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By C. A. Hartnagel

The production of petroleum in New York state in 1942 amounted to 5,410,000 bbl. This represents an increase over each of the four preceding years, and, with the exception of 1937, is the largest production since the peak of over six million barrels in 1882, shortly after the discovery of oil in Allegany County. The oil industry in New York is virtually stabilized and there has been little or no exploration for new sources of supply in recent years. The boundaries of the present fields were established more than 30 years ago and since then no additional reserves have been found. Deep drilling for natural gas during the past 10 years both within and without the oil fields has failed to find oil in quantity below the present producing sands of Devonian age. Oil Since 1882, the New York fields have produced approximately 140 million barrels of oil. Of this amount, 65 million was produced up to the year 1919 by ordinary pumping methods; that is, without flooding by water or the use of other secondary-recovery methods. In the year 1919, when a law was passed making the use of water-flooding methods legal, the oil production, of the state amounted to only 851,000 bbl. In the 23 years since 1919, there have been produced in the state 75 million barrels, or 10 million more than in the previous 37 years, which included the period of flush production. Price.—On March 26, the price of crude oil was raised from $2.75 to $3.00 per barrel, hut this increase had little or no effect on the number of wells drilled. In the Allegany field, which accounts for three fourths of the state's oil production, 1259 wells were completed in 1942, 7 less than in 1941. The smaller number of wells drilled was not only due to the shortage of labor but also was caused by the fact that in water-flooding territory a definite program for drilling wells is followed, and the flush period of an oil well is not usually reached in time to profit by any given price increase. Daily Production.—All of the oil wells in New York state, some 21,000 in number, are small producers. The average daily oil production is 7/10 bbl. per well. Under good flooding conditions, a maximum production of 5 to 8 bbl, a day is reached about one year after the flood begins to act on the well. In 5 years the daily production usually is less than one barrel. Although most of the oil in New York is produced by flooding, an estimated 2500 old wells produce without flooding. Some of these produce as little as 1/20 bbl. a day and are pumped only at intervals. The small producing Wells are mostly on lands leased to the larger oil companies and later will be included in a flooding program. At present, they are pumped only because the owners of the land insist on the royalty. Government Aid Needed.—Because of its good lubricating qualities, New York oil is one of those for which there is an insistent and growing demand, but even at its present price of $3.00 per barrel, many of the operators are producing at a loss.

Jan 1, 1943

By J. A. Richardson

J. A. RICHARDSON[t]-Mr. Stuckey's account of the geology of the kaolins of North Carolina suggests that they exhibit some features similar to those of the Malay Peninsula. About one half of British Malaya (some 26,000 square miles) is occupied at the surface by igneous rocks, of which granite and its associated derivatives constitute fully 90 pct. The average granite is a biotic variety. It is predominantly potassic but carries, in addition, notable amounts of soda and lime so that albite and albiteoligoclase are generally common constituents of the rock. The minor derivatives from granite "magma" likewise include potassic and sodi-potassic members. Decomposition of the granitic rocks has led to the formation of kaolin in many localities, but, in general, the deposits are rather small and, at present, none is exploited commercially. True primary or residual kaolins (as defined by Stuckey; p. 48 Vol. 173, p. 2 TP 2219) have been developed from feldspar veins, pegmatites and from other granitic differentiates which cut through the larger masses of granite; they are usually small and narrow. Kaolin has been derived from aplite, quartz porphyry, quartz micro-porphyry, felsites and microgranites belonging to a complex known as the Chegar Perah-Pulai intrusion in northwest Pahang and southwest Kelantan. The bulk of the component rocks has been more or less strongly sheared. Kaolin-like end products have been formed from both sheared and nonsheared varieties. Kaolin has been formed also by the complete breakdown of trondjemite flanking the northwestern and western margins of the complex Bukit Ranjut syenite intrusion in northwest Pahang. These trondjemites range in composition from essentially quartz-free albiteplagioclasites to quartz-free microcline rocks; they are, in effect, the syenitic homologues of white sodic and potassic granite-aplites. The granites in Malaya have generally undergone more or less severe surficial weathering leading to the production of a spongy and rather cellular or a soft "mealy" material (sandy clay) down to depths of about 100 ft; in places the granite is strongly decomposed down to 150 ft. There has occurred, therefore, widespread partial kaolinization, but, in general, no extensive deposits of good quality kaolin seem to have been formed by these processes of decay. Frequently, the katamorphic processes of desilicification have advanced beyond the production of kaolin towards the formation of lateritic end products. The more important Malayan kaoiins seem to be confined to the western alluvial plains, particularly in areas either where granite crops out at surface, or where it forms the bedrock of the alluvium. There are numerous excellent examples around Gopeng and elsewhere in the Kinta Valley of Perak. Such kaolins belong to Stuckey's groups of secondary deposits, and, like them, these also may be bedded. Field observations have established the existence of what for want of a better name is referred to by the Malayan Geological Survey as "granite wash." It is believed that granite, undergoing extensive deeply penetrative and probably rather rapid destruction under humid tropical weathering (temperature around 90°F; rainfall about 100 in. per annum; high humidity) "flowed," was moved by "creep" and, in addition, probably slumped in large waterlogged masses of sandy clay from topographic salients onto the adjacent flats. Either during this period of transmigration or later when the sludged granite detritus, after having more or less come to rest, was attacked by flowing water, differential sorting and erosion took place with the result that the resorted debris was moved and later deposited as bedded alluvium. In this way, banded "granite wash" was produced comprising alternating bands of clayey sand (resorted granite sludge) and white

Jan 1, 1947

By E. F. Raffo

THE design of the Chuquicamata concentrator offered an unusual combination of problems, all of which had, in one way or another, a definite effect upon the final arrangement of all the equipment and necessary facilities. First of all, a remotely located project of this magnitude required careful selection of equipment with an adequate supply of replacement parts and in some cases necessitated stand-by spare equipment ready for immediate use. Fortunately, in the course of determining the proper flowsheet, in itself a prime factor in the concentrator design, it was possible to conduct numerous grinding and flotation tests at the Anaconda and Andes concentrators. As a result of these tests made under actual operating conditions, most of the major equipment best suited for treating the Chuquicamata ore efficiently was decided upon while preliminary design layouts were still being made. The design has been based primarily on the treatment of sulphide ore crushed to about 3/4 or 1 in. However, until the new sulphide crushing plant is completed it will be necessary to handle ore crushed to 3/8 in. in the existing oxide crushing plants and to some extent a mixed ore which has had its oxide copper content removed by leaching. This mixed ore, known as leached residue, will contain a certain amount of sulphuric acid from the leaching process which necessarily requires special protective measures. Several factors influenced the final location of the Concentrator. The only established fact known during preliminary design layout was the elevation of the dumping floor over the fine ore bin. This was predetermined by the desired level railroad grade across part of the existing tailing and leached residue dumps from the oxide plant to the bin. Next it was necessary that a site be chosen that would not require an excessively long haul from either the present oxide plant or the mine and in addition provide a fairly constant natural ground slope sufficiently wide to accommodate the length of the ultimate concentrator. The site as finally chosen to meet the foregoing restrictions provided a natural ground slope of 6 to 7 pct. This was excellent, especially in a region subject to earthquakes, as it assured that foundations of heavy equipment and footings for building and crane columns would rest on cut or undisturbed ground. However, this shallow slope, although having some definite advantages in the overall consideration of plant layout, resulted in the need for extensive studies in the general arrangement of equipment. Wasteful and excessive vertical drops had to be eliminated in the

Jan 1, 1952

By John R. Keith

Lead and zinc contents of elm, maple, and oak trees and of soils in which the trees grew were determined in samples from the Upper Mississippi Valley mining district, and from an area outside but contiguous to the district. The lead and zinc analyses of plant materials served to differentiate the mineralized and nonminer-alized areas. Stems contained about two times as much lead and about two to five times as much zinc as did leaves taken from the same stems. Lead and zinc analyses of all soil horizons could be used to differentiate between the two areas. Soils at most sample sites in the mineralized district seemed to be more reliable than plants for geochemical prospecting and are a more practical sampling medium. Soil samples from areas with thick loess cover, however, failed to indicate the presence of mineralized bodies whereas tree samples from the same sites contained anomalous amounts of lead and zinc. It is suggested that in areas of thick loess cover, plant samples could be used in prospecting for lead and zinc. The important part that plants play in the cycling of elements through the soil horizons is well-known. The concentration of elements in the upper soil horizons by the action of plants has been called the Goldschmidt enrichment principle.' The fact that plant species vary greatly in their ability to concentrate particular elements in their tissues is also well-known.' Agricultural literature has dealt extensively with relationships of the major nutritive elements in soils and plants3 The extent, however, to which plant actions cycle the minor nutritive and the nonmetabolic elements in the soil has not been adequately investigated. Lead is non-nutritive and toxic to plants, whereas zinc is a micronutrient in low, but toxic in high, concentrations.4 This report presents the results of a biogeochemical study concerned with the amounts of lead and zinc in some plants and soils that occurred both within and adjacent to the Upper Mississippi Valley lead-zinc district. Field work for this study was done in the summers of 1961 and 1962. SAMPLE TARGETS AND CHEMICAL ANALYSES Geologic and physiographic descriptions of' the district have been made by Heyl 5 and several other authors in a series of U.S. Geological Survey (USGS) reports and quadrangle maps.6-I Sedimentary rocks, mostly dolomite, limestone, shale, and sandstone, of Cambrian, Ordovician, and Silurian age predominate in the area. Mineral deposits are of the vein and cavity-fill type; the principal ore minerals are galena and sphalerite. Vegetation of the area was characterized by Curtis1 6 as southern hardwood forest interspersed with patches of oak savanna and prairie. Plant genera typical of the eastern United States Temperate Zone, including ash, cherry, dogwood, elm, hackberry, hickory, maple, oak, and other woody and herbaceous genera, are present. The natural vegetation has been greatly modified by agricultural, logging, and mining operations. The species of the original flora are present, but not in their former relative abundance. Soils of this district belong, for the most part, to the Brunizem and Gray-Brown Podzolic great soil groups. The Brunizems are the soils developed under prairie vegetation.'' The Gray-Brown Podzolics are the soils developed under deciduous forests,18 They are derived from both bedrock and from loess parent materials, and

Jan 1, 1970

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By F. E. Bedale

STUDY of several severe mine explosions that occurred during the winter of 1907 led to the belief that coal dust was a definite explosion hazard. The Consolidation Coal Co. was a pioneer in the early work to' eliminate this hazard and to combat the dangers incident to poor ventilation and the presence of methane. Maintenance of high humidity throughout the mines was one of the first measures adopted, later to be superseded by rock-dusting. In the twenty years from 1907 on, many safety measures were developed and adopted, and several prizes were won, though the practice was crude by present-day standards. But in 1927, the company went to work in earnest, and organized a Department of Safety with adequate personnel and with authority to enforce all approved rules for the prevention of accidents and the preservation of company property. All plans for new construction work, changes in mining systems, and proposals for the purchase of new mining equipment had to receive the approval of this Department. A set of operating- and safety standards was written to govern the activities of all employees, and physical examinations for them were required.

Jan 1, 1938

By George S. Rice

INTERESTING developments going on in European coal mines look to: (1) increasing mechanization; 12) concentration of mining; (3) improvement in safety appliances; and (4) studies in bettering roof supports and. incident to this, studies of the way a roof behaves under different conditions of support and methods of mining and back-filling. Mechanization and concentration off mining is being adapted to a considerable extent in Great Britain coincident with the development of conveyors. Belt conveyors, shaker-pan conveyors, and a combination of the two are used. For pitching seams not steep enough for the coal to run in station-chutes, the. shaker-pan conveyor is more popular. For large outputs from a contiguous face, the double-unit system of face layout-that is, the conveyors meeting at a common point-is favored.

Jan 1, 1934

?Aaronson, Alfred E , Oil Producer, Pres, Tuloma Oil Co Tulsa, Okla '18 ||Abad, Leopoldo F , College of Mm , Univ of California Berkeley, Cal '23 ||Abadilla, Quirico A , Geol Dept, Cia Mexicana de Petroleo "El Agmla," S A , Edifico "El Aguila," Esq, Caller Aurora y Comercio, Tampico, Tamps , Mexico '20 ||Abbey, Robert Graham, Student, Case School of Applied Science Cleveland, O '21 Abbott, Clarence E, Gen'l Supt, Iron Mines Div, Tenn Coal, Iron & R R Co, 1405 Minnesota Ave , Bessemer, Ala '04 Abbott, Louis Benjamin, Chief Engr , Consolidation Coal Co , Elkhorn Div Jenkins, Ky '15 Abbott, Robert R , Met Engr , Peerless Motor Car Co Cleveland, Ohio '12 ||Abbott, Wells Osborn, Student, Univ of Arizona Tucson, Ariz '23 Abe, Asaka Ashiya, Mukogun, Hmyogoken, Japan '14 Abeel, George H, Jr , Cons Min Engr 4612 W 17th St , Los Angeles, Cal '18 Abel, Albert E , Asst to Pres , Valley Mould & Iron Corpn Sharpsville, Pa '19 ?Abeloff, Miran S 322 W 45th St, Ness York, N Y '21

Jan 1, 1923

By L. H. JEFFRIES

In the mining operations of The Canyon Corp., Deadwood. S. D., the use of mucking machines has been of definite advantage. The type used is that which depends upon the traction of the motor-driven wheels to crowd the dipper into the muck pile, while the car in which the muck is received is coupled to the mucker and moves forward or back with it. This general arrangement is entirely satisfactory on a substantially level track. The ores which occur in this district are a replacement of dolomite and the beds are now tilted to an inclination of about 7-1/20. In our operations it is necessary to drive inclines down-hill on this slope, which makes a difficult hand-mucking job even if water were not present. With these facts in mind, an attempt was made to use the mucking machine for driving down the in- cline, but as expected, the traction was insufficient to handle the machine and the necessary car. The means used to correct this condition will perhaps impress most readers as too complicated and cumbersome to effect any economy, but the idea may be of service to some who have a similar condition. After considering several s

Jan 1, 1940

By S. T. Wlodek

The surface and subscale oxidation reactions were followed by means of continuous weight-gain and metallographic techniques over the range 1600" to 2200°F (871° to 1204 °C) for up to 400 hr. Full identification of all scale and subscale reaction products was obtained by electron and X-ray diffraction. At or below 1800°F (982°C) a linear rate of reaction (QL = 46.0 kcal per mole) governed the oxidation process, extending for up to 100 hr at 1600°F (871 "C). During linear oxidation the surface scale consisted of an amorphous SiO2 film overgrown with Cr 2O 3 and NiCr204. This initial linear process was followed, and above 1800°F completely replaced, by two successive parabolic rate laws (Qp = 60 and 57 kcal per mole). This parabolic reaction involved the formation of a complex scale consisting of Cr2 O3 and smaller amounts of NiCr2O4. Parabolic oxidation appeared to coincide with the disruplion of the silica film present during linear oxidation and was followed by subscale (internal) oxidation of crystobalite and NiCr2O4. The balance between the subscale and surface oxidation reactions controls the oxidation of this commercial alloy. The amorphous silica film appears to result in the linear rate and diffusion through Cr2O3 is the more likely rate-limiting step during parabolic oxidation. THE oxidation of a multicomponent composition is a complex phenomenon not presently amenable to a rigorous classical interpretation. Nevertheless, even a qualitative understanding of the scaling and subscale reactions that occur in a commercial composition can illuminate the reactions that limit its high-temperature stability in an oxidizing environment. This study of the oxidation of Hastelloy Alloy X presents the first of a series of studies with the above approach in mind. Hastelloy X exhibits one of the best combinations of strength and oxidation resistance available in a wrought, solution-strengthened, nickel-base alloy. Although during long time exposure some precipitation of M6C and M23C8 carbides as well as a complex Laves phase occurs, the amounts are probably small enough to have no appreciable effect on the chemistry of the matrix. Radavich has identified the oxidation products on Hastelloy X oxidized for 5 min to 10 hr at 1115°F as NiO and the NiCr2O4 spinel. Oxidation for 5 to 15 min at 1500°F produced a scale of spinel, NiO, and a rhombohedra1 phase, probably Cr2Os. Sannier et 2. have reported continuous weight-gain data for Hastelloy X at 1650" and 2010°F and internal-oxidation measurements after 150 hr at 2010°F. In addition, much of the data on binary Ni-Cr alloys recently reviewed by Kubaschewski and okins' and Ignatov and Shamgunova4 as well as studies of binary Ni-Mo alloys5 are also pertinent to the oxidation of this composition. EXPERIMENTAL Continuous weight-gain measurements and metallographic measurements of subscale reactions were the main experimental techniques used in this study. X-ray and electron diffraction backed up by a limited amount of electron-microprobe analysis served to characterize the nature of the scale- and subscale-reaction products. Two heats of commercial sheet of the composition given in Table I and identified as A and B were used in the bulk of this study. Internal-oxidation measurements were made on a third heat of material in the form of a 0.5-in.-diam bar. In order to assure homogeneity, all heats were reannealed 4 hr at 2175°F prior to sample preparation. weight-Gain Measurement. All specimens (1.5 by 0.4 by 0.03 in.) were abraded through 600 paper, electropolished, and lightly etched in an alcohol-10 pct HCl solution. An electrolyte of 150 cu cm H,O, 500 cu cm HsPO4 (85 pct conc), and 3 g CrO3 at a current density of 0.9 amp per sq cm or a solution of 10 pct HaW4 in alcohol used at 4 v and 0.3 amp per sq cm was used for electropolishing. The resultant surface exhibited a finish of 3 ± 1 p rms. Continuous weight-gain tests were made at 1600°, 1700°, 1800°, 1900°, 2000", and 2200°F on auer' type balances capable of recording a total weight change of 110 mg with an accuracy of k0.1 mg. All tests were made in air dried to a dew point of -70°F and metered into the 2-in.-diam reaction

Jan 1, 1964

By R. F. Slack, J. C. Draper, J. A. Younkins

In 1959, the Duquesne Light Co. was faced with the problem of moving more than 40 million tons of bituminous coal in Greene County, Pa. The coal had to travel a distance of about 4 ½ miles from a new mine portal in Rocky Run Valley to the company preparation plant and loading facilities on the Monongahela River. After careful study of this immense transport problem, Duquesne decided that the belt conveyor system described in the following pages offered the best solution. This system, now in operation, is probably one of the four or five longest permanent overland conveyor systems in existence. The general layout of the entire system is shown in Fig. I. A detailed description of the salient features of the system is presented below. THE MINE The valley of Rocky Run where the system begins is very, very narrow and very steep. It was chosen as the site of Duquesne's new mine portal because it was the only place on this particular property where the Sewickley coal seam cropped out. The valley also had two other advantages. First, it fronted on a state road and, second, it offered a site where some 4 million gallons of water could be stored for use in dust-allaying sprays needed in the mine. The company elected to construct a drift mine in the valley because this would require neither the hoisting time of a shaft mine nor the walking time of a slope mine. As part of the mine construction project, the valley was leveled by relocating its stream and filling its center with material cut from each side.

Jan 11, 1966

By Thomas Henderson

Strategically, intermaterials competition is important to companies involved in raw materials extraction, smelting, refining or other processing, semi-finished and finished goods fabrication, and end-use consumers. In other words, just about every company involved in manufacturing. The array of materials covered by other papers in this session is testimony to the subject's breadth. In trying to provide an over- view of such a broad topic, the purpose has been to relate some of the common characteristics of the markets, economics and technology of raw materials, and to describe some of the ways in which Intermaterials competition impacts the markets, economics and technology of raw materials.

Jan 1, 1982

By J. C. Melrose, W. R. Foster, J. G. Savins, E. R. Parish

Many differences can be imagined between gas-oil flow in which the gas is supplied at the face of the core and gas-oil flow in which the flowing gas was originally dissolved in the oil. If capillary pressure characteristics and flow requirements control gas saturation distribution, the gas would be expected to be located at preferred sites within the porous medium as determined by pore sizes. On the other hand, during solution gas drive the gas first appears as bubbles through a nuclea-tion process. Nothing in self-nucleation theory specifies at which sites the first bubbles should be formed. In all probability they will be randomly distributed throughout the porous medium. Furthermore, it is not at all certain that even at low rates of production the gas will redistribute itself after nucleation to the channels normally occupied by gas in simple gas flow. Stewart, et at, have shown that at least for some limestone samples, oil recoveries could not be predicted for all rates of production using any one set of relative gas and oil permeabilities. An important factor in controlling recoveries during solution gas drive was the rate of bubble formation, higher rates giving higher recoveries. Stewart, et al, attributed the increase in recovery to a better distribution of the gas phase in heterogeneous limestone samples than is obtained by simple external gas drive. Differences in recovery from these causes were not reported for sandstone cores. In the experiments to be reported here, oil recovery, pressure and producing GOR history were measured during solution gas drive for a 5-ft sandstone core. The results were compared with predictions from the Muskat method for computing solution gas-drive behavior using external gas-drive relative permeability. The effects of changing the rate of production and oil viscosity were studied. At high laboratory rates of average pressure decline, two observations were made which would not have been predicted by Muskat's depletion theory: (1) oil recovery increased with increasing rate of production for a given viscosity oil, and (2) oil recovery increased with increasing oil viscosity for a given high rate of production. Both of these observations are explained as consequences of diffusion control of gas saturations superimposed on the normal gas-oil flow requirements, Fur- thermore, discontinuous gas phase flow appears to be significant during solution gas drive. The laboratory tests were performed at rates of average pressure decline many times greater than the maximum possible rate of average pressure decline in an actual oil field. It is, therefore, not possible to draw any direct conclusions regarding the effect of rate on recovery for the solution gas-drive mechanism under actual field conditions. However, at the lower laboratory rates, recoveries were nearly independent of rate and could be predicted by the Muskat method, using external gas-drive relative permeability data. These results suggest that at normal oilfield rates the effect of rate on recovery for the solution gas-drive mechanism is negligible. EXPERIMENTAL PROCEDURES The core material for the pressure depletion studies was Bandera sandstone from an outcrop in the Mid-Continent. This sandstone was selected because of its low permeability (about 10 md), which would permit the development of substantial pressure gradients in the corn at moderate flow rates. The core was 5-ft long and 2-in. in diameter. Its properties are listed in Table 1. Relative gas-to-oil permeability ratios were measured by an external gas-drive method2. The results are shown for a short 2-in. core and for the 5-ft core in Fig. 8. Oils used in the pressure depletion experiments were kerosene and a highly refined white oil (standard white oil No. 3) with gas-free viscosities of 1.8 and 25 cp, respectively. The gas was a naturally occurring methane from Gough field, Inglewood, Calif. The oil viscosities, gas solubilities and formation volume factors are plotted as functions of pressure at 75°F in Figs. 1 and 2. Methane viscosities and compressibilities were obtained from the literature"'. A core mounting was required which could withstand up to 2,500 psi internal pressure. This was obtained by first encasing the core completely in a plastic resin (Scotch Cast, manufactured by Minnesota Mining & Manufacturing Co.) The plastic covered core was then inserted into a steel pipe equipped with screw caps so that the plastic coating could be pressured from

By N. A. Gokcen

THE data of Chaudron,1 Tonosaki,2 and Collins³ on the thermodynamic properties of MOO, disagree widely. These authors, by using essentially similar methods, studied the following reaction: 1/2MoO2(s) + H2(g) == i/2Mo(s) + Ph,o H2O(g),K = —----- [1] Ph2 Tonosaki used a vacuum system consisting of a furnace containing MOO, and a water saturator whose temperature was kept at 25 °C with a thermostat. After repeated evacuation, hydrogen was admitted slowly into the system. The experiments were based upon the fact that at a constant furnace temperature and constant partial pressure of water, the total pressure of gas mixture over MOO? + Mo is constant. Any attempt to vary the pressure by external forces would vary only the amounts of MoO2, and Mo, after which the pressure should return to the equilibrium value in accordance with the equilibrium constant of reaction 1. The actual value of K was determined from the total equilibrium pressure (sum of Ph2o + Ph2) at each temperature. The total pressure of gas was varied within the range of 80.2 to 125.4 mm Hg, within a temperature range of 645" to 823 °C. The results were summarized as log K = 0.9413 — 1444.6/T for reaction 1. The ratio of H2O/H2 was considered to be uniform in spite of the presence of a thermal gradient across the static gas phase. It was shown by Rosenqvist and Cox,' however, that in somewhat similar circumstances the error resulting from thermal diffusion may be large. Collins³ improved Tonosaki's method by attempting to avoid thermal diffusion errors. His equilibrium data were obtained at 700°, 800°, and 900° C, and the result for reaction 1 was expressed as log K = 1.258 — 1822/T. The purpose of this investigation was to study reaction 1, avoiding the thermal diffusion errors, and to obtain equilibrium data in a considerably wide temperature range for the reliable extrapolation of the resulting thermodynamic functions. Experimental Procedure The diagram of the apparatus is shown in Fig. 1. Tank hydrogen was passed through a tube containing platinized asbestos at 425 °C in order to convert a trace of oxygen into water vapor. The flow rate was carefully controlled with a capillary-type flow-meter B and a bubbling column D. The gas was then presaturated sufficiently at P and led into a condenser system immersed in a thermostat, controlled to within ±0.002C. The temperature of the thermostat was measured with a thermometer calibrated against a certified standard. The temperature of P was adjusted to avoid the condensation of unduly large amounts of water as judged from the rate of flow out of the capillary tube M. Argon was purified by passing it through magnesium chips kept at 630°C. After passing through the flowmeter B', it was mixed with moist hydrogen emerging from the thermostat. The resulting gas mixture was then led into the hot zone of the furnace through the heated glass tubing and the silica tube S, thus insuring the same ratio of PII2o/Ph2 from the condenser to the reaction chamber, thus avoiding thermal diffusion. The entire gas system was of all-glass construction, except at the magnesium train. The furnace comprised a glazed alumina tube over which a 15-in. platinum coil was wound. The lower end of the alumina tube was tightly closed with a brass bottom and a silicone rubber gasket, and the upper end with two glass disks, each with a hole of 1/16 in. in diameter, through which the gas mixture escaped into the atmosphere. A back-up coil of kanthal wire facilitated the temperature control of the furnace. A coil of annealed molybdenum strip of 99.99 pct purity, 0.005 in. thick and 0.050 in. wide, and weighing 17 g was hung in the furnace with a 0.010-in. platinum wire attached to one end of a sensitive analytical balance. The temperature of the furnace was measured with a Pt-Pt-10 pct Rh thermocouple checked against a standard. The experimental procedure consisted of heating the gas purification trains, adjusting the gas flow rates, attaining a constant thermostatic temperature, flushing the entire system for 2 hr while heating the furnace to well above the expected equilibrium temperature and then cooling it at a rate of 0.3°C per min during which time the change in the weight of molybdenum was observed on the balance. For a given thermostatic temperature, i.e., a constant Prr,o/Px,, molybdenum oxidized upon cooling slightly below the equilibrium temperature. The procedure was then repeated by heating the furnace and thus reaching a temperature slightly above the equilibrium value. The average of the two temperatures, differing by 2" to 3"C, was considered to be the true equilibrium temperature. In order to determine the stoichiometric composition of the oxide phase present in this investigation,

Jan 1, 1954

By Thomas J. Menner

Jan 1, 1983

By David E. Nicholas

INTRODUCTION In this paper, a numerical process for selecting a mining method, with the emphasis on underground mass mining techniques, such as caving, induced caving, and stoping, is proposed. In the past, selection of a mining method for a new property was based primarily on operating experience at similar type deposits and on methods already in use in the district of the deposit. Then, the chosen method was modified during the early years of mining as ground conditions and ore character were bet- ter understood. Today, however, the large capital investment required to open a new mine or change an existing mining system make it imperative that the mining methods examined during the feasibility studies and the method actually selected have a high probability of attaining the projected production rates. Although experience and engineering judgment still provide major input into the selection of a mining method, subtle differences in the characteristics of each deposit, which may affect the method chosen or the mine design, can usually be perceived only through analysis of measured characteristics. The parameters that must be examined when choosing a mining method include: 1) geometry and grade distribution of the deposit; 2) rock mass strength for the ore zone, the hanging wall, and the footwall; 3) mining costs and capitalization requirements; 4) mining rate; 5) type and availability of labor; 6) environmental concerns; and 7) other site-specific considerations. This paper encompasses a detailed look at the first two parameters since they, plus mining costs, have the greatest impact on the selection of a mining method. The proposed method selection process is for a project where drilling has defined sufficient geologic reserves, but little or no underground development has been done. Since each deposit has its own characteristic geanetry/grade distribution, and rock mechanics properties, mining method selection should be at least a two-stage process. In Stage 1, the deposit is described in terms of geometry, grade distribution, and rock mechanics properties. Using these parameters, the mining methods can be ranked to determine which are most applicable; they can then be considered in general terms of mining and capitalization cost, mining rate, type and availability of personnel, environmental concerns, and other site-specific considerations. In Stage 2, the most likely mining methods are costed out, based on a general mine plan. Mining and capitalization costs are used to determine a cut-off grade fran which a minable reserve can be calculated; economic comparisons can then be made to determine the optimum mining method and economic feasibility. During the mine planning phase of Stage 2, rock mechanics information would be used to provide realistic estimates of underground opening size, amount of support, orientation of openings, and caving characteristics, and open pit slope angles. If ground control or operational problems should be encountered with the methods being considered, modifications could be made. Although planning on paper extends start-up time, it is cheaper to err on paper than to find the error after mining has begun. METHOD SELECTION - STAGE 1 The main purpose of Stage 1 is to select those mining methods which should be considered in greater detail. The simplest way to do this is by defining those characteristics required for each mining method and then determining whether the characteristics of the deposit are suitable. However, no one mining method is so restrictive that it can be used for only one set of characteristics, as indicated by the classification system proposed by Boshkov and Wright (1973). In the mining method selection proposed, geometry, grade distribution, and rock mechanics characteristics are ranked according to their acceptability for ten general mining methods.

Jan 1, 1981

By E. R. Hoskins

Results of a primary stress measurement investigation conducted on the 1100 copper ore body of the Mt. Isa mine, Mt. Isa, Queens., Australia, are presented in this chapter. The investigation started in September 1965, after discussion with the engineering staff at Mt. Isa, and consisted of both underground or in-situ tests at Mt. Isa and laboratory tests on Mt. Isa rocks and stress measurement techniques at the Australian National University in Canberra. Underground stress measurements were made in and adjacent to the 1100 copper ore body. The primary stress measurement site was in J 32 west crosscut on 14 Level. Approximate coordinates on the mine grid are 3000 north, 1650 east. The vertical depth below ground surface at this point is approximately 2180 ft. Some preliminary measurements were made in M 30 west crosscut on 13 Level, coordinates 3000 north, 2000 east, depth beneath the surface approximately 1990 ft. Additional flatjacks were installed on 11 Level in L 30 west crosscut, coordinates 3000 north, 2200 east, and depth beneath the surface approximately 1580 ft. All of the sites were in rock that from a mining engineering viewpoint must be considered extremely sound. The 13 and 14 Level crosscuts were approximately 14 ft on a side and nearly square and the 11 Level crosscut was approximately 10 ft on a side. All of the crosscuts had been standing open for several years and no artificial support of any type was necessary. Extraction openings hundreds of feet on a side have been made in this mine in similar rock and these have remained open also without additional support (see Davies 1 for a recent summary of mining techniques and conditions). The geology of the Mt. Isa mine has been most recently described by Bennett.2 Briefly, Mt. Isa is a silver-lead-zinc and copper mine in a series of Precambrian siltstones, shales, and bedded carbonates that strike north-

Jan 1, 1972

By Irving Crosby

ELECTRICAL prospecting by potential methods has been applied to mining problems for some years and determinations of the depth to bed rock have been made, but so far as is known it has not been used previously in the solution of civil engineering problems, such as investigating the foundations of dams. During the past winter and spring the senior author has been engaged as consulting geologist in studying the site of a large darn on the upper Connecticut River and has had the satisfaction of introducing electrical prospecting to foundation problems. The valley across which the dam is to be built is underlaid by a deeply buried preglacial gorge which is filled with glacial deposits. An extensive drilling campaign was in progress, but the work was unusually slow and expensive on account of the number of boulders in the overburden. The necessity of speeding up the work was pressing and it was decided that much of the needed information could be obtained by geophysical methods. After investigating the various methods and conferring with the company engineers and with the consulting engineer, Albert Crane, a contract was made with the Schlumberger Electrical Prospecting Methods, and an observer with apparatus was sent to the dam site the first of April, 1928. The problem was to determine the depth to rock at various points in order to prepare a contour map of the bed-rock surface and outline the buried gorge in the vicinity of the dam site. Over one hundred determinations of depth have been made and the bed-rock topography, is now known in considerable detail. In addition, some information has been obtained about the nature of the overburden. The drilling has been continued; in fact, two additional drills have been put to work, and the electrical determinations have been checked at several points. At one of these the depth to rock given by the electrical work was 142 ft. and the drill found rock at 147 ft., an error of less than 4 per cent. The electrical work has also been checked by the geological study, and it has been found that when the electrical determinations are in agreement with the geological expectations they are probably reasonably accurate. When the electrical interpretation of the sub-surface topography is incon-

Jan 1, 1928

By D. G. Westlake

Polycrystalline tensile specimens of various Zr-0-H alloys have been tested at 298°, 178°, and 77°K. Solute oxygen and hydride precipitates in quenched alloys made individual contributions to the yield strength at 0.2 pct strain which combined to produce a resultant strength increment, a,., Ductility changes which were ohserved can he interpreted in terms of the various oxygen and hydrogen concentrations, testing tem -peratures, and dispositions of the hydride. ADDITIONS of oxygen in solid solution were known to increase the yield and tensile strengths of polycrystalline zirconium as early as 1951.' More recently, the critical resolved shear stress (CRSS) for prism slip in zirconium single crystals was also shown to be affected by the solute oxygen impurity.' This latter work also demonstrated that large increments of strength could be contributed by the finely dispersed zirconium hydride precipitates that are present in quenched Zr-H alloys.3 It was concluded that the combined strengthening due to alloying could be expressed by where to is the increase in the CRSS due to solute oxygen alone and TH is the increase due to finely dispersed hydride precipitates. Eq. [I] is analogous to one used to express the combined strengthening effects of work hardening and neutron radiation damage.4 Eq. [1] was verified only indirectly and for only small amounts of the impurities—up to 0.14 at. pct 0 and 0.63 at. pct H. The present investigation was undertaken to obtain a more direct verification of the validity of the form of Eq. [1] for this system and also to determine the combined effects of oxygen and finely dispersed hydride precipitates on the tensile strength and ductility of polycrystalline zirconium. EXPERIMENTAL PROCEDURE Tensile specimens were machined from the same rolled billet of Kroll zirconium used in the earlier study.' These measured 38 by 4.7 by 0.5 mm and had 10-mm gage lengths which were 2.8 by 0.5 mm. Each specimen was ß-annealed in vacuo at 1173°K for 15.5 hr and a-annealed at 1073°K for 4 hr to D. G. WESTLAKE, Member AIME, is Associate Metal l ur-gist, Metallurgy Division, Argonne National Laboratory, Argonne, III. Manuscript submitted July 17, 1964. IMD______________ give an equiaxed structure with grain diameters averaging 0.06 mm. Oxygen was added by allowing the metal to react with a known quantity of oxygen during the 0 anneal and known quantities of hydrogen were added during the a anneal. Each alloy was encapsulated in Pyrex under vacuum, annealed at 873°K for 4 hr, quenched into ice water, and polished by immersion in a solution of 46.75 vol pct H2O, 46.75 vol pct concentrated HNO3, and 6.5 vol pct HF (49 pct) at 298°K. Special heat treatments given to a few specimens are described in the results below. Tensile tests were done on an Instron machine and were begun within 20 min after the quench, except where specified otherwise. Tests at 298°K were in air, at 178°K in acetone, and at 77°K in liquid nitrogen. All tests were at a strain rate of 8x sec-1. RESULTS AND DISCUSSION Yield Stress at 298°K. The compositions of alloys and the corresponding yield stresses (0.2 pct strain) are given in Table I. A plot of the yield stresses of the oxygen alloys, A, B, C, and D, indicates that varies linearly with CO1/2, where Co is the oxygen concentration, Fig. 1. This is in accord with Fleischer's6 theory for solution strengthening if the oxygen atoms do not cluster, or the cluster size remains constant with increasing oxygen concentration. In Fig. 1, it appears that if one could prepare some oxygen-free zirconium its yield stress would be very low. Therefore, we shall assume that for the oxygen alloys is equivalent to O0, the strength increment contributed by the presence of oxygen. The relationship between0.2and Co is expressed by 0.2 = 31.3 CO1/2, when the yield stress is in kg per sq mm and the concentration is in at. pct. Each of the hydrogen alloys, Al, A2, A3, and A4, contained 0.081 at. pct 0 as an impurity. In Fig. 1, it appears that this small amount of oxygen makes a significant contribution to the strength which cannot be ignored when we evaluate the contribution of the finely dispersed hydride. Let us assume the validity of the following equation: a0.2 = (a2o+a2R)1/2 [2] which is analogous to Eq. [I] for single crystals, and calculate values of UH for the hydrogen alloys by using the experimental values of 0.2 and o (0.081 at. pct) = 8.9 kg per sq mm. For 0.36 at. pct H, oH = 6.47; for 0.72 at. pct H, OH = 11.30; for 2.16 at. pct H, OH = 19.4; and for 3.60 at. pct H,

Jan 1, 1965