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By E. J. Koval

Practical miscible displacement processes will be characterized by lingering of the solvent into the oil. The lingering process is brought on by viscosity differences, and can be accentuated by channeling and longitudinal dispersion. The effects of these factors on the efficiency of unstable completely miscible displacements are combined in what is called the K-factor method. This method, analogous to the Buckley-Leverett method, predicts recovery and solvent cut as a function of pore volumes of solvent injected. Experimental data are included and show excellent agreement with theory for a wide variety of sandstone cores and viscosity ratios. INTRODUCTION Theoretical considerations,1 laboratory experiments,2 and pilot tests3 lead to the conclusion that miscible displacements in the field will be unstable In an unstable miscible displacement, the solvent fingers through the oil. This fingering leads to early breakthrough of the solvent and an extended period during which both oil and solvent are produced. For such a system, there appear to be four principal factors which bring about or accentuate the effects of instability: longitudinal dispersion (including geometrical effects), channeling, viscosity differences and gravity differences. Other factors, such as diffusion and flooding rate, can also influence the effects of instability, but at the flooding rates considered in this report, circa 15 ft/D, they are unimportant. Longitudinal dispersion can be thought of as a spreading of the solvent front caused by the presence of microscopic inhomogeneities. Channeling of the solvent occurs when a porous medium has macroscopic inhomogeneities; i.e., gross permeability variations. Viscosity differences lead to fingering of the less viscous solvent. This difference in viscosity accelerates the growth of fingers along paths previously developed because of permeability variations. Gravity differences lead to overriding by the usually less dense solvent. Al- though gravity effects are generally small at a flooding rate of 15 ft/D, they would, nevertheless, unnecessarily complicate the interpretation of flooding experiments. For this reason, all the experiments reported herein were done with matched density fluids. Fingering and the resultant poor areal sweep were recognized early as the dominant influences on the efficiency and the economics of miscible displacement processes. Much research effort has been spent on ways to minimize fingering and increase areal sweep such as the use of graded viscosity slugs4 or water slugs. Some researchers attempted to work out ways to prevent fingering completely; i.e., to achieve a stable displacement through gravity or latitudinal dispersion stabilization. l Others did not attempt to control fingering but obtained an economic process by merely recycling the solvent and sweeping pattern by pattern.6 During this period, all aspects of fingering came under close scrutiny. Some researchers reported on how fingering looks2 and how it is affected by viscosity ratio,7 geometry,7 and slug size.2,* Peaceman and Rachfordg suggested a mathematical approach to the prediction of unstable miscible displacements in relatively homogeneous sand packs, but their work cannot be extended conveniently to heterogeneous systems. Hence for a heterogeneous system, no method is presently available for predicting solvent cut and recovery as functions of pore volumes of solvent injected. The purpose of this investigation was to attempt to fill in the gap in our knowledge concerning the prediction of performance of unstable miscible dis placements. Necessarily, the system selected for study was a relatively simple one. The restrictions placed on the system were: (1) The system was linear; (2) The solvent was miscible in all proportions with the oil in place; (3) The solvent was continuously injected into the porous medium; (4) Gravitational effects were eliminated by matching densities and (5) All the flood rates were high and constant at 15 ft/D to avoid any small rate effect and to minimize any diffusion effects. To simplify and to indicate that both longitudinal dispersion and channeling arise from permeability variations, the effects which they cause or influence have been termed heterogeneity effects. The use of

By G. B. Glass

In 1996, 92.5% or 233 Mt (257.4 million st) of coal produced in Wyoming came from the Wyodak coal zone in the eastern Powder River Coal Field. While the mined portions of the zone consist of one to three beds of coal with an aggregate thickness between 15 and 33.5 m (50 and 110 ft), the thickest expression is 61 m (200 ft). The delivered coal from this zone averages 0.33% sulfur, 5.12% ash, and 19.98 MJ/kg (8,590 Btu/lb). This coal zone occurs at or near the top of the Paleocene Fort Union Formation, which is overlain by the Eocene Wasatch Formation. There is also production from two coal beds in the lower portion of the Wasatch at the southern end of the basin. Production of coalbed methane derived from the Wyodak coal zone is currently an estimated 10.2 billion cu ft a year. This production is occurring just west and downdip from the coal mines and is increasing each year.

Jan 1, 1998

By H. H. Podgurski, H. E. Knechtel

Annealed Fe-Al alloys do not react readily to form AlN when held at 500ºC in NH3-H2 gas mixtures, but do so upon the introduction of dislocatims. Nuclea-tion of the nitride phase occurs on dislocation sites. In turn, the growth of the aluminum nitride particles causes the ferrite phase to yield plastically, generating more dislocations for the nucleation process. The nitride phase extracted from an Fe-2 pct A1 alloy nitrogenated at 500°C was identified as stoichio-metric aluminum nitride with a hexagonal crystal lattice. THIS investigation reveals the role that dislocations play in initiating and sustaining the nitriding reaction in Fe-A1 alloys. As early as 1931 the work of Meyer and Hobrock1 suggested that the initiation of the nitriding reaction could involve a nucleation controlled process. Recently Bohnenka2 depicted the gas-phase nitriding process below 600°C as one of mixed control limited by nitrogen penetration through the surface, by nitrogen diffusion, by aluminum diffusion, and by nucleation of the nitride phase, Fig. l(a). In our research in a comparable alloy (0.57 pct Al) at 575ºC, we have observed a nitrogenation which we feel is better described by Fig. l(b). In the case of a 2 pct-A1 alloy partially nitrided at 500°C we propose the profiles shown in Fig. l(c). For a complete and accurate description of the process, a concentration profile of the dislocation density in the test specimen would be needed. EXPERIMENTAL Nitrogenization was conducted between 500" and 575°C in a variety of NH3-H2 gas mixtures on three Fe-A1 alloys: 1) zone-refined iron + 0.16 i 0.2 pct Al—levita-tion melt, 2) zone-refined iron + 0.57 0.02 pct Al— levitation melt, 3) plastiron + 2 pct Al—melted by induction heating. To demonstrate the effect of dislocations on reactivity, both cold-worked and annealed samples were investigated. All nitrogenation rate studies were conducted gravimetrically with a gold-plated invar balance4 contained in a gas-flow system. To avoid contamination of the specimens in the reaction zone of the system, the reaction chamber was constructed of high-purity dense alumina. The activity of nitrogen was fixed by specific NH3-H2 gas mixtures whose compositions were continually monitored by calibrated thermal conductivity gages and checked by chemical analysis. Variations of ± 0.1 pct NH3 could easily be detected by both methods. Throughout this paper the activity of nitrogen is defined as PN3 /PH23/2 , where PNH3, and Ph2 are partial pressures in atmospheres. Electron transmission, density measurements, and chemical analyses were made on specimens before and after nitrogenating in order to reveal structural and chemical changes. Similar studies as well as X-ray diffraction studies were conducted on nitride extractions from the nitrogenated 2 pct-A1 alloy. These extractions were obtained by the use of an anhydrous bromine-methyl acetate solution which dissolves the iron and leaves the insoluble nitrides as a residue. Hardness profiles were obtained on cross-sections of partially nitrided specimens to demonstrate the extent of nitriding through the thickness of the specimens. RESULTS AND DISCUSSION The nitrogen activity in the NH3-H2, atmospheres was never allowed to reach a level capable of producing iron nitride (Fe4N). Hence, the term nitriding as used in this paper refers only to the formation of aluminum nitride whereas nitrogenation refers to the total uptake of nitrogen regardless of how it is distributed throughout the alloy. The weight increases observed during the initial stage of a nitrogenating treatment are due primarily to the solution of nitrogen in the ferrite phase, particularly when starting with annealed specimens. Because this initial nitrogenation rate in the case of the 0.57 pct A1 alloy, see Figs. 2 and 3(a), was most rapid the weight change that occurred thereafter might be attributed to the nitriding reaction with the exception of a small weight increment due to the irreversible pickup of oxygen by aluminum. The oxygen (<70 ppm) came from traces of H2O and 0, in the hydrogen and ammonia gases. On the basis of discrepancies between total weight increase and the increase in the nitrogen content of the sample as determined by chemical analysis, it was estimated and later established by activation analysis, that as much as 200 ppm of oxygen were taken up by a fully nitrided Fe-0.57 pct A1 specimen at 575°C. Most of the oxygen could have been picked up from the nitriding atmosphere on the surface of the samples during cooling to room temperature. Even 50 ppm of water in the gas phase will become oxidizing to iron before the sample has cooled to room temperature. The lack of reactivity* of these alloys in the annealed

Jan 1, 1970

GWILLIAM H. CLAMER (Philadelphia, Pa.).-About 16,000 tons of tin is used annually in the production of white metals. The real reason for using the so-called genuine babbitt, which is a high-tin base babbitt, perhaps in the majority of cases, is that tin has always been. a high-priced metal and people naturally believe that what they pay a high price for must be good. Mr. Jones has shown that the common metal, lead, is a superior base for most purposes for Which babbitt metal is used. The addition of lead is often looked upon as adulterant to genuine babbitt, and specifications for this reason rigidly restrict the lead content. A sub-committee of the Society for Testing Materials, at the meeting in June, presented tentative specifications on babbitt metal, in which were included twelve formulas, containing from zero to: 20, per cent. of tin, and several containing from 65 and 92 per cent. Babbitts. containing

Jan 1, 1919

By R. J. Lloyd

INTRODUCTION The Mount Isa Mine is a large modern underground mine located in North-Western Queensland, Australia. Two ore types are mined and treated separately. Currently silver-lead-zinc ore is mined at the rate of 2.6 million tonnes per annum and copper ore at the rate of 4.6 million tonnes per annum. The silver-lead-zinc orebodies are mined by sub level open stoping and cut and fill methods whilst all copper ore is extracted by various forms of sub level open stoping. The mining operations are highly mechanised and utilise diesel powered load-haul-dump units of various sizes, two and three boom face drilling jumbos, electric-hydraulic rotary-percussive ring drills, in-the-hole hammer drills, rotary blasthole drills, raiseborers, battery/trolley locomotives, a large fleet of rubber tyred service and supply vehicles plus light rockdrills mounted on airlegs, rail mounted overshot shovels and battery and diesel powered locomotives. All stoping production ore is extracted with diesel powered load-haul-dump units. The ore is tipped into orepasses which connect to the main rail haulage levels. On the haulage levels electric battery/trolley locomotives are used to pull rakes of bottom dump or side dump trucks which feed the ore into primary crushers. Following two stage crushing the ore is conveyed to the shaft and hoistod to surface in bottom dump skips. Each ore stream has its own rail transport system, crushing stations and hoisting shaft feeding separate concentrators. This paper will consider some aspects of the role of equipment and machinery in the growth of the mine, will discuss in some detail the machinery and equipment in use today and will conclude with a look at some possible developments in mining equipment for the future, with particular reference to the Mount Isa Mine. A summary of mining equipment in use at Mount Isa is given in the Appendix. THE PAST Silver-lead-zinc ore was discovered at Mount Isa in 1923 and within a very short time many small mines were being developed. However the rich surface ore traces did not continue and by 1928 Mount Isa Mines Limited had acquired sufficient leases to plan for a large scale mine to exploit the low grade orebodies. At this early stage it was recognised that the mine would need to adopt the most mechanised mining techniques then available in order to overcome its twin disadvantages of low graaes and extreme remoteness. The Company introduced bar mounted drifters and overshot shovels to replace hammer and tap and hand shovelling in development work and by 1930 headings were advancing at the rate of 30 metres per week. In keeping with the concept of mechanisation the early engineers designed the Man and Supply shaft with a very large, for those days, cage compartment measuring 3.8 m x 3.4 m. This shaft was the main means of access into the mine for 36 years until the K57 shaft (now R62 shaft) was commissioned in 1966. Man and Supply is still in regular use. Production commenced in 1931 at an annual rate of approximately 660 000 tonnes of silver-lead-zinc ore. The mine was mechanised to an extent which was previously unknown in Australia; whilst the concentrator, smelter and power house utilised the most modern and efficient equipment available. Even so, the mine was very near to closure on several occasions and did not return a dividend to its shareholders for many years. Production levels remained at or slightly below 660 000 tonnes until a separate copper ore streani and mill was commissioned in 1953, doubling the existing production levels. The first underground crushers were installed at this time. Production rates expanded in the years between 1953 and 1973. By 1960 production reached 2740 000 tonnes of ore. In 1965-66 total production was 3650 000 tonnes and in 1973 the production rate plateaued at the current level.

Jan 1, 1981

By Vinod K. Namilakonda

In situ synthesis of discontinuously reinforced aluminum alloy metal matrix composites was investigated. Aluminum alloy (A356: Al-7 wt% Si) was reinforced with dispersed AlN particles. Thermodynamic calculations were made for the composition determination of various phases formed during bubbling of NH3 gas into the Al alloy melts in the temperature range 1173-1473 K. Experiments were carried out for in situ production of AlN particles with the reaction between Al melt and NH3 gas in high temperature resistance furnace at 1473 K. The products were characterized using SEM, EDS and XRD. Effect of bubbling time of NH3 gas on the AlN formed in the composites was studied. The amount of AlN reinforcement in the Al alloy composites was observed to be in the range of 5 to 35 wt%. Good agreement was observed between the experimental and thermodynamic calculated data on the composition of AlN in Al alloy composites.

Jan 1, 2006

By F. B. Petermann

Two-stage electric smelting tests, using small submerged arc furnaces, were made on high-iron manganiferous materials from the Pioche District, Nev., and the Cuyuna Range, Minn. The object was to reduce sufficient iron in the first stage to give a slag with a favorable Mn to Fe ratio for smelting to ferromanganese in the second stage. The nodulized flotation concentrate from the Pioche District contained 29.2 percent Mn and had a Mn to Fe ratio of 2.3:l. When smelted with coke and silica or alumina as flux, more than 90 percent of the manganese was recovered in slags with Mn to Fe ratios of 10: :24:1. The Pioche slag was smelted to ferromanganese using coke and lime- stone or dolomite as flux. The ferromanganese was high in silicon but other- wise met specifications. Recovery of manganese to ferromanganese in this step was 73 to 78 percent. Overall recovery of manganese was about 70 percent. The sintered flotation concentrate from the Cuyuna Range material contained 12.9 percent Mn and had a Mn to Fe ratio of 0.3:1. When smelted with limestone as flux, about 83 percent of the manganese was recovered in a slag with a Mn to Fe ratio of about 6:1. The Cuyuna slag was smelted to ferromanganese with coke and limestone for flux. Recovery in this step was 61 percent of the manganese in a high silicon ferromanganese. Overall recovery of manganese was about 50 percent.

Jan 1, 1963

By David Berberick

"Hecla Limited’s No. 4 Shaft at the Lucky Friday Mine is the largest capital expenditure in the corporation’s 128-year history and is considered the deepest shaft in the United States. This paper will discuss the project’s history and lessons learned from the initial conceptual design work in 2007 through project commissioning in 2017. Change management was key for the owner, contractors, and vendors in delivering the operational and organizational success on the project. The No. 4 Shaft has extended the life of the Lucky Friday Mine by 21 years to 2039. INTRODUCTION The Lucky Friday Mine of Hecla Mining Company (Figure 1) is the oldest active (77 years) and deepest mine in the Coeur d’Alene mining district, producing approximately 725 tonnes (800 tons) per day from narrow, vertical, silver, lead, and zinc veins. Primary access into the mine is via the Silver Shaft, a 5.5-m (18-ft) diameter, concrete lined shaft sunk to an original depth of 1,890-m (6,200-ft). Current production is from the Gold Hunter vein system, approximately 1,525-m (5,000-ft) northwest of the Silver Shaft.Cut and fill stoping is currently employed at depths to 2.3 km (7,500 ft) below ground surface, making it one of the deepest mines in the western hemisphere. No. 4 Shaft is a 5.5-metre (18-foot) diameter, concrete lined circular shaft, constructed as a 1.2 km (3,860 feet) deep winze in the Lucky Friday Mine. The 2.4 m (8 ft.) diameter, double drum, double clutch, Siemag hoist and associated headworks are situated at the 4760 level, top station at 4940 and shaft stations at 5900, 6500, 6580, 7500, 7580, 8300, 8380, 8500, and 8620. Shaft stations reference depth below Silver Shaft top station. The 4 Shaft was designed, constructed and commissioned over ten years, between 2007 and 2017. Basic design work was completed by Nordmin Engineering, followed by detailed design completed by Cementation Canada. Shaft construction was awarded to Cementation USA and lateral development was initiated in early 2008. Over the course of the next nine years, the project was suspended and restarted twice, dealt with the loss of a miner, added scope including shaft depth and mine infrastructure facilities, experienced squeezing ground that required liner replacement, changed excavation section, and dealt with a final alignment that was not within design requirement for verticality. Hecla and Cementation were successful in delivering a facility that provides fully functional access from collar at 4940 level to shaft bottom at the 8620 level. Two active mining levels below the collar, 5900 and 6500 are accessed. In addition, the shaft provides access to the future mining level at 7500, skip loading facility at 7580 level, station and loading pocket cutouts at 8300 and 8380 levels, spill pocket at 8500 level, and shaft bottom pumping facility at 8620 level (Figure 2)."

Jan 1, 2019

By Robert I. Jaffee, Bruce W. Gonser, Eugene M. Smith

There has been little investigation of the gold-germanium binary alloys. Haughtonl does not mention any work on the system. In a review article commemorating the fiftieth anniversary of the discovery of germanium, E. Einecke2 mentioned a gold-germanium eutectic at 24 atomic per cent Ge (10.4 weight per cent) melting at 359°C., but did not give the source of his information. From measurements of electrical conductivity, J. 0. Linde3 estimated that at least 0.6 weight per cent germanium is soluble in gold. The recent availability of germanium in this country in commercial quantities has led to interest in its properties and its alloys. This study is part of a program of investigation on germanium alloys; work on ternary alloys involving germanium and gold will be reported later. In this investigation, the composition and melting point of the Au-Ge eutectic reported by Einecke has been found to be substantially correct. Eutectic composition has been placed at I2.0 per cent germanium, however. An alloy of 10.4 per cent Ge, Einecke&apos;s value, was found to have a few gold dendrites in it. The cutectic temperature has been determined as 356.0 6 ± 0.5°C., slightly less than the temperature of 359°C. reported by Einecke. The liquidus curves and the solid solubility of ger- manium in gold have been determined, and are shown in Fig. I. Materials and Alloying The germanium used in this investigation was prepared by the Eagle-Picher Lead Co. from pure germanium oxide, by reduction in clay crucibles with sodium cyanide. The regulus is quite pure, assaying over 99.9 per cent Ge. High-purity gold obtained from Handy and Harmon was used. Alloying was carried out in graphite crucibles in an atmosphere of purified hydrogen. The temperature was brought up to 1050" to iioo°C., and the metal was allowed to freeze in the crucible. Prepared in this way, the gold dendrites in hypoeutectic alloys were somewhat segregated at the bottom and sides of the ingot, and on the hyper-eutectic side the lighter primary germanium tended to gather at the top. Most of the alloys were made from virgin metal. Low-germanium heats were made by alloying pure gold with the eutectic. Melting losses were consistently a few milligrams or less, and the composition could safely be taken from the intended analysis. The 30 per cent and 50 per cent germanium heats for thermal analysis were made in part from remelted higher germanium heats, and they may be OR composition as much as one per cent. Because of the high cost of materials, the weight of heats for thermal analysis were 15 to 20 grams and heats for metallo-graphic, X-ray, and hardness specimens were 2 to 4 grams.

Jan 1, 1945

By Jan M. Peter

The Early Norian Windy Craggy massive sulfide deposit is within the allochthonous Alexander terrane of the Insular tectonic belt in extreme northwestern British Columbia (Figure 1). Host rocks are the Middle Tats Volcanics, part of an informally-named volcano-sedimentary succession of mixed graphitic argillites and mafic pillowed and massive volcanic flows and sills of Upper Triassic age (Figure 2) that have suffered only lower greeschist-facies metamorphism. The volcanic rocks are light rare earth element (LREE)-enriched and display a variably developed back-arc subduction signature. Major and trace element compositions of the host argillite show that it contains an appreciable hydrothermal component, as evidenced by low high Fe and Mn contents. These sediments therefore record steady, continuous hydrothermal activity punctuated by intermittent turbidity currents sweeping into the basin. The deposit consists of at least two discrete sulfide lenses (the North and South Sulfide Bodies), as shown in Figure 3, a level plan at the elevation of the exploration workings. Published reserves at 0.5% copper cut-off are 297.4 million tonnes grading 1.38% Cu, 0.08% Co, 0.21 g/t Au and 3.9 g/t Ag (as of December 1991). However, the deposit is larger than this and further drilling is required to delineate it completely. Mineralization comprises massive sulfide, stringer/stockwork, and finely bedded to laminated sulfides and exhalites. Stockwork mineralization consists of sulfide-quartz-chlorite veins with attendant chlorite and silica alteration. Mineralization consists predominantly of massive pyrrhotite andlor pyrite with lesser chalcopyrite and magnetite. Figure 4 is an oblique section through the North Sulfide Body, the most northerly sulfide mass shown in Figure 3, which shows the distribution of the pyrite and pynhotite-rich massive mineralization and the stringer zone. The North Sulfide Body contains appreciable sphalerite, whereas the South Sulfide Body contains only trace amounts. Gangue minerals include quartz, chlorite, calcite, siderite, ankerite, stilpnomelane and magnetite. The minor hydrothermal exhalite sediments are comprised of chert, carbonate, and minor

Jan 1, 1993

By Francisco Belmino Romero, Walney Silva Araújo, Vitória Lima de Souza, Edilberto Rocha Silveira, Adriana de Oliveira Sousa Leite

O uso de inibidores verdes na corrosão do aço carbono SAE J403 1020 configura-se como uma alternativa econômica e ambientalmente acessível, pois, em relação aos inibidores comerciais geralmente utilizados, são de fácil obtenção, já que são adquiridos a partir de vegetais, portanto, não oferecem risco ao meio ambiente. Assim, o presente trabalho tem como objetivo avaliar o uso das folhas de aroeira como inibidor de corrosão para o aço carbono SAE J403 1020, o qual é bastante utilizado na indústria e possui baixa resistência à corrosão. A escolha da aroeira se deu devido ao fato de a planta ter origem no Brasil e abundância no Nordeste do País, principalmente no Ceará. Para análise, foram utilizados como meio corrosivo: H2SO4 0,1 M, NaCl 3,5% e NaCl (ph = 4,0); as concentrações do inibidor foram: 50, 200, 500 e 1000 ppm. Em relação aos ensaios eletroquímicos, utilizou-se: Potencial de Circuito Aberto (OCP), Espectroscopia de Impedância Eletroquímica (EIE) e Curvas de Polarização Linear (CP). Dessa forma, por meio da análise das técnicas mencionadas, conseguiu-se obter o comportamento do inibidor em relação ao material e ao meio, bem como avaliar qual o melhor meio para a utilização da aroeira como inibidor verde de corrosão.

Mar 1, 2019

By Maurizio Massimo

Danieli developed the “3Q” suite of process control technologies to address three critical principles: delivering the highest product Quality, in the required Quantity, with Quickness of response required for today’s casting machines. Ever increasing market demands for productivity and slab quality necessitate the latest advances/improvements in equipment design, process technology, and control software for all types of slab casters, and Danieli boasts a unique experience that extends from ultra high-speed thin slab casters to extra high-thickness conventional slab casters. The “3Q” system is monitored and controlled by a unique, intelligent HMI incorporating. Touch screen technology and Operator Assistant tablets. Operator Assistant is a tablet based system that guides the operator through the plant process, so that in the event of a process change or an alarm the operator is automatically presented with the choice of possible solutions / actions. The comprehensive range of technological packages cover all aspects of caster design and automation control, from feeding of liquid steel into the tundish to the final slab leaving the caster.

Aug 17, 2017

By J. P. Nielsen, H. Margolin, E. Ence

The Ti-Ni phase diagram has been investigated up to 68 pct Ni with iodide titanium base alloys by metallographic, X-ray, and melting point methods, and from 68 to 90 pct Ni by examination of as-cast structures of sponge titanium base alloys. NVESTIGATION of the nickel-rich portion of the I Ti-Ni phase diagram was first reported by Vogel and Wallbaum in 1938.&apos; This work was subsequently extended to lower nickel contents by Wallbaum&apos; who indicated the possibility of a eutectic reaction for nickel contents below 38 pct. Long et al.3 studied the titanium-rich portion of the phase diagram and found eutectic and eutectoid reactions below 38 pct Ni. However, the temperature of the eutectic indicated by Long et al. was considerably lower than that suggested by Wallbaum. Long and his coworkers synthesized their alloys by powder metallurgical techniques and encountered oxygen and/or nitrogen contamination. Thus the diagram which was obtained did not represent binary alloying conditions. However from these results the features of the binary diagram were predicted. At Battelle Memorial Institute4 the Ti-Ni diagram was investigated up to approximately 11.5 pct Ni with sponge titanium alloys. The range of temperatures used was not sufficient to define the eutectoid temperature or composition. The data of Wallbaum2 and Long et al.8 were of particular interest for the present study, and although the work was originally concerned with the region below 40 pct Ni, the investigation was extended to higher nickel contents in an attempt to resolve the differences between these workers. Experimental Procedure Preliminary work on the Ti-Ni system was carried out with duPont Process A sponge titanium alloys to reduce the amount of subsequent work to be done with iodide titanium base alloys. The sponge titanium used contained 99.71 to 99.77 pct Ti, 0.1 pct Fe and 0.005 to 0.009 pct Ni. The iodide titanium obtained from the New Jersey Zinc Co. contained 99.9 to 99.95 pct Ti. Nickel used with sponge titanium was 98.9 pct pure. The high-purity nickel alloyed with iodide titanium was cobalt-free with approximately 0.05 pct C and was obtained through the courtesy of the International Nickel Co. The 15 g sponge titanium charges for melting were prepared by compacting in a die or by placing the weighed portions of nickel and titanium directly into the furnace. Iodide titanium charges were made by drilling holes in the as-received rod and inserting the nickel or by wrapping the nickel in sheet. Sponge titanium alloys containing from 0.2 to 90 pct Ni and iodide titanium alloys containing 0.2 to 68 pct Ni were prepared by these methods. In addition to these alloys several 1/2 1b sponge titanium alloys were supplied by the Allegheny Ludlum Co. The charges were melted in an arc furnace under an argon atmosphere. The procedures used were similar to those reported in the literature5,&apos; and the furnace has been described.&apos; Except for iodide titanium alloys with 40 to 68 pct Ni (see section on copper contamination), each alloy was melted for 1 min, then either turned over or broken before re-melting for an additional minute. Currents of 200 to 400 amp were used depending on the melting point of the alloy. Prior to heat treatment, alloys containing less than 14.5 pct Ni were hot-forged at 750°C. With the exception of alloys in the homogeneity range of the compound TiNi, alloys of higher nickel contents could not be hot-forged. Heat treatment of iodide titanium base alloys was carried out in argon-filled quartz capsules which were broken under water at the conclusion of heat treatment to quench the specimens. Temperatures were controlled to ±5oC and annealing times up to 48 hr were used. For melting point determination, specimens were placed in carbon crucibles which were in turn en-capsuled in argon-filled quartz capsules. The start of melting was determined by rounding of corners and by metallographic examination. Complete melting was considered to have occurred at that temperature at which the specimen assumed the shape of the crucible. Specimens were prepared for metallographic examination by mechanical polishing or by an electrolytic procedure." For alloys containing up to 80 pct Ni Remington A etch7 50 pct glycerine, 25 pct HNO,, 25 8 HF) was used. For higher nickel alloys aqua regia and Carapella&apos;s etch (5 g FeCl,, 2 ml HNO,, and 99 ml methyl alcohol) were employed. Specimens to be exposed for powder patterns were prepared by filing, by breaking specimens in a

Jan 1, 1954

By A. J. Duncan, Bath M. D., E. R. Rudolph

Some factors influencing gold recovery by gravity concentration by M. D. BATH,. M.A.Se. (Met.) British Columbia (Visitor) A. J. DUNCAN,. B.Sc. (Hons.) Rhodes, M.S.A.C.I. (Visitor) E. R. RUDOLPH,* M.Se. (Eng.) Rand (Member) Recent studies have shown that the overall extraction of gold from Witwatersrand ores is increased as the proportion of gold reporting to the amalgamation circuit is increased. Testwork has therefore been carried out with a view to improving recovery by gravity concentration. In this paper, the history of gold recovery by gravity concentration is reviewed, and methods and machines at present in use are briefly described. The results of plant and laboratory tests are presented, and the limitation of existing concentrators is discussed. A pilot plant is being built for the testing of improvements to the existing concentrators and of several new devices, particularly the Reichert-cone concentrator.

By R. D. Haynes

Coal, gas, and oil in many geological formations of the Eastern United States are intimately associated but have traditionally been recovered independently of each other. Oil and gas wells, however, are often spaced sufficiently close to cause the subsequent loss of up to 50 percent of the recoverable coal in the form of pillars or barriers left for mine safety purposes. A long-range planning technique was devised for exploiting oil and gas wells simultaneously with the mining of coal to recover the maximum fuel potential of the formations. Planned plugging of oil and gas wells, determined partly from production predictions via decline curve analysis, is coordinated with a mine development plan to achieve maximum coal recovery and reduce overall production losses of gas and oil due to plugging of the wells.

Jan 1, 1973

"Hematite pellets were electro-deoxidized to iron in molten sodium hydroxide to produce iron containing about 10-wt % oxygen. The anodic reaction was the evolution of oxygen on an inert nickel anode. Before the electrochemical reaction, there is a reaction between sodium hydroxide and ferric oxide to form sodium ferrite and, in order to compensate for the reduction in sodium oxide content, 1 wt % sodium oxide is added to the melt. In a laboratory cell, the cell voltage was 1.7 V with a current density of 5000 A m-2, a current efficiency of 90% and an energy consumption of 2.8 kWh kg-1.Introduction The iron and steel industry contributes about 2.3 bn tonnes/annum of the carbon dioxide to the atmosphere, which is about 10% of the total, and some believe that, in order to prevent a catastrophic rise in the earth’s temperature, carbon dioxide emissions must be reduced by 50% come 2050. However, by 2050, it is likely that the iron and steel industry would have grown significantly so that the production may have risen to 2 or 3 bn tonnes/year so in order to meet the 50 % reduction on today’s figure, the carbon dioxide may have to be reduced to 25% or less, which is a very demanding requirement. There are probably three ways of reducing the carbon dioxide output:1. By capture and sequestration, but this may be technically very demanding.2. Reduction using hydrogen in place of carbon."

Jan 1, 2009

The U.S. Bureau of Mines, Department of Labor, and United States National Committee of the CIE hosted the TC4.10 Mine Lighting Committee Meeting and Second International CIE Mine Lighting conference at the MSHA Academy in Beckley, WV from October 12-16, 1981. This was the first time the meeting and conference had been held in the United States. The TC4.10 Mine Lighting Committee, chaired by Dr. Adam Peretiatkowicz of Poland, was attended by 15 committee members and six guests representing eight countries. Discussion during the meeting covered mine lighting instrumentation, open cast (surface mine) lighting, and lighting sources. The committee also reviewed a draft of proposed lighting terminology and suggested changes to make the definitions more acceptable to representatives of the nations present. The mine-lighting conference featured 20 papers presented by 18 speakers representing six countries. The papers covered all phases of mine lighting, including research, hardware development and application, personnel acceptance, and requirements for system maintenance. The conference attracted 126 attendees and also included displays by eight lighting hardware manufacturers and four mining machine manufacturers. Papers from the conference are being published by the Government Printing office, with copies presented to the conference attendees. Copies will be available to others interested through the NTIS. During the course of the TC4.10 meetings and conference, committee members toured three mining equipment assembly plants, and six of the foreign representatives visited a longwall installation in an underground mine. Planning and administrative services for the meeting and conference were handled by Bituminous Coal Research, Inc., under USBM Contract H0318038.

Jan 1, 1982

By N Stockton

Some preliminary results on the secondary leaching of previously heap leached gold ore are presented. Alternative lixiviants for gold comprising chlorine, thiourea and thiosulphate were compared with cyanide using bottle roll tests. Chlorine was subsequently selected for column leaching tests. Column tests at 1.0, 0.1 and 0.01 gL-1 Cl2 at pH 2 were conducted. The results indicated that about 23 per cent of the gold was leached over 45 days using 1 gL-1 Cl2 at pH 2. The experiments conducted proved that the chlorine/chloride system has good potential for further extraction of gold and silver from the existing cyanide heap leach residues and a process for secondary leaching of gold has been proposed. Further testwork is recommended to extend conditions, sample types, and the duration of leaching period.

Jan 1, 2000

By S. X. Shi, M. Tan, Y. Q. Li, D. F. Han, Z. B. Han

Bauxite resources in China are mainly diaspore ore with the characteristics of low ratio silicon to aluminum. In order to obtain high quality of alumina product, flotation process is generally used to extract the aluminum from diaspore ore. Small aeration rate is needed for bauxite flotation, which requires air to be controlled accuracy. In addition, productive froth from bauxite flotation is too sticky to effects froth transportation. Air-forced XCF/KYF flotation machines are helpful for aeration control, while self-suction GF flotation machines benefits froth transportation. Zhongzhou aluminum plant of China Aluminum Group (Chinalco) uses 7 series of XCF/KYF /GF combined flotation machines to process bauxite. Air distribution, suspension characteristics, froth flow velocity and Al2O3 recovery were investigated, to evaluate the performance of flotation machines in one series and determine recovery contribution of each cell to the entire flotation circuit.The air distribution degree varies from 3 to 4, which is enough for bauxite floatation process. Froth in the tank could flow smoothly with a flow velocity of 2.4m/min. The recovery of the first KYF-40 flotation cell is the largest. Flotation concentrate of Al2O3 is up to 64.29%, the total recovery is about 75%, and the ratio of Al/Si has reached 10.4.

Jan 1, 2015

By Thomas J. Peters

In 1987, at the request of the U.S. Bureau of Land Management, the U.S. Bureau of Mines studied the 15,174-acre Sleeping Giant study area, including the 6,112-acre Wilderness Study Area (MT-075-111), in order to evaluate its identified mineral resources. The study area is located in Lewis and Clark County, MT, about 30 miles north of Helena. Sleeping Giant Mountain trends west-northwest from the Missouri to the eastern crest of Little Prickly Pear Canyon (the route of Interstate 15), and has several summits over 6,000 ft. Geologically, the area is in the Disturbed Belt, and a large block of Middle Precambrian metasedimentary siltstone and underlying shale has been thrust northeast over younger, sedimentdry volcanic and intrusive rocks.

Jan 1, 1988