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By Martin Coryell

Ehrenberg was the first to publish and direct attention to the peculiarities of the diatomaceous deposit of Richmond; but how and by whom he was supplied with the material for his microscopical investigations, is not generally known; but we may presume W. B. Rogers, formerly State Geologist of Virginia, furnished the material, and may have given the geological position or horizon. OF these facts we have no available evidence, nor have we any estimate of the thickness of this stratum of diatomaceous or infusorial earth. At the request of a member of this Institute, I have investigated the subject, and offer the following facts as the result, derived principally from information supplied by Mr. John Ott, of the Southern Fertilizing Company, Dr. William Taylor, State Chemist, C. Peticolas, microscopist, Captain Robert H. Temple, civil engineer, Major C. M. Bolton and Mr. C. M. Smith, civil engineers, and formerly engaged in constructing the tunnel under Church Hill, in Richmond. The city of Richmond is generally described as being built on seven hills. This is not the fact, geologically speaking. Originally, there was an immense plateau of land, about two hundred feet above tide, in which a deep excavation was made from the west to east by the James River cutting down to the granite. Opening southward and meeting the valley of the James are various ravines and gullies, in the principal of which are Shockoe and Gillies Creeks. The formation is tertiary, subdivided into all the classified divisions, but at the present moment no one has been found to make this section a special study. Rogers's early classifications and diagrams, with all their imperfections, are the only published records, and the printed editions are long since exhausted. The city of Richmond is really situated in these ravines of the James, Shockoe, Gillies, and their connections. The profile marked A, Plate V, represents the section of the plateau

By Rodney P. Smith

The diffusivity of carbon (0.1 wt pct C) in Fe-Nz alloys (0 to 100 pct Ni) has been determined for the temperature range 860° to 1100°C. As a function of nickel content, the diffusivity has a maximum near 60 pct Ni (the maximum diffusivity being about 1.3 times that in the absence of nickel); the activation energy has a maximum between 40 and 50 pct Ni and a maximum between 80 and 90 pct Ni. The difference between the minimum activation energy and that in iron is about 3000 cal pev g-atom; Do has a minimum between 40 and 50 pct Ni and a maximum between 80 and 90 pct Ni. The results cannot be rationalized by an approximate thermodynamic treatment. THE diffusivity of carbon has been determined in a number of iron alloys over a limited concentration range. It seemed desirable to investigate a system which allows an extended range of alloy composition within a single-phase region. The Fe-Ni system is ideal in this respect, in that all alloys from 100 pct Fe to 100 pct Ni are fee in a convenient temperature range.&apos; The carbon diffusivity was determined by a decar-burization method. The experimental procedure was identical with that used to determine the diffusivity of carbon in y Fe-Co alloys.2 The experimental data are given in Table I. A small correction (order of a few percent) has been made to the measured carbon loss to correct for the carbon lost from the ends of the cylinders.&apos; Since the diffusivity of carbon varies with carbon content the measured diffusivity is an average value for a carbon content between zero (surface) and that at the center of the sample at the end of the decarburization periods. In making the correction in D to 0.1 wt pct C it is assumed that the measured D corresponds to the arithmetical mean of the carbon content at the surface and at the center of the sample at the end of the decarburization period.3 Since this correction is small (<4 pct in D) and since for our decarburization times the changes in carbon content at the center of the sample was small the mean carbon content could have been taken as half the initial value. It is further assumed that the change in D with carbon content for the alloys is the same as that for the diffusion of carbon in iron. From the data of Wells, Batz, and Mehl4 and of smith5 the correction of D from the mean carbon content to 0.1 wt pct C is 0.3 (0.1 - mean wt pct C). The results for iron are given in Ref. 2. Within the experimental error log Do.l%C for each alloy is a linear function of 1/T; the constants for the equation determined by the method of least squares are given in Table I. The deviations of the experimental points from the least-squares line are of the order of 2 pct in D. A comparison of our results for the diffusivity of carbon in nickel with those of other investigations is shown in Fig. 1. The lower curve in Fig. 1 is a linear extrapolation of values calculated* from the equation of Diamond6 for the relaxation time (temperature range 100° to 500°C). The results indicate a small increase in the activation energy over the temperature range 100° to 1400°C; however, it is difficult to say whether the change in Q is real or experimental error. Certainly the change in Q is less than the variation of 5 kcal per g-atom in the diffusivity of carbon in a iron.6 The experimental data for all the alloys are plotted in Fig. 2. As a function of nickel content the diffusivity has a maximum near 60 wt pct Ni at all temperatures investigated and possibly a minimum between 80 and 90 wt pct Ni for temperatures below 1000°C. The activation energy, Q, and log Do are plotted as a function of the nickel content in Fig. 3. Due to the limited temperature range of our experiments neither Q nor Do can be determined precisely; the activation energies appear to be consistent to ±0.3 kcal per g-atom; however the deviation from the absolute values may be considerably larger, see Table II. The Do values probably have little significance. The solid line for Do in Fig. 3 represents the values required to reproduce the experimental values for D when Q has values represented by the upper solid line The diffusivity of carbon may be expressed in terms of the mobility B22, the activity coefficient r2,

Jan 1, 1967

By John G. Tate, George W. Nicolson, James L. Bruce

For more than 25 years modern mining has been carried on in the Island of Cyprus, Mediterranean Sea, by the Cyprus Mines Corp. of Los Angeles, Calif. The general features of these operations have been described.&apos; From 1935 to June 30, 1940, top slicing in the Mavrovouni mine produced in excess of 2,000,000 long tons. The rate for the last part of this period was approximately 50,000 tons per month. Description of Ore Bodies The shape Of the Ore body now being top sliced in the Mavrovouni mine is an inclined irregular pyritic pipe or chimney of large dimensions) extending from unknown depths below the bottom level at elevation 250 ft below sea level) to the highest level at elevation 400 ft above sea level. The present surface above the ore body lies between 550 and 650 ft elevation. Undoubtedly the Ore body extended to surface and was Mined by the ancients by opencast and underground methods. The old pits were gradually filled up by erosion until now they show as slight depressions in the topography. The largest area of the Ore body is at elevation 170 ft below sea level, where the horizontal section is almost elliptical, with a width of 600 ft and a length of 1000 ft. Below this elevation the shape of the ore body continues to be elliptical but with reduced dimensions. Above this elevation it rises flatly towards the south with a substantial depression in the upper-side of the chimney and irregular ridges or wings rising on each side. It gradually grows smaller as it rises. The ore body is a pyritic replacement of andesitic extrusives which occur as flows and pillow lavas. With the exception of the upper portions of one or two small offshoots rising from the main ore body, all ore is completely surrounded by andesite. Most of the andesite lying above the ore body is covered by marl and Clay sedimentaries of Miocene and later age, The Miocene is presumed to be older than the ore mineralization. That portion of the andesites which lies between the sedimentaries and the ore body has been broken by slumpage, and weathered, and iron-Stained by the descending meteoric waters, It is a mass of crushed material that caves and follows downward readily as the top-slice stopes retreat. The main ore body is a heavy cupreous pyrites mass assaying on the average about 4.0 pet Cu; 48.0, S; and 42.0, Fe. In places the copper drops below 1.0 pet. on the borders of this heavy sulphide ore body lie disseminated cupreous pyrites ore deposits of large dimensions. some portions of these appear to be closely related to the heavy sulphide ore, as the ratio of copper to pyrite is about the same. The greater part of the disseminated ore, most of which lies immediately to the east of the main Ore body, contains less than 0.10 pct Cu and about 25.0 to 30.0 pct S. &apos;This probably can

Jan 1, 1949

R. Schuhmann, Jr. (Purdue University)— Fulton and Chipman&apos;s results on rate of silica reduction from slags are analogous in many was to the results of Parlee, Seagle, and Schuhmann10 on rate of alumina reduction from alumina crucibles. Both investigations have given comparably low rates of reduction and slow approaches to equilibrium. Accordingly, we may hypothesize that the rate-determining step is the same in both kinds of experiments; that is, oxygen diffusion across the stagnant boundary layer on the liquid-metal side of the interface between the liquid metal and the oxide phase (slag or solid oxide). I suggest that silica reduction involves the following consecutive steps: I) At the slag-metal interface: SiO2(slag) Si+ 20 II) Transport of oxygen from slag-metal to gas-metal interface: a) diffusion across liquid-metal boundary layer at slag-metal interface. b) convection within the body of liquid metal. c) diffusion across boundary layer at metal-gas interface. 111) At the metal-gas interface: C +O- CO (gas) Iv) At the graphite-metal interface: C (graphite) -C At steelmaking temperatures it is reasonable to assume that equilibrium is attained in all three chemical reactions (I, 111, and IV) right at the respective interfaces. Convection within the stirred liquid metal (step IIb) is also rapid. Transport of Si and C should be orders of magnitude easier than transport of 0, because of the relatively high concentrations of Si and C. Accordingly, we might expect the overall reaction rate to be determined by boundary-layer diffusion of oxygen, either IIa or IIc. Fulton and Chipman&apos;s demonstration that bubbling CO through the system had no major effect on reaction rate indicates that IIc is not the slowest step. Therefore, it becomes logical to estimate the maximum rate for step IIa and to compare this theoretical estimate with Fulton and Chipman&apos;s experimental data. If oxygen diffusion across the liquid metal boundary layer at the slag metal interface (step IIa) is rate-determining, In this equation, dn sio, /dt is the rate of silica reduction in moles per sec,A is the area of slag-metal interface in sq cm, Do is the diffusivity of oxygen in sq cm per sec, 6, is the boundary layer thickness in cm, c,* is the oxygen concentration right at the slag-metal interface in moles per cubic cm, and co is the oxygen concentration in the body of the liquid metal, also in moles per cubic cm. Equilibrium data" on the silicon deoxidation reaction in liquid iron and steel at 1600°C indicate that the oxygen contents of the liquid metal in Fulton and Chipman&apos;s experiments at 1600°C probably fell in the range of 0.5 x10-3 x10-3wt pct. That is, the maximum conceivable value of co*-co for the system under consideration was on the order of 10"5 mole oxygen per cubic cm. On the basis of previously published data,1O,11 it is estimated that Do/0 will fall somewhere in the range from 10-3 to 10-1 cm per sec. The surface area A in Fulton and Chipman&apos;s experiments was approximately 20 sq cm, and the weight of metal involved was about 500 grams. Combination of all these figures with the above rate equation leads to an estimate that the rate of silica reduction should fall within the range from 0.002 to 0.2 wt pct Si per hr. This estimate is consistent with the experimental data. For example, Fulton and Chipman&apos;s Fig. 2 shows a change of about 0.3 pct Si in 10 hr, corresponding to an average rate of 0.03 pct per hr. According to the proposed hypothesis, increasing the temperature will increase the reaction rate ill two ways: 1) by increasing oxygen diffusivity and 2) by increasing the oxygen concentration (oxygen solubility) in the liquid metal. The combination of these two effects accounts for the high value of the observed activation energy. Summarizing, I propose that the rate of silica reduction, like that of the carbon-oxygen reaction, is diffusion controlled and that low oxygen concentration in the liquid metal is the major factor accounting for the very low observed rates of silica reduction. John Chipman (author&apos;s reply)—The authors thank Professor Schuhmann for his interesting contribution. His proposed explanation may well prove to be the correct one. There is clearly a need for much further experimental work on this problem, and further research is in progress.

Jan 1, 1961

By T. C. Campbell

Copper is extracted from the Chuquicamata oxide ore by a hydro-metallurgical process. The ore is crushed to 3/8-in. sizing, and leached with a sulfuric acid electrolyte. Chlorine is precipitated. and the ferric iron reduced in the enriched electrolyte, after which the copper is recovered by electrolysis with insoluble Chilex and lead-antimony anodes, the spent electrolyte being returned to leaching. Cathodes are melted and refined in market furnaces and cast into commercial wire-bar and cake shapes. Sulfuric acid for the process is supplied by the brochantite in the ore. Water for washing the ore is advanced through the solution system and finally after cutting by electrolysis to from 6 to 16.0 grams per liter is completely stripped by the cuprous chloride method and run to waste. The cuprous chloride so obtained plus that resulting from precipitation of chlorine from strong solution is dissolved in ferrous chloride brine and the copper cemented on scrap iron. The cement so obtained is in part used to reduce ferric iron in electrolyte, and as it is of exceptional purity is also furnace refined and cast into an exceptional quality of fire-refined copper. Fig. 1 is flow sheet of the operation. The reduction plant is composed of seven divisions as follows: (1) crushing, (2) leaching, (3) tailings disposal, (4) dechloridizing, (5) sulfur dioxide treatment, (6) electrolytic tank house, (7) smelting and melting. A brief description of each of these divisions will be given. Crushing Ore that is frequently blocky and up to 5 ft. in section is crushed in four stages to 3/8-in. mesh. The first stage of crushing is called primary and the next three stages are known as secondary crushing. The first stage is accomplished with two 60-in. Superior McCully gyratory breakers by which the ore is broken to 9 in. The second coarse gyratory breakers crush the ore to 3-in. sizing. There are seven No. 10 McCully gyratories in this service, for one of which recently has been substituted a No. 7 cone crusher. The third and fourth stages of crushing are accomplished by fifty 48-in. disk crushers. In the third stage 14 disk crushers break the ore to 1 in., and in the last stage 36 disk crushers turn out the final product of 0.371-in. sizing. Reference to the flow sheets of Figs. 2 and 3

Jan 1, 1934

By C. A. Herbert, C. M. Young

WithIn the past few years, considerable development has been made in the coal-mining industry in Illinois and Indiana and it is the purpose of the authors to record its most important phases. Perhaps the two most striking features are the entry into the producing fields of certain large consumers of coal and the magnitude of some of the new operations. Mines are now being equipped by the Chicago & Northwestern R. R., the Chicago, Burlington & Quincy R. R., the Standard Oil Co., of Indiana, and the Union Electric Light & Power Co. of St. Louis. Besides, the U. S. Steel Corpn. has increased its coal-mining activities by commencing operations on its large holdings northeast of Benton, Franklin CO., Ill. Until recent years, the supply of coal has commonly been adequate and prices have generally been favorable, especially to large consumers, as the magnitude of their purchases enabled them to obtain satisfactory quotations from producers. In fact, most consumers of coal have believed that they could buy their fuel more satisfactorily in the open market than they could produce it themselves. The entry of the United States into the war was accompanied by various disarrangements of industrial conditions, and a demand for coal in excess of the supply seemed likely to be experienced for an indefinite period. Under these circumstances it seemed that the greatest assurance of a supply of fuel was its production by the operation of mines. In the case of the Chicago & Northwestern R. R. and the U. S. Steel Corpn. the production of coal has been carried on for many years with satisfactory results, but the C. B. & Q. R. R., the Union Electric Light & Power Co. and the Standard Oil Co. of Indiana had not been producers of coal, at least in this district. Apparently the most impressive feature of the newest mines is their capacity, for in some cases it is planned that production shall be in the neighborhood of 1000 tons per hr. There are mines now in this field having nearly that output, but these were not planned for such output; in fact, when the shafts were sunk and the machinery installed it was generally supposed that the daily outputs would be only in the neighborhood of 4000 to 5000 tons. Incidentally, there has risen

Jan 1, 1920

By D. L. Katz, M. J. Rzasa

The derivations of three methods of computing the static pressure gradients in natural gas wells have been presented to show the assumptions made. Charts were developed from which the pressure gradients may be read when the well-head pressure, the well fluid gravity, depth, and the average well temperature are given. A chart for estimating the well fluid gravity from the condensate content and separator gas gravity is included. The effect of the increased average well temperature after flow on the calculation of the static pressure gradient is discussed. Introduction Reservoir pressures have been calculated from well-head pressures for gas wells for many years. As the pressure measurements become more accurate, the need for a reliable calculation of the static pressure gradient often arises. This paper will develop the several methods for computing the pressure gradient in gas wells and make a comparison between them. The method of calculating static pressure gradients in common use is Eq. I (ref. I) or its counterpart which includes a factor for the deviation of the gas from the ideal gas law. P2 - P1 = P1(e0.0000347GX _ I) Eq. [I] in which P1 = pressure at well head, lb. per sq. in. abs. PP = pressure at bottom of well, Ib. per sq. in. abs. G = gas gravity X = depth of well, ft. An alternate method is to compute the average density of the gas in the well and multiply by the well depth in a manner similar to that used for liquid gradients. Derivation of Formulas A static pressure gradient is a special case of the general fluid-flow equation. Consider a pound of fluid flowing in a vertical column from point I to point 2, Fig. I. By an energy balance for fluid flowing from I to 2: U1 + P1V1 + u12/2g + X1 + q = U2 + P2V2+ u22/2g+X2 +W [2] in which U = internal energy, ft-lb. per lb. P = pressure, lb. per sq. ft. V = volume, cu. ft. per lb. u = average linear velocity, ft. per sec. X = height above datum, fl. q = heat absorbed by fluid, ft-Ib. pcr lb. W = work done by system, ft-lb. per lb. An energy balance on the fluid itself gives: U = ?TdS - ?PdV + etc. [3] in which T = temperature, deg- R- S = entropy, ft-lb. per deg. R. per lb.

Jan 1, 1945

By Cyril Wells, Robert F. Mehl

THE practical importance of a knowledge of the rates of diffusion of carbon and of alloying elements in steel has often been pointed out 1,13 This importance ties chiefly, though by no means only, in the study of carburizing and in the study of reactions involved in the eutectoid inversion in steel The present paper is one of a series in which the rates of diffusion of carbon and of alloying elements in y-Fe and their mutual effects are measured with precision The present paper reports the results of meas¬urements of diffusion coefficients D of manganese in -y-Fe over wide ranges of temperature, and concentration, and of the effect of carbon and impurities and of grain size upon these diffusion coefficients [$] Data on the rate of diffusion of manganese in y-Fe have been reported by Fry2 and by Paschke and Hauttmann 3 Claussen4 calculated the diffusion coefficient D from Fry&apos;s data as 3 X 10-10 sq cm per sec at 960°, which will be shown to be incorrect by many orders of magnitude Paschke and Hauttmann gave a rough value of 9 6 X 10-8 sq cm per sec at 1400° Neither of these workers considered the variation of D with concentration and with temperature, and neither studied the effect of carbon METHOD AND MATERIALS Method -The method employed is identical with that used earlier 5 Owing to the volatility of manganese and the consequent loss of manganese from the outer surfaces of the welded diffusion bars, these surfaces were discarded The diffusion anneals were performed in argon, except one (test 18) in which hydrogen was employed for purposes noted below The temperatures reported are accurate to about 3° and constant within I°C , the diffusion times are reported to the nearest half hour The specimens were furnace-cooled after the diffusion anneal at about 2° per min Layers, usually 0 004 in thick, parallel to the welded interface, were machined from the specimens after the diffusion anneal and analyzed for manganese to 0 01 per cent Mn, the distances of these layers from the interface was measured to ±0 0002 in The concentration-penetration curves obtained were analyzed by both the Grube and the Matano solutions e The D values calculated by these two methods do not differ appreciably 5 The Grube solution was used except in tests in which the effect of concentration upon D was greater than 50 per cent, in such cases the Matano method, which explicitly provides for the variation of D with concentration,1 is the more reliable In general, the number of layers machined from the zones of variable composition were large enough to permit the use of the average layer composition as that of the center of the layer In 3 of the 30 tests, however, this is not justified, and the D values

Jan 1, 1941

By Hans Lundberg

Jan 1, 1927

By Chester A. Rowland

INTRODUCTION Comminution is generally a feed preparation step for subsequent processing stages; exceptions being when a final product such as aggregates, specification sand, Portland Cement, and similar products is produced. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding -- primary and secondary), natural or manufactured non-metallic media (pebble milling) or manufactured metallic media -- steel rods, steel or iron balls. This chapter will discuss general mill design and the specific design and application of the following types of tumbling grinding mills. [Overflow Rod Mills Figure 1 Peripheral Discharge Rod Mills Figures 2 and 3 Compartment Mills Rod and Ball Figure 6 Ball Figure 6a Pebble Mill Figure 6b Overflow Ball Mills Figure 8 Diaphragm (Grate Discharge) Ball Mills Figure 9] I. GENERAL MILL DESIGN A. Liners The interior surface of grinding mills exposed to grinding media and/or the material being ground are protected from wear and corrosion by rubber, metallic, a combination of rubber and metallic, or non-metallic wear resistant materials. B. Drives Economics at the time of plant design and mill purchase determine the drive to be used. The simplest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinion-shaft by either an air clutch or flexible coupling.

Jan 1, 1982

By John Mullen

Impingers TWO types of midget impinger apparatus are used at Climax. The Mine Safety Appliances Company&apos;s midget impinger (MSA) and the Climax Molybdenum Company&apos;s midget impinger (CMC) (Fig. I). The Leiman Brother&apos;s No. 25, positive displacement pump is used in the CMC impinger instead of the four-cylinder radial pump that is used in the MSA impinger. With the MSA apparatus it is possible to obtain the proper vacuum by adjusting the needle valve on the vacuum chamber, while with the CMC impinger the proper vacuum is obtained by varying the rate at which the pump is cranked. The Leiman Brother&apos;s No. 25 pump is built to displace 0.1 cu. ft. of air per min. when operated at 145 r.p.m. In the CMC impinger this pump is so geared that the required crank speed is approximately 59 r.p.m. Both types of impinger apparatus are equipped with a water gauge mounted on top of the case and they are calibrated and operated in reference to the gauge. The MSA impinger flasks and nozzles are used with both types. Calibrating Impingers A calibrated precision Sargent wet-test meter (Fig. 2) is used in calibrating the impinger apparatus. The apparatus is set up as for taking a sample in the field, except that the hose from the outlet side of the meter is attached to the impinger nozzle. One revolution of the large hand of the meter indicates 0.1 cu. ft. of air displaced. The impinger pump is cranked for one minute, therefore, at a rate that will cause the large hand of the meter to make one revolution. The gauge reading is noted, then the impinger is cranked at this rate for 10 min. and is checked for error. If the meter indicates that more or less than I cu. ft. of air has been displaced, the impinger must be cranked—ither slower or faster according to the error—for another 10 min. and again checked for error. This operation is repeated until the proper gauge reading for a sampling rate of 0.1 cu. ft. per min. has been determined. This gauge reading is used in all sampling operations until the impinger is again calibrated. An impinger is calibrated at least twice each week. MlCROPROJECTOR A Bausch and Lomb model B micro-projector is used (Fig. 3). It consists of a substantial metal base supporting the light source and a microscope. Illumination is provided by a Bausch and Lomb 4.5-amp. arc lamp with mechanical feed. A glass water cell filled with a heat-absorbing and color-filter solution is placed between the arc lamp and the microscope. Two types of solution have been found satisfactory— a 2 per cent solution of copper sulphate or a 2 per cent solution of picric acid and distilled water. This percentage may be varied to give the desired color to the light reflected upon the screen. Any standard microscope may be used. A Bausch and Lomb microscope with

Jan 1, 1946

By E. C. Harder

AT a Meeting of the New York Section, on May 23, 1918, the sole subject of discussion was the nature and occurrence of iron ores in certain parts of the world.? Owing to the importance of this subject, it has been deemed&apos; advisable to publish the remarks made at that meeting in the form of a paper, for presentation at the Milwaukee Meeting, instead of in the usual place, under News of Local Sections. The remarks related to the following districts: Brazil, by E. C. Harder, of the U. S. Geological Survey; Scandinavia, by Waldemar Lindgren, of the U. S. Geological Survey; Cuba, by C. M. Weld, of the U. S. Bureau &apos;of Mines; Southern Europe, by A. C. Spencer, of the U. S. Geological Survey; China, by H. Foster Bain, of the U. S. Bureau of Mines; Alsace-Lorraine, by Sidney Paige, of the U. S. Geological Survey. BRAZIL By E. C. Harder During the years immediately preceding the war, the iron ores of Brazil were attracting the attention of iron operators both in Europe and in the United States. This was clue to the large quantities of practically undeveloped ore available, and to its great purity. It is reasonably safe to state that had capital not been diverted from the exploitation of these deposits to war purposes, Brazilian iron ores would probably now be offered in European and American markets. It is certain that after the war these ores will play a prominent part in the economic reconstruction of Europe and perhaps in the development of America.

Jan 9, 1918

By N. Van Wingen

EFFective sand permeabilities can be ascertained from core analysis if the laboratory data are compensated to allow for the presence of connate or residual water. Such adjustments can be made by applying empirically derived correction factors. Effective sand permeabilities can be estimated also from analytical studies of field depth-pressure measurements. In one method the time rate of pressure build-up after a well is shut in from steady-state flow is required, and in the other the direct measurement of a steady-state rate of flow and the corresponding equilibrium pressure. The results of the investigation discussed in this paper demonstrate that these methods for determining mean effective sand permeabilities can be correlated. Introduction In recent years many investigators have demonstrated the practical value of an accurate knowledge of the physical and dynamical characteristics of reservoir sands containing oil and gas. Of these factors, which serve to completely define a porous medium as a container or carrier of fluids, the dynamical characteristics are of particular interest to the production engineer because they afford a direct measure of the rate at which the oil or gas content of a reservoir will become available. The constant that defines dynamically the porous medium as the carrier of a homogeneous fluid in viscous motion is the "permeability." Coke-analysis Method The generally accepted practice in com-piling permeability data has been to analyze in the laboratory core samples taken in the course of drilling through the productive horizon, and to obtain the mean measure of the productive ability or permeability for the interval open to production by graphical integration of the permeability profile. An average permeability ascertained in this manner, however, is useful to the production engineer only for comparisons of wells drilled in the same horizon, as permeabilities determined in the laboratory are measures of only the ability of a porous medium to transmit a single fluid. As a result of many investigations, it has been shown, however, that under actual reservoir conditions water is coexistent with the petroleum and that its presence greatly reduces the permeability of the sand to oil. Empirical relationships showing the effect of the presence of any immiscible fluid A (such as water) in a porous medium on the permeability of the medium to a fluid B (such as oil) have been established separately by Hassler,1 Dunlap,2 and the writer (Fig. I),3 so that the average specific or dry-core permeability as determined in the laboratory can be corrected to give the true measure of the ability of the horizon penetrated to transmit oil, provided the percentage of pore space occupied by connate water is known accurately.

Jan 1, 1942

By N. Van Wingen

EFFective sand permeabilities can be ascertained from core analysis if the laboratory data are compensated to allow for the presence of connate or residual water. Such adjustments can be made by applying empirically derived correction factors. Effective sand permeabilities can be estimated also from analytical studies of field depth-pressure measurements. In one method the time rate of pressure build-up after a well is shut in from steady-state flow is required, and in the other the direct measurement of a steady-state rate of flow and the corresponding equilibrium pressure. The results of the investigation discussed in this paper demonstrate that these methods for determining mean effective sand permeabilities can be correlated. Introduction In recent years many investigators have demonstrated the practical value of an accurate knowledge of the physical and dynamical characteristics of reservoir sands containing oil and gas. Of these factors, which serve to completely define a porous medium as a container or carrier of fluids, the dynamical characteristics are of particular interest to the production engineer because they afford a direct measure of the rate at which the oil or gas content of a reservoir will become available. The constant that defines dynamically the porous medium as the carrier of a homogeneous fluid in viscous motion is the "permeability." Coke-analysis Method The generally accepted practice in com-piling permeability data has been to analyze in the laboratory core samples taken in the course of drilling through the productive horizon, and to obtain the mean measure of the productive ability or permeability for the interval open to production by graphical integration of the permeability profile. An average permeability ascertained in this manner, however, is useful to the production engineer only for comparisons of wells drilled in the same horizon, as permeabilities determined in the laboratory are measures of only the ability of a porous medium to transmit a single fluid. As a result of many investigations, it has been shown, however, that under actual reservoir conditions water is coexistent with the petroleum and that its presence greatly reduces the permeability of the sand to oil. Empirical relationships showing the effect of the presence of any immiscible fluid A (such as water) in a porous medium on the permeability of the medium to a fluid B (such as oil) have been established separately by Hassler,1 Dunlap,2 and the writer (Fig. I),3 so that the average specific or dry-core permeability as determined in the laboratory can be corrected to give the true measure of the ability of the horizon penetrated to transmit oil, provided the percentage of pore space occupied by connate water is known accurately.

Jan 1, 1942

By G. F. Comstock

In the literature of metallography there is a large amount of material describing the various non-metallic inclusions found in iron and steel, and the appearance of sulfides, silicates, oxides, or alumina in steel under the microscope is fairly well known. The inclusions found in non-ferrous metals are, however, not so well known, at least judging from published writings, and the author has been interested in examining small castings of copper, brass, or bronze with certain inclusions purposely mixed with the metal, to see what was the characteristic appearance of each kind of inclusion, and if they could be readily distinguished from one another. The method employed in most cases consisted in overheating and oxidizing a small charge of copper, then cooling it somewhat, and adding a suitable amount of the element whose oxide it was desired to observe. This element was stirred in well and given time to react with the oxygen in the copper, then the melt was poured in a sand mold in the form of a cylinder about 21/2 in. (63 mm.) in each dimension, with a sprue about 11/2 in. (38 mm.) in diameter and 3 in. (76 mm.) high on top of the cylinder, the sprue forming the only riser. Samples for examination were cut sometimes from the center of the casting, and sometimes from the sprue or riser. A few of these samples were cast in chill molds. The appearance of sulfides was studied in an alloy containing 10 per cent, each of tin and lead, with which a flux of plaster of Paris was used. The appearance of foundry sand was studied in several alloys, one of them having sand stirred into the metal intentionally, and others having cut into the molds badly when poured, so that the castings were spoiled by included sand. As all this work was done on alloys having copper as the chief ingredient, the natural starting point was copper oxide, the appearance of which in a micro-section is well known. Cuprous oxide is soluble in molten copper, but separates out in freezing and forms a eutectic containing 3.5 per cent. Cu20, or 0.39 per cent. oxygen. If there is less

Jan 1, 1919

By Merton C. Flemings, Harold D. Brody

Analyses that include diffusion of solute in the solid phase are formulated to describe solute redistribution in dendritic solidification of metallic alloys. The analyses are based on conditions that include negligible undercooling before nucleation of solid phases, negligible increase of solute in advance of the tips of growing dendrites, complete diffusion within the liquid over distances the order of dendrite spacings, and a plate-like dendrite morphology. The classical nonequilibrium freezing equation, Cs = kCo(I - fS)k-l accurately describes solute redistribution between dendrite arms for the solidification processes considered, provided diffusion in the solid is negligible. To account for the effect of diffusion in the solid, an analytic expression is given which is similar inform to the classical expression. 172 addition, a numerical-analysis procedure is etnployed to examine in more detail the effect of diffusion both during and after solidification. The analyses are intended for application to solidification of castings and ingots to describe a) final solute distribution after solidification and cooling to room temperature (microsegregation) and b) local fraction solid as a function of ternperatcre within the solidifying casting or ingol. SOLUTE redistribution in cellular and dendritic solidification has been discussed quantitatively in recent papers&apos;-3 and in a series of earlier work.4-6 However, none of the analyses developed have permitted close correlation of theory with final solute distribution (microsegregation) observed in castings and ingots. In this paper, discussion is given of approximations which may reasonably be made in mathematical treatment of microsegregation, and analyses are presented based on these approximations. Numerical results are given for Al-Cu alloys. In a second paper the analyses will be compared with exeriment.7 CLASSICAL ANALYSIS The classical quantitative treatment of solute redistribution within a closed "volume element" has been derived repeatedly, for example by Gulliver,4 Scheil,&apos; and Pfann.&apos; This "classical nonequilibrium solidification equation" is written for constant partition ratio: where Cs = interface composition of the solid when the weight-fraction solid within the "volume element" is fs (weight fraction, wt pct); k = equilibrium partition ratio; Co = initial alloy composition within the volume element (weight fraction, wt pct). Assumptions of Eq. [11 are: 1) There is negligible undercooling before nuclea-tion, or from curvature or kinetic effects. 2) There is no mass flow in or out of the volume

Jan 1, 1967

By Wilbur C. Helt

The United States is self-sufficient in bituminous coal and lignite resources to meet the high range of energy demands forecast for the remainder of this century and beyond. The known or proved recoverable reserves of energy from coal in the United States are estimated to be 88% of all fuel reserves, whereas the energy reserves for oil are estimated to be 276, gas 376, shale oil 5% and uranium 2% (as now used). The demand for all forms of energy used in the United States is increasing at a rate of 3% to 4% per year. Presently, coal is competitive on a supply and cost basis with other fuels in conversion to electrical energy and as a source of energy for converting raw minerals to metals and nonmetals, but it is not competitive in transportation and other uses that demand the high flexibility that oil and gas enjoy because of their form. Coal conversion to fluid forms is essential to make it interchangeable with gas and oil to meet the markets they now occupy. Essentially, the chemical constituents that make up coal are the same hydrocarbons that make up oil and gas; it is a matter of reconstituting the proportions of hydrocarbons to make coal interchangeable with the other fuels. The production of coal has fluctuated widely between the peak production of 631 million tons in 1947 to a low of 403 million tons in 1961. Throughout this period, the total value of the production, f.o.b. mines, has averaged more than $2 billion annually. Beginning with 1962, there has been a steady upward trend in coal production with only minor fluctuations that were mainly caused by labor disputes. Output for the latest typical year, 1970, was 588 million tons. Production in 1971 was affected by a major strike and dropped to 548 million tons. Exports of coal from United States to foreign countries have increased steadily and developed into a substantial annual market for over 71 million tons per year. About 56 million tons of this is suitable for metallurgical purposes, and the remaining 15 million tons is exported for electric power generation. The largest single-industry demand for coal is in the electric generating industry, Fig.1. Roughly, the production of electricity has doubled every

Jan 1, 1973

By W. M. Weidel

While electric lamps both of the cap and hand type are being introduced into many mines requiring the use of safety lamps, the oil-burning safety lamp is still used in the great majority of cases, and even where the former are used the latter must continue to be employed when testing for gas. It is not the intention to discuss here the relative merits of the two types, but simply to give a comparison of the illuminating effect or candlepower of several kinds of ordinary safety lamps used for general work and for testing. Few data have been published in recent years on this subject and because of this it is hoped that the following information may be of some interest and value. The lamps tested were in all cases of standard size and in good condition, all practically new, and were selected from a collection of about 100 with a view to obtaining results on as many different common types as possible. With the exception of the gasoline-burning lamps, five different oils were used in each lamp to determine whether particular oils were suited to particular lamps and whether there was any great difference in the illuminating power of the different oils, or whether some oils would maintain their illuminating power better than others throughout a working period of 8 hr. No attempt was made to determine the amount or cost of oil burned by each lamp, but the condition of the lamp, wick, glass, and gauzes was noted after each test. Description of Apparatus Used The candlepowcr was determined with a Weber photometer, Fig. 1, which consists of a tube A containing a movable milk-glass screen adjustable by the milled screw B. The distance of this screen from the standard light at C is read on the scale attached to A. The standard light consists of a benzine lamp so arranged that the height of the flame may be adjusted exactly by observing it against a scale in the lamp box through an opening in the front. The lamp to be tested is placed at D so as to illuminate a milk-glass screen in the holdcr at E. At the intersection of the axes of tube A and sight-tube F are placed two 45&apos; prisms with their surfaces in contact so that the light from D may pass through with-

Jan 1, 1917

By J. R. Campbell

A dozen years or so ago, the general superintendent of our company, now the president, Mr. W. H. Clingerman, asked me to study the coal-washing problem. This work brought me into contact with the best washery men of the country. I investigated many kinds of coal-cleaning schemes, made numerous tests, and collected voluminous data, yet I confess frankly that I have had little practical washery experience and that little has been at the Middlefork washery of the United States Fuel Co. at Benton, Franklin Co., Ill., which was put in operation last fall. While this washery installation may not be the "last word," I am quite sure it embodies some long steps in advance of the older designs and constructions. From the washery man&apos;s standpoint, the removal of pyrites or marcasite, a disulfide of iron, is the one vital problem, and the one productive of the most beneficial results, so far as elimination of sulfur is concerned. Sulfates and so-called organic sulfur occupy quite a sec-. ondary position and are generally lost sight of in coal washing. The elimination of sulfur in coal in the past has usually been accomplished by jigging in water, but lately concentrating tables have been successfully employed. The flotation process has been applied to fine coal, but is hardly considered practicable. Coal washing is rapidly coming into its own, due to the depletion of low-sulfur coals for coking purposes, and we predict that soon the same careful attention will be given to the design and construction of coal washeries that is now being given to byproduct plants and other allied industries. Heretofore, coal washing has been considered only an incidental and necessary evil instead of the big problem it is—one worthy of the most careful study by men well trained along technical lines and capable of delving into the fundamental principles. Preparation oF Coal It is usually necessary to crush the mine-run coal before sending it to the jigs, the degree of crushing depending on the occurrence of the sulfur and other impurities in the coal and the type of jig over which it is to be sent. The feldspar jig requires finer and more uniform crushing

Jan 1, 1920

By G. Das, S. V. Radcliffe

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

Jan 1, 1969