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By Thomas S. Lovering

Hot springs, geysers, steam-emanating fumaroles- these creatures of nature were until recently little more than geological curiosities, but today they are exciting the interest of engineers around the world for it has been shown that such hyperthermal areas can be profitably harnessed by man to provide heat and power for his communities. Although water from hot springs was used centuries ago to heat some Korean dwellings, is being used today by one-third of Iceland's population for similar purposes, and is finding industrial applications in New Zealand and elsewhere, the most dramatic use of geothermal energy has been its conversion into electricity. The largest installation at the present time is that at Larderello, Italy, where the Count of Larderello began to experiment in the production of electricity from geothermal steam in 1904. Eight years later after a quarrel with the local power company, the Count installed his first steam turbine, a unit of 250 kw capacity.

Jan 9, 1965

By Thomas Elliott

It is now clearly evident that 1963 is witnessing a record year for prospecting and exploration in British Columbia and Yukon Territory, as substantiated by the following facts: ? About 60 large mining companies and many smaller ones are active in the search for new ore deposits. ? More than 500 prospectors and exploration geologists have been in the field this season. ? It is estimated that $10 million has been spent this year on the search for new ore deposits. ? Claim staking in B. C. to July 31, totalling 14,250, is up 2000 over 1962. ? Helicopter companies report a record total of 25 aircraft chartered to mining groups, with additional numbers used in oil-gas exploration. ? Leading assayers state that business with the industry is double that of the previous year. ? Diamond drillers report an increase of approximately 25% in the amount of contracts now on hand. The question is often asked, "Why is mining so active in Canada's Pacific Northwest?"

Jan 12, 1963

By Homer S. Frankhouser

Materials may be dredged, processed, stored and loaded for shipment at a single site in water depths ranging from 250 to 1200 ft. Solution mining may also be accomplished in deep sea situations. And, open air subsea mines with direct surface access may be developed and worked under the ocean floor. These concepts for deep sea mining are the results of extensive testing of individual components of the proposed systems by Dravo Ocean Structures, Inc., a subsidiary of Dravo Corp. A basic structure in these systems is the articulated platform. The platform consists of a gravity- type base and a buoyant cylindrical column which is constructed at an onshore location, towed to the offshore site and installed in a matter of days without use of heavy equipment. The base is also buoyant for towing purposes. This allows for gradual flooding to the point of negative buoyancy for positioning and anchoring to the ocean floor.

Jan 1, 1971

POSITIONS VACANT (Under this heading will be published notes sent to the Secretary of the Institute by members or other persons.) General manager .for a pyrite mine in Virginia, with experience in the mining and particularly in the washing of pyrite, required. Salary at start, $250 per month with perquisites. No. 28. Assayer and cyanide man for a Costa Rica plant. Salary $100. No. 29. Mine manager to develop gold-mining property in Witwatersrand, Transvaal. Salary £2,000 to £4,000, according to the experience and ability of the engineer. No. 31. Mine superintendent for a mine in Arizona containing copper, gold, silver ore. Will have entire charge of operation and development. Salary to start, $200 per month with perquisites. No. 32. .Mining engineer to make complete report and suggest necessary machinery for development and operation of a placer property; also to examine a copper property. Both properties in Peru, South America. No. 34.

Jan 4, 1915

By G. A. Savanick

Borehole (slurry) mining, where a tool incorporating a water-jet cutter and a downhole slurry pump mines phosphate through a borehole, was field tested in northeastern Florida as part of a US Bureau of Mines program to increase the domestic availability of critical minerals. About 1.5 kt (1700 st) of phosphate ore were mined at 23.6 t/h (26 stph) from a depth of 76 m (250 ft). The borehole was filled with water during mining for roof support and to assist slurry pumping. Underwater cutting was enhanced with an air-shrouded water jet. Mined cavities were backfilled to prevent subsidence after abandonment. Roof subsidence during mining was avoided only when the cavity was filled with water. Environmental monitoring before, during, and after mining included hydrological studies and measurements of water quality. Significant water quality changes occurred only in the zone being mined and were confined to the immediate area of the borehole.

Jan 1, 1986

The pellet plant, built by Dravo Corp, in Groveland, Mich., for the Hanna Mining Co., is now in full operation. Although using jasper, a type of iron ore with low iron content (approximately 28% Fe), the plant has a rated capacity of 1,250,000 tans of pellets per year. In its four years of existence, the nearby concentrating plant has seen its capacity increase from 700,000 tons to 1,500,000 tons a year as part of a major expansion program at Groveland. Utilizing the Dravo-Lurgi continuous grate process, the pellet plant produces %a-in. pellets of more than 60% iron content. The fully-hooded pelletizing grate, largest in the U.S., is 9 ft. 10% in. wide and 223 ft. 1% in. tong over 34 windboxes, with an effective hearth area of 2195 sq. ft. Drying, firing, and cooling functions are all carried out on this single piece of equipment.

Jan 4, 1964

By John Chipman, Wayne L. Worrell

Using recent thermodynamic data for the carbides and oxides of tantalum, columbium, and vanadium, the stable solid phases aboue 1300°K and at 1 atm CO(g) pressure in each M-C-O system have been determined. Pourbaix-Ellingham diagrams have been constructed and are used to estimate the minimum temperatures necessary to obtain these metals by carbothermic reduction under reduced pressures. TANTALUM, columbium, and vanadium react with carbon to form the M,C and MC carbide phases. The M,C carbides have narrow homogeneity ranges, but the MC phases possess considerable ranges of homogeneity.' The free energies of formation of the MC carbides were recently measured, and those of MC and M2Ccarbides have been estimated., (The notations MC and MC denote, respectively, the carbide compositions at the carbon-rich and at the metal-rich boundaries of the homogeneous MC phase.) In this paper, these carbide data are combined with thermodynamic data for the oxides to predict the stable phases in the Ta-C-O, Cb-C-O, and V-C-O systems at specified temperatures and oxygen pressures. M-C-O PHASE DIAGRAMS The solid phases in equilibrium above 1300°K for each M-C-0 system are shown in Figs. 1, 2, and 3. Above 1300°K, the vapor would consist entirely of carbon monoxide; and these phase diagrams were obtained by specifying the vapor pressure of CO(g) to be 1 atm. The temperature at which any two solid phases become unstable is specified along the two-phase equilibrium lines in each diagram. To simplify the diagrams, all solid phases in the Ta-C-O and Cb-C-O systems are represented as stoichiometric compounds. However, the homogeneity ranges of VO and VC are indicated, because they form a complete series of solid solutions.3 It also should be noted that the homogeneity range of

Jan 1, 1964

By G. Horvay, J. G. Henzel

Nomograms and charts are provided which permit rapid determination of the mold-casting interFace temperature and the speed of solidification when a semiinfinite ingot is cast into a semiinfinite mold. The physical properties of the mold, the frozen metal, and the liquid metal are not assumed to be equal, but they are assumed to be temperature independent. The results are correlated with charts prepared by Paschkis and Hlinka—using Columbia University's Analog Computer—for the freezing of finite ingot slabs when a constant surface temperature is impressed, and with experimental cooling curves obtained by Pellini in sand and iron molds for 7 in.-sq steel castings. x = coordinate directed to right of zero (freezing direction) y = coordinate directed to left of zero (melting direction) e = position of freezing front (in x direction) ? = position of melting front (in y direction) $ = temperature (function of position and time) ? = fixed temperature 2L = width of finite slab Y = a fixed position in the semiinfinite mold, or the width of a finite mold (see Section 7) Properties y = density, lb per cu ft c = specific heat, Btu per lb deg F k = conductivity, Btu per ft hr deg F X = heat of fusion, Btu per lb k = k/Yc = temperature diffusivity, sq ft per hr* b - = = - heat diffusivity, Btu per sq ft deg F Vhr Subscripts 2 = pertaining to the still-liquid metal (case of freezing), or the medium which furnishes the heat (case of melting) 1 = pertaining to solidified metal (case of freezing), or to molten metal (case of melting)* 0 = pertaining to mold (case of freezing), or to the still-solid metal (case of melting) f = pertaining to fusion condition (eg.,?f - fusion temperature) i = pertaining to the location x = 0 (the mold-ingot "interface")

Jan 1, 1960

By J. Chipman, T. P. Floridis

The equilibrium of gaseous H2O-H2, mixtures with liquid iron is used to establish the activity coefficient of dissolved oxygen as log f o = — 0.20 [%O]. Discrepancies in earlier work are explained or eliminated. Additions of cobalt, molybdenum, tungsten, and platinum increase the activity coefficient of oxygen whereas copper and gold decrease it. THE importance of dissolved oxygen as a refining agent in liquid steel has led to a number of studies of its thermodynamic properties. It is the purpose of this paper to examine the effects of a number of alloying elements on the activity of oxygen in the steel bath. As a preliminary to the study of the effects of alloying elements, it seemed desirable to examine certain minor discrepancies in the published data on the properties of oxygen in simple iron-oxygen solutions. The activity and free energy of oxygen in liquid iron were studied by chipman1 in 1933 by means of the reaction h2(g)+-o = H2O (g);K; = PH2O/PH2[% O] [1] The data were in fairly good agreement with calculations from results on the similar reaction CO (g) + O- = CO2 (g); K2 =P CO2/P CO [% O] [2] which had been obtained by Vacher and Hamilton2 as an incidental part of their study of the reaction of carbon with oxygen dissolved in the melt. It was found that the equilibrium ratio k1 was not a true constant but varied with the oxygen concentration in the manner shown in Fig. 1, which is copied from the original paper. This early observation has been refuted in subsequent work but is confirmed in the present paper, and an explanation of the discrepancy can now be presented. In Fig. 1, the two points on the extreme right were disregarded in drawing the

Jan 1, 1959

By T. J. Koppenaal, M. E. Fine

The critical resolved shear stress of aCu-A1 single crystals has been investigated as a function of composition, testing temperature, and strain rate. The strength, and its temperature and strain rate dependency increase markedly with aluminum content. The solid solution strengthening at 4.2°K is much larger than at room temperature. The activation energy for the low temperature dislocation cutting process increases from about 0.17 ev for a 0.5 at. pct A1 alloy to about 0.50 ev for a 14 at. pct A1 alloy. This result is attributed to an increase in the width of extended dislocations due to the known decrease in the stacking fault energy. In addition, the activation volumes at 4.2 and 77°K decrease with aluminum additions. Concerning the solid solution strengthening in the 296° to 373°K range, no definite conclusions are reached, but it is suggested that subgrains and cutting of subgrains are important. In a previous paper,' the authors reported that adding aluminum to copper produces two types of strengthening: a yield point effect and an increase in the lower yield stress. This behavior is illustrated in Fig. 1 by shear stress-shear strain curves of furnace-cooled single crystals of pure copper and Cu-14 at. pct A1. The yield point effect, At, was thoroughly studied previously1 and shown to be associated with short-range order. The present paper reports an investigation of the lower-yield stress in aCu-A1 single crystals as a function of composition, testing temperature, and strain rate. EXPERIMENTAL PROCEDURE Single cryski1 tensile specimens were prepared in the same manner as previously described.l Briefly,

Jan 1, 1962

By S. R. Maloof

In 1945, Hollomon' showed that after plastic yielding and prior to necking under simple tension, both ferrous and nonferrous materials are approximated by an equation of the following form: where, a = true stress K = strength coefficient 5 = true strain n = strain hardening exponent If a material is stress free, a plot of log o vs log 6 would be expected to be a straight line, the slope of which is the strain-hardening exponent (n) and the intercept on the ordinate axis for 6 = 1. is the logarithm of the strength coefficient K. It can be easily shown that the strain-hardening exponent in the above equation is equal to the total uniform elongation prior to necking. Thus, one can determine the total uniform elongation prior to necking from such a plot and, as pointed out by C. W. MacGregor, 2 it is the amount of strain prior to necking which determines the extent to which a material may be deep drawn. In the view of the current interest in the fabrication of beryllium shapes, it was thought that a knowledge of the strain-hardening exponent of cross-rolled beryllium sheet would be of general interest. Load-strain data after yielding and prior to necking was taken from the work of Greenspan,3 converted to true stress and true-plastic strain, and plotted on log-log paper as shown in Fig. 1. The value of Young's modulus E for beryllium was taken as 43 . l06 psi. The slope of this plot gave a value of 0.212 for the strain-hardening exponent (n) and the intercept on the ordinate axis a value of 105,000 for K. The total uniform elongation is thus 21.2 pct, in excellent agreement with the value estimated by Greenspan (about 20 pct). Furthermore,

Jan 1, 1960

By Masayuje Otagawa

The mining methods adopted in Japanese mines are less known to the mining world than those of other countries, owing to the geographical remoteness, but they present many features of interest to mining engineers. The Ashio copper mine is at the town of Ashio, in Tochigiken, about 109 mi. (175 km.) from Tokyo by the Imperial Government railroad. The area of the concession owned by the Furukawa Mining Co. is 5000 acres (5023 ha.). The mine was formerly divided into three, viz. Hon-ean, Kotaki, and Tsudo, which are distant 3 mi. (4.8 km.) from each other as shown in Fig. 1. However, in 1914, when the writer was general manager, the three mining plants were amalgamated into one with the result of simplifying the working projects and increasing the operating efficiency. The mine was visited by the Institute party in November, 1911, when K. Inouye was the general manager. Shortly afterward, the writer was appointed manager, and, in 1915, he was succeeded by Dr. K. Asano. At the end of 1917, K. Kibe became general manager and, after two years, was succeeded by the present manager, I. Sugimoto. For some time Dr. J. Kojima was acting general manager. It is necdless to say that the gentlemen mentioned are all members of the Institute, as is also Baron T. Furukawa, president of the company. Since the discovery of the enormous masses, called "kajika" or "bullhead" deposits, which the writer has described elsewhere,' the mining methods have been varied to meet the different conditions presented by these deposits. The Ashio Ore The country rock of this mine consists of the middle Chichibian slate and quartzite into which has been intruded a post-Tertiary plagio-liparite neck more than 2 mi. in diameter, in the shape of an inverted cone which narrows downward to a relatively small diameter in depth.

Jan 1, 1923

By Masayuje Otagawa

The mining methods adopted in Japanese mines are less known to the mining world than those of other countries, owing to the geographical remoteness, but they present many features of interest to mining engineers. The Ashio copper mine is at the town of Ashio, in Tochigiken, about 109 mi. (175 km.) from Tokyo by the Imperial Government railroad. The area of the concession owned by the Furukawa Mining Co. is 5000 acres (5023 ha.). The mine was formerly divided into three, viz. Hon-ean, Kotaki, and Tsudo, which are distant 3 mi. (4.8 km.) from each other as shown in Fig. 1. However, in 1914, when the writer was general manager, the three mining plants were amalgamated into one with the result of simplifying the working projects and increasing the operating efficiency. The mine was visited by the Institute party in November, 1911, when K. Inouye was the general manager. Shortly afterward, the writer was appointed manager, and, in 1915, he was succeeded by Dr. K. Asano. At the end of 1917, K. Kibe became general manager and, after two years, was succeeded by the present manager, I. Sugimoto. For some time Dr. J. Kojima was acting general manager. It is necdless to say that the gentlemen mentioned are all members of the Institute, as is also Baron T. Furukawa, president of the company. Since the discovery of the enormous masses, called "kajika" or "bullhead" deposits, which the writer has described elsewhere,' the mining methods have been varied to meet the different conditions presented by these deposits. The Ashio Ore The country rock of this mine consists of the middle Chichibian slate and quartzite into which has been intruded a post-Tertiary plagio-liparite neck more than 2 mi. in diameter, in the shape of an inverted cone which narrows downward to a relatively small diameter in depth.

Jan 1, 1923

Executive Committee. CHESTER F. LEE, Chairman. C. R. CLAGHORN, Vice-Chairman. WC. BUTLER. M. K. LODGERS. Joseph DANIELS, Secretary-Treasurer. Meeting, Jai. 25, 1913.-Dean Milnor Roberts read a paper on The Equipment and Curricula of the College of Mines, University of Washington.

Jan 4, 1913

United Engineering Trustees, Inc. H. G. MOULTON, 40 A. L. QUENEAU, '38 ALBERT ROBERTS, '39 The Engineering Foundation GEORGE D. BARRON, '40 A. L. QUENEAU, '38 Engineering Societies Library Board H. K. MASTERS, '38' A. B. PARSONS, Ex-officio F. F. SHARPLESS, '37

Jan 1, 1936

By A. S. Malicsi, I. Iwasaki

The removal of hydrophobic coatings of flotation collectors from iron ores becomes of interest when a duplex flotation process is considered for upgrading, when a pelletizing process is considered for a concentrate floated with a fatty acid or a soap collector, or when a disposal of froth products from cationic silica flotation is of environmental concern. Ozone can oxidize organic compounds rapidly, thereby removing the hydrophobic coatings of flotation collectors. Ozone is widely used for treating and purifying drinking water, waste water treatment, and for chemicals processing (Murphy and Orr, 1975; Rice et al., 1980). Its uses in metallurgical operations, however, are very sparse (Allegrini et al., 1970; Chernobrov and Rozinoyer, 1975; Ishii et al., 1970; Iwasaki and Malicsi, 1985; Matsubara et al., 1978). Yet, its high reactivity and the absence of potentially hazardous byproducts become of interest in destroying flotation reagents adsorbed on mineral surfaces or remaining in mill water for recycle or for discharge. Duplex Flotation A duplex flotation process, as applied to oxidized iron ores, would involve a fatty acid flotation of iron minerals followed by an amine flotation of the siliceous gangue from the rougher iron concentrate. Such a process has been used in the Florida phosphate fields. Fatty acid coatings cannot be removed as readily with a simple acid or alkali treatment from iron oxide surfaces as from Florida phosphates. A combination of reagents, such as lime and quebracho, lime and alkali phosphate, or sulfuric acid and oxalic acid, has therefore been proposed. In a previous article (Iwasaki et al., 1967) , the use of activated carbon was found to be effective in removing fatty acid coatings both in the duplex flotation and the pelletizing processes. The use of ozone offers another approach to the removal of fatty acid coatings from iron oxide surfaces. To investigate the possible application of the duplex flotation process, a specularite ore from Michigan analyzing 36.5% iron was used. A 600-g (1.3-1b) sample was ground in a laboratory rod mill together with 250 g/t (0.5 lb per st) of sodium silicate to -150 µm (-100 mesh). This was transferred to a Fagergren laboratory flotation cell, and deslimed four times at 20 µm (quartz equivalent). The deslimed pulp was transferred to a laboratory conditioner, diluted to 40% solids, and conditioned with 250 g/t (0.5 lb per st) of soda ash and 250 g/t (0.5 lb per st) of oleic acid. The conditioned pulp was then transferred back to the Fagergren cell, floated until barren of froth, and the rougher froth product was returned to the cell and cleaned. The results are presented in Table 1. The cleaner concentrate at this point analyzed 45.3% Fe. The cleaner concentrate coated with fatty acid was transferred to a 2-L (0.53-gal) beaker. While the pulp was agitated with a glass T-stirrer, ozone was bubbled into the agitated pulp for 15 minutes at a rate of 10 mg/min (0.00035 oz per min) ozone (250 g/t or 0.5 lb per st 03 feed). It was observed that the pulp ceased to froth after about 10 minutes. The amine flotation of siliceous gangue from the ozonated pulp was carried out first by conditioning with a dextrin, a commonly used starch depressant for iron oxides. This was followed by flotation with a stage addition of an ether amine at increments of 100 g/t (0.2 lb per st). Three stages were required to float the siliceous gangue to near completion. The three froth products were combined and cleaned twice. When the cationic flotation Rougher, Cleaner 1 and Cleaner 2 cell products were combined, an iron concentrate analyzing 64.5% iron was obtained at an overall iron recovery of 72.8%. Pelletizing Fatty acid flotation concentrates have been pelletized successfully in northern Michigan mills. But at other locations, fatty acid coatings on iron flotation concentrates proved so undesirable in agglomeration that other methods of concentration had to be sought. For example, a sinter mix containing iron ore concentrates upgraded by fatty acid flotation resulted in decreased productivity. This occurred because the micropellets of particles with the hydrophobic coating are less tolerant of moisture. Thus, the bed permeability is lost (Beebe, 1965). The agglomeration of concentrates obtained by the fatty acid flotation alone, and the hydrophobic coatings destroyed by ozonation or by the duplex flotation process, is not expected to cause any difficulty since the surfaces of the concentrates would be hydrophilic. Removal of the fatty acid coating with activated carbon, indicated by the loss of floatability, was shown to restore the decrepitation temperature of wet balls during drying cycle (Iwasaki et al., 1967). Disposal of Cationic Silica Flotation Froths Recent demands of iron blast furnaces place the silica content of the magnetic taconite pellets at about 5%. Conventional process for magnetic taconite involving fine grinding and magnetic separation often produces magnetic concentrates analyzing in excess of 5% silica. This is due to the presence of the middling grains of siliceous gangue and magnetite. Cationic silica flotation of magnetic taconite concentrates (DeVaney, 1949) may be used to reduce the silica content. But the amine coating on siliceous gangue becomes of environmental concern when the flotation tailings are discarded in tailing ponds.

Jan 1, 1986

By John Chipman, Carl F. Floe

The system molten copper-oxygen-sulphur is of interest from both the practical and theoretical standpoints; practically, because oxygen and sulphur play an important role in the commercial production of copper, and theoretically because it offers a chance to study the reaction between an oxide gas, SO2, and a molten metal. Reactions of this type, such as that between CO or CO2 and molten iron, are important in many metallurgical operations. Again, the determination of any gas-metal equilibrium is of theoretical interest and value insofar as it contributes to the general development of the laws for solution of gases in metals. This problem is being given an increasing amount of attention in modern metallurgy. This paper is the first of a series and presents the results of a determination of the solubility of sulphur dioxide in molten copper at temperatures from 1100" to 1500ºC., and pressures from about 20 mm. of mercury to somewhat above atmospheric. Review of Previous Work The solubility of gases in most liquids has been found to obey Henry's law, which states that for a fixed temperature the amount of gas dissolved is directly proportional to the pressure. However this law does not hold for liquid metals. Instead, it appears that for many gases, particularly the simple diatomic ones such as O2, N2 and H2, the amount dissolved in liquid metals is directly proportional to the square root of the pressure. This is known as Sieverts' law. It may be recognized as a special case of Henry's law, which applies when the solute gas is dissociated into atoms or when it reacts with the solvent metal to form solute molecules containing single atoms of the gaseous elements, as, for example: 2M+G2=2MG K=(MG)2 /M2 X PG2 [I] which, if the solubility is not too high and the concentration of metal therefore nearly constant, can be simplified to: MG = K1 vPG2 [2 Following the same reasoning, it would be expected that the solubility of SO2 in copper would be proportional to the cube root of the pressure, the reaction being as follows: 6Cu + SO2 = Cu2S + 2Cu2O [3] K= Cu2S X (Cu2O)2 i - (Cu)6(Pso2) [4] or, neglecting the change in metal concentration and considering that the concentrations of sulphur and oxygen are both proportional to the volume V of gas absorbed: K1 =V3SO2 /Pso2 or

Jan 1, 1941

By Carl F. Floe, John Chipman

The system molten copper-oxygen-sulphur is of interest from both the practical and theoretical standpoints; practically, because oxygen and sulphur play an important role in the commercial production of copper, and theoretically because it offers a chance to study the reaction between an oxide gas, SO2, and a molten metal. Reactions of this type, such as that between CO or CO2 and molten iron, are important in many metallurgical operations. Again, the determination of any gas-metal equilibrium is of theoretical interest and value insofar as it contributes to the general development of the laws for solution of gases in metals. This problem is being given an increasing amount of attention in modern metallurgy. This paper is the first of a series and presents the results of a determination of the solubility of sulphur dioxide in molten copper at temperatures from 1100" to 1500ºC., and pressures from about 20 mm. of mercury to somewhat above atmospheric. Review of Previous Work The solubility of gases in most liquids has been found to obey Henry's law, which states that for a fixed temperature the amount of gas dissolved is directly proportional to the pressure. However this law does not hold for liquid metals. Instead, it appears that for many gases, particularly the simple diatomic ones such as O2, N2 and H2, the amount dissolved in liquid metals is directly proportional to the square root of the pressure. This is known as Sieverts' law. It may be recognized as a special case of Henry's law, which applies when the solute gas is dissociated into atoms or when it reacts with the solvent metal to form solute molecules containing single atoms of the gaseous elements, as, for example: 2M+G2=2MG K=(MG)2 /M2 X PG2 [I] which, if the solubility is not too high and the concentration of metal therefore nearly constant, can be simplified to: MG = K1 vPG2 [2 Following the same reasoning, it would be expected that the solubility of SO2 in copper would be proportional to the cube root of the pressure, the reaction being as follows: 6Cu + SO2 = Cu2S + 2Cu2O [3] K= Cu2S X (Cu2O)2 i - (Cu)6(Pso2) [4] or, neglecting the change in metal concentration and considering that the concentrations of sulphur and oxygen are both proportional to the volume V of gas absorbed: K1 =V3SO2 /Pso2 or

Jan 1, 1941

By M. J. Pool, P. J. Spencer, J. R. Guadagno

THE present work was carried out as part of a general program to study the magnitude and concentration dependence of exothermic heats of solution of various metals in liquid tin. Thermodynamic data for the Te-Sn system are very limited, particularly for dilute alloys. Indeed, apart from the prior heat of solution determinations of Pool,1 the only other published values are those of McAteer and seltz2 who carried out electromotive force measurements on solid alloys of compositions from 75 to 95 at. pct Te at temperatures between 543° and 668° K, of Colin and Dro-wart3 who made mass-spectrometric studies of the vapor in equilibrium with solid SnTe, and of Robinson and Bever4 who determined the heat of formation of SnTe by solution calorimetry. The present work thus provides further thermodynamic information for this system, specifically in a temperature range where the solubility of tellurium is very limited. The liquid metal solution calorimeter used in this investigation has been described previously.' No further discussion of its design or operation will therefore be given here. For each series of measurements a tin bath consisting of approximately 70 g of 99.99 pct pure Sn was carefully weighed out and inserted in the calorimeter which was maintained at the desired temperature to within ± 1/4°c. Tellurium, of 99.999 pct purity, was supplied by Atomergic Chemetals Co., a division of Gallard-Schlesinger Chemical Manufacturing Corp. Approximately 0.0006 to 0.001 g-atom specimens of tellurium were used for each experimental run and these were added to the calorimeter from a temperature of 273" K. Heat of solution values were calculated by analysis of a strip-chart recorder plot of the heat effect accompanying the solution of each sample in the tin bath. The heat capacity of the latter was determined at regular intervals during a series of tellurium drops by using pure tin or tungsten calibration samples, the heat contents of which were obtained from the standard compilation of Hultgren et al5 The composition of the bath was calculated after each drop from a knowledge of the weights of tin and tellurium added. Values of the heat of solution of tellurium at 725" and 775° K are listed in Table I together with the corresponding tellurium contents of the alloys. The data are also plotted as a function of tellurium concentration in Fig. 1, which includes the equivalent data obtained by pool' at 750° K. It can be seen that the heat of solution remains

Jan 1, 1969

By Murray F. Hawkins

Horner1 and van Everdingen8 ave shown that the pressure drop within the wellbore, as a result of having produced the well at a constant rate q for time t, where t is sufficiently large, is: van Everdingen observed that better agreement between theory and well performance can be obtained if, instead of assuming the permeability is kc everywhere about the well, it is assumed the permeability near the well-bore is substantially reduced as a result of drilling, completion and/or production practices. In order to account for the additional pressure drop he introduced the dimensionless quantity S, the skin effect factor, so that Eq. 1 becomes: Eq. 2 might have also been obtained as follows. Assume a zone of altered permeability k, exists about the well out to a radius r and beyond that the unaltered, external permeability k. The additional pressure drop required to overcome this skin of reduced permeability may be calculated with sufficient accuracy using the incompressible flow equation; for Browns-combe and Collins1 have shown almost no difference between compressible and incompressible steady-state flow, in the vicinity of the wellbore, and the small volume of fluid in the vicinity of the wellbore makes unsteady-state mechanics unnecessary. Then. The sign of this skin pressure drop will be positive or negative depending upon whether the altered permeability k, is smaller or larger, respectively, than the external permeability k... Adding the pressure drop of Eq. 3 to Eq. 1 to find the total pressure drop: Comparing Eq. 4 with Eq. 3 it is seen that the skin effect may be defined by: be determined from pressure build-up tests1,3,4,5,6,7,8 The average permeability kavg including the altered and external permeabilities, can be determined from PI tests, and may be defined approximately on the basis of steady-state flow, as was done by Thomas' in defining the damage factor, by: The productivity ratio is the ratio of the average to the external permeability, kavg/ke or This equation shows that the productivity ratio and the skin effect are not uniquely related, because of the uncertainty in the drainage radius, and also, in many instances, the wellbore radius. Fortunately, they enter in the logarithm. The curves of Fig. I are plots of Eq. 8 for values from 100 to 50,000. The points represent the skin effects and corresponding productivity ratios from a large number of well tests, taken from Fig. 15 of Ref. 5 or Fig. 13 of Ref. 9. While some of the data

Jan 1, 1957