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Elected Died 1884 *ADDY, MATTHEW 1896 1894 *ALLFN R B 1896 1874 *AMIOT, H 1892 1886 *AMSLER, CARL 1894 1895 *ÅNGSTROM, CARL 1901 1895 *ARMSTRONG, JOHN F 1898 1882 *ARNOLDS, HUGO 1886 1875 *ASHBURNER, CHARLES A 1889 1872 *ASIIBURNER, WILLIAM 1887 1883 *ATKINS, C M , JR 1885 1897 *AYERS, WILLIAM 1898 1885 *BABCOCK, GEO H 1893 1884 *BAILEY, EDWARD 1889 1880 *BAILEY, JACKSON 1887 1900 *BALLOU, FRANKLIN 1903 1886 *BARNES, F W 1892

Jan 1, 1910

By Philip W. Henry

There is a difference of opinion among engineers on the subject of depreciation in general, and still more on its application to any given case The committee which was appointed by the American Society of Civil Engineers to make a report on Valuation of Public Utilities, states "There is no subject connected with valuation about which there are more diverse views than those relating to depreciation;" and in the discussion which took place upon the presentation of the preliminary report of this committee in January, 1914, that part of the report dealing with depreciation called forth more discussion than any other. From this discussion it was evident that the word "depreciation" was not always used in the same sense, so it is quite important that some definition be adopted. As this paper relates to the proper method of treating depreciation by corporations or individuals operating oil properties, both of which must make a return of annual net income to the government in connection with the income tax, it will be in order to adopt the definition of the Internal Revenue Bureau. Under date of Mar. 29, 1910, the Commissioner of Internal Revenue of the Treasury Department stated "Deduction on account of depreciation of property must be based on lifetime of property, its cost, value and use." Again, on Form 1035, "Return of Annual Net Income" for miscellaneous corporations (Section 2, Act of Congress approved Oct. 3, 1913), a deduction from gross income is allowed for "Total amount of depreciation for the year." This item is defined more particularly as follows: "The amount claimed under Item No. 5 (b) for depreciation should be such an amount as measures the loss which the corporation actually sustains during the year in the value of buildings, machinery, and such other property as is subject to depreciation on account of wear and tear, exhaustion, or obsolescence. The amount taken credit for on this account in order to be allowable should be so entered on the books as to constitute a liability against the assets of the corporation. The amount claimed under this item should not cover losses in the value of stocks and bonds. Decrease in the book value of securities owned, so far as such decrease represents a decline in the actual value of such securities, should be deducted under item 5 (a) of the return. "Where depreciation of physical property is made good by renewals, replacements,

Jan 1, 1915

By W. S. Reid

In March, 1938, E. P. Fleming, metallurgist for the American Smelting and Refining Co. inaugurated an investigation into the possibilities of recirculating the gases from Dwight-Lloyd sintering machines operating on lead charge, with the twofold object of concentration of the SO2 content and reduction in volume of total gas produced. The possibility of recovering a commercial grade of SO2 gas from D & L machines operating on lead charge had previously been considered by several investigators. Early History of Recirculation The Selby Smelter Commission Report, published by the Bureau of Mines in 1914, contains a chapter by A. E. Wells regarding results obtained at Selby, wherein some of the richer gas was recirculated through a hood over the pallets. Oldright and Miller of the U. S. Bureau of Mines had also made tests at Trail, B.C., and at Kellogg, Idaho. R. C. Rutherford, while at the Chihuahua, Mexico, Smelter of the American Smelting and Refining Co., in May, 1937, proposed recirculation of D & L gases to decrease the volume of gas handled by the baghouse. At none of these plants, however, was the operation commercialized. In July, 1938, Mr. Fleming, in correspondence with the Selby Plant, inquired regarding the possibility of obtaining 6 pct SO2 gas from the Selby D & L machines. At that time, the writer advised that there was slight possibility of obtaining 6 pct SO2 gas without re- circulation, but believed that it was possible with recirculation, and that experimental work toward that end should be tried at some plant where spare D & L machines were available. The foregoing statement was based on the following information then available— 1. Tests on Selby first-over machines showed 2.28 pct SO2 from first windbox and 1.03 pct SO2 from second windbox, and corresponding figures for second-over charge of 0.81 pct SO2 for first windbox and 0.08 pct SO2 for second windbox. 2. Oldright and Miller (US. Bureau of Mines) in 1932 at Bunker Hill, on 42 in. X 22 ft machines found: a. First-over charge—Maximum SO2 concentration (leaving cake) of 9.5 pct. b. First-over charge—SO2 concentration of over 8 pct from the 4 ft to the 12 ft points beyond the front dead-plate and that the concentration then dropped rapidly. c. That approximately 80 pct of the total sulphur eliminated on the second-over machines occurred during the travel of the pallets from the 1 ft to the 6 ft distances from the dead-plate. d. That approximately 94 pct of the total sulphur eliminated on the second-over machines occurred over the first windox. 3. Oldright and Miller (US. Bureau of Mines) in 1932, at Trail, on 42 in. X 50 ft machines found: a. First-over charge—SO2 varied from 2.0 pct to 5.5 pct (leaving cake) from the 12 ft to 28 ft points from the deadplate. b. That the average SO2 increased from 1.0 pct at 7 ft from dead-plate to maximum of 3.3 pct at 20 ft, then dropped to 1.5 pct at 40 ft. c. That on a 22 ft, second-over machine with an 11 in. bed the SO2 varied from 4.5 pct to 6.5 pct from the 2 ft to the 8 ft points from the dead-plate. d. That on the final roast, the SO2 concentration also varied across the pallets; that is, 2 1/2 pct at the side, increasing to 6.0 pct, 5 in. in, and to 7.0 pct at 10 in. to 20 in. in, then decreased vice versa at the opposite side. e. Concluded that most of the sulphur on a 22 ft machine was removed over the first 7 ft of the first windbox; therefore, they partitioned the first windbox so that the exit gas from the second 4 ft section was returned to the surface of the pallets over the fist 7 ft section, and during a seven-day trial the gas from the 4 ft section averaged 2.4 pct SO2, while the recirculated gas from the first 7 ft section only increased to 3.8 pct SO2. (Excess suction over the 7 ft section to prevent escape of SO2 laden gas from the 4 ft section caused dilution by air.)

Jan 1, 1950

By Pablo Hadzeriga

Great Salt Lake brine was subjected, in laboratory scale, to conditions simulating solar evaporation. Solid phases and the variation in composition of the liquid phase throughout the potassium salts crystallization zone are presented. These experimental results are discussed and compare with stable equilibrium data of the quinary system Cl-, SO4, Na+, K+, Mg + + and H2O. The Great Salt Lake brine in Utah is a vast reserve of several potential chemicals which has attracted much attention in the recent past. From the standpoint of an ore reserve, it may be said that the lake contains a low grade ore which has to be concentrated by evaporation of water. Because of the fairly favorable weather conditions in the Great Salt Lake area, it seems reasonable to consider solar evaporation as an economical means of concentration. From the view of phase chemistry, the main constituents of Great Salt Lake brine make it belong to the five component system: C1-, SO4=, Na+, K+, Mg++ and H2O. Because of their low concentration, the presence of other elements may be regarded as having little effect on the equilibrium of the five main components. Equilibrium data for the quinary system of our interest has been available for many years. Using these data, it is possible to calculate how the brine of the lake will concentrate when subjected to solar evaporation. In practicing solar evaporation, the weather factor and supersaturation conditions cannot be ignored, but their introduction into calculations using equilibrium diagrams is almost an impossible task. With the experimental work presented in this paper, an attempt is made to compare the stable equilibrium data of the quinary system as applied to the case of the Great Salt Lake brine and the dynamic equilibrium when that brine is subjected to continuous solar evaporation. Sea water brine belongs to this quinary system when only the major constituents are considered. The study of this system has been carried on since the last decades of the nineteenth century. The reason for these intensive studies was to give a geological explanation to the underground potash deposits. AS a result, many theories were presented trying to relate the equilibrium data with the geological formation, paragenesis and metamorphism of these deposits. Two main approaches were followed for the equilibrium studies of this quinary system. One of them, and the best known to us, was the school started in Germany in the last century by J. H. Van't Hoff. The other one, which started about the same time, was the school of N. S. Kurnakov in Russia. Both approaches were different and as a consequence generated opposite theories about the geological formation of potash deposits. The Van't Hoff school studied the quinary system of sea water under careful laboratory control, resulting in stable phase diagrams. During these studies, Van't Hoff recognized the difficulties in crystallizing some double salts.' ". . . . we unexpectedly found that some compounds whose formation in the solutions under investigation was possible at 25°C, nevertheless totally failed to put in an appearance. These, not to mention the calcium salts, were leonite, MgSO4 . K2SO4 4 H2O; kainite, KC1-MgSO4 . 3 H2O; and kieserite, MgSO4 .H2O. Even exceedingly slow crystallization with the addition of the compounds themselves did not remove the condition of supersaturation in the case of these bodies." However, those salts (leonite, kainite, and kieserite) were included in the stable phase diagram at 25°C as shown in Fig. 1 using D'Ans data.3 The figure also shows the sequence of crystallization of the Great Salt Lake brine according to the stable diagram. In later studies by the Van't Hoff school, especially by D'Ans,3 Karsten,4 Autenrieth,5,6 attention

Jan 1, 1968

By H. D. Kessler, R. J. Van Thyne

Phase diagrams of the titanium-rich portion of the ternary systems from 0 to 10 wt pct Al and 0 to 1 wt pct 0, N, and C were determined. Micrographic analysis of annealed high purity arc melted alloys was the principal method of investigation and was supplemented by X-ray diffraction. TITANIUM-ALUMINUM alloys exhibit excellent properties, particularly for elevated temperature use. For this reason, the Materials Laboratory, Wright Air Development Center, sponsored an investigation of the effect of the interstitially soluble contaminants, oxygen! nitrogen, and carbon, on the Ti-A1 system. Using high purity arc melted alloys and micrographic analysis of annealed samples as the chief method of investigation, titanium-rich partial phase diagrams were determined for the ternary systems. Experimental Procedure Materials: The titanium used in the preparation of the alloys was iodide crystal bar (99.9+ pct pure) produced by the New Jersey Zinc Co. and Foote Mineral Co. The aluminum was obtained from the Aluminum Co. of America in the form of sheet. The given analysis was: Si, 0.0006 pct; Fe, 0.0005; Cu, 0.0022; Mg, 0.0003; Ca, <0.0006; Na, <0.0005; and Al, 99.99. High purity titanium dioxide purchased from the National Lead Co. was used in the preparation of the oxygen-bearing alloys. The spectrographic analysis of this material was as follows: SiO,, 0.07 pct; Fe,O,,, 0.002; A1,0,, <0.001; Sb,O,, <0.002; SnO,,, <0.001; Mg, <::0.001; Cb, <0.01; Cu, 0.0004; Pb, 0.002; Mn, <0.00005; W, <0.01; V, <0.002; Cr, <0.002; Ni, <0.001; and Mo, <0.002. Special spectroscopic graphite rod purchased from the National Carbon Co. was used in the preparation of the Ti-A1-C alloys. Alloy Preparation: Alloy ingots weighing 10 grams were melted in a nonconsumable tungsten electrode arc melting furnace using water-cooled copper hearths and a helium atmosphere. The arc was struck on a tungsten stud in the copper block to minimize contamination from the electrode. Details of the techniques have been reported.&apos;, &apos; Each ingot was inverted and remelted a minimum of four times to insure homogeneity, without opening the furnace. Remelting small control ingots of iodide titanium under similar conditions resulted in no measurable hardness increase, indicating contamination-free melting conditions were used. The alloy charges and resultant ingots were weighed to the nearest milligram. Only small weight losses were obtained upon melting, indicating that the actual compositions were very close to the nominal compositions. Check analyses substantiated this. Oxygen and nitrogen were added during ingot preparation as master alloys. An oxygen master alloy containing 25 wt pct 0 was prepared by melting iodide titanium and pressed titanium dioxide (40 pct 0) powder. The Ti-25 pct 0 alloy is more easily handled during weighing and charging than the pressed TiO, powder. Chemical analysis indicated that titanium-nitride received from an outside source contained large amounts of impurities (2.7 pct Ca). Therefore, a master alloy containing approximately 12 pct N was prepared by melting nitrided sponge titanium. As the sponge melted with difficulty, the alloy was diluted by addition of titanium to lower the melting point. The chemical analysis of the final alloy was 6.69 pct N. Both master alloys are friable and were used as —8 +16 mesh lumps resulting from crushing the ingots. NIelting the Ti-A1-0 and Ti-A1-N compositions five times produced homogeneous ingots. The preparation of homogeneous carbon-bearing alloys caused considerable difficulty. Alloys containing less than 1 pct C were prepared using spectrographic graphite rod, 1/8 in. diameter by 1/8 in. long. However, the graphite did not dissolve even with long melting times and many remelts. The ?/s in. pieces of graphite were used rather than powder because the low density powder is troublesome in arc melting.

Jan 1, 1955

By B. J. Kochanowsky

The idea of a portable crusher is not new. Many such crushers are available but they are small and designed for construction work. For many years the author has suggested, both in this country and in Europe, the building of larger portable crushers intended expressly for use in quarries or open pits. Although not applicable under all conditions, there are mining operations where a mobile crusher arrangement could be more profitable than the facilities now used. The primary use of a portable crusher, i.e., a crusher mounted on crawlers or tires, in the rock and mining industries is to reduce costs by permitting the substitution of conveyor belt haulage for truck or track haulage. The usual sequence of operations in surface mining is drilling, blasting, loading, haulage, and crushing. Haulage is normally accomplished by truck or track-mounted cars, the latter method being used for the longer distances. However, by using a portable crusher in the pit, the sequence of operations would be changed so that the crushing stage would occur before haulage (Fig. 1). Such a sequence would permit the use of conveyors to replace the more expensive truck or track haulage methods. Since most quarry and open pit operations normally require a crushing stage, the only additional costs incurred will be due to the investment required to purchase or construct a mobile arrangement for a crusher. But this factor has to be weighed against the advantages to be gained by conveyor haulage. As shown in Fig. 2, transportation of material by belt conveyor over short distances is less expensive than by truck. The inclination of the belt has no effect on belt speed; consequently, the hourly tonnage moved remains the same. Conversely, the output rate of trucks as expressed in tons or ton-miles per shift decreases proportionally to the haulage speed, which is considerably slowed by the steepness of the road (Fig. 3, left). Although maximum possible grades and maximum economic grades of haulage are greater for a belt than for a truck (over the same total lift), the longer haulage distances favor the use of trucks. Although power consumption for hauling on a grade increases for both conveyances, the rate of power consumption increases faster for trucks than for conveyor belts (Fig. 3, right). Since the output rate and related fixed costs are affected by the travel speed, total haulage costs with trucks would increase with the grade more rapidly than the similar costs of conveyor belts (Fig. 4). Travel distance, road grade, speed, size and number of pieces of equipment, efficiency of operation, and many other factors affect such haulage costs. In general terms it can be said that the shorter the distance, the steeper the grade, and the greater the output, the more advantageous the belt becomes in comparison to truck or track haulage. In addition to potential cost savings in haulage procedures, a portable crusher would allow better utilization and performance of shovels. Loading operations would not be interrupted as often by the necessity of waiting for cars or trucks. Unfortunately, the application of belts in open pits for haulage from bench sites is generally not practical under existing conditions because a belt fed directly by a mechanical shovel can be torn, damaged, or worn out quickly by the large rock fragments falling on it during loading. However, by using a mobile crusher this situation can be avoided. As shown in Fig. 1 (b), the shovel feeds rock into the crusher located behind it. The crushed material is initially transported by an extensible and/or movable belt, thence by a longer stationary conveyor to the plant where the material is subjected to further treatment by secondary crushing, screening, etc. The first-mentioned conveyor, needed to bridge the distance between the shovel and the stationary conveyor, is necessarily variable in length owing to the continuous movement of the shovel and the desire to keep the stationary belt at a safe distance from the bench during blasting operations. The remarkable part of mobile crusher operations is the extra-ordinarily high output per man-shift, the low maintenance and power requirements for haulage, and the increased output of the loading shovel. A cement quarry which has been using a portable crusher and conveyor since 1956 requires only three men to operate the shovel and crusher and to transport the crushed rock by belt from the quarry face to the screening plant. If truck haulage

Jan 1, 1961

By George Smith

IN discussing present-day business and industrial troubles we easily drop into the habit of clinical diagnosis. Talk of this kind, with its emphasis on suspicious symptoms and abnormal tendencies, makes us all morbid. As an English essayist has put it, we may thus induce a melancholy, temper that simply in-vites the ministrations of social and political quacks. It is far better-because nearer the truth-for us to recognize that our condition is in no sense diseased, that our body politic, social and economic, is suffering merely from growing pains; that America is not de-cadent-indeed, that any lapses from staid normality are traceable to the, excesses of youth. Our chief .trouble, then, is only that our industrial state is still young; it may be that we have been growing too fast. The growth in the industrial development of our country from pioneer days to the present time has been marvelous. Too many of us, however, take all this as a matter of course-we credit this record of American progress to our American genius alone. Too often we overlook the fundamental relationship between resources and progress-that man is aided and even guided by Nature. The fact is, social advance, indus-trial progress, national prosperity all have their ma-terial basis in natural resources. In furnishing the needed mineral raw materials to the great manufac-turing industries of the United States there has been no halt, but rather a general acceleration. No other phase of American activity can show-equal advance. As the decades have passed, the winning of mineral wealth has progressed from the role of a stimulus for western exploration to that of the foundation for nation-wide business. In 1890 the Sherman law was passed; since then the population of the United States has about doubled. Since then, too, thevalue of the annual product of the mines of this country has increased better than eight-fold-from 650 millions to 51/ billions of dollars. Since then the four largest branches of the mineral industry, which supply the iron and copper, the bituminous coal and petroleum-the bone and sinew and the energy of all other American industries-have increased their annual output nearly ten-fold as to value and six-fold as to quantity. Such a growth in the volume of mineral output must be taken as a true measure of the ex-ceptional service rendered by the mines of the United States. Not only is this a much larger country than it was 38 years ago, but its growing mineral industry has met demands that have been out of all proportion to the growth of the population.

Jan 8, 1928

By Benjamin F. Tillson

THE past seven years, and 1937 in particular, have witnessed the return of New Jersey iron mining to a place of importance. Following the World War period, little mining was done for several reasons. The properties were operated by small merchant furnaces whose demand was not sufficiently large to warrant large-scale methods with their attendant low operating costs. Under those conditions, the ore produced could not compete with that from the Great Lakes. However, when on Nov. 1, 1.929, the Alan Wood Steel Co., of Conshohocken, Pa., leased the Scrub Oak and Washington properties at Mine Hill and Oxford respectively, they made radical changes in the mining and milling processes which greatly reduced the cost of production and pointed the way to a renewed interest in mining in New Jersey. The following factors have been advanced as influencing the reopening of New Jersey mines:

Jan 1, 1938

By William A. Mitchell

AN X-ray diffraction pattern for "calcined cold precipitated ferric oxide" is reproduced diagrammatically along with data for other iron oxides by R. C. Mackenzie.1 This pattern, which shows spacings higher than those for any other form of iron oxide, was obtained from material prepared and photographed at several different times over a period of weeks, but later attempts to reproduce the results were unsuccessful the pattern of hematite being obtained every time.2

Jan 9, 1953

By B. H. Dunshee

This paper is not intended to be a technical discussion of square-set framing as used in mines, but merely a short description of the different kinds of framing that have been used in the Butte mines, and what has been decided upon, after years of practical experience, as being the best, taking into consideration the physical condition of the veins and country rock, and also the supply of timber available for mining purposes. Butte is, essentially, a square-set district. Virginia City, Nev., was the first mining camp to use this method of timbering, and to Philip Deidesheimer, who went to the Ophir mine in 1860, is given the credit of solving the problem of square-set framing, where large bodies of ore were stoped and the timbers were supposed to hold the ground from caving without any filling. That he did his work well is shown by the fact that the main features of his framing are in use to-day. In the Butte mines the conditions are such that this method of timbering is generally used. We do not, of course, depend upon the timbers any more than is necessary. The worked-out stopes are filled with waste rock as close behind the miners as possible and not interfere with mining operations. The waste rock is obtained from development work on the different levels or from the surface. In the earlier days of mining in Butte there were numerous independent companies and naturally there were different ideas as to the best way to frame the square sets, although the general features were the same. Timber was abundant for mining purposes and the common practice was to use sawed lumber, generally 10 by 10 in. or 12 by 12 in. square. Some of the mines used for girts 6 by 10 in. or 8 by 12 in. timber, depending upon whether they used 10-in. square or 12-in. square timber for posts and caps. This avoided the necessity of framiug the ends of the girts, and also required less timber for the

Jan 1, 1914

PRESIDENTS DAVID THOMAS 1871 R W RAYMOND 1872-1874 A L HOLLEY1875 ABRAM S HEWITT 1876 T STERRY HUNT 1877 ECKLEY B COXE 1878 - 1879 WILLIAM P SHINN 1880 WILLIAM METCALF 1881 RICHARD P ROTHWELL 1882 ROBERY W HUNT 1883 JAMES C, BAYLES 1884-1885 ROBERT H RICHARDS 1886 THOMAS EGLESTON 1887 WILLIAM B POTTER 1888 RICHARD PEARCE 1889 ADAM S HEWITT 1890 JOHN BIRKINBINE 1891-1892 H M HOWE 1893 JOHN FRITZ 1894

Jan 1, 1923

By C. R. Johnson, R. A. Greenkorn, R. E. Haring

Three confined five-spot miscible displacements at unity, favorable, and unfavorable mobility ratios were conducted in a shallow, water-saturated sandstone of Pennsylvan-ian age near Chandler, Okla. These studies, plus associated laboratory experiments, were designed to measure miscible displacement performance in a controlled natural system, using known scaling criteria to develop an approach to modeling the heterogeneous field system. We have concluded from these studies that: (I) displacement efficiency in the field is a pronounced function of mobility ratio, indicating that miscible fingering observed in simple laboratory models occurs in the field: (2) field displacements can be quantitatively predicted by scaled laboratory models if the degree and location of field permeability variations are preserved in the models: and (3) arbitrary simplifications of heterogeneity will not necessarily predict observed displacement efficiency, and the simpler the model, the more optimistic the prediction. INTRODUCTION Many processes to achieve miscible displacement of reservoir oil by injected fluids have been conceived and field tested by the oil industry. Among the better known are high-pressure gas, enriched gas, and LPG banks. The simplest form of miscible displacement—one fluid miscibly displacing another fluid of different viscosity but the same density—has been studied extensively in homogeneous laboratory models. Observations of unstable fingering have been made which explain the significant decrease in displacement efficiency as the mobility ratio (ratio of displacing fluid mobility to displaced fluid mobility) increases. General industry experience with field tests of miscible displacement projects, mostly LPG banks, has been premature solvent breakthrough and lower than predicted production rate increases. These results have been attributed to either unstable fingering, unusual or unexpected permeability stratification, or both. Miscible displacement data in a controlled natural system have never been reported, however. Also, it has not been shown that properly designed and constructed laboratory models quantitatively predict field-scale behavior. The purpose of the combined field and laboratory ex- periments reported in this paper is twofold. The first was to measure miscible displacement performance at different mobility ratios in natural rock approaching field size under precise, controlled conditions. The second purpose was to utilize known scaling criteria plus several approaches to heterogeneity to model the field. Comparison of model and actual field results should then determine whether or not the laboratory phenomena (manifested by miscible displacement efficiency) are exhibited in large, natural rock systems. We carried out our program by first locating a shallow, water-saturated reservoir whose rock properties were representative of oil-bearing reservoirs. Detailed reservoir description by core analysis and interference testing showed the field site to be heterogeneous. A sequence of controlled, aqueous-phase miscible displacements was conducted at unity, favorable and unfavorable mobility ratios. A central, confined pattern was used to obtain the displacement data. A laboratory program using sand-packed models was conducted to determine the modeling criteria necessary to simulate field behavior of miscible displacement in a heterogeneous system. SCALING THEORY The detailed derivations and descriptions of the scaling laws that apply to laboratory models of reservoirs are adequately described elsewhere, so the following discussion will be restricted to facets of importance in this study. For a displacement in which one liquid miscibly displaces another, the following dimensionless groups are required to have the same numerical value in the model as in the field: The model also must be geometrically similar to the field, be spatially oriented the same as the field (same dip angle), and have the same initial and boundary conditions as the field (same initial fluid saturations and same injection-production well arrangement). When these conditions are satisfied, the theory predicts that at dny dimensionless time (pore volumes of produced fluids) the dimen-sionless flow potential and dimensionless fluid concentrations will be identical at all dimensionless spatial locations within the model and the field. If this prediction is correct, then the local dimensionless velocities must be identical, thus the instantaneous fraction of displaced fluid produced and the cumulative recovery expressed as fraction of original fluids in place must be identical at all dimensionless times. The theory outlined above has been obtained by either dimensional analysis or inspectional analysis of differential

Jan 1, 1966

By John Tyssowski

IN 1925, Harry Howard Stout, then metallurgist for Phelps Dodge Corporation, while investigating the cleaning of cathode copper by various gases at elevated temper-atures below the melting point of the metal, observed in the laboratory that particles of pure copper heated in a closed tube filled with hydrogen seemingly "stuck together." By further investigation he found that there was actual crystal growth across the boundaries of the particles of pure metal; in other words, there was a complete new crystal realignment and growth throughout the newly formed copper mass. It was coalescence, not adhesion; the several parti-cles had become one unit. From this dis-covery has been developed the coalescence process, whereby copper of cathode purity is squirted out in semifabricated form from an extrusion press, by the appli-cation of about 30c0 tons pressure on a 10-in. piston. Stout decided that cathode copper pro-duced by electrolytic deposition was the logical source of the pure copper required for the coalescence process, but as there was no way of reducing the ordinary tough commercial cathodes to the small particles required he was confronted by a serious problem at the outset. Under his direction, this problem was solved by William H, Osborn and Harry Howard Stout, Jr. They developed the process and technique for producing cathode copper in brittle, friable form, and to them must be given the greater part of the credit for this phase of the development program as well as for that relating to the gas cleaning of the cathode particles.

Jan 1, 1940

By J. W. JESSUH

TO certain people the name of Northern Rhodesia brings only a vague recollection of a distant country somewhere in Africa; to others, it means a big game territory and the opportunity for excellent shooting; to some, a land of sunshine and peace; but to mining men the name signifies the land where the latest discoveries of copper ore are astonishing the mining world and bid fair, in a few years, to make Northern Rhodesia one of the largest copper- producing countries. There is a fascination about this part of the globe that is somewhat similar to southern California, and at the same time there is that feeling of mystery peculiar to a tropical climate. Perhaps it is the isolation, the fear of wild animals, or the influence of an undefinable something which cannot be expressed.

Jan 1, 1931

The Pinto Valley, Arizona, plant of Cities Service Company is one of the newest copper producers in the United States, having started production in 1975. It is located a few miles west of Miami, Arizona, on the site of the Castle Dome Mine, a World War II operation which was shut down and its mill later moved and renamed the Copper Cities plant. The Pinto Valley concentrator uses six 18 ft (5. 5 m) diameter by 21 ft (6.4 m) long, 4000 hp, single-stage ball mills, taking crushed ore from four 7 ft (2. 13 m) standard and eight 7 ft shorthead crushers. It has a design capacity of 40,000 stpd but has higher tonnage capability for an expanded concentrator. The copper porphyry-ore is mined by two 15 yd and one 20 yd shovels (11.5 and 15.3 cu m) and is transported to the primary crusher by 150-ton trucks. The dump pocket is relatively shallow (14 ft or 4. 27 m deep); and because of the width necessary for the 150-ton trucks, it was designed to use the ore itself instead of manganese steel as an abrasion- resistant lining, although a heavy steel cap protects the dump pocket top edge. The pocket has a capacity of 300 tons of ore. Water sprays under the dumping positions direct a fog toward the truck being dumped for excellent dust control. Breaking of oversize is done by a hydraulic impact breaker mounted on a hydraulically operated articulated arm situated at the side of the dump pocket. Again because of the shallow pocket, no splitter is used to protect the spider of the 60 in. by 89 in. Traylor gyratory crusher, which is equipped with a hydroset mechanism. Crushed ore at 8 in. (203 mm) drops into the 360 st surge chamber lined with AR plate and then is fed directly to the 60 in. (1.52 m) stock- pile conveyor by an 84 in. by 20 ft (2. 13 m by 6.1 m) four-speed apron feeder at a rate of up to 4000 stph. Early in 1977, Cities Service will replace the four -speed electric drive with a hydraulic pump and motor with infinitely variable speed. They are also studying the replacement of the apron feeder with an 8 ft wide (2.44 m) belt feeder similar to those at Similkameen and Palabora. Peak crushing capacity would then be increased to 5000 stph. The AR plate in the crushed ore pocket was installed in large, rectangular sheets which required much less embedded metal to hold it in place than would the more conventional, smaller arid square pieces of cast manganese steel. For duty in the crushed ore pocket, where wear is not severe, the AR plate is a satisfactory material. The new Palabora Figure-8 crusher also uses this design.

Jan 1, 1978

By Felix B. Shay

The Agua Fria mine is in eastern Honduras at Lat. 140" 06&apos; N., Long. 86" 36&apos; W. Danli, the seat of the local municipal government, is the nearest town. The property is approximately 44 miles east of Tegucigalpa, the terminus of a truck road from the Pacific coast. The mine lies at an altitude of 3300 ft. with surrounding elevations as high as 6000 feet. For 20 years the oxidized croppings of the Agua Fria vein had been worked intermittently when in 1927 an attempt was made to open the property. A shaft was sunk 300 ft. and considerable ore was blocked out, but before a mill could be installed financial difficulties arose and the mine reverted to the original owners. Erle Y. Halliburton, of Los Angeles, purchased the property in 1934. At that time a careful consideration of all factors involved showed that air transport was the only practical method of moving the equipment and machinery necessary to commence production. The other alternative was by muleback, which required sectionalizing all equipment to a maximum load of 200 lb., and during the rainy season this type of freighting was subject to serious delays. It was decided to purchase a trimotor Ford powered by three Pratt and Whitney Wasp B motors developing 400 hp. each. There were several reasons for the choice: (1) this type of plane was available at a reasonable price, owing to its replacement by faster models on most of the airlines in the United States; (2) the ship was capable of carrying a pay load of 3000 Ib. over the required distance; (3) since this airplane had been widely used there would be no difficulty in securing competent personnel or in purchasing spare parts. It was necessary to rebuild the plane for cargo servicc. The windows were removed and the fuselage was entirely covered. A hatch was cut in the top of the fuselage to receive heavy crane loads and pieces of freight that were too large or too long to go through the rear door. The seats were removed and the interior of the plane was lined with Duralumin sheets. A heavy-duty landing gear was installed and the floor was reinforced. The nearest site for an airport was two miles north of the mine. The necessary grading was done by a small track-type tractor with a wheeled grader, which was brought in overland. Six weeks was required for bring-

Jan 1, 1940

By Felix B. Shay

The Agua Fria mine is in eastern Honduras at Lat. 140" 06&apos; N., Long. 86" 36&apos; W. Danli, the seat of the local municipal government, is the nearest town. The property is approximately 44 miles east of Tegucigalpa, the terminus of a truck road from the Pacific coast. The mine lies at an altitude of 3300 ft. with surrounding elevations as high as 6000 feet. For 20 years the oxidized croppings of the Agua Fria vein had been worked intermittently when in 1927 an attempt was made to open the property. A shaft was sunk 300 ft. and considerable ore was blocked out, but before a mill could be installed financial difficulties arose and the mine reverted to the original owners. Erle Y. Halliburton, of Los Angeles, purchased the property in 1934. At that time a careful consideration of all factors involved showed that air transport was the only practical method of moving the equipment and machinery necessary to commence production. The other alternative was by muleback, which required sectionalizing all equipment to a maximum load of 200 lb., and during the rainy season this type of freighting was subject to serious delays. It was decided to purchase a trimotor Ford powered by three Pratt and Whitney Wasp B motors developing 400 hp. each. There were several reasons for the choice: (1) this type of plane was available at a reasonable price, owing to its replacement by faster models on most of the airlines in the United States; (2) the ship was capable of carrying a pay load of 3000 Ib. over the required distance; (3) since this airplane had been widely used there would be no difficulty in securing competent personnel or in purchasing spare parts. It was necessary to rebuild the plane for cargo servicc. The windows were removed and the fuselage was entirely covered. A hatch was cut in the top of the fuselage to receive heavy crane loads and pieces of freight that were too large or too long to go through the rear door. The seats were removed and the interior of the plane was lined with Duralumin sheets. A heavy-duty landing gear was installed and the floor was reinforced. The nearest site for an airport was two miles north of the mine. The necessary grading was done by a small track-type tractor with a wheeled grader, which was brought in overland. Six weeks was required for bring-

Jan 1, 1940

By R. B. Snow

THIS paper describes a technique of polishing and etching specimens of open-hearth furnace slags or hearth aggregates for identification of the crystalline constituents-lime (CaO), tricalcium silicate (3CaO•SiO2), dicalcium silicate (2CaO-SiO2), monticellite (CaO•MgO-SiO2), or forsterite (2MgO•SiO2), with especial emphasis on the mineral merwinite (3CaO-MgO.2SiO2). With proper standardization, this identification does not require the use of the petrographic microscope. The composition of basic open-hearth slags and furnace bottoms falls, almost without exception, within systems containing CaO, MgO, "FeO," MnO and SiO2, in which the number of basic molecules so greatly exceeds the orthosilicate ratio (two molecules of base to one of silica) that free basic oxides, and combinations between them such as aluminates or ferrites, are present in cooled specimens. Orthosilicates of (CaO + MgO) are the most common in such specimens, since in nearly all cases, except premelt slags, the molecular ratio of (CaO + MgO) to SiO2 is more than 2 to r. When sufficient lime is available it combines with the silica to form dicalcium silicate (2CaO•SiO2), which contains little, if any, MgO, FeO or MnO in solid solution whereas the latter oxides combine to form the oxide solid solution known as periclase. If the lime present is insufficient to form dicalcium silicate (2CaO•SiO2) it combines with MgO to form either merwinite (3CaO•MgO.2SiO2) or monticellite (CaO-MgO•SiO2); these minerals take little if any FeO or MnO into solid solution and the remaining MgO, FeO and MnO combine as periclase. This generalization seems to be valid for basic slags and furnace bottoms, since minerals such as CaO-MnO•SiO2 and CaO-FeO-SiO2 are found only in slags in which the lime-silica ratio is less than 2 and are not observed in specimens from furnace bottoms. The identification of crystalline constituents in such materials, especially of fine crystals in the groundmass, is difficult under the petrographic microscope. They are often masked by their neighbors because of their small size in relation to the thickness of the thin section and because of the presence of opaque or colored constituents. The indices of retraction and the optical sign of the mineral are sometimes difficult to determine because of the small size or because of twinning or of inclusions within the crystal. Moreover, the positive identification of merwinite (3CaO•MgO.2SiO2) from its optical properties is usually difficult in the presence of dicalcium silicate (2CaO•SiO2). CaO, MgO, 3CaO•SiO2 and 2CaO•SiO2 in open-hearth slags have been identified for a number of years in the U.S. Steel Corporation Laboratory by the usual

Jan 1, 1947

By Arnt Solbu, Jomar Thonstad

The reaction between CO, and aluminum in cryolite-alumina melts in contact with aluminum has been studied by passing CO2 over the melt. In unstirred melts a homogeneous reaction between dissolved metal and dissolved CO2 was observed. In stirred melts in which convection was induced by bubbling argon through the melt, the dissolved metal apparently reacted mainly with gaseous CO2. The rate of formation of CO increased slightly with increasing depth of the melt, and it did not depend on whether CO2 was passed over or bubbled through the melt. The rate of formation of CO increased with increasing area of the metal/melt interface and with the application of anodic current to the metal. It is concluded that the dissolution of metal into the melt is the rate-determining reaction. THE current efficiency in aluminum electrolysis is determined by the rate of the recombination reaction between the anode gas and the metal: 2A1 + 3CO2—A12O3 + 3CO [1] as originally stated by Pearson and waddington.1 The occurrence of this reaction in cryolite-alumina melts in contact with aluminum was first verified experimentally by Schadinger.2 Thonstad3 has shown that the reaction may proceed further to give free carbon: 2A1 + 3CO— A12O3 + 3C [2] Normally only a few percent of the CO formed undergoes such reduction. The mechanism of these reactions has not yet been clarified. Aluminum, as well as CO,, is soluble in the melt. The solubility of aluminum in cryolite-alumina melts at around 1000°C corresponds to 75 x 10- 6 mole A1 per cu cm,4 while that of CO2 is only 3 x 10-6 mole CO, per cu cm.5 Taking into account the stoichiometry of Reaction [I], the ratio between dissolved aluminum and dissolved CO2 available for the reaction in a saturated melt is about 40. Therefore, as will be shown in the following, the reaction probably mainly occurs between gaseous COa and dissolved aluminum. The dissolved aluminum presumably consists of subvalent ions of aluminum and sodium.4&apos;6 Since the interpretation of the present results is not dependent upon the nature of this solution, the dissolved metal will be designated solely as Al+ in the following. The reaction can then be divided into four steps: A) dissolution of metal, e.g., 2A1 + Al3 — 3A1+ [3] B) diffusion of dissolved metal through a boundary layer; C) transport of dissolved metal through the bulk of the melt; D) Reaction [1]. If dissolved CO, takes part in the reaction, three additional steps embodying the dissolution and transport of CO2 must be added. schadinger2 observed, when bubbling CO2 through the melt, that the rate of formation of CO (in the following designated rfco) did not depend on the distance from the metal surface. The results also indicate that the rate of bubbling did not affect the rfco. When passing CO, over the melt, Revazyan7 found that the loss of metal did not depend on the depth of the melt above the metal or on the flow rate of CO2, and concluded that Step A is rate-determining. In an unstirred melt, however, Gjerstad and welch8 found that the rfCo decreased with increasing depth of the melt, indicating that step C was rate-determining. It thus appears that the rate control of the process depends on the experimental conditions, particularly on the convection. In the present measurements the reaction has been studied in unstirred as well as in stirred melts. EXPERIMENTAL AND RESULTS The experiments were carried out at 1000°C in a Kanthal furnace with a 10-cm uniform temperature zone (±0.l°C). The melts were made up of "super purity" aluminum (99.998 pct), hand-picked natural cryolite, and reagent-grade alumina. In experiments where alumina crucibles were used, the alumina content in the melt was close to saturation (13.5 wt pct9); otherwise it was 4 wt pct. Pure Co2 (99.85 pct) was passed over the melt, and the exit gas was analyzed for CO2 and CO by the conventional absorption method.3 From the weighed amount of CO (as CO2) the rfco was calculated as the number of moles of CO formed per min per sq cm of the surface area of the melt. The amount of carbon formed by Reaction [2] was not determined. As already indicated the rfco is much higher than the rfC, by Reaction [2]. Since the rfC probably is proportional to the rfco, the measured rfco should then the proportional to, but slightly lower than, the total rate of Reactions [I] and 121. In general the scatter of results obtained in duplicate measurements was ±5 to 10 pct, while within a given run a precision of ±3 to 5 pct was obtained. The various crucible assemblies that were used will be described below. Measurements in Unstirred Melts. When carrying out aluminum electrolysis in small alumina crucibles. Tuset10 observed that after solidification the lower part of the electrolyte was gray and contained free metal, while the upper part near the anode was white and contained no metal. One may test for the presence of free metal by treating with dilute hydrochlorid acid.

Jan 1, 1969

By L. W. Emery

Undergrorlnd combustion operations were initiated in a 60-acre Bartlesville sand "shoe-string" reservoir in Allen Connty, Kans., in 1956. Tests in separate patterns were conducted using various co~nbinations of air and recycle gas to propagate combustion fronts from the injection toward the producing wells. These patterns were made up of 6 injection and 20 prodrrcing wells Gas and liquid prorluctiorz from each pattern was measured on an individual-well basis, and comparisons were made between the three patterns to ascertain the relative effects of injected gas composition on production behavior. Breakthrough of the combustion front at a well was characterized by an increase in water production from the well followed by an increase in bottomhole temperatrrre to approximately 250" F. After burning fronts had broken through at five producing wells, operations were terminated in 1960. From the total project approximately 79,000 bbl of oil were produced during thermal operations at a cumulative produced GOR of 23 Mcf/bbl. No appreciable change in the character of the produced crude was observed. Combustion in the reservoir was maintained with injected gas compositions ranging fronz 6 per cent oxygen in recycle gas to 100 per cent air. lnjectiotz of large quantities of recycle gas resulted in higher producing GOR&apos;s from offset wells than were measured from a pattern into ~vhich straight air ~vas injected. The air required to move the combustion front through I acre-ft of reservoir was computed to be 20 MMscf. This valrre was found to be relatively independent of the quantities of recycle gas injected. The recovery efficiency from the swept area was esti~izated to be about 59 per cent. Areas swept were similar in shape to tlzose obtained with a laboratory potentiometric model. Samples of sund taken from behind the burning front by coring indicated almost total oil removal from the sand. Petrographic analysis of the core samples indicated that the sand had been heated to peuk temperature of rlbout 1,200" F. No rignificant difference in peak temperature was forrnd in two areas where compositions of injected gas were quite different. Compression costs for thermal recovery were estimated to be $1,20/bhl of produced oil. INTRODUCTION The use of the "forward combustion" process as an oil recovery method has received a great deal of attention. This method involves ignition of the formation in an injection well, followed by propagation of a combustion front through the reservoir. Combustion is maintained by the injection of an oxygen-containing gas to react with reservoir hydrocarbons. As the flame front progresses through the reservoir, oil and formation water are vaporized, driven forward in the gaseous phase and recondensed in the cooler part of the formation. In turn, the condensed fluids push oil into the producing wellbores. Completed field tests of the process were first reported by Kuhn and Koch,&apos; and by Grant and Szasz.&apos; Results from other tests have since been reported by Walter,3 by Moss, White and McNeil,&apos; and by Gates and Ramey." ach of these tests essentially utilized a single injection well surrounded by four or more producing wells. Sinclair Research, Inc., elected to do field experimental work using a number of test patterns in a single field in order that comparisons between various operating schemes could be made. The site selected and purchased in 1955 for this experimental work was a 60-acre Bartlesville sand reservoir located in Allen County, Kans. Combustion operations were initiated in mid-1956. Between that time and termination of the project in mid-1960, combustion fronts were propagated from injection wells to producers in three separate well patterns, using different mixtures of air and recycle gas. The test was terminated before sweep of the three patterns was complete so that information about the effect of combustion on the swept areas could be obtained by coring. Results from the test in the form of injection and producing well performance have been carefully recorded, and these form the general basis for this paper. DESCRIPTION OF RESERVOIR The reservoir in which the combustion tests were conducted is a Bartlesville sand "shoe-string", typical of a number of small reservoirs in Southeastern Kansas. Average reservoir characteristics are shown in Table 1. Fig. 1 is an isopachous map of the producing sand showing the reservoir to be approximately 400-ft wide and 2,500-ft long. Maximum net productive sand thickness is 21 ft. Fig. 2 shows a typical core analysis obtained by coring with water-base mud. The reservoir has no appreciable dip and is closed on the sides by degradation of sand into shale. The main body of sand is heavily laminated with shale stringers, which are not continuous between wells. The main reservoir is overlain by 30 to 40 ft of laminated low-permeability sand and shale streaks. No information is available on the original properties of