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(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who called at the Institute headquarters during the period Oct. 10; 1916 to Nov. 10, 1916. Edwin C. Bloomfield, Montreal. Roy E. Hoffman, Hannibal Mo. A. H. Collbran, Seoul, Korea. E. C. Harder, Cienfuegos, Cuba. Eugene Coste, Calgary, Canada. B. Jaeger, London, England. Fred Crabtree, Pittsburgh, Pa. Herbert H. Lauer, Philadelphia, Pa. P. St. John Dixon, London, England. Frank C. Laurie, Los Angeles, Cal. A. E. Drucker, London, England. 0. T. Lempriere, Melbourne, Australia. F. N. Flynn, Clifton, Ariz. W. C. Phalen, Washington D. C. Justice Grugan, Edwards, N. Y. William H. Rea, Pittsburgh, Pa. Fred A. Hale, Jr., Goodsprings, Nev. Thor Warner, Kingman, Ariz. P. S. Haury, Seoul, Korea. Arthur P. Allen has resigned his position with the Calumet & Hecla and is now connected with the Highland Mining Co., Ashcroft, B. C., Canada. H. Foster Bain, editor of the Mining Magazine of London, has resigned his position.

Jan 12, 1916

Tars Committee is the result of a desire of the authorities in charge of the publications of the four national engineering societies to co-operate in this matter. The members of the Committee are the following: Charles WarreN Hunt, Secretary of the American Society of Civil Engineers. D. S. Jacobus, Vice-President of the American Society of Mechanical Engineers. Joseph Struthers, Assistant Editor of the American Institute of Mining Engineers. CaRy T. Hutchinson, Chairman of The Editing Committee of the American Institute of Electrical Engineers. This Committee has held several meetings; it seemed advisable, at the outset, to limit its discussions closely to the general subject of abbreviations. Further, it seemed best to formulate a few general rules to be followed in making abbreviations, rather than to compile a list of forms to be recommended. The Committee decided to limit the subject more narrowly by considering only abbreviations to be used in the text, or general reading-matter, arid not those to be used in special matter,

Jan 1, 1905

FROM north to south activity is picking up through- out the Labrador Trough-already shown to be one of the world's great iron ore provinces. Center of current activity and interest lies in the mid-southern part of the trough as drills, equipment, and . . . money start to pour into the area near Mt. Wright and Wabush Lake. First major announcement in the area came early this spring when J&L and CCI outlined an estimated billion-ton deposit for "future" operations. Later impetus came from rumored railway for US Steel's Quebec Cartier projects in the Mt. Reed area slightly to the south. As this issue went to press US Steel's request for bids on railway and hydroelectric construction, plus IOCC's decision to step up the time- table on its program in the Wabush Lake area east of Mt. Wright, triggered a flurry of other announcements involving many of the top U. S. and Canadian iron ore firms. Do not look for most of these place names in the average atlas . . . they aren't there yet, but they soon will be.

Jan 6, 1958

By AIME AIME

THE quarters provided for the session on Milling Methods, Monday afternoon, Feb. 17, were filled to capacity. Galen H. Clevenger presided and notables, such as A. W. Fahrenwald, A. O. Gates, F. A. Thomson, W. L. Zeigler and others who had not graced a milling meeting in New York in recent years, or per- haps even never before, were actually present and spoke "in person." The first paper, by A. W. Fahrellwald (T. P. 275), entitled "Classifier Efficiency ; An experimental Study," was summarized by the author, who clearly showed the inconsistency of determining the efficiency of a classifier by means of sizing tests, especially where the material classified consists of a mixture of minerals of different specific gravities. In place of sieve analysis the procedure recommended involves the use of a glass settling-tube which is considered to give ideal classification and thus serve as a standard against which mill classification will be compared and relative efficiency calculated. This procedure sounds logical. Perhaps the greatest difficulty may lie in securing an ideal classifier acceptable to every one.

Jan 1, 1930

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By Robert S. Shoemaker

Recently we read a short article which stated that Marcona Corp. was studying the use of 25,000-hp autogenous grinding mills for an iron ore project in Alaska. As if that weren't enough, they were going to dispense with the primary crusher and feed run-of-mine ore to the grinding mills which were to be driven by wrap-around seven-cycle motors. We see also that Bougainville Copper Co. will be grinding 90,000 tpd of ore in only eight 4250- hp ball mills. At Bougainville too, their float section will be the first large one to be completely equipped with the new 350-cu ft flotation cells. After a half century of relatively peaceful growth in the minerals industry, we are now seeing almost exponential advances in both equipment and plant size. At the same time, we are seeing a constant blending of hydro and pyrometallurgical techniques and processes into minerals beneficiation until the sample grind-'em-up-and-concentrate-'em plant is becoming a rarity.

Jan 1, 1970

What do you think are the major problems that the mining industry is facing? While I don't have any intensive knowledge of the various problems that plague our industry, I know that they are all part and parcel of a more complex and fundamental question: What must we do to make mineral extraction successful, safe, and profitable? Obviously, as mining people, we must continue to expand our technical know-how and to make better use of it. In today's context, however, that is not enough. We must also improve the "public image" of the mining industry. Why? Because the key word in the term "public image" is: the public. Many people in our country and in Congress, for instance, are against strip mining-and some are even against underground mining. Pressures from the public, from voters, have affected past, present, and new legislation. And government controls can be a major problem in any industry, let alone mining. At first, people ask government to intervene, act, or resolve problems, even in management and labor contract negotiations. But later, after Washington gets involved, there are clashes between departments and personalities, and eventually there is resentment against the interference with mining operations or our daily lives.

Jan 3, 1975

By J. S. Levine, M. Prats

A homogeneous and uniform cylindrical reservoir containing oil and gas is fractured vertically on completion and is produced at a constant bottom-hole pressure. The fracture has an infinite flow capacity, is of limited lateral extent and is bounded above and below by the impermeable strata defining the vertical extent of the reservoir. Results show that such a fractured reservoir can be represented by a reservoir of circular symmetry having very nearly the same production history. The well radius of this circular reservoir is about 1/4 the fracture length and is essentially the same as that obtained previously for a single fluid of constant compressibility. At the same value of cumulative oil production, gas-oil ratios of fractured reservoirs producing at constant terzinal pressure are larger than those of reservoirs having no fractures. This leads to more inefficient use of the reservoir energy in fractured wells and results in lower reservoir pressures for the same cumulative oil production. The reduction in operating life due to fracturing a reservoir is not as great as that for a slightly compressible fluid. This diflerence can be accounted for by the lower reservoir pressure in the fractured reservoir and its adverse effect on the average mobility and compressibility of the oil. As anticipated, the reduction in operating life increases czs the reservoir permeability decreases. The type of results presented in this report can be used to determine the economic attractiveness of fracture treatments per se, to setect the initial spacing to be used in developing a field, and to compare the relative merits of fracturing available wells and infill drilling. INTRODUCTION The effect of vertical fractures on a reservoir producing either an incompressible or a compressible liquid has already been discussed in the 1iterature.l,2 Those results indicate that the production history of such a reservoir is essentially the same as that of a circular reservoir having an effective well radius of approximately one-fourth the fracture length. The present work reports on the effect of a vertical fracture on a reservoir producing two compressible fluids —oil and gas—by solution gas drive. Because of the empirical nature of the PVT and relative permeability data used to obtain the performance of such reservoirs, results can only be obtained numerically and with the aid of high-speed computers. Since reservoirs lose their radial symmetry when fractured vertically, pressure and saturation can no longer be given only in terms of distance from the well. Two coordinates (such as x and y) must now be used to describe the pressure and saturation within the reservoir, and, since we are dealing with compressible fluids, time is also a variable. Thus the solution of a vertically fractured reservoir requires finding two unknowns (pressure and saturation) in two space variables (say x and y) and in time (t). Since no means are readily and generally available for solving such problems at the present time, we have used the results of previous work1,2 to approximate the effect of a vertical fracture on a reservoir producing both oil and gas by depletion. The purpose of the present wmk, then, is to investigate the possibility of using available numerical techniques (limited at the moment to one space variable) to study the two-space-variable flow behavior resulting from a vertical fracture. Results obtained in the course of this investigation are also reported and discussed. Input and output data of the numerical methods used are given in practical units: BOPD, feet, psi, cp, and md. Results are discussed fist in terms of specific reservoir and crude properties and geometries. Later, dimensionless parameters are introduced in order to extend results to different values of some of the reservoir and fracture properties. IDEALIZATION AND DESCRIPTION OF THE FRACTURED SYSTEM It is assumed that a horizontal oil-producing layer of constant thickness and of uniform porosity and permeability is bounded above and below by impermeable strata. The reservoir has an impermeable, circular, cylindrical outer boundary of radius r,. The fracture system is represented by a single, plane, vertical fracture of limited radial extent, bounded by the impermeable matrix above and below' the producing layer (reservoir). It is assumed that there is no pressure drop in the fracture due to fluid flow. 1 indicates the general three-dimensional geometry of the fractured reservoir. Gravity effects and the effects of differential depletion resulting from variations in hydrostatic head (pressure) will be neglected. Thus, the flow behavior in the fractured reservoir is described by the

By R. I. Jaffee, H. R. Ogden, D. J. Maykuth

IN THE past year, Jaffee and Campbell&apos; and Finlay and Snyder2 reported on the mechanical properties of titanium-base alloys, some of which were in the same ranges of composition as are covered in this paper. In this paper, evidence confirming that given by Finlay and Snyder on the effects of carbon, oxygen, and nitrogen on titanium will be presented; and, in addition, new data will be given on the effects of these elements on the flow properties and phase transformation of titanium. Materials and Preparation of Alloys The preparation and general properties of iodide titanium have been adequately described elsewhere.&apos; , As-deposited iodide titanium rod, prepared at Battelle, of Vickers hardness less than 90 was employed as the base metal in the present work. This was the same material as that used by Finlay and Snyder.2 The probable analysis reported by them for standard quality metal holds here also: N 0.005 pct, 0 0.01 pct, C 0.03 pct, Fe <0.04 pct, A1 <0.05 pct, Si <0.03 pct, and Ti 99.85 pct. Carbon was added in the form of flake graphite supplied by the Joseph Dixon Crucible Co. Oxygen was added in the form of c.p. grade TiO, powder, produced by J. T. Baker Chemical Co. Nitrogen was added in Ti3N4 powder, supplied by the Remington Arms Co. Individual ingots weighed 7 or 8 g. Carbon, oxygen, or nitrogen was added by placing the corresponding powder in a capsule made from as-deposited iodide titanium rods and melting the capsule with the balance of the charge. The charge was are-melted with a tungsten electrode on a water-cooled copper hearth under a partial vacuum of very pure argon (99.92 pct minimum). Melting was practically contamination free. Vick-ers hardness increases of less than 10 points were normal for unalloyed iodide titanium control melts. Nitrogen analyses of are-melted iodide titanium showed a nitrogen content of 0.005 pct, about the same as is present in the as-deposited rod. No tungsten pickup was found in a melt of iodide titanium analyzed for tungsten. Weight losses in melting nitrogen-free alloys were very small and varied consistently from nil to 0.015 g (0 to 0.2 pct). This permitted the use of nominal composition for these alloys. Chemical analyses made for carbon, which can be analyzed conveniently by combustion methods, justified this procedure. Where nitrogen was added, considerable splattering took place. Here it was necessary to analyze for nitrogen by the Kjeldahl method. The ingots were hot rolled at 850°C to about 0.045 in. thick. After hot rolling, the strips were descaled by mechanical grinding, and then given a cold reduction of 5 to 10 pct to insure a uniform thickness throughout the length of the specimen. The edge strips and the tensile strips were annealed in a vacuum of 1x10-4 mm Hg pressure for 3 1/2 hr at 850°C and furnace cooled. Methods of Investigation Hardness Measurements: At least five Vickers hardness measurements were taken using a 10-kg load on each sample in the following conditions: (1) top and bottom of each ingot, (2) top and bottom surface of as-rolled and annealed sheet, and (3) on cross-section of annealed sheet and all quenched specimens. Tensile Tests: Tensile tests were conducted on Baldwin-Southwark testing machines having load ranges of 600 or 2000 lb. Tests were made on 1-in. gauge-length specimens, 3 1/4-in. overall length, 1/2 in. wide, 0.040 in. thick, with a reduced section 1 1/4 in. long and 0.250 in. wide. Two SR-4, A-7 strain gauges, one mounted on each side of the specimen, were used to measure the strain over a limited range to determine the modulus of elasticity. After the modulus of elasticity readings had been taken, load vs. strain readings were taken, using only one strain gauge, at increments of 0.0001 in. until the yield points were passed and then at 0.001-in. increments to the limit of the strain-gauge indicator (0.02 in.). Strain readings above 0.02 in. per in. were taken every 0.01 in., using dividers to measure the strain between the 1-in. gauge marks until the maximum load had been reached. Crosshead speed, when using the SR-4 gauges, was 0.005 in. per min, and, when using dividers, 0.01 in. per min. Flow Curves: Flow curves were determined using the true stress-true strain data obtained during the tension test. The usefulness of this type of information has been dealt with very adequately elsewhere by L. R. Jackson,&apos; J. H. Hollomon,6 and many others. Flow curves of true stress vs. true strain could be converted to the more conventional cold-work curve of 0.2 pct offset yield strength vs. percentage of cold reduction by means of the transformation, 1/1 = 1/1-R, where R is the fraction reduction in cold working. Thus, the true strains corresponding to percentage reduction can be calculated, and the 0.2 pct offset yield strengths scaled off the — 6 curve by taking the true stresses corresponding to the values of 6 + 0.002 strain. Heat Treatment: For the transformation studies, the alloys were heat treated in a horizontal-tube furnace using a dried 99.92 pct argon atmosphere, and quenched into water. Essentially no contamination was found after several hours of heat treatment at temperatures up to 1050°C. Metallography: Specimens were prepared in the

Jan 1, 1951

By J. W. Spretnak, D. A. Chatfield, G. W. Powell, J. R. Eifert

In nzost cases, the driving force for a lattice transformation is produced by supercooling below the equilibriunz transformation temperature. The interfnce reaction in isothermally annealed, multiphase diffusion couples may involve a luttice transformation which also requires a driving force. Direct experinzental evidence has been obtained for the existence of the driring force in the form of a supersaturated phase at the aocc)-0@cc) interface in Cu:Cu-12.5 ult pct A1 couples; the super saturation is equivalent to an excess free energy of approximately 3 cal per mol at 905. A tentatiue interpretation of the dynanzic situation a1 the interface based on the free energy-composition diagram is proposed. THE presently accepted theory of diffusion in multiphase couples1 states that there will be a phase layer in the diffusion zone for every region which has three degrees of freedom and which is crossed by the diffusion path in the equilibrium phase diagram. For binary systems, this restriction excludes all but single-phase fields and, for ternary systems, only one- and two-phase fields are included. In addition, Rhines"~ as well as other investigators3 6 have reported that the compositions of the various phases adjacent to the interfaces are, for all practical purposes, the compositions given by the intersections of the diffusion path with the solubility limits of the single-phase fields of the equilibrium phase diagram. Some studies of the rate of thickening of these intermediate diffusion layers indicate that the thickness of the layer changes para-bolically with time, or: where x is the position of the interface relative to an origin xo, t is the diffusion time, and k is a temperature-dependent factor. crank7 shows mathematically that, if the compositions at an interface are independent of time and the motion of the interface is controlled by the diffusion of the elements to and from the interface, then the segments of the concentration penetration curve for a semi-infinite step-function couple will be described by an equation of the form: hence, Eq. [l] follows from Eq. (21 if the interface compositions are fixed and if the motion of the interface is diffusion-controlled. Although the concept of local equilibrium being attained at interfaces has assumed a prominent role in the theory of diffusion in multiphase couples, experimental evidence and theoretical discussions which challenge the general validity of this concept have been reported in the literature. arkeen&apos; has stated that strict obedience to the conditions set by the equilibrium phase diagram cannot be expected in any system in which diffusion is occurring because diffusion takes place only in the presence of an activity gradient. Darken also noted that it is usually assumed that equilibrium is attained locally at the interface although the system as a whole is not at equilibrium, the implication being that the transformation at the interface is rapid in comparison with the rate of supply of the elements by diffusion. ISirkaldy3 indicates agreement with Darken in that he believes the concept of local equilibrium is at best an approximation because the motion of the phase boundary requires that there be a free-energy difference and, hence, a departure from the equilibrium composition at the interface. Seebold and Birks9 have stated that diffusion couples cannot be in true equilibrium, but the results obtained are often in good agreement with the phase diagram. The initial deviation from equilibrium in a diffusion couple will be quite large because alloys of significantly different compositions are usually joined together. Kirkaldy feels that the transition time for the attainment of constant interface compositions (essentially the equilibrium values) will be small, although in some cases finite. Castleman and sieglelo observed such transition times in multiphase A1-Ni couples, but at low annealing temperatures these times were quite long. Similarly, ~asing" found departures, which persisted for more than 20 hr, at phase interfaces in Au-Ni and Fe-Mo diffusion couples. Braun and Powell&apos;s12 measurements of the solubility limits of the intermediate phases in the Au-In system as determined by microprobe analysis of diffusion couples do not agree with the limits reported by Hiscocks and Hume-Rothery13 who used equilibrated samples. Finally, Borovskii and ~archukova&apos;~ have stated that the determination of the solubility limits of phase diagrams using high-resolution micro-analyzer measurements at the interfaces of multiphase couples is not an accurate technique because of deviations from the equilibrium compositions at a moving interface; diffusion couples may be used to map out the phase boundaries in the equilibrium diagram, but the final determination of the solubility iimits should be made with equilibrated samples. The purpose of this work was to investigate the conditions prevailing at an interface in a multiphase diffusion couple and to compare the interface compositions with those associated with true thermodynamic equilibrium between the two phases. Microanalyzer techniques were used to measure interface compositions in two-phase Cu-A1 diffusion couples annealed at 80@, 905", and 1000°C for various times.

Jan 1, 1969

By J. F. Emig

The Burns Harbor Plant was designed to produce and process large ingots into plate, hot rolled sheet, cold rolled sheet and tin plate. Rimmed ingots, in particular, are poured 96" to 106" high and weigh up to 85,000 lbs. During the latter part of 1970, a quality problem was encountered involving white line "sliver" defects on the surface of cold rolled, low carbon, rimmed steel sheet. Sheets containing these defects are not desirable for exposed automotive items because of the possibility of opening during forming or becoming more apparent after forming. Metallographic work located and identified two basic types of inclusions within the white line defects: (1) complex oxides consisting largely of iron and manganese oxides and (2) scale-type iron oxide. The presence of these oxides strongly suggested that the-white line defects were resulting from an unsound, unclean rim zone.

Jan 1, 1972

By John W. Wagner, Robert K. Willardson

Single crystals of Pbl-xSnxTe have been grown from nonstoichiometric, cation-rich melts with the objective of producing as-grown, bulk material containing carrier concentrations ranging from 1016 per cu cm to 1018 per cu cm. Three specific crystal composi-tions were investigated in detail; x = 0.00, 0.10, and 0.17. Pull rates of from I to 3 mm per hr were used. Single crystals were successfully pulled from melts containing as little as 30 at. pct Te. Hall coefficients, resistivities, and carrier mobilities of these materials were determined. The relationship between the composition of the melt and the carrier concentration in the as-grown crystal has been studied for the three crystal compositions of interest. BULK single crystals of Pb,-,Sn,Te have previously been grown from stoichiometric melts.1,2 Such crystals are p-type and have relatively high carrier concentrations ranging from -9 x 1018 per cu cm for PbTe to -8 x l020 per cu cm for SnTe at 77°K. These carrier concentrations result from deviations from stoichiometry (lead vacancies) in the as-grown crystals. Since lower carrier concentrations are desirable for electrooptic device applications, these crystals are usually subjected to long-term, isothermal anneals.&apos; This paper reports on the growth of Pbl-xSnxTe single crystals from nonstoichiometric melts with the primary objective of producing as-grown material containing relatively low (1016 to 1018 per cu cm) carrier concentrations and also reports on the general characteristics of these crystals. The phase relationships in the Pbl-xSnxTe systems are such that materials solidifying from nonstoichiometric, cation-rich melts will have smaller deviations from stoichiometry than materials grown from stoichiometric melts. Fig. 1 is the T-x phase diagram for PbTe in the vicinity of the stoichiometric composition.3 A crystal grown from a stoichiometric melt will solidify at a melting point maximum at which the solid will contain -0.5002 atom fraction of tellurium. PbTe single crystals grown in our laboratories from stoichiometric melts have carrier concentrations of 9 x 10" per cu cm, indicating that the excess tellurium in the crystals is as expected from this phase diagram. However, growth of PbTe from a lead-rich melt will result in material having a more nearly stoichiometric composition. Although the addition of Sn to the melt shifts the solidus curve further toward the tellurium-rich side,4 the general discussion given for PbTe applies to the Pbl-xSnxTe systems as well. EXPERIMENTAL Single crystals of Pb1-xSnxTe have been grown in our laboratories from nonstoichiometric, cation-rich melts using the Czochralski technique and boric oxide liquid encapsulation. The details of the growth apparatus and growth technique have been reported in a previous paper on growth of these alloys from stoichiometric melts.2 In the present study, three specific crystal compositions were investigated in detail; x = 0.00, 0.10, and 0.17. Growth of alloy crystals from nonstoichiometric melts requires considerable care, and good quality single crystals were obtained in this study only by optimizing the mechanical and thermal stability of the growth system and by using pull rates of from 1 to 3 mm per hr (Pbl-xSnxTe crystals are easily pulled from stoichiometric melts at rates of from 5 to 10 mm per hr). The liquid encapsulation technique was found to yield near-ideal conditions, since the B2O3 layer increased the thermal stability at the growth interface, permitted easy attainment of near-ideal thermal gradients, and dampened vibrations at the melt surface. The hygroscopic character of the B2O3 was a slight problem and vacuum heat treating was necessary to completely remove the water from the boric oxide. During initial growth, the seed diameter was reduced and a narrow neck (1 to 2 mm diam) of several millimeters length was grown. The latter steps were found to be necessary for the growth of single crystals, i.e., if either of these two requirements were

Jan 1, 1970

By Frederic Dewey

HAVING shown1 the difficulty of assaying so-called cyanide bullion and the extreme variations often found in the results, an investigation was undertaken to discover if possible the causes of these variations by a series of tests upon the specific action of various metals during cupellation. Zinc.-In order to test the effect of zinc, 48 assays were made in sets of 6, 2 rows of 3. The charge for the middle cupels consisted of 500+ mg. Au, 1,125 mg. Ag, 10 mg. Cu. To the end cupels 50, 75, 100 and 125 mg. Zn were added, and a set of each proportion was cupelled with 4 and 8 grants Pb. The cupels were assayed and showed the following gold contents in milligrams, the figures being arranged as the cupels stood in the muffle: TABLE 1 50 Mg. Zn 75 Dig. Zn 100 Mg. Zn 125 Mg. Zn 4 grams, Pb 0.2S 0.25 0.2S 0:29 0.27 0.34 0.33 0.28 0.41 0.31 0:30 0.36 0.30 0.33 0.34 0.31 0.31 0.29 0.35 0.37 0.42 0.41&apos; 0.34 0.38 8 grams, Pb 0.40 0.43 0.42 0.38 0.36 0.43 0.35 0.34 0.41 0.33 0.37 0.37 0.54 0.17 0.59 0.44 0.43 0.43 0.38 0.39 0.41 0.57 0.38 0.46 _ In many instances the zinc does not appear to affect the cupel gold, but various exceptions are apparent, generally in the cupels in the back row. With 100 and 125 mg. Zn, in some cases the Zn burned with its characteristic flame before it could alloy with the other metals. Also, less fusible oxidized lumps were observed during the cupellation and some cupels showed a scum. In order to lessen the premature burning out of the Zn, an alloy of 2 parts Ag to 1 part. Zn was used. A uniform charge of 500,+ mg. Au,

Jan 7, 1917

By Nathaniel Arbiter

THE separation of minerals by flotation can be regarded as a rate process, with the extraction of any one mineral determined by its flotation rate, and the grade of concentrate by the relative rates for all the minerals. So regarded, the significant variables for the process are those that control the rates. These variables are of two types, the first describing the ore and its physical and chemical treatment prior to flotation and the second characterizing the separation process in the cells. This paper will examine the variation in rates for a group of separations, will show that a simple rate law appears to govern, and will consider the relation of the control variables to the rates. The use of rate constants for evaluation of performance and efficiency will be discussed. Flotation involves the selective levitation of mineral and its transfer from cell to launder. The flotation rate is the rate of this transfer. It may be defined by the slope of a recovery-time curve for any cell in a bank, or at any time in batch operation. The objective in flotation rate study is an equation expressing the rate in terms of some measurable property of the pulp. This can be either the concentration of floatable mineral in weight per unit volume1,2 or a relative concentration, which will be a function of the recovery." A rate equation for an actual flotation pulp will contain at least two constants, both to be determined from the data. One of these, the initial concentration or proportion of floatable mineral, is not necessarily equal to the feed assay because of nonfloatable oversize or locked particles." The other, a rate constant, is a measure of proportionality between the rate and the pulp property on which the rate depends. The value of the rate constant will be determined by the values of all variables which control the process and will be changed by significant changes in any of them. It is, therefore, a direct measure of performance. Where recovery or grade change continuously with flotation time, the rate constant will be independent of time and will characterize the entire course of the separation. Development of Rate Equations Rate equations can be developed either by analysis of the mechanism of the process or by direct fitting of equations to recovery-time data. Sutherland&apos;s attempt by the first method&apos; suggests that the effect of particle size variation on the rate complicates the derivation of a simple equation applicable to an ore pulp. A further problem with an ore is the concentrate grade requirement, which usually involves a variable rate of froth removal. Thus the final rate for any cell may depend on the froth character and froth height, as well as on the pulp composition. This does not imply that each cell cannot reach a steady state2 in which the rate will depend ultimately on pulp composition. The second method is the fitting of rate equations consistent with the necessary boundary conditions* to experimental recovery-time curves. On the assumption that under constant operating conditions the flotation rate is proportional to the actual or relative concentration of floatable mineral in the pulp, a generalized rate equation may be expressed as follows: Rate = Kcn [I] where K is the rate constant, c is some measure of the quantity of floatable mineral in the pulp at time t, and n is a positive number. In previous rate studies, the value of n has been taken as 1, either by direct assumption," or as a result of the hypothesis that bubble-particle collision is rate determining.&apos; A first order equation results, which after integration in terms of cumulative recovery R, leads to Loge A/A-R = Kt [2] The quantity A is the maximum possible recovery with prolonged time under the conditions used. No conclusive proof for the validity of this equation in flotation has been advanced. The evidence cited in its support consists entirely in the demonstration that it appears to apply to a limited number of recovery-time curves."&apos; It will be shown subsequently that this procedure is not sufficient to establish the order of a flotation rate equation. The possibility that the equation may be of higher order therefore requires examination. If, in particular, the exponent in eq 1 is assumed to be 2, then after integration there results R = A2Kt/1 + AKt [3] with K again a rate constant and A the maximum proportion of recoverable mineral. Eq 3 may be

Jan 1, 1952

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By C. S. Matthews, H. C. Lefkovits

The performance of depletion-type reservoirs at the stage in which the gas is at such low pressure that gravity is essentially the sole driving force has been investigated both theoretically and by use of scaled models. A simple theory is developed which completely explains the scaled model experiments. It also agrees with some field data. When the range of production rates to be considered is less than 25, the theory predicts that the production rate decline curve for homogeneous reservoirs in which there is a free surface will be of the hyperbolic type with n = 2: The theory gives a physical significance to the parameter y. INTRODUCTION This investigation was undertaken to determine a method for prediction of performance of reservoirs where gravity is the sole driving agent. Depletion-type reservoirs fall into this category after the pressure has fallen to some low value. Some examples of this type are found at Healdton (&apos;Oklahoma), Oklahoma City (Oklahoma), Midway-Sunset (California), and East Coalinga (California). Generally the production rate per unit thickness is quite low in such reservoirs, with the conse- quence that this stage is sometimes called the stripper stage. However, per well production rates may be of the order of 50 B/D because of the considerable sand thickness open in some of the reservoirs. In the past, predictions of behavior of such reservoirs have been based on empirical extrapolation of the rate of production to the economic limit. Difficulties arise when new wells are drilled in the field or when production is artificially curtailed, since in such cases the decline rate of individual wells or leases changes and no clear trend can be detected. It was hoped that the present study would lead to improved methods for evaluation of such reservoirs. The study was initiated by building a scaled model for a particular field and by running a number of experiments on this model. Later a simple theory was developed which explained the laboratory results completely. The theory led, as shown below, to a hyperbolic-type decline equation which had been used previously&apos; in a completely empirical manner. THEORY The General Case The problem to be treated here is the following: We are given a homogeneous cylindrical porous reservoir with exterior radius r, wellbore radius r., and height h,, containing, from the bottom up to a height hit oil, some gas, and interstitial water (see Fig. 1 A). The portion of the reservoir between K, and h, is filled with gas at low pressure, with residual oil, and with interstitial water. If at a time t = 0, the level of the oil in the wellbore is lowered to a height h,, what is the production rate from the reservoir at a time to The configuration within the reservoir at some instant

Jan 1, 1957

[SATURDAY, FEBRUARY 16 10 am to 5 pm Council of Section&apos; Delegates SUNDAY, FEBRUARY .17 1 pm Student Relations Committee 2 pm Board of Directors 2:30 pm MIED-Mineral Economics Instruction 5:30 pm MIED-Cocktails and Buffet Supper 7:30 pm MIED-Mineral Education 8 pm IMD-Program Committee]

Jan 1, 1952

By R. H. Card

THE geophysicist is interested greatly in the resistivities of different formations or parts of the earth&apos;s crust; sometimes he is interested in a single figure in the nature of an average, or what may be called the effec-tive resistivity, of all the formations or parts of the crust at a given location to a certain depth. Conversely, the engineer whose work is concerned with the coordination of power and communication lines in such a manner as to avoid electrical interference between these two types of facilities is concerned very much with this effective resistivity; to a lesser extent he is interested in the resistivities of individual forma-tions. In connection with his work the coordination engineer often estimates the effective resistivity; of the earth from measurements of the mutual impedances of lines; in other instances he derives front measure-ments of a different type the resistivities of different layers of the crust and synthesizes these to obtain an estimate of the effective resistivity. In connection with problems concerning the coordination of power and telephone lines, measurements from which earth resistivities can be estimated have been made at many points in the United States by the American Telephone and Telegraph Co. and its associated companies, in cooperation with various power companies and with railroad companies having electrified sections of line. Other such measurements have been made by the Joint Subcommittee on Development and Research of the Edison Electric Institute and the Bell Telephone System. A study has been made of the correlation of these data with the ages and types of materials of the strata of the earth&apos;s crust at each of the test locations. The purpose of this study has been primarily to provide information to assist in predicting the order of magnitude of effective resistivities to be used in the preliminary consideration of problems con-cerned with the coordination of power and communication lines. To make the results of such a study, most useful, it is essential that the geo-logical data to be secured when such a prediction is wanted be readily obtainable. Data on the ages of strata and the types of materials com-posing geological structures can usually be obtained, therefore particu-

Jan 1, 1937