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By Thomas R. Mager

An in.t!estigation has been made of the equilibrium conditions at 1200°C in the reaction between hydrogen sulfide gas and sulfur dissolved in Fe-Si alloys From this the equilibrium constant, activity coefficient, and activity of sulfur in solution were calculated. A number of studies of the equilibrium of sulfur with iron and iron alloys have given closely agreeing results from which the activity and free energy of the dissolved sulfur may be found. Sherman, El-vander, and chipman1 discussed the significant researches of dilute solutions of sulfur in liquid iron prior to 1950, and the results of this study indicated that the relationship between the ratio of PH2S/PH2 in the environment and the percentage of sulfur in solution is not a linear one. Morris and williams2 studied the equilibrium conditions in the reaction between hydrogen sulfide gas and sulfur dissolved in liquid iron and Fe-Si alloys, and reported that silicon dissolved in iron has a pronounced effect on the equilibrium conditions. They found that the activity of sulfur in iron is increased by the addition of silicon. At a silicon content of 4 pet the activity coefficient of sulfur was about twice that for sulfur dissolved in pure iron. Sherman and chipman3 investigated the chemical behavior of sulfur in liquid iron at 1600°C through the study of the equilibrium: H2 + S = H2S; K = PH2S/PH2 . 1/as [1] From the known equilibrium constant of the reaction between H2, H2S, and S and the experimental data, the activity of sulfur in the melt was determined. They found that the activity coefficient of sulfur defined as fs = as/%s is increased by silicon and decreased by manganese. Morris4 and Turkdogan5 also reported that manganese decreases the activity coefficient of sulfur in liquid iron and iron-base alloys. A recent technique of sulfur analysis developed by Kriege and wolfe6 of the Westinghouse Research Laboratories permits an accurate sulfur analysis of 0.5 * 0.2 ppm in the range of 0.1 to 3 ppm, whereas in the range of 3 to 50 ppm the accuracy is ±1 ppm. This technique of sulfur analysis was utilized in this experiment. Previous unpublished data reported that sulfur analysis by the combustion technique was not accurate below 20 ppm. EXPERIMENTAL PROCEDURE Five 5-lb ingots of high-purity Fe-Si were prepared. Three of these ingots were prepared without the addition of manganese but with a variation of silicon contents from 2 to 4 pet. The remaining two ingots contained 3 pet Si with the addition of manganese. Ingots were made at each of three silicon levels: 2, 3, and 4 pet. No alloys were made with less than 2 pet Si since below approximately 1.8 pet Si the binary alloy exhibits a to ? transformation. The two additional ingots of 3 pet Si-Fe were made at each of two manganese levels: 0.20 and 0.50 pet. To minimize the effects, if any, of impurities on the activity of sulfur on Si-Fe, the best metals available were used for melting. All ingots were vacuum-melted in magnesium oxide crucibles. After obtaining samples for chemical analyses, the ingots were processed. This consisted of hot rolling and subsequently cold rolling the alloys. Each ingot was hot-rolled at 1000°C, reheating between every pass to minimize grain growth. All heating was done in a protective argon atmosphere. The slabs were hot-rolled to strips 50 mils thick. After hot rolling, all the material was pickled to remove the scale formed on the surface of the strip during hot rolling. The material was then cold-rolled to 12-mil strips. Single strips of the material used in this experiment were hydrogen-annealed at 1200°C for 16 hr in an alumina tube. Chemical analyses of strips M-1, M-3, M-4, M-7, and M-8 are given in Table I. Sulfur, silicon, and manganese analyses were made from the millings from the cold-rolled 12-mil strips. The oxygen analyses were made from slugs of the as-cast material. The hydrogen sulfide used in these experiments was supplied from cylinders containing a mixture of argon and 1 pet hydrogen sulfide. The parts per million of hydrogen sulfide were determined from the analysis of the exit gas of the annealing furnace during each anneal. The flow rate of hydrogen was approximately 1 liter per min in all anneals. The

Jan 1, 1964

By D. F. Hewitt

Page Introduction..............................305 Geologic features of eastern Oregon....................306 Geology of the Sumpter quadrangle....................306 Rocks older than the ore deposits...................307 Rocks younger than the ore deposits.............. . Lodes of the Sumpter quadrangle..................312 Areal distribution.......................312 Geologic distribution.................... . Major features of the veins...................313 Mineralogy and texture of the veins.............. . Ore deposits zonally disposed around the Bald Mountain batholith . 317 Summary of mineralogy and texture.............336 Zonal relations......................337 Ore deposits zonally disposed around Greenhorn Mountain batholith. 340 Zonal relations..................... . Relation of textural features to zonal arrangement of veins.....341 Summary of zonal relations.................. . Comparison with similar districts................. . Outlook for the district......................344 Introduction In an earlier report on the Sumpter quadrangle1 it was briefly stated that the quartz lodes of that region might be separated into several groups or belts on the basis of the kind and abundance of the minerals of which they are composed, and that these belts were disposed as areal zones around the principal intrusive mass of the region. When the opportunity to revisit the region was offered in 1929, as a result of a cooperative agreement between the State Mining Board of Oregon and the United States Geological Survey, it seemed advisable to pay particular attention to this feature of the lodes, in the hope that if the idea of zonal relations was confirmed it might serve as a guide in exploitation.

Jan 1, 1931

By Patrick Corser, Mitch Usibelli

Introduction Coal mining in Alaska's interior, specifically in the Healy area, began as early as 1918 with the construction of the Alaska Railroad. Mining was originally limited to underground operations but has expanded to entirely surface operations. In 1943, the Usibelli Coal Mine was formed and started developing Alaska's first surface mine east of Suntrana (Usibelli Coal Miner, 1984). Production from the local coal deposits has steadily increased and, in 1978, surface mining of Poker Flats was initiated (Fig. 1). Currently, a 25-m3 (33-cu yd) walking dragline strips two coal seams, using an extended bench on the second pass. In addition, a fleet of trucks and shovels are used for coal removal and some limited overburden stripping. In 1984, a contract was signed between Usibelli Coal Mine and Sun Eel Shipping Co. in 1984. Since then, production has nearly doubled to more than 1.3 Mt/a (1.5 million stpy). This article will discuss geotechnical constraints on mining within the steeply dipping coal deposits that exist within the Poker Flats mining area. Specifically, the article will describe how the mining operation retriggered an historic landslide on the No. 5 coal seam (Fig. 2). And the article tells how a mine plan was developed that allowed the coal to be safely removed without inducing additional movement. Regional geology The coal-bearing group in the Nenana coal field is of Tertiary Age. It is overlain in some areas by several thousand feet of Tertiary gravels - the Nenana Gravels. In areas mined by surface methods, the Nenana Gravels have been eroded off, and up to 30 m (100 ft) of quaternary outwash gravels overlay the coal-bearing formations. The coal-bearing group is divided into five formations: Healy Creek, Sanctuary, Suntrana, Lignite, and Grubstake (Wahrhaftig, 1969). Lignite Creek lies on the north limb of a west plunging anticline. This has brought the Suntrana coal-hearing formations near enough to the surface to allow surface mining. Mining is presently in progress on the south side of Lignite Creek in the Poker Flats area. The coal-bearing formation is cut off to the south by a fault having perhaps several thousand feet of vertical displacement, with the upthrust side to the north. South of this fault, Nenana Gravels are exposed on the surface. The Suntrana Formation contain the minable reserves at Poker Flats. This formation is a repeated sequence of poorly consolidated pebbly sandstone near the bottom, grading through a silty fine sandstone to a footwall clay unit immediately below a coal seam cap. The footwall clays are high plasticity clays to silty clays. It has been reported that they contain 30% to 50% montmorillonite (Usibelli Coal Mine Inc., 1982). There are six coal seams in the Suntrana Formation, No. I (the lower seam) through No. 6. Only the top four seams are currently exposed. No. 3, No. 4, and No. 6 seams are the only mined seams. The No. 5 seam is very thin or not present. Portions of the undisturbed Suntrana Formation are overlain by up to 15 m (50 ft) of Quaternary outwash gravels or recent landslide rubble. The surface is overlain by a very thin layer of muskeg and isolated areas of permafrost. In many areas, the outwash gravels are found immediately below the surface muskeg. Numerous landslides have been documented along the north facing slopes of Lignite Creek (US Geological Survey, 1970, and Wahrhaftig, 1958). These appear to be surficial solifluction or skin flow types of landslides. In addition, deep-seated structurally controlled slides are also evident on both the north and south sides of Lignite Creek. Structural features Premining aerial photographs (Fig. 3) of the Lignite Creek slopes in the Poker Flats area indicate substantial evidence of deep-seated landsliding. The landslides noted in Fig. 3 are both inside and outside of the current mining area. Surface mapping and geologic exploration indicate that the coal seams are dipping out of the slopes within the noted slide areas. It is suspected that, historically, these landslides were triggered by undercutting of the toe of the slopes by Lignite Creek. And sliding it thought to have taken place on one or more of the clay beds underlying the coal seams (Golder, 1985). The slide areas are characterized by semicircular head scarps and slumped topography. Based on the premining photographs, these slides do not appear to have been recently active. However, they are expected to be in a state of only marginal stability. Extensive coal exploration indicates that the primary structural feature within the

Jan 1, 1989

By G. M. Brownell

Introduction The Flin Flon mine is located on the boundary line between the provinces of Manitoba and Saskatchewan at a point about four hundred miles northwest of Winnipeg. A branch of the Canadian National railway provides the town of Flin Flon with direct transportation facilities. The property was staked in 1914 and, after a considerable amount of exploration; the Hudson Bay Mining & Smelting Company was organized and brought the property into production in November, 1930 (1). Mining is carried on by open pit and underground methods at the rate of about 4,400 tons per day. Production for the year 1934 was: 99,334 ounces of gold; 1,335,319 ounces of silver; 37,486,054~pounds of copper; and 49,427,280 pounds of zinc; in addition to small amounts of cadmium and selenium. The Flin Flon deposit is one of the largest low-grade ore deposits in Canada with complete hydro-electric and metallurgical plants for the production of blister copper and electrolytic zinc. Previous Work The development of the Flin Flon mine into one of the major mining operations in Canada has created a need for a more detailed geological study of the area. Practically all of the information published to date concerns the general surface features. It is the purpose of this paper to deal more specifically with the ore-body and its relationship to the enclosing rock as revealed by studies made possible through recent underground development. The earliest published report on the area is by E. L. Bruce (2), who made a reconnaissance geological investigation of the region for the Geo-logical Survey of Canada in 1914 and established the main. petrological units. Several years later the area immediately around Flm Flon was mapped in greater detail by F. J. Alcock (3). In the summer of 1934, the area of Missi sediments to the north and east of the mine was studied by J. W. Ambrose. His report has not yet appeared in print but his findings are of importance.

Jan 1, 1935

By M. Kalin, H. D. Sharma

INTRODUCTION The inactive uranium mill tailings investigated in this study are located in two mining districts, Elliot Lake and Bancroft, Ontario, Canada. The sites exhibit a mixture of surface features consisting of dry areas, with or without vegetation and areas covered with water. On the edges of the water bodies, indigenous vegetation has invaded the tailings beaches [ Typha] spp. (Tourn.) L. a dominant plant on these tailings beaches, has been studied for the uptake of radium-226 and lead-210 (Kalin and Sharma, 1981b) from the tailings. It was found that most of the radium-226 remains in the roots of the plants, and that the solubility of radium-226 in control soil differs from that in tailings. The uptake of radium226 by vegetation and other biota is related to the solubility of the element in water. Factors controlling the solubility of radium-226 in uranium mill tailings are of interest in assessing the environments effects of these wastes. Rusanova (1962) found that the soluble or extractable amount of radium from the soil is inversely related to the total concentration of calcium and magnesium in the soil. Experimental work has clarified some aspects of the leachability of radium from uranium mill tailings (Levins, et al., 1978; Wiles, 1978, and others). Halvik, et al., (1967) studied the effects of pH and chemical composition of surface water on the liberation of radium from uranium mill tailings and uranium ore. He found that an increase in pH, up to a value of pH 9, decreases the amount of radium released from the tailings and the ore. A positive effect of calcium chloride was noted on the leachability of radium. Benes (1981) reviewed the physicochemical forms of radium and its migration in water. Based on experimental work, he identifies primary factors which determine leaching of radium from uranium mill tailings. The ratio of the volume of the leaching solution to the weight of the leached sample; the composition of the leached solids and the leaching solution, and finally the pH of the leaching mixture are of importance. He emphasized that a scarcity of field data exists, which would relate experimental work to the actual situation in the tailings ponds. Inactive tailings ponds in Ontario are 16 to 23 years old, and processes which are of importance in evaluating the long term effects of uranium mill tailings in the environment can be studied. The objective of this work was to investigate the leachability of radium-226 from the tailings under field conditions. MATERIALS AND METHODS Sample Collection A description of inactive tailings sites in Ontario, where the tailings and the water for this study were collected, can be found in a report by Kalin (1981). [Typha] spp. specimens were excavated from 16 different locations along with the tailings attached to the roots of the plants. A tailings sample from the surface area, around the plant, was secured before the excavation. Tailings samples were also collected from dry areas free of vegetation in the vicinity of the wetland stand. Surface samples (depth 0-5 cm) and samples at a depth of 20-25 cm were secured. Water was collected from the shallow tailings beaches around the vegetation stands. Sample Preparation The roots of the plants were washed free of tailings with distilled water. The resulting thick slurry of tailings was allowed to settle in the wash basin for five minutes and the supernatant water was decanted. The remaining saturated tailings were brought to 400 ml volume with distilled water in a beaker. The slurry was mixed with a magnetic stirrer for two 24-hour periods. After the first 24-hours, the slurry was allowed to settle for 20 minutes and the supernatant water was removed. The remaining sediments were suspended again in distilled water (400 ml total mixture) and leached for a second 24hour period. The tailings slurries from the different locations had a solid to liquid ratio, which ranged from 0.5 to 1.4 grams of dry tailings material per millilitre. The variations in the ratios are the result of differing fractions of coarse and fine tailings on the sites. In the final leachates, the solid to liquid ratios of the samples were 0.17 (±0.1) g/ml. The pH of the surface water and the leachate was determined with an I.L. Portomatic pH meter. All the tailings samples were homogenized in a mortar and brought to dryness at 75 to 85°C. Approximately 0.5 g of tailings were

Jan 1, 1981

By W. S. Potter

The proportions of manganese and carbon in manganese steel are familiar to all…

Jan 1, 1915

By V. F. Swanson

A petrographic study was made of the type and sources of mineral contamination occurring in Florida phosphate rock products. Emphasis was placed on contaminants in the screened rock (+1 mm), since this portion of the product undergoes no physical beneficiation except for scrubbing to remove slimes. Since iron and aluminum contents of the finished product affect the salability, special emphasis was placed on identifying types and sources of these contaminants and relating chemical composition to visual characteristics. Wavellite derived from the overburden was the major aluminum contaminant with some aluminum being included as secondary clay balls. Iron occurred as phosphates in the overburden and as colloidal oxides included in the phosphate pebbles. Chert and limestone contamination derived from the Hawthorne formation underlying the phosphate matrix.

Jan 1, 1972

By W. M. Dressel

A simple field test for detecting the presence of germanium in rocks, coals, and flue dust has been developed by the Bureau of Mines. The method consists of dissolving germanium from the samples by either acid or fusion treatment methods, adding a few drops of hematein or phenylfluorone dye, and extracting the resulting colored complex with carbon tetrachloride. INTRODUCTION Germanium is obtained commercially as a byproduct of zinc production, as a coproduct of base-metal refining from germanium-bearing mineral, and as a recovered product from the ash of some coals. The growing use of germanium in the electronics industry has increased demand for this semiconducting metal to the extent that an intensive search for new sources of supply is necessary.

Jan 1, 1962

By A. L. Mular

Optimization strategies of possible interest to mineral processing are reviewed and typical strategies initiated in processing plants to date are discussed. Both On-Line and Off- Line optimization procedures have been employed by industry, assisted either by operators or by digital computers. In every case, dollar benefits have accrued.

Jan 1, 1976

By R. A. Metz, Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (1), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profitability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus minimum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12). Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-lowgrade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1992

By Masakazu Takano, Koji Suzuki, Naotoshi Shinkawa

"Cement deep mixing (CDM) method successfully applied to a marine construction project in northern Vietnam is reported. The method is classified as wet method, in which cement slurry is injected into the ground. Thickness of clayey deposit subjected to improvement is from 20 m to 28 m, and the total volume of CDM improvement exceeds 460,000 m3. This is the first massive application of CDM method in the region. Three CDM barges, each equipped with four mixing shafts, were brought to the site in order to finish CDM work within a period of seven months. Cement content used in the work was 180 kg/m3 and water-cement ratio of slurry was 100%. Unconfined compression strength of stabilized soil was 1,200 kPa to 1,600 kPa in average with coefficient of variation from 0.3 to 0.4. Statistics analysis of the strength with these values gave percentage of lower strength less than 10%. Trial use of needle penetration test performed to study strength profile of a CDM column is also reported.INTRODUCTIONThe site of Lack Huyen Port Infrastructure Construction Project is located in the sea adjacent to Cat Hai Island, offshore Hai Phong city which is around 100 km east of Hanoi, the capital of Vietnam. Left side of Fig. 1 shows the location of the site. The project contains development of container terminal, dredging of waterway, construction of sand protection dyke to prevent waterway from being buried, and bridge construction to connect the island to the mainland.Joint venture of Penta Ocean Construction and Toa Corporation was awarded Package 6 of the project, in which reclamation of container terminal area, retaining wall construction and terminal road construction were included. Plan of the site is presented in right side of Fig. 1. The site work has been started in June 2013 and is scheduled to finish within 4 years."

Jan 1, 2015

By Desmond M. Levins

"Australia has almost twenty per cent of the Western World's low-cost uranium reserves, located mostly in the Alligator Rivers region of the Northern Territory. At present, only one uranium mill is operating in Australia, but a number of new mills are planned for the early 1980s.Details are given of Australian uranium mining and milling proposals and the measures taken to minimize their environmental impact. Major factors affecting environmental impact are discussed, including treatment of liquid wastes, water management, control of radon and other airborne releases, and disposal of tailings. IntroductionAustralia is entering a new phase in uranium mining and milling activity . In the first phase of operations, between 1954 and 1971, about 10,000 tonnes of yellowcake was produced at five locations. Most of the production came from the mines at Rum Jungle in the Northern Territory and Mary Kathleen in Queensland. In the 1960s, the demand for uranium slumped, and much of the yellowcake produced at Rum Jungle was stockpiled by the Australian Atomic Energy Commission. The early uranium processing operations in Australia were typical of overseas practice at that time. Environmental problems were generally given low priority and pollution control measures were unacceptable by today's standards. There was little or no recycling of process water, acid waste streams were not neutralized and tailings dams were poorly designed. At Rum Jungle, oxidation of sulphides in waste heaps, tailings dumps and ore stockpiles subsequently led to destruction of vegetation and heavy metal pollution in a nearby river.In the early 1970s, major uranium deposits were discovered in the Alligator Rivers region of the Northern Territory. Significant orebodies were also located in the Great Artesian Basin in South Australia, in calcrete formations in Western Australia and at several locations in Northern Queensland. According to latest available estimates. Australia has about 160/0 of the Western World's uranium recoverable below $US 80 per kg U. Mining these orebodies will require pollution control measures more stringent than practised previously in Australia."

Jan 1, 1980

The Mineragraphic Section of the Council for Scientific and Industrial Research has had occasion from time to time to examine rocks, ores and concentrates from the Yatukoula goldfield, Fiji. The results of the examinations have been supplied as Mineragraphic Reports and have been supplementary to various investigations carried out by the mining companies on ore treatment, or by visiting geologists on the nature of the rocks and lodes. The presence of tellurides was determined in 1935, though incorrectly as nagyagite. The general absence of lead, except in traces, was soon afterwards discovered, and it came to be recognised as the mines were opened up that the chief telluride minerals on the field were sylvanite and native tellurium. Specimens of the tellurides have been provided from time to time by E. G. Theodore, Managing Director of the Emperor Mines Ltd., and of the Loloma (Fiji) Gold Mines N.L., and further material was recently obtained, as the result of a visit to the field at the invitation of Mr. Theodore.The aim of this paper is to describe the occurrence of the tellurides in these specimens. The work has been carried out as part of the programme of research of the Mineragraphic Section.THE VATUKOULA GOLDFIELDThe Emperor, Loloma and Dolphin mines are situated at Vatukoula, six miles, south of Tavua, on the north coast...

Jan 1, 1949

By R. Grierson, L. J. Bonis

The high-temperature Strength oF TD nickel has been observed to be dependent upon the previons thermal and mechanical history of the material. Variations in both the level and the anisotropy of strength have been observed. 01 this paper- these variations are correlated with the storing of annealing resistant elastic strain energy in the matrix of the TD nickel. An x-vay line -broadening tecknique is used to measure the maLrTis elastie strain. THE inclusion of a finely dispersed second phase into a ductile matrix has long been recognized as an extremely effective method of strengthening the matrix both at high and at low homologous temperatures. It has been found, however, that the factors which determine the high-temperature strength are not the same as those which are important at low temperatures. Below 0.5 Tm the size and distribution of the second phase particles are of prime importance in determining the strength,')' while above this temperature the strength is mainly dependent upon the previous thermal and mechanical history of the alloy,3-7 This paper is primarily concerned with explaining the response of the high-temperature mechanical strength of one of these alloys (DuPont's TD nickel) to various thermo-mechanical treatments. It will be shown that this response is not associated with the occurrence of any form of dislocation substructure within the matrix of the alloy. It has been found, however, that a correlation does exist between the elastic strain level in the matrix and the previous thermomechanical history of the alloy and that the observed changes in elastic strain level parallel the measured changes in high-temperature strength. It therefore must be concluded that variations in high-temperature strength are a direct result of the variations in elastic strain level. MATERIAL TD nickel contains approximately 2 vol pct of Tho2 in an unalloyed nickel matrix. It is formed, as a powder, by a chemical technique and this powder is compacted to form ingots which are then extruded to give 21/2-in.-diam rod. Rod of smaller diameter is prepared from the as-extruded rod by swaging. In the studies reported in this paper, 1/2-in.-diam rod was used. This rod received an anneal of 1 hr at 1100°C prior to being used in any of these studies. EXPERIMENTAL TECHNIQUES Two methods were used to examine the structure of the nickel matrix of the TD nickel. These were: 1) transmission electron microscopy; 2) the analysis of the position and profile of X-ray diffraction lines obtained using the nickel matrix as the diffracting media. To prepare thin foils for electron-microscopical examination, slices of TD nickel approximately 0.050 in. thick were cut from the as-received 1/2-in.-diam rod. These were then chemically polished down to 0.045 in., rolled to 0.009 in., given a predetermined heat treatment, and thinned, using a modified Bollman technique, to provide the foils for observation. All observations were carried out at 100 kv, using a Hitachi HU-11 electron microscope. Specimens of the undeformed rod were prepared by grinding down the 0.050-in.-thick slices to approximately 0.015 in. and then thinning chemically and electrolytically to give the thin foils. The X-ray specimens were prepared by rolling 0.375-in.-thick rectangular blocks down to 0.075 in. The surfaces of the rolled material were ground flat, chemically polished to remove the layer disturbed by the grinding, and given a predetermined anneal in an inert atmosphere. They were then ground lightly to check their flatness and given a final chemical polish prior to being examined. The X-ray diffraction line profiles were measured using an automated Picker biplane diffractometer. A special specimen holder was built to allow a more accurate and reproducible positioning of the specimen. The line profiles were determined by carrying out intensity measurements at intervals of either 1/30 deg or 1/60 deg over a range of 3 deg on either side of the nickel peaks of interest. A piece of pure nickel which had been recrystallized to give a large grain size was used as a standard to give the X-ray line profile generated by a strain-free matrix. The analysis of the X-ray diffraction line profiles is a modification of that due initially to Warren and Aver-bach8and has been described elsewhere.3 This analysis gives a measurement of two parameters associated with the structure of the nickel matrix. These parameters are: 1) the size of the coherently diffracting domains within the nickel matrix; 2) the magnitude of the elastic strains in these domains. Both of these parameters are first determined in terms of a Fourier series. These series are obtained from other Fourier series which describe the measured profile of the X-ray diffraction lines. Thus, for both the coherently diffracting domain size and the elastic strain level, it is possible to plot Ft (the Fourier coefficient) against t (the term in the Fourier series), where t can be expressed in terms of a distance L and the Fourier coefficient Ft(S) (associated with elastic strain level) can be expressed in terms of the root mean square strain (e2)1/2. Thus a plot of (F 2)1/2 vs L can be obtained. Plots of this type are shown graphically in Figs. 6 and 8. Interpretation

Jan 1, 1968

By Terry J. Close

In 1992, as part of the U.S. Bureau of Mines Idaho Land Assessment Program, the Western Field Operations Center investigated the mineral resources of the 30,300-hectare Elk Summit study area. The area is in the Clearwater National Forest, adjoining the Selway-Bitterroot Wilderness in Idaho County. Idaho. The study area is underlain by granitic and metasedimentary rocks. Associated with these rocks are road metal (stone) resources, black sand mineral placer resources, and thermal water resources. Also present are abundant sand and gravel. Two of the five sites examined during the Bureau study have road metal (stone) resources, one has black sand mineral resources, one has thermal water resources. The fifth site is a black sand mineral placer occurrence. There is no evidence of petroleum or coal resources. Idaho County records indicate that 38 mining claims, mostly black-sand placer, have been located in the study area, Twenty-three of the claims are current. Workings consist of a few pits and trenches at the placer, and two borrow pits. The only mineral commodity produced from the area was about 1,500 tonnes of stone excavated from the Savage Pass and Swamp Creek borrow pits along the Elk Summit road, north of Hoodoo Lake. The stone was used to surface the road; these stone resources will continue to be used as needed.

Jan 1, 1993

By John C. F. Randolph

This establishment being now in full working order, it has seemed of considerable professional interest to collect together, in a concise form, the various points as to its plan, method of dressing, and equipment. The data contained in this paper have been largely drawn from the able paper of E. Kutscher, constrnoting engineer of the works, contained in the Zeitschrift fur Berg-HÜtten-und Salinenwesen im Preussischen Staate, 1873, supplemented by some personal memoranda made while the works were under construction and by memoranda on the works appearing in various German technical papers from their completion up to date. This is probably the largest dressing works in the world devoted to the beneficiation of argentiferous lead ores, and it is doubtful if any establishment for the concentration of mineral of any description can compare with it in size. As showing the final practice adopted in an old and very conservative mining region, in which the concentration of ores has been a matter of constant practice and experiment for a very long period, these works are of peculiar interest. The fact that the Clausthal ores strongly resemble a very large and abundant class of ores found in the United States has always attracted the attention of American mining engineers and mine owners to the methods employed in dressing and smelting them. The study of their methods of concentration of ores has, however, always been full of difficulty,

Jan 1, 1879

By R. M. Smailer

Detailed description of the development of the FIOR process1 and, subsequently, other technical papers relating to the operation and start up of the FIOR plant2,3 are available in the technical literature. The Fluidized Iron Ore Reduction (FIOR) process plant constructed in the Matanzas region of Venezuela represents the first commercial application of this technology. The FIOR process was developed by Exxon Company during the 1950's and 1960's; and the facility in Venezuela was designed and constructed by Davy McKee with the plant start up in 1976. While there is a minority government ownership in the company, FIOR de Venezuela operates as a private sector company with the majority of the ownership being private capital.

Jan 1, 1981

By Lyden, Joseph P.

THE Picher Mining Field, fig. 1, which lies between Baxter Springs, Kansas, and Commerce, Okla., is the most intensely mineralized and the largest zinc-lead ore producing area in the Tri-State District of Missouri, Kansas, and Oklahoma. It has a vast lateral extent of underground workings that is unequaled for the observation of the geology of zinc-lead ore bodies in sedimentary horizons; and thousands of churn drill holes in the unmined areas supplement the underground observations with useful data for the interpretation of the geology. This paper is confined to the ore deposits of the Picher mining field. The reader is referred to the bibliography for the general geology of the Tri-State district which has been ably covered by many geologists. The writer was co-author of several publications on the Tri-State district.- This paper is necessarily a repetition of parts of those publications but contains a few changes and additions on the deposition of the minerals. The parts of this paper concerning mineralization maps and pipe slump-structures are not covered in the above publications.

Jan 1, 1950

By Philip Bucky

THE general conception of a mine model is that of a three-dimensional object representing the mine workings, the orebody and the country rock of a particular property. Its chief uses have been to make it easier to understand the mine and to serve as a basis for the arguing of legal questions. Fayol endeavored to use mine models in a different way and to all students of subsidence "Fayol's dome theory" is of decided interest. We have more recently the suggestion of Young and Stoek of the University of Illinois, who1 suggest the use of models as a means of studying the problem. Prof. Henry Briggs, of the University of Edinburgh and Herriott Watt College, also makes use of models to determine the effects of bending at the edges of extensive workings.2 It is interesting to note that civil engineers are taking a decided interest in the application of models and the principles of dynamic similarity to the solution of their problems. The Proceedings of the American Society of Civil Engineers for 1930 contain an elaborate article by Benjamin F. Groat,3 who says, in his opening remarks: "Every engineer should know and understand that a model constructed in accordance with Newton's theory of similarity is not only the means for reproducing the complexities of flow, or other mechanical action, on a small scale, but that models may be made the best means of designing works of great magnitude. In the present state of knowledge, a properly constructed and properly tested model will answer the difficult questions of hydrodynamics much more quickly and accurately than the most profound mathematical analyses." A study of P. W. Bridgman's book on Dimensional Analyses leads one to arrive at the same conclusion.

Jan 1, 1931

By J R. Hernandez-Aguilar, C O. Go, C Acuna

The gas dispersion properties that are the subjects of this communication include three measured parameters, gas superficial velocity (gas rate), gas holdup and bubble size, plus derived parameter, bubble surface area flux. Over the past ten years sensors to measure these parameters have been introduced and those originating from McGill University are briefly described. The sensors have seen extensive use in plant campaigns and experiences on four cell types at three sites are discussed. The impact of frother in controlling bubble size by retarding coalescence is illustrated, introducing the need to understand chemistry/machine interactions in interpreting gas dispersion results. An increase in Sauter mean bubble size (D32) with superficial gas velocity (gas rate, Jg) was found, which followed the empirical relationship where Do ~ 0.5 - 0.6 mm and C and n are fitted parameters dependent on the system (machine and chemistry). Following this trend in Db - Jg, consistent bubble surface area flux (Sb) - Jg trends are also observed. The trend in Sauter mean versus number mean (D10) is used to help interpret the nature of the distribution; as the means approach numerically, the distribution is narrowing. Analysing gas holdup versus gas rate shows both a dependence on location in the cell as well as cell position in a bank, factors which must be accounted for when comparing data from different plants and circuits. The instrumentation and methodologies being developed are presenting new opportunities for incorporating cell hydrodynamics into metallurgical diagnostics. A case study at one plant serves as illustration.

Jan 1, 2005