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By H. E. Montgomery

Engineering trends for some years have been toward higher temperatures, both in process work and in steam generation. Progress in design has been dependent upon the development and evaluation of suitable alloys. Elementary knowledge of the limitations of copper-base alloys indicate their unsuitability for the higher ranges of operating temperatures; say, for steam to 950°F. Emphasis upon the more severe conditions perhaps is a reason why more complete evaluations have not been made for copper-base alloys over their range of suitability. However, the few studies made show clearly the need for more accurate evaluations. Some copper-base alloys long looked upon favorably because of superior room-temperature properties are not as desirable as some others perhaps not so "strong" and "hard," as measured by usual acceptance tests. Load-carrying capacity and structural stability arc two important factors. Structural stability as influenced by temperature, by presence of corrosive media, or by the effects of stress itself, frequently is a limiting factor. An early reference to creep properties of the cast copper-base alloys is that of Brown1 (Admiralty gun metal, Cu 87.5, Sn 10.1, Zn 2.3 and Pb 0.08). His tests were similar to present-day stress-rupture tests without measurement of creep elongrttion. Loading was in steps with usually 24 hr. between steps, and continued until failure. He advocated a "useful strength value" of 61/2 tons (I4,560 Ib.) per sq. in. at 572°F. (300°C.). A specimen broken in 1/2 hr. at 572°F. had a strength of 21 tons (26,880 Ib.) per sq. in. The strength values as determined by 1/2-hr. tests and those by the longer stress-rupturc start to differentiate at 392°F. (200°C). This is the temperature at which time at load began to play an important part. Bolton, Price and Hebemann2 give creep curves on 88-10-2 (A.S.T.M. B-60, CU 88, Sn 10, Zn 2) and steam or valve bronze (A.S.T.M. B-6I, Cu 87.50, Sn 6.00, Zn 3.75, Pb 1.70) at 500°F. Direct comparison of the two alloys under 7500 lb. per sq. in. stress showed the steam bronze (A.S.T.M. B-61) superior in creep strength to the 88-10-2 (A.S.T.M. B-60). The steam bronze shows no creep after 700 hr. (1540-hr. test). The 88-10-2 shows a rate of about 7.0 per cent in 100,000 hr. over the last 700 hr. of the test. Creep curves on the steam bronze are also given at the same temperature (500°F.) with loads of 6500 lb. per sq. in. and 8000 lb. per sq. in. These indicate "creep strengths" (creep rate within I per cent in 100,000 hr.) in excess of 8000 lb. per sq, in. for this alloy. It was found that higher percentages of tin are detrimental to creep resistance. This is caused by the presence of or formation during test of considerable alpha-delta eutectoid in the higher tin alloys. Bolton also states that the high

Jan 1, 1945

By C. C. Herrick

The vapor pressure of indium has been measured by the torque-effusion technique, as a function of temperature between 1102o and 1422oK. For liquid indium, the vapor pressure (in atmospheres) can be represented by the equation -R In P = 5.782 x104/T - 25.29from 1000o to 1422oK. Extrapolation yields a calculated normal boiling point of 2329°K and a heat of vaporization at 298 oK of 58.09 kcal per mole by the third-law method and of 58.39 kcal per mole by the second-law method. Calculations indicate that the free-energy functions of Hultgren are to be preferred to those of Guruich. BECAUSE of the potential uses of liquid metals, it is important that convenient experimental methods be developed which will lead to a sufficient body of thermodynamic data so that reasonable estimates of properties of liquid metals can be made. Indium is particularly well-suited to investigation, since it remains liquid over a long temperature interval, it alloys readily, suitable containers for it are easily procured, and relatively few investigations of its thermodynamic properties in the region above 1000°K have been made. PREVIOUS RESEARCH The first determination of the vapor pressure of indium was made by Anderson,' using the Knudsen method and silica-lines stainless-steel cells. Kohlmeyer and spandau2 determined the normal boiling point, and a mass-spectroscopic investigation has been made by Lyubimov and Lyubitov.3 (Unfortunately, only an equation representing the data of the latter investigators has been available to us.) Hultgren et al.4 and Gurvich5 have published discordant tables of the free-energy functions for indium. EXPERIMENTAL METHOD The torsion-effusion principle has been described amply in the literature.8,7 With this method, one observes the angular deflection induced by effusion of vapor from a multihole effusion cell which is suspended from a filament producing a small rotational restoring force. The pressure within the effusion cell can be computed from the relation ziaifiai where k is the torsion constant of the filament, 6 is the observed angular deflection, and ai, fi, and qi are the area, force factor, and moment arm of the ith effusion orifice. The force factor f is analogous to the Clausing factor,' and corrects for the reduction in effusive force attributable to the finite geometry of the effusion orifice. Schultz and Searcy' have tabulated the force factors for various tube geometries. EXPERIMENTAL APPARATUS The physical arrangement employed in this study is shown in Fig. 1. The furnace elements are six Globar heat rods with 15-in. hot zones. Temperature is maintained to * 1°K by a Beck mechanism driving two 26-amp Powerstats in parallel. A Pt (Pt, Rh 10 pct) thermocouple serves as the sensing element. Four inner tubes act as vibration isolators; the air pressure required for maximum isolation was determined empirically, and is maintained by a solenoid valve. The tungsten fiber, 9 by 0.002 in., is joined to an 0.008-in. tungsten wire or graphite rod via a machined lightweight chuck.

Jan 1, 1964

By Robert J. Linney

IN the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the " 21 " Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. 1, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously developed sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel to footwall and hanging wall to make a bench, spalling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig. 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore; by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1945

By Robert J. Linney

In the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the "21" Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. I, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously devel-oped sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel te footwall and hanging wall to make a bench, spdlling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore, by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1946

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By M. H. HARRINOTON

THE Transactions of the Institute afford abundant evidence of the general recognition by mining engineers of the importance of a safety-lamp which will not only give warning of the presence of fire-damp, but will also indicate the proportion in which it is present.' Mining-men are only too familiar with the frequent and sudden changes that constantly occur at the working-face of a gaseous mine. Nothing illustrates more forcibly the pregnant condition of the strata enfolding the seam and the seam itself, than the fact that a circulation of perhaps 150,000 cu. ft. of air may contain 2 or even 3 per cent. of marsh-gas, representing a volume of from 3,000 to 4,000 cu. ft. of this gas pouring-out each minute from the natural strata into the mine-workings. It would, indeed, be strange if this great outflow of gas were always constant, and it creates no surprise that such earth-breathings, as they may be called, are spasmodic and irregular in their wane and flow. There are other causes that act to vary the gaseous condition of mine-air. Gas has a tendency, often very manifest in mine-workings and airways, to stratify and, at times, to travel long distances in veins or streams. It is common to catch and lose the gas in the lamp when making a test. At times two lamps, held side by side in an air current, will indicate different gaseous conditions of the passing air. When testing for gas, the lamp may suddenly fill with flame, in a mine-chamber, where a moment before a careful test showed the presence of but 1 per cent. of gas. It is sometimes remarked that there is little need of detecting a smaller percentage of gas than is revealed by the common

Jan 1, 1906

By A. T. Aldred

HIS note reports the existence of several new scandium intermetallic compounds of the A2B and AB stoichiometries where the A element is scandium and the B element is from group VIII or IB of the periodic table. Alloys were arc melted from high-purity materials using 2-g charges. Typical lot analyses for the materials used are given in Table I. Chemical analyses were not made on the alloys as there was little or no weight loss on melting. After annealing the specimens for 14 days at 600°C and water quenching, metallographic and X-ray techniques were used to determine the occurrence of the various phases. Table II lists the relevant data for the A2B phases. The Ti2Ni structure,2 has ninety-six atoms in a face-centered cubic cell, whilst the Al2Cu structure is body-centered tetragonal.3 The occurrence and stability of the various A2B type phases in transition metal alloy systems has recently been considered by Nevitt.4 The Ti2 Ni-type phases occur over a relatively narrow range of radius ratios from 1.14-1.26 with A partners mainly from the titanium group and B partners predominantly from the cobalt group. This suggests that both a favorable relative size of the atoms and a favorable electron concentration are necessary conditions for the existence of this structure type. A12Cu-type phases involving Ti group elements as A partners are formed with B partners from the cobalt and nickel groups, predominantly the latter, and have a higher radius ratio (1.27-1.29) than the Ti2Ni-type phases. The occurrence of both Sc2Pd and Sc2Co can be rationalized in terms of these considerations but SczNi is anomalous because of its high radius ratio. The radius ratios are calculated from the Goldschmidt CN12 radii and it might be postulated that the scandium atom is reduced in size in the compound. Unfortunately the apparent atomic size of scandium in the compound cannot be calculated from a knowledge of the structure and lattice dimensions. The crystal structure of the Sc2Au compound could not be identified from the powder diffraction pattern, but it is interesting to note that it is isostructural with a series of A2Au phases where A is a rare earth element.5 Pertinent information concerning the AB phases found in this investigation is given in Table 111. All the phases have the ordered bcc CsCl structure. The occurrence and stability of CsC1-type phases have recently been discussed by Dwight,7 who concludes that the relative sizes of the atoms do not appear to be a controlling factor (the radius ratios of all the known phases vary between 0.985 and 1.439) but that the occurrence of the phase is predominantly dependent on the relative positions of the A and B partners in the periodic table. As a measure of the stability of a given CsCl phase it has been customary to define a lattice contraction on alloying, DAB — dAB, where DAB is the AB interatomic spacing as calculated from the Goldschmidt CN8 radii and dAB is the AB interatomic distance in the compound (= v3/2 ao). Values of the parameter DscB -dScB for the compounds found in this investigation are given in column 5 of Table III and are plotted in Fig. 1. If the lattice contraction does in fact reflect the stability of a phase then Fig. 1 shows two very interesting trends when the position of the B partner in the periodic table is considered. Firstly, within any given group the stability increases in order on going from the first to third long period. Secondly, within any given period the stability increases in order as the B partner is taken from the copper, nickel, and cobalt groups. In the case of the second long period, the lattice contraction reaches a maximum at ScRh and then decreases to ScRu. The same trend could

Jan 1, 1962

By LUIS EAIYLNN SALAS

IF the ordinary wet method be attempted for silver-bullion containing tin, much trouble is experienced, varying with the amount of tin present. Even with a percentage as low as 0.05, the end-point is masked by a persistent turbidity, while with amounts ranging from 0.5 per cent. upward, the determination of the exact end-point is impossible, owing to the finely divided or colloidal metastannic acid resisting all efforts to cause it to settle, or to give a clear supernatant liquid. The object of the following experiments was to seek a remedy for these difficulties, and to find the conditions whereby the Gay Lussac method might be applied directly to these bullions. Material containing tin is met with frequently in bullion from Mexico or Bolivia; and tin is carried (sometimes in large quantities) in jewelers' sweeps. 1. A bullion was prepared containing Ag, 74; Cu, 25; and Sn, 1 per cent. A weighed amount was treated by the humid assay-method as if it were an ordinary bullion free from tin. Ten cc. of HNO3 (1.21) was added, and the bottle was heated until all traces of red fumes had been expelled. Before heating, the solutions were moderately turbid; and after heating, the turbidity greatly increased. One hundred cc. of normal NaCl solution was added and the bottle was shaken vigorously at intervals during 30 min. The turbidity persisted. The bottle was allowed to stand over

Mar 1, 1912

By Benjamin Gi ldersleeve

Uniquely colored, thin-bedded quartzite is quarried between Crossville and Crab Orchard in Cumberland County, Tenn. It is produced in all sizes up to the limits of transportation from beds usually ranging in thickness from less than 1/2 in. to 6 in. The stone is very resistant to wear and weathering and is marketed throughout the country for many different purposes in exterior and interior construction. The annual production exceeds 36,000 tons, valued at nearly $1,000,000.

Jan 1, 1950

By D. S. Sanders

IN the four mills of the Cerro de Pasco Copper Corp. in Peru, some 3000 tons of complex sulphide ores are treated daily, with four kinds of concentrates produced: copper, lead, zinc, and pyrite, each containing values in silver, gold, bismuth, and some other metals. Copper occurs in the various ores as chalcopyrite, enargite, tetrahedrite, tenantite, bournonite, and chalcocite; lead as galena; and zinc as sphalerite and marmatite. Small amounts of silver and tungsten minerals are also present. The gangue consist mainly of shale, quartzite, calcite, rhodochrosite, pyrite, and pyrrhotite. In general the values are freed with a relatively coarse grind, about 70 prr cent minus 200 mesh being the finest to which the pulps are ground. In Morococha and Casapalca however. sphalerite particles much finer than this contain inclusions of chalcopyrit r. At Cerro de Pasco the silver and gold are present in all sizes of the ground pyrite.

Jan 1, 1945

The Library is a cooperative activity of the American Institute of Mining and Metallurgical Engineers, the American Society of Civil Engineers, the American Institute of Electrical Engineers and the American Society of Mechanical Engineers. It is administered as a public reference library of engineering and the allied sciences. It contains 133,000 volumes and pamphlets, 6700 maps and receives currently most of the important periodicals in its field. In order to place the resources of the Library at the disposal of those unable to visit it in person, the Library is prepared to furnish lists of references to engineering subjects, copies or translations of articles, and similar assistance. Charges sufficient to cover the cost of this work are made.

Jan 1, 1933

By D. Harper

The National Pocahontas Mine in Wyoming County, WV, has been developed in the Pocahontas No. 3 coal. During 14 months of ventilation surveys, the locations of large rockfalls, many in areas infrequently visited by mine personnel, were mapped. Several dozen rockfalls were later studied in detail. Kettlebottoms, rider seams, root-penetrated roof strata, slicken sides, and shales that spall are common in areas with a high density of rockfalls. Large-scale crossbedding and micaceous and carbonaceous films in sandstones are less common. Density of rockfalls in any area is related to the time elapsed since mining, thickness of overburden, and orientation of entries and panels. Some linear features overlie areas where rockfalls are densest. Most rockfalls occur during summer.

Jan 1, 1983

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 U308 /t) (0.80 lb U308) 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.

Jan 3, 1978

Edward Dyer Peters, the only child of Henry Hunter Peters and Susan Barker Thaxter, was born in Dorchester, Mass., June 1, 1849. From his father he was a descendant of the Peters family of Ipswich and Andover, Mass., whose earliest ancestor settled in Essex County in 1659; while upon his mother's side he descended from several old Hingham families, as well as from the Quineys and from Governor Bradford. Between the ages of fourteen and sixteen he was a student at the Episcopal School for boys in Cheshire, Conn. Near the school was an old tin mine, and he used to spend his Saturday afternoons exploring it. When Edward was sixteen, his father sold his property in South-borough, and for the next three years lived in Europe. Sept. 26, 1865, Edward entered the Royal School of Mines at Freiberg, Saxony. In the summer of 1867 he joined his family on Lake Geneva, and spent a part of the season in that vicinity, going to Chamounix and taking other trips. Extract from a letter to his father: Elbingerode, Harz, Aug. 30, 1868: I left Freiberg with my friend Lilienthal, on Wednesday the 19th of August, reached Leipzig in the evening and spent the night in that city: the next morning we took the cars and in the afternoon reached Eisleben, in Prussia, which was to be our first stopping-place; there are immense copper-mines and smelting-works at this place, and the method of working the mines is very peculiar ... The next morning we got up at five and spent the whole day in the smelting-works where I obtained a number of drawings of the furnaces they use here and a good many valuable data. Saturday morning we got up at four o'clock, and descended a mine about three miles from the town; the Director of this mine was one of the most intelligent and agreeable men that I ever met; he accompanied us in the mine and explained everything that was new to us. The system of mining at Eisleben is peculiar to this town, and is not used at any other place in the world; the copper ore comes in narrow, flat veins eighteen inches high, at the utmost, consequently the men work lying flat on their left side and using the pick with both hands: the ore is carried out of the side passages into the main ones, in very low waggons about twelve inches high, and four by two feet; the boys who pull these waggons have a board buckled to their left thigh and their left arm; they pass their bare foot through a strap in the waggon and wriggle along on their side dragging the waggon after them, with their foot, almost as fast as a man can walk above ground. In the afternoon we walked over to Hittstadt, about twelve miles, and remained there Sunday and Monday. Monday night, at ten o'clock we went out to one of the smelting-works, to see the refining of the copper; this is only done at night, so we worked there from ten in the evening till eleven the next morning, and came home all blackened up and with blistered hands, for our skin was not quite so tough as that of the workmen: at noon the same day we took the stage to Ascherls-leben, and then the train to Stassfurt, where there are immense salt-mines, as well as factories for the manufacture of bromium, which latter process is, however, kept a profound secret, so we had very little hope of seeing anything but the salt mines; we were talking with our landlord about our disappointment, that evening, when he told us that the director of the Bromium works was sitting in the next room with no company but his glass of beer, so we went over to him, introduced ourselves as Americans, and actually succeeded in persuading him that it could do his business no earthly hurt if he showed us his works; at last he consented, after making us give our word of honor

Jan 1, 1919

By K. Sallmann

Spurred by a variety of factors, coal flotation is making headway among the preparation plants of West Germany. The author gives some statistics on German coal flotation plants and information on the properties of the feed and quality of the derived products. Types of machines and reagents used, dewatering practices, tailings disposal, and thickening operations are covered. There is special emphasis on the many-sided problems that confront German coal preparation engineers. The coal mining industry of West Germany is concentrated in three coalfields: the first on the Ruhr and the Rhein rivers, the second on the Saar river, and the third near the frontier between Germany and the Netherlands, around the towns of Aachen and Erkelenz. The total run-of-mine production of these three coalfields amounts to 736,000 tpd which come from 145 collieries. There are 121 washeries, 43 of them with a flotation plant. The total throughput of these 43 washeries is 280,000 tpd, out of which 26,000 tons (9.3 pet of the washery feed coal) are cleaned by flotation. The capacity of the individual flotation plants varies within relatively wide limits, the average being 40 tph and the capacity of the largest flotation plant being 120 tph. During the last few years, the number and size of flotation plants have been steadily increasing, although flotation must be looked upon as an expensive and rather complex cleaning process. The considerations which have led to the widespread application of flotation may be summarized as follows: 1) In making coking coal, it is seldom possible to add uncleaned fines to the coke oven charge, as their ash content is too high. If, however, the ash content of the fines is reduced to a maximum of 7 or 8 pet, they can be admixed to the washed and crushed small sizes without difficulty. This means, naturally that revenue is increased since the price of coking fines is always higher than the price for dust or uncleaned wet fines. 2) To prevent silicosis and pneumoconiosis, water infusion and spraying of the coal are practiced much more today than ever before. As a result, moisture content of run-of-mine coal has markedly increased and the washery feed coal contains more and more slurry instead of dry dust. The price of slurry, how- ever, is very low, and in many cases it is impossible to sell filtered raw slurry with a moisture content of 20 to 25 pct and, at the same time, 20 to 25 pct of ash. Reduction of the ash content of these slurries improves the possibility of dewatering them and, in this way, also enhances the marketability of this product. 3) The new severe laws and regulations against pollution of air and rivers make it necessary to de-dust coal better than in the past and to reduce the quantity of the waste water from coal preparation plants and their solids content to a minimum. 4) Even if there were no compulsory reasons to clean the fines, flotation will often lead to an increase in the overall yield. Although not the only application, it appears that the treatment of coking fines is the primary field for the flotation process in coal preparation. It must not be overlooked, however, that a series of arguments may be advanced against flotation: 1) Flotation is an expensive process because, in addition to the cleaning operation itself, the dewatering of the froth and the disposal of the tailings is very costly. 2) Operation of a flotation plant requires well trained personnel. 3) Filtered froth has about a 20 pct moisture content and, therefore, if mixed with the cleaned small coal it causes an undesirable increase in water content of the coke oven charge. For this reason it is necessary to take additional measures for dewatering washed small coal and this, of course, entails additional expense. 4) Disposal of flotation tailings involves very difficult problems, particularly in the highly industrialized regions. An investigation covering 32 flotation plants of Western Germany shows the following composition and properties of their feed:

Jan 1, 1961

By B. E. Eakin, A. L. Lee, M. H. Gonzalez

Expeximental viscosity and density data of lour natural gases are presented for temperatures from WO to 340F and pressures from 100 to 8.000 psia. A correlation is also discussed and results reported. INTRODUCTION This investigation is one of several efforts by the authors to provide viscosity data for pure hydrocarbons and mixtures. Results of some pure hydrocarbons and their mix tures have been presented." Several predictive methods and correlations have also been reported.'," This paper presents the experimental viscosity and density data of four natural gases and confirms a correlation in a previous study. APPARATUS The viscometer was described previously. 2,6 A bank of stainless steel pycnometers was included to determine density in conjunction with the viscosity measurements (Fig. I). DATA AND MATERIALS The natural gases were furnished by the Atlantic Richfield Co. (Samples 1 and 21, the Continental Oil Co. (Sample 3) and the Pan American Petroleum Corp. (Sample 4). Table 1 shows the composition of these gases obtained by mass spectrometer analysis. The pure component viscosity data used in the correlation have been published."".'"', '" Experimental and calculated viscosity data for the natural gases are presented in Tables 2 through 5. CORRELATION Starling and Ellington"' have reported several semi-empirical expressions based to a certain extent on the theory of viscosity advanced by Born and Green.' The final expression presented by Starling and Ellingto'n is: i'. (micropoise) -- /t,, exp rX(7') p ..... (I) Eq. 1 was modified by Lee et 01." to represent mixture and pure component data simultaneously. This equation has the form: F = K exp [XpY] :.........(2) where K=- (7.77 i 0.0063M)Ti 122.4 + 12.9M +T .................(3) ' ' ' X= 2.57 + ---1914.5+ 0.0095M .... (4) Y = 1.11 -r 0.04 X ........(5) Over the pressure and temperature range studied in this investigation, this equation represents the data on methane, ethane, propane, n-butane and four methane-n-butane mixtures with a standard deviation* of ? 1.89 per cent. Density data by Sage and Lacey" were used to fit the equations. Eq. 2 gives reliable values of the viscosity of light hydrocarbons without prior knowledge of experimental viscosity values, but accurate density data must be available. Eq. 2 was used to predict viscosity values for the natural gases studied in this paper. The particular set of parameters contained in Eqs. 3 through 5 gave values that reproduced experimental data within 25 per cent. Needless to say, in using Eq. 2 better density values will give better viscosity results. However, those engineers who do not have accurate density values at hand or the time, or such facilities as a library, computer, etc. readily available, may feel uneasy in using Eq. 2. Therefore, the authors sought the easiest but by no means best density prediction method, which was reported by Kaf 30 years ago. A set of density values, which were calculated based on Kay's method, and a generalized compressibility factor chart were used to fit Eq. 2. The results are:

Jan 1, 1967

By Nicholas J. Grant, W. Michael Yim

The creep-rupture properties of nickel, in as-prestrained or prestrain-polygonized condition, were studied at 1300°F and 4000 psi, and also at 700°F and 26,000 psi. An improvement of strength was noted in both the as -prestrained and the prestrain-polygonized tests at each temperature. The as-prestrained tests, however, showed superior creep resistance to the prestrain-polygonized ones. From the results of X-ray and etch-pit studies, the strengthc gain was interpreted in terms of dislocation density and the stability of preformed substructures. COLD work is known to improve the strength properties of metals at low temperatures. At high temperatures, however, the strength improvement is often lost if gross structural changes, such as re-crystallization, take place prior to or during the test. Even in the absence of recrystallization, a prestrained metal may undergo structural changes— polygonization and the formation of well-defined subgrains—which importantly effect the resultant properties. The relationship between substructure and creep properties has received considerable attention in recent years. The work of wood1 indicated that the creep strength of aluminum and its alloys was closely related to the subgrain size. Hazlett and Hansen,2 and Ancker, Hazlett, and parker,3 working with nickel and nickel alloys, had shown that small amounts of prestrain at room temperature followed by a recovery anneal at 800°C (1470°F) improved the strength at both low and high temperatures. The strength gain was attributed to an increase in the number of subgrain boundaries as a result of the prestrain-polygonization treatment. More recently Coldren and Freeman4 studied the effect of substructures on the creep rate of nickel at 1550°F. Their conclusions indicated that the substructure size as measured by the number of boundaries was the factor that controlled the creep rate. In addition, the perfection of the subgrain boundary has been indicated to have an important role on the strength.' On the other hand, there are data which disagree with the above observations. Grant, Chaudhuri, Silver, and Ganow6 found no creep strength improvement by a prestraining treatment. In reviewing the role of subgrains on high-temperature creep, Shepard and Dorn7 concluded that the contribution of the subgrains to high-temperature creep strength was a minor one. Guard,8 in a recent study on poly-gonized nickel, has cast doubt on the validity of the postulate that subgrain size controls the creep rate. In the majority of the previous studies, the subgrain boundaries have been singled out as the most important factor in determining creep strength; however, there is no firm theoretical ground for the postulate that the subgrain boundaries act as effective barriers to dislocation motion.B910 Little effort has been directed toward the study of the relationship between creep strength and other structural variables, such as dislocation density, lattice strain,

Jan 1, 1963

By R. J. Stevens

The Roan Antelope Smelter commenced operations in October, 1931. As originally designed, its equipment consisted of one reverberatory furnace, 120 X 25 ft, two Peirce-Smith converters 12 X 20 ft, and one straight line casting machine. Since that time the following additions have been made: One reverberatory furnace 120 ft X 28 ft ('934) One reverberatory furnace 95 ft X 28 ft (1943) Two 12 X 20 ft Peirce-Smith converters (1934 and 1938) One 13 X 30 ft holding furnace (1938) One straight line casting machine and 2 ladle tilting quadrants (1934) The metallurgy of the smelting process is comparatively simple. However, there are several features of operation which are uncommon to most copper smelters: (I) The high grade of concentrate treated; this has always been about 50 pct copper. (2) The high flux burden on the charge. This is necessary because of the deficiency of bases in the concentrate and the high alumina content of the concentrate. (3) The frequent necessity of adding coal to the furnace charge to control the matte grade. (4) The high grade of matte produced. This has led to unusual converter problems. From the commencement of operations, the ore mined has come from the Roan Basin orebody. The concentrate produced from this ore consists mainly of chalcocite and shale with small proportions of other copper minerals. As mining operations extend to the Roan Extension orebody, the proportions of bornite and chalcopyrite in the concentrate will increase with a corresponding decrease in the copper content of the concentrate. Because of the concentrate grade the capacity of the reverberatory furnaces in terms of tons of copper produced per ton of coal burned is high. This ratio is usually over 2.2, and on especially favourable runs has reached 2.6. Reference to Table I shows that the iron content of the concentrate is unusually low. Only part of this iron is available for slag formation, the balance going to the matte. To make up this deficiency in bases, limerock is added to the furnace charge. This flux ensures a slag that is of the correct silicate degree, is suficiently fluid, of low specific gravity, and it counteracts the thickening effect of the alumina present. An ample supply of flux is obtainable from Ndola, 22 miles distant by rail. Although the reverberatory slag has a high copper content, the slag fall is low and the copper loss in the slag now rarely exceeds 1.0 pct of the total copper charged. The blister copper shipped is of such purity that only one fire refining operation has been necessary to produce a wirebar with properties comparable to electrolytic copper. Bismuth has always been the most troublesome impurity and consequently the problem of its control in the finished product has been of prime importance. Practically none is removed in the reverberatory furnace. The bulk of the elimina-

Jan 1, 1949

By R. J. Stevens

THE Roan Antelope Smelter commenced operations in October, 1931. As originally designed, its equipment consisted of one reverberatory furnace, 120 X 25 ft, two Peirce-Smith converters 12 X 20 ft, and one straight line casting machine. Since that time the following additions have been made: One reverberatory furnace 120 ft X 28 ft (1934) One reverberatory furnace 95 ft X 28 ft (1943) Two 12 X 20 ft Peirce-Smith converters (1934 and 1938) One 13 X 30 ft holding furnace 6 (1938) One straight line casting machine and 2 ladle tilting quadrants (1934) The metallurgy of the smelting process is comparatively simple. However, there are several features of operation which are uncommon to most copper smelters: (1) The high grade of concentrate treated; this has always been about 50 pct copper. (2) The high flux burden on the charge. This is necessary because of the deficiency of bases'in the concentrate and the high alumina content of the concentrate. (3) The frequent necessity of adding coal to the furnace charge to control the matte grade. (4) The high grade of matte produced. This has led' to unusual converter problems. From the commencement of operations, the ore mined has come from the Roan Basin orebody. The concentrate produced from this ore consists mainly of chalcocite and shale with small proportions of other copper minerals. As mining operations extend to the Roan Extension orebody, the proportions of bornite and chalcopyrite in the concentrate will increase with a corresponding decrease in the copper content of the concentrate. Because of the concentrate grade the capacity of the reverberatory . furnaces in terms of tons of copper produced per ton of coal burned is high.. This ratio is usually over 2.2, and on especially favourable runs has reached 2.6. Reference to Table , shows that the iron content of the concentrate is unusually low. Only part of this iron is available for slag formation, the balance going to the matte. To make up this deficiency in bases, limerock is added to the furnace charge. This flux ensures a slag that is of the correct silicate degree, is sufficiently fluid, of low specific gravity, and it counteracts the thickening effect of the alumina present. An ample supply of flux is obtainable from Ndola, 2 2 miles distant by rail. Although the reverberatory slag has a high copper content, the slag fall is low and the copper loss in the slag now rarely exceeds 1.0 pct of the total copper charged. The blister copper shipped is of such purity that only one fire refining operation has been necessary to produce a wirebar with properties comparable to electrolytic copper. Bismuth has always been the most troublesome impurity and consequently the problem of its control in the finished product has been of prime importance. Practically none is removed in the reverberatory furnace. The bulk of the elimina-

Jan 1, 1947