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By Th. Hassel

The application of magnesium as an absorbable biomaterial demanded the consideration of the alloy composition, the mechanical properties and the corrosion a special adapted material development in terms of the human biological system. Magnesium for technical applications mostly comprised alloying elements up to 10 wt%. Thereby often metals (e.g. rare earth and aluminum) with high toxic potentials are used to optimize the mechanical properties. To minimize toxic reactions during the degradation of a absorbable magnesium alloy, the Mg-Ca system is being selected and investigated because both elements belong to the mineral nutrients. Rods and small tubes are produced by extrusion of gravity die casting billets (Ca-content 0.4 to 4.0wt%). The positive grain fining influence of calcium and its improvement of the corrosion behavior are additional advantages. The microstructure, the mechanical properties and the corrosion attitude are investigated. The results show that an optimal composition of micro alloyed Mg-Ca and ideally processing parameters lead to 16% strain and a tensile strength up to 185MPa. Electrochemical investigations and measurements of the hydrogen production demonstrate the influence of the Ca-content to the corrosion behavior of the investigated alloys.

Jan 1, 2006

By Curtis L. Graversen, J. H. Hamilton, John C. Nixon, John R. Lewis

RECENTLY there has been much speculation concerning the advantages of using oxygen enriched air or pure oxygen in pyrometallurgical processes. The advantage of using oxygen in the iron blast furnace and elsewhere in ferrous production metallurgy has been given careful consideration and, as a result, several experiments have been tried and a number of papers have recently appeared on this subject.1.2 The use of oxygen in nonferrous smelting likewise poses interesting problems, therefore any experimental data obtained in this area should likewise be of general interest.3 Ashcroft4 in 1933 claimed to have obtained improved results when copper concentrates were roasted in pure oxygen. However, his' paper gives neither data nor experimental details. A few years ago a research program was started in this laboratory which had for its objective the determination of the mechanism of oxidation of metallic sulfides under varying conditions. To date we have studied the oxidation of galena, 5 sphalerite, and chalcocite. It is a well known fact that copper sulfide minerals are oxidized slowly by weathering agents, so that the sulfides of copper which find their way into a tailing pond are gradually changed into watersoluble and acid-soluble compounds. Commercially, the oxidation of copper sulfide minerals is accomplished by roasting them at elevated temperatures in air using a suitable (usually multiple-hearth) roaster. Most copper smelters are equipped with such roasters. In this phase of the study we were especially interested in determining the effect of temperature on: (I) the air oxidation of chalcocite and (2) the oxidation of chalcocite using commercial oxygen. The amount of sulphur removed as sulphur dioxide, the percentages of water-soluble copper, and of acid-soluble copper were determined for the air oxidation tests. In the tests using commercial oxygen, the percentages of sulphur dioxide, water-soluble copper, acid-soluble copper, water-soluble sulphur and acid-soluble sulphur were determined. For these experiments the temperature was varied from 250 to 700°°C. EXPERIMENTAL Preparation of Chalcocite (Cu2S) Natural chalcocite is not pure. Minerals such as bornite, covellite, and chalcopyrite may be present. It was decided, therefore, to prepare some chalcocite in our own laboratory. An intimate mixture of sulphur and finely divided metallic copper in the ratio of 4 g of copper to I g of sulphur was prepared. To this was added a slight excess of

Jan 1, 1948

By Roger P. Ashley

The main problem in geochemical exploration for gold ore bodies in the 40 km2 hydrothermally altered area at Goldfield is to decide which of the many silicified ledges exposed are most likely to yield deposits with additional subsurface exploration. We conclude that the most efficient way to search for exploration targets is to collect rock samples from the ledges and analyze for the following elements in decreasing order of priority: gold, lead and silver, bismuth, and mercury. These elements form relict hypogene aureoles restricted to the ledges and are not appreciably redistributed during supergene leaching.

Jan 1, 1978

By Waddah Akili

This paper describes the steps taken in planning, developing, and executing a case study/ case history course in geotechnical/ foundation engineering at an international university. The paper sheds light on: how a “workable” format for the course was arrived at; the organization of the course; and the results of evaluating the effectiveness of this approach versus traditional lecturing. Problems and challenges that could arise when offering the course for the first time are also addressed. Embedded in this experience and its related protocols are: the emphases on engineering design and the practice, teamwork and leadership development, organizational management, and oral and written communication skills. The paper concludes by confirming that discussions, through an open forum, are judged to be superior to traditional lectures in improving critical thinking, cultivating desirable personal attributes, and acquiring problem-solving skills.

By J. Güntner

Most copper/cobalt ores from the Central African Copperbelt contain sulphidic compounds, which are not extractable in direct leaching. Roasting provides the possibility to transfer the valuable components into leachable compounds. In an operation window defined by temperature and off-gas composition it is possible to selectively oxidize iron sulphides to haematite and copper/cobalt sulphides to sulphates. Selective sulphating of the valuable minerals is an important condition for the subsequent leaching stage. The dissolved copper and cobalt are recovered by solvent extraction and electrowinning. This technology has been part of an established process for many decades. The concentrates from recent projects on the Copperbelt have lower sulphur contents, and frequently higher copper contents, than those from older projects. The paper shows the consequences of this trend for plant operation and explains useful counter actions by means of case studies. Autothermal combustion is ensured even at very low sulphur contents. Off-gas treatment has so far been based on SO2 removal through sulphuric acid production. The generated acid is used to make up acid losses in the hydrometallurgical plant section. This well-proven process combination is still the preferred route if the feed sulphur content is high enough; however, some cases may require different off-gas cleaning concepts without compromising on clean air quality. Roasting in a fluid-bed furnace remains the core of the technology. Fluid-bed roasters are easy to operate and provide an excellent control of the calcine quality. Different feeding and heat-recovery systems are used to provide a tailor-made solution for each project. Outotec built it first sulphatizing roaster in Zambia about 30 years ago. At that time, the roasting of metal sulphides was already one of the main technologies in the company. Outotec has developed roasting solutions for copper, gold, zinc, lead, molybdenum, and pyrite ores. The continuously expanded and upgraded research facilities provide an excellent basis for developing sustainable, energy-efficient, and economic solutions for copper and cobalt production in Africa.

Jan 1, 2012

By J. Güntner

Most copper/cobalt ores from the Central African Copperbelt contain sulphidic compounds, which are not extractable in direct leaching. Roasting provides the possibility to transfer the valuable components into leachable compounds. In an operation window defined by temperature and off-gas composition it is possible to selectively oxidize iron sulphides to haematite and copper/cobalt sulphides to sulphates. Selective sulphating of the valuable minerals is an important condition for the subsequent leaching stage. The dissolved copper and cobalt are recovered by solvent extraction and electrowinning. This technology has been part of an established process for many decades. The concentrates from recent projects on the Copperbelt have lower sulphur contents, and frequently higher copper contents, than those from older projects. The paper shows the consequences of this trend for plant operation and explains useful counter actions by means of case studies. Autothermal combustion is ensured even at very low sulphur contents. Off-gas treatment has so far been based on SO2 removal through sulphuric acid production. The generated acid is used to make up acid losses in the hydrometallurgical plant section. This well-proven process combination is still the preferred route if the feed sulphur content is high enough; however, some cases may require different off-gas cleaning concepts without compromising on clean air quality. Roasting in a fluid-bed furnace remains the core of the technology. Fluid-bed roasters are easy to operate and provide an excellent control of the calcine quality. Different feeding and heat-recovery systems are used to provide a tailor-made solution for each project. Outotec built it first sulphatizing roaster in Zambia about 30 years ago. At that time, the roasting of metal sulphides was already one of the main technologies in the company. Outotec has developed roasting solutions for copper, gold, zinc, lead, molybdenum, and pyrite ores. The continuously expanded and upgraded research facilities provide an excellent basis for developing sustainable, energy-efficient, and economic solutions for copper and cobalt production in Africa.

Jan 1, 2012

By Andy Fortin, André Vien, Christophe Bouché, Jeff McKay

"The technology and equipment design of the Twenty-first Century permit us to exploit low-grade ore bodies with milled tonnages reaching 300,000 tpd. Because of their substantial operational costs (energy, wear parts, etc) and precarious economic situations, the operations of today must maximize their metallurgical performance.The implementation of an Expert System becomes a logical avenue to pursue in view of increasing plant performance.Many operations, following the implementation of an Expert System, meet the objectives they determined before its integration; such as increased tonnage, stability, decreased costs, etc. However, they often cannot develop the full potential of the system due to certain factors that have been neglected or omitted during the implementation of the system.Three main elements have been identified during the implementation phase at INMET Mining Troilus Concentrator: 1- Instrumentation & Control 2- Software 3- Integration. Even though they have been clearly identified, these elements are often neglected due to lack of resources or lack of comprehension on the part of the operations personnel. This document will demonstrate, based upon our experience at Troilus and that of Metso Minerals, that we can identify and handle many factors which influence the profitability of the Expert System, for both short and long-term process control improvements.INTRODUCTIONThe Inmet Troilus Mine is an open pit mine, situated 175 km north of Chibougamau in the province of Québec, Canada. The mill, which started production in November of 1996, was originally designed to treat a gold – copper ore at a rate of 10,000 tpd.In early 1998, the mine plan called for a pit pushback and decreasing head grades for the following two years. With the price of gold nearing decade lows, the Troilus project was facing a major economical setback. There was a sense of urgency to reduce operating costs below the market price of gold. The mill team presented a project to management calling for an increase in mill throughput while minimizing capital expenditures and operating costs. This idea was supported by the existing mine plan, which called for a mine cut-off of 0.7 g/t Au and stockpiling low grade 0.5 g/t material for future milling."

Jan 1, 2003

By S. J. Rothman, L. T. Lloyd

S. J. Rothman and L. T. Lloyd (Argonne National Laboratory)— Hagel's paperz0 discusses the question of the mechanism for self-diffusion in bcc transition metals. Recent work21"23 has led us to agree with him in reabandoning the ring mechanism in favor of a defect mechanism, most likely vacancies. On the other hand, we think that Hagel's attempt to group together all self-diffusion in bcc transition metals is wrong. We prefer the scheme of Peart, Graham, and omlin," who separate them into two groups. One group contains the anomalously behaved metals P titanium,', y uranium, and perhaps zirconim.' The other group contains a iron," this alloy was held at 1800°C for 2 hr before quenching, its structure was all a. B. C. Giessen, I. Rump, and N. J. Grant (Authors' reply)—We very much appreciate the discussion by Deardorff and Kato regarding the a to /3 transformation temperature of hafnium. In the W-Hf program, as well as in a current study of the system Cr-Hf, standard precautions were exercised against contamination, including vacua of less than 10"5 mm Hg and a weight control check of all specimens to insure an impurity rise of less than 100 ppm. Following receipt of the above discussion, a revision was made in the annealing process. Arc-cast specimens were etched with a 10 pct HF-90 pct HNO solution until 1 mm of surface, including a possible oxide film, was removed. Annealing was done in a hafnium container which provided a long diffusion path for gaseous contaminants. Finally, the heat treating time was shortened. Structural changes were noted to occur quite slowly. A 99.5 at. pct Hf alloy showed 70 pct /3 after 30 min at 1700°C and was 100 pct /3 after 10 min at 1715°C. After 24 hr at 1740C, the alloy was found to be 10 pct/3. This change is due to contamination; it was estimated that it corresponds to an increase of 60°C in the transformation temperature over a 24 hr period. The corresponding change in 30 rnin would be slight. Using the same hafnium as in the above publication, and three new alloys at 98, 99, and 99.5 at. pct Hf, the transformation temperature was indicated to be 1715" 15°C. Since this hafnium contained 4.4 at. pct Zr, a correction of 40 5C, based on an assumed approximately linear transformation temperature curve, must be made for pure hafnium. This yields a value of 1755 20°C in good agreement with Deardorff and Kato's value. Because the slope of the boundary is small, this permits a good extrapolation, and we regard the W-Hf alloys as well suited to check the Hf transformation temperature. niobirn,'' and molybdenum,22 whose activation energies correlate with the melting temperature and whose Do values follow Zener's predictions.28 We have no quarrel with Hagel's placing chromium into the second group; we disagree with his interpreting our data to make uranium fit into that group. Hagel considers that the measurements of self-diffusion in uranium below about 950°C represent combined volume and grain boundary diffusion. This contention is based on the curvature of the penetration plot for our 891°C run,4 which also has a nearly straight line plot of log c vs x. Hagel has extrapolated these characteristics to our 803.5C run and fitted his Arrhenius plot for self-diffusion in y uranium to our four highest temperature points only. We think this is wrong for the following reasons.

Jan 1, 1963

By Jens Markowski

The EU Directive 2000/53/EC on end-of life vehicles (ELVs) states that in 2015 at least 85 wt.-% of a vehicle has to be reused or recycled. This result in the obligatory need to separate additionally metallic and other components of the automotive shredder residues (ASR), so that they can be professionally prepared for recycling and reuse further. Currently the iron-containing fines of ASR are deposited as it is not possible to process in blast furnaces due to its particle size (< 4 mm), composition and material properties. Additionally problems are also caused in handling and transportation (dust) of the material. This work focuses mainly on the ASR materials utilization in blast furnaces. Material was further separated into two consecutive fractions by carrying out repeated crushing and magnetic separation twice. The two fractions have significantly higher iron content (up to 75 wt.-%) compared to the initial material mixture. The materials were assessed for their usage in blast furnaces based on the possible pelletisation with and without binders, stability of pellets, shape and size of pellets, etc. Pelletisation of the material was carried out with a single mould pellet press. Further lignite coal was added to the mixture as additive which improved the pelletisation behavior and the mechanical stability of pellets. Lignite coal was chosen as a reducing agent in the blast furnace process contributing towards additional savings in energy supplied. The experiments reveal that the production of pellets from iron-containing fines is possible with a lignite coal admixture of about 15 wt.-%. These pellets displayed good mechanical stability and can be used in blast furnace processes in recovery of the recyclable metals which are currently not recycled. The recycling of not yet recyclable metals would reduce the deposition and can be easily adopted into the economic cycle by their utilization. Keywords: Automotive shredder residue (ASR), FE-fines, pelletisation

Sep 1, 2012

By Joseph J. Wallace

1. From the standpoint of present production and recoverable reserves, there are two beds of major importance in Allegheny County -Pittsburgh and Upper Freeport -and one of lesser importance -Redstone. 2. Known measured and indicated reserves in all beds, based on a minimum thickness of 14 inches and 1,800 short tons per acre-foot of coal in place, are estimated at 832 million short tons as of January 1, 1953. Of this total, 768 million short tons is in beds 28 inches and more thick. Areas in some beds were omitted from the estimate because available data relative to the bed characteristics are too meager for making an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1953

By John S. McCartney

This paper focuses on the seasonal thermo-mechanical response of two energy foundations installed at the new Denver Housing Authority Senior Living Facility in Denver, Colorado. The two drilled shaft foundations, which have diameters of 36 inches and depths of 45 ft and are end bearing in the Denver shale formation, include a closed-loop series of polyethylene tubes through which fluid is circulated to exchange heat with the surrounding ground. The tubes within the foundations are coupled together with the conventional ground source heat pump system for the building and were installed to evaluate the heat exchange response of the foundations as well as any thermo-mechanical soil-structure interaction effects. A series of vibrating wire strain gages and thermistors were installed at different depths in the foundation to record the thermo-mechanical strain and temperature at different depths. Also, pipe-plug thermocouples are used to measure the inlet and outlet temperature of the heat exchanger fluid entering each of the foundations, which is necessary to evaluate the total heat exchange for the foundations and the efficiency of the heat pump. This paper will describe the energy foundation construction processes, the strain distribution and ambient temperature profiles during building construction and the seasonal strain and temperature distributions during heating and cooling of the building. The data indicates that typical heat exchange associated with heating and cooling operations leads to strains which are within acceptable limits, and that the foundations can provide a clear option for absorbing the base heating and cooling load for the building.

By Mahmoud Shakeri1

Various types of thin-walled structures have been frequently used as energy absorbers to mitigate the adverse effects of impact and hence protect the vehicle structure under consideration when subjected to different loading conditions. Compression tubular structures that buckle in a progressive manner, can be used to advantage, providing inexpensive and versatile, but nevertheless weight and stroke efficient energy absorbers. In this paper quasi-static nonlinear finite element simulations are performed to study the effects of number of corners of the cross-section, wall thickness, plastic hardening rate of material, foam-filler and load angle on deformation mode and mean crushing force of tube. Crush strength increases as the number of corners of the cross-section increases, though it almost saturates for the number of corners beyond 11. Higher plastic hardening rate stabilizes the collapse pattern. Foam-filled tubes seem to have higher critical load angles in oblique loading and display a more stable and gradual decrease in mean load over the transition region. Substantial decrease in the initial peak load and dynamic load at any a given deflection was observed as the load angle increases. The investigation of bitubal arrangements and Multi-cell tubes reveals that these may be preferable to monotubal specimens due to the presence of the inner profiles. In addition, a comparison is conducted between the numerical and experimental results of the mean forces of some special angle elements. Good agreement is obtained.

Oct 1, 2011

By S. G. Proctor, D. L. Baaso, W. V. Conner

A double crucible system was developed for I-g scale plutonium reductions. The equipment consists of an inner MgO crucible, an outer MgO crucible, and a stainless steel pressure vessel. The reduction charge is PIaced in the inner crucible and the annulus between the inner and outer crucibles is filled with a mixture of calcium and iodine. The exothermic reaction between the calcium and iodine in the annulus supplies the heat required for complete reaction of the reduction charge ana&apos; good metal coalescence. Metal yields of 80 to 85 ,pct were obtained from I-g scale reductions and yields as high as 97.5 pct were obtained from 0.5-g scale reductions. The system was used to reduce charges containing as little as 0.1 g of Pu resulting in metal yields up to 90 Pct. THIN metal foils of highly enriched plutonium isotopes are used as targets for cross section and other measurements. The isotopes are separated in the calutrons at Oak Ridge and are very expensive to prepare. Often, only 1 or 2 g of material are available thereby emphasizing the need for a method of preparing these small quantities of metal. Procedures for 1-g scale plutonium reductions have been described, but these procedures require elaborate or expensive equipment. For example, the procedure described by Anselin et al.1 requires an inert atmosphere glovebox and induction heating equipment. The procedure described by Baker2 also uses induction heating equipment to obtain the recommended heating rates and liner temperatures. Baker&apos;s procedure also requires high purity PuF4 with very little PuO2 present. The procedure described in this paper resulted from a search for a simple and inexpensive method for making 1-g scale plutonium reductions. EXPERIMENTAL Equipment. The equipment required for the double crucible system consists of a pressure vessel, two MgO crucibles, a MgO crucible lid, and a resistance-heated, vertical crucible furnace. The crucibles were slip cast from high purity MgO and were supplied by the Coors Porcelain Co. and the Norton Co. The pressure vessel and lid were fabricated from 316 stainless steel. Materials. The PuF4 used for this study was obtained from the P1utoniu:m Metal Production Department at Rocky Flats. The PuF4 was prepared in a con- tinuous hydrofluorinator by reacting Pu02 with HF at 650°C. The isotopic composition of the plutonium was approximately 93 pct 239PU, 6 pct 240PU, and 0.5 pct 241PU. Some of this PuF, contained less than 1 pct Pu02 (Batch No. 6 and 7), while other batches contained up to 15 pct Pu02 (Batch No. 3). The chemical analyses of all the PuF4 used for this study is given in Table I. The calcium was 99 pct pure, AEC grade, and only that fraction which would pass through a 20 mesh screen was used. The I2 was USP grade re-sublimed I2 which was ground before being used. Procedure. Various charge compositions and methods for loading the double crucible system were tested. The optimum conditions for 1-g scale reductions are described below. The double crucible system was loaded as shown in Fig. 1. The outer crucible was placed in the pressure vessel and the annulus between them was filled with MgO sand. The inner crucible was placed in the outer crucible, supported by a layer of MgO sand in the bottom of the outer crucible. A mixture of 5.2 g of Ca and 25 g of I2 was placed in the annulas between the two crucibles. The upper portion of the annulas was filled with MgO sand. Next, a layer of a mixture of calcium and I, equal to 20 wt pct of the calcium and I2 used in the main charge was placed in the bottom of the inner crucible. The PuF, to be reduced was mixed with a 30 pct excess of calcium and 1 mole of I2 per mole of Pu and this mixture was placed in the inner crucible. The charge was topped with a layer of a mixture of calcium and I2 equal to 20 wt pct of the calcium and I2 used in the main charge. The pressure vessel was sealed with a flat copper gasket and was purged by alternately evacuating and filling with argon. The purge valve was closed and the vessel was placed in a vertical crucible furnace which was preheated to 950°C. The furnace was turned off after 20 min of heating and the vessel allowed to cool. Experience has shown that. this amount of heating is sufficient to assure complete reaction of the charge. RESULTS AND DISCUSSION The double crucible system has been used to produce plutonium metal on a 0.1- to 1-g scale. Reduction

Jan 1, 1970

By Gerhard W. Heinrich, Gerald V. Schroeder, Daniel A. Murphy, Daniel F. Suarez, Jamie B. Beauchamp

"Washwater 'Finger' sprays were installed and tested in the copper/nickel separation circuit at Strathcona Mill to improve flotation capacities and related metallurgical performances. The new spray design features submerged sideways introduction of wash water into the flotation froth through nozzles at the end of vertical pipes which are suspended from a header above the froth. This maximizes the effective cross sectional area of flotation vessels while maintaining effective froth cleaning and a desirably high density of the concentrate slurry. The plant tests gave increases of copper concentrate production rates of 17% to 20% for conventional cells and 16% to 20% in column flotation at target grades of 29% to 30% copper and of less than 0.5% nickel. Basic considerations and concepts of tests and evaluation procedures are described.IntroductionMeeting flotation concentrate grade targets at maximum production rates and recoveries is a challenge all too familiar to operators of base metal ore concentrators. Frothwashing is an option to improve grade/recovery performance by minimizing entrainment of impurities into the concentrate, which is commonly known since the sixties (Kaya & Laplante, 1989). It has been recommended for the flotation of sulphides and industrial minerals alike (Kaya & Laplante, 1989; Miller, 1969; Heinrich, 1989) and has found its ultimate application in flotation columns (Finch & Dobby, 1990). The washwater addition in a flotation column is most evenly distributed over the column's cross-sectional area thus ensuring almost complete displacement of feed water from the froth by a countercurrent washwater flow -provided, a positive bias is maintained. Therefore, flotation columns are ideal for cleaning flotation. However, there is usually less room for boosting the concentrate production capacity by increasing airflow or slurry level then with conventional cells because of the smaller cross-sectional area available."

Jan 1, 1994

By Kenneth A. Rogers, Richard F. Hill

Introduction A number of processes for deriving more convenient liquid and gaseous fuels from coal are commercially available or in final stages of development. Generally, synfuels from coal processes have efficiencies ranging from 50 to 70%. Considerable engineering design and development work are necessary to bring forth commercial plants using the various technologies-but technological breakthroughs are not expected or needed. Gasification While a variety of solids handling, gas contacting, and gas quality upgrading techniques may be employed in the various gasification processes, the primary goal is the same-to convert a sulfur-containing solid to an essentially sulfur-free gaseous product. The gaseous product may be a mixture of hydrogen and carbon monoxide with nitrogen. Definitions are not rigorous, but high-Btu gas is generally over 900 Btu/scf(33 534 kJ/m3) and low-Btu gas is generally below 350 Btu/scf(13 041 kJ/m3). Medium or intermediate gas falls between these values. High-Btu production is more difficult, as it requires additional upgrading of a low- or intermediate-Btu gas stream to produce a commercially acceptable product. The carbon/ hydrogen ratio is adjusted either by rejecting a portion of the carbon, increasing the hydrogen, or a combination of both. The chemical processing steps include production processing gas containing carbon oxides and hydrogen by the reaction of carbon and steam (gasification); reaction of hydrogen and coal to form methane (a hydrogasification process); reaction of hydrogen and carbon monoxide to produce methane (methanation); and the removal of impurities. The gasification processes share a number of common techniques. The coal received at the plant site must be further cleaned and crushed or ground before it can enter the gasifier. Conventional cleaning methods are used to remove extraneous materials such as metal, rocks, etc. For some processes the coal must be finely ground so it may be fluidized or entrained in a flowing gas stream. For other processes crushing and sizing only are required. Agglomerating coals may be pretreated by heating, mild oxidation with air, or other procedures to assure smooth gasifier operation and to avoid reactor plugging. Nonagglomerating coals Such as lignite and subbituminous coal normally do not require pretreatment. Feeding a solid material such as coal to a higher pressure reactor is a difficult task. Dry feeders using a series of pressurized lock hoppers, or wet systems in which a slurry is pumped into the reactor, are used. The gasification step is the heart of the process. Here, the solid feed material is contacted with synthesis gas (carbon monoxide and hydrogen). The coal is devolatilized and a portion of the fixed carbon converted to methane. The gas-solids contacting methods used include fixed beds, fluidized beds, moving beds, entrained beds, and free-fall systems. The gasification process, overall, is endothermic and the heat required to maintain reaction rates is usually supplied by burning a portion of the coal feed. External heat sources such as electric resistance heating, molten salt or solid heat carriers, and the introduction of a hot hydrogen stream also have been tested. Following the gasification step the particulate materials and heavy hydrocarbons contained in the reactor effluent are removed by a water quench and solids separation system. If low-Btu gas or synthesis gas is the desired product, the raw gas is scrubbed to remove carbon dioxide and sulfur compounds and then is used directly. However, if high-Btu gas is desired, methanation is necessary. In most gasification processes, the hydrogen content of the gasifier

Jan 2, 1980

By Sneyd R. J, Walton M. C

In early 1987 Minproc acquired a license from the inventor to engineer and construct acid plants incorporating the Fattinger technology. The process offers a simple and reliable method of producing sulphuric acid from metallurgical off gases at significantly lower capital and operating costs than the more familiar contact process. The process also offers the simultaneous production of nitric acid, if the NOx content of the off gases is high enough. A development of this latter process has other applications within the minerals industry, particularly in conjunction with refractory gold ore preoxidation.

Jan 1, 1989

By A. S. Macfarlane

Over the last few years, and especially the last few months, issues relating to Corporate Governance in industry in general, and the Minerals industry in particular, have become increasingly topical, as a result of a number of scandals and failures. The development of International and National Codes for the reporting of Mineral Resources and Reserves began some years ago, and has been followed by more stringent Listing Rules for the Securities Exchanges, as well as more standardised formats and protocols for financial reporting of assets and activities. More recently, some countries have adopted Codes for the Valuation of Mineral Properties, which have focussed on standardisation of valuation approaches to be adopted, and the qualifications of people deemed competent to sign off such valuations. This process has not been without criticism, the main criticism being the proliferation of local standards and regulations, which are often difficult to apply and regulate, or which are too prescriptive for multinational mining companies to deal with. In South Africa, these concerns have been recognised, while at the same time a need has been identified to establish a Code which could be applied to South Africa, and the African continent as a whole. Such a development would necessitate that the Code satisfies the needs of the continent, but in a global context. This global context has already been set by the activities of the International Organisation of Securities Commissions, the International Valuations Standards Committee and the International Accounting Standards Board. This paper explains the progress and process that has been established in South Africa for this purpose, and highlights the key issues which are open to public debate, so that the process is inclusive and transparent, whilst at the same time it is consistent with international time frames, and provides a deliverable product that meets the desired objectives.

Jan 1, 2004

By C. D. Anderson

A variety of processing technologies exist for recovery from both primary and secondary sources of rhenium. Currently, there are no known primary rhenium deposits; thus, the method in which primary rhenium is produced is dependent on the commodity of which it is a byproduct, e.g., copper, molybdenum, uranium, etc. In addition, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing. This paper presents a review of both primary and secondary processing technologies for the recovery of rhenium.

By Nicholas J. Grant, Rolf Nordheim, Bill C. Gissen

The system Cb-Re was examined in detail utilizing pure metals, careful melting techniques, and heat treatments. Metallographic and X-my methods were utilized for phase identification. In addition to soli-dus determinations, the composition limits and mode of formation of the intermetallic compounds 0 and X were determined. The binary constitution diagram Cb-Re has been the subject of a number of investigations.&apos;-7 Two intermetallic phases have been described and tentative diagrams have been proposed. It was the aim of this investigation to obtain a complete and more accurate diagram through the use of purer alloys and improved techniques. The diagram worked out by Knapton7 was published after our experiments were concluded. EXPERIMENTAL METHODS The starting materials were Re powder of 99.95 pct purity, supplied by the Chase Brass and Copper Co., and Cb rondelles of 99.6 pct purity, supplied by the Electro Metallurgical Co. Typical analyses are given in Table I. The alloys were prepared from compacted Re powder and Cb rondelles; for critical compositions master alloys were used. The melts varied from 5 to 50 g and were melted nonconsumably in an Heraeus arc furnace, under ti-tanium-gettered laboratory grade argon. It was necessary to invert the buttons and remelt, followed by crushing and further remeltings, the number of which depended on the nature of the fracture of the alloy. The weight of each alloy was carefully checked after each melting operation. It was found that significant weight losses occurred only when loose Re powder was used. These losses amounted to a maximum of 3 wt pct in such instances, but did not occur after the charge was once molten, regardless of the composition. Typical loss of weight after 4 remelts, including crushing, amounted to 0.2 wt pct. Oxygen pickup during successive melting cycles did not appear to be significant, since the weight gain values never exceeded 0.01 wt pct. Control of the composition of the alloys was not too difficult; however, the problem of homogenization was a far greater one. Optimum alloying conditions were finally achieved by a combination of master alloy production and repeated crushing and remelting cycles. The compositions of the alloys utilized to determine the constitution diagram are listed in Table 11. The weight balance method was used ultimately as the sole analytical method, since repeated checks indicated that it was accurate to 0.2 pct, if the weight losses were small. The results were double-checked by a metallographic study of the homogeneity of the samples, since this was presumed to be the more serious factor. Heat treatments at 1200°C or lower were accomplished by sealing the specimens in Vycor tubes, heating in a Globar furnace, followed by air cooling. This was considered to be an adequate quench because of the very low temperatures involved. For temperatures of 1200" to 1900°, alloys were homogenized and heat treated in a tungsten-filament resistance vacuum furnace. From 1900" to 2750" a tantalum-tube resistance furnace was utilized, operating under a vacuum of 105 mm Hg. The temperature measurements up to 1200" were made by means of platinum-platinum 10 pct Rh thermocouples, with an accuracy of 5°C. For temperature measurements greater than 1200°C, optical pyrometers were used. They were calibrated against a standardized instru-

Jan 1, 1962

By J. M. Richards

For most of the 20th century optical petrography has been the primary petrological and mineralogical tool used to characterize coal. The development of quantitative SEM-based techniques, e.g. QEMSCAN®, for coal began only about a decade ago. The application of these techniques for coal lagged behind other commodities, but they are currently being developed with the aim to provide ?one complete analysis? for coal. Quantitative SEM-based techniques are supplemented by quantitative X-ray diffraction (XRD). Recent indications are that these more modern techniques cannot replace the ?standard? petrographic and chemical evaluations, but rather complement them where and when required. The great advantage of quantitative SEM-based techniques is that they are very rapid, with the result that large volumes of samples can be processed on a routine basis. This is ideal for coal type identification, since the results can be used in the creation of ?intelligent? composites. This can lead to the more speedy evaluation of coal deposits by reducing the number of samples on which detailed metallurgical and characterization test work is required, without an increase in the overall statistical error of the resource model. Coal petrography, however, remains important for the prediction of the coking characteristics of certain coals and coal products. As a consequence it is therefore important that any coal laboratory be able to produce data with confidence. This requires strict quality control and assurance protocols that adhere to international standards. Keywords coal petrography, coking strength, scanning electron microscopy, QEMSCAN®.

Nov 1, 2013