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By D. E. Wolfson, D. A. Reynolds, F. W. Smith

IN 1948 the U. S. Bureau of Mines began a three-phase program to evaluate the extent and quality of U. S. coking coal: 1) a factual appraisal of known recoverable reserves in beds of mineable thickness, 2) a study of preparation characteristics, and 3) a study of carbonization properties. A previous paper presented the methods used in assessing reserves.' Although this investigation is only partly completed, data on coking properties of the Pittsburgh district coals are now available. The results are presented in this article. When coals are selected for manufacture of metallurgical coke more factors must be considered than for any other major use, as no basic industry requires coal of such high quality and uniformity. This is particularly true in the Pittsburgh district, where blast furnace practice is based on use of cokes from high quality coals. Although it is true that inferior coals are carbonized in other industrial areas, both here and abroad, the Pittsburgh coals are the best economically obtainable in the required tonnages, even though they are not the most desirable for metallurgical use. In the general evaluation of suitability of coals for coke making, particular attention is given to: 1) reserve, 2) ash and sulfur content, 3) inherent properties of forming strong, well fused coke, 4) expanding characteristics, and 5) blending properties. Items 1 and 2 are covered in detail in the first two phases of the survey, and coking behavior is considered in items 3 through 5. Problems is Coal Evaluation: Although not all coal beds of immediate or potential use to the coking industry in the Pittsburgh area have been tested for their coking properties, enough tests have been made to permit certain generalized statements. Virtually all the coals that will be used in this district are from the Appalachian region, and except for high oxygen coals from outcrops or stripping operations, they are coking coals. Hence the evaluation problem is not so much that of distinguishing between coking and noncoking coals as attempting to measure quantitatively differences in the coking and expanding properties of the better coking coals. Quantitative differentiation between good coking coals, is complex, since differences in their coking abilities are frequently of about the same magnitude as the precision of the test. Small-scale tests are generally unsuitable for measuring these slight differences, and there has been gradual acceptance of pilot-scale tests. Test Methods Employed: All the data reported here were obtained in two units known as the Bureau of Mines-American Gas Association (BM-AGA) carbonization apparatus and the sole-heated expansion oven. The BM-AGA unit, which consists essentially of a cylindrical steel retort about 18 in. diam and 26 in. high, carbonizes enough coal (180 lb) to allow evaluation of the coke by standard methods of the American Society for Testing Materials. The BM-AGA carbonization procedure and methods of testing the gas, liquid, and solid products have been described.' Expanding characteristics of the coals were determined in the sole-heated expansion oven," wherein the coal charge, heated from the bottom only, is carbonized under a constant applied force and the linear expansion or contraction is measured. Although results in this oven have never been quantitatively correlated with the expansion behavior of coals in conventional slot ovens, they are useful in assessing expansion behavior. In this report discussion will be confined to results obtained from these pilot-scale units, although the following data are available for most coals: proximate and ultimate analyses, free swelling index, agglutinating value, and plasticity. Sources of Coals Tested: In selection of coals for coke manufacture, economic considerations are as important as quality, and each coke plant balances these two factors to meet requirements of the most satisfactory coke at lowest price. Although these economic factors may in some instances outbalance small differences in coke quality, they are usually specific to each coke plant and cannot be given in any general statement. Recognizing that coal and plant operating costs are beyond the scope of the survey, the USBM evaluates the coking quality of various coals without regard to their competitive economic status. For the purpose of this article, coals of interest to the Pittsburgh district are arbitrarily defined as

Jan 1, 1958

By M. J. Mullikin, J. B. Conway

Equations capable of representing combined first -and second-stage creep behavior include the Andrade, Cottrell-Aytekin, Garofalo, and de Lacombe equations along with a third-degree polynomial in t1/3. Employing some recent high- temperature creep data for arc-cast tungsten the effectiveness of these expressions was evaluated. It was found that, in this instance, the most accurate representations were afforded by the third-degree polynomial in t1/3 and the de Lacombe equation. Also noted was the fact that the Andrade equation gave results just slightly inferior to those obtained with the third-degree polynomial, definitely better than those due to the Cottrell-Aytekin equation, and slightly superior to those furnished by the Garofalo equation. EXPRESSIONS designed to describe combined first -and second-stage creep behavior must possess certain specific characteristics. At time zero the equation must reduce to a strain value corresponding to the instantaneous deformation on loading. Then in the early moments following load application transient-creep behavior must be described to reflect a creep rate which constantly decreases with time. And finally at larger time values linear (or at least nearly linear) strain-time behavior must be accounted for to yield the steady state or minimum creep rate. Some equations, such as the one based on the parabolic creep law, yield fairly satisfactory results in transient analyses but are not effective at all when combined first- and second-stage data are to be analyzed. Other equations give equally accurate results in both instances although in general the complexity of the equation is slightly increased. For example the Andrade1 equation: has found fairly general application in expressing first-stage creep behavior. Furthermore the contribution of the exponential term becomes more and more prominent as the time increases and causes this equation to exhibit a point of inflection at large values of time. With a proper selection of equation constants the curve can be made fairly flat in the region of the inflection point indicating a desirable equation form for combined first- and second-stage creep analysis. Another equation form with the proper characteristic is that of Cottrell and Aytekin:2 e =a +bt1/3 +ct [2] Obviously as time increases the first derivative of this expression approaches a constant value to suggest application to combined first- and second-stage creep analysis. Similarly a third-degree polynomial in t1/3: e -a + bt1/3 + ct2/3 +dt [3] has the same characteristics. Expressions of this type have been shown3 to be very effective in the analysis of transient-creep behavior and the fact that the first derivative has the same characteristic as Eq. [2] suggests use in the combined analysis.

Jan 1, 1967

By F. A. Forward, J. Halpern

A new process for extracting uranium from ores with carbonate solutions is described. Leaching is carried out under oxygen pressure to ensure that all the uranium is converted to the soluble hexavalent state. By this method), alkaline leaching can be used successfully to treat a greater variety of ores, including pitchblende ores, than has been possible in the past. The advantages of carbonate leaching over conventional acid leaching processes are enhanced further by a new method which has been developed for recovering uranium from basic leach solutions. This is achieved by reducing the uranium to the tetravalent state with hydrogen in the presence of a suitable catalyst. A high grade uranium oxide product is precipitated directly from the leach solutions. Vanadium oxide also can be precipitated by this method. The chemistry of the leaching and precipitation reactions are discussed, and laboratory results are presented which illustrate the applicability of the process and describe the variables affecting leaching and precipitation rates, recoveries, and reagent consumption. THE extractive metallurgy of uranium is influenced by a number of special considerations which generally do not arise in connection with the treatment of the more common base metal ores. Perhaps foremost among these is the very low uranium content of most of the ores which are encountered today, usually only a few tenths of one percent. A further difficulty is presented by the fact that the uranium often occurs in such a form that it cannot be concentrated efficiently by gravity or flotation methods. In these and other important respects, there is evident some degree of parallelism between the extractive metallurgy of uranium and that of gold and, as in the latter case, it has generally been found that uranium ores can best be treated directly by selective leaching methods. It is readily evident that this parallel does not extend to the chemical properties of the two metals. Unlike gold, which is easily reduced to metallic form, uranium is highly reactive. It tends to occur as oxides, silicates, or salts. Two ores are of predominant importance as commercial sources of this metal: pitchblende which contains uranium as the oxide, U3O51 and carnotite in which the uranium is present as a complex salt with vanadium, K2O-2UCV3V2O5-3H2O. These ores may vary widely in respect to the nature of their gangue constituents. Some are largely siliceous in composition, while others consist mainly of calcite. Sometimes substantial amounts of pyrite or of organic materials are present and these may lead to specific problems in treating the ore. Further complications may be introduced by the presence of other metal values such as gold, copper, cobalt, or vanadium whose re- covery has to be considered along with that of the uranium, or whose separation from uranium presents particular difficulty. In general, there are two main processes for recovering uranium in common use today.'.2 One of these employs an acid solution such as dilute sulphuric acid to extract the uranium from the ore. A suitable oxidizing agent such as MnO, or NaNO, is sometimes added if the uranium in the ore is in a partially reduced state. The uranium dissolves as a uranyl sulphate salt and can be precipitated subsequently by neutralization or other suitable treatment of the solution. The second process employs an alkaline leaching solution, usually containing sodium carbonate. The uranium, which must be in the hexavalent state, is dissolved as a complex uranyl tricarbonate salt, and then is precipitated either by neutralizing the solution with acid or by adding an excess of sodium hydroxide. The latter method has the advantage of permitting the solutions to be recycled, since the carbonate is not destroyed. This is essential if the process is to be economical, particularly with low grade ores. With each of these processes, there are associated a number of advantages and disadvantages and the choice between using acid or carbonate leaching is generally determined by the nature of the ore to be treated. In the past, more ores appear to have been amenable to acid leaching than to carbonate leaching and the former process correspondingly has found wider application. With most ores, acid leaching has been found to operate fairly efficiently and to yield high recoveries. One of the main disadvantages has been that large amounts of impurities, such as iron and aluminum, sometimes are taken into solution along with the uranium. This may give rise to a high reagent consumption and to difficulties in separating a pure uranium product. Excessive reagent consumption in the acid leach process also may result

Jan 1, 1955

By Lewis A. Riley

I desire to submit, for the consideration and information of the members of the Institute, the following data, drawings, and tables, showing what I believe will be interesting information with regard to the diamond drill, and its uses in developing mineral lands. I would first say that the results given have been obtained during the past two years by means of two drilling machines belonging to the Lehigh Valley Coal Company, operating on their coal lands in the Mahanoy, Lehigh, and Wyoming regions. 1st. A tabulated statement, giving location of borings, date, progress, total cost, itemized cost per foot, etc. 2d. Section showing strata bored in Nos. 5 and 10 holes at Mt. Carmel, see Plate VI. 3d. Section showing strata bored in No. 19 hole at Hazleton, see Plate VI. 4th. Photograph and description of drill No. 2. From the first it will be seen that the total number of test-holes bored up to May 1st of this year is 24; 9 of these are in the Mahanoy region, 6 in the Lehigh region, and 10 in the Wyoming region, in all of which cores have been taken for the whole distance. The total distance bored is 9902 feet. The majority of the holes have been put down for the purpose of proving the lower veins of the coal series, and have had to encounter the harder rocks of the coal formation. Much of the distance has been through the lower conglomerates, going in some cases through the coarse egg conglomerate, the foundation of the coal deposit, and to the green sandstone and red shale which underlie it. These lower sandrocks and conglomerates are among the hardest known to geologists, and present a formidable resistance to the exploring drill. This fact must be taken into consideratiotl in comparisons of cost of drilling, and when considered will still further improve the very favorable results I am able to show in my statement of costs. In illustration, I will cite the work of one bit when boring in the upper or softer rocks of the Wyoming region, where 280 feet were obtained without resetting, while in the Lehigh and Mahanoy regions, in the lower and harder rocks, bits set with the same quality of diamonds would only bore from 3 to 20 feet before being worn out. In this connection it may

By Louis A. Riley

I DESIRE to submit, for the consideration and information of the members of the Institute, the following data, drawings, and tables, showing what I believe will be interesting information with regard to the diamond drill, and its uses in developing mineral lands. I would first say that the results given have been obtained during the past two years by means of two drilling machines belonging to the Lehigh Valley Coal Company, operating on their coal lands in the Mahanoy, Lehigh, and Wyoming regions. 1st. A tabulated statement, giving location of borings, date, progress, total cost, itemized cost per foot, etc. 2d. Section showing strata bored in Nos. 5 and 10 holes at Mt. Carmel, see Plate VI. 3d. Section showing strata bored in No. 19 hole at Hazleton, see Plate VI. 4th. Photograph and description of drill No. 2. From the first it will be seen that the total number of test-holes bored up to May 1st of this year is 24; 9 of these are in the Mahanoy region, 6 in the Lehigh region, and 10 in the Wyoming region, in all of which cores have been taken for the whole distance. The total distance bored is 9902 feet. The majority of the holes have been put down for the purpose of proving the lower veins of the coal series, and have had to encounter the harder rocks of the coal formation. Much of the distance has been through the lower conglomerates, going in some cases through the coarse egg conglomerate, the foundation of the coal deposit, and to the green sandstone and red shale which underlie it. These lower sandrocks and conglomer¬ates are among the hardest known to geologists, and present a formidable resistance to the exploring drill. This fact must be taken into consideration in comparisons of cost of drilling, and when considered will still further improve the very favorable results I am able to show in my statement of costs. In illustration, I will cite the work of one bit when boring in the upper or softer rocks of the Wyoming region, where 280 feet were obtained without resetting, while in the Lehigh and Mahanoy regions, in the lower and harder rocks, bits set with the same quality of diamonds would only bore from 3 to 20 feet before being worn out. In this connection it may

Jan 1, 1877

By R. G. Wayland

SINCE the installation at Kropfmuehl, Bavaria, of a modern flotation concentrator in 1938, the flake and fine graphite from the Passau area can now be delivered in about any normal specified carbon content of any size range up to a flake averaging about 0.7 mm. The graphite finds a wide German and export market for crucible manufacture, pencil leads, dry cells and other uses. A controversy over the origin of the graphite deposits is being resolved in favor of syngenesis rather than epigenesis. The syngenetic theory is newly supported by the yet unpublished work of Hartmann of the Bavarian Geological Survey. Development work and exploration for graphite in the area may be changed in direction as the syngenetic theory is accepted. Crystalline graphite is produced in the area east of Passau near the junction of Germany, Austria and Czechoslovakia, as shown in Fig. 1. This is the only graphite area of importance in Germany, and Gra-phitwerk Kropfmuehl AG is the only operating firm in the area. The plant and mine are located at Kropfmuehl near Hauzenberg, Kreis Wegscheid, about 10 miles east of Passau. A narrow gage railway from the mine connects with the German Railway at Schaibing Bahnhof. The Pfaffenreuth mines date from about 1730. Until the early 20th century mining operations were carried out in a haphazard fashion. During World War I graphite mining and milling was increased, since it had to cover almost all of the crucible needs of the Central Powers. Between the wars some 11 mines were operated by two large and several small companies, but under the Nazis these were consolidated by 1938 into the Kropfmuehl enterprise. Kropfmuehl built a modern flotation mill to treat its own ores and small amounts of custom ores and tailings from the area. Since Graphitwerk Kropfmuehl AG was an I.G. Farbenindustry subsidiary, it has been under Military Government Property Control and probably will be sold to private German capital. Geology The country rock of the graphite area is part of the "kristallines Grundgebirge," the series of old gneissic and schistose rocks that constitutes the bed rock of most of the Bohemian basin and rims the Sudeten-land. The gneissic rocks of the graphite area are considered to have been metamorphosed during the Carboniferous period. They are bordered on the north by granite stocks and penetrated by numerous smaller granite and pegmatitic intrusive rocks, as shown in Fig. 1. The gneiss is classed as a micaceous, coarse-grained cordierite gneiss by most investigators. It is much metamorphosed by the granite, particularly in the north near the larger granite bodies. Interbedded in the gneiss are the graphite seams and lenses, and also beds and lenses of crystalline limestone containing disseminated graphite in noncommercial quantities. The gneiss, together with the included graphite and limestone seams and lenses, is cut and displaced by a number of granite sills of medium to fine grain and by a large number of diorite lamprophyre dikes and a few syenite-pegmatite dikes. The lamprophyre dikes are of various mineral compositions and textures, but many are banded and richly impregnated with pyrite; while the syenite-pegmatite dikes are coarse-grained with good crystals of titanite, pyroxene, uralite and other green amphiboles. Most investigators and the miners speak only of diorite and granite dikes cutting the graphite seams. The diorite dikes are later than the granite and some of the faulting, as is evident from Figs. 1 and 2. Individual graphite seams and lenses may be mined for thicknesses of several feet up to several scores of feet, and for distances of several hundred feet. The aggregate thickness of a series of some 20 seams of graphite, limestone and interbedded gneiss at Kropfmuehl is stratigraphically about 450 ft. Laterally, the graphite in a seam may pinch out or grade into crystalline limestone. Graphite crystals also are found disseminated in the gneiss itself, although in unmineable concentrations. The faults that cut the graphite seams carry graphite for some feet or tens of feet away from the seams, apparently mechanically. Similarily, the graphite lenses themselves often contain mechanically introduced inclusions of wall rock, probably from flowage during folding. Graphite crystals make up 10 to 30 pct of the fresh, mineable graphite lenses at Kropfmuehl, averaging about 20 to 25 pct after hand-sorting by the miners. In weathered lenses, the graphite concentration is said to be as high as 50 pct. The associated primary and hydrothermal minerals are dom-inantly feldspar and calcite, plus quartz, pyrrhotite, pyrite, biotite and occasional garnet, hornblende, sphalerite and galena. Associated secondary minerals include kaolin, nontronite, mangano-oxide-silicates (mog), adularia and chlorite. The superimposed suite of siliceous cementation minerals present consist largely of opal, chloropal, chalcedony, jasper, and hyalite. The kaolin is of special interest since it too was mined as early as 1730 and was used in the well-known Nymphenburg porcelain from 1756 on. The kaolin is derived from the gneiss and the syenite pegmatites. The crystals of graphite vary in size within a given seam, but in seams more than a mile away from the granite on the north the average crystal-linity is less coarse, lowering the commercial value. The lenses in the Kropfmuehl-Pfaffenreuth area are the most developed, and are the only ones with deep workings now accessible. Other similar crystalline graphite lenses are known from older workings at Habersdorf, Oberoetzdorf, Ficht, Diendorf, and

Jan 1, 1952

By R. S. Olson

An analysis of the wash thickener-two stage filtration circuit was undertaken to quantitatively describe the circuit performance. On the basis of information developed in this study, the factors of major importance in reducing soluble loss are: filter wash, wash zone size, filter capacity, filter feed pulp dilution, blow back, filter cloth, water to repulp, water to primary wash, water to feed pulp and thickener underflow. The wash thickener-two stage filtration circuit has been used in many hydrometallurgical plants to separate solute from leach solids. As many as 20 to 30 plant operating variables determine the performance of the wash circuit and hence the soluble loss. The complex relationships of these variables, with each other and with soluble loss, make the evaluation of their effect on mill performance difficult. An analysis of this type of circuit was undertaken to quantitatively describe the circuit performance. THE CIRCUIT The diagram in Fig. 1 represents the operation of the wash thickener-two stage filter circuit. This circuit provides for counter current, cross current or for a combination of countercurrent-cross current operation. The feed pulp stream, J, is intended to include any solution or water, other than primary and secondary filtrate, which enters the wash thickener. Except for filter washing, the operations in the circuit are simple mixing operations which can be described by material balance. The concentration of solute in the filtrate and cake solution after application of wash liquor cannot be described by simple material balance. The displacement washing equation, derived by Rhodes1 and extended by Choundhury and Dahlstrom,2 appears to best describe this case of miscible displacement in porous media. In a useful form, it is

Jan 1, 1963

By P. J. Potter

PRACTICALLY all tin-base babbitt metals used in engine bearings are made to customers' specifications, which are many and varied. The copper ranges from 3 to 8 per cent. and the antimony from 4 to 13 per cent.; generally, the babbitt with lower copper content will contain from 4 to 8 per cent. antimony and that with higher copper will have from 7 to 13 per cent. antimony. The allowable lead content varies from 0.20 to 2.00 per cent. If a solder is used as a bonding material instead of tin, the resulting material in the finished bearing will have a higher lead content, but it would not be enough to bring the lead above specificational limits. Impurities such as iron, arsenic and bismuth should be determined in the tin before using to insure a uniform and high-grade product. These points must be considered when segregating and grading the waste that comes from the finishing operations on lined bearings.

Jan 1, 1929

By George L. Lyle

WITH the advent of World War II, the need for agricultural products rose phenomenally and caused a similar increase in demand for plant foods of which phosphate rock is one of the more important raw materials. By the close of the war or shortly thereafter, all of the operating companies were expanding their production facilities. Many companies installed modern bulk material handling facilities. Designs varied, but were generally of the overhead trestle type of storing conveyors. American Cyanamid studied these various installations but was not satisfied with the conventional system because of high initial cost, increased maintenance problems in the way of painting structural steel that might be submerged in stored materials, and because of slightly increased operating costs. As a consequence of these economic analyses, conferences were held with the various conveyor manufacturers to develop a better method of storing and reclaiming. After considerable engineering study, one conveyor manufacturer suggested the unique system which was installed in 1950.

Jan 8, 1954

By C. T. Sheeran, J. C. Franklin

A commercially available microprocessor-based system capable of monitoring 512 channels has been interfaced with monitors for radon, working level, air velocity, and fan power. The basic system utilizes both 280 and 8080 microprocessors in a desktop central processing unit (CPU). CPU hardware includes keyboard, video display (CRT), and disk drive; a printer is used to keep permanent data records. Signals from all channels are transmitted to the computer in digital form where they are processed for alarm status. Software developed for the system allows for audiovisual alarms in the event of low and high readings, rate change, change of state, or communica- tion failure. Up to six channels can be continuously displayed on the CRT for current readings. Shift reports and trend logs may be generated to help the ventilation engineer determine average working levels and ventilation effectiveness. Additional software permits the operator to program command sequences which may be used to automatically restart fans after a power outage.

Jan 1, 1982

By Gerhard Bockstiegel

Kuczynski's formula has been derived for the case of nonspherical particles. TWO formulae of Kuczynski's type have been derived, one describing the increase in tensile strength, the other describing the progress of shrinkage of a powder compact. It has been strength,shown that the exponents of all three formulae each contain two magnitudes of different physical characters, viz, the geometrical factor a and the kinetic factor ß. The interrelationships between the three exponents are stated. SOME years ago Kuczynski1 experimentally showed that the radius, x, of the area of contact between very small spherical metal particles and a metallic block is related to the time of sintering, t, by the following equation x = constant tk [11 where k has the value 1/5 or 1/7. Assuming that the metal particles were perfect spheres and the metallic block was perfectly flat, he derived the foregoing equation from theoretical considerations of the process of material transport in metals, and he showed that exponent k is different for different mechanisms of transport, e.g., k = 1/2 for viscous flow (according to Frenkel2), k = 1/3 for evaporation and condensation, k = 1/5 for volume diffusion, and k = 1/7 for surface diffusion. From this Kuczynski concluded that the mechanism of transport was either volume diffusion or surface diffusion, depending on whether the value of k, as found in his experiments, was 1/5 or 1/7. Subsequently. Cabrera8 corrected Kuczynski's calculations with regard to surface diffusion, showing that the theoretical value of exponent k is 1/5 for both volume and surface diffusion. He supposed that the different experimental values of k were due to slight differences in the shape of the metal particles. An exponential relationship similar to the aforementioned was found by Okamura, Masuda, and Kikuta,4 Masuda and Kikuta, and Takasaki8 when studying the rate of shrinkage on powder compacts during sintering. The authors measured the shrinkage by means of the fraction w = Vp — V./Vp — V,,,, where V,, is the volume of the green compact, V, is the volume of the sintered compact, and V,,, is the volume of the compact in its densest state. This fraction, w, they found, is related to the time of sintering, t, by the equation w == constant tm. [21 Further, Bockstiegel, Masing, and Zapf7 observed that the tensile strength, s, of sintering iron powder compacts can also be related to the time of sintering, t, by an equation of the foregoing type, i.e., s = constant tn. [3] For exponent n the values 0.28 (S=2/7) and 0.35 ( 2/5) were obtained, and the authors pointed out that there might exist a simple interrelation between exponent n as found in their experiments and exponent k in Kuczynski's equation. The authors supposed that 2k = n, since the strength of adhesion between a metal sphere and a block (as in Kuczyn-ski's experiments) must approximately be proportional to their contact area, p. x2. Theoretical Considerations This paper is an attempt to correlate the fundamental experiments of Kuczynski's type with the results obtained with powder compacts as represented by Egs. 2 and 3. In particular, the paper is to show how the rate of sintering is influenced by the geometry of the sintering particles and by the type of material transport. As the geometry of particles conglomerating in a powder compact is very complex, some simplifying assumption has to be made, of course, in order to adapt the problem to mathematical treatment. In the following paragraph a suitable simplification is introduced, and Kuczynski's formula is derived for the case of nonsphcrical particles. Relation Between Area of Contact and Sintering Time—As the face of contact between two particles in a sintering powder compact is not necessarily a circle (as in the case of spheres sintering to a block), Kuczynski's formula is modified as follows: Let the perimeter of the face of contact be described by means of polar coordinates R, 4, as shown in Fig. la, so the area of contact, f, is determined by f= 1/2 . S112p[R(Æ) ]2 dÆ [4] Then, let the two particles be intersected by a plane perpendicular to area f. The intersection is shown in Fig. lb. According to the nomenclature in this figure, the distance, h, between the surfaces of the two particles is a function of T and Æ: h = h(r,Æ). For the particular case of spherical particles, as in Kuczynski's theory, this function becomes: h = constant r2. It shall be assumed here that in the close neighborhood of their

Jan 1, 1957

The Institute conducts jointly with the American Society of Civil Engineers, American Society of Mechanical Engineers and American Institute of Electrical Engineers, certain activities as listed below, and is joint owner with them of certain properties. United Engineering Trustees, Inc.* When Mr. Carnegie made his gift which made possible a home for the engineering societies, a separate corporation was organized to act as a holding company and to manage the common property. This was long known as the United Engineering Society but the name has now been changed to one which more nearly reflects the real function of the corporation. It holds title to the Engineering societies Building and to various funds placed in its charge and is managed by a Board of Trustees of which three are chosen by each of the four societies, American Institute of Mining and Metallurgical Engineers, American Society of Mechanical Engineers, American Institute of Electrical Engineers and American Society of Civil Engineers, which own equal and undivided interests in the property. The land and building cost two million dollars. The interest of each of the Founders societies, aside from especial trust funds, is now valued at $493,352. The A. I. M. E. trustees are Arthur S. Dwight, J. V. N. Dorr and H. A. Guess. George D. Barron will succeed Mr. Guess as trustee at the end of this year. The President of the Board is Francis Lee Stuart and the Secretary, Alfred D. Flinn. The Engineering Foundation The Engnieering Foundation was established in 1814 as a result of a gift of $200,000 made by Ambrose Swasey. It is intended not only "for the Furtherance of research" along broad lines but also "for the advancement of" a wide variety of other activities of benefit to engineering as need arises. The endowment fund, which now amounts to $630,000, is administered by the United Engineering Trustees, Inc. Through The Engineering Foundation board, on which the Institute is now represented by J. V. N. Dorr, John M. Callow and Galen H. Clevenger, the income of the endowment, approximately $28,000 a year, and funds contributed for special projects are appropriated for the purposes intended. H. C. Bellinger will succeed Mr. Callow on the board at the end of this year. Alfred D. Flinn is Secretary- and Director. Researches of wide interest in many lines have been aided by Engineering Foundation. . A grant from it to the Mining Methods Committee assisted in the preparation of Volume 72 of the Transactions. For several years the Foundation has aided in support of the study of Blast-furnacc Slags by R. 8. McCaffery at the University of Wisconsin, sponsored by the Institute. The Foundation has raised over $230,000 to support the Alloys of Iron Research endorsed by the Institute and George B. Water-house is serving as Chairman and Director of the Iron Alloys Committee which has been brought together to direct this work. The Foundation has also on recommendation of the Institute made a grant in aid of the studies of Mining and Strata being conducted at Columbia University by P. B. Bucky, which promise information of great usefulness relative to mine subsidence and support of ground. John Fritz Medal Board of Award The medal was established in 1902 and lias since, been awarded not oftener than annually by a hoard of sixteen members, four representatives each from the American Societies of Civil, Mining and Metallurgical, Mechanical, and Electrical Engineers.

Jan 1, 1930

By John H. Melvin

THE Omaha District, Corps of Engineers, has been doing subsurface exploration work for a number of years, both by contract and with its own forces. Certain practices and procedures concerning the recovery of large diameter cores have been developed which will be described in this paper. GEOLOGY Most of the drilling to date has been in connection with structures on the main stem of the Missouri River above Sioux City, Iowa. A wide variety of subsurface conditions has been encountered. The bedrock formations appear in regular order from the oldest to the youngest, traveling upstream; or in other words, the general regional dip is to the northwest. The oldest formation is the Carlile shale, a compaction or noncemented waxy shale of Cretaceous age. Because of its uncemented, soft nature this shale has a tendency to grind or wash away when drilled. It also has a tendency to rebound or expand when the overlying pressures are removed. A 25-ft bed of sharp, friable sandstone called the Codell member is present near the top of the Carlile shale. This member is a cemented sediment, but in drilling, many of the sand grains are broken loose and exert extensive abrasive action on the core barrel and bit. Above the Carlile shale is the Niobrara chalk, a soft calcareous deposit with shaly horizons and some thin beds of bentonitic clay. The average thickness of this deposit is about 145 ft. It is soft enough to be sawed or bored with ordinary woodworking tools, yet is compact enough to be used as building block. The Pierre shale, the formation above the Niobrara chalk, is over 800 ft in thickness and is the most widespread material in the area. It is a dark colored, compaction shale containing numerous thin bentonitic clay seams and at certain horizons, hard manganese concretions. The shale has numerous faults and in addition, along the valley walls, has undergone considerable - slumping. Joints, fault planes and slickensided surfaces are common features encountered in drilling and add to the difficulties of high core recovery. East of the Missouri River the bedrock is generally covered with from 100 to 200 ft of glacial till, composed of clays, silts, sands, gravels and boulders. The valley itself is filled with alluvial deposits, mostly of the sand and gravel sizes, and contains a buried preglacial or interglacial channel over 100 ft in depth. The overburden deposits require sampling in connection with engineering studies and must also be cased before the underlying bedrock can be cored. The diameter of the required overburden samples can thus have an indirect effect on the size of core recovered from the bedrock. West of the river, which in most cases is the western boundary of glaciation,

Jan 1, 1947

By John H. Melvin

The Omaha District, Corps of Engineers, has been doing subsurface exploration work for a number of years, both by contract and with its own forces. Certain practices and procedures concerning the recovery of large diameter cores have been developed which will be described in this paper. Geology Most of the drilling to date has been in connection with structures on the main stem of the Missouri River above Sioux City, Iowa. A wide variety of subsurface conditions has been encountered. The bedrock formations appear in regular order from the oldest to the youngest, traveling upstream; or in other words, the general regional dip is to the northwest. The oldest formation is the Carlile shale, a compaction or noncemented waxy shale of Cretaceous age. Because of its unce-mented, soft nature this shale has a tendency to grind or wash away when drilled. It also has a tendency to rebound or expand when the overlying pressures are removed. A 25-ft bed of sharp, friable sandstone called the Codell member is present near the top of the Carlile shale. This member is a cemented sediment, but in drilling, many of the sand grains are broken loose and exert extensive abrasive action on the core barrel and bit. Above the Carlile shale is the Niobrara chalk, a soft calcareous deposit with shaly horizons and some thin beds of bentonitic clay. The average thickness of this deposit is about 145 ft. It is soft enough to be sawed or bored with ordinary woodworking tools, yet is compact enough to be used as building block. The Pierre shale, the formation above the Niobrara chalk, is over 800 ft in thickness and is the most widespread material in the area. It is a dark colored, compaction shale containing numerous thin bentonitic clay seams and at certain horizons, hard manganese concretions. The shale has numerous faults and in addition, along the valley walls, has undergone considerable slumping. Joints, fault planes and slickensided surfaces are common features encountered in drilling and add to the difficulties of high core recovery. East of the Missouri River the bedrock is generally covered with from 100 to 200 ft of glacial till, composed of clays, silts, sands, gravels and boulders. The valley itself is filled with alluvial deposits, mostly of the sand and gravel sizes, and contains a buried preglacial or interglacial channel over 100 ft in depth. The overburden deposits require sampling in connection with engineering studies and must also be cased before the underlying bedrock can be cored. The diameter of the required overburden samples can thus have an indirect effect on the size of core recovered from the bedrock. West of the river, which in most cases is the western boundary of glaciation,

Jan 1, 1949

By RUSSELL C. FLEMING

THE important advances that have been made of recent years in mining and milling methods and in mechanical equipment at mines need no re- telling, but there has been a remarkable growth in one type of equipment which has gone nearly unnoticed. That is the increasing use of storage-battery locomotives underground. Since the war the number of these locomotives and their field of use have increased tremendously; their growing acceptance represents an al- most complete reversal of opinion from an earlier period when they were generally looked upon as in- adequate for the strain of mine service. At the time when storage-battery powered units were first making their bid for recognition, the use of power in haulage had already passed through several phases; hand tramming, rope haulage, compressed air locomotives, even gasoline locomotives, had all been used, ending finally in the general adoption of the electric trolley locomotive for practically all power-haulage requirements. It was, and is, conceded that the trolley loco- motive met more adequately the varied and severe conditions of mine service than any previous type of haul- age unit. Subsequent developments have had to endure comparison and surpass in performance that type of haulage before changes would be considered.

Jan 1, 1930

By P. D. White, J. T. Moss, J. S. McNiel

Results of experimental work on the in situ combustion process were first published in this country in 1953' when Kuhn and Koch described results of a three-well test in Jefferson County, Okla. Shortly there-after Grant and Szasz2 described field studies in Nowata County, Okla., in which several modifications of the basic combustion process were investigated by injection of various oxidizing gas mixtures other than air. In the past five years speculation concerning this method of oil recovery has continued, with a number of interesting papers on various aspects of in situ combustion appearing in the literature3,4,5,6,7. However, only one of these' provided actual engineering information on pattern-type field experiments. The primary objectives of the Oklahoma three-well test were to check operation of specially designed electrical ignition equipment and evaluate problems associated with initiation of combustion in a natural oil sand reservoir. Immediately following successful completion of this test in the Pontotoc sand, plans were activated for a second field experiment to be conducted in the same sand a short distance away. The objective of this second field test1 as to establish a relationship between field observations and information derived from laboratory experimentation and computations. Secondary objectives of the test were to develop operating tech- niques and define operating problems, and to obtain information that would assist in the economic and engineering appraisals of the process. Careful planning was required since this test was designed with the hope of providing basic information on the characteristics and operation of a combustion drive. It was necessary to include test facilities which would not only permit precise control and measurement of the process variables but would also provide sufficient flexibility of control to insure that variables in the combustion process could be studied during the course of the test under a wide range of conditions. In order to accomplish these objectives elaborate instrumentation and several control wells were included in the design of the test installation. This type installation was felt to be well suited for an experimental study, but was in no respect intended as a counterpart of a commercial-scale operation. PRELIMINARY LABORATORY WORK An important objective of the field experiment, as stated previously, has been to obtain reliable data on operating variables which might be compared with similar data obtained in laboratory combustion tests. The value of successful correlations between quantities that can be measured in the laboratory and those obtained in a full-scale commercial venture has been well established for other industrial processes. The difficulty of following a process underground should make such correlations even more essential here. Also. approaching the field test phase of development of in situ combustion, it was considered desirable to precede design of test equipment with laboratory combustion studies under a variety of conditions to establish likely ranges of variables. These included temperature, air injection rate, frontal advance rate and produced gas composition.

By P. J. Shenon

ENROLLMENT in geological and mining engineering curricula is declining at an accelerated rate despite the greatest need for trained men ever extant in the minerals industry. Industrial and military demand is mounting, but the number of freshmen selecting the mineral field continues to fall. Estimates on the needs of industry range as high as 30,000 new engineers a year. The current deficit is more than 60,000 engineers less than the 350,000 to 450,000 which eventually will be needed. The indisputable fact is that the colleges are turning out fewer and fewer engineers despite the greatest enrollment in colleges and universities ever experienced in the United States. In 1950 a record 52,000 young men stepped out of the confines of ivy covered walls with engineering degrees in their hands. By 1951, however, the number dropped to 41,000 and present enrollment indicates a national graduating class of only 25,000 for 1952. No letup in the drop is forecast. About 19,000 can be looked for in 1953 and 1954 may reach an unhappy 12,000. It becomes clear that something must be done to attract high school graduates to engineering. One immediate possibility could be to make the course burden carried by the engineering student somewhat lighter. The prescribed curriculum in many schools is such that the student takes the path of least resistance, and instead of training for an engineering future, studies for a vocation which will allow him to learn and at the same time get at least a nominal enjoyment out of college life. Review geological and mining curricula of 20 colleges and it will be found that the engineering student is a veritable pack mule compared to a lad taking liberal arts or some other non-technical program of study. The curriculum for geological engineering at one school calls for 202 semester hr, with almost 23 hr carried per semester. Multiply this figure by three hr, the minimum supposedly to be devoted to a credit and you get 69 hr per week. With a bare minimum of 84 hr for sleeping and eating, about two hours a day remain for recreation. However, the load of other schools investigated is about 19 hr. The University of Utah requires 238 quarter hr for graduation with a degree in geological engineering, while requiring only 183 quarter hr for baccalaureate degree from University college, Utah's liberal arts school. It can be stated with a measure of surety that the same proportions exist in other universities. The first step would be for ECPD to review its requirements for mining and geological engineering. It must recognize that mining and geological engineers operate in a specialized field, as do other types of engineers. Although a geological engineer may not design a bridge, as pictured by the ECPD Committee on Engineering Schools, his field of design calls for similar engineering precision, a knowledge of materials, construction methods, economic considerations, and financing. Six schools have been accredited by the ECPD. What is the basis for approval and can the requirements be modified and still be kept in line with the needs of the geological engineer? Course work from school to school varies with the exception of mathematics, chemistry, and physics. Even in those courses the not inconsiderable variation lends dubious creditability to the mean. One accredited school requires 7 1/3 semester hr of chemistry, compared with 24 hr required by another, making an average for the six schools of 17 1 /3 hr. Required credit hr in mechanics ranges from 4 to 18 and in surveying from 2 to 15. Several non-accredited schools require more hr than do the accredited schools in some courses. Why is the engineering student forced to carry such a back-breaking load? The answer is of course fairly obvious. He is irrevocably set apart from the rest of the student body because of the nature of his life's work. He is training for a place in a world where technology is becoming increasingly involved. He must be prepared to do a job now-and not later. Mining and geological engineering require the same essential backgrounds as other engineers, and more. The "more" is a knowledge of mining methods, metallurgy and geology for the mining engineer. The geological engineer must know in addition, mineralogy, petrography, and geophysics. The load is compounded finally by the addition of liberal arts courses. Should anything be done to relieve the situation? Today's engineer must be a whole man, capable of handling the tools of communication and with an understanding of the economics of industry. He must be able to write clear simple English, and he must be man who can think from some other position than bent over a work table. He must be aware of the history of his country and to some extent that of the world. Not all schools share this view. Only two of the accredited schools require history courses. However, five of the non-accredited schools make it mandatory. Four accredited and five of the nonaccredited schools require economics. Courses in mathematics, physics, and chemistry are fundamental in engineer training. The average for the accredited schools could serve as a guide in

Jan 1, 1952

By B. A. Wilcox

ThERE are several reports in the literature which indicate that both solid-solution hydrogen and hydride precipitates can promote low-temperature em-brittlement of tantalum.1-3 For example, Imgram et al.2 showed that hydrogen additions in the range of 140 to 580 ppm promoted a ductile-to-brittle transition at slightly above room temperature in both wrought and recrystallized tantalum. This contrasts with the observation that tantalum of norma1 purity is very ductile, even at liquid-nitrogen temperature. Imgram et al. also found that hydrogen additions caused a deterioration of notch tensile properties of tantalum. The purpose of this communication is to present some limited evidence that hydrogen can also cause fatigue embrittlement of tantalum. The material used in this investigation was wrought, stress-relieved (1000°C, 1 hr), are-melted tantalum rod (5/8-in. diameter) containing the following impurities in weight percent: 0.005 C, 0.007 02, 0.0005 H2, 0.003 N2, 0.08 Cb, 0.01 W, and 0.01 Fe. Hydrogen charging was done in an atmosphere of flowing dry hydrogen for 3 hr at 600°C, followed by an homogenizing anneal in purified argon for 15 min at 600°C, and cooling in an argon atmosphere. Typical micro-structures of the hydrogen-free and the hydrogen-charged materials are shown in Fig. 1(a) and Fig. 1(b), respectively. Fig. 1(b) shows the appearance of hydride precipitates which, in many cases, formed preferentially parallel to the rod axis. The hydride dispersion was uniform across the diameter of the specimen. The fact that the precipitate was a tantalum hydride was tentatively confirmed by X-ray diffraction on a metallographically prepared surface. After hydrogen charging, chemical analyses by vacuum fusion revealed the hydrogen content to be 0.0295 wt pct. In addition, the oxygen and nitrogen contents were slightly raised to 0.0095 O2 and 0.005 N2.

Jan 1, 1964

By Thomas V. Falkie, Robert Stefanko

Recently The Wall Street Journal featured a series of articles titled "The Dirty Work-Brutal, Mindless Labor Remains a Daily Reality for Millions in The US.-Mining Coal, Shoveling Slag, Gutting Hogs Pays Bills But Can Drain Spirit, Hope." The article was devoted to describing how tough and dangerous is mining (and some other jobs) and how rough life is for miners and their families. Similarly during the past few years some psychiatrists and sociologists have devoted increased attention to the attitudes and life styles of miners. Unfortunately most of their approaches have been neither scientific nor objective, and very little good clinical research has been done on the subject.

Jan 1, 1971

By Heribert Wiedemeier, Ali Khan

The tetnperature-conposition phase diagram of the syste,n MnS-CdS has been investigated over a range fro)n 600° to 1000° C. The small degree of solid solution on the MnS rich slde cornpared to the large CdS rich single-plzase region could be esplained by the difference in strzrctures of the cotnponents and the relative sizes of the metal ions. The change in lattice parameter as a function of cot)zposition is linear atzd jollous Vegavd's rule within experittzental limits of error. Mns and CdS have gained considerable attention in recent years because of their interesting solid state properties. In order to investigate such properties and thermodynamic quantities of mixed compositions of transition metal chalcogenides, it is necessary to establish their phase diagrams. Two earlier publications are concerned with the MnS-CdS system. Schnaase' found complete miscibility when MnS and CdS were co-precipitated in aqueous solutions at room temperature. However, there is evidence that the mixed crystals were not stable over the entire range of composition. roger's' investigation of the ternary system ZnS-CdS-MnS at 900" C indicates a solution of 48 mole pct MnS in CdS. During the course of this work which covers a temperature range from 600" to 100O° C, emphasis has been placed on the use of high-purity materials and well-defined modifications of starting products. EXPERIMENTAL Starting Materials. The stable green modification of MnS with NaCl structure was prepared by direct fusion of the elements and by chemical transport reaction. Stoichiometric amounts of manganese of 99.99+ pct purity and of sulfur of 99.999 pct purity from Gallard-Schlesinger Corp. were sealed in an evacuated quartz tube at a pressure of 10"5 mm Hg and annealed for 48 to 60 hr at 800°C. The quartz tubes used throughout this work were outgassed under high vacuum at a temperature of 90OU to 1000° C for 12 to 18 hr. For the synthesis of MnS by chemical transport3 stoichiometric ratios of elemental manganese and sulfur and small amounts of iodine were sealed under vacuum in a quartz tube, which was placed in the temperature gradient, 900/800-C, of a two zone furnace. Single crystals of MnS in form of platelets and octahedra were deposited at the cooler end of the tube. The manganese content was determined by EDTA titratin. A typical result of several analyses was 63.05 pct Mn which agreed well with the computed value, 63.14 pct. Commercially available CdS of 99.9+ pct purity, a mixture of the cubic and hexagonal forms, was converted into the hexagonal high-temperature modification by annealing CdS under vacuum in quartz tubes at 700° C for about 36 hr. As a second method, cadmium sulfide was transported in a temperature gradient, 70O° /650" C, with in vacuo sublimed iodine from the hot to the cooler zone of the tube. Large sized crystals in form of hexagonal columns and some platelets were obtained. An average result of the cadmium analysis by EDTA titration4 was 77.93 pct in good agreement with the computed value, 77.81 pct. Only the NaCl modification of MnS and the wurtzite modification of CdS were used for the solid solutions. Preparation of Solid Solutions. Finely powdered MnS and CdS samples were thoroughly mixed in molar ratios and compressed into pellets of about 6 mm diam and 2 mm in height at a pressure of approximately 13,500 atm.

Jan 1, 1969