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By David P. Shoemaker, Clara B. Shoemaker

It has been shown for an important class of complex transition intermetallic compounds (a, P, R, 6, and p phases) characterized by "normal" coordination [CN12 (icosahedral), CN14, CN15, CN16/ that interatomic distances nay be calculated to a good approximation as the sum of characteristic atomic radii. Two radii, one for major ligands and one for minor ligmds, are specified for each atom, except in the case of CN12 where only a miaaor-ligand radius is specified. The same appears to be true of transition-metal phases of simpler struc-ture: Laves phases (CN12, CN16), and p-tungsten phases (CN12, CN14). In the case of known examples of the more complex phases, a simple rule is given which specifies these radii. However, only a fraction of the known examples of the simpler phases obey this rule closely. To include the latter phases the rule may be modified by considering the radii as linear functions of the weighted average of the Pauling CN12 radii of the two kinds of atonzs, with the radii weighted according to the over-all chemical composition of the alloy. With very few exceptions interatomic distances for both tlze complex and the simpler transition phases can b$ predicted with this modified rule to within 0.06A. ManY intermetallic compounds are known of composition A,By, in which A is a transition element to the left of the manganese column in the periodic table and B is a transition element in or to the right of it. Frequently the coordination numbers (CN) found in these compounds are CN12 (icosahedral), CN14, CN15, and CN16 (called "normal" coordinations by Frank and Kasperl). Well-known examples are the cubic and hexagonal Laves phases which have CN12 and CN16, and the 0-tungsten (CrsO) phases which have CN12 and CN14. In the more complicated (often ternary) phases, such as the a phase,2 the Beck phases p3 and R~, the 6 phase,5 and the p p atoms occur with CN12, CN14, CN15, and (except for a) CN16; in many cases several crystallographically independent atoms of one particular CN occur in the asymmetric unit. A large number of independent interatomic distances are found in these complicated phases, varying from 20 in the a phase to 94 in the 6 phase. These distances show a large spread; they vary, for example, from 2.358 to 3.278A in the 6 phase. In our analysis of these distances we found that in each of these compounds every atomic position can be characterized by either one or two radii. The CN12 positions are characterized by a single radius, The higher coordinated positions are characterized by two radii, namely: the CN14 positions by 4 in the direction of the twelve "5-coordinated" ligands3 (called 'minor" by Frank and Kasperl) and by r:, in the direction of the two "6-coordi-nated" ligands (called "major" by Frank and Kasper); the CN15 positions by r15 for the twelve minor and r:, for the three major ligands; the CN16 positions by rlE for the twelve minor and r:, for the four major ligands. We have expressed the experimentally determined interatomic distances in observational equations as the sums of the appropriate pairs of these characteristic radii and the value of these radii have been determined by the method of least Squares. Despite their wide range, the interatomic distances could then be predicted by the sums of these atomic radii with an average deviation in any one compound of 0.06A or less. The results are summarized in Table I. Inspection of the radii thus obtained shows that in the structures in Table I the radii (in A) are given to a first approximation by the simple relationship: Where CN is the coordination number (12, 14, 15, or 16), and A = 1 for major ligands and = 0 for minor ligands. The interaLomic distances can be predicted within about 0.1A by sums of these atomic radii. Another phase belonging in this group with CN12, 14, 15, and 16 is the y phase & B7, in which A is molybdenum or tungsten and B is iron or cobalt. Recently the M%C phase has been refinedE and the observed distances also agree well with those calculated with Eq. [I]. (In the original determination of the structure of W6FeV7 the F$(II)-W(II1) distance was erroneously given as 2.84A, but we have recalculated it fro? the published parameters and found it to b? 2.57i4, in good agreement with the value of 2.6A predicted with Eq. [I.].) Many binary transition alloys are known to crystallize with the simpler structures having "nor-

Jan 1, 1964

By H. E. Cook

The Ising model for the internal energy of a binary alloy has been used to obtain a general equation for the critical resolved shear stress of partially ordered crystals. The equation expresses the stress in terms of the Warren "alphas " and can be used to estimate the variation in strength with order without the assumption, present in the original formulation of this problem in terms of domain size, that order is complete within each domain and that the domains are of ungorm size and shape. In addition, it is the general equation, according to the Ising model, for strengthening by short-range order. Two applications of the equation are considered: One is an estimate of the variation in strength of CkAu with long-range order. The other is an estimate of the variation in strength of FeCo with quench temperature. Reasonable agreement is found with the variations reported in the literature. When the internal energy of an alloy crystal depends upon the distribution of solute, the strength of the crystal will also depend upon it because a portion of the applied stress for plastic deformation will be: where V is the volume of the crystal and E(E) is the energy change associated with the solute redistribution caused by the plastic strain, E. We expect T to equal zero for a crystal having a random arrangement of solute because the arrangement would remain random after plastic deformation. Likewise, we expect it to equal zero when the crystal is perfectly ordered because the motion of paired dislocations found in such crystals does not disrupt order. However, when short-range order exists or when long-range order is incomplete, plastic deformation will decrease the amount of order and additional work, proportional to the ordering energy, will be expended. Fisher' estimated T for crystals having short-range order by assuming an interaction energy between neighboring atoms and estimating the change in the number of unlike neighbors as a dislocation moved through the crystal. (His analysis was limited and several workers2"6 have since given more complete ones.) Fisher minimized the importance of a strengthening mechanism of this type for paired dislocations in a structure having long-range order. ~ottrell,' however, pointed out that T could be appreciable for ordered crystals having antiphase domains. He attributed the strengthening to the increase in surface energy of the domains as they were cut by paired dislocations. Ardley,' in his test of Cottrell's theory, found that r for Cu3Au crystals obeyed the equation: for 1 > t where 1 is the domain size, t is the domain wall thickness, and y is the surface energy of an antiphase boundary. His experiments represent the classic confirmation of the strengthening mechanism proposed by Cottrell. However, the assumptions involved in using Cottrell's theory are valid only for large domain size in CU~AU,~"~ i.e., when Eq. [2] reduces to: For small domains, ~linn~ has questioned Ardley's assumption that order was complete, and, indeed, Stoloff and ~avies" fpnd it incomplete until a size of approximately lOOA was reached. Even when the order within a domain is complete, it is not obvious how one determines the appropriate value for I in a structure where domains vary in size and are irregularly shaped. The purpose of this paper is to estimate T without restrictions upon the degree of order and domain shape. Our major assumption will be the use of a generalization of the model proposed by Bethe" (the Ising model) for the internal energy. This will in fact allow us to combine the theories for strengthening by short-range order and by antiphase domains into a single, general formalism. We will use the results to estimate the variation in strength of Cu3Au crystals with long-range order8 and the variation in flow stress of FeCo crystals with quench temperature.12'13 INTERNAL ENERGY For simplicity, we restrict our considerations to those binary solid solutions which can be described as an arrangement of atoms on a Bravais lattice. An atom site will be indexed by three numbers (PI, pz, p3) determined by the vector: from the origin fixed at atom (0, 0, 0) to the atom site where a', an, and a, are the lattice translation vectors. We write: For the energy of the crystal where pi(p) is the probability (either zero or one) of finding an atom of type i (i = 1, 2) at site (p), which is shorthand for (pl, pz, p3) and pj(p + r) is that for an atom of type j (j = 1, 2) at the site (p + r), which is shorthand for (pl + rl, The coupling parameter, resents the energy associated with the pair Pi(p), Pj(p + r). The crystal is assumed large enough so that surface effects can be neglected; therefore, trans-lational and inversion symmetry require the coupling parameters to obey the relations

Jan 1, 1969

By G. M. Schwartz, D. M. Davidson

THE Duluth gabbro outcrops containing sulphides of copper, nickel, and iron are located on both sides of State Highway No. 1 an airline distance of 8.5 miles southeast of Ely in northeastern Minnesota. The region of known sulphide occurrences includes parts of sections 5, T. 61 N., R. 11 W., and parts of sections 25, 26, 32, 33, and 34, T. 62 N., R. 11 W. These sections, given in Fig. 1, are all in Lake County, Minnesota. Part of the area, which lies entirely within the Superior National Forest, is shown on the topographic map of the Ely quadrangle. The original discovery was made in 1948 when a small pit was opened in weathered gabbro rubble for use on a forest access road. A shear zone had caused unusual decomposition in this glaciated area, and the resulting copper stain was noted by Fred S. Childers, Sr., an Ely prospector, who began searching the outcrops along the base of the intrusive. He was joined in further exploration by Roger V. Whiteside of Duluth. In the summer of 1951 a small diamond drill was moved into the area and a hole 188 ft deep was drilled, passing through 11 ft of glacial drift into sulphide-bearing gabbro. This paper is a preliminary report on the geology of the newly discovered ore. The Duluth gabbro is one of the largest known basic intrusives and may be defined as a lopolith.1 It extends northeastward from the city of Duluth as a great crescent-shaped mass that intersects the shore of Lake Superior again near Hovland, 130 miles to the northeast, see Fig. 2. The distance around the outside of the crescent is nearly 170 miles. The form of the intrusive is simple at Duluth where it ends abruptly north of the St. Louis River; at the east end, however, the gabbro splits into two elongated, sill-like masses separated mainly by lava flows and characterized by minor irregularities. The outcrop reaches a maximum width in the central part where it is about 30 miles across, and a maximum thickness of about 50,000 ft. It may be significant that the sulphides occur at the base of the thickest part. The lopolith has segregated into rock types ranging from peridotite to granite. The most abundant types are olivine gabbro, gabbro, troctolite, anorthosite, and granite. Of lesser importance quantitatively are peridotite, norite, pyroxenite, magnetite gabbro, and titaniferous magnetite. Grout estimates that two-thirds of the gabbro at Duluth is olivine gabbro. Variations in the percentages of plagioclase, augite, olivine, and magnetite-ilmenite constitute the only essential differences found among the basic rock types. The predominant mineral is plagioclase, mainly labradorite. Plagioclase and olivine seem to have crystallized early, and the olivine rich rocks, usually troctolite, are found in the lower part. Segregations of titaniferous magnetite are abundant near the base of the gabbro along the eastern part and also occur far above the base. These have recently been described in detail by Grout' Near the top, segregation has produced a gradation to granite, or "red rock," as it is known locally. This consists of quartz, red feldspar, and hornblende. The red rock forms a. zone with a maximum width of nearly 5 miles but is quantitatively unimportant from Duluth northward for 35 miles. In Cook county, where the gabbro splits, each of the two sill-like masses has a red rock top somewhat thicker in proportion to the gabbro below than in the main central mass. The intrusive ranges from coarse to medium in grain size and from granitoid to diabasic in texture. Throughout much of the Duluth gabbro in Minnesota banding and foliation are well developed, as Grout has emphasized! The bands are mainly a result of variation in the percentage of minerals, as in troctolite with alternating bands high in olivine and in plagioclase. A few bands may consist largely of one mineral, as is true of some segregations of magnetite. Many of the banded rocks show a clearly developed parallelism of platy plagioclase crystals, and both banding and foliation are believed to conform to the floor of the lopolith. Throughout its extent in Minnesota the Duluth gabbro dips east and south toward Lake Superior. It is generally believed to extend beneath Lake Superior and is found as a smaller mass exposed along the north side of the Gogebic district in Wisconsin and Michigan. The dip at and near the base ranges along most of its length from 20 to 40°, but at places the internal banding dips even more steeply. The dip of the upper part is much less, and if it is assumed that the flows along the north shore of Lake Superior are a dependable indication, it does not exceed 15º. The formations shown in Table I which are intruded by the gabbro range from Keewatin to Middle Keweenawan in age. They present a significant picture. At the top, the gabbro and its accompanying

Jan 1, 1952

By G. M. Schwartz, D. M. Davidson

THE Duluth gabbro outcrops containing sulphides of copper, nickel, and iron are located on both sides of State Highway No. 1 an airline distance of 8.5 miles southeast of Ely in northeastern Minnesota. The region of known sulphide occurrences includes parts of sections 5, T. 61 N., R. 11 W., and parts of sections 25, 26, 32, 33, and 34, T. 62 N., R. 11 W. These sections, given in Fig. 1, are all in Lake County, Minnesota. Part of the area, which lies entirely within the Superior National Forest, is shown on the topographic map of the Ely quadrangle. The original discovery was made in 1948 when a small pit was opened in weathered gabbro rubble for use on a forest access road. A shear zone had caused unusual decomposition in this glaciated area, and the resulting copper stain was noted by Fred S. Childers, Sr., an Ely prospector, who began searching the outcrops along the base of the intrusive. He was joined in further exploration by Roger V. Whiteside of Duluth. In the summer of 1951 a small diamond drill was moved into the area and a hole 188 ft deep was drilled, passing through 11 ft of glacial drift into sulphide-bearing gabbro. This paper is a preliminary report on the geology of the newly discovered ore. The Duluth gabbro is one of the largest known basic intrusives and may be defined as a lopolith.' It extends northeastward from the city of Duluth as a great crescent-shaped mass that intersects the shore of Lake Superior again near Hovland, 130 miles to the northeast, see Fig. 2. The distance around the outside of the crescent is nearly 170 miles. The form of the intrusive is simple at Duluth where it ends abruptly north of the St. Louis River; at the east end, however, the gabbro splits into two elongated, sill-like masses separated mainly by lava flows and characterized by minor irregularities. The outcrop reaches a maximum width in the central part where it is about 30 miles across, and a maximum thickness of about 50,000 ft. It may be significant that the sulphides occur at the base of the thickest part. The lopolith has segregated into rock types ranging from peridotite to granite. The most abundant types are olivine gabbro, gabbro, troctolite, anortho-site, and granite. Of lesser importance quantitatively are peridotite, norite, pyroxenite, magnetite gabbro, and titaniferous magnetite. Grout estimates that two-thirds of the gabbro at Duluth is olivine gabbro. Variations in the percentages of plagio-clase, augite, olivine, and magnetite-ilmenite constitute the only essential differences found among the basic rock types. The predominant mineral is plagioclase, mainly labradorite. Plagioclase and olivine seem to have crystallized early, and the olivine rich rocks, usually troctolite, are found in the lower part. Segregations of titaniferous magnetite are abundant near the base of the gabbro along the eastern part and also occur far above the base. These have recently been described in detail by Grout.' Near the top, segregation has produced a gradation to granite, or "red rock," as it is known locally. This consists of quartz, red feldspar, and hornblende. The red rock forms a zone with a maximum width of nearly 5 miles but is quantitatively unimportant from Duluth northward for 35 miles. In Cook county, where the gabbro splits, each of the two sill-like masses has a red rock top somewhat thicker in proportion to the gabbro below than in the main central mass. The intrusive ranges from coarse to medium in grain size and from granitoid to diabasic in texture. Throughout much of the Duluth gabbro in Minnesota banding and foliation are well developed, as Grout has emphasized.V he bands are mainly a result of variation in the percentage of minerals, as in troctolite with alternating bands high in olivine and in plagioclase. A few bands may consist largely of one mineral, as is true of some segregations of magnetite. Many of the banded rocks show a clearly developed parallelism of platy plagioclase crystals, and both banding and foliation are believed to conform to the floor of the lopolith. Throughout its extent in Minnesota the Duluth gabbro dips east and south toward Lake Superior. It is generally believed to extend beneath Lake Superior and is found as a smaller mass exposed along the north side of the Gogebic district in Wisconsin and Michigan. The dip at and near the base ranges along most of its length from 20 to 40°, but at places the internal banding dips even more steeply. The dip of the upper part is much less, and if it is assumed that the flows along the north shore of Lake Superior are a dependable indication, it does not exceed 15". The formations shown in Table I which are intruded by the gabbro range from Keewatin to Middle Keweenawan in age. They present a significant picture. At the top, the gabbro and its accompanying

Jan 1, 1953

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually average about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment. Surface deposits contain sulphur in the same range of concentrations as the salt dome deposits, i.e., from 10 to 50 pct sulphur, associated with various gangue materials such as silica, limestone, and gypsum. The pirincipal distinction, then, does not lie in the percentage of sulphur contained in the ore, but in the geological nature of the deposit. A recent study' of the world sulphur supply situation estimated 1950 sulphur production in the free world countries at 5.6 million long tons, of which the United States produced 5.2 million tons, or 93 pct of the total. While America's domestic needs alone would have been covered by national production, about 1.4 million tons were exported during the same year. Despite all the steps which are being taken to restrict use of elemental sulphur and to force the fullest possible development of alternate sulphur sources here and abroad, the deficit in elemental sulphur production will probably increase with time. As a result of intensive prospecting for oil throughout the Gulf Coast area discovery of significant new salt domes is held unlikely. With the growing scarcity of sulphur and what appears to be an inevitable rise in price, recovery from deposits not amenable to Frasch-process mining assumes greater economic importance. Untapped Reserves The most important deposits in this category are located in Sicily, where elemental sulphur occurs in Miocene limestone and gypsum formation. Sulphur content of these ores ranges from 12 to 50 pct with an estimated average of 26 pct. Although quantitative estimate of these reserves is not available it is held that they exceed 50 million tons of sulphur. Similar deposits occur also on the mainland which contribute about one-third of Italy's total current annual production of 230,000 tons, but these are known to be nearing exhaustion. Significant surface deposits of volcanic origin are located in South America, Japan and western United States, silica being characteristic gangue con-stituent. The largest of these deposits are in South America. More than 100 extend over a zone 3000 miles long, paralleling the west coast of South America. 'Total sulphur content of these deposits has been estimated to be as high as 100 million tons. The main islands of Japan also possess at least 40 known volcanic sulphur deposits with probable reserves of 25 to 50 million tons.' Prospected reserves in western United States might amount to 2 million long tons, principal deposits being located in the northwestern part of Wyoming, southern Utah, and eastern California. Volcanic deposits of lesser importance are found around the Mediterranean, in Turkey and Greece, and in Africa, Egypt, Abyssinia, and Somaliland. Beneficiation Methods Different methods of beneficiation have been used in these various locations. In Italy the Calcarone kiln and Gill regenerative furnaces are used exclusively. Both utilize heat liberated by burning part of the sulphur in the ore to liquify or vaporize the remaining sulphur, which is recovered by solidification or condensation. The Calcarone kiln is of conical shape, generally 35 ft in diam at base and 18 ft high. A kiln of 25,000 cu ft capacity burns for about two months and yields about 200 tons of sulphur. The Gill furnace consists of a series of chambers with domed roofs. Sulphur is burned and melted in one chamber at a time and the hot combustion gases are used to preheat the ore charge in the subsequent cell. These furnaces operate on a cycle of 4 to 8 days. The recovery yield of both systems is about 65 pct. Sulphur losses amount to 25 pct through the combustion to sulphur dioxide; about 10 pct is retained in discarded calcines. Ores containing less than 20 pct are not considered suitable as furnace feed. These methods are not only wasteful because of the low recovery obtained, but represent a serious atmospheric pollution problem. Sulphur produced ranges from 96 to 99 pct purity and thus does not match Texas or Louisiana sulphur. Owing to the present shortage, sulphur in the Middle East sells

Jan 1, 1954

By S. E. Szasz, E. J. Moore, B. F. Whitney

An accurate value of formation water resistivity R, is essential in calculating formation porosity and fluid saturation from electrical well logs. In the cases where R, has not been measured directly, it must be obtained from other data, e.g., the SP curve. This paper deals with another approach: how to calculate R, from the chemical analysis of the formation water. INTRODUCTION It is known that the resistivity of aqueous solutions of pure salts depends on their concentration and on the temperature; the concentrations are given in MPL (mg of solute per liter of solution), or sometimes in ppm (mg of solute per kg of solution): MPL = ppm X specific gravity. Values for different pure salts are available in the literature, but not for solutions of mixtures which are of practical interest. The major component of the dissolved material in almost all formation waters being sodium chloride, it is customary to express the resistivity of formation waters in terms of equivalent sodium chloride concentration, i.e., the concentration of a solution of pure NaCl which has the same resistivity at a given temperature as that of the formation water under consideration. Thus, the problem of calculating R, from the chemical anaylsis can also be stated as how to convert the other constituents of the solute into equivalent NaCl concentration. Salts dissolved in water are at least partly dissociated into ions, and do not conserve their identity. If known amounts of several salts are dissolved in water, the solution does not necessarily contain the same salts in the original proportion, but perhaps some other combination of the ions, along with free ions in solution. This is why the chemical analysis of formation waters is often given in terms of ions, as if all dissolved salts were completely dissociated. Our problem then boils down to how to convert the concentrations of the various ions to equivalent concentrations of Na' and C1-. Dunlap and Hawthorne' have proposed to convert the concentration of all other ions to equivalent Na' and C1-concentrations by means of constant multipliers; e.g., 0.95 for Ca"; 2.0 for Mg"; 0.27 for HCO 3-; 0.5 for SO, -, etc. Their factors were based on measurements made at 68F on 26 formation water samples from the Texas Gulf Coast, ranging in concentration from 1,500 to 75,000 ppm. The Dunlap method is widely used in electric log interpretation, and is often extrapolated beyond its original concentration range. A comparison of R, values calculated by this method and values actually measured on formation water samples has shown large discrepancies, especially at higher concentrations. Therefore, two new methods were developed at Sinclair Oil Corp.'s Tulsa Research Center to calculate equivalent sodium chloride concentration from the chemical analysis of formation water samples. FUNDAMENTAL CONSIDERATIONS The resistivity of a solution, or its reciprocal the conductivity, at a given temperature is determined by the charge, concentration and mobility of the ions actually present. Monovalent ions such as Na' or C1- always carry the same charge. Compounds of polyvalent ions, however. may show incomplete dissociation, e.g., NaSO; + Na' instead of SO,-- + 2Na'. This happens especially in more concentrated solutions. Only very dilute solutions are completely dissociated, as assumed in the chemical analysis report. At higher concentration, the degree of dissociation depends not only on the nature and concentration of the particular salt under consideration but also on the nature and concentrations of the other solutes. Mobility of the ions depends on the viscosity of the solution. It also depends on the degree of hydration of the ions, which in turn is a function of the nature and the charge of the ions and also of the amount of free water available per ion, i.e., the total ionic concentration. The net effect is that the conductivity increases slower than proportional to the concentration, even if a solution contains only one salt such as NaC1, and is different for different salts (Fig. 1). Conductivity can even decline with a further increase in concentration, e.g., if additional salt is little dissociated but ties up some of the free water and/or causes an increase in viscosity. In solutions containing more than one salt, the contribution of one salt to the total conductivity depends not only on the fractional concentration of this same salt, but also on the concentration of all other solutes. A perfect method would give the conductivity or resistivity of a solution as a function of the concentrations of all solutes present. This is so complicated as to be impractical, and a simpler method must be found which is of acceptable accuracy. The Dunlap method, on the other hand. is too simple because it askmes that at any concentration the contrih-

Jan 1, 1967

By Niels Hansen

The substructure formed by drawing at room temperature in aluminum (99.5 and 99.998 pct purity) and in recrystallized aluminum aluminum-oxide products containing from 0.2 to 4.7 wt pct of aluminum -oxide was examined by transmission electron microscopy, and the flow stress of the drawn materials was measured by tensile testing at room temperature. A sub-grain structure was present after a reduction in area by drawing of 10 to 20 pct, and the subgrain size was observed to decrease with increasing deformation. The tensile data show that the increase in flow stress (0.2 pct offset) by drawing from 10 to 95 pct depends on the reduction in area, not on the composition of the materials. Dispersion strengthening and subgrain bowzdary strengtlzening contribute to the flow stress, and these strengthening processes have been found to be linearly additive. The flow stress (0) can be related to the subgrain size &) by the Petclz relation = uo + k . t, where go is dependent on the composition of the products and k is approximately the same far all materials. THE microstructure of dispersion-strengthened aluminum aluminum-oxide products consists of small oxide plates distributed in an aluminum matrix. The matrix structure depends on the manufacturing history, and in hot-worked as well as cold-worked products the matrix is divided by subgrain (or dislocation) boundaries. For hot-extruded products it has been shown1 that dispersion strengthening and subgrain boundary strengthening are linearly additive, and the flow stress (0.2 pct offset.) at room temperature has been related to the subgrain size (ts) by a Petch equation,2,3 s = so + k . ts-1/2, where s0, increases with increasing oxide content. For cold-worked products containing subgrains no systematic work has been reported, and it was the aim of the present study to examine the microstructure and the relationship between the flow stress and the subgrain size for such products. The behavior of' aluminum aluminum-oxide products depends on the purity of the aluminum matrix, and aluminum of the matrix purity (99.5 pct) was included in the investigation. The literature contains few data about the behavior of this impure aluminum, and aluminum of a higher purity (99.998 pct) was therefore also examined. As regards the relationship between the flow stress and the subgrain size in cold-worked dispersion-strengthened products, no systematic work has been reported. For aged cold-worked structures containing fine precipitates (Fe-Mo carbide) a Petch relation has been found,4 and it has been shown that the k value NlELS HANSEN is Head, Metallurgy Department, Danish Atomic Energy Commission, Research Establishment Riso, Denmark. Manuscriot submitted January 9, 1969. IMD is approximately the same as in iron, whereas the s0 value is higher owing to the presence of the precipi-tates. Investigations of metals such as tungsten,' ferrous metals,4,6 and molybdenum7 cold drawn or swaged at room temperature have shown that the flow stress can be related to the subgrain size by a Petch relation when ts is taken as the subgrain size perpendicular to the direction of deformation. For aluminum no work has been reported on the relationship between the flow stress and the subgrain size after deformation at room temperature, whereas for aluminum tensile strained at different temperatures in the range -183" to 375°C a Petch relation has been found by taking ts equal to the subgrain size.' In the present study two aluminum materials (99.998 and 99.5 pct) and three aluminum aluminum-oxide products (containing 0.2, 1.0, and 4.7 wt pct oxide) were drawn at room temperature to reductions in area from about 10 to about 95 pct. The structures were studied by transmission electron microscopy, and the flow stress (0.2 pct offset) was measured at room temperature. EXPERIMENTAL Materials. The materials are given in Table I together with the chemical analysis. The three aluminum aluminum-oxide products were manufactured from aluminum powder that had been compacted and Table I. Chemical Analysis of Materials Al203 Fe SI Material wt pct wt pct wt pct 99.998 pct* - 0.0004 0.0012 Aluminum 99.5 pctt - 0.36 0.16 Aluminum AlMD13† 0.2 0.16 0.12 Aluminum- -Oxide AlMD105† 1.0 0.26 0.18 Products SAPISML960s† 4.7 0.22 0.19 *Other impurities: O.0004 pct max each of Cu and Zn (supplier's analysis). †Other impurities: 0.03 pct max Cu. 0.02 pct max each of Mn, Mg, Zn, Ti. Table 11. Mean Diameter of Aluminum-Oxide Particles in Extruded and in Cold--Drawn Aluminum Aluminum-Oxide Products Mean Diam. of A1203-Plates* Material State A AlMD105 Extruded 540 Cold Drawn 97 pct 510 SAP ISML 960 Extruded 770 Cold Drawn 95 pct 820 *The standard deviation of the mean is approx. 5 ±pct.

Jan 1, 1970

By Robert A. Rapp, Ronald L. Pastorek

Solid-state electrochemical measurements by three alternative experimental procedures were made with the cell FeO, Fe3O4 |Zro.85Cao.15O1.85 |Cu| Zr0.85CaO.15O1.85 | FeO, Fe304 to establish the solubility and diffusivity of oxygen in solid copper in the temperature range 800" to 1030°C. The solubility of oxygen in solid copper and the diflusivity of oxygen in solid copper Dgu = 1.7 X 10-2 exp(-16,000/RT) Cm2/sec were determined and confirmed in alternative experiments. The enthalpy of solution of oxygen in solid copper equals —10,000 cal per mole; the partial excess entropy of the oxygen atoms in the Cu-O dilute solution is approximately the same as that found for interstitial atoms in other metals. The diffusivity of oxygen in solid copper is consistent with that expected for an interstitial atom. RELIABLE values for the saturation solubility N(s) and diffusivity DO of oxygen in solid copper have not been unambiguously established in the literature. Following three early determinations by others,1"3 Rhines and Mathewson4 reported that the solubility of oxygen in solid copper increased from 0.007 at. pct 0 at 600°C to about 0.015 pct at 1050°C. Phillips and skinner,, using essentiially the same analytical procedure, reported that the solid solubility increases from 0.0018 at. pct 0 at 550°C to about 0.0075 pct at 1050OC. The only previous value for the diffusivity of oxygen in solid copper was reported by Ransley.6 Ransley deoxidized Cu-Cu2O alloys in an atmosphere of carbon monoxide gas to yield a solubility-diffusivity product. He used the solubility data of Rhines and Mathewson to calculate the diffusivity values. Another method for obtaining the solubility-diffusivity product (N(s) DO) is by measuring the widths of internal-oxidation zones in copper alloys as reported by Verfurth and Rapp.7 However, the calculated N(S)Do products depend upon the alloy content of the specimen, so that the internal oxidation of copper alloys does not follow ideal internal oxidation kinetics. As a result, unequivocal values for the N(s) DO product were not obtained by this procedure. A solid-state coulometric titration technique similar to that employed in this work was introduced by C. Wagner8 to study the dependence on silver activity of the Ag/S ratio in silver sulfide in the temperature range of 160" to 300°C. Similar experiments have been carried out by C. Wagner and co-workers9-11 to study the stoichiometry range of silver and copper tellurides, cuprous sulfide, and cuprous selenide. Numerous authors have carried out electrochemical measurements with a solid oxygen-ion-conducting electrolyte to determine the solubility and/or diffusivity of dissolved oxygen in several liquid metals.12-l6 Rickert and Steiner17,18 have used solid-state electrochemical measurements to determine the diffusivity of oxygen in solid silver from 760" to 900°C. Two different cell geometries were used. In the cell of linear geometry Fe, FeO | ZrO2 + (CaO) | Ag + [0 (dissolved)] [1] oxygen diffused from the interior of the silver electrode to the silver/electrolyte interface where the oxygen activity had been lowered from a fixed initial value to practically zero by the application of voltage to the cell. The diffusivity of oxygen in solid silver was determined from the solution of the diffusion equation and the time dependence of the cell current. However, this determination of the diffusion coefficient depended upon a knowledge of the solubility of oxygen in solid silver. A cylindrical geometry was used for the cell Pt, O2(Po2 = 0.21 atm) | ZrO2 + (CaO) | Ag + [0 (dissolved)] [II] which also allowed the diffusivity of oxygen in solid silver to be determined. These values were in agreement with other available data.l9 Recently, Raleigh20,21 used a method involving the measurement of diffusion-limited currents in a cell involving the AgBr solid electrolyte to determine the diffusion coefficient of silver in Ag-Au alloys at 400°C. Diffusivity values on the order of l0-14 sq cm per sec were measured in the alloy composition range 10 to 60 at. pct Ag in a single experiment. From numerous electrical conductivity and galvanic cell measurements,9'22"26 the solid solution Zr0.85 Ca0.15 O1.85 has been established as an electrolyte with predominant oxygen ion conduction over a wide range of intermediate and high oxygen activities. For interrelating the thermodynamics and the kinetics of the dissolution of oxygen in solid copper in this investigation, a galvanic cell was constructed with FeO-Fe3O4 as the reversible reference electrode, the Zr0.85Ca0.15 O1.85 electrolyte, and a pure copper specimen under-saturated in oxygen as the other electrode. THEORETICAL ANALYSIS Three variations of a high-temperature electrochemical technique were used in this study to provide two determinations each of the solubility and diffusivity

Jan 1, 1970

By J. W. Powell, F. H. Spedding

A complete review of the use of chelating agents in the sepa ration of rare earths by ion-exchange is given as well as a concise description of the recent pilot-plant operations of the Ames Laboratory. The two chelating agents which show the greatest promise are ethylenediamine-N,N,N',N'-tetraacetic acid and N'-(2-hydroxyethyl)ethylenediarnine-N,N,N '-triacetic acid. The first successful separations of rare earths by ion exchange were reported in a collection of papers which appeared in the November, 1947, issue of the Journal of the American Chemical Society.1'9 Some of this work was performed at the Ames Laboratory of the A.E.C., the remainder at the Oak Ridge National Laboratory. The processes developed at Oak Ridge, as well as some of the early Arnes methods, employed 5 pct citric acid-ammonium citrate eluant at low pH and were carried out on either H+-state or NH:-state resin beds. These techniques were successful for the separation of small quantities of either naturally occurring or radioactive rare earths and are still used for the isolation of rare-earth activities from fission products. Concentrated citrate is not economical, however, for use in moderate or large-scale rare-earth separations. For this reason, the Ames Laboratory turned its attention to lower concentrations and higher pH's in order to make more effective use of the eluting agent.10-14 Although elutions have been performed successfully over a wide range of conditions, 0.1 pct citrate at a pH of 8.0 is highly recommended for use on H+-state resin beds.15,16 Elu-tion with 0.1 pct citrate in the pH range from 5 to 9 brings about the separation of the constituent rare earths into a series of flat-topped elution bands which progress down the resin bed, head to tail, without actually pulling apart as do the rare-earth peaks which develop when trace quantities are eluted with 0.25M (5 pct) citrate at low pH's. Because of this, elution with 0.1 pct citrate at pH 8.0 do not produce pure rare earths unless sufficient quantities are present to provide developed bands which are at least several inches long on the columns. Although other eluting agents have proved more effective than citrate solutions, articles concerning the use of citric acid for separating rare earths still appear in the literature occasionally. For example, Ketelle and Boyd17 reported some further studies on the separation of rare earths with 5 pct citrate in 1951. They used 270 to 325 mesh Dowex-50 columns at 100°C and a pH of 3.28. vickery18 compared the effectiveness of a number of eluants for the separation of rare earths on NH+4 state Dowex-50 in 1952. He found that citric acid was more efficient than acetic, malic, tartaric, and aminoacetic acids. In 1953, Mayer and Freiling18 reported that citrate was inferior to malate, glycolate, lactate, and EDTA for the resolution of Sm-Eu and Y-Tb mixtures on 250 to 500 mesh, NH+4-state, Dowex-50 at 87°C. pinta20 used 5 pct citrate in the pH range 2.8 to 3.4 to obtain some rare earths for analytical work the same year. Trombe and Loriers21 reported the use of citric acid in their laboratory to separate rare earths in kilogram quantities. Lariers and Quesney22 reported some separations with citrate in 1954. They used 5 pct citrate at pH 2.8 to separate yttrium and the yttrium-group rare earths from the cerium-group rare earths. briers23 used 5 pct citrate at pH 3.2 and 90°C as late as 1956 to isolate thulium in fair purity. For the separation of tracer quantities of rare earths, Mayer and Freiling19 have recommended pH 5.00, 0.24M lactate at 87 °C on 250 to 500 mesh, NH+4-form Dowex-50. Freiling and Bunney24 have also employed lactic acid for the separation of fission-product rare earths. Cunninghame, Size-land, Willis, Eakins, and Mercer25 have reported a 4-hr separation of Y, Eu, Sm, Pm, Nd, and Pr with 1M lactic acid at pH 3.25 on Zeokarb-225 at 87°C. Stewart, et al.,26 reported separation factors for rare earths distributing between Dowex-50 and 0.25M glycolic acid. stewart27 also reported a 30-61 separation of Y, Tm, Er, Tb, and Lu tracers with buffered, 0.25M glycolic acid containing 0.05 pct Aerosol OT. The resin bed was 400-mesh Dowex-50 (X12). The pH of the ammonia-buffered eluant was 3.5. Various aminopolyacetic acids have also been used to obtain varying degrees of separation of the rare earths. In 1951, Fitch and Russell28 investigated iminodiacetic acid (IDA) and nitrilotriacetic

Jan 1, 1960

By W. S. CAYPLESS, H. O. Hofman, E. E. HARRINGTON

I. INTRODUCTION. LIME-ROASTING is a term proposed by Ingalls 1 for the operation of forcing air under pressure through a mixture of galena and lime at the kindling-temperature with the object of oxidizing lead and sulphur and of fritting or fusing the charge. If, finely-divided galena were treated in this manner without the addition of lime, the heat set free by the oxidation of part of the lead and the sulphur would be sufficiently great to fuse undecomposed sulphide, and thus stop desulphurization. Besides the chemical action that the addition of lime, limestone or gypsum to the charge may have, the admixture has the physical effect that it keeps the particles of galena separated from one another and accessible to the oxidizing effect of the air. At present, three methods of lime-roasting are carried out on a working-scale, the Huntington-Heberlein, the Carmichael-Bradford and the Savelsberg2 processes. In the last, which interests us here, an 8-ton charge is made up of galena, limestone and perhaps some siliceous or ferruginous flux; the whole is crushed to pass a screen with 3-mm. holes and moistened with 5 per cent of water. It is fed gradually into a bowl-shaped converter, 6.56 ft. in diameter, supported by trunnions attached to a truck. On the .bottom the converter has a grate with blast-inlet beneath. In starting, the truck with the converter

Jan 1, 1907

By G. C. Wallick

Since the appearance of the paper, "Solution of the Equations of Un-steady State Two-Phase Flow in Oil Reservoirs," by W. J. West, W. W. Garvin, and J. W. Sheldon,' a two-fold investigation of this subject has been carried out. One objective of the investigation has been to deter-mine the feasibility of solving such problems 7on a medium-size com-puter such as the Datatron*, and the other objective has been to in-vestigate the application of such cal-culations to experimental and theo-retical petroleum reservoir research. In the first Datatron calculations, the fluid and rock properties published by West, et al, were used, together with the published equations describ-ing the system. Details of the formu-lation not given in the original paper are discussed in the Appendix to this note. Reference should be made to the subject paper for the complete equations and defintion of symbols. An unexpected result of this in-vestigation was the discovery that the linear solution published by West was in error. Thus, in addition to describing the Datatron solutions and to discussing certain numerical diffi-culties which will be encountered if one uses the published method of solution, the purpose of this note is to indicate the nature of this error. LINEAR FLOW Since the linear case requires a minimum amount of scaling, a fixed-decimal point Datatron program was written for the one-dimensional flow problem and an attempt was made to duplicate the solution described by West. In the case described, fluid was produced at a constant rate, Q, until such time as well pressure reached 0.04. Production was then continued at constant pressure. From the constants and curves given by West it was determined that the ini-tial constant production rate could be approximated by Q = 0.007. An ini-tial dimensionless time step ?t = 0.434 X 10 - "as used, and each suc-cessive time step was doubled until a value of = 0.444 was reached. This constant interval was then used for the remainder of the solution. In subsequent solutions, several varia-tions in the time schedule were em-ployed, including smaller time steps and slower rates of increase in the time steps. In all cases, almost identical results were obtained regardless of the time schedule employed. However, as described below, it was noted that the time schedule had some influence on the rate of convergence of the solutions. As a check on the accuracy of the solution, the cumulative production at each time step was calculated using the two methods described in the Appendix. Satisfactory agreement was observed with the differences in these two values of the order of two parts in 50,000. It should be noted that the mass balance check as described is of questionable value, particularly with regard to the well pressure and saturation. This is especially true in the radial solution where pressure and saturation values near the wellbore would make only a negligible contribution to the numer-ical integration. It is believed, how-ever, that such a comparison is ot value in determining the over-all accuracy of a solution. In comparing the Datatron solu-tion with that published by West it was discovered that in the later stages of depletion, the pressures near the well declined more rapidly in our solution than in the West solution, and that the limiting well pressure of 0.04 was reached at an earlier time than that originally reported. It thus became evident that it would be impossible to duplicate the production schedule described by West and a constant rate of production was maintained until the well pressure was equal to 0.0. A representative comparison of the results published by West with those obtained in this investigation is shown in Fig. 1, which is a plot of GOR as a function of cumulative recovery. These two curves should be in agreement until a cumulative recovery is reached which corresponds to a well pressure of 0.04 — for the Datatron solution, a recovery of approximately 5.6 per cent. Actually, a major disagreement is evident. Subsequent correspondence

Jan 1, 1958

By J. D. Wood, J. T. Smith

ALLOYS with a dispersed second phase in a metallic matrix are generally much stronger than the matrix itself. Plastic deformation in dispersion-strengthened alloys is usually confined to the matrix phase when recovery processes are active, while in the absence of recovery both phases may yield.' The alloy system studied in the present research was WC-12 wt pct Co and consisted of noncoherent WC particles dispersed in the cobalt matrix. Some particle-to-particle contact existed but not enough to produce a continuous WC skeleton. The microstruc-ture of the WC particles was characterized by very straight edges, forming a trapezoidal shape in any plane of polish. Previous investigations with WC-Co alloys at room temperature have shown that fracture of the WC particles occurs in transverse rupture testing.' Room-temperature slip was reported for WC particles after indentation for hardness measurements.3 Elevated-temperature deformation of WC particles in a WC-12 pct Co alloy was suggested by recent electron microscope studies of specimens deformed at 900' to 1000°C.4 In highly deformed alloys, the WC edges were serrated in contrast to the usual straight or smooth appearance. WC-12 pct Co and WC-15 pct Co alloys have been previously studied under elevated-temperature com-pressive-creep conditions by the present authors. Electron microscope studies of two-stage replicas from deformed specimens showed no evidence of slip or fracture of the WC particles. These specimens were brought to temperature and allowed to equilibrate prior to the application of the creep load. It was believed that the load-application rate, a crosshead speed of 0.005 in. per min on an Instron universal testing machine, was sufficiently low that recovery within the cobalt matrix was sufficient to limit the deformation to this matrix. A series of experiments was performed to evaluate the influence of loading rate on the deformation of WC-Co alloys. A WC-12 pct Co alloy was selected for these determinations. The average WC particle size was 4.45 p with an average linear separation between particles of 0.59 p. The selected temperature was 800°C and was monitored with a Chromel-Alumel thermocouple attached to the specimen. Testing was conducted in an argon-atmosphere chamber to prevent oxidation of the WC-Co specimens. This chamber was mounted on an Instron universal testing machine equipped to apply the load at a fixed rate. Each specimen was loaded to 110,000 psi compression stress at 0.05 and 0.5 in. per min. The loading rate was monitored prior to insertion of the test chamber and was found to be almost precisely the nominal rate selected. The specimens were raised to temperature and held to equilibrate with the surroundings, and then the load was applied and held for 4 hr to duplicate the exposure time utilized for the creep specimens. The time to reach full load at a crosshead speed of 0.005 in. per min was some 500 sec and was reduced to 50 and 5 sec as the loading rate was increased to 0.05 and 0.5 in, per min, respectively. The model developed by Ansell,' when recovery processes do not occur, considers that fracture or deformation of the dispersed particles is necessary to relieve back stresses on dislocation sources and allow dislocations piled up against particles to sweep out in the matrix to cause plastic deformation; he further states that, even at elevated temperatures, the dispersed-particle deformation is necessary for yielding in the absence of recovery. For the case of straight dislocation segments piled up against a straight barrier, such as the straight-sided WC particles, the shear stress, 7 exerted on a particle is: where h is the spacing between particles (0.59 p), a is the applied stress (110,000 psi), p, is the shear modulus of the matrix (6.7 X lo6 psi at 80O°C), and b is the Burgers vector of the matrix dislocation. From Eq. [I], the shear stress, 7, exerted on the WC particles when no recovery occurs is of the order of 6 X lo6 psi at 800°C. The limiting stress, F, that will

Jan 1, 1969

By W. W. Eckles

This paper is presented as a suniniary report of the use of well gas composition correlations obtained from mass spectrometer recordings as a means of identification and determination of reservoir continuity. Conventional methods for detecting composition differences are en-pensive, elaborate, and difficult to obtain. This excludes the use of extensive composition data for most applications. During recent years the mass spec-trometer has come into general use as an analytical tool in petroleum refineries. The use of mass spectrometer composition patterns ir~ characterizing or "finger-printing" the produced gas from a reservoir, presents a novel method for correlatitlg gas samples from well to well. The mass spectrometer provides a trace similar to an electric log. having peaks which represent the abuntlnnce of certaitz hydrocarbons in the well gas sample. Without going further into the detailed analysis the idea has been advanced that these traces or patterns could be used as a means of identifying a particular natural gas. This theory has proven to be essentially correct. The mass spectrometer pattern method is simple and cheap as COi?7pared to other standard methods. It greatly facilitates the solution of reservoir and geological problems in which correlation of well gas Com-positions is a factor. Specific field applications have been made. This paper concerns the results obtained in 465 individual gas analyses from 35 fields and 77 res- ervoirs. In a number of cases it has been found that such data have been extremely valuable in the determination of reservoir continuity. In at least one case, the method was a valuable contribution in tracing a reservoir from sand to sand in a coinplex fnulted field involving n11rnerous gas reservoirs. Field applications are presented to illustrate the possibilities of the method at the present stage of developnient and to stimulate the ernployrnent of this new approach by geologists and petroleum engineers in the industry. INTRODUCTION Identification of producing horizons and the determination of reservoir continuity are often a problem in those areas where dome structures and highly faulted sands are encountered. To complicate the picture further, there may be numerous sands, one on top of the other which dip and diverge in different directions. Even though it may be possible to develop some solutions to the preceding problems on the basis of the geological and reservoir data on hand. it is readily recognized that substantiating data based on independent methods would he extremely valuable. It has been found through field studies that correiation of well gas composition can be used to advantage in the geological and engineering study of a complex reservoir identification problem. METHOD FOR COMPARISON AND CORRELATION OF WELL GAS SAMPLES Since a large number of well gas samples are required in a gas identification or reservoir con- tinuity problem, it is necessary that a method of analysis be employed which can detect differences in composition readily and inexpensively. Although detection would be possible by means of low temperature fractional distillation (POD) analpsis, the method requires a relatively large sample and is comparatively slow and expensive to run. The mass spectrometer affords an inexpensive and precise analysis of small well gas samples taken at the surface which are about 1/300 as large as a POD sample. These samples can be obtained by regular field personnei and shipped to the spectrometer for analysis. It is, therefore, a practical approach to the problem. A typical record from the mass spectrometer is illustrated in Fig. I. The peaks on the record represent the abundance of ions produced from the different hydrocarbon molecules making up the gas sam-ple. In well gas comparisons only five peaks are employed, since the other peaks are formed from the same gas molecules and furnish no additional information. The five peaks used represent the abundance of methane, propane, ethane, butane and heavier, and oxygen. They are referred to respectively as the 16, 29, 30, 43, and 32 peaks. The oxygen peak is used to correct for air content in

Jan 1, 1958

By E. S. Machlin, W. R. Yankee

IN a previous study1 of the effect of heating com-mercial titanium in air on its subsequent friction coefficient against other metals, as well as itself, it was found that the friction coefficient markedly decreased from a value of about 0.7 to about 0.3. A tentative explanation was given that surfaces normally produced at room temperature are not contaminated sufficiently to prevent seizing or welding of the titanium to the softer mating metals. The latter tend to cleanse themselves during rubbing over the harder titanium. It was thought that the lack of a contaminant protective film on the titanium was due to the high solubility of titanium for oxygen and nitrogen and hence an inability to form a contaminant oxide or nitride. This explanation requires the ratio of the surface absorption rate to the diffusion rate to become much lower at room temperature than it is at high temperatures. In order to check the phenomenon further, commercial titanium specimens were nitrided or oxidized at 800°C for 20 hr in flows of prepurified N2 and 01, respectively, at about 1/2 in. H2O above atmospheric pressure. Friction runs were made in argon using a freshly cut copper hemisphere (cut in argon) on surfaces cut successively into the diffusion layers in the titanium (cut in argon) using the techniques described in a previous publication.' DPH values (100 gram load) were made as a function of depth into the diffusion layer using a Tukon tester. Also, micrographs were taken at separate cross sections to indicate the diffusion layers. The results obtained are presented in Figs. 1 and 2, which show the "static" friction coefficient vs hardness for the nitrided and oxidized specimens, respectively. A separate measurement of the friction coefficient of clean copper vs iodide titanium also was made. From results reported in the literature' giving the oxygen and nitrogen contents as functions of the hardness, cross plots were made showing the friction coefficients as functions of the amount of interstitial solute. These plots are given in Figs. 3 and 4. From micrographs of the diffusion layers and the phase diagrams, it was deduced that the data in Figs. l through 4 correspond to the single phase a region. The points observed on the compound regions have been excluded from the figures. It is apparent that nitrogen or oxygen in solution in the a titanium markedly affects the friction coefficient against a softer mating metal. Discussion of Results These results are extremely interesting from both a practical and theoretical viewpoint. The theoretical implications will be discussed first. According to Bowden, the friction coefficient should be given to a good approximation by the relation where a is the fraction of contact area that has welded, a is the shear strength of weaker component, and H is the hardness of softer component. Using this relation alone, it is difficult to understand the results because none of the terms should be affected by a variation in the oxygen or nitrogen content of the harder and stronger metal, titanium. Even if the ratio of S/H for titanium is used in Eq. 1, the ratio has been shown to be independent of oxy-gen or nitrogen content.' If a more rigorous equa-tion is used combining Eq. 1 and a result given pre-viously" for the case where welding is absent, then the relation obtained is µ = a S/H + (1-a) a W aß/H where a is constant and Waß is the work of adhesion between the two metals comprising the friction couple. This relation states that if a is less than 1/2 or so, variation in the work of adhesion Waß between copper and the titanium should affect the friction coefficient markedly. It is reasonable to expect that the work of adhesion will depend on the oxygen or nitrogen content of the titanium. Available data4 show that clean metals and oxides have much lower works of adhesion than the same metals against the

Jan 1, 1955

By H. R. Peiffer

Hardening of soft metals can be accomplished by dispersing finely divided hard particles in them. The dispersing of finely divided alumina in silver in the presence of oxygen yields a high strength material which is unusual in that its mechanical properties above the melting point of the continuous Ag-O alloy matrix are similar to other solids. The tensile strength is studied for two of these alloys, one of which contains 15 pct by weight alumina and the other 20 pct by weight alumina. The average fracture strengths above 960°C of these alloys were found to be 0.4 x106 dynes per sq cm and 3.8x106 dynes per sq cm respectively. The strengths appenred to be independent of temperature above 960° C. The creep behavior of the 20 pct alumina material was studied above 960°C. The initial creep rate, 6 , of this material can be represented by where s is the applied stress and E the activation energy for the process. This energy is of the order of 1.7 to 2.1 ev. HARDENING of soft metals by the addition of finely divided, hard particles has resulted in the production of materials with excellent creep resistance and with inhibited recrystallization.' It has been demonstrated that such composites have useful strength up to temperatures very near the melting point of the soft metals, but one might expect that the strengthening by the dispersed particles ceases to be of importance above the melting point of the softer phase. This, however, is not so.2 In this paper the strengthening of a liquid Ag-O solution by the presence of dispersed alumina is discussed and experimental data concerning the mechanical properties of such a material at temperatures above 960°C presented. EXPERIMENTAL PROCEDURE Specimens were prepared using powder metallurgical techniques. Finely divided silver oxide was mixed in the appropriate proportions with "Linde B" alumina. For these experiments the specimens contained 15 or 20 wt pct alumina in pure silver.* These alloys were used in these experiments since they were the ones that could be handled readily enough to yield experimental data. Silver oxide was used in preference to silver because it is readily available in a very finely divided form and because the presence of excess oxygen in the silver is important to the properties2 of the material. The powders were mixed wet, using alcohol as the medium, in an ordinary food blender until a uniform mixture was obtained. They were then dried and heated to reduce the silver oxide to silver. Only a fraction of the finely divided silver remains unoxi-dized upon removal from the oven since the finest particles of silver oxidize immediately upon contact with air. The remaining fine silver served as a binder during pressing. The mixture was then pressed at 20 tons per sq in. into shoulder-grip (reduced section 1/4 in. by l/4 in. by 1 1/4 in.) and rectangular bar specimens (l/4 in. by 1/4 in. by 2 in.). The green specimens were heated very rapidly in a globar furnace to 1000°C, held at temperature untic they had reached theoretical density and then furnace cooled. The strength of the two compositions at temperatures above 960 °C was determined by pulling shoulder grip specimens inside a furnace mounted on a tensile machine. Specimens were gripped by means of wires wrapped around the shoulders of the specimen. Temperatures were measured by means of a chro-mel-alumel thermo-couple placed in the vicinity of the specimen. Control of temperature (within ±°15C) was accomplished by means of a Weston Recorder-Controller. Before loading, specimens were held at temperature for approximately 15 min in order to insure uniform specimen temperature. The low loads were measured by a special load cell designed by Baldwin Lima Hamilton. All testing was performed on a FGT testing machine of the same com-pany. Creep studies were performed in bending utilizing rectangular specimens of the 20 pct alumina alloy. Specimens supported at both ends by knife edges of inconel were placed in a globar furnace, heated quickly to temperature, and the deflection of the beam measured optically as a function of time. The weight and size of the specimen were predetermined and the maximum stress on the beam calculated from

Jan 1, 1962

By Henry R. Piehler

Continuous seven- and nine teen -filament close-packed silver-steel filamentary composites mere tested in tension. For purposes of comparison, the tensile behavior of the composite was predicted from the measured properties of the individual com-ponents. It was assumed that the axial strain is the same in both components and that the contribution of each component to the composite flow stress is proportional to its volume fraction. Good agreetnent was obtained between the observed and predicted tensile behavior. However, the composite elongation at fracture was about twice that observed when the individual steel wires were tested alone. Composites in which the filaments were widely spaced failed by the consecutive fracture of the filaments. In composites with closely spaced filaments, a composite instability preceded frachcre. These effects are explained in terms of a lateral restraint to the necking of the filaments which develops in the deforming composite. TWO-COMPONENT or composite materials are increasingly being developed and used as structural materials. Of all the composite geometries used, the filamentary configuration has proved to be the most successful. Polymer-bonded fiber-glas, the most common of the filamentary composites, has been investigated most extensively to date. One explanation of the behavior of fiberglas1 suggests that the load transfer between the glass fibers and the polymer matrix occurs primarily through friction. Bond separation between the fibers and the matrix is presumed to occur at very low stresses. Subsequently, only frictional bonding can provide for the transfer of load from the matrix to the fibers. The force normal to the fiber-matrix interface which gives rise to this friction is assumed to arise from the matrix shrinkage which occurs during solidification. The behavior of metallic filamentary-composite materials would be expected to differ from that of fiberglas in several significant aspects. Residual stresses resulting from differences in thermal contraction can be kept to a minimum by proper heat treatment. A strong wetted bond is formed between the two phases. In brazed joints,2 for example, the interphase bond is sufficiently strong that the weaker material will fail before separation occurs at the interface. Most metal filaments can undergo appreciable amounts of plastic deformation before failure. Failure by plastic instability or necking frequently occurs in metallic filaments, but not in glass fibers tested at room temperature. Differences between the behavior of fiberglas-polymer and metallic filamentary composites have indeed been observed in composites containing various types of metallic filaments in a silver matrix.3,4 At strains of the order of 10-2 pct, all the deformation was accommodated by the silver matrix. Hence, the strength of the composite depended only on the fiber concentration and not on the nature of the fiber. At higher strains, the strength of the composite increased when filaments with higher work-hardening rates were used. Surface slip line observations on a silver-mild steel composite that had been stretched 4 pct showed that an appreciable amount of deformation occurred in the filaments. Work on tungsten-copper filamentary composites5 has shown that the ultimate tensile strength for varying filament fractions followed a linear mixture rule. Assuming that the axial strains in both filaments and the matrix are equal, this linear mixture rule can be expressed as: The subscripts c, f, and m refer to the composite, filaments, and matrix, respectively. A and V are the area and volume fractions of each component. sc and of are the ultimate tensile strengths of the composite and the filaments. sm is the flow stress of matrix at a strain equal to the elongation at failure of the filaments. However, the tungsten filaments can undergo only a limited elongation before they fracture without an appreciable reduction in area. A composite containing more ductile filaments might indeed deviate from this linear mixture rule for the ultimate tensile strength, since filament failure by necking might be arrested by the matrix. SPECIMEN PREPARATION Specimens were prepared from 0.8 pct carbon steel piano wires of 0.015 in. initial diameter. The wires were given a thin nickel flash and a silver plate varying in thickness from 0.015 to 0.007 in., depending on the filament fraction desired. In order

Jan 1, 1965

By John E. O’Rourke, Kevin M. O’Connor, Pamela H. Rey

INTRODUCTION The resurgence of coal mining activity in the United States, brought on by the spiraling costs of fossil fw1 energy in the Seventies, has come at a time of intense public concern for the quality of the environment. Notwithstanding pressure on our economy to develop alternate sources of fue1 energy to the import of oil, the legislatures of several states have reacted to public concern over the environment by passing strict regulations aimed at con- trolling the subsidence effect s of underground mining. Agencies of the federal government charged with assistance to the mining industry, including the Department of Energy and the Bureau of Mines, have sponsored a number of instrumentation and field measurement projects aimed at the development of subsidence prediction models that can aid the mine operator's task of subsidence control. There are good empirical models developed in Europe for subsidence prediction, but they were made possible by a large body of mining-induced subsidence data collected there over a long period of time. No com- parable subsidence data base exists in the United States, and consequently empirical modeling of subsidence is not a realistic approach for our near term needs. Moreover, the geologic and topographic diversity of the several coal regions in the United States is expected to necessitate the development of several empirical models, each one expected to be relevant to its own region. Because of the time and costs that are likely to be involved in an empirical modeling approach, it is considered more expedient and cost effective to develop a general, mechanistic model for subsidence prediction purposes. In order to develop such a model, it is necessary to investigate and quantify the mechanics of the subsidence process from the mine level up to the ground surface. The series of projects discussed in this paper are designed to achieve this objective and include the following work: (1) the identification of geotechnical instrumentation that will pro- vide mine level overburden and surface subsidence data. (2) a field demonstration of selected instruments, and (3) documentation of case histories for complete subsidence mechanics, using the demonstrated and preferred instruments. An identification of feasible instrumentation and monitoring techniques was completed by Woodward-Clyde Consultants (WCC) in 1977 (O'Rourke and others). This paper discusses a demonstration of those instruments at a mine in Utah, and at two subsequent projects, currently underway at longwall mines in Colorado and West Virginia. 'he latter two projects when complete will provide documented case histories of subsidence mechanics. The process of optimizing subsidence instrumentation and monitoring techniques to the conditions encountered during installation and monitoring for these underground mines is shown to be an evolving one, and one which has had some notable successes to date. INITIAL DEMONSTRATION The initial design and demonstration of selected monitoring systems was carried out at the SUFCO No. 1 mine, near Salina, Utah. The instrumented panel was approximately 152 m wide. 640 m long. and was 290 m to 320 m deep. The mined height of coal. seam averaged 2.4 m. The mining method was room and pillar using continuous mining machines. This method allowed some monitoring of the supported condition during development, and eventually allowed monitoring of a caved system when both chain pillars and room pillars were extracted on retreat. The two instrument systems shown on Table 1 were selected from the earlier feasibility report for demonstration at SUFCO No. 1. Collectively, the two systems, one for a fully-supported mining method and the other for a fully-caved method, incorporate most of the instrumentation to be found within all five systems listed in the earlier feasibility report. The instrumentation includes surface, subsurface and mine monitoring installations. All of the SUFCO instruments selected to meet the specifications of the general instrument types listed on Table 1 were manually operated. That is, data from the installed system could only be obtained while a person was there to physically observe or operate the system readout. Automatic data recording equipment was available for some installations, but the objectives were to keep the systems as simple as possible for the demonstration project. A complete description of the surface, sub- surface and mine level instruments, and the demonstration project results are given in WCC (1982), and selected features are discussed in this paper.

Jan 1, 1982

By J. Chipman, H. R. Larson

The data previously reported for the quantity as a function of oxygen pressure at 1550°C have been used to compute the activities of Fe, FeO, Fe2O3, and COO in slags of the ternary system. Activities of the first three have been obtained also for two quasi-ternaries involving fixed CaO:SiO2 ratios. IN a previous paper1' the authors reported the results of an investigation into the effects of oxygen pressure on the composition of various simple slags analogous to some of those which are found in steel-making practice. The ratio of ferric iron to total iron was studied at 1550°C in iron oxide melts to which lime, magnesia, lime-plus-silica, and other oxides were added. The oxygen pressures involved those of air, carbon dioxide, and carbon dioxide plus carbon monoxide in several proportions. Although very low oxygen pressures could not be used, the slag-metal equilibrium studies of Fetters and Chip-man' permitted extending the results to slags in equilibrium with iron. For the ternary system CaO-FeO-Fe,O, the oxygen pressure-composition relationship has been determined from zero percent lime to lime saturation over an oxygen pressure range from air to that represented by equilibrium with liquid iron. Lime and silica were added to iron oxide in the ratios 0.54, 1.306, and 2.235 to form three quasi-ternary systems which were also studied over the entire region of liquid melts at 1550°C. Ternary Gibbs-Duhem Equation Wagner" has developed a method by which the activities of two components of a ternary system can be calculated if' the activity of the third component is known throughout the composition range being considered. The fundamental form of the Gibbs-Duhem equation for ternary systems is N, d In a, + N2 d In a, + N3 d In a3 — 0. Wagner has developed a usable form of this equation by introducing the term y = N3/(N3 + N1) and rearranging the equation to give (? In a1/?N2)x = -y/(1-N2)2 (? In a2/?y) x2 N2/1-N2 (? In a2/?N2) y To apply this equation to the slag system CaO-FeO-Fe,O,, the activity of one of these components must be known. However, the only activity which is known from the experimental data is the activity or pressure of oxygen in the gas phase with which the slag is in equilibrium. The activity of oxygen in the slag can be defined as the square root of the oxygen pressure. In order to use oxygen as one component, the composition of the slags must be converted to the basis Fe-O-CaO. Oxygen, exclu-sive of that contained in CaO, then becomes com-ponent 2 in Eq. 1, Fe is selected as 1 and CaO as 3. Then y = Ncao/(Ncao + NFe). Eq. 1 then becomes (? In a fe/? Nx) y = -y/(1-No)2 (? In ao/?y) xp - No/ 1- No (? In ao/ O No) y. In order to evaluate Eq. 2, the boundary conditions must be known. The obvious choice of a standard state for iron is to assign slags in equilibrium with iron an activity of one. Then In a,., or log aFe, which is substituted for convenience, is determined by integrating along a line of constant y from the slag in equilibrium with iron to the composition at which log a,, is to be determined. Mathematically this can be expressed as log aFe (Nlog y) = log a'Fe + (? log a Fe/? No) dNo where the primes indicate equilibrium with liquid iron and a'Fe = 1. When Eq. 3 is integrated along a line of constant y, the following is obtained: log a Fe (No, y) = - y No ?/ No ? y ( log ao/(1- No)2) No d No - No/1-No d log ao. Lime-Iron Oxide Slags Iron Activity: The oxygen pressure of lime-iron oxide slags is shown in Fig. 1 as a function of j, de-fined as j = Fe+++/(Fe++ + Fe +++) for various constant mol percentages of lime. The j values at 0 to 60 pct lime were then determined for oxygen pressures of 1, 10-1, etc. to 10-" atm. For purposes of calculation, the line for zero percent lime was extrapolated to

Jan 1, 1955

By P. R. Sperry, P. A. Beck, J. Towers

Dahl and Pawlek1 found that electrolytic copper develops extremely coarse grains at 1000°C after about 90 pct reduction by rolling. This coarsening occurs only under conditions of penultimate grain size, deformation, and alloying which lead to the "cube" recrystallization texture.l,2,3,5 The peculiar angular shapes and straight grain boundaries of the coarse grains were noted by several investigator.1,4,5 On the other hand, coarsening in Fe-containing aluminum or in AI-Mn alloys8 does not depend on a "cube" (or any well developed) recrystallization texture. It is true that increasing deformation by rolling, and, therefore, an increasingly well developed re-crystallization texture, are associated with decreasing incubation periods of coarsening.6-7-8 Nevertheless, coarsening readily develops in aluminum even after only 30 pct reduction by rolling, where the recrystallization texture is very weak.6,8 Also, coarsening was observed by Jeffries9 many years ago in sintered thoriated tungsten, which presumably has no preferred orientation. In all these cases coarsening is associated with grain growth inhibition by a dispersed second phase.8,9 The annealing temperature has to be suficiently high to overcome the inhibition at a few locations. But if it is too high, growth starts at many points, and the resulting grain size becomes much smaller.9 Normally, the coarse grains are more or less equiaxed, and the boundaries have a typical ragged appearance.6.8 Cook and Macquarie4 demonstrated that, in addition to the texture-dependent coarsening previously found at 1000°C,l electrolytic tough pitch copper may also coarsen at 800°C after 50 pct reduction by cross rolling. The coarse grains formed under such conditions have rounded shapes and ragged boundaries, like those in aluminum. When the annealing temperature is higher, the tendency for their formation decreases. All these observations suggest that the coarsening at 800°C is associated with inhibition by a second phase. Actually, coarsening at 800°C after 50 pct reduction by cross rolling was observed only in tough pitch copper,4 which contains Cu2O particles. On the other hand, the texture-dependent 1000°C coarsening occurs in both tough pitch and oxygen-free copper;4 it does not appear to depend on the presence of a dispersed second phase. However, the interpretation of the 800°C coarsening in Cu after 50 pct rolling as an inhibition-dependent process, similar to the coarsening in A1-Mn alloys, is somewhat weakened by the fact that this coarsening was reported4 to occur only after cross rolling, and not after straight rolling. It was, therefore, decided to re-examine this question. A 1 in. diam electrolytic tough pitch copper rod, No. 2 hard drawn, was annealed for 20 min at 700°C, rolled to 0.5 in., annealed 10 min at 700°C, and straight rolled to 0.064 in. It was then given a penultimate anneal of 20 min at 500°C and it was cut into four sections, which were given final reductions by straight rolling as follows: A 30 pct reduction of area B 50 pct reduction of area C 70 pct reduction of area D 90 pct reduction of area Specimens cut from the four sections were finally annealed at 800°C in an oxidizing atmosphere. Strip A remained fine grained up to 10 hr, but the specimen annealed 12 hr consisted of only 2 large grains. Strip B had a few scattered large (1/2 to 3/4 mm) grains after 1 min, although the balance of the specimen consisted of fine grains of about 0.02 mm. After 5 min there were several 10 to 15 mm grains present, and after 1 hr strip B was completely coarsened. The coarse grains had the same characteristics (see Fig 1) as those obtained by Cook and Macquarie at 800°C after cross rolling. Strip C had several grains of 0.05 to 1 mm after 1 min, but it was still largely fine grained after 12 hr. After 48 hr it consisted entirely of grains of about 0.5 to 4 mm, with an extraordinarily large number of twin bands. Strip D remained com- pletely fine grained after 4 hr at 800°C. These results indicate that, in the deformation range of 30 to 70 pct reduction, the incubation period for coarsening as well as the rate of growth and the final size of the coarse grains decreases with increasing deformation. Similar

Jan 1, 1950

By W. S. Cavender

The Round Mountain mining district, Nye County, Ne- vada, was discovered in 1906 on claims owned by Lewis D. Gordon. Initial mining operations uncovered gold veins of spectacular richness, and within a few days of discovery, Gordon sold his controlling interest for some $87,000. From this sale emerged the Round Mountain Mining Co., predecessor of Nevada Porphyry Gold Mines, Inc., the latter destined to become the major property owner in the area. Vein mining in the district continued sporadically into the early 1930s, yielding 9.3 Mg (330,000 oz) of gold plus substantial silver credits from approximately 626 kt (690,000 st) of ore. In addition to the lode deposits, the early miners recognized the placer potential in the alluvial fan material accumulated around the west and north sides of Round Mountain itself. Intermittent placer operations were carried out for a number of years, and in the 1940s and 1950s, Round Mountain Gold Dredging Co. worked the placers under a lease from Nevada Porphyry Gold Mines. The last placer operation terminated in 1959 when it, like some of its predecessors, proved uneconomic. Total placer production for the district is estimated at 3657 km (4 million yd) of gravel containing 59 Mg (210,000 oz) of gold and possibly 2.0 to 2.3 Mg (70,000 to 80,000 oz) of silver. Round Mountain is a small hill situated on the east flank of the Toquima Range in central Nevada. The hill is com- posed of relatively flat-lying Tertiary rhyolitic ash flow tuffs, which overlie Paleozoic metasediments and Cretaceous granites. Throughout the surrounding Round Mountain mining district, most of the known gold ores occur in the tuffs, although the metasediments and granites are also mineralized. Mineralization is structurally controlled, principally by a series of northwest-trending shears and shattered zones. Vein, stockwork, and disseminated ores occur, usually containing simple quartz-pyrite-gold mineral assemblages. The gold itself is electrum, having a silver content of 30% to 40%. In September, 1967, Elwood Dietrich, a prospector and mine promoter, obtained a purchase option on the 4452 ha (1 1,000 acres) of mineral rights held at Round Mountain by Nevada Porphyry Gold Mines. The original option had a buy-out price of $1 million and was established through Dietrich's friendship with officers of Nevada Porphyry. In April, 1968, Dietrich conveyed his option to Ordrich Gold Reserves Co., a partnership created by a group of west coast investors, mostly employees of the airline industry. There- after, Ordrich invested considerable funds in trying to test and develop the property, but soon recognized the need to seek financial and technical support from the mining industry. In December, 1968, Dietrich contacted Wayne Cavender, then Regional Geologist, Southwest, for Copper Range Exploration Company (CRX) in Tucson, Arizona, and made a data presentation. Shortly thereafter, Cavender was appointed Manager of Exploration and Chief Geologist for Copper Range Co. (parent company of CRX), New York City, and he asked C. Phillips Purdy, CRX Regional Geologist, Northwest, to make an initial property examination. Purdy's one-week field study took place in March, 1969, and resulted in a recommendation that CRX pursue its investigation of the property. The presence of low-grade gold mineralization in both the alluvial gravels and in the bedrock was verifiable, but the placer was deemed to have the greater immediate economic mining potential. At that time, gold was in the $1.41/g ($40 per oz) price range. Working from Purdy's information, Cavender decided to attempt acquisition of the property, and the first in a long series of negotiations was initiated with Ordrich. Basically, CRX felt that the placer had a promising potential for several reasons, including (I) past operators had recovered free gold but not the gold contained in the pebble fraction of the gravels; (2) past operations appeared to have been ineffectively designed or managed and not costefficient; and (3) the price of gold appeared to be poised for an upward move. Negotiations with Ordrich were prolonged and difficult, with CRX competing against several ma* mining companies, but finally an agreement was reached, effective June I, 1970. Gold was then back to $35. It is believed that, in

Jan 1, 1985