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By Richard D. Miller

The helium resources base of the United States was estimated by the Bureau of Mines to be 1,040 Bcf as of January 1, 1987. These resources are divided into four categories in decreasing degree of assurance of their existence: (1) helium in storage and in proved natural gas reserves, 265 Bcf, (2) helium in probable natural gas resources, estimated at 228 Bcf, (3) helium in possible natural gas resources, estimated to be 320 Bcf, and (4) helium in speculative natural gas resources, 227 Bcf. These helium resources are further divided into depleting and nondepleting, with the helium in storage being in a separate classification. The depleting resources are those associated with natural gasfields that are, or will be, produced for the natural gas they contain. Almost all of the helium in potential (probable, possible, and speculative) natural gas resources is included in this classification. These depleting resources are estimated to contain 951 Bcf of the total helium resource base. Helium resources contained in nondepleting natural gas resources, i.e., those in shut-in, abandoned, or otherwise nonproducing natural gasfields, are estimated to total 53 Bef. There is 36 Bcf of helium in storage in the Federal Government-owned Cliffside Gasfield near Amarillo, TX.

Jan 1, 1988

By Dick Whittington, W. R. Taylor

GENERAL In mid-1974, a representative of ARCO Petroleum Company walked into our Austin offices and asked for a permit for ARCO's first Clay West Mine site. This was the first in situ uranium mining permit issued in the United States. That initial permit, issued in January 1975, was for three acres. The four years since have been characterized by growth and change. Today there are 16 permits approved for approximately 1,540 production acres issued to five companies at eight sites in four counties. There are approximately seven additional permit applications being processed or on the verge of being submitted for an estimated 2,000 additional acres of production. As a rough gauge of the financial scope of this activity, at the typical 40 production wells per acre and an estimated $2,000/well, operators hale installed or committed $122 million for production wells. This does not account for the exploration and development wells installed to delineate the ore or the cost of processing facilities for extracting the uranium. HISTORY OF URANIUM MINING IN TEXAS Information I have concerning early in situ uranium mining activities is "sketchy" but the apparent first on-site test of in situ uranium mining was conducted between 1967 and 1971 at the Dueterstat site, north of Yorktown in Gonzales County, by City Service. Others ran tests at the Bruni site in Webb County and the Burns and Clay West sites in Live Oak County in the mid-70's. During this period, Texas Water Quality Board personnel in the central and district offices maintained a general surveillance of these various test activities. Regulation of in situ uranium mining was formalized with the issuance of a permit in January 1975 with the advent of ARCO's initial production scale (3 acre) effort in the Clay West area at the Moser mine site. Since 1975, the accelerating rate of mining and exploration activity is noticeable. There were 16 exploratory rigs operating early in 1977. This year, the rig count is averaging 46--a 287% increase. HISTORY OF LEGISLATION Paralleling this mining activity was state and federal legislation which subjected in situ mining to environmental regulation. The following is a brief chronology of the legislation that is directly or indirectly related to uranium mining. As most of you know, disposal wells are an integral part of most in situ uranium mines operating in Texas. They are used both during production and during aquifer restoration following production. The wells are regulated by the Texas Injection Well Act, enacted in 1961 by the Texas Legislature. In 1967, the Water Quality Act established the Texas Water Quality Board, and provided for the regulation of water quality by the so-called permit system. In 1969, the Solid Waste Act was enacted by the Texas Legislature. The legislation provides for control of wastes from manufacturing, mining, and agriculture. This Act regulated uranium surface mining activities until 1975 when the Texas Surface and Mining Reclamation Act placed responsibility with the Railroad Commission for uranium open-pit mining and surface mining of coal and lignite. In 1972, the Federal Water Pollution Control Act (Public Law 92-500) was passed. This legislation forms the legal base for the NPDES (National Pollutant Discharge Elimination System) and the subsequent point source effluent guidelines and standards. The November 6, 1975 issue of the Federal Register published interim final rules for Best Practical Treatment levels for effluent discharges from ore mining and dressing operations. Litigation was immediately initiated by eleven companies against these proposed rules, and they are currently being reconsidered for reissuance in 1979. In 1974, the Federal Safe Drinking Water Act (Public Law 93-523) was enacted. A part of this legislation is the Underground Injection Congrol Program. The regulations implementing this program have undergone various public hearings and it is expected that they will be published in the Federal Register before the end of 1978. In 1975, the Texas Surface Mining and Reclamation Act (Senate Bill 55), providing for the Railroad Commission's authority to require reclamation of surface mines, was passed. In 1977, an amendment to Senate Bill 55 was introduced that proposed to place regulation of in situ mining of coal, lignite, and

Jan 1, 1979

By C. G. Hunkin

In situ, or in-place leaching is concerned with the extraction of metals from natural mineral disseminations without removal from their geologic location which, in many cases, amounts to the reversal of the deposition process accompanied by a presumed increase in velocity of the processes, if the project is to be profitable. Solution mining processes involve removal of minerals such as sulfur, common salt, and potash, generally leaving cavernous voids, in contrast to in situ leaching which term is applied to the selective dissolution of a mineral from its gangue without sensible disturbance of the formation. Where such a process can be operated, the advantages are manifold and have been so recognized by miners from the earliest days of mining. Recent technological developments have encouraged a reassessment of the status of in situ leaching, and it now appears possible to extend the application of the method from a few isolated special situations to many uranium and copper deposits. Anaconda's interest in the techniques stems from long experience with vat and heap leaching, 'dump leaching, and natural leach activities associated with several established mines. Increasing activity through-out the industry has been noted, particularly in leaching low grade pillars and the peripheral zones of operating mines, often by the introduction of leach solutions via boreholes drilled from surface. Reactivation of dormant mines] using similar techniques has been successful. Over the long term, environmental improvement activity will lead to excess sulfuric acid production, even to a disposal problem. The cost of leach solutions and reagents must be reexamined in this light. The mineral industry now classifies as uneconomic deposits of readily soluole minerals whose potential value, on a per acre basis, may De several times that of many currently profitable oil field reser-

Jan 1, 1971

By Richard L. Stein

INTRODUCTION Increasing a miner's chance of surviving a disaster requires advance planning for emergencies, adequate training of personnel, and the provision of proper survival equipment. It is extremely important that fires and toxic gases be detected early, and that warning systems such as stench, audible and visual alarms, and communications systems be provided to allow timely escape. However, situations do occur where timely warning is not provided and immediate escape is not possible, regardless of the precautions that are taken. Under such circumstances, a combination of new per¬sonal protective devices and old techniques can provide protection that is adequate to save a miner's life. SELF-RESCUE EQUIPMENT Self-rescuers are devices which provide a short-term supply of respirable air under the potentially lethal con¬ditions following a mine disaster. Functioning either as filtration devices for the elimination of carbon monoxide (CO) or as oxygen supplies, self-rescuers allow an endangered miner a limited amount of time in which to effect an escape or reach a place of temporary safety. Carbon-Monoxide Self-Rescuers For almost 50 years, filtration self-rescuers have protected miners' lives following explosions or fires. Designed to eliminate carbon monoxide from inhaled air, the first of these devices was approved in the 1920s and has evolved into the two units that are approved today-the Mine Safety Appliance (MSA) W-65 and the Draeger 810. Both of these units rely on the cata¬lytic conversion of very toxic carbon monoxide to rela¬tively safe levels of carbon dioxide (CO,). As shown in Figs. 1 and 2, both of these filtration self-rescuers operate in the same manner. On inhalation, the mine air passes through both coarse and fine filters that remove the dust, preventing the dust from coating the chemical beds or entering the miner's mouth. The air then passes over a drying agent that removes water vapor; this is required to prevent the water vapor from poisoning the catalyst. Subsequently, the dried air passes over the Hopcalite catalyst, where carbon monoxide is converted to carbon dioxide. Then, the air passes through a heat exchanger that reduces the temperature of the inhaled air. Finally, the air is inhaled by the user. Air exhaled by the user passes through the heat ex¬changer and exits the device through an expiratory valve. Both units of this type are subject to certain limita¬tions: 1) The units do not protect the user against oxygen¬deficient air; when air is inhaled through this type of self-rescuer, the device only removes the carbon mon¬oxide. If the mine air contains less than 15% oxygen, anoxia is inevitable. Symptoms of anoxia include dizzi¬ness, shortness of breath, quickened pulse, and deeper and more rapid respiration while the victim is at rest. During heavy exertion such as would be expected dur¬ ing escape efforts, a 15% oxygen level can cause loss of consciousness. 2) The units do not protect against excessive levels of carbon dioxide. Available data on carbon monoxide and carbon dioxide concentrations following explosions or fires are limited, but there are indications that the carbon monoxide concentration can rise to 2% and the carbon dioxide concentration can reach 5 or 6% of the mine air (by volume). In some situations, the in¬haled air could contain up to 6 or 7% carbon dioxide. At a carbon dioxide concentration higher than 2%, breathing patterns can be affected adversely. At a con¬centration of 6 or 7%, the effects include severe respira¬tory distress, with unconsciousness resulting from exer¬tion such as that needed for escape activities. 3) The units do not protect against high inhalation temperatures. The catalytic oxidation of carbon mon¬oxide to carbon dioxide is an exothermic reaction that evolves a large amount of heat. As the carbon monoxide concentration increases, the temperature of the inhaled air also increases. Available data show that above 1.5% carbon monoxide, the air inhaled through some self¬rescuers can be as high as 90°C (194°F). At those

Jan 1, 1982

By W. K. Connor, O&apos

The Bureau of Mines is investigating the Co and Mn resource of the West-Central Arkansas District, including field investigations, mineralogical characterization, beneficiation, and hydrometallurgical tests. The deposits occur within the Ouachita Mountains and are con-trolled structurally, occurring along the axes of anticlines making up the mountain ridges. Brecciated novaculite is cemented with a mixture of Mn oxides, primarily cryptomelane and psilome-lane. Cobalt is concentrated within a particular Mn oxide phase, lithiophorite. Chemical analyses on head samples reported up to 25 pct Mn and 0.17 pct Co. Magnetic separation yielded products with up to 41 pct Mn and 0.22 pct Co at recoveries of 95 pct Mn and 93 pct Co.

Jan 1, 1991

By Chonghun Han, Tae-Hwa Choi, Young-Hwan Chu

"A statistical optimization method is proposed to maintain stable operation in a blast furnace with stave coolers. The objective of this study is to find the operating condition which maximizes productivity as well as minimizes the heat load on the stave coolers. In this method, we first analyzed historical data and calculated the heat flux of stave coolers and the burden distribution by mathematical models. In this approach, we determined the variables to optimize and manipulated variables to adjust these response variables. Next, we collected the necessary data for the data variables after preprocessing such as removal of noise and outlier. Finally, we constructed empirical models describing the patterns of response variables in terms of manipulated variables by partial least squares regression method, and then these models were used as objective functions for entire optimization problem. By using appropriate optimization algorithm, we determined the compromised optimum operating condition considering all response variables from the past operating conditions.IntroductionBlast furnace operation which is accompanied by complex physical and chemical phenomena is an essential process for steel industry. In this process, iron ore which is charged into the blast furnace from the top goes downward. It is reduced to iron by CO gas and is melted by the heat of combustion of coke and pulverized coal. The pig iron is tapped at the bottom in liquid phase after removing the slag, which floats on the pig iron.The cooling equipment of the blast furnace wall is changed from the plate type to the stave cooler recently. In order to cool the shell uniformly and eliminate shell spray cooling, Pohang No.2 blast furnace of POSCO (Pohang Iron & Steel Company) was equipped with 11 rows of stave coolers. These staves are only 350mm-thick including the 170 mm-deep refractory on the hot face."

Jan 1, 2001

By Wiraseranee. Chompunoot, Kazuki Morita, Toru H. Okabe

"Aiming at the development of more effective recycling processes, dissolution behavior of rhodium into molten slag was investigated. Using a pure rhodium crucible and Na2O-SiO2 slag system, equilibration experiments were carried out for 18 h at 1423-1623 K under various oxygen partial pressures of 0.005-1 atm with varied slag basicity. Rhodium solubility increased with increasing oxygen potential, slag basicity, and temperature. Rhodium dissolved into slag as rhodate ion ([RhO2]-) by the endothermic reaction: Rh(s) + 3/4O2(g) + 1/2O2-(in slag) = [RhO2]-(in slag). In the dissolution reaction, the “rhodate capacity” of slag is newly-defined. The correlation between rhodate capacity and optical basicity is reported.IntroductionDue to its high chemical resistance and outstanding catalytic properties, rhodium (Rh), one of the platinum group metals (PGMs), is utilized in various applications such as catalysts, molten glass vessels and containers, and electronic devices. Owing to the rarity of Rh resources in nature, recycled Rh from scraps such as spent automobile catalytic converters is currently considered as an important source of Rh supply [1]. According to the recycling practices, Rh is currently mainly recovered from scraps along with other PGMs such as platinum (Pt) and palladium (Pd) by a dry process (e.g., smelting), a wet process (e.g., leaching and subsequent separation of PGMs from aqueous solution), or a combination of both processes [2]-[4]. Among these processes, smelting is commonly applied because PGMs can be effectively separated from a relatively large amount of various impurities coexisting in scraps. In the smelting process, Rh and other PGMs are collected in the collector metal (e.g., iron and copper), and impurities in scraps are removed into slag [4]-[6]. The process may be carried out at a temperature range of 1573-1973 K, in either an oxidizing or reducing atmosphere. Due to the high oxidation resistance at high temperatures, Rh is expected to be collected efficiently in the metal collector rather than being oxidized and forming RhOx in a molten slag. However, Rh loss into slag does occur in this process [7]."

Jan 1, 2011

By Richard Kunter

INTRODUCTION Additional work has been done on the CIP circuit at Creede, and a brief description of this work is presented in this paper. DREDGE The original dredge for the Bulldog was built by Al Hack, Inc. of Denver. The dredge cutter head is mounted on a boom made of 6-inch diameter pipe which is raised and lowered by a cable-pulley system. Maximum cutting depth is 8 feet. Slurry from the dredge flows through a floating 6-inch diameter pipe supported on 55-gallon drums which runs from the dredge to the shore and then to a trash screen and pumping system. See Figures 1 and 2. The cutter head, suction pump, winches for dredge movement, and all other operations are controlled from a small operations house located toward the rear of the dredge. All of the basic dredge equipment is mounted on a chassis about 8 ft x 16 ft in size. Additional buoyancy was originally provided by sealed 55-gallon drums supported along the sides of the chassis by a small wooden framework. Additional framework and barrels on the front of the chassis on either side of the cutter boom provided additional stability, particularly when digging in hard material. The dredge was quite stable, but two problems were apparent. For one, it could have been sunk by someone from the shore firing a rifle into all of the barrels. In prime deer-hunting and sportsman country

Jan 1, 1983

By Frédéric Flament, Luc Lachance

"The important task of metallurgical accounting that used to be done manually is being more and more automated. Currently available solutions can generate accounting reports without human intervention or almost. However, the practical conditions for fully delivering the promised benefits of those solutions are sometimes overlooked either by end-users or by integrators. The mere fact of filling reports with values satisfying several constraints such as mass conservation constraints does not, by itself, ensure a high quality metallurgical balance. Since managers make decisions based on these reports, it is logical that the metallurgical balance performance be monitored and retrofits be made if necessary. Such retrofits may involve modifications to the measurement strategy or the reduction of the spatial resolution of the material balance model. This requires an estimability and redundancy analysis (ERA) tool to assess the feasibility of the modified model. Since determining by hand the estimability and redundancy of a set of non measured and measured variables is not trivial, even for a medium-sized mineral processing plant, a dedicated ERA tool brings valuable assistance to the metallurgist in charge. The present paper aims first at explaining how a state of the art metallurgical balance is a key enabler for the adoption of a smart process plants paradigm (SPPP) in mineral processing. Then, the fundamentals behind the building of an ERA algorithm are introduced and illustrated by two examples. Finally, the role of an ERA tool in achieving state of the art metallurgical accounting and how it contributes to delivering additional value are discussed.INTRODUCTIONThe Smart Process Plants Paradigm (SPPP)Mineral processors have seen an increase lately in external pressures toward operating performance improvements. Indeed, the globalization of the mining business is a key driver to becoming more productive, more responsive to market changes, to improving corporate governance and reporting standards. Furthermore, mineral processors are also expected by their neighbouring communities to continuously improve their sustainability score. These current pressures differ somewhat from the past ones in leading to a larger than ever number of conflicting constraints that plant managers must concurrently satisfy while maximizing plant profitability. It is in turn not surprising that this context requires actions involving more real-time collaborative actions between more people in more places than ever before."

Jan 1, 2011

By Elli Nurminen

Sulphate decomposition was studied in the conditions of a typical copper flash smelting heat recovery boiler. The aim was to determine the kinetics of the copper sulphate decomposition reactions and to find out if they are likely to take place in the process conditions prevailing in an industrial HRB. The information is essential in CFD-modelling of the reactions taking place in the boiler and for optimising the process. Thermal decomposition of sulphates was studied by thermogravimetry. The experiments were conducted with a commercial copper sulphate, a partially sulphated copper oxide and an actual process flue dust. The main variables studied were temperature and gas composition of a N2-O2-SO2-mixture. The samples were characterized with SEM, optical microscope and XRD. Temperature affected the decomposition rate strongly but the gas composition had a minor effect on the kinetics. The decomposition was fastest in the partly sulphated oxide and slowest in commercial copper sulphate. Keywords: Sulphate decomposition, flash smelting, heat recovery boiler, flue dust

Jan 1, 2004

By E. M. Stiles

THE Wellington Lake power plant of the Consolidated Mining and Smelting Company of Canada, Limited, about 25 miles from Goldfields, Saskatchewan, is the second hydro-electric project operating in that Province, the first being the Island Falls plant of the Churchill River Power Company which serves the operations of the Hudson Bay Mining and Smelting Company at Flin Flon, Manitoba. The Churchill River project develops 90,000 h.p. at 56-ft. head; the Wellington Lake plant 3,300 h.p. at 75-ft. with pro-vision for an additional 3,300 h.p. unit. Churchill River power is trans-mitted at 110,000 volts over a 58-mile line of steel towers to Flin Flon and an additional 45 miles over a wooden pole-line to Sherritt-Gordon. Our steel tower-line is 22 miles long, and the voltage is 60,000; this choice be-cause certain such equipment was available from Trail. The diversion of 15,000? second-feet at 56-ft. head for 90,000 h. p. at Churchill river dwarfs our 500 second-feet flow at 75-ft. head for 3,300 h.p. In both cases, the main obstacles to be overcome during plant construction were the lack of transport facilities and the short open season. Let it be noted here that this paper deals mainly with those phases of plant construction which were particularly interesting to the writer. With-out these variations from the ordinary routine, this is just another power project whose story is better told by the accompanying charts and pictures. To many members of the Institute it is still just that; but I am writing for those who, like myself, have been too seldom responsible for major under-rakings at an appreciable distance from a railway siding. When it was planned in early 1937 to erect a 1,000-ton mill at our Box mine, both hydro and Diesel power were discussed. With ten years' ore in sight, even an expensive hydro project, such as ours turned out to be, produces kilowatts materially below the 2-cent operating cost of a Diesel plant using fuel-oil at 14 cents a gallon.

Jan 1, 1940

By G. M. Kintz

From 1895 to 1932, inclusive, 1,0163/ men were killed in and about New Mexico coal mines. Of these, 473 (approximately 47 percent) were killed in the 17 explosions listed in this paper. The second largest explosion of record in the United States, in so far as fatalities are concerned, occurred in New Mexico. The data in this circular were taken from publications of the United States Bureau of Mines and published annual reports of the State inspector of coal mines of New Mexico. The tables included are complete only with regard to major disasters. The Bureau of Mines considers as a major disaster one in which 5 or more men are killed. A few minor disasters are included, however.

Jan 1, 1934

By R. P. Hackl, P. G. West-Sells

Carbonaceous gold ores are particularly attractive for leaching by ammonium thiosulfate as the gold thiosulfate complex is not significantly adsorbed by the organic carbon present in the ore. A technically and economically feasible process for treating these ores with ammonium thiosulfate has been developed. In this process, gold is extracted by leaching the ore with ammonium thiosulfate, the solution is separated from the ore by counter - current decantation, the gold is recovered by precipitation with ammonium sulfide, and finally the gold containing product is upgraded by reacting the elemental sulfur present in the precipitate with ammonium sulfite. The precipitation filtrate is recycled back to the grinding step to close the solution balance. Reagent consumption is not excessive. Besides precipitating gold, the precipitation step removes most deleterious solution impurities and converts a portion of the polythionates back to thiosulfate. This paper discusses the optimization of the leaching, precipitation, and upgrading steps. The results of a 10 kg/h pilot plant campaign are also discussed.

Jan 1, 2005

By S. Nitta

The present study deals with a new approach in calculating the size of Tower Mills. The method is derived from the application of grinding kinetics theory on the analysis of by-gone operational data. It requires only information on the capacity of the mill to be designed, the product fineness and the Bond?s work index to give an accurate and simple prediction of Tower Mill characteristics. An example illustrates the different steps of calculation.

Jan 1, 2014

This paper describes an interactive personal computer program, ENVIRON, which simulates cooling and ventilating systems on gold mines. An example of the practical application of the program is given and proposed future developments are discussed. The program consists of two main modules, HEATFLOW and VENT- FLOW. HEATFLOW calculates the heat loads and the effect of air coolers, while VENTFLOW estimates the air flow distribution. As the calculation of the heat loads requires a knowledge of the air flow and vice versa, the two modules are run iteratively until the solution converges. The program is written in Pascal and runs on any IBM-compatible personal computer.

Jan 1, 1987

By Rodrigo Perito Cardoso

Low-temperature nitriding of stainless steels is applied to obtain wear resistant surfaces without losing its corrosion resistance. The treatment consists in introducing nitrogen into the steel surface avoiding chromium nitride precipitation. The corrosion resistance of martensitic stainless steels is expected to be improved if the treatment parameters are correctly selected. It can be related to ?Fe2-3N and a'N phase formation on the treated surface. In the present work the influence of the pretreatment and treatment temperature on corrosion resistance of plasma nitrided AISI 420 steel samples is studied. Annealed, as-quenched and 400ºC-tempered samples were nitrided at 300, 350, 400, 450 and 500°C, for 6 h at 600 V. The pressure was 3 Torr and the gas mixture was composed of 70%N2, 20%H2 and 10%Ar, at a gas flow rate of 3.34×106 Nm3s-1. Plasma was generated using a square-wave pulsed DC power supply and the sample temperature was controlled by the duty cycle. The modified layers were characterized by optical microscopy, X-ray diffraction and microhardness measurement. The corrosion resistance was studied by electrochemical characterization, by means of Open circuit potential and Anodic polarization curves. The maximum surface hardness was achieved at 400°C. This temperature seems to be near the limit temperature for low-temperature treatment aiming to improve or at least to keep unaltered the corrosion resistance in relation to that verified for non-treated steel samples. Results confirm that the pretreatment and treatment temperature significantly affects the corrosion resistance of the treated surface.

Aug 8, 2018

By A. C. Mollison

In November of 1983 a fire inside the autoclave caused the curtailment of operation of the Kidd Creek pressure leaching facility. Subsequent repairs and recommissioning of the autoclave in 1984 allowed Kidd Creek to increase its production of zinc as originally planned. This paper describes the operation of this process from 1984 to 1988 and details major mechanical work carried out on the vessel in 1.988:

Jan 1, 1990

By M Mohammad Khani

An economic block cave is dependent to the characteristics of the rock mass at stages of caving. These are cave initiation, propagation and ore flow. In this research, the rock engineering system (RES), first established by Hudson (1992) to solve complex rock engineering problems, is introduced and verified against the well-established caving parameters established by Laubscher. There are two types of factors influencing caving and these are natural (cannot be controlled) and induced (operator controlled). A RES based interaction matrix is developed by determining the rating parameters required to reproduce the caving database using Laubscher’s caving parameters. This provides the relative importance of the parameters in the interaction matrix. After discussing the rock mass caving process and identifying all the effective parameters, an interaction matrix is introduced to study the cave initiation of rock masses. The expert-semi-quantitative (ESQ) coding method is utilised to assign the numerical value to each parameter’s interaction. Thereafter, the interaction intensity and dominance of each factor in the system is determined to classify these factors. Finally, all the parameters affecting caving operation and their weights on the RES system are identified. This new application of RES in block caving is very efficient and useful for any stage of caving studies and helps experts to make decisions more accurately.CITATION:Mohammad Khani, M, 2016. The application of rock engineering systems to block caving to identify the effective parameters, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 853–860 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By M. T. Ityokumbul

The use of yield index, the product of the yield and gangue rejection, for the optimization of separation processes is presented. The characteristics of the yield index are contrasted with separation efficiency calculations and applied to laboratory froth flotation and washability data. Data from continuous froth flotation tests conducted to "optimize" froth flotation using process optimization software were also evaluated with the yield index procedure.

Jan 1, 1997

By Ray Y. Lin, Randall A. Blue

"Joining of silicon carbide with infrared using a mixture of Si-30wt.%C as the brazing material has been investigated. Joining was performed at a temperature of 1500 °C in a flowing argon atmosphere for processing times ranging from 5 to 60 seconds. The joining filler materials include thin film silicon obtained by thinning down silicon wafers to about 70 /!m and graphite powder of a 15 /!m average particle size. Joint strengths were determined with a specially designed shear-in-compression jig for testing shear strengths of brittle materials. Cross section examinations of the joined specimens with optical microscopy, scanning electron microscopy and transmission electron microscopy have shown that all joints are essentially void free with excellent wetting between the base SiC and the brazing material. Silicon carbide in the B form has been identified as the reaction product in the joining zone with a transmission electron microscope (TEM) and selected area diffiaction (SAD) analysis. At 1500 °C with a joining time of 60 seconds, the joint has consisted of primarily SiC. With such a high SiC content in the joint, the joint strength is comparable to those by other joining techniques, such as diffusion bonding, as reported in the literature. A kinetic model for the formation of SiC from liquid silicon and carbon is proposed to explain the observed joint microstructural evolution.IntroductionSilicon carbide is of great interest for high temperature structural applications because of its excellent high temperature strength, wear resistance and anticorrosive characteristics [1]. Even though the interest in SiC is high, the use of the material has been limited by the high cost and difficulties involved in forming components of complicated shapes [2]. It is these limitations that necessitate the development of effective joining techniques that are cost effective and able to provide strong joints."

Jan 1, 1996