Search Documents

By W. W. Vaughn, R. H. Barnett, E. E. Wilson

Soon after the search for uranium ores on the Colorado Plateau began in earnest, thousands of feet of drill core ranging from 1 1/8 to 2 1/8 in. diam became available for study. Although significant advances had been made in the technique of quantitative gamma-ray borehole logging, instrumentation was in the development stage, and complete reliance could not be placed on gamma-ray logs alone to interpret quantitatively the meaning of radioactivity in a drillhole. A method that would be faster than chemical analysis and still give reproducible and reliable results for various drill core sizes was desirable to provide additional information on the enormous footage of drill core being accumulated. A solid phosphor scintillation drill core scanner was designed and constructed. Basically the instrument was developed to measure radiation from a drill core which would not be clearly recorded by a gamma-ray logger using a Geiger tube as the sensitive element. Such data would be beneficial in constructing isorad maps to delineate ore-bearing zones. A calibration in the range 0.01 to 0.1 pct eU3O8 was provided; above 0.1 pct eU3O8 gamma-ray logs were available and were being used to calculate grade and tonnage of ore reserves. The core scanner, however, has been used to estimate equivalent uranium content of ore-grade materials containing as much as 2.2 pct eU308 with an accuracy of ±10 pct, the sample being in the form of a BX drill core. Actually, an apparent calibration of eU308 vs counts per unit time is a straight line with a slope that is a function of the sensitive element and the geometry of the counting assembly. A true calibration that will show the expected departure from a straight line is due principally to the random nature of the pulse from a radiation source and the nonlinearity of the electron circuitry.

Jan 6, 1959

By Henry M. Howe

The data which are collected in Table 1 show that the ferrite of low-carbon steel and of electrolytic iron, like the network of hypo- and hyper-eutectoid carbon steel, inherits, either absolutely or relatively, the grain size of the mother austenite from which it is born in cooling through the transformation range. This evidence is so abundant and concordant as to cast doubt on Prof. Jeffries'l hypothesis of what might be called reversed inheritance, by which he explains Mr. Ruder's observation2 that a grain in a "slug of pressed pure electrolytic iron powder" was coarser, perhaps 1,500,000p2, aftcr a 3-hr. exposure to 1000" C. than after a 3-hr. exposure to 1250-1300°, when the grain size was about 35,000µ2. His explanation implies a natural tendency of coarser austenite grains to yield finer ferrite ones than spring from finer austenite grains, so that the inheritance is reversed. What seems to me a simpler explanation is that the enormous grains which followed the 1000" heating were not inherited from the mother austenite, but were formed by the process of natal coarsening, by means of which Stead and Carpenter developed grains of this same order of magnitude in this same kind of material, on heating to 950" or to 1000°, the temperature used by Mr. Ruder, provided the exposure was not unduly prolonged, and it does not seem to have been in Mr. Ruder's case. That their coarsening was natal and not through inheritance is shown first by the fact that their grain coarseness was out of all proportion to their relatively low temperatures and short periods, and hence was very far in excess of the grain size which the mother austenite must have had, and second and more cogently by the fact that the shape of these coarsened grains was evidently determined by the temperature distribution which existed during and immediately after the Ar3 transformation, that during the passage from the gamma to the non-gamma state. In discussing this subject, we must remember that coarse ferrite grains, whether those formed in cooling from the molten or those induced by

Jan 1, 1918

By A. E. Zawadzki

A progress report is presented on the development of a process for the removal of sulfur dioxide from flue gases. Catalytic oxidation of the sulfur dioxide in flue gases, with the production of recoverable sulfuric acid, has been demonstrated to be technically feasible. Problems associated with the process are discussed. The need for an inexpensive and efficient process for controlling sulfur dioxide emissions from utility plant stacks has been accentuated by the ever-increasing public demand for elimination of all man-made air pollutants. I would like to describe the results of an investigation of the development of such a process; more important, I would like to present a discussion of the problems associated with the process which remain to be solved. The research work on the BCR catalytic gas phase oxidation process has been jointly sponsored by the Association of Edison Illuminating Cos., Edison Electric Inst. and Bituminous Coal Research, Inc., since 1958. However, work on SO2-pollution abatement was initiated by BCR in 1953. Tieman and Clymer,1 working with ozone and simulated flue gases, evaluated a gas phase oxidation process for removing SO2 from flue gases. Their experimental data indicated that the use of ozone to oxidize SO2 in flue gas concentrations, with subsequent recovery of H2SO4, did not appear to be economical due to the large ratio of ozone to sulfuric acid produced (2.6 to 1). The cost of the ozone would be 6 to 10 times the cost of sulfuric acid produced. As a result of the ozone study, work on a catalytic oxidation process was developed. The following reactions are the basis for the catalytic oxidation process Reaction [I] normally occurs in the combustion of sulfur-bearing fuels. At the high temperature found in the combustion chamber, 95% to 99% of the sulfur in the fuel is oxidized to SO2 and 1% to 5% is oxidized to SO3. The concentration of SO2 in the combustion gases depends on the combustion efficiency and the quantity of excess air used. A commonly used rule of thumb is 0.1 percent by volume SO2 in the flue gas for each 1% sulfur in the fuel. Reaction [2] proceeds at a wide range of temperatures. The rate of the reaction is rapid above l,000°F; however, the equilibrium conversion drops rapidly above this temperature. 2 Reaction [3] occurs rapidly to completion at temperatures below 400°F; above this temperature some SO3 exists in equilibrium with H2SO4. Complete dissociation of H2SO4 occurs above 650°F.3 The simplicity of the proposed process provided added incentive to the development and evaluation of the various component unit operations. The advantages of the process were considered to be as follows: 1) no raw materials were required, 2) the unit operations were not radically new; and 3) a salable product, sulfuric acid, would be produced. The successful development of the process was contingent upon the solution of several major problems: 1) The development of a catalyst which would provide high catalytic activity at reasonable conditions of gas flow and temperature without causing an unacceptable increase in pressure drop. 2) The development of a high temperature dust collection system to provide clean flue gases to the catalytic reactor. 3) The complete recovery of the sulfuric acid produced in the process; the effluent gas must contain less than 20 ppm by volume of sulfuric acid. As of this date, the complete solution to these problems has not been achieved. EXPERIMENTAL Catalyst Development: The use of platinum and vanadium catalysts for the production of sulfuric acid by the treatment of dry gases containing 5% to 20% by volume sulfur dioxide has been known for many years. The application of the basic idea of catalytic oxidation of sulfur dioxide in flue gas concentrations was

Jan 1, 1965

By F. W. Bradley

MY first Nevada City mining reminiscence is one of seeing Capt. Thomas Mein, over 52 years ago, in the old Wyoming mill on Deer Creek about a mile below the town of Nevada City. Captain Mein was then superintendent of the Wyoming mine, where the old fashioned Cornish pump was used, as well as hurdy-gurdy water wheels for power. This mill was one of the typical, up-to-date mills of that time and consisted of stamps, fed by Chinese hand feeders, who also cobbed the ore down to a small enough size to be fed into the mortars. Stamps at that time were of light weight and crushed about one ton to one and a half tons to the stamp, the pulp going over plates, then into sluices with self-raising gates. When a sluice became full the pulp was shut off and directed into the parallel sluice, also with a self-raising gate. The concentrate thus collected in the sluice was shoveled out and carried to Cornish buddles and there cleaned up, preparatory to being sent to a local custom chlorination works.

Jan 1, 1929

By F. L. Clerc

THE lead and zinc region of SW. Missouri is interesting, not only by reason of the value of its output, which ranges in the neighborhood of ten million dollars a year, but even more because of the facilities which it offers for the study of certain forms of ore-deposits, of which the loci and genesis are somewhat obscure. Although the mineral species found in the region are few and of common occurrence, crystallization has taken place on a generous scale, and unusual forms abound in many combinations, presenting examples of pseudomorphism, paragenesis, metasomatic replacement and the action of mineral-depositing and mineral-dissolving waters. For the mineralogist, lithologist, and specialist in certain lines of metamorphism, it has peculiar attractions, while the systematic geologist will perhaps find its chief claim to attention in the simplified problem it presents, in the study of ore-deposits. In addition to the phenomena above enumerated, it presents instances of well-developed comb-structure in the ores, fissures and faults, and slickensides-features usually associated with extensive movements in the earth's crust, and believed by some authorities to prove the deep-seated origin of the ores. In connection with these occurrences, the absence of all igneous rocks from this region is noteworthy and significant. For that reason, the simplified case presented by the Joplin region, from which these rocks have been eliminated as a factor, is specially valuable. Having had unusual opportunities to watch the development of this district, I have been deeply interested in the bearing of the revelations of the Joplin mines upon questions of universal importance to the science of ore-deposits ; and it was to my great regret that the project of a meeting of the American Institute of Mining Engineers at Joplin, proposed in 1892, was

Mar 1, 1907

By B. M. Loring, W. H. Baer, G. M. Carlton

It is a well known fact that age-hardening alloys remain in a supersaturated, or partially supersaturated, condition only for limited periods of time at temperatures below the solvus. In order to develop optimum properties by subjecting the quenched alloy to precipitation or age-hardening treatment at a specified temperature, it is necessary that the cooling rate be sufficiently high to maintain the required degree of supersaturation of the solid solution. In practical applications the size and shape of the specimens obviously have considerable influence on the rate of cooling and there is need for more information on the effects of various rates of cooling. Cooling rates of quenched aluminum alloys generally have been established by the type of quenching medium, the temperature of the medium and the dimensions of the specimen. Except for the work of Roth3 little quantitative information is available on the measurement of cooling rates of aluminum alloys and the significance of the influence of various cooling rates upon the mechanical properties. The purpose of the present investigation was to study the use of the Jominy test for aluminum alloys (Table I). The material for this investigation consisted of extruded rods 1 1/4 to 136 in. in diam. The chemical compositions were as follows: The modified L-type Jominy specimen with dimensions shown in Fig I was selected as the most convenient means for obtaining a wide range of cooling rates. The standard Jominy specimen was not satisfactory because its cooling rate was not sufficiently high. The four flat surfaces machined at right angles along the length of the L-type specimen to a depth of 0.015 in. facilitated the hardness measurements. The specimens were machined from the extruded stock with no intermediate forging. New specimens were used for each aging experiment. The specimens were given solution treatments of 2 hr duration at recommended temperatures in a Hoskins furnace (Table 2). The temperature of the furnace was maintained constant to 3°C. No protective atmosphere was necessary because the thin film of oxide that formed on the surface of the specimens had no effect on the hardness measurements. AT

Jan 1, 1949

By Grover J. Holt

In order to provide a clear picture of the development of the sink-and-float process of concentration as applied to the iron ores of Minnesota, a few pertinent facts should be brought out concerning the ores and the products of concentration. Iron ore is primarily a cheap commodity handled in immense quantities. The price of iron ore at present, based on a natural iron content of 51.5 per cent, varies from $2.55 to $2.70 per gross ton f.0.b. cars at the mine or washing plant in Minnesota. Last year 92,000,000 gross tons were shipped from the Lake Superior district (Minnesota, Michigan, and Wisconsin) to the furnaces. Concentrated iron ores have no greater value per unit of iron content than ores that can be mined and shipped direct from the shafts or open pits. For years only the higher grades of concentrating ores could be treated and sold in competition with the immense reserves of direct shipping ore available. These higher-grade "straight wash" ores required only the simplest treatment; namely, the washing out of fine sand from coarse ore particles. As the exhaustion of the " straight wash " ore is reached, rock is found replacing part of this sand, and the ore no longer responds to the former sizing treatment. Substantial tonnages of this marginal (jig ore) remained to be treated. Because of the limitations placed on concentrates by the furnaces—namely, low price of product and minimum amount of fines—iron-ore concentration oil has been and must continue to be a one-pass process. Only an increase in price or a change in structure of concentrates allowed by the ore consumer will encourage departure from this practice. Faced with the necessity of handling large tonnages, with the limitations mentioned, iron-ore research men have found their search for new and more efficient methods of treatment seriously handicapped and removed from the field of fine crushing. Sink-and-float treatment appeared to offer the best field in coarse concentration and has been proved by several plants to be well adapted to our problem. The use of ferrosilicon as a medium solid was decided upon. In our present practice, the ferrosilicon metal, carrying about 83 per cent iron, 15 per cent silicon and 2 per cent carbon and nonmagnetics, is crushed to pass 65 mesh and used without further sizing. When mixed with water the specific gravity of the medium may be controlled to within 0.02 per cent above or below the required gravity. Table I shows examples of a change in solid content of the medium for varying gravities. the I is based on clean ferrosilicon, whereas some contamination is always present in the mill circuit, tending to

Jan 1, 1947

By Grover J. Holt

In order to provide a clear picture of the development of the sink-and-float process of concentration as applied to the iron ores of Minnesota, a few pertinent facts should be brought out concerning the ores and the products of concentration. Iron ore is primarily a cheap commodity handled in immense quantities. The price of iron ore at present, based on a natural iron content of 51.5 per cent, varies from $2.55 to $2.70 per gross ton f.0.b. cars at the mine or washing plant in Minnesota. Last year 92,000,000 gross tons were shipped from the Lake Superior district (Minnesota, Michigan, and Wisconsin) to the furnaces. Concentrated iron ores have no greater value per unit of iron content than ores that can be mined and shipped direct from the shafts or open pits. For years only the higher grades of concentrating ores could be treated and sold in competition with the immense reserves of direct shipping ore available. These higher-grade "straight wash" ores required only the simplest treatment; namely, the washing out of fine sand from coarse ore particles. As the exhaustion of the " straight wash " ore is reached, rock is found replacing part of this sand, and the ore no longer responds to the former sizing treatment. Substantial tonnages of this marginal (jig ore) remained to be treated. Because of the limitations placed on concentrates by the furnaces—namely, low price of product and minimum amount of fines—iron-ore concentration oil has been and must continue to be a one-pass process. Only an increase in price or a change in structure of concentrates allowed by the ore consumer will encourage departure from this practice. Faced with the necessity of handling large tonnages, with the limitations mentioned, iron-ore research men have found their search for new and more efficient methods of treatment seriously handicapped and removed from the field of fine crushing. Sink-and-float treatment appeared to offer the best field in coarse concentration and has been proved by several plants to be well adapted to our problem. The use of ferrosilicon as a medium solid was decided upon. In our present practice, the ferrosilicon metal, carrying about 83 per cent iron, 15 per cent silicon and 2 per cent carbon and nonmagnetics, is crushed to pass 65 mesh and used without further sizing. When mixed with water the specific gravity of the medium may be controlled to within 0.02 per cent above or below the required gravity. Table I shows examples of a change in solid content of the medium for varying gravities. the I is based on clean ferrosilicon, whereas some contamination is always present in the mill circuit, tending to

Jan 1, 1947

By P. Ramachandrarao, T. R. Anantharaman

CONSIDERABLE interest has lately been evinced as regards the nature of chill zones in solidified metals and alloys. Biloni and Chalmersl were the first to show through their study of segregation substructures in some nonferrous alloys that the formation of the dendritic region may be preceded by a diffusionless solidification process, viz., predendritic solidification. Subsequently Bower and Flemings2 tried a special solidification technique wherein the liquid was drawn into a narrow copper mold with polished surfaces and concluded that convection might trigger some multiplication mechanism leading to grain refinement. Using uncoated and partially coated surfaces for molds in the Bower-Flemings technique as well as in conventional molds, Biloni and Morando3 have very recently brought out the importance of the rate of heat extraction on the mechanism of solidification. We report here on the surface structures and substructures observed in a Sn-1 wt pct Pb alloy solidified at even higher cooling rates (riz., l05 to l07 ºc per sec) than those employed by earlier workers. The alloy was made by melting tin (99.99+ pct) and lead (99.9+ pct) in a silica tube under vacuum. Small quantities of the homogenized alloy (0.5 to 1.0 g) were then quenched (or splat-cooled) from the melt by the "gun technique" developed by Duwez et al.4 The thickness of the resultant foils varied from specimen to

Jan 1, 1970

By J. C. Coughlin, M. C. Flemings, T. Z. Kattamis

Experiments on Al-4.5 pct Cu alloy shou the dendrite arnz spacing increases with time during isothermal holding in the liquid-solid region. This coarsening of the dendrite structure is then shown to occur at roughly the same rate during solidification as during isothermal holding. Final dendrite arm spacing in both types of experiments depends primarily on coarsening kinetics. It is concluded that driving force for the coarsening is reduction of liquid-solid sur/ace area. Calculations based on two simple models show qualitative apreement with experinzent. 1 HERE has been considerable interest in recent years in determining effects of solidification variables on dendrite arm spacing in cast alloys. The interest has resulted primarily from the clear evidence that 1) the degree of homogeneity which can be obtained in a cast alloy with commercially practical homogeni- zation treatments depends on dendrite arm spacing, and 2) mechanical properties of the heat-treated casting depend on the homogeneity obtained.'-' In wrought alloys, microsegregation from the original ingot is carried to some extent into the final wrought product, and affects mechanical properties in similar fashion Experimental evidence available indicates that dendrite arm spacing in cast aluminum alloys is influenced only by cooling rate (or "local solidification timev*) and alloy content.Fig. 1, for example,

Jan 1, 1968

By K. Mukherjee

A quasi-thermodynamic calculation is presented for estirt~ating vacancy Parameters in fcc metals. For fcc metals the equilibrium vacancy concentration at the melting point can be approximated by a constant mole fraction. The Gibbs free energy of formation of the vacancies can then be expressed as a simple function of the absolute melting temperature alone. From this description the formation energy Ef, the vibrational entropy Sv, and the formation volume of a vacancy can be calculated in terms of known thermodynamic parameters. The calculated values are in good agree-rnent with available experimental results. IT appears that many of the physical parameters associated with lattice defects can be empirically related to thermodynamic melting parameters.'-7 Besides the implicit correlations between melting phenomenon and point defects, the melting temperature itself at atmospheric pressure can be used as a reference temperature for the purpose of comparison of vacancy concentrations in a given crystal class. The concentration at the melting point then corresponds to the maximum vacancy super saturation in the crystalline state at a fixed pressure. For pure fcc metals such as Au, Cu, Al, Ag, Pb, and Pt the mole fraction of vacancies at the melting point8 is of the order of 10*"3 to 10"1. The equation expressing the equilibrium vacancy concentration at the melting point contains three parameters: the formation energy, the vibrational entropy, and the absolute melting temperature, see Eq. [2]. Since these three parameters cover a wide range of values8 in the above-mentioned metals, it can be assumed that in other fcc metals, for which no reliable data are available, the mole fraction of vacancies at the melting point most likely lies in the range 10~3 to lO"4. This is then the basis of the calculations described in this article. Within this range the mixing entropy can be linearized and the vibrational entropy, the formation energy, and the vacancy formation volume can be calculated in terms of known thermodynamic constants. 1} LINEARIZATION OF MIXING ENTROPY The Helmholtz free energy of a crystal containing n vacancies at the melting point can be writteng as where Ef, T,, k, and S, are the formation energy of a vacancy, the absolute melting temperature, Boltz-mann's constant, and the vacancy formation entropy, respectively. After applying Stirling's approximation in Eq. [l] and minimizing with respect to n where (n/N), is the equilibrium vacancy concentration in mole fraction at the melting temperature at 1 atm pressure. If instead of setting (6~/6n) = 0 immediately after Stirling's approximation in Eq. [I], the configurational partition function is simplified, then some useful and significant results can be obtained by utilizing the experimental inforrnation After applying Stirling's approximation and rearranging is plotted as a function of (n/N) in Fig. 1. The points, not all of which are shown, are fitted by a straight line substituting from Eq. [5] into Eq. [I] and minimizing with respect to n combining Eqs. [6] and [2] It must be emphasized that Eq. [7] is only a crude approximation based on an oversimplified model of

Jan 1, 1967

By H. Conrad

The strain hardening associated with the creep of magnesium single crystals at room temperatu.Je was investigated by shear tests in which the direction of stressing was reversed a number of times after creep in a <100> direction. The strain hardening was strongly anisotropic; the largest amount occurring in the direction of flow and the least 180 deg to this. The time law for creep changed from a parabolic type to logarithmic for the first and subsequent reversals of stress. For a given stress the creep strain decreased significantly for the first and second reversals and then only slightly for subsequent reversals. The logarithmic creep constant a changed in a similar manner. A linear relationship was observed between the creep strain for a given reversal and the stress. The slope for the third reversal yielded a strain hardening coefficient 12 to 16 times greater than that observed for unidirectional flow. Al-though no recovery was observed upon resting at room temperature, the hardening resulting from stress reversals could be completely removed by heating to 450" C. PREVIOUS work1.&apos; indicated that the creep rate of magnesium single crystals tested in the temperature range of 78" to 364°K decreased with time. This was attributed to an increase in internal stress resulting from an accumulation of dislocations at internal obstacles. The object of the present investigation was to further evaluate the nature of this hardening by conducting tests in which the direction of stressing is changed after previous creep in shear in a given slip direction. Previous to 1950 three investigations indicated that a Bauschinger effect3 occurred for single crystals as well as polycrystals brass in tension compression,

Jan 1, 1960

By Harold L. Walker

ENGINEERING educators have recently been discussing the advisability of extending the undergraduate curricula to five or six years, and a plan has also been proposed requiring a preliminary period of study in the liberal arts college before admission to the engineering college. Extension of the undergraduate curricula has been suggested as a means of more fully covering the subject matter in the major field of study. Interest in this proposed plan of requiring a preliminary period of study in the liberal arts college has been heightened through a bill recently introduced in the legislature of the State of New York. This plan intends to give a more professional status to engineering, and credit for these additional years of study would be allowed to those engineers applying for professional licensure. As a result of the interest in engineering curricula and due to the impetus given by these proposed plans a general reexamination of curricula has been conducted by a large number of schools with the stated purpose of determining if the present curricula are meeting the aims and scopes of engineering education.

Jan 1, 1940

By R. C. Wilmoth, R. D. Hill, R. B. Scott

A cost breakdown is given for reclamation and mine drainage abatement conducted by the Federal Water Quality Administration&apos;s demonstration project near Elkins, W. Va. Costs for construction of control measures have been developed and include backfilling strip mines, sealing subsidence areas, reconstructing stream channels, and installing mine seals to prevent the entrance of water and air.

Jan 1, 1973

By AIME AIME

EARLE EDWARD SCHUMACHER, the new Chairman of the Institute of Metal, Division. is well known to the metallurgical profession. His election a, Chairman is the culmination of fifteen years&apos; service to the Institute, beginning in 1930 when he was elected Member. Since that time he has served as member of the I.M.D. Data Sheet Committee. 1935-40; Papers and Publication Committee. 1936; Executive Committee. 1937-1910: Papers and Programs Committee, 1937-1938. (Vice-Chairman, 1939-40, Chairman. 1941-12): Annual Award Committee. 194142. (Vice-Chairman, 19431: Vice-Chairman, LM.D., 1943-44; Chairman Publication Committee, 1943-44; member Powder Metallurgy Committee), 1944. Pew other men have contributed so Much time anti effort in so many way- to the interests of the metals branch o4 the A.I.M.E.

Jan 1, 1945

By C. Scala, M. Gondouin

Published laboratory data have established that very significant streaming potentials can exist across mud cakes subjected to pressure differentials such as exist between a mud column and formation fluids. Since shale is similar in kind and in properties to a well-formed mud cake, it appeared possible that streaming potentials might also arise across shales. A practical consequence would be that the streaming potential contribution to the SP would be smaller than might otherwise be assumed since the SP is measured with respect to the "shale bare line". The resultant contribution would be the difference between the streaming potential across the shale and that across the mud cake. In order to investigate this possibility, a laboratory apparatus was devised for measuring streaming potentials under pressure conditions normally encountered in wells. Reported data on streaming potentids across mud cakes were verified. More important, appreciable streaming potentials were found across shale samples. As in the case of mud cakes, the shale potentials varied with the concentration and nature of the impregnating electrolyte. Dependence on pressure was In a different rela-tion than that for mud cakes, possibly because of the lower compressibility of shales. The differences in streaming potentials between the shales md mud cakes were generally small, although they might vary considerably in practice depending on the relative electro-kinetic properties of the particular shale-mud cuke combination. A "poorly eflective" shale with a "very eflec-tive mud cake" would produce the largest negative addition to the electrochemical component of the SP. Results of field experiments with pressurized wells were considered in the light of the laboratory results. INTRODUCTION Field experiments wherein the pressure of the mud column in a borehole was varied have often indicated the presence of a streaming potential, but its magnitude was not easily predictable and its sign sometimes changed from one formation to another in the same well.&apos; In the interpretation of SP logs from standard surveys, nevertheless, the correct value of R,, is very frequently obtained, within the limits of field accuracy, using only the simple electrochemical expression of the SP provided that proper corrections are made for bed thickness, ion activity, etc.&apos; Accordingly, it was apparent that the streaming potential component of the SP in a borehole was considerably more complex than commonly assumed and clarification was indeed necessary The close similarity in nature and constitution of a shale to a well-compacted mud cake suggested that a streaming potential might also exist across shales as pointed out by M. P. Tixier in 1953 and as recently emphasized by Wyllie." e practical consequence would be that the streaming potential across the mud cake would be canceled, at least in part, on electrical logs since the SP is recorded with respect to the shale potential as a base line. This appeared to be a plausible explanation for many of the effects observed but no supporting data were available. In 1955 preliminary experiments at low pressures were conducted in our laboratories which supported the hypothesis of the shale streaming potential and formed the basis for later studies at borehole pressures.* EARLY LOW PRESSURE EXPERIMENTS The early experiments were conducted primarily to determine the possibility of a streaming potential across shales. The shale used was a compact type designated as No. 7 from Reagan County, Tex. (Permian). Cut samples were saturated for over one month in .2 to 2.0-N NaCl solutions. In order to eliminate all electrode polarization effects, a dynamic method was used. A transient pressure of about 1 atm was applied across

H. H. Kellogg (Columbia University)—The accurate measurements of the equilibrium gas ratios for the reaction: Pb(1) + H2S = PbS(c) + H2 [I] reported in this paper are a distinct contribution to the literature, and the authors are to be congratulated. I wish to comment, however, on the correlation of the data which is presented. The authors show a straight line correlation of log H2S/H2 vs 1/T, Fig. 1, and are satisfied with this for calculation of ?Fo even when extrapolated to low temperature (427 °C). Heat capacity data for the constituents of this reaction are available (at least to 900°K), however, and ACP for reaction [I] is finite and positive in the temperature range studied. Simplification of the heat capacity equations yields: ?CP= 2.71 + 1.65 x 10-3T (298°-900°K) This positive value of ?CP should result in a small but noticeable curvature of a plot of log K vs 1/T. That the experimental data fail to show such a curvature may be traced to another oversimplification of the author&apos;s treatment of their data. They assume that the activity of liquid lead is unity for all temperatures, so that

Jan 1, 1961

By Joseph E. Worthington

The present-day Northumberland gold mine is one of the deposits generally characterized as a Carlin-type occurrence. It lies at the crest of the Toquima Range in Nye County near the center of Nevada. Gold mineralization occurs in two modes, in argillaceous and in silicified limestones, and is generally very fine-grained or micron gold. The Northumberland district has had a history that is typical of many western gold districts: minor production prior to World War II and intermittent exploration thereafter until a combination of geological insight, improved economics, and the advent of heap leach technology created the Northumberland gold mine. Nye County, Nevada, was sparsely populated and little explored in the early years of the settlement of the west; the Northumberland district was not established until 1866. Initial interest was in silver and the district operated as a very small producer for the next seventy years. The disseminated gold occurrences in silicified limestones were recognized and Northumberland Mining Co. was organized to develop the property in the late 1930s. Northumberland Mining Co. actually conducted drilling operations (over 200 drill holes) and mined from small open pits in the silicified limestones. They ultimately produced almost 936 kg (33,000 oz) gold before being shut down by War Production Board Order L-208 in 1942. After World War II gold mining activities were essentially nil for over a decade due to the poor economics of gold production. The property was, however, a known gold producer and attracted recurrent exploration attention. About 45 holes were drilled under the direction of Peter Joralemon for private interests between 1959 and 1963. Next Kerr McGee drilled about 25 holes during 1963 and 1964. Some- what later, in 1968, Homestake drilled 20 holes. The property was then acquired by Idaho Mining Co. which drilled about 30 more holes between 1972 and 1974. By this time the Northumberland mine was becoming somewhat shopworn with over 300 holes drilled. Interest in gold prospects was increasing substantially in Nevada, how- ever, due to rising gold prices in late 1974, and several companies were interested in continuing exploration at Northumberland. In 1975 Cyprus Mines Corp. was successful in obtaining a joint venture arrangement with Idaho Mining Co. for further exploration and development of the property. The overall Cyprus exploration program was under the direction of James G. Hansen, Vice President Exploration. The geologist recommending acquisition of Northumberland was Peter E. Chapman who reported to Joseph E. Worthington, Manager of U.S. Exploration for Cyprus. The basis for selection of the Northumberland mine as an exploration target for Cyprus by Chapman was essentially prior knowledge of regional and 16cal geology and of the mine. Exploration for the next few years was directed by Chapman under the supervision of Worthington. During 1975 and 1976 rotary and check core drilling were conducted that indicated that a substantial Carlin-type or disseminated, low-grade gold deposit occurred in two separate bodies. Drilling was based on geologic mapping and rock chip geochemical sampling. Both ore zones were reflected at the surface as gold and arsenic anomalies in rock chips. Heap leach tests attempted in 1977 were aborted by a flash flood, but were completed in 1978. Engineering studies occupied the next couple of years until the property achieved production in the fall of 1981. It is now producing by open-pit mining with gold recovery by heap leaching and cyanide extraction at a rated capacity of approximately 2722 t/d (3000 stpd) ore. Metal recovery has been projected (probably conservatively) at 5 10 kg/a (18,000 oz per year) gold and 1.6 Mg/a (59,000 oz per year) silver. Reserves are reported to be adequate for ten to fifteen years of production. REFERENCES Anon., 1981, "Gold in Nevada," Span Magazine, Vol. 21, No. 3, Standard Oil Co., pp. 6-9. Koschman, A.H. and Bergendahl, M.H., 1968, "Principal Gold- producing Districts of the United States, Professional Paper 610, US Geological Survey, p. 193. Kral, V.E., 195 1, "Minerals Resources of Nye County, Nevada, Bulletin, Vol. 45, No. 3, Geology and Mining Series 50, Nevada University.

Jan 1, 1985

By J. D. Allen

In the coal preparation plant at U.S. Steel Corp.&apos;s Concord mine in Alabama, the product moisture was reduced from about 20% to 9% through the development of a workable flowsheet for dying the clean coal. Some of the high points of the conversion of the Concord washer from vibrating screen dewatering to centrifuge drying are presented in this article. The coal preparation plant at the United States Steel Corp.&apos;s Concord mine near Bessemer, Ala., began operations in June, 1950. The basic cleaning circuit consisted of a Chance cone for cleaning the 2-in. x 3/8-in. fraction and concentrating tables for cleaning the -3/8-in. coal fraction. Dewatering of the final product was accomplished with 35-mesh vibrating screens. However, because Concord coal is very friable with 85% passing 3 mesh and 25% passing 28 mesh (Fig. I), a considerable quantity of fines passed through the dewatering screens. To recover these fines, the dewatering screen effluent was sent to two 70-ft-diam hyd rose para tors from which the underflow returned to the dewatering screens and the overflow went to a 175-ft-diam thickener for water clarification, as illustrated in Fig. 2. As might be expected, there was considerable recirculation of the -35-mesh coal in this system. This degraded the coal and led to higher moistures. It also resulted in a fines bleed-off of plant water and thickener underflow to a fine-coal settling pond. In 1961, vacuum filters were installed to supplement the screens and to recover the fine coal lost in the bleed-off. The above system resulted in a product moisture of about 20%. Further reduction to approximately 11% was accomplished by drainage from railroad cars during transit to the coke plant. Control of the moisture at unloading was poor because of variations in operating levels and railroad car availability, but since the Concord coal constituted a relatively small percentage of the over-all coal mix at the coke plant, its high moisture and fluctuations did not seriously hamper coking operations. Early in 1964, however, the proportion of Concord coal was increased with a resulting increase in the moisture of the total coal mix, and this caused difficulties in the coke plant. The increased participation of Concord coal also brought about a shortage of railroad cars which raised the moisture content to a point where the coal became almost impossible to handle. In an effort to alleviate this problem, U.S. Steel initiated a research program to develop a flowsheet for reducing the moisture to approximately 10% and to determine whether the resulting product would be satisfactory for the coke plant. Under this program, an experimental flowsheet for evaluating both a vibratory centrifuge system and a solid bowl centrifuge system was drawn up as shown in Fig. 3. A horizontal vibratory centrifuge was installed to reduce the moisture in the dewatered coarse coal, and a four-cell, flotation machine was installed to clean the -65-mesh portion of the 3/8-in. x 0 table raw coal feed. Several shortcomings in the above circuits were apparent. For one thing, large quantities of repulping water were required. Furthermore, since about half of the flotation feed was of a size consist that could have been cleaned on the tables, these circuits would result in an unnecessary increase in reagent costs and facilities. Maintenance of the hydro separators and pumps for pumping the underflow to the filters,

Jan 1, 1969

By Jules E. Jenkins

IN the past quarter century the seismograph has played an increasingly important role in evaluating vibratory effects transmitted to adjacent communities by industrial blasting operations. In this period research has advanced through two major phases—from falling pin seismometers to three-component recording seismographs. The initial phase of this research is now regarded as the era of recording, during which the seismograph merely recorded occasional events. It has emerged from this stage into the era of control and analysis and in this later period has been recognized as a full-fledged industrial tool. Although the falling pin machines of earlier years indicated probable amplitudes of industrial blasting vibrations, they gave no reliable measure of such displacements, nor did they provide essential information concerning direction or frequency of motion. The first contribution of the three-component recording seismograph was the weight-for-distance formula developed by the U. S. Bureau of Mines. This empirical formula enabled users of commercial explosives to determine beforehand the probable vibratory effects of a blast when the weight of the explosives charge and the distance from the shot point were known. This formula has withstood the test of time and is still applicable for blasts detonated by instantaneous methods. The second major contribution of the three-component recording seismograph followed the introduction in 1945 of short period or millisecond delay detonation. This might be described as a two-stage contribution, the first of which revealed a 50 pct reduction in the vibration magnitude when millisecond detonation was used rather than instantaneous discharge. This second part of this double contribution led to the discovery of the vibration dip phenomena, see Fig. 1. In Fig. 1 the horizontal coordinates indicate the displacements amplitude in inches, and the vertical coordinates, the explosive weight in pounds. The top curve is that of the U. S. Bureau of Mines for instantaneous detonation. Its curvature conforms to the two-thirds power of the explosive weight as provided in the weight-for-distance formula; hence the vibratory effects increase in proportion to an increase in the explosive charge. If the other curves shown below are temporarily disregarded, it may be considered that the entire area below this top curve represents both useful and waste energy. As mentioned previously, when millisecond delay detonation was first investigated it was found that in general the U. S. Bureau of Mines curve could be decreased by about 50 pct, or to the position of the middle curve. At the outset it was assumed that this curve would likewise increase in some proportion to the explosive charge, as did the top curve, that is, the curve would indicate a continuous upswing. As seismographic data accumulated this assumption was found to be erroneous in part. It was discovered that

Jan 1, 1957