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By H. L. Couch, J. D. Lubahn

In decarburized mild steel, the strain hardening arising from 1/2 pct strain can he partly recovered by subsequent heating, even though re crystallization or grain growth does not occur. This recovery varies from 5 pct in an hour at room temperature to 70 pct at 800°C. The variation with temperature was qualitatively similar to that reported for copper, but occurred at lower temperatures. Except for the initial portion (Bazdschinger effect), the stress-strain curve following recovery can be superimposed on the original curve by shifting it to the left. In other words, the tensile behavior is essentially the same after straining and heating as if there had been less strain and no heating. The importance of this observation in understanding simultaneous deformation and recouery is explained. AN earlier investigation1 on copper showed that strengthening due to strain hardening was augmented by heating at moderate temperatures, but was diminished by heating at higher temperatures, Fig. 1. Auxiliary experiments1 indicated that recovery and strain aging are independent processes, which may proceed simultaneously, and whose effects are additive as indicated in Fig. 1. In order to study recovery by itself, similar tests were made on decarburized mild steel, where strain-aging effects are known2 to be absent. MATERIAL, TESTING PROCEDURE The specimens were made from rimmed, hot-rolled SAE 1020 steel 0.040 in. thick, by stamping out rectangular blanks and end holes, Fig. 2, and pack milling the reduced section. These were wet-hydrogen treated2 for 16 hr at 720°C to remove carbon and nitrogen completely and then heated 1 hr at 850°C in vacuum to stabilize the grain size. Auxiliary ex- periments showed that heating up to 800°C will not cause grain growth. They were prestrained 5 to 7.5 mils per in. using the special grips shown in Fig. 2. Such small strains are less than the "critical" strain for recrystallization and, therefore, will not cause recrystallization, even at high temperature. The recovery treatment after prestraining and before the retesting consisted of heating for 1 hr in wet hydrogen at the desired temperature. In one test, Huggenberger gages were used to define more accurately the loading and unloading curves. Fig. 3 illustrates the creep during the early stages of unloading, the hysteresis due to cold stretching, and removal of hysteresis by heating.3 RESULTS The load-deflection curves had ratner widely differing shapes for different specimens, both before and after the recovery treatment. Some specimens showed perfectly smooth curves, Fig. 4, while others exhibited an inflection more or less suggestive of a yield point (Fig. 4 shows all extreme case). Some curves were much steeper than others; Fig. 5 shows two extremes. Some additional ex-

Jan 1, 1960

By Robert S. Moehlman

Is recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. Cost oF Drilling Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 30 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $I per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft.. as more time is consumed in pulling rods' 3. Cost of preparing drilling stations, getting air underground to stations? etc' 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946

By R. S. Moehlman

IN recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. COST OF DRILLING Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 50 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $1 per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft., as more time is consumed in pulling rods. 3. Cost of preparing drilling stations, getting air underground to stations, etc. 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946

Since production curtailment for other than engineering reasons is gradually disappearing, and more and more wells are now producing at capacity and showing declining production rates, it was considered timely to present a brief review of the development of decline-curve analysis during the past three or four decades. Several of the commoner types of decline curves were discussed in detail and the mathematical relationships between production rate, time, cumulative production and decline percentage for each case were studied. The well-known loss-ratio method was found to be an extremely valuable tool for statistical analysis and extrapolation of various types of curves. A tentative classification of decline curves, based on their loss ratios, was suggested. Some new graphical methods were introduced to facilitate estimation of the future life and the future production of producing properties where curves are plotted on semilogarithmic paper. To facilitate graphical extrapolation of hyperbolic-type decline curves, a series of decline charts was proposed, which will make straight-line extrapolation of both rate-time and rate-cumulative curves possible. Introduction During the period of severe production curtailment, which is now behind us, production-decline curves lost most of their usefulness and popularity in prorated areas because the production rates of all wells, except those in the stripper class, were constant or almost constant. While production-decline curves were thus losing in importance for estimating reserves, an increasing reservoir consciousness and a better understanding of reservoir performance developed among petroleum engineers. This fact, together with intelligent interpretation and use of electric logs, core-analysis data, bottom-hole pressure behavior and physical characteristics of reservoir fluids, eliminated a considerable part of the guesswork in previous volumetric methods and put reserve estimates, based on this method, on a sound scientific basis' At the same time, a number of ingenious substitutes were developed for the regular production-decline curve, which made it possible to obtain an independent check on volumetric estimates in appraisal work, even though the production rates were constant. with the now steadily increasing demand for oil to supply the huge requirements of this global war, proration for reasons of than prevention of underground waste is gradually disappearing. More and more wells are, or will be, producing at capacity or at their optimum rates, as determined by sound engineering practice. With this trend, the character of producing wells seems to regain, more or less, its "individuality," and the old and familiar decline curve appears to have had a comeback as a valuable tool in the hands of the petroleum engineer. It may be timely, therefore, to retrace the development of decline-curve analysis in the past by presenting a brief chronological review of bulletins and papers published during the past three or four decades, which have

Jan 1, 1945

By J. J. Kanter

The choice of a nonferrous metal or alloy for a given Application is frequently predicated upon a consideration of properties other than merely the capacity to withstand stress. When ability to withstand stress is the dominant engineering requisite of the material to be used, economic considerations seldom lead to the choice of nonferrous metals and alloys. In most instances where a nonferrous metal is used it is because of some special quality of either corrosion resistance, electrical properties, thermal properties, antifriction, or fabrication advantage that commends its application. The mechanical properties, as a rule, present an incidental engineering consideration entirely secondary to the special quality deciding its use. Indeed, great ingenuity has often been exhibited by engineers and metallurgists in overcoming the mechanical shortcomings of a nonferrous base material. in order to take advantage of some outstanding property. . With this thought in mind, we might better appreciate some of the problems in stress design. encountered with nonferrous metals and alloys pertinent to a symposium on their creep characteristics. Our considerations must necessarily be far from complete because of the broad range of materials in the nonferrous category for which creep properties assume a technical interest. Creep deformation of these metals can be subject for study at temperatures ranging from ambient to incandescent, depending upon which one is under consideration. Whereas the steel metallurgist thinks of creep as an essentially high-temperature engineering consideration, the same is not necessarily true in the field at hand. The diversity of the creep problems in the nonferrous field is perhaps accountable to some extent for its virginity as a symposium subject. Each research has necessarily been sponsored by some organization or group interested either in promoting the use of a certain nonferrous base or in obtaining data toward the solution of a mechanical difficulty encountered in applying a material indispensable for its special attributes. Perhaps as many techniques of creep testing have been used as there are base metals implied under "nonferrous." Some of the creep problems that have occasioned researches are associated with steam-pressure vessels such as tubes, drums, castings, and forgings of copper-base and nickel-base alloys, cable sheathings of lead and lead-base alloys, electrical transmission lines of copper and aluminum, engine parts die-cast from alloys based on zinc, aluminum, copper— soldered and brazed joints-—as well as numerous duties at extremely high temperatures met with alloys based on cobalt, nickel, tungsten, molybdenum, platinum, and other precious or semiprecious metals. Postwar technology will have the benefit of much creep research on nonferrous metals, which for reasons of security cannot be discussed on this occasion. Suitability OF Metals As already intimated, the bulk of the constructional metal that ordinarily is

Jan 1, 1945

By M. C. Udy, P. C. Rosenthal

The use of minute boron additions to steel has been given considerable attention in recent years. Comparisons made between boron-free and boron-containing heats of otherwise identical analysis have indicated, beyond a doubt, the potent effect of boron for increasing hardenability. The actual hardenability increase has been expressed as a multiplying factor, or in terms of the percentage of other alloying elements boron might replace. The present study presents the results of a systematic survey of the use of boron as an addition to two types of alloyed steels and as a replacement for various percentages of molybdenum in these steels. Both cast and wrought steels of the two basic compositions were prepared. Molybdenum, because of its scarcity at the time the investigation was started, was selected as the element to be replaced by boron. Attention was also given to the effects of boron on other properties of the steels, particularly their notched-bar toughness, to investigate other advantages or limitations of this element as an alloy addition. Experimental Work Steels Investigated The two basic compositions used in this survey of boron as a replacement for molybdenum are given in Table I. The cast steels were higher than the wrought steels in silicon content to improve the casting quality. Table i.—Nominal Composition of the Two Base Alloy Steels composition. Per Cent. steel Mn Si Ni Cr Xi-Cr-Mo Wrought 0.28-0.70-0.15-o.45—o.45-0.32 0.90 0.25 0.55 0.55 Cast 0.28-0.70-0.35-0.45-o.45- 0.32 0.90 0.45 0.55 0.55 hln-&lo Wrought 0.28-I .go-o.0.15-0.32 1.70 0.15 Cast o. 28- I.40- 0.35- 0.31 1.70 0.45 a Sulphur was 0.040 per cent maximum and phosphorus 0.030 per cent maximum. Each of these base steels, both cast and wrought, was made with molybdenum contents of o, 0.10, 0.20, 0.30, or 0.40 per cent, the total comprising 20 steels. In addition, each of these 20 compositions was prepared with boron additions in amounts of o.oo15, 0.003, and 0.006 per cent as ferroboron; and in amounts of 0.002 and 0.005 per cent as two commercial intensifier alloys, designated alloys A and B. The boron-treated steels totaled 100, and the entire series 120. Test Procedure Each steel was melted as an individual heat in a 50-lb., magnesia-lined induction furnace. Ingot iron scrap or mixtures of ingot iron and steel. scrap were used for melting stock. Deoxidation was carried out in the furnace with 0.10 per cent aluminum. The aluminum addition was

Jan 1, 1948

By R. E. Mussler, F. E. Block, T. T. Campbell

An exploratovy study was made to deternzine the feasibility of preparing beryllium by the metallother-. mic reduction of beryllium oxide. The procedure involved heating a relatively nonvolatile metal reductant in contact with beryllium oxide in a vacuum. Beryllium vapors formed in the reaction were condensed as a coherent deposit on a heated surface. Reductants tested included tanthanun, yttrium, thorium, Th-Mg, and Zr- Ti alloys, and the hydrides of lanthaniir~z, yttriunz, and zirconiur?7,. Test ten7pevatures ranged from 1350'to 175O' C, and reaction times ranged frot 30 to 100 lzr. In most cases metal recoveries were lola, although yields close to 90 pct of theoretical were obtained with latzthanum, yttrium, and zirconium reductants. In most cases metal deposits were contanzinated by beryllides, /ls/cally of the Beln M or Bel3 M type. BERYLLIUM has a variety of unique properties including light weight (sp gr 1.85), low thermal neutron absorption cross section, high scattering cross section, excellent dimensional stability, high strength to weight ratio, high melting point (1285"C), high specific heat, and high thermal conductivity. Interest in beryllium as a structural material in aerospace and nuclear-reactor applications is based on these properties. In general, the potential of beryllium has been only partially realized, largely because of the brittleness of poly crystalline forms of beryllium, difficulties associated in its fabrication, and joining problems encountered in its use in structural applications. Several excellent books and monographs are available to those interested in the occurrence, extractive metallurgy, and utilization of beryllium.1"* Although an extensive amount of research on beryllium has been in progress during the past 10 years, much of the work has been concerned either with the so-called "ductility problem" or with the development of methods of fabrication and evaluation of properties of various commercial or purified grades of metal. A recent survey of the literature indicates that only limited efforts are presently being directed toward developing new methods for reducing beryllium compounds to metal. Two major commercial methods now used to produce beryllium are the magnesium reduction of beryllium fluoride and e lec trow inn ing of beryllium from fused sodium chloride-beryllium chloride salt baths. Only a limited amount of work has been done on the direct reduction of beryllium oxide. Major factors which complicate this method are the great stability of beryllium oxide and the reactive nature of beryllium metal. The stability of the oxide necessitates extremely high reduction temperatures and vacuum operation to liberate beryllium metal vapor and cause reactions of the following type to proceed: XBeO(s) +YM(s)-MYOX(s) + XBe(& As early as 1939, roll' used this reaction for the direct reduction of beryllia by zirconium and titanium, and later in 1952 ~a~el~ studied the reduction of beryl-

Jan 1, 1967

By M. C. Udy, P. C. Rosenthal

The use of minute boron additions to steel has been given considerable attention in recent years. Comparisons made between boron-free and boron-containing heats of otherwise identical analysis have indicated, beyond a doubt, the potent effect of boron for increasing hardenability. The actual hardenability increase has been expressed as a multiplying factor, or in terms of the percentage of other alloying elements boron might replace. The present study presents the results of a systematic survey of the use of boron as an addition to two types of alloyed steels and as a replacement for various percentages of molybdenum in these steels. Both cast and wrought steels of the two basic compositions were prepared. Molybdenum, because of its scarcity at the time the investigation was started, was selected as the element to be replaced by boron. Attention was also given to the effects of boron on other properties of the steels, particularly their notched-bar toughness, to investigate other advantages or limitations of this element as an alloy addition. Experimental Work Steels Investigated The two basic compositions used in this survey of boron as a replacement for molybdenum are given in Table I. The cast steels were higher than the wrought steels in silicon content to improve the casting quality. Table i.—Nominal Composition of the Two Base Alloy Steels composition. Per Cent. steel Mn Si Ni Cr Xi-Cr-Mo Wrought 0.28-0.70-0.15-o.45—o.45-0.32 0.90 0.25 0.55 0.55 Cast 0.28-0.70-0.35-0.45-o.45- 0.32 0.90 0.45 0.55 0.55 hln-&lo Wrought 0.28-I .go-o.0.15-0.32 1.70 0.15 Cast o. 28- I.40- 0.35- 0.31 1.70 0.45 a Sulphur was 0.040 per cent maximum and phosphorus 0.030 per cent maximum. Each of these base steels, both cast and wrought, was made with molybdenum contents of o, 0.10, 0.20, 0.30, or 0.40 per cent, the total comprising 20 steels. In addition, each of these 20 compositions was prepared with boron additions in amounts of o.oo15, 0.003, and 0.006 per cent as ferroboron; and in amounts of 0.002 and 0.005 per cent as two commercial intensifier alloys, designated alloys A and B. The boron-treated steels totaled 100, and the entire series 120. Test Procedure Each steel was melted as an individual heat in a 50-lb., magnesia-lined induction furnace. Ingot iron scrap or mixtures of ingot iron and steel. scrap were used for melting stock. Deoxidation was carried out in the furnace with 0.10 per cent aluminum. The aluminum addition was

Jan 1, 1948

By J. C. Detweiler

THE Jacksonville plant of Humphreys Gold Corp., operating under leases from National Lead Co. and Rutile Mining Co. of Florida, recovers rutile, ilmenite, zircon, and monazite from an ancient beach deposit containing about 2.5 pct of these minerals in combination. The location of the plant is unique, being situated about ten miles from the center of Jacksonville, Fla., surrounded by a thriving suburb. Two four-lane highways pass within 1 ½ miles of the plant. An expressway now under construction will pass over the tailing pile, within 500 ft of the mine office. The elevation at the plant site is approximately 50 ft above sea level. This is one of two mines producing titanium minerals and zircon in Florida. Another mine is also operated by Humphreys at Starke, about 50 miles southeast of Jacksonville. The Jacksonville deposit extends about six miles in a north-south direction, averaging approximately one half mile wide and twenty ft thick. Values begin at the surface over most of its length and ter-

Jan 1, 1952

By Bernard S. Borie

THE U-A1 binary system has been studied by Kaufmann and Gordon.' They have shown that three intermetallic compounds occur in the system: UAl², UAl², and a third compound tentatively identified as UA15. UA1³ is simple cubic, a, = 4.26A. Its unit cell contains one formula weight, and it is iso-morphous with AuCu³. UA1² is face-centered cubic, a, = 7.72A, with eight formula weights per unit cell. Rundle and Wilson have shown that it is iso-morphous with Cu²Mg.² A Debye-Scherrer pattern of the Al-rich compound indicated a more complex structure than either UA1² or UA1³. The following is a report of an investigation of the crystal structure of this compound and a determination of its stoi-chiometric formula. A melt of uranium (99.9 pct pure) and aluminum (99.9 pct pure) was prepared which contained about 25 pct U by weight. It was cooled from 950" to 650°C at a rate of 500" per hr and then taken from the furnace and air cooled. An X-ray diffraction pattern of the ingot showed only lines of aluminum and the Al-rich compound. Filings from the ingot were reacted with a solution of sodium hydroxide, producing a fine black powder, the diffraction pattern of which showed no traces of aluminum. A comparison of the Debye-Scherrer pattern of our sample with the data published by Kaufmann and Gordon1 showed that it is identical with the Al-rich phase which they reported. Single crystals were obtained by cooling the melt more slowly (about 50°C per hr). After reaction with sodium hydroxide, the crystals appeared as small black needles attached to the ingot surface. Cell Size, Space Group, and Uranium Positions A series of Weissenberg photographs for rotation about two of the crystallographic axes, taken with copper radiation filtered through nickel foil, revealed an orthorhombic cell of dimensions: a = 4.41 ± 0.02A b = 6.27 ± 0.02A c = 13.71 ± 0.03A Reflections of the type hkl were observed only if: h + k + I = 2n, of the type hkO only if: h = 2n and k = 2n, of the type h01 only if: h + 1 = 2n, and of the type Okl only if: k + 1=2n. Hence, the cell is body-centered with an 001 glide plane. The evidence for the existence of the glide plane is the observed extinction of 23 possible re- flections of the type hkO with at least one index odd. There are only two space groups consistent with these data: C"" — I2ma and D28 2h — Imma.³ he possible sets of atomic positions for these space groups are: 12ma: (0 0 0; ½ ½ ½) + 4: (a) x 0 %; x % % (b) x ¼ z; x ¾ Z 8: (c) x y z; x y z; x, ½ + y, Z x, M - y, z Imma: (0 0 0; ½ ½ ½) + 4: (a) 0 0 0; 0 ½ 0 (b) 0 0 ½; 0 ½ ½ (c) ¼ ¼ ¼; ¾ ¼ ¼ (d) ¼ ¾ ¾ ; ¾ ¾ % (e) 0 ¾ Z; 0 ¾ z 8: (f) x 0 0; x 0 0; x 1/2 0; x ½ 0 (g) ¼ Y ¼; ¾ Y ¾; ¾ y ¼; 1 y ¼ (h) 0 y z; 0; ;; 0, ½ + y, z; 0, l/2 - Y, Z (i) x 1/4 z; x ¾ z; ¼ z; x ¾ z 16: (j) xy z; x y z; x y z; x y z; x, ½ + y, Z; x,½ - y, z; x, ½ - y, z; 11 1, + y,Z Though density measurements on different samples were found to vary somewhat, they were all within the range 5.7 * 0.3 g per cu cm. If it is assumed that the unit cell contains four formula weights of UAl,, it is 6.5 g per cu cm. If there were as many as eight uranium atoms per unit cell, the

Jan 1, 1952

By H. Margolin, E. Ence

Two versions of the Ti-Mn binary diagram have been published recently.' , -0th diagrams show two compounds in the region between 40 and 70 wt pct Mn, but disagree as to the reaction in which these compounds are involved. The investigation of the manganese-rich portion of the Ti-Mn binary diagram was undertaken at New York University as part of an investigation of ternary systems with titanium and manganese in order to determine which of the proposed diagrams is correct. A number of alloys between 31 and 70 pct Mn were prepared in a manner similar to that described by Cadoff and Nielsen; except that low currents (110 amp in argon atmosphere) and long melting times (up to 20 min) were used to prepare the alloys. Compositions were determined from weight-loss data, assuming all loss as due to manganese. Heat treatments were carried out with relatively large pieces, since there was a tendency for manganese to be lost during heat treatment. When X-ray diffraction data was to be obtained, the central parts of heat-treated specimens were used to make filings or powders which were not heat treated. Specimens were heat treated from the as-cast state. Heat-treatment times were as follows: 1150°C, 1 day; ll00°C, 2 days; 1000°C, 5 days; and 900°C, 15 days. It should be noted that equilibrium was not attained at 900°, 1000°, and in one case, at 1150°C. Metallographic specimens were electrolytically polished and strain etched by a technique described elsewhere.' At least four, and possibly six, phases were found in the region investigated. On the basis of information available at this time, a completely self-consistent diagram cannot be constructed, and therefore data are presented only for those phases which have been identified by both X-ray and microstruc-ture. Of the compounds detected, the one highest in manganese is the Laves phase, TiMn,, which is hexagonal with a MgZn, type of structure.V he diffraction data for TiMn2 from a 70 pct Mn alloy annealed at 1150°C are shown in Table I. The c/a ratio and parameters of TiMn2 agree well with those of Wall-baum hnd Rostoker et al. The values obtained here are c/a = 1.641, a = 4.825A. Another compound is the y-phase which is estimated to contain about 60 pct Mn. The y-phase has a structure which is identical to that of TiMn, with c/a = 1.639, a = 4.906Å. The fact that y has a Laves-type lattice would suggest that TiMn, has a range of solubility which extends from the composition of TiMn2 (69.6 Mn) to that of y (60 pct Mn). However, this could not account for microstructures which, in the 60 to 70 pct Mn region, show several phases in both the as-cast state and after annealing at 1150°C. Fig. 1 shows a two-phase structure of a 60 pct Mn alloy annealed at 1150°C. The phases present are y (white) and e (dark). If y were TiMn, then, according to Rostoker et al.,' the second phase of Fig. 1 should be P-Ti. The diffraction data for the 60 pct Mn alloy of Fig. 1 are shown in Table I. The starred d-values are from the e-phase and these lines do not correspond to those of &Ti. Comparison of as-cast and 1150°C microstructures of the 60 pct Mn alloy indicates that e precipitated from y. Since y has a

Jan 1, 1955

By William L. Langdon

PRELIMINARY COMMENTS Recent years have been difficult for the mining industry, as reported in the most recently published Minerals Yearbook (Volume III), the estimated value of world crude mineral production has been generally declining since an historical peak in 1979 (16).* The costs of mineral extraction have, however, been rising rapidly. Individual project costs are, in general, exceeding the growth rate of mining companies, while operating and technical risks continue to increase (16). Risks resulting from currency flotation and political actions in the host country are also on the increase. Labor difficulties (such as those experienced by Phelps Dodge) and the Reagan Administration's recent rejection of a proposal by the International Trade Commission for higher tariffs and/or quotas on copper imports have also contributed to the mining industry's difficulties. The net result of these factors has been financial difficulties for mining companies with respect to paying for their current financial obligations as well as trouble finding new financing. It may be that one of the main factors in determining the future of the mining industry will be centered upon capital structure, i.e. how its assets are financed. Historical Perspective on Capital Structure In The Mining Industry As discussed by Isreal Borenstein in (4) and Thomas Navin in (7), mining firms typically have used a financial model based upon high financial leverage, i.e. large amounts of debt relative to equity. This has been especially true for American firms developing orebodies discovered overseas. Problems sometimes occurred for those who were forced to commit themselves to completion guarantees in order to obtain bank loans. Mines under these circumstances must be brought into operation regardless of market conditions. Politicians in overseas countries also may use the high profits required by these high risk ventures to argue for nationalization of mines. Coupled with the frequent use of high leverage, mining companies traditionally have paid out a higher percentage of earnings in the form of dividends than manufacturing firms. This is significant, in that such funds paid out in dividends are not available for other uses. Hence, a greater reliance upon debt and/or additional stock issues than would ordinarily be the case for manufacturing firms. Thomas Navin has stated in (7), that the theory behind the high dividend payout was that, "most mining investors thought of themselves as investing in a wasting asset, one that would gradually deplete over time until it ultimately became valueless." Investors, therefore, required a receipt, in the form of dividends, of a percentage of current earnings and a return on original capital. This view of reality, apparently, has not changed in some quarters, even though modern mine management ordinarily replaces depleted mines with new ones. It may be that this practice of high dividend payout coupled with relatively high financial leverage should be reexamined. Current capital structure theory could prove interesting and useful for those concerned with current conditions. Capital structure patterns and related practices in this industry are not, in many cases, congruent with those thought by some to be theoretically optimal. Prelude to Capital Structure Discussion With a few notable exceptions (such as the Guggenheims and Anaconda's chief operating,. officer in 1971) mining companies' chief executives and boards of directors have not been persons with financial backgrounds. This, one could conjecture, might have accounted for the relatively nonsophisticated nature of capital structures of mining companies. In order to appreciate the possibilities with respect to alternative influence on a firm's performance, one must be acquainted with the basics of financial theory. This paper is an attempt to familiarize the reader with (1) capital structure in general, (2) the usual components of this structure, (3) the general theory behind the use of financial leverage, and

Jan 1, 1985

By H. W. Mead, C. E. Birchenall

SELF-DIFFUSION in pure gold has been studied by Sagrubsky,1 McKay,2 Gatos and coworkers,3,4 and Okkerse.5 Sagrubsky used a sectioning technique but reported results far lower than any of the others. The work seems to be carefully done, and it appears that a factor of 60 must have been omitted in his calculations or the units misstated by an equivalent amount. McKay used a surface activity method, and Gatos and Kurtz, following Gatos and Azzam, applied autoradiography to the measurement of activity on an oblique section through the diffusion zone. These two investigations agree well with each other but give results below those reported here. Okkerse employed a very careful sectioning technique. Chemical interdiffusion in Au-Ag alloys has been measured under several concentration conditions. Average values of the diffusion coefficient for the range of 0 to 18 atomic pct Au were obtained by Braune.6 since the diffusion coefficient probably varies markedly over this range, the values may not be too useful. Also, a plot of these data suggests that grain boundary contributions may have been appreciable at the lowest temperature employed. Ebert and Trommsdorf7 give average values for 0 to 9 atomic pct Ag, and Johnson8 reports average values for 45 to 55 atomic pct Au. Johnson also measured silver and gold self-diffusion coefficients in 49.2 atomic pct Au alloys, and self-diffusion in pure silver. The self-diffusion of gold in pure silver has been determined by Jaumot and Sawatsky.10 Other measurements which do not cover a range of temperature or are of low precision are not mentioned in this account since a review by Kubaschewski is available." This contribution reports diffusion coefficients for self-diffusion in gold, self-diffusion of gold in Au-Ag alloys, and chemical diffusion in Au-Ag alloys. Although it had been intended to obtain silver self-diffusion data in the alloys it now appears unlikely that these measurements can be done soon, so a table has been worked out for a self-consistent scheme of diffusion coefficients which should not be too far from the true values. Experimental Details Experiments on pure metals were conducted using fine gold of 99.99+ pct Au and silver of the same purity. Alloys were prepared by melting the two metals together in a graphite crucible in vacuo and allowing the slug to solidify in the crucible. After machining to a cylinder about 0.85 in. diam each slug formed one half of a diffusion couple. Compositions were determined later by analyzing chips taken from the extremities of the diffusion zones. The gold used in the preparation of alloys was sometimes pure gold but was normally either in the form of clean pickled alloy from previous experiments or gold recovered by oxalic acid precipitation from warm aqueous solution. Chips from the various stages of experimentation were dissolved in aqua regia and the solution taken to low bulk, diluted with water, and again taken to low bulk before diluting for the precipitation. This precipitated gold contained a small amount of silver as the only impurity detectable by spectrographic analysis. Disks for use in diffusion couples had opposite plane faces abraded on emery paper as far as zero grit, were annealed for 12 hr at 900°C in argon, and had one face of each disk polished to 000 emery paper. Radioactive gold was plated at about 2 ma per sq cm from a small volume of cyanide solution containing 1 mg of inactive gold as a carrier. Specimens , were freshly abraded on 4/0 emery paper immediately prior to plating. In the case of the active couples disks of a similarly prepared but unplated alloy or pure metal were placed on top and welded to the active disk. With the concentration couples disks of different compositions were welded. Welding was performed after a pressure of 5,000 lb had been applied at

Jan 1, 1958

By P. T. Carter, A. B. Murad, H. B. Bell

The distribution of manganese and oxygen between molten iron and FeO-MnO-SiO2 slags not saturated with SiO2 has been determined and used to calculate activities of MnO and SiO2 in MnO-SiO2 slags. These were found to be intermediate between the corresponding activities in FeO-SiO2 and CaO-SiO2 slags. THE equilibrium constant for the reaction between molten iron containing manganese and slags containing FeO and MnO, namely: FeO(slag) + Mn = MnO (slag) + Fe(1iquid) is most conveniently represented by: (MnO) Kmn== (FeOHMn] [Fe] is omitted, since for dilute solutions of manganese in iron, afe is approximately constant. When activities are used for (MnO), (FeO) and [Mn], Kmn is a true equilibrium constant and will apply to all types of slags, but when weight percentages are used, the values obtained for Kmn for different slags vary considerably. The equilibrium between iron containing manganese and relatively pure FeO-MnO slags has been investigated by several workers;1-4 the most recent investigation reported is that of Chipman, Gero, and Winkler.4 Their results, corresponding to slags containing up to 4 pct SiO2, were represented by the equation: 6440 log KMn = +2.95 In this case both slag and metal may be assumed to be ideal solutions, therefore the values of Kmn so derived represent the true constant for the manganese reaction. As the molecular weights of MnO and FeO are almost equal, weight concentration of these oxides may also be used in the above expression. Acidic oxides such as SiO2 and P2O5 affect the activities of FeO and MnO differently, so that when total concentrations by weight of FeO and MnO are used in the above expression for Kmn the values obtained are no longer constant. These will be denoted by K'mn. Körber and Oelsen5 investigated the manganese reaction using silica-saturated iron-manganese-silicate slags and found: 7940 logK'mn =3.172 This corresponds to Kmn values much higher than those obtained for pure oxide slags. The results of Körber and Oelsen, and Chipman, Gero, and Winkler are summarized in Table I. In the present investigation it was decided to study the manganese reaction using iron-manganese-silicate slags not saturated with silica, and which fell mainly between the orthosilicate join 2FeO . SiO2 - 2MnO.SiO2 and the FeO-MnO side of the ternary system FeO-MnO-SiO2. It has been shown" that the phases present in solidified slags of this part of the diagram are (FeO, MnO) and (2Fe0.Si02, 2Mn0-SiO,) solid solutions. Such an investigation would be of interest in deoxidation and also gives information on activities in liquid slags. This equilibrium has not as yet been investigated successfully. Hilty and Crafts7 attempted to examine this reaction in a rotating crucible furnace, but were unable to obtain suitable slag samples. In the present investigation one series of experiments was carried out in a rotating crucible furnace, because the success of Taylor and Chipman8 with such a furnace in their study of oxygen distribution between molten iron and (CaO, MgO) — FeO-Si0, slags suggested that it would minimize contamination of the slag by the crucible material. A second series was carried out using a wire-wound resistance furnace. Rotating Crucible Furnace Experiments In all cases the metallic charge consisted of 800 to 1000 g Armco iron containing 0.02 pct C, 0.03 pct Mn, and traces of Si. Manganese and silicon additions consisted of electrolytic manganese (99.9 pct)

Jan 1, 1953

By G. W. Doepel, W. P. Sibley

A three-dimensional, multilayer, mathematical model has been developed for predicting performance by mis-c.ible displacement. Areal and vertical coverage, total gas-oil ratio, propane-oil ratio and oil-recovery factor for the multilayer system are computed as functions of hydrocarbon pore volumes of solvent injected. The technique has been programed for use on the IBM-7090 digital computer and was used in calculating perfortnance for continuous miscible displacement of a five-spot, 50-layer .system. Results reported are based on the use of published laboratory data for a single, uniform five-spot layer. However, the technique is sufficiently generalized to calculate performance of any system of layers of any pattern configuration and mobility ratio, provided data are available to describe perfortilance of a single layer. The technique has also been adapted for predicting oil recovery resulting from various gas-driven slug sizes. The results presented for both continuous and slug-type injection show, that permeability variation has 0 significant effect on oil recovery but that solvent composition, within the range studied, has a lesser effect on oil recovery. INTRODUCTION Experience from many of the miscible-phase displacement projects currently in operation indicates that the injected solvent flows preferentially through the more permeable zones in the reservoir. This preferential solvent flow in the high-permeability streaks results in a reduction of vertical coverage (C,) and an attendant decrease in oil recovery. The multilayer technique presented in this paper uses areal coverage (C.,) data for a single layer from the literature and from laboratory measurements of displacement efficiency (herein referred to as microscopic coverage, C,,) to calculate areal coverage (C,) and vertical coverage (C,) for the multilayer system. Oil recovery for the multilayer system is obtained as the product of (Ca) (Cm) (Cv). It is assumed that C, in a single layer is 100 per cent; therefore, (C,) (Cm) is a measure of oil recovery from that layer. Data presented in the literature for a single layer generally deal only with areal coverage and/or microscopic coverage —but seldom specifically with oil recovery. Nevertheless, this technique is flexible in that either Ca and C,, of an individual layer or a direct measurement of oil recovery for the layer may be used. In developing this technique, the problem arose concerning a method of determining the relative amounts of solvent that will enter each layer during any increment of solvent injected. To solve the problem, a dimensionless equation was derived for expressing solvent injected into each layer as a fraction of total injection. A finite-difference method was applied in integrating the equation to determine oil recovery for the system of layers as a function of pore volumes injected. The technique subsequently was programed for solution on a digital computer. The method was applied to predict performance of a deep California reservoir under continuous miscible displacement and has been extended for calculating oil recovery as a function of slug size for gas-driven slug injection. The technique may be applied to any petroleum reservoir that can be adequately represented as a layered system, by considering permeability variation,1-3 permeability range and average permeability. Each layer is then assigned its appropriate permeability, porosity and hydrocarbon saturation. RESERVOIR VOLUME CONTACTED BY SOLVENT It has been recognized for some time that high microscopic coverage (C,) can be obtained by injecting a miscible agent, such as propane, into an oil reservoir. These high values of C,,, have been demonstrated in the laboratory and are normally between 80 and 100 per cent, depending on solvent and reservoir oil composition as well as on reservoir temperature and pressure. In miscible-phase displacement, because there is no interface between solvent and oil, effective permeability to solvent is the same as that to oil. (This assumes that oil saturation has not changed and that water is immobile.) Consequently, mobility ratio (M) simply reduces to that of viscosity ratio . Perrine4 and Habermann5 have shown that large

By Stuart S. Forbes

Changes and additions made to the Canadian Copper Refiners Ltd. electrolytic refinery between 1949 and 1955 are reviewed. The effect of high current density on current efficiency and section work is discussed. OPERATING and physical changes which have been made to the electrolytic tank house of Canadian Copper Refiners Ltd. suggest a review of present conditions. Previous papers1 a have described the general arrangement. Since then, expansion, closer anode spacing, and higher current densities have increased annual cathode capacity from 112,000 to 182,000 tons. A recent extension to the tank house, begun in the spring of 1954 and completed in August 1955, was undertaken to handle the monthly production of about 3,000 tons of anodes from Gasp6 Copper Mines. The anticipation of these additional receipts and large stocks of unrefined copper caused by the increase in custom smelting at Noranda made it necessary to increase production without immediately available additional cell capacity. Accordingly, from December 1954 through March 1955, the tank house was operated successfully at a current density of 24.0 amp per sq ft. Again, as of September 1955, one of the two tank house circuits is being operated at this high density. Although other refineries may operate at current densities higher than 24 amp per sq ft, none is doing so with anodes which contain silver and gold content as high as those being refined at Canadian Copper Refiners Ltd. General Arrangement The original building consisted of two parallel bays, each 660 ft long and 60 ft wide, with a 24 ft pump bay extending along the west side of the building. A previous expansion, completed in 1939, was to the west of the original building and consisted of two 60 ft parallel bays 240 ft long. The length of these two bays was extended in 1954 to 540 ft, adding one stripper and 12 commercial sections to No. 2 electrical circuit. Space is still available for an additional six commercial sections and one stripper section. Construction on these seven sections started in September 1955, and will be completed early in 1956. The tank house capacity will then be 206,000 tons of cathodes per year. In the two tank houses there is a total of 900 cells distributed as follows—commercial, 468; stripper, 36; total, 504 in the No. 1 tank house (No. 1 circuit): commercial, 360; stripper, 36; total, 396 in the No. 2 tank house (No. 2 circuit). Section arrangement consists of nine cells to a tier and two tiers or 18 cells to a section. Anodes, weighing 620 lb each, are cast with a Baltimore groove and are refined by the multiple system. Solution enters each cell through a bottom inlet at one end and overflows at the top of the opposite end. Rate of solution flow through the cells is controlled by valves at each section to 4.5 gpm. Electrical power to the tank house is now provided by nine 675 kw motor generator sets. Each set consists of a 135-12 v 5000 amp dc generator, driven by a 980 hp 2300 v synchronous motor. When No. 1 circuit is operated at 22,000 amp, five sets, at 4400 amp each, are connected in parallel. As the maximum allowable current on No. 2 electrical circuit has recently been set at 18,000 amp, only four sets are connected in parallel on No. 2 circuit. A tenth set, now being installed, will be reserved as a spare, for use in either circuit as required. Recent Changes In 1949, the annual capacity of the tank house was increased from 112,000 to 132,000 tons by closing the

Jan 1, 1957

By W. R. Kanne, P. R. Strutt, R. A. Dodd

A study has been made of the creep behavior of ß-NiAl of stoichiometric composition in the temperature range 475" to 775°C. Single crystal tensile specimens were deformed under a constant applied load. From slip line studies and tensile axis rotations it was concluded that the crystals deformed by pencil-glide on (100) and (110) planes with a (100) slip direction. Electron microscope thin-film studies showed evidence for profuse cross-slip, and the dislocation Burgers vector was determined to be (100). There was no tendency for the dislocations to lie in definite slip planes, although a strong alignment of segments of edge dislocations was frequently seen. The activation energy for creep in the temperature range 475" and 775°C was judged to be much lower than that for creep at temperatures above 775°C. It is possible that the rate-controlling process is the thermally activated cross-slip of screw dislocations. In recent years considerable interest has developed in the possible use of high-melting-point intermediate phases for high-temperature applications necessitating good creep resistance. Ordered alloys such as Ni3A1, modified by alloying additions, have received particular attention, culminating in the recent development by Pratt and Whitney of single crystal turbine blades.' The intermediate phase NiAl has considerable theoretical interest from a creep standpoint because not only is it ordered up to the melting point (-1650°C at stoichiometry) but it contains a high concentration of structural vacancies on the aluminum-rich side of stoichiometry. These vacancies are known to exert a strengthening effect under some conditions,' but in addition they might be expected to affect high-temperature creep rates through dislocation climb. The present paper deals with the slip line and dislocation structures developed during creep of stoichiometric NiAl between 475" and 775C, and with possible attendant creep mechanisms. MATERIALS AND TEST TECHNIQUES NiAl ingots of stoichiometric composition were prepared by induction melting 99.996 pct A1 and 99.98 pct Ni (kindly provided by the Kaiser Aluminum Co. and the International Nickel Co., respectively) in 100 cc pure alumina crucibles in a helium atmosphere. No chemical melting flux was used in order to avoid con- tamination of the alloy. Each ingot was slowly direc-tionally solidified to avoid porosity and to produce very large grains up to 2 in. in length. In this way, ingots were obtained which were suitable for sectioning into single crystal creep specimens. Chemical analysis indicated that all alloys were within the range 49.6 0.3 at. pct Ni. For the present investigation, identical rectangular single crystal blanks were cut vertically from one particularly large crystal in an ingot using a series of three diamond cut-off wheels with appropriate spacers. Thus, all sliced crystals had the same orientation with the tensile axis parallel to the long side of the rectangle; the initial orientation of the tensile axis is shown as S, in Fig. 1. The blanks were homogenized at 1250 to 1300°C for 48 hr and subsequently cooled slowly to room temperature. A gage length of rectangular cross section and about 0.75 in. in length was prepared on each specimen using an electrolytic lathe and a dilute nitric/phosphoric acid electrolyte. The gage length cross section was 0.250 0.0005 in. by 0.040 ± 0.0005 in. Holes to accommodate tensile grips were electrolytically machined using a rotating acid jet. After creep testing, this same technique was used to cut discs for transmission electron microscopy. Thus, all shaping was done electrolytically, the only exception being that (122) sections were spark-machined from some speci-

Jan 1, 1970

By Benjamin Smith Lyman

PAGE I. ANCIENT HISTORY,........... 57 Accepted Fables ; Babylonian Mapping ; First Surveying. II. COMPASS,.............59 Chinese Invention; Marco Polo; First European Compasses ; Early Knowledge of Variation ; Plotting with the Compass ; Hanging-Com pass ; Supposed Gravitational Error ; Merit of the Compass. 111. Telescope,............63 Origin (Roger Bacon, Porta, Digges, Lippershey, Galileo) ; Improvements (Micrometer and Cross-Hairs, Rittenhouse's First Telescope, Platinum Cross-Wires, Telescopic Sights, Reflecting Telescope, Achromatic Lens, Kellner Lens in Erecting Telescope, Inverting Telescope). IV. Theodolite,.............81 Origin ; Derivation of the Name. V. TranBit,.............86 First Surveying Transit; Edmund Draper. VI. Instrument-Parts,..........89 Ramsdenl a Dividing Engine ; Fineness of Graduation ; Conical Gradua tions ; Full Vertical Circle ; Leveling-Screws ; Shifting Tripod-Heads ; Hoffman-Harden Tripod-Head ; Heller & Brightly's Improvements. VII. OTHER INSTRUMENTS AND APPLIANCES,......100 Sunflower; Plummet-Lamp; Chains and Tapes. VIII. SCOTT'S TACHYMETER,.........104 IX. Nomenclature and Classification,.......105 Names; Grouping. Surveyors are much indebted to Mr. Scott for so vigorously attacking the subject of the origin and history of mine-surveging instruments, the compass, the telescope, the transit and other apparatus. He is not afraid of the dark, nor yet foolishly bashful in broad daylight. He does not recoil from the obscurities of the earliest times and the most difficult historical points; and, with commendable public spirit, he makes his own invention known to his fellow-workers, and challenges their criticism. He has, moreover, elicited the praiseworthy emula-

Jan 1, 1902

By Davenport, Edmund S.

THE specific aim of this work has been to study certain forms of internal deterioration which occur in tungsten filaments when subjected to high temperatures under various conditions, and to determine the cause of such deterioration. This would naturally lead to a consideration of methods of control or means of prevention of the phenomena. The deterioration was first noticed as a peculiar condition in the grain boundaries of certain tungsten filaments which had been burned t for long periods in lamps. In these cases it was noted that the grain bound aries contained a number of minute black dots which were at first thought to be non-metallic inclusions. It was demonstrated later that these dots or pits, as they were designated, could be readily produced by burning filaments in an atmosphere of flowing hydrogen. This method was convenient and adequate for the purpose, and was adopted for most of the work discussed in this paper. The evidence accumulated to date indicates that these pits arc actually small voids between the grains. The purpose of this paper is to, present the information on this interesting subject gathered during a period of about 3 years. As applied to tungsten, the subject appears to be of considerable importance, experience having shown that this pitting is frequently associated with poor performance of filaments in lamps.

Jan 1, 1927

By AIME AIME

THE iron and steel activities of this meeting opened on Monday morning with the steel melting session," with G. B. Waterhouse in the chair and A. L. Feild as vice-chairman. N. A. Zeigler's paper, on "Gases Extracted From Iron Carbon Alloys by Vacuum Melbing," created considerable interest and the discussion in general was most favorable. H. W. Gillett and Louis Jordan of the Bureau questioned the method of analysis for oxygen and C. H. Herty, Jr., and J. M. Gaines, Jr., I , expressed their surprise at the large amount of oxygen with the high ,carbon alloys. Whether the gases found existed as gas in blowholes or as oxide inclusions was also a moot point. A paper by C. D. King, on "Basic Open-Hearth Yields," was an extremely interesting and valuable contribution. It was discussed by C. E.Kinney and A. L. Feild and numerous questions were asked throughout the presentation. Mr. King then presented a critical analysis of the losses together with novel methods of calculating them. A paper by C. H. Herty, Jr., and J. M. Gaines, Jr., on "Unreduced Oxides in Pig Iron and Their Elimination in the Basic Open-Hearth Furnace," followed. The question of silicates and oxides in pig iron created much discussion, notably by Henry Hanson and W. J. MacKenzie, who gave figures on chipping cost as a function of blast-furnace practice.

Jan 1, 1929