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By Thomas Brownfield

The chronicle of the oil industry in Oklahoma in 1939 is one of declining production bolstered by strenuous efforts to find new pools or new producing horizons in the older, highly exploited, relatively shallow areas. Production, which averaged 608,000 bbl. daily in 1937, the post-depression peak, declined to 463,000 bbl. daily in 1938 and continued downward to 421,000 bbl. daily for the year 1939. While the state could produce somewhat more oil if demand warranted it, production in most areas is at as great a rate as could be justified on the basis of maximum efficient operation. Outlet for Oklahoma crude has been seriously affected by the increase in production in the Illinois Basin. Crude movements from the Kansas-Oklahoma area to the northeastern consuming centers averaged 314,000 bbl. daily in 1939 as compared with 397,000 bbl. in 1938. Oklahoma producers bore the brunt of this loss of market. The general production decline was accentuated by the August, shutdown of 15 days, precipitated by the action of a major crude purchaser posting a 20-cent reduction in the price of crude. The Oklahoma Corporation Commission, in collaboration with other members of the Interstate Oil Compact Commission, ordered a complete shutdown of all producing wells in order to prevent the premature abandonment of stripper wells, which might result from a prolonged period of production at prices below operating costs. The reduction in price was forestalled, most major purchasers failing to meet the cut in view of the shutdown, and the company initiating the move rescinded its action. Fear of the shutting in of oil wells, especially those making considerable quantities of water, is generally unfounded in the light of a superficial examination of the records. Oklahoma's unprorated pools, which produced 154,000 bbl. daily in May, declined to 152,000 bbl. in June and 143,000 bbl. in July. Immediately following the shutdown, the same pools produced 150,000 bbl. in September, 145,000 in October, and 145,-000 in November. Thus, while the half month's production loss in August has not been made up, the wells apparently hare produced at a greater rate since the shutdown than immediately preceding it.

Jan 1, 1940

By P. E. Fitzgerald, Dana G. Hefley

An apparatus has been developed which permits selective acidizing of producing formations and determination of the relative permeability of a formation by fluid injection. This apparatus, known as the "Electric Pilot," is an electrical fluid-interface locator, the electrical circuit of which is completed when the elec-tcodes in the well are in contact with a conductive fluid, but is not completed when immersed in a nonconductive fluid. Results are given for a number of typical acidizing applications in which the interface locator was used to control the acid. Description of the method used in making a permeability survey and the data obtained are also included. Introduction Early in the development of the acidizing industry it was recognized that one of the most important problems to be solved in the chemical treatment of any well was the determination of what part of the exposed section in the well should be treated and the selection of the technique that would insure treatment of the desired section. One of the first techniques adopted for assisting in proper placement of the acid, which is still widely used, is the method of confining the acid to the producing section by means of a column of oil, which is known as the Carr1 method of treatment, named for its inventor. In the Carr method it is not possible to selectively acidize a specific horizon in a multiple pay zone, so many improvements and modifications of the basic Carr technique were devised to accomplish this result. In order to overcome the possibility of the acid entering an undesirable section at the bottom of the well, organic gel plugs2 were introduced into the bottom of the well, which served to halt the entrance of the treating solution into these zones.3 These chemical gels later liquefied because of bacterial action and were removed from the well with the produced fluid. In fields where a gas cap existed, as in the Goldsmith field of Ector County, Texas) it was imperative that a treating technique be evolved that would confine the acid to the oil-producing zone below the gas cap. By so doing the oil-productive capacity of the wells could be improved without greatly increasing the gas production and this would lower the gas-oil ratio of the well. ~h, technique employed4 to achieve this effect utilized two pumps, one for pumping the chemicals down the tubing, the other to simultaneously force oil down the annular space to confine the acid to the oil pay, thus preventing the chemicals from entering the upper gas-bearing zones. Experience has indicated that gas wells and certain types of oil wells respond better to ''stage treating.'' This technique utilizes the introduction of a certain volume of acid and then allows the well to ''clean up99 before introduction of the next and usually successively larger stages. Where such methods are employed steps must be taken to keep a disproportionate share of the acid from entering the most permeable and previously treated Zones. A special tool

Jan 1, 1944

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,' of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in terms of mass per unit time by w Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L Slope of energy grade line by s Mass per unit volume of solids by m8. Mass per unit volume of liquid by mi Mass per unit volume of mixture by 11 We then have as an expression for the solids load: w = pqmm The work done per unit mass, p, on the mixture to move it along the pipe the

Jan 1, 1947

By Charles White Merrill

Money is one of the most .pervasive elements in human life. The compensation for a workman's daily efforts is expressed as a wage and is measured in money. What an individual may consume depends largely on his income which also is expressed in terms of money. If he would start an enterprise its size would be determined in large part by the amount of money available for capitalization and its success would be measured in net income expressed in money. In short, the most widely used common denominator for comparing merchandise, real property, jobs, enterprises, gross national product, or personal success is money. Despite this use of money by almost everyone past infancy to express his appraisal of the in- numerable values he sees around himself, the complexity of the device called money makes its nature and workings a mystery to most users. Some of the difficulty in understanding money results from the fact that it has three distinct but interrelated functions: (1) a medium of exchange, (2) a unit of account, and (3) a store of value. In many statements about money the discussion moves from one function of money to another without the speaker or the listener recognizing the changes. For example, the housewife says "We took the money we had saved up to make the down payment and after twelve monthly payments the television will be ours." When this housewife spoke of "saving up the down payment" she referred to money in its function as a store of value; when she spoke of "the down payment" she was using money as a medium of exchange; and when she referred to the "monthly payments" she spoke of money as a unit of account. Another cause of confusion in understanding money is its paradoxical nature in that it represents both the power to produce and the power to consume. For example, when the rich uncle dies one nephew may regard his inheritance as an opportunity to buy a farm and leave the ranks of the wage earners to become an independent producer, whereas the other nephew may regard his portion as the means of living in complete leisure as a pure consumer.

Jan 1, 1959

By Robert G. Woolery, S. Ramachandran, James A. Weber, Donald J. Hansen

Union Carbide began looking seriously at heap leaching in 1971. At that time some 1.6 million tons of mineral averaging 0.40 kg U3O8 /t) (0.80 lb U3O8) were stockpiled at various sites around the Gas Hills, Wyoming, uranium facility. Several alternatives for economically exploiting these reserves were considered and heap leaching seemed the most attractive. At the time, Western Nuclear Inc. was operating a heap leaching project nearby in which the low-grade ore was leached with sulfuric acid, the product liquor treated in an on site solvent extraction circuit, and the stripped product liquor re-acidified and recycled. The operators were most generous in allowing us to visit their site and in answering our questions. They were also helpful in discussing the problems associated with heap leaching. Their experience confirmed our belief that it was possible to operate a heap leach profitably on material too low grade for conventional milling. Based largely on what we had learned from Western Nuclear's experience, we undertook a program involving a laboratory study to determine the characteristics of the mineral stockpiles, as related to heap leaching, and to conduct a pilot heap leach adjoining our Gas Hill mill. Basically the pilot heap construction and leach procedure were similar to those described in the literature l, 2 with certain modifications to permit detailed monitoring of the operation. In May 1972, the decision was made to construct a test heap. The UCC plant at Gas Hills, Wyoming was selected as the most appropriate site for the following reasons: (1) The necessary equipment for heap construction was readily available. (2) Stockpiles of mineral from the various mining sites were on hand: (3) Construction of the heap could be made on an abandoned tailings pile minimizing problems associated with environmental approval. (4) Proximity of the mill would allow product streams to be absorbed into existing circuits. (5) Necessary utilities were available. (6) Tailings water from the existing circuit could be used for leaching. A program for the pilot heap leach was developed with the following primary objectives. (1) To confirm the data obtained in bench-scale tests particularly on uranium recovery, acid usage, and ore permeability. (2) To evaluate plastic liners used for base preparation. (3) To investigate heap configuration variables. (4) To determine heap leach amenability of the different mineral sources. (5) To establish the economics of heap leaching low-grade ores.

Jan 1, 1979

By Thomas Brownfield

The chronicle of the oil industry in Oklahoma in 1939 is one of declining production bolstered by strenuous efforts to find new pools or new producing horizons in the older, highly exploited, relatively shallow areas. Production, which averaged 608,000 bbl. daily in 1937, the post-depression peak, declined to 463,000 bbl. daily in 1938 and continued downward to 421,000 bbl. daily for the year 1939. While the state could produce somewhat more oil if demand warranted it, production in most areas is at as great a rate as could be justified on the basis of maximum efficient operation. Outlet for Oklahoma crude has been seriously affected by the increase in production in the Illinois Basin. Crude movements from the Kansas-Oklahoma area to the northeastern consuming centers averaged 314,000 bbl. daily in 1939 as compared with 397,000 bbl. in 1938. Oklahoma producers bore the brunt of this loss of market. The general production decline was accentuated by the August, shutdown of 15 days, precipitated by the action of a major crude purchaser posting a 20-cent reduction in the price of crude. The Oklahoma Corporation Commission, in collaboration with other members of the Interstate Oil Compact Commission, ordered a complete shutdown of all producing wells in order to prevent the premature abandonment of stripper wells, which might result from a prolonged period of production at prices below operating costs. The reduction in price was forestalled, most major purchasers failing to meet the cut in view of the shutdown, and the company initiating the move rescinded its action. Fear of the shutting in of oil wells, especially those making considerable quantities of water, is generally unfounded in the light of a superficial examination of the records. Oklahoma's unprorated pools, which produced 154,000 bbl. daily in May, declined to 152,000 bbl. in June and 143,000 bbl. in July. Immediately following the shutdown, the same pools produced 150,000 bbl. in September, 145,000 in October, and 145,-000 in November. Thus, while the half month's production loss in August has not been made up, the wells apparently hare produced at a greater rate since the shutdown than immediately preceding it.

Jan 1, 1940

By H. M. O?Harra

Experiments have shown that the use of enriched air would be of particular benefit in the roasting of zinc ores for the manufacture of sulfuric acid. Enriched air increases capacity of furnace, decreases fuel consumption, and increases SO2 content of roaster gas. THE work here described had for its purpose the procuring of data from which some quantitative estimate might be made of the results obtainable by using oxygen-enriched air in roasting zinc ores on a commercial scale. The principal metallurgical advantages of using enriched air in roasting zinc ores would be: 1. The rate of roasting would be increased, with consequent gain in the capacity of the roasting furnace. 2. As less air would be required for roasting, the volume of hot gases leaving the furnace and the heat carried out of the furnace as sensible heat in these gases would be less per ton of ore roasted; partly for this reason and partly because of the increased quantity of heat generated in the furnace by the larger amount of ore that could be roasted, the consumption of fuel by the furnace would be lessened; and by the use of air sufficiently enriched with oxygen the necessity of using fuel might be entirely obviated. 3. Roaster gas having a higher SO2 content could be produced; this would make possible greater capacity and more economical operation of the sulfuric-acid plant. Certain phases of the application of enriched air to roasting can be worked out only by experimenting with a furnace of commercial, or at least semicommercial, size. Thus the precautions necessary to secure proper distribution of heat in the furnace;1 the volume of enriched air,

Jan 3, 1924

ALTHOUGH iron ores occur in many parts of Sweden the two principal deposits are those at Grängesberg (see accompanying map) and at Kiirunavaara-Gellivare. Both of these deposits are con-trolled by a holding company, Trajikaktiebolaget Grängesberg-Oxelösund, which sells annually about one-half of, the iron in ore traded in world markets and has a fleet of 24 seagoing ore boats, with a total dead weight capacity of 172,500 tons. Because of its pre-eminence in international iron ore trade a brief outline of its operations will be of interest. ORE DEPOSITS OF GRÄNGESBERG The Grängesberg deposits have been known since the Middle Ages, but up to the middle of the nineteenth century the ore was transported from the mines to the furnaces in winter in. sleighs. Railway building began in 1853, and in 1873 the complicated structure of capital investment was tottering, but the basic bessemer process created a demand for ore high in phosphorus, as this is, and refinancing on the initiative of Sir Ernest Cassel put the principal operation on a firm footing. There are numerous property owners in the district, but they are all controlled by a joint administration board, and 90 per cent of the ore is produced by a subsidiary of the Grängesberg company from proper-ties which it either owns or leases. The ore, a mixture of magnetite and specular hematite, is in lenticular bodies, dipping about 70 E. in. fine-grained gneiss, and contains about 61 per cent Fe and 1 per. cent. P. There are two principal veins; the total ore reserves to a depth of 500 meters are estimated at 150,000,000 tons. As seen in the accompanying illustration, the ores were first worked in open cuts, but eventually underground shrinkage stoping was adopted. Difficulties were encountered in the support of the workings and from the freezing of the ore in the stopes, so slicing was adopted, although block-caving is used in one place. Most of the mining is now done at a depth "of 150 to 190 meters. The present annual output is about 1,400,-000 tons of crude ore, which yields about 1,000,000 tons of concentrate after being subjected to magnetic con-centration and jigging.

Jan 1, 1927

By W. L. Falke, A. A. Cochran

A basic problem in connection with manganese is to find economical ways to utilize domestic resources. As a part of its program to conserve domestic mineral resources and to reduce dependence of foreign sources, the Bureau of Mines had conducted extensive research to broaden manganese technology and to find economical, technically-feasible processes for recovering manganese from these low-grade resources. This paper analyzes some of the major difficulties which have been encountered in previously proposed processes for treating low-grade manganese ores and discusses a process which may circumvent many of these difficulties. Vast deposits of low-grade manganese ore exist in the U.S., yet domestic production is insignificant compared to the more than 2 million tons of ore imported yearly into this country.' A basic problem in connection with manganese is to find economical ways to utilize domestic resources.2 As a part of its program to conserve domestic mineral resources and to reduce dependence on foreign sources, the Bureau of Mines has conducted extensive research to broaden manganese technology and to find economical, technically feasible processes for recovering manganese from these low-grade resources. Although a great many processes have been proposed for treating domestic manganese resources, a failing of many of them is that too many steps are involved for economical operation.' Some of the major processes that have been proposed were reviewed by Norman and Kirby.3,4 Chlorination processes have become increasingly useful as a means of recovering valuable metals from raw materials, because the relative cost of chlorine (Cl2) and hydrogen chloride (HC1) has decreased through the years, and because the corrosion resistance of materials of construction has been improved. The objective of these processes is to form metal chlorides that are more volatile or water-soluble than are the corresponding oxides or sulfides; these compounds then can be separated from the gangue or other minerals by volatilization or leaching. Several chlorination processes for the recovery of manganese have been investigated by the Bureau of Mines5-8 and others; the major chlorination processes were reviewed in 1963, and their main features were compared.4 In all these processes, the simultaneous leaching or volatilization of iron chlorides with manganese chloride (MnCl2) is a problem, and the more iron that is present, generally the greater the problem. The leaching step is not highly selective, and the product of this step is, at best, suitable only for the production of ferromanganese or for further processing to remove iron. The product of the volatilization step is, in all cases, a mixture of iron and manganese chlorides that contains too much iron even for ferromanganese production, and must undergo additional processing steps for iron removal. THERMODYNAMIC CONSIDERATIONS The literature indicates that the desired chlorination reactions will proceed at practical rates at temperatures near 1200°K (927°C). Free energy changes at 1200°K for the critical steps of the reduction and chlorination reactions9-11 follow: Most of the research on chlorination processes for manganese recovery has been done with HC1 rather than with C12, for the following three main reasons: 1) Recycling of the chlorinating agent, which is economically vital, appeared to be easier for HC1

Jan 1, 1968

By P. E. Fitzgerald, Dana G. Hefley

An apparatus has been developed which permits selective acidizing of producing formations and determination of the relative permeability of a formation by fluid injection. This apparatus, known as the "Electric Pilot," is an electrical fluid-interface locator, the electrical circuit of which is completed when the elec-tcodes in the well are in contact with a conductive fluid, but is not completed when immersed in a nonconductive fluid. Results are given for a number of typical acidizing applications in which the interface locator was used to control the acid. Description of the method used in making a permeability survey and the data obtained are also included. Introduction Early in the development of the acidizing industry it was recognized that one of the most important problems to be solved in the chemical treatment of any well was the determination of what part of the exposed section in the well should be treated and the selection of the technique that would insure treatment of the desired section. One of the first techniques adopted for assisting in proper placement of the acid, which is still widely used, is the method of confining the acid to the producing section by means of a column of oil, which is known as the Carr1 method of treatment, named for its inventor. In the Carr method it is not possible to selectively acidize a specific horizon in a multiple pay zone, so many improvements and modifications of the basic Carr technique were devised to accomplish this result. In order to overcome the possibility of the acid entering an undesirable section at the bottom of the well, organic gel plugs2 were introduced into the bottom of the well, which served to halt the entrance of the treating solution into these zones.3 These chemical gels later liquefied because of bacterial action and were removed from the well with the produced fluid. In fields where a gas cap existed, as in the Goldsmith field of Ector County, Texas) it was imperative that a treating technique be evolved that would confine the acid to the oil-producing zone below the gas cap. By so doing the oil-productive capacity of the wells could be improved without greatly increasing the gas production and this would lower the gas-oil ratio of the well. ~h, technique employed4 to achieve this effect utilized two pumps, one for pumping the chemicals down the tubing, the other to simultaneously force oil down the annular space to confine the acid to the oil pay, thus preventing the chemicals from entering the upper gas-bearing zones. Experience has indicated that gas wells and certain types of oil wells respond better to ''stage treating.'' This technique utilizes the introduction of a certain volume of acid and then allows the well to ''clean up99 before introduction of the next and usually successively larger stages. Where such methods are employed steps must be taken to keep a disproportionate share of the acid from entering the most permeable and previously treated Zones. A special tool

Jan 1, 1944

By R. N. Orava, L. A. Rice

One of the difficulties encountered in the preparation of titanium specimens for transmission electron microscopy is the formation of a surface hydride phase during the thinning operation at ambient temperatures. This necessarily destroys contrast, obscures features of interest, and often leads to misleading results. By cooling the polishing solution to temperatures below 0°C this problem has been overcome in the past.' However, since the cooling procedure is somewhat inconvenient, a technique was developed at this laboratory for thinning titanium at room temperature. The material examined was A-70 grade commercial-purity titanium purchased from the Crucible Steel Co. of America. The mechanical and thermal treatments, described elsewhere,* yielded a recrystallized grain size of 4 to 5 p. A slice, 0.5 mm thick and 1.5 mm wide, was cut from a bulk specimen2 using a diamond cut-off wheel. The sample was then coated with paraffin wax. Two concentric holes, about 0.03 in. in diam, one at each broad face, were scribed through the wax to the surface of the metal. Subsequent polishing for several minutes resulted in a preferentially thinned or "dimpled" region. Thereafter, the wax was removed and the entire specimen polished to the point of breakthrough at the base of the dimple. Polishing was effected chemically in a solution comprised of 30 pct HF and 70 pct HN03 by volume, kept at room temperature (20" to 25OC). The specimens were mounted in a stereo holder and examined at 100 kv in a Philips l00 B Electron Microscope. A micrograph from a foil prepared in this way is shown in Fig. 1. Several points are noteworthy. This electron micrograph shows neither evidence for the formation of a hydride phase due to specimen preparation nor, moreover, any obvious surface contamination in electron-transparent areas of a freshly prepared foil. Localized variations in foil thickness observed in some specimens seem to be a characteristic of this chemical polish but do not seriously detract from foil quality. With care, the lifetime of a good specimen exceeded 2 hr in the beam. On the other hand, the exposure of a sample to air at atmospheric pressure, when still warm from the beam, clearly resulted in the deterioration of a formerly clean area just previously in the field of view. An example in Fig. 2 of a contaminated region reveals features, and "spots", not unlike those observed in much purer material by Yamane and Ueda. Consequently, one can conclude that: 1) titanium can be thinned satisfactorily in a solution at room temperature for transmission electron microscopy; 2) large-scale precipitates, >50A in diam, are absent in commercial-purity, recrystallized, furnace-cooled, A-70 titanium containing about 3800 wt ppm interstitials and the same amount of iron; 3) care must be taken not to attribute surface contaminants to precipitates. The authors are grateful to the Franklin Institute for financial support and permission to publish.

Jan 1, 1967

By Jack A. Kanz, Samuel R. Shortes, E. C. Wurst

Zinc diffusion into GaAs from the vapor phase through reactively sputtered SiO, films Izas been inrestigated for various zinc pressures at 1000"C. Diffusions were carried out in evacuated quartz ampules with elemental zinc and with various Zn-Ga alloys. Sputtered films have proved effective in lowering surface concentrations and preventing surface deterioration during diffusion. Approxin.ate complenzentary error-function impurity distributions were found when surface concentrations were reduced to approximately 3 -—4 x 10" atoms cm'3 or less with dilute zinc sources or by diffusion through SiO, films. Zinc difhsions in GaAs tlzrough SiO, films produced more planar diffusion fronts than were observed with uncoated zuafers under identical conditions. A) Surface Deterioration During Diffusion. Device technolorn-- in compound semiconductors has been hampered by several difficult problems not normally encountered with silicon and germanium. 111-V and 11-VI compound semiconductors generally have volatile constituents at temperatures used for diffusion. To achieve specific junction depths in reasonable lengths of time, or to obtain particular impurity distributions and doping levels, it is often necessary to use temperatures which cause surface deterioration and erosion of the wafer. The resultant surfaces are generally unsuitable for devices. The type and extent of surface damage depends to varying degrees on the particular environment of diffusion and on the particular impurity in question. The most common type of deterioration results from simple dissociation of the gallium and the arsenic during diffusin.' The deterioration process is significantly enhanced by thermal gradients in the diffusion zone, usually resulting in vapor transport and condensation at the cooler portion of the system. The resulting damage sometimes appears as a general haziness over the surface or an array of pyramidal etch pits characteristic of the particular orientation. Occasionally, erosion of the surface occurs uniformly over the wafer so that surface loss is not readily apparent. Workers in GaAs usually resort to the sealed-ampule type of diffusion system3"5 rather than the more conventional open system used for silicon and germanium. Another type of deterioration results from direct interaction of the impurity with the surface. The form of deterioration may vary from surface alloying, as in the case of zinc, to direct chemical attack and vapor transport of the reaction products, as in the case of sulfur. Surface alloying with zinc occurs at higher temperatures and source vapor pressures.4 The process is largely controlled by the amount of free gallium on the surface of the wafer in equilibrium with the arsenic lost to the system. The group-VI elements, sulfur, selenium, and tellurium, are commonly used donor impurities for gallium arsenide. These elements are characterized by high vapor pressures and high chemical

Jan 1, 1964

By Louis E. Toth, Alan W. Searcy

A modification of Le Claire's microscopic model for self-diffusion is developed in a form suitable for prediction of activation energies for diffusion in disordered substitutional solutions as well as in pure metals Bonding is considered as a localized interaction, and the energy of bonding between atoms of different types is taken as the arithmetic mean of the energies in the pure elements. The activation energies for vacancy formation and migration in substitutional alloys are shown to depend on the empirical constants developed for self-diffusion when the equations are adjusted for the mole fractions of the two elements. The differences in activation energies between alloys and pure metals are calculated and compared with the experimental differences. MaNY different correlations have been proposed for use in predicting activation energies for self-diffusion or diffusion in dilute alloys. Usually theories for self-diffusion cannot be applied to alloys nor can theories for dilute alloys be satisfactorily adapted to self-diffusion. We present here a correlation scheme which can approximate activation energies for diffusion, vacancy formation, and atom migration in substitutional-alloy systems of any concentration. The activation energy for self-diffusion QD has been correlated to the melting temperature Tm,1, 2 the sublimation energy L,,l the heat of fusion Lf,3 crystal structure, and valence.4 Working with mechanical parameters Le claire5 proposed that the activation energy for vacancy migration Qm could be related to appropriate shear moduli. He also proposed that the activation energy for vacancy formation 62, is proportional to Ls. THEORY We consider that Le Claire's expression for 4, is satisfactory. But an alternate expression for Qm in terms of the bulk modulus is just as satisfactory from a theoretical point of view for the prediction of the energy of vacancy migration in pure metals and can more easily be used for alloys. In order for Qm to be a minimum, not only the moving atom but also adjacent and nearby atoms must undergo distortions. These distortions are neither isotropic as is implied by a correlation with the bulk modulus nor unidirectional as implied by a correlation with a shear modulus, but are complex and multidirectional. A complete analysis in terms of the bulk modulus and all the shear moduli is not yet available. In a first approximation in which only one parameter is used, there is no obvious theoretical reason to prefer the shear modulus over the bulk modulus. The bulk modulus is a much more convenient parameter for use in predicting Qm. Shear moduli are known for less than half of the metallic elements while bulk moduli are usually available. Furthermore, because of the directional dependence of shear moduli, their estimation for alloys appears inherently more difficult than is the estimation of bulk moduli. We assume that the bulk modulus Bs reflects an average value of the shear moduli such that the strain energy at the point of maximum distortion of the lattice during vacancy migration can be expressed as

Jan 1, 1964

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,' of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in terms of mass per unit time by w Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L Slope of energy grade line by s Mass per unit volume of solids by m8. Mass per unit volume of liquid by mi Mass per unit volume of mixture by 11 We then have as an expression for the solids load: w = pqmm The work done per unit mass, p, on the mixture to move it along the pipe the

Jan 1, 1947

By Philip S. Morse

THE Australian lead-smelting plants began to use charges carrying high zinc percentages somewhat earlier than was common with American plants. When lead smelting first started in Australia the immense deposits of oxidized lead at Broken Hill were available but later these oxidized ores became exhausted and were replaced by sulfides. This was before the days of flotation and these sulfides when concentrated carried about 1 per cent. of zinc to 6 per cent. of lead. The works of the Sulphide Corpn. at Cockle Creek was perhaps the pioneer in high-zinc slags but the Port Pirie plant of the Proprietary Co. was a close and very able second and has continued the work to the present time. The Cockle Creek lead furnaces remained in operation until shortly after the war when most of the Broken Hill companies combined and took over the Port Pirie plant to treat their concentrates. As it was the practice both at Cockle Creek and Port Pirie to carry 20 per cent. or more of lead on the charge, it was found that under these conditions of large amounts of zinc in the charge, good roasting was absolutely necessary and matte production was unwise. The Hunting-ton-Heberlein system of roasting was introduced and later the Dwight & Lloyd process. In this way, by roasting the sulfide portions of the charge as thoroughly as possible and making no matte, a method of smelting was developed where the slag over long periods carried as high as 20 per cent. ZnO without serious furnace troubles, and slags carrying even up to 33 per cent. ZnO have been reported. The zinc slags were not very different in appearance from the ordinary slags produced in American practice except that they gave off more smoke and fume on tapping. The slag usually ran hot and freely, coming from the tap in a full stream. Normally there was little crust on the sides of the furnace and the tuyeres gave no trouble though they were always rather dark. At the works of the Sulphide Corpn. it was the custom, on account of the labor situation, to shut down the furnaces during the Christmas holidays, thus giving a campaign of about 12 months. The furnaces at the end of the run usually showed no crust on the sides, though crusts always occurred at the back and front ends. A furnace campaign of 22 to 23 months has been made at this plant where both sides of the

Jan 1, 1929

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By P. R. Kalischer

In the production of alloys by powder metallurgical processes it is often necessary or desirable to include one or more cornponents that tend to form very stable oxides. Included in this group of metals are aluminum, manganese, magnesium, silicon, and a few others. In many cases it is not necessary for the sintered compact to have full density and the other physical characteristics of a cast alloy and in that event no undue precautions need be taken. If, however, it is required that the sintered part be entirely homogeneous, and have maximum density and physical properties, some method must be found to reduce the surface films of oxide, which are always present on the metals, and for the metals mentioned above this is always extremely difficult, if not impossible. since the production of high-density alloys is of particular interest to powder metallurgists, some experiments were undertaken to devise methods of reducing the refractory oxides. For this work, alloys of iron, nickel, and aluminum were chosen because aluminum oxide is one of the most stable oxides and because these alloys in the right compositions form excellent permanent magnets. All of the pertinent properties of the magnets could be accurately and delicately measured by density measuremerits and magnetic tests of sufficient scope to give a complete demagnetization curve. From such a curve the maximum coercive force, residual magnetism, and energy products could be obtained and compared with the same properties for a cast material. Any lack of homogeneity, density, or solid diffusion would be immediately apparent as a decrease in the magnetic properties. Nature and Removal of Surface Barriers The early work of the present investigation was based upon published processes.* In the published information it was claimed that by intimately mixing iron, nickel, and iron-aluminum alloy (50 Per cent Fe, So Per cent Al) powders, compressing such a mixture into the desired shape and size, and then sintering (heating to a temperature below the melting point of the component metals) in hydrogen at 1200°C. or higher, a sound, homogeneous article could be made having the desired physical size, shape, and strength, and good magnetic characteristics. However. this early work showed clearly that something in the powder compress was inhibiting diffusion and the process was almost entirely unworkable. In the initial phase of the investigation samples of iron powder, nickel powder, and iron-aluminum alloy powder were pressed into separate bars and sintered in hydrogen. Both the iron and nickel powders sintered

Jan 1, 1941

By Walter Miller

During the second year of America's active ia in the war the inain objectives of the petroleum-refining industry were again to provide the four most important product needs for war: 100-octane aviation gasoline, toluene for T.N.T. production, high-quality lubricating oils for the needs of aviation and the armed forces, and petroleum chemicals for synthetic rubber manufacture. All other materials supplied by the petroleum industry took an inferior priority ranking, depending on the degree of essentiality. To accomplish the above objectives the industry initiated and carried on a building program during 1942 and 1943 which will put those two years down in petroleum-refining history as the greatest construction years of all time, a record which will hold for decades to come. Led by an estimated total of $ 900,000,000 going into 100-octane and contributory plant investment, it is easy to picture a total of two billion dollars and more going into additional refining-industry plant investment during 1942 and 1943, most of it privately paid for, but a substantial part covered by R.F.C. Government money financing. Some of the new plants will not be completed until 1944, but a large proportion of the program was finished and put into operation, adding substantially to the output of the four principal products during 1943- lne most important and spectacuiar activity of the refining industry is of course the 100-octane gasoline program necessitated by the fact that nearly 90 per cent of the requirements of the United Nations is being furnished by the United States. Recent P.A.W. figures mention $goo,-ooo,ooo as the cost of the construction looking to increasing the output of this all-important munition. Thirty-four major plants have been completed, 11 of them during the last quarter of the year. Thirty-eight more will be completed and put into operation during the first four or five months of 1944. Twenty-two additional plants authorized during 1943 are already well under way and scheduled for production late in 1944. Early in 1942 the aim was to treble the capacity for making 100-octane gasoline then existing; later, plans were to make it five times as great, to match the increased warplane building program; and toward the end of the year the sights had been set at a still higher but unannounced target, representing six to seven times the early 1942 figure. Comparative figures issued at the end of 1943 enable one to guess a program by the middle of 1944 of eight to nine times the former production, with an ultimate production in 1944 of eleven or more times the early 1942 output. It became obvious in 1943 that building of new units would be materially slowed because of delay in getting construction materials and that they would reach operation at considerably later dates than

Jan 1, 1944

By James Pollock

The possibility of colloidal deposition of cinnabar has been neglected. In opalite deposits cinnabar exists in particles within or near the colloidal size range. Colloidal processes have been admitted to be operative in hot springs and low-temperature deposits similar to most cinnabar deposits. Mercuric sulphide sols are stable to elevated temperature and pressure, and in nature probably are chiefly affected by the electrolytes in solution, dilution of colloid, and presence of protective colloids as silica. The stability of mercuric sulphide sols to electrolyte is measured, and shown to be greatly increased by dilution and by the protective action of silica. Mercuric sulphide sols follow the Schulze-Hardy valency rule, therefore the precipitating power of complex mixtures of electrolytes can be computed roughly. The stability of mercuric sulphide sols is shown to be greater than the equivalent precipitating power of the electrolytes in a hot spring depositing mercuric sulphide. It is concluded that finely divided or earthy types of cinnabar are colloidally deposited. Possibility of Collodial Deposition While colloidal solution has been suggested as a means of transport for a number of minerals, such as the sulphides of antimony, arsenic, and iron, it has been neglected in connection with mercury. This is surprising because there exists positive evidence that cinnabar is present in certain deposits in particles of colloidal or near colloidal size. Colloidal transport could result from either the rapid dilution or neutralization of alkaline sulphide solution bearing mercury coming in contact with deep ground waters, or from colloidal solution near the original source and subsequent flow to the point of deposition unchanged. That the dilution of a mercuric alkaline sulphide solution will precipitate mercury sulphide in colloidal form has been demonstrated a number of times and is easily confirmed. The dilution hydrolyzes the alkaline sulphide Na2S to the corresponding hydroxide and hydrosulphide NaOH and NaHS, and the diminution of the concentration of alkaline sulphide reduces the solubility. It is not the contention of this writer that colloidal deposition is the mode of transport of mercuric sulphide in all mercury deposits, but that the stability of a mercuric sulphide sol, as demonstrated below, in waters that are similar to those that have deposited mercuric sulphide necessitates consideration of this possibility in many deposits, especially those of the hot-spring and so-called opalite type. Evidence of Colloidal Deposition Ross1 has shown that most deposits of quicksilver are formed at low temperatures and pressures, and probably by solutions that have relatively simple chemical compositions. In most cases the solutions are probably similar to those observed in hot springs, and in two instances—at Steamboat Springs, Nevada, and Sulphur Bank, California—hot springs have been reported by Becker as actually depositing cinnabar.2 Opalite deposits were believed by early investigators to be surface phenomena

Jan 1, 1949

By W. A. Boyer

If the safety factor of a mine hoisting rope is checked for the lowest depth, is the rope then safe for all levels? The answer here is no. A new set of volues is proposed. HOISTS for metal mines are seldom designed for one particular depth. They are intended for an ultimate load and depth with a given speed but are first used to hoist from a shallower depth, gradually working to the ultimate as development progresses. Sometimes the mine is not worked on the upper levels because of the kind and grade of ore, the mining of which depends on market conditions. In this case the hoist may be used in the same working shift to lift ore, for example, from the 1200 and the 4400-ft level. This means that the safety factor of the rope should be considered not only when it hoists from the greatest depth, but also when it hoists from the upper levels. According to present safety factor standards, a hoisting rope can have ample safety margin when hoisting from the lower or lowest level and still be shy of sufficient safety margin when hoisting from the upper levels. This will be shown by curves presented in this paper. The safety factor of a hoisting rope as defined by the U. S. Bureau of Mines' is the ratio of the static load to the tensile strength of the rope. It is calculated by dividing the breaking strength of the rope* In Fig. 1 curves A and B are plotted against the vertical suspended distance, which is the abscissa, and the factor of safety, the ordinates. Curve C is a smooth curve drawn through the midpoints of the steps of curve A and would more closely represent the minimum factor of safety for new rope for any particular depth. Curve D is a smooth curve drawn through the midpoints of the steps of Curve B and would more closely represent the minimum factor of safety when ropes should be discarded for any particular depth. If a new rope were applied with a minimum factor of safety, then the vertical distance or ordinate under curve D for any particular depth, divided by the vertical distance or ordinate under curve C, would represent the minimum percent of remaining area intact that could be allowed before the rope would be discarded. Also, that section of the ordinate between curve C and curve D is proportional to the section of the area of the new rope that could be allowed for wear and broken wires. Considering that point at 2500 ft on the chart, the ratio do to co would be the minimum percent of rope area remaining at the point of discard. The ratio of that portion of the ordinate cd to co would represent maximum percent of area of a new rope available for wear and broken wires.

Jan 1, 1955