was: silver, from 4.1 to 4.8 per cent.; cobalt, from 6.9 to 8.2 per cent ; nickel, from 3 to 4.7 per cent ; and arsenic, from 30.9 to 34.6 per cent. The ore which we treated consists chiefly of smaltite ; In our laboratory-work

Jan 1, 1907

By G. Sachs, A. W. Dana, C. C. Chow

A great number of the investigations on the plastic flow of metals have been concerned with the establishment of a "universal" stress-strain relation. In such a relation some stress function when plotted against a strain function should yield identical curves for the various stress states. In the first investigation of this type, Ludwik and Scheu1 plotted the maximum shearing stress as a function of the maximum principal strain. Later Ros and Eichinger2 introduced two universal stress-strain relations, the one relating the maximum shearing stress to the maximum shearing strain, and the other relating a stress invariant, suggested by von Mises and Haigh, to the corresponding strain invariant. (In more recent investigations the stress and strain invariants are frequently supplemented with some factor to render their meaning more lucid.) A further suggestion which has not attracted appreciable attention is that by Baranski³ who used stress and strain deviators. The most common means of experimentation to determine the relation between stress and strain consists in subjecting thin walled tubes to combined internal pressure and axial tension.4a,4b,4c This method allows the study of plastic flow under stresses which are variable in two directions. However, the plastic flow which can be obtained in this manner is comparatively small, being limited by either tension failure or instability. For copper,'. only the relation between maximum shearing stress and maximum shearing strain yielded good agreement. On the other hand, tests on a stee14b and on an aluminum alloy4c. resulted in systematic deviations if any of the discussed universal stress-strain relations were used. It would seem, therefore, that the agreement mentioned above for copper is only incidental and explained by its high rate of strain hardening compared to that of other metals. Much larger strains than experienced in the tube tests can be obtained by subjecting a thin membrane of a ductile metal, which is restrained at its periphery, to a uniform hydraulic pressure. The thin sheet forms a deep bulge before it fails. The stresses and strains in such a bulge increase with increasing distance from the edge of the clamping "die," the maximum stresses and strains occurring at the pole (crown) of the bulge. While the stress and strain states are determined by the contour of the bulge, the absolute magnitude of the stresses and strains depends upon the hydraulic pressure. The bulge contour is in turn correlated with the geometry of the die opening. The deformation and fracture characteristics of circular bulges, that is, bulges formed with circular clamping dies, have been the subject of numerous experimental and analytical investi-gations.5,6,7 It has been shown that plastically deformed circular bulges develop large and comparatively uniform strains before failure by instability"6b,6c,6d and closely assume a spherical shape.6d Also the distribution of strains across the contour of the bulge is dependent on the metal being investigated and is correlated with, but cannot be predicted from, the metal's stress-strain characteristics. On the other hand, oblong or elliptical bulges, that is, bulges formed with elliptical clamping dies, are not as susceptible to analytical analysis and have not been investigated to the extent that circular bulges have. The few available data6c,7c indicate that stress states are obtained at the poles of the bulges, varying between plane strain and balanced biaxial tension, depending upon the geometry of the die opening. In this paper, the strain state and curvatures exhibited by three bulge shapes, a circular and two elliptical bulges, Fig 1, are analyzed experimentally using methods described in previous publications.6a,6c An attempt is made to derive the stress-strain relations for these bulges, which represent strain states in which the ratio of the two positive principal strains varied between 1.0 and 0.35. In addition, tension tests yielded data for a value of —0.5 for this strain ratio. Such an analysis should indicate the applicability of the various laws correlating stress with strain to the stress and strain states occurring in bulged shapes. Definitions and Nomenclature The definitions of the major stress and strain quantities used in this paper are as follows: s1, s2, s3 = principal normal stresses Sl > s2 > S3 t = shear stress e = conventional (unit) strain e = In (1 + e) El, E2, E3 = principal natural strains 7 = shear strain The maximum shear stress: , _ S1 — S3 lmax = 2 Frequently, the flow stress, s1 — s3 = 2lmax rather than the maximum shear stress is used.

Jan 1, 1950

By D. H. Gieskieng

THE advent of curved jaw crusher wearing plates made an approach other than segmental layout analysis desirable for prediction of capacities. For some time it had been known that the drawing board capacities of crushers using these plates had to be considerably modified by complicated experience factors to achieve agreement with results. Because these apparent capacities could be readily increased severalfold by minor crushing chamber shape changes, it was necessary that the utmost precaution be taken in predicting capacities of jaw plates modified for nonchoking, special wear characteristics, or any other reason. To this end the laboratory and field tests outlined by the author in a previous paper1 were made on Blake-type jaw crushers. The results of these tests were summarized in a simple first degree equation applicable to crushers using either straight or curved jaw plates. This equation first outlines the maximum capacity potential of a given crusher, then reduces this figure in accordance with installation circumstances by means of a realization factor. It was found subsequently that this equation, with the addition of an eccentric throw factor, is applicable to standard types of single-toggle or overhead eccentric jaw crushers as far as maximum capacity potential is concerned. However, these crushers have realization factor curves somewhat different from those outlined for the Blake type. While this paper is concerned principally with standard type single-toggle crusher capacities, the evaluation of data obtained with these machines is simplified by comparative reduction to the 10 x 7 in. Blake-type equivalents upon which the summary of the preceding paper was made. Convertibility of data from one type of crusher to the other also tends towards confirmation of both. The agreement of these data is sufficient to be considered complimentary. Consequently the feed factors, f, previously reported for Blake crushers are slightly adjusted to an average with the single-toggle crusher results. Blake-type equation: C = f.d.w.y.t.n.a.r [I] Single-toggle type equation: C = f.d.w.y.t.n.a.e.r [21 where C is the capacity in short tons per hour through the crusher, f is a feed factor, dependent upon the presence of fines in the feed, and the surface character of the jaw plates used. Values of f : Smooth Plates Corrugated Plates With normal fines 0.0000414 0.0000319 Fines scalped out 0.0000368 0.0000252 Large pieces only . 0.0000312 0.0000215 d is the apparent density of the broken product in pounds per cubic foot. (If the true specific gravity of the feed is known, 40 pct voids may be assumed and d becomes 37.4 times sp gr). w is the width of crushing chamber in inches. y is the openside setting of the crusher, in inches. In the case of corrugated jaw plates it is measured from the tip of one corrugation to the bottom of the valley opposite. t is the length of jaw stroke in inches at the bottom of the crushing chamber. It is the difference between open and close-side settings. n is rpm, or crushing cycles per minute, a is the nip-angle factor. It is unity for 26" and 3 pct greater for each less nip-angle degree. A nip-angle of 20" has an a value of 1.18, and an angle of 30" has an a value of 0.88, see Fig. 1. r is the realization factor. It is unity for perfectly uniform choke feeding and usually less for actual operating conditions according to the method of feeding used and the probabilities of hang-ups involving the size of feed and crusher opening. Approximate values are given by the curves in Fig. 2. These values are further reduced by intermittent feeding, e is the throw or diameter of gyration of the single-toggle crusher eccentric in inches. As evident in Fig. 1A, variation of feed size will generally have little effect on nip-angle if both jaw plates have flat areas. Jaw plates having continuous curvature, as in Fig. 1B will have different nip-angles, depending upon the size of feed. For test work as described in this paper this effect was accounted. For general compilation of capacities for average feeds it is suggested that the nip-angle be taken at the various settings computed, at an arbitrary level, such as is indicated in Fig. 1C. Data Evaluation To bring the Blake and single-toggle type crusher capacity test results to common terms for evaluation, all data are converted to terms of 10 x 7 in. Blake-type performance at conditions of 100 lb per cu ft, 10 in. chamber width, 250 rpm, 0.65 in. stroke, 3-in. openside setting, and 18° nip-angle. (The nip-angle of the 10x7 in. Blake is 18° at 3-in. setting.) The single-toggle crusher performances are also divided by the eccentric throw to bring this effect to unity. As outlined,' laboratory and field tests made on Blake-type crushers ranging from 10 x 7 in. to 60 x 48 in. were summarized along the foregoing conditions of speed, stroke, etc. This resulted in groups of data which correspond to feeds with fines, feeds

Jan 1, 1952

By H. J. McQueen, G. C. Kuczynski

ThE useful properties of k6-Alfenol and the thermenol alloys (summarized by J. E. Nachman and W. J. Buehler') have encouraged research into those alloys which, having compositions near 25 at. pct Al, 75 at. pct Fe, undergo an ordering transformation between 500° and 550°C. The equilibrium diagram of the iron-rich iron-aluminum alloys was first determined by Bradley and Jay.2,3 They distinguished three structures: the disordered, which is bcc; FeAl, which is simple cubic with one Fe atom and one Al atom per lattice site; and Fe3Al, which is fcc with three Fe atoms and one Al atom per lattice site. This last structure may be conceived as eight body-centered cubes with Al atoms at the centers of alternate ones, so that Al atoms have only Fe as first and second nearest neighbors. Later, Osawa and Murata4 determined the boundaries of the different superlattice regions in the iron-rich Fe-Al solid solution by dilatometric and X-ray methods. Two of the six phases, which they found between 0 and 50 at. pct Al, were never identified by other workers. A. Taylor and R. M. ones,' concurrently with and independently of this research, redetermined this diagram by measuring lattice parameters at temperature. Further reference will be made to their results. EQUILIBRIUM DIAGRAM A) Experimental Procedure—The alloys were prepared from NRC gas-free iron (99.9 pct Fe) and 99.99 pct Al. They were melted in alumina' crucubles under vacuum in an induction furnace. If the ingots were unsound or showed more than a 3 pct weight loss, they were rejected. The ingots were homogenized at 1100°C for 24 hr before hot-rolling at 700" to 900°C from 1 in. to 0.25 in. thick. It is interesting to note that alloys containing more than 28 at. pct Al were able to take an 80 pct reduction of area in the ordered FeAl region. The dilatometer specimens and specimens for chemical analysis were cut from these plates. A Leitz dilatometer was used to measure the coefficient of linear expansion as a function of temperature. The specimens were slowly cooled from 900" to 300°C where they were annealed for 10 hr before slowly cooling to room temperature. The expansion was recorded photographically as the alloys were heated and cooled at 1.5oC per min between 100" and 1000°C. The magnetic measurements determining the Curie points were performed on the dilatometer specimens in the dilatometer, by using direct current in the furnace windings both to heat the specimen at 1.5°C per min and to serve as the primary coil. The regular specimen holder was replaced by a thin alundum tube on which was wound, as a secondary, a closely spaced coil of fine copper wire. A reversal of current in the furnace windings produced, in the secondary, an emf which was proportional to the induction in the specimen, and which could be detected on a galvanometer. In this way, it was possible to determine the Curie points of alloys to within * 5°C. B) Results and Discussion—The coefficients of linear expansion, as obtained from the dilatometer curves, are plotted as a function of temperature and composition in Fig. 1. Since in many cases the heating and cooling curves were superimposed, it was decided to measure slopes of heating curves only. Definite high peaks, due to the transformations, were observed on each curve between 400" and 520°C, and, for alloys containing more than 23 at. pct Al, a second set of smaller peaks at temperatures above 660°C. The sharpness of the peaks indicates that ordering to Fe3Al is a first-order reaction. The compositions and the temperatures of these peaks are listed in Table I. When these temperatures are plotted as a function of the composition, the result is the constitution diagram presented in

Jan 1, 1960

By G. E. Steudel

The study of the possibility of effecting a lower cost in the manufacture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the consideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improvements, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph,5 whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality. Large slag volumes result in higher coke consumption and cost of disposal. Any attempt to lower slag quantity must be accomplished

Jan 1, 1940

By G. E. Steudel

The study of the possibility of effecting a lower cost in the manufacture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the consideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improvements, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph,5 whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality. Large slag volumes result in higher coke consumption and cost of disposal. Any attempt to lower slag quantity must be accomplished

Jan 1, 1940