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By Thomas L. Kesler

IF all of the 14 million tons of kaolin produced in Georgia through 1949 had been mined from a single deposit 20 ft thick, it would represent a mined-out area of less than 1 sq mile. This measure of depletion is in such strong contrast with the large size of the area shown in Fig 1, from which most of the output has come, that there has developed an optimistic outlook concerning future supplies of kaolin equal in grade to that already mined. Dozens of mines are widely but not evenly distributed within this area, which consists mostly of parts of Twiggs, Wilkinson, and Washington Counties. There are scarcely half a dozen kaolin mines elsewhere in the state that have been appreciably productive, although the areas in which they occur may have good possibilities. Most of the efforts to find new deposits, therefore, will be confined to the area shown in Fig 1, for it is here that the industry has established its facilities, and prospecting is encouraged by relatively numerous outcrops and opened deposits of kaolin. The Ounulgee and Oconee Rivers drain the area. They and their main tributaries have flat, swampy flood plains covered with alluvium and for the most part thickly wooded. The divides between these streams are intricately dissected by the minor tributaries, which are spring-fed in steep-walled, cirquelike heads. Gullying and undercutting are active along the streams but are partly controlled by unusually dense vegetation. The mean altitude is about 350 ft above sea level and locally the relief ranges up to 200 ft. The area is near the inner margin of the Coastal Plain, and is underlain by Cretaceous and Tertiary formations. Its general location is shown in black in Fig 1. The division between the Coastal Plain and the Piedmont Plateau is actually a very sinuous line, the fall line, marking the extent to which erosion has stripped the Coastal Plain sediments from the southeastward- sloping surface of the crystalline rocks of the Piedmont.

Jan 1, 1952

By J. H. Wiggins

Granting that about 600,000,000 bbl. of light oil will be produced in the United States this year and taking the Bureau of Mines' statement that 71/2 per cent. of the total crude production will be lost before the finished gasoline is shipped from the refinery, the volumetric evaporation will be 45,000,000 bbl. of the cream of the gasoline. This is a ConseNative figure and is equivalent to 1,890,000,000 gal. The real value of this liquid is at least 150 per cent. of the value of an equal volume of motor fuel. At only 10 c. per gal., the value lost is $189,000,000; at 15 e. per gal. the loss is $283,000,000. If 750,000,000 bbl. of light and heavy oil are produced here this year and the average price per barrel is $1.333, the gross return to producers will be $1,000,000,000 and the net returns about $200,000,000, if the cost of production is only $1 per barrel. Thus, the evaporation bill is about 20 per cent. of the gross income and is equal to the net income of the crude-oil producers. Two other comparisons with popular phases of the petroleum industry will help give an adequate conception of the evaporation problem. These are: cost of preventing evaporation compared to cost of cracked gasoline and volume of gasoline evaporated compared to volume of gasoline manufactured from natural gas. When gasoline is manufactured from heavy oils by the cracking process, there are two important divisions of the total cost of the product: the manufacturing cost, such as plant cost, operation, maintenance, etc., and the cost of the barrel of heavy oil, from which the cracked gasoline is made. In preventing evaporation loss the cost first mentioned enters into the total cost, but the second cost is eliminated; that is, no charge is necessary for stock from which to manufacture the saving. A detailed analysis of these two costs on a large scale shows that the net income from a dollar spent in preventing evaporation loss from the lease through the refinery is twice the net income from a dollar spent in cracking. Of

Jan 1, 1924

By D. I. Cameron

A graphical method has been developed and tested for separating the effects of grain boundary and lattice diffusion in polycrystalline materials. The method is based on the assumptions that for unidirectional diffusion from a plane source of radioactive material, the penetration of activity far into the specimen from the source is due solely to grain-boundary effects, and 2) pure grain-boundary penetration is described by a linear plot of log activity vs the perpendicular distance x from the source. The former is based on the fact that at any temperature the grain-boundary diffusion coefficient is always very large relative to the lattice diffusion coefficient; and the latter is based on the derivation of Fisher.1 Using these criteria along with appropriate penetration data (in this case, data for the diffusion of Cb95 in pressed and sintered UO2) one first plots the data as log activity vs x2 and as log activity vs x, as shown by curves 1 and 2, respectively, in Fig. 1. A tangent line to curve 2 for large x (curve 3) is then transferred to the x2 plot as curve 4. The subtraction of curve 4 from curve 1 yields curve 5 which represents only lattice diffusion and may therefore be related to the diffusion coefficient by using the appropriate solution to Fick's second law. It should be noted that the method has inherent inaccuracies for small and large values of x2 and only the subtraction at intermediate values can be used. For the illustrated specimen the calculated diffusion coefficient is 3.4 X 10-11 sq cm per sec at 1445°C while data using a single crystal of UO2 at the same temperature yielded a value of 2.9 x 10-11 sq cm per sec.2 The same method of subtracting grain-boundary diffusion effects to obtain lattice diffusion coefficients has been successfully used to determine self-diffusion coefficients of Th230 in polycrystal- line thoria.

Jan 1, 1962

By J. E. Lawver, R. R. Beebe, J. L. Wright

This paper describes a separation coefficient Ks useful as a "measure of goodness" to be used in factorial and response surface methodology. The use and limitations of the separation coefficient is discussed, and a computer program for calculating the coefficient is presented (in Fortran). The application of the coefficient is illustrated in terms of electro-dynamic separation of Mesabi semitaconites. Present Mesabi wash ores and gravity concentrates with an iron to silica ratio of only 5 or 6 to 1 are no longer readily salable, for the steel industry now demands concentrates with an iron to silica ratio of at least 10 to 1 and a structure comparable to pellets. THE ELECTRODYNAMIC-FLOTATION PROCESS Wash ores and gravity concentrates consist largely of hematite, goethite, and quartz. Several methods of beneficiation are being tested at the University of Minnesota Mines Experiment Station with a view to improving the salability of these materials. A process that looks particularly promising from an economic view is the combination electrodynamic-flotation process shown in Fig. 1. This process is designed to upgrade gravity concentrates to a 5% silica content and to provide fines — from the flotation circuit —to permit agglomeration. Although electrodynamic concentration of specularite is an established and well-understood process, the electrodynamic concentration of Mesabi goethite-hematite gravity concentrates presents interesting and enigmatic problems. The gravity concentrates are by no means assemblages of clean, free mineral particles. Instead, the concentrates to be retreated consist of a great variety of middlings. More or less complete liberation would require grinding to minus 325 mesh, at which size only flotation could be expected to function, and even then grades and especially recoveries would not be satisfactory. A close examination of the gravity concentrates shows that grinding to 100% minus 48 mesh will produce a product which is approximately 70 weight % plus 325 mesh (see Fig. 1). The remaining 30 weight % can be deslimed and floated by conventional methods, but the resulting iron concentrate is

Jan 1, 1964

By T. S. Lundy, J. I. Federer

A graphical method has been developed and tested for separating the effects of grain boundary and lattice diffusion in polycrystalline materials. The method is based on the assumptions that for unidirectional diffusion from a plane source of radioactive material, the penetration of activity far into the specimen from the source is due solely to grain-boundary effects, and 2) pure grain-boundary penetration is described by a linear plot of log activity vs the perpendicular distance x from the source. The former is based on the fact that at any temperature the grain-boundary diffusion coefficient is always very large relative to the lattice diffusion coefficient; and the latter is based on the derivation of Fisher.1 Using these criteria along with appropriate penetration data (in this case, data for the diffusion of Cb95 in pressed and sintered UO2) one first plots the data as log activity vs x2 and as log activity vs x, as shown by curves 1 and 2, respectively, in Fig. 1. A tangent line to curve 2 for large x (curve 3) is then transferred to the x2 plot as curve 4. The subtraction of curve 4 from curve 1 yields curve 5 which represents only lattice diffusion and may therefore be related to the diffusion coefficient by using the appropriate solution to Fick's second law. It should be noted that the method has inherent inaccuracies for small and large values of x2 and only the subtraction at intermediate values can be used. For the illustrated specimen the calculated diffusion coefficient is 3.4 X 10-11 sq cm per sec at1445°C while data using a single crystal of UO2 at the same temperature yielded a value of 2.9 x 10-11 sq cm per sec.2 The same method of subtracting grain-boundary diffusion effects to obtain lattice diffusion coefficients has been successfully used to determine self-diffusion coefficients of Th230 in polycrystalline thoria.

Jan 1, 1962

By John D. Hanks

INTRODUCTION Chevron's Uranium Mill is located near Panna Maria, Texas; 70 miles southeast of San Antonio. Designed by Kaiser Engineering, the Mill will process a nominal 2500 dry T.P.D. of uranium bearing ore containing 15% uncombined moisture. Earl Torgerson, San Mateo, California, is the consulting metallurgist on process design. Feed to the plant consists of a mixture of high, medium and low grade sandy day ore; the average grade of the ore will be 0.7% throughout the life of the project. The ore is delivered to the mill via truck and stored by type in individual piles on a flat storage area. The ore is fed by conveyor to a semi-autogenous grinding mill. The SAG mill discharge slurry is pumped either to a storage tank or to the first of five mechanically agitated leach tanks where both H2SO and NaC103 are added. Following leaching, the slurry is mixed with thickener No. 2 overflow before being pumped to a six-thickener, countercurrent decantation circuit where the solution containing uranium is separated from the leach residue. The residue is washed essentially free of solubilized U308 values at the sixth thickener and discharged to an adjacent tailings pond. The first thickener overflow, containing approximately 0.4 grams U308 per liter, is filtered for clarification and sent to the liquid ion exchange (solvent extraction) section. The pregnant aqueous solution is mixed with organic solvent containing amine on which the complex uranyl sulfate ions are absorbed. The immiscible aqueous and organic solutions are mixed and separated in each of the four stages of solvent extraction. The final pregnant organic is directed to a stripping section. A strip solution containing (NH4)2S04 (ammonium sulfate) and NH4C1 (ammonium chloride) is contacted and separated from the organic in each of the four mixer-settler units. The strip solution is mixed with NH3 and the uranium precipitates along with trace amounts of sulfate, chloride, and ammonia. After washing in a thickener, the uranium precipitate or yellowcake is centrifuged and dried in a multiple hearth roaster. Overflow from the yellowcake thickener is recycled back to the stripping section of the solvent extraction circuit. The dried yellowcake concentrate contains more than 98% U308; diluents include H20, ammonia, chloride, etc.. Impurity concentratons in the product are sufficiently low, after drying, to permit direct shipment to refinement installations. ORE RECEIVING Ore from one or more mines will be stored on a pad adjacent to the uranium processing mill. Ore stored in the area may total 200,000 tons or more, which is equivalent to over a two month treatment reserve for the mill. In addition to providing surge capacity, the proposed storage facility permits natural oxidation of the ore and affords an area which, in turn, improve plant U308 recovery and reduces consumption of oxidizing reagents. The uranium bearing ore is segregated at the storage area into several distinct types. One group can be identified by its sandy, clay-like matrix and will be distinguished by its U308 concentration of high, medium, or low. Other ores contain substantial quantities of carbonaceous shale, typical of ore in the area. Feed is recovered from any one or combination of piles, in accordance with operating requirements and recovery factors. The run-of-mine ore, at 15% moisture, is transferred to a stationary, 24-inch by 24-inch grizzly. Undersize material falls into a 280-ton surge hopper located below grade, while oversize material is removed for preliminary size reduction. A sump-pump is located at the reclaim hopper to recover excess mositure and ore spillage for processing. Plant feed is continuously drawn from beneath the hopper at an average rate of 120 tons per hour by means of an apron feeder. The hopper is equipped with a low level safety system to protect the apron feeder assemble. The ore is transferred to a belt conveyor, elevated and discharged into a semi-autogeneous grinding (SAG) mill. Water addition at the discharge point is automatically controlled by a feed rate, moisture analyzer system located on the belt conveyor. GRINDING Minus 24-inch ore, at a rate of 120 mosit tons per hour and cyclone underflow at a rate of 137 moist tons per hour are combined with a controlled 186 gallons per minute fresh, warm plant process water. The mixture, at an average 68% solids, is fed to the SAG mill. The viscous mass is subjected to grinding at a temperature moderately above ambient resulting from use of the 140ºF well water. The 16.5'x5.0' Marcy Mill is driven by a 500 H.P. A.C. Motor. The mill will rotate at 73.4% of critical speed or 14.06 R.P.M. Eight percent of the mill's volume will be occupied by steel grinding balls; 22% by pulp. The mill is designed to rotate in either direction, in order to obtain maximum lifter and liner life. The SAG mill is equipped with a trommel having ½ -inch slots for removal of oversize material at its discharge and the large or oversize material is collected and either recirculated periodically or discarded. The undersize slurry is diluted to 64% solids with the addition of hot well water or mine water to the mill discharge sump. The flow is automatically controlled utilizing data from a gamma density meter. The slurry from the mill discharge sump is pumped to hydrocyclones to classify the particles. Underflow from the cyclones, at 80% + 28 mesh, is recycled to

Jan 1, 1979

By A. Gittens

In two recent papers stevens13,14 has compared the methods used by various investigators to calculate the contribution of grain boundary sliding to the total creep strain. In previous work Gittins15 has shown that the strain due to grain boundary sliding Egb may be estimated from the lateral, d, and vertical, v, displacements of a longitudinal marker line as it crosses a grain boundary. He obtained the formula, where nL is the number of grains per unit length measured on the specimens before creep, and \p and ?2 are, respectively, the angle between the grain boundary and the stress axis in the plane of the surface and normal to the surface (see Ref. 13, Fig. 1). The subscript L indicates that observations are made in a longitudinal direction parallel to the stress axis. It was generally found more convenient to measure the grain size of a specimen before creep; this was necessary to characterize the specimen. Stevens13,14 however prefers to define Egb in terms of NL, the number of grains per unit length after creep. Since nL - NL(l + €1) where cr is the total strain, Eq. 1 11 becomes The quantities d and ?, can be determined directly from the marker lines, while v and ?2 are found by sectioning normal to the surface. In the interior when symmetry considerations suggest that v = d and ?1 - ?2 the strain due to grain boundary sliding may be estimated from a suitable internal longitudinal marker line by the equation To determine cgb at the surface using either Eqs. |1| or [2] requires a large number of observations, particularly if many estimates of cgb are required, and also it is impossible to determine the variation of cgh with creep strain on the same specimen because the specimen must be sectioned to determine ?2. Gittins15 found that for a number of materials the sliding vectors d and v were related to each other so that provided one was measured the other could be estimated. Also during creep, grain boundaries tend to migrate to meet the surface at angles close to 90 deg,16 so that the term nL(v/tan?2)L was comparatively small when 4,- 90 deg. By determining the frequency distributions of ?2 on control specimens this term can be estimated without sectioning the creep specimen. In this way an empirical equation based on Eqs. |1| and |2| enables cgb to be determined with reasonable accuracy with fewer observations of the angles and sliding components. For B brass, the relationship between the sliding vectors indicated that ceb = 0.85nLVL . In Table I of Stevens paper this equation is stated to be incorrectly derived. However it should be pointed out that this is an empirical equation derived from Eq. |1|, based on a knowledge of the relationship between d and v and on the frequency distributions of ?1 and ?2 for a particular material under given experimental conditions. It was used only when these relationships were known, to reduce the number of observations required to estimate cb. Therefore the comment "derivation incorrect does not apply. Gittins did in fact make use of Eqs. |1| and 131. Author's Reply I must apologize to Dr. Gittins for the inadvertent misrepresentation of his work. Other slight errors in Table I have been pointed out to me. Langdon (No. 12 Table I), Gifkins (No. 7, Table I), and possibly other workers did not measure the number of boundaries per unit length after creep (A') but the corresponding quantity bejove creep (n). As Dr. Gittins points out, a completely analogous set of equations to those given in the paper can be derived by the same method: using the number of boundaries per unit length before instead of after creep. The basic equations corresponding to Eqs. [2] and 171 of the paper, from which all the others can be derived, are

Jan 1, 1968

By E. P. Sadowski, L. P. Trudeau, C. R. Cupp

The apparent high radiation resistance of two varieties of maraging steel is described and an indication is given of some phenomena that require further study. Two aspects were included in this work: a) effects of neutron irradiation on the room-ternperature tensile properties of 18 Ni (250) maraging steel; b) effects of neutron irradiation on the room-temperature tensile properties and on the impact resistance of 12 pct Ni-5pct Cr-3 pct Mo maraging steel. Subsize tensile and impact test specimens were used. The 18 pct Ni steel was irradiated to a neutron fluence of 3.3 x 10 n per sq cm (>1 mev). The 12-5-3 variety was irradiated to two neutron fluence levels: 1) 1.5 x 10 per sq CTTL (>I mev) 2) 3.1 x 10'9nper sq cm (>1 mev) In all respects both varieties showed unusually high resistance to change in strength and ductility after neutron irradiation. There was no noteuorthy over-aging in fully aged specimens. The aged 18 pct Ni variety showed a 4 pct drop in yield strength upon irradiation with little or no decrease in ductility. The aged 12-5-3 variety showed a 5 pct increase in yield strength on irradiation with a very slight decrease in ductility. The impact transition of the 12-5-3 variety only increased 35 to 45 Co upon irradiation, and the shear energy fracture value for the subsized specimens decreased only half as much as expected frorn experience with conventional steels. THE development of new reactor concepts and nuclear technology now depends to a large extent on the availability of new alloys capable of operating in stringent reactor environments. One of the most important new types of metallurgical material that has been developed in recent years is the maraging steels. This type of alloy is notable for its remarkable retention of toughness at high strength levels. In the course of a research program which incorporated ambient temperature radiation effects studies to determine the mechanisms of strengthening of 18 pct nickel maraging steel,' it became evident that the resistance of the alloy to neutron-induced changes in mechanical properties should be studied. This paper is to present evidence for the apparent high radiation resistance of maraging steel. The experimental program was in two parts: a) Effect of neutron irradiation on the room-temperature tensile properties of 18 Ni (250) maraging steel. b) Effect of neutron irradiation on the room-temperature tensile properties and on the impact resistance of 12 pct Ni-5 pct Cr-3 pct Mo maraging steel, a cobalt-free alloy. EXPERIMENTAL PROCEDURE 18 Ni (250) Maraging Steel. Tensile test specimens were machined to the dimensions shown in Fig. 1 from $-in.-diam hot-rolled bar of the composition shown in Table I. All specimens were austenitized for 1 hr at 1500°F (815°C) in argon and water-quenched. Half of the specimens were then mar aged for 3 hr at 900° F (482°C) in argon. The specimens were irradiated in a fast neutron transformer rod at approximately 120°F (50°C) to a neutron fluence of 3.3 x 10 10 n per sq cm (>1 mev). The irradiation was performed in the NRX reactor of Atomic Energy of Canada, Ltd., at Chalk River, Ontario, Canada. In all cases, the thermal neutron flux was three times as high as the reported (> 1 mev) fluxes. The room-temperature tensile tests were performed with an Instron tensile testing machine operating at a crosshead speed of 0.05 in. per min. The loading frame of the machine, mounted in a hot cell, was remotely controlled. The specimens were tested either

Jan 1, 1968

By S. W. Parr

FRom a study of sulfur with reference to its specific combination in coal, published as University of Illinois Bulletin No. 111, 1919, it is now possible to determine the various forms of this constituent with a good degree of accuracy. This fact suggests two questions: (1) Are the same analytic methods applicable to coke? (2) Is there any correlation between the forms of sulfur as it occurs in coal and the residual sulfur found in the finished coke? The answer to the second question might be helpful in answering a still further question as to what chemical route the sulfur travels during the carbonization process. While the special studies along these several lines are by no means complete, a few preliminary statements are herewith presented. They may, at least, prove suggestive of further possibilities from a continuation of research along these lines. There is no reason, chemically, why the analytic methods devised for coal should not apply to coke and a fair amount of experience in connection with such use affords confirmation of this statement. The procedure in the case of coal may be briefly outlined as follows: Combinations of sulfur as sulfate, or sulfides other than FeS2, are readily dissolved without disturbing either the pyrite sulfur or the sulfur that is in organic combination by means of dilute hydrochloric acid— 1 part of concentrated HC1 to 10 parts of water or approximately 3 per cent. acid. Digestion at 60" C. for 24 hr. insures the complete removal of the sulfur thus combined. In freshly mined coal the amount of such sulfur is very small; weathering processes, however, may very materially increase the amount of sulfate present. Sulfur in the form of iron pyrites, or marcasite, may be quantitatively removed without oxidation of the organic sulfur as follows: A dilute solution of nitric: acid is prepared having 1 part of concentrated HNO3 to 3 parts of water or with an approximate specific gravity of 1.12. The coal sample, ground to 100 mesh, is digested with this solution at room temperature for 3 or 4 days, filtered and evaporated to dryness, taken up with water, and acidified with hydrochloric acid. It is now ready for precipitation with barium chloride in the usual manner. If sulfate is present in the original sample to an appreciable amount, it should be removed by the procedure outlined above. It is not necessary here to give the methods for determining the two

Jan 1, 1920

By J. Chipman, N. J. Grant, H. L. Bishop

In a recent review1 of the iron-oxide activity of simple open-hearth type slags containing lime, magnesia, silica, and iron oxide, it was established that activity values were lacking in the range of slag composition most interesting to open-hearth operators. That is, laboratory experimental results were not available for simple slag compositions having basicity ratios ranging from about 1.5 to 4.0 and iron-oxide concentrations ranging from about 10 to 30 pct. However, in order to estimate iron-oxide activities for the range of slag compositions of interest to open-hearth operators, activity values were calculated using ionic concepts of slags by following the treatment of Flood and Grjotheim.2 A study of these data indicated that the desired slag compositions were probably not obtained due to the limited solubility of the basic constituents lime and magnesia. Thus, it was decided to make additional heats equilibrating molten carbon-free iron with slags containing MnO, P2O5, and A12O3 in percentages commonly found in open-hearth refining slags, and to examine the liquid slag field over a greater composition range than that of the simple slags studied previously. The ranges of the constituents used in the present study are summarized as follows: Metal Pct 0 0.070-0.201 PctMn 0.005-0.142 Slag Pet CaO 24.45-41.90 Pct MgO 3.23- 9.21 Pct MnO 1.13- 9.62 Pct FeO 20.08-48.7 Pet Fe2o3 2.23-10.14 Pct SiO, 3.51-28.50 Pct P2O5 0.23- 3.65 Pct A12O3 0.43- 3.38 The experimental procedure consisted of equilibrating the slag and metal in an induction furnace provided with a carbon-arc slag heater as described previously by Winkler and chipman.3 The data obtained were used to establish both the iron-oxide activity values and the activity coefficients for FeO and MnO in the open-hearth type slags. RESULTS AND DISCUSSION I) Iron-Oxide Activity of Open-Hearth Type Slags—The iron-oxide activities of various iron-oxide contents of the open-hearth type slags, as calculated from the oxygen content of the metal by the method of Taylor and Chipman, 4 are plotted against the molar ratio of bases to acids in Fig. 1. The data represent slag-metal studies carried out in the temperature range 1530" to 1700°C. For the purpose of this study, the constituents CaO, MgO, and MnO are assumed equal on a molar basis in their effect on the iron-oxide activity; and, PO2,5 is calculated as an equivalent to SiO2 and A12O3 on a molar basis. The total iron content of the slag is calculated to FeO and is represented by the symbol (FeO) The points presented in Fig. 1 represent the results of this investigation for iron-oxide slags containing between 1 and 8 mol pct MnO and less than about 2 mol pct each of A12o3 and PO2,5 The curved lines shown in Fig. 1 represent the results of the summary1 of the iron-oxide activities of simple slags of the system (CaO + MgO) - SiOz - (FeO),. The solid portions of the curved lines represent experimental data, and the broken portions represent iron-oxide activities calculated from ionic concepts of slags. The iron-oxide values of the open-hearth type slags of this study are in good agreement with the values of the simple iron-oxide slag system. Only a slight modification is indicated. The results of this study suggest that the line for 20 mol pct (FeO), should be lowered slightly in the basicity range 1.5 to about 3.

Jan 1, 1959

By F. Clerf

BLAST-FURNACE practice in France is determined more or less by the character of the ores used. Some French ores are siliceous and others are calcareous, therefore by proper burdening a self-fluxing mix can be used. This is not always clone, as, for instance, at the Caen furnaces of the Société Métallurgique de Normandie, where the Soumont ore of Normandy is used, which is a carbonate. This ore is crushed and calcined at the mines, yielding a product that is porous and well sized for blast-furnace consumption. The lines of the Caen furnaces are given in Fig. 1. Originally these furnaces had 16 tuyeres; six are plucked and these tuyeres are set in the lower part of the bosh. Operating data on these furnaces are given in Table 1.

Jan 1, 1937

By C. M. Wayman, R. Bullough

The selection rules fir twinning in fcc, bcc, and hcp lattices as established in Part I are used to predict the positions, relative to the matrix (untwinned) reciprocal lattice, of reflections due to double diffraction from the twins. These doubly diffracted reciprocal lattice points, which often coincide with ordinary matrix reciprocal lattice points, are deduced from kinematical considerations. 1. INTRODUCTION. A Bragg diffracted beam under certain conditions may be rediffracted to produce a third beam i .e., the incident beam, the first diffracted beam, and the rediffracted beam). In such a case an initial incident beam is diffracted by a set of planes; the diffracted beam may then be rediffracted by a second set of crystallographically different planes. The initially diffracted beam acts as an incident beam for the second set of planes, thus resulting in a twice-diffracted beam. This twice-diffracted beam may be regarded as a singly diffracted beam from a set of lattice planes whose structure factor is zero. In order to find the reciprocal lattice point corresponding to the doubly diffracted beam, it is customary1,2 to combine vectorially the component reciprocal lattice vectors which head to the extra reflections. For example, hcp metals may exhibit the forbidden (0001) reflections.* As Fig. 1 shows, the (0001) reciprocal lattice point results from: (1010) + (1011) = (0001) Similarly, combination of the beams (1011) and (1010) gives rise to: (1011)- (1010) = (0001) As Pashley and stowell2 have pointed out, even if in fact the double diffraction points lie on the Ewald sphere, they may not actually represent observed doubly diffracted beams because the initially diffracted primary beam, i.e., the source, must also correspond to a reciprocal lattice point which lies on the sphere of reflection; however, the reciprocal lattice point corresponding to the planes causing the rediffraction need not lie on the sphere. A little consideration of the Ewald construction in Fig. 2 makes this argument clear. In Fig. 2 the incident beam is represented by the wave vector ko. This beam is initially diffracted by the lattice planes (hkl) which are represented by the reciprocal lattice vector r(hk1) The reciprocal lattice point (hkl) of course lies on the Ewald sphere in order that the Bragg condition be satisfied. The initially diffracted beam is represented by the wave vector kl. If now kl is rediffracted by a second set of planes (h'k'l') resulting in k2 the lattice point (h"k"l") must lie on the Ewald sphere if the double diffracted beam k2 is to be observed. It is clear that (h"k"l") is the vector sum of (hkl) and (h'k'l'), and that (h'k'l') need not in fact lie on the Ewald sphere. In the usual1,2 analysis the origin (000) of the reciprocal lattice is shifted to the point (hkl). However the translation (000)— (hkl) automatically shifts the point (h'k'l') to the Ewald sphere, i.e., (h'k'l')— (h"k"1"). It is to be seen that (hkl), the source reflection, must lie on the Ewald sphere; if it does not the translation (000) — (hkl) places the "new" origin off the sphere, not satisfying the Ewald construction and hence the Bragg condition. The point (h"k"1") of course must lie on the sphere if it corresponds to an observed doubly diffracted beam. It is to be noted that crystals having the hcp structure, as just discussed, are intrinsically capable (i.e., without lattice defects such as twins) of double diffraction, as contrasted to fcc and bcc crystals which are extrinsically capable (requiring lattice defects or reoriented lattice regions). The "allowed" bcc reciprocal lattice points (which comprise an fcc lattice) are (110), (2001, (211), (220), (310), (2221, (321), (4001, and so forth. The combination

Jan 1, 1967

By Denton R. Wieland, Harvey T. Kennedy

The frequency of bubble formation is measured for oils from the East Texas field and the Slaughter field in cores from these fields. East Texas oil was also tested in a Slaughter core. The data checked previous results in that a definite supersatrtration may exist without the formation of bubb1es in periods of observation totaling 158 hours. Measured frequencies varied from zero at supersaturations below 14 psi to 3.1 bubbles/sec/ cu ft of rock at 40 psi supersaturation INTRODUCTION The formation of bubbles is essential in a solution gas-drive reservoir in the displacement of oil. In formations where the basic pore structure consists of fine pores or fractures connected to larger pores or crevices, bubble formation in the fine pores may be the most effective means by which the oil may be displaced, even if secondary methods are applied. The most efficient oil displacement would occur where at least one bubble would form in every pore. As the pressure decreased, the bubbles would expand, due to the gas diffusing into the bubbles, and displace the oil toward the lower pressure area around the wellbore. Kennedy and Olson' have shown that the total number of bubbles formed in a reservoir is primarily dependent upon the rate of pressure drop, the rate of diffusion of gas through oil, and surface area of the rock. The surface tension of the liquid and other characteristics of the rock may also exert an important effect, which would give a different frequency to different rocks and oils and various combinations of different rocks and oils. The purpose of this paper is to present frequency measurements on two rocks and two oils. A hydrocarbon liquid is at equilibrium with gas when it contains the maximum volume of gas in solution that it can have at the prevailing temperature and pressure. If the temperature is maintained constant and the pressure is reduced below the equilibrium pressure, the liquid is supersaturated to the extent of the difference in the two pressures until the gas comes out of solution as a free phase. As shown by previous work, and confirmed in the present investigation, the number of bubbles formed per second increases with an increase in supersaturation. Kennedy and Olson showed that a tenfold increase in the rate of pressure decline resulted in a tenfold increase in the number of bubbles formed per cubic foot of reservoir rock, e.g., for decline rates of 0.1, 1, and 10 psi/day, the number of bubbles formed per cubic foot would be 40, 400, and 4,000, respectively. For another hypothetical system, Stewart, Hunt, and Berry' estimated that for pressure declines of 0.1, 1, and 10 psi/day, the number of bubbles formed would be approximately 10, 100, and 1,000 cu ft of reservoir rock. Their value of bubbles formed is lower than given by Kennedy and Olson; however, the order of magnitude agreed satisfactorily. The rate of diffusion will also affect the number of bubbles formed. If the diffusion coefficient is high, the gas in solution in the oil surrounding a single bubble will tend to diffuse into the bubble rather than forming other bubbles. On the other hand, the formation of more bubbles will tend to occur if the diffusion coefficient is very low. In his thesis dated Aug., 1953, Wood" reports on the frequency of bubble formation in a system consisting of oil and rock from the Rangely field. In the present work, the apparatus of Wood was redesigned to allow greater precision, especially at low supersaturations, where bubble growth is slower, and detection is more difficult.

Jan 1, 1958

By J. G. Patchett, R. W. Rausch

Fresh waters and the presence of clay in many Rocky Mountain and West Coast sands require special methods of log analysis. Archie's saturation equation requires addition of a shale correction term, and the SP equation must also be modified to account for clays. Suitable equations were developed several years ago, but have not been widely used due to the algebraic complexity. A computer-oriented method has now been developed to overcome this problem. The basic shaly sand equations are rearranged in four different ways to permit solution for various sets of available input data. Essential to application of the method is the correction of observed SP values to those that would be observed if the resistivity of the formation waters were exactly interchangeable with the activity. A graphic method for doing this is given. Where conditions require consideration of the effect of clay in the sands, the method presented has been found to improve the accuracy of water-saturation determinations. INTRODUCTION Log interpretation in many Rocky Mountain and West Coast basins is complicated by rapid vertical and lateral changes in water resistivity. Calculation of formation water resistivity from the SP curve becomes difficult in zones that contain clay, since changes in SP deflection may be due to changes in either clay content or water salinity. In hydrocarbon-producing reservoirs, the problem is further complicated because hydrocarbon saturation also reduces the SP.1 A log interpretation system using computers has been developed to provide a solution to this problem, based on equations proposed by de Witte.2 Four different simultaneous solutions of de Witte's equations have been made. Each solution method uses a different set of input data as independent variables. Thus, a choice of solution method is possible, depending upon the logs run and the availability of other data. Two of the solutions do not require a knowledge of water resistivity. This system is intended to be used primarily in multiple sandstone-shale sequences of low and moderate resistivities where the principal contaminant in the sandstones is clay. However, where sufficient regional data are available, interpretation in single-zone sandstone reservoirs can also be improved by using the method. THEORY AND HISTORY OF SHALY SAND ANALYSIS The log interpretation formula originally proposed by Archie3 in 1941 is applicable only to rock-fluid systems wherein the rock has negligible electrical conductivity. In 1949, Patnode and Wyllie4 showed that if the rock itself can be considered conductive due to the presence of clay, a different calculation approach is necessary. During the following years, this problem was investigated at great length, as was the related problem of the effect of rock conductivity on the SP.5-11 These investigations established functional relationships between SP, resistivity, water saturation and water resistivity for such a formation. Refs. 2 and 12 provide summaries of these studies. Unfortunately, practical use of these relationships required that water resistivity be known independently from the SP. Although log interpretation methods for rock systems containing clay were proposed at that time,' they were not generally accepted for routine use. There are three principal reasons for this. First, in many field situations involving high-salinity water, rock conductivity may be neglected (even if present) without introducing appreciable error. This may be seen by considering the following expression for waier-saturated rock.' 1/R2=1/R1+1/FRn....(1) where 1/R, is conductivity due to clay. As Rw becomes small, I/FRw becomes much greater than 1/R, which may be neglected. Where 1/R, may be neglected, the sandstone is called clean. If the term may not be neglected, the sandstone is termed dirty or shaly. For resistivity purposes, the classification between clean and shaly sands then depends not only upon the conductivity due to shale in the sand, but also upon the resistivity of the associated water (shale is used here to mean surface condition due to disseminated clay). A sand of given conductivity might safely be treated as clean in association with high-salinity water, but would require shaly sand methods if associated with fresher waters. Shaly sand methods are not required in many areas having saline waters; but in Rocky Mountain and West Coast sands having relatively fresh waters (often more than 0.3 ohm-m resistivity at formation conditions), the shaly sand methods are needed. Errors Rw calculations from the SP due to the presence of shale are likewise related to water salinity. In saline water formations drilled with fresh mud, the ratio of mud filtrate resistivity to water resistivity is high, the SP is large and the presence of shale can introduce large errors in water resistivity calculated by the conventional method. When the resistivity ratio is low, the errors are smaller. At zero SP, no error would result from shale. Thus, from the SP viewpoint, a given rock could be shaly if associated with a saline water, and clean in association with a fresh water, which is the opposite of the resistivity-oriented definition above.

By Harold Margolin

The Ti-Al system certainly merits the investigative attention it has been receiving and this latest contribution by Blackburn is therefore to be welcomed. The titanium-rich end of the phase diagram shown here, as well as the diagram by crossley,17 Clark et a1.,18 and Tsujimoto and Adachi,29 reveal no other intermediate phase than Ti3A1, although a previous work by Ence and Margolin20 had reported an additional intermediate phase. Since then, Buerger Precession pictures of single crystals of 10.5, 11.5, and 12.5 wt pct A1 were made at New York university21 and the extinctions were found to be typical of a D6h a Ti structure. Some superlat-tice spots and/or streaks suggesting the presence of an Mg3Cd structure, i.e., Ti3Al, were also found. Since these specimens had been quenched from the y region20 and had revealed a Ti, it was concludedz1 that y does not exist and the structures interpreted as m + y could not be so considered. The absence of normal two-phase behavior in these structures had been indicated in the paper. Accordingly, the two-phase region y + 6 becomes a + Ti3Al and there is some agreement between the results of Blackburn and of Ence and Margolin20 as to both shape and existence above 900°C of an a + Ti3Al field. Crossley 17 proposed that Ti3Al is a line compound which decomposes congruently at 900°C. It is difficult to discount entirely the fact that clear two-phase structures, varying in amount with position in the two-phase field, could be detected above 900°C by Ence and Margolin.20 Precipitation during quenching could not have been so regular and the twin-phase syndrome17 does not apply. The heat treatment and quenching technique was essentially the same as was used by Crossley, whereas Blackburn, who used a rapid quenching procedure, also showed a two-phase a + Ti3A1 field above 900°C. If the two-phase field exists above 900° C, it is easy to picture a two-phase field which could account for the data of Crossley below 900°C without invoking a line compound existence of Ti3A1. All that would be necessary would be an 0 + Ti3Al/Ti3Al boundary which curved sharply to higher aluminum contents below 900° C. It is clear that the controversy must be solved by techniques which obtain data at elevated temperatures without quenching. It is conceivable that the difference in width of the a + Ti3N fields reported by Blackburn and by Ence and Margolin is due to the longer annealing times used by the latter investigators. The criterion for establishment of equilibrium in our work20 was the

Jan 1, 1969

By W. A. Boyer

MINE hoisting ropes can be loaded to capacity only when the strength of each component is exactly known. Characteristic curves provide this information. When load and rate of acceleration are specified for an individual rope, the characteristic curve assumes a predetermined shape. The factor of safety, is the value drived by dividing the breaking strength of the rope by the total stress on the rope where it goes over the sheave wheel. Total stress is made up of the weight of skip and cage, weight of ore, weight of rope between sheave and rope clevise or attachment, friction load, and acceleration force. Bending stresses are neglected in this discussion, as sheaves and drum di- ameters are assumed to be large enough to reduce these stresses to a negligible value. Fig. 1 gives a family of characteristic curves for 13/8-in. improved plow steel hoisting ropes of 6 x 19 or 6x21 construction with hemp core. The curves will hold for all rope sizes of this specified quality when they are subjected to an acceleration rate of 1.608 ft per sec2. As acceleration rates of most deep shaft hoists are close to this value, this family of curves is presented as a standard. Deviation is slight for other rope sizes. Note that for 0, which is the most important, the deviation is within the width of the line on the graph. In a comparison of values for ropes of 5/8-in. and 21/4-in. diam the deviation ranges from a maximum 1.2 pct at 0 ft to 0.6 of 1 pct at 5000 ft. The value of 5.55 is the same for all ropes at 2500 ft. The tabulated columns to the left of the curves are values of total connected load for each size of rope from 5/8-in. to 21/4-in. diam inclusive. With each of these connected loads the safety factor characteristic curve for the rope will be as indicated.

Jan 1, 1957

By F. S. McNicholas

IN the writer's opinion the term "block caving" has been rather loosely applied to a general principle of mining, and therefore detailed classification of caving methods and draw, together with the characteristics required for each method, will fill a desired need. Selection of caving method should be made only after a study of all governing conditions, such as size of orebody, lateral and vertical extent of ore-body, attitude, treatment demands, possible resultant pressures, and cost and profit comparison of methods. Successful caving methods can be classified under six main headings: (1) Full-protected block caving, (2) semi-protected block caving, (3) unprotected block caving, (4) full-panel caving, (5) half block-panel caving, (6) block-panel caving. The methods and factors governing choice of method are shown in Table I. Full-protected Block Caving Full-protected block caving consists of develop-ing, undercutting, and drawing to exhaustion a block of ore of predetermined size without coordination of mining or draw with adjoining areas. The block is mined and drawn as an independent unit (Fig. 1). Full-protected block caving demands that mining and draw cannot be done against unconsolidated capping in adjoining caved and exhausted blocks. Two sides of block are against solid. Caved capping must be allowed to consolidate for a suitable period, usually not less than one year. Pillars may be left in lieu of consolidated capping (Fig. 2). The advantages of full-protected block, caving are: (1) A definite area is developed, mined, and drawn as a unit so that all work may be completed, the block drawn to exhaustion in the minimum length of time, and the area then abandoned. (? Production is not restrained because of coordina1, on with adjoining areas. (3) Records of draw and caved reserve of ore are good. Upon abanwnment of the block, accountability for production is closed out with no carry-over to adjoining areas. The disadvantages of full-protected block caving are: (1) Side dilution from the draw adjacent to the consolidated capping may be serious, particularly where capping does not consolidate sufficiently in the allotted time. (2) Boundary weakening is needed on the solid sides of the block except in weak rocks. (3) Maximum cantilever and compression effect to aid caving and fragmentation may not be secured since the draw should proceed from the solid to the consolidated capping to minimize side dilution. However, undercutting from the loose (caved capping) to the solid with a slight initial draw in the same manner will develop cantilever effect. (4) A period of at least one year, more is usually desirable, must be allowed for caved capping ,to consolidate before mining the adjoining block, depending upon the characteristics of the capping. (5) Capping may not consolidate sufficiently in the time that can be allowed. Full-protected block caving is indicated where: (1) The ground is weak and heavy and workings

Jan 1, 1952

By E. L. Dougherty, J. W. Sheldon

A numerical method for computing the dynamical behavior of fluid- fluid interfaces is described. Results of studies made to assess the accuracy and economy of the method on a computer are reported. It is concluded that it is practical to obtain digital computer solutions for fluid- fluid interface problems with two space dimensions. PROBLEM DEFINITION We consider two-dimensional flow in a rectangle as shown in Fig. 1, where a fluid-fluid interface divides the rectangle into two regions. The pressure is specified along the lower boundary (the injection boundary) and upper boundary (the withdrawal boundary). There is to be no flow of fluid across the side boundaries, which are walls of symmetry for the flow, and px = 0 along these boundaries. The fluids are incompressible, the absolute permeability and porosity is constant, and each fluid has its own constant mobility. Thus, the pressure satisfies Laplace's equation in each region. Across the interface, pressure is continuous and the normal component of fluid flux is 1. These conditions are shown in Fig. 1. The velocity of a point on the interface in the direction of the normal from Fluid 1 into Fluid 2 is U,/ Q, where Un is the normal component of volumetric flux at the point on the interface, i.e., For brevity, we call the problem defined above "Problem A". Problem A is a simple example of a dynamical fluid-fluid interface problem. However, the same numerical methods which we use to solve Problem A can be extended to the solution of a large number of fluid-fluid interface problems which are mathematical models for various oil recovery processes,l hence our interest in Problem A. We define the interface by prescribing the x, y coordinates of a finite set of points. These points are called "interface mesh points" and are illustrated by triangles or asterisks in the figures. Fig. 2 is a procedural (flow) chart displaying the main steps in the solution. We comment briefly on some of these steps in this section and then give further details in subsequent sections. For mobility ratio greater than one it is well known that Problem A is unstable.*2 Small amplitude,

Jan 1, 1965

By W. F. Baldwin

The timing circuit generally employed in Carter Reservoir Behavior Analyzers consists of a synchronous motor driven cam which actuates a microswitch. A mechanically actuated timer of this type has several inherent disadvantages, among which are the following: 1. The first time interval may not be of uniform length, depending on the exact position of the cam at the conclusion of a previous run. 2. The positioning adjustments on the microswitch lever arms are critical and require periodic readjustment. 3. The mechanical system is subject to wear which will ultimately impair its accuracy. It was therefore considered desirable to devise an alternate method which would eliminate these difficulties without introducing others that would be equally troublesome. This note describes the design of an electronic timing device which has been incorporated in the analyzer constructed by the Field Research Laboratories of Magnolia Petroleum Co. The electronic timer circuit is built around a simple neon glow tube relaxation oscillator, the complete circuit diagram of which is shown in Fig. l. The operation of the circuit is as follows: Condenser C, is charged to the critical firing voltage of the neon glow tube VR-75 by battery B1 through resistors R1 and R2. When the neon glow tube fires, the condenser C, is discharged, The charging cycle then repeats itself,

Jan 1, 1950

By George C. Fryburg

The oxidation of platinum at high temperatures (above 800°C) is controlled by boundary-layer diffusion, except at the lowest pressures. The rate of oxidation is determined by the rate of diffusion of oxide through the boundary layer away from the platinum. This is in contrast to the situation in tungsten and molybdenum oxidation, where the rate is controlled largely by the diffusion of oxygen through the boundary layer towards the metal. The difference in mechanism is a consequence of the much lower rate of oxidation of platinum. Evidence is presented that supports this mechanism for the oxidation of platinum. SOME years ago we reported1 a study of the oxidation of platinum in ordinary oxygen in the temperature region above 800°C, where the oxide is volatile. All the data available up to that time had been obtained with platinum crucibles or plates heated in tube furnaces. The results were widely variant and highly dependent on the flow rate of gas through the furnace: the faster the gas flow rate, the higher the measured oxidation rate. These results indicated to us2 that reaction equilibrium was being approached in the furnace and that the rate was diffusion limited. To minimize these effects, we employed1 electrically heated ribbons placed in a cooled bulb. We were primarily interested in the low-pressure region, but some experiments were performed at pressures up to 1 atm. We found a peculiar pressure dependency which was explained on the basis that the rate of oxidation was proportional to the pressure at all pressures, but that with increasing pressure an increasing fraction of the volatilizing oxide was back-reflected to the ribbon by the surrounding gas molecules. The oxide molecules are not stable at the temperature of the platinum and are decomposed on impact, the platinum being re-deposited on the specimen. Because of the limitations of space, this explanation was offered summarily without detailed support. Recently, Jaffee and associates3, 4 have suggested that the pressure dependency of our data up to 0.5 torr could be explained on the basis of a Langmuir-type adsorption of molecular oxygen followed by dissociation. Since the correct explanation of the effect of pressure on the oxidation rate is a necessary requirement to the development of the basic mechanism of the oxidation reaction, we felt that the back-reflection theory should be presented in more detail. This is the object of the present paper. RESULTS AND DISCUSSION A typical set of results showing the pressure dependency of the oxidation reaction is presented in Fig. 1, where we have plotted, on logarithmic scales, the rate of oxidation, R, vs the pressure, p. For comparison some results of Krier and Jaffee4 and of Raub and plate5 are also shown. These data were taken at atmospheric pressure using furnaces. The effect of reaction equilibrium within the furnace is apparent from the much smaller values that these observers obtained: Raub and Plate In O2 found a rate of 3.1 µg cm-2 hr-1 compared to our value of

Jan 1, 1965