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By Scott D. Thomson

The client for this project specializes in the use of state-of-the-art technology to fabricate and assemble composites and metal-bonded structures for commercial and military aircraft programs. One of the key components for this operation at the client's plant is a highly precise computer-numerical-controlled router known as the Shoda Machine. This machine tool was installed at the plant in July 1998 for the manufacture and inspection of aircraft parts, and it is extremely sensitive to deflections. To monitor the occurrence of deflections, an internal laser system continuously checks the machine calibration. Deflections in excess of 5/1000-inch in the foundation on which the machine rests require costly periodic re-leveling and recalibration of this tool. Therefore, a significant problem arose when measured deflections during machine operation were as large as 20/1000-inch. In addition, the foundation was cracked in at least three places as a result of elastic foundation movement caused by dynamic loading as well as shrinkage and loss of base support. Deciding on a remediation procedure to alleviate or eliminate these problems involved the following important considerations: (1) achieving the client's need for reducing deflections to less than 5/1000-inch, (2) minimizing the repair cost as much as possible, and (3) reducing the time required for implementation of the foundation remediation solution. After evaluating the advantages and disadvantages of several alternatives, jet grouting was selected as the most suitable solution. The jet grouting program provided columnar elements of soilcrete to remediate the Shoda Machine foundation.

Jan 1, 2003

By K. J. Fair, H. W. Smith, D. R. Francis, J. M. Toguri

"Bench-scale batch milling and screening tests were carried out on secondary aluminum dross 10 produce an aluminum concentrate. To complement these tests, an analysis of a commercial plant was also performed by sampling dross at various stages in the plant's milling and screening circuit. These studies have indicated that benefits of increased capacity, grade and recovery of aluminum metal can be realized by addition of screens before the ball mill, longer grinding period and by grinding the smaller size fractions more efficiently.To test the predictions, a plant test was performed on the milling of100tons of secondary dross. The recycling of the large size fraction (+1/8 in.) of aluminum dross through the impact mill rather than through the rate limiting ball mill resulted in a comparable aluminum grade. The ball milling of the _1/ 8 in. + 10 mesh size gave a dross grade of71% AI. A lower feed rate to the ball mill resulted in a higher final grade. With minor modifications, secondary dross can be processed in 22% of the time it takes with the present circuit without a sacrifice in grade.IntroductionSecondary aluminum dross is a mixture of metallic aluminum, aluminum oxides and salts (alkali halides). Any comminution process deforms or slowly grind s the malleable metal phase and reduces the size of the friable, brittle, oxide-salt phase relatively fasterr(1).The milling of aluminum dross is performed to upgrade the metallic aluminum fraction. Figure I shows a flowsheet for the milling of secondary dross as presently practised at Metals and Alloys Co. Ltd., Toronto, Ontario. At point A, large acceptable grade material is hand picked for refurnacing, the unacceptable off-grade material is sent to the sweat furnace and the remaining, + 2 in. fraction is recycled through the 3.9 ft by 4.1 ft impact mill. For simplification, three ball mills are shown. However, in reality there is only one, a 5 ft by 10 ft ball mill used in a semi-batch mode. The 2 in. + 3/8in. and 3/8 in. + 10 mesh products are remilled in the ball mill. The -10 mesh material is discarded as waste."

Jan 1, 1987

By G. R. Peterson

To determine the availability of lead and zinc from demonstrated resources, the Bureau of Mines evaluated 235 mines and deposits in 31 market economy countries. Of the 235 mines and deposits evaluated for this study, 186 were evaluated as zinc operations, 30 as lead operations, and 19 as copper operations. Demonstrated lead-zinc resources of market economy countries in 1981 were approximately 4.3 billion metric tons (t) of ore containing 221 million t of zinc and 97 million t of lead. Of these amounts, approximately 154 million t of zinc and 70 million t of lead are estimated to be recoverable. The analyses indicate that demonstrated resources in market economy countries should be sufficient to satisfy projected demand for primary lead and zinc through the balance of the century. The U.S. lead industry will continue to have a comparative advantage over the rest of the world industry, barring any drastic increase in the cost of compliance with pollution control regulations. It appears that the comparative disadvantage faced by the U.S. zinc industry will probably intensify owing to the relative quantity of lower cost zinc resources in other countries, especially Canada, Australia, Mexico, and Peru.

Jan 1, 1985

A study of the accessory opaque iron-titanium oxide minerals of the Oiles Complex, central Australia, was an Important complement to an investigation of the magnetic properties of the Complex. These minerals constitute no more than 10 per cent of the rocks sampled, and generally less than 1 per cent of the ultramafic rocks ("magnetite" concentrations were not sampled for magnetic investigation).Titanomagnetite is the main opaque mineral present, although ilmenite exsolution has been widely developed. High-temperature oxidation of these produced mainly titanohematite, pseudobrookite and rutile. Both exsolution and high-temperature oxidation developed along specific crystallographic directions (along {111}), and in Irregular patches. Low-temperature oxidation is generally not well-developed, indicating the small degree of weathering in the rocks sampled.Although the Oiles Complex is layered, no apparent overall preferred orientation of the iron-titanium oxide minerals was observed, other than crystallographically controlled alignment of very fine grains within pyroxene crystals. Other textural features of the opaque oxides are restricted to symplectic intergrowths mainly with the ferromagnesian silicates.

Jan 1, 1971

By R. C. Milici

Historical production data and reserve decline rates can be used both to predict future coal production rates and to estimate when the reserve will be depleted. Reserve estimates inferred herein from historical production data and decline rates are commonly lower than estimates made from geologic data. The latter include significant amounts of lower quality (deeper, thinner, higher sub) coal which may be mined when the better deposits are depleted.

Feb 24, 1997

By T. T. Chen, W. Petruk

The mineralogy of the feed, concentrates and tailings from the Heath Steele mill in New Brunswick was studied in detail to determine the liberation and recoveries of the major ore minerals, and to characterize the occurrence and distribution of the trace elements. The ore milled during August 1977 contained an average of 4.34 wt % Zn, 1.25 wt % Cu, 1.64 wt % Pb and 90 gr/ton Ag. The main trace elements in the ore are silver, cadmium, indium, cobalt, antimony, bismuth, arsenic, mercury, tin and gold. Silver is carried in solid solution in tetrahedrite-tennantlte, galena, Ag-Sb sulfosalts and Pb-Bi-Sb sulfosalts. Approximately half of the silver is recovered in the lead and copper concentrates and the remainder is lost in the tailings and Zn concentrate. As the major silver-bearing minerals also contain Sb and Bi, recovering silver by ore dressing would involve recovering elements that are troublesome in extraction metallurgy. Cadmium and most of the indium and mercury occur in solid solution in sphalerite and consequently are recovered in the zinc concentrate. Some indium is also recovered in the copper concentrate in a second unidentified indium-bearing mineral. Cobalt and arsenic occur principally in arsenopyrite, cobaltite, a Co-Fe sulfarsenide, pyrite and pyrrhotite, and follow these minerals into the tailings. Recovery of the valuable element cobalt, therefore, would involve recovering the environmentally hazardous element arsenic. Most of the tin and gold report to tailings. The ore was ground to about 80% minus 270 mesh. At this grind, 71 wt % of the sphalerite, 66 wt % of the chalcopyrite and 59 wt % of the galena were liberated. Recoveries of free sphalerite, free chalcopyrite and free galena in the Zn, Cu and Pb concentrates were 85 wt %, 76 wt % and 80 wt %, respectively; recoveries of unliberated sphalerite, unliberated chalcopyrite and unliberated galena in the respective Zn, Cu and Pb concentrates were 48 wt %, 5 wt % and 7 wt %. Some of the unliberated sphalerite and chalcopyrite grains in the tailings are large and would have been liberated at a finer grind.

Jan 1, 1980

By Philip Bonanno

The first major project in Boston's new access to Logan Airport was an Immersed Tube Tunnel which extended 3,600 ft. from South Boston to East Boston across Boston Harbor. Twelve tube sections, 300 ft. long x 80 ft. wide x 40 ft. high, were lowered into an excavated trench which was then backfilled. The depth of the trench at mid-channel was approximately 100 ft. below mean high tide. At the East Boston end the tube terminated close to the present shoreline. At the South Boston end the tube terminated approximately 600 ft. inland from the shore line where a ventilation building was required. The depth to subgrade from existing ground at the ventilation structure was approximately 85 ft. The soil profile at this location consisted of 50 ft. of miscellaneous rubble fill, 25 ft. of Boston blue clay, and 10 ft. of glacial till. A small amount of weathered argellite existed at the north end of the structure. Part of the Morrison Knudsen, Interbeton, J.F. White, Joint Venture contract to construct the Immersed Tube Tunnel (ITT) was the construction of a cofferdam at the tube/ventilation building interface. The ventilation building was to be built as part of another contract. The construction of the cofferdam and the excavation to the building subgrade was the responsibility of J.F. White Contracting Company.

Jan 1, 1994

Provide the mining industry with a simple, accurate and readily available means of measuring the gas flow from pipe outlets such as gob holes Approach The Bureau of Mines conducted laboratory studies to establish correction or method factors for anemometers when used to measure gas flow from pipe outlets. Using the Anemometer Anemometers are used by the mining industry to make air velocity measurements in mine entries. Although the instrument has no appreciable influence on the airstream in mine entries, especially since the cross-sectional area of an entry is much larger than the cross-section of the anemometer, correction or method factors are used to correct the center line anemometer velocity to average velocity. This adjustment is made because the air velocity in a mine entry is highest in the center and lower at the sides. Anemometers can be used to measure gas flow from small diameter pipe outlets; however, since the instrument diameter is much larger than one-sixth of the pipe diameter, it will influence the air-stream from these small outlets. Resulting velocity measurements are erroneously high by as much as 30 percent. Method factors must be applied to such anemometer measurements taken at pipe outlets. Through a research investigation the Bureau of Mines established a set of method factors for center-line anemometer readings taken at the discharge end of 4-to 8-inch diameter pipes. These method factors were combined with method factors which were developed by the National Coal Board of Great Britain for 12-to 30-inch diameter pipes. The anemometer was mounted on the end of a short rod and held against the pipe for support except on the 4-inch pipe where the instrument was started and stopped remotely to prevent hand interference with the air-stream during measurements. Method factors were determined by comparing the anemometer measurements at the pipe outlet to the true average airflow which was simultaneously determined upstream by venturi measurements. Consequently, a method factor is nothing more than the ratio of the true velocity to an anemometer measurement.

Jan 1, 1987

By N. T. Lowe

The National Institute for Occupational Safety and Health (NIOSH), Spokane Research Laboratory (SRL), investigated a means to evaluate the potential impact of surface conditions above underground mines on emergency response activities. Recent events have highlighted the need to conduct critical emergency operations on the surface above underground work areas. Accessibility and usability of the surface, such as the ability to drill holes or deploy seismic monitoring or through the earth communications systems, are an important consideration for emergency response planning. This paper identifies critical surface conditions such as: surface waters, valley fill, highways, railroads and significant structures that must be considered in emergency response planning and in the conduct of emergency operations as they relate specifically to the seismic detection of trapped underground miners. Results of case studies indicate a need to consider the impact surface conditions may have on emergency response capabilities.

Jan 1, 2009

By Budiselic C, Hefter G

The uptake of gold by activated carbon from acid chloride solutions involves the reduction of gold(IIl) to metallic gold. Deposition occurs only on the surface of the carbon at localized active sites at first, and subsequently at less active areas. The rate of uptake depends on the surface area of activated carbon, which confirms results from a recently published study (de Siegel and Soto, 1984). Increased rates of uptake were observed at higher temperatures and lower chloride concentrations due to the decreased stability of AuCI4-, the major gold(IIl) species present. Recoveries of gold approached 100 per cent for a contact time of 60 minutes at 20°C in a well stirred solution initially 5 ppm in gold(III) and containing 0.5 per cent w/v coconut-shell carbon. The presence of 0.25 Mnitric acid or 100 ppm iron(IIl) had no detectable effect on either rate or efficiency of gold uptake.

Jan 1, 1985

As far as is known at the present time, important ore occurrences at Mount Isa are restricted to the Urquhart Shale. These occur at Mount Isa itself, and some 13 miles north along strike at the Northern Leases.The primary silver-lead-zinc and copper orebodies occur as adjacent but discretc entities in lithologieally distinct host rocks. The copper orebodies occur in an envelope of "Silica Dolomite", whereas the silver-lead zinc orebodies occur in shale typical of the Urquhart Shale in general.At both Mount Isa and the Northern Leases, secondary lead orebodies occur at thc surface representing primary stratiform silver-lead-zinc sulphide accumulations in the fresh rocks at depth.Secondary copper orebodies are known only at Mount Isa, where they have formed in favourable environments, but unlike the secondary lead orebodies, not in the positions of the former primary mineralization that gave rise to them. The secondary copper is localized in shale remote from the point where the copper was released into solution by weathering. It was derived from very low grade fracture fillings and disseminations of chalcopyrite in the outcropping or suboutcropping crest of the "Silica Dolomite".Primary silver-lead-zinc orebodies arc oxidized without significant transport or secondary enrichment, but secondary copper orebodies arc characterized by considerable transport and extreme secondary enrichment.The original water table stood at approximately 1,150 ft (mine datum) and for the majority of the mine area controlled the base of complete oxidation fairly closely.An oxidized zone approximately 175 ft deep is general in the Urquhart Shale.Structures affecting ground conditions have influenced oxidation and leaching of the silver-lcad-zinc orebodies, and have had effects vital to the formation of secondary copper orebodies.

Jan 1, 1966

By B. W. Gandrud

This is the third report published by the Bureau of Mines concerning the washability of Black Creek coal. Both of the former reports5/6/ dealt with the Black Creek coal at the Bradford mine, Dixiana, Ala., near the northeastern extremity of the Warrior field, whereas the present report is based on float-and-sink tests of coal from what is presumably the Black Creek bed near the southern extremity of the field. Although the Yolande Coal & Coke Co., owners of the property at Yolanda No. 6 slope, consider this opening to be on the Black Creek bed, the authors do not feel that there is as yet positive proof that this correlation is correct. There has been little or no development of the Black Creek bed in the Yolande Basin in the past, and a posy-tine identification of the bed at Yolande No. o will have to await further exploration and development. The Yolande No. 6 is an exploration slope 3/4 mile north of Rockcastle, Ala. The slope is approximately 400 feet from top to bottom on a dip of 18° As fax as the authors can ascertain, coal has never been produced commercially from the Black Creek bed in the southwestern part of the Warrior field. Exploration work in the past has indicated that the Black Creek coal in. this area is extremely friable and high in sulfur. Extreme friability would naturally discourage its' use as a domestic and railroad fuel, and high sulfur content would be a disasvantage in its use for coking coal. The present study of this coal has been made with particular reference to sulfur elimination on the assumption that it would make a good coking coal provided enough of the sulfur could be eliminated by washing.

Jan 1, 1939

The mining industry to the lay mind is one which appears peculiarly dangerous, partly because the accidents which take place are usually of a very tragic nature, and partly because other, industries claim victims more from disease than fromviolence, and their fatal nature is not advertised to the same degree so that in spite of this idea of unusual risk, it is a fact that mining will compare favourably, as regards vital statistics, with nearly all other trades requiring special training.Neverthelessthere are a larger number offatal and serious accidents occurring year by year than one would like to see, and could the deaths which take place by accident be eliminated, it is safe to say that mining would become one of the most healthy of all species of skilled labour. If we examine the accidents appearing yearly in the reports of the Mining Departments of the various colonies, it will be at once admitted that many of them need riot, and in many cases ought not to, have occurred.It is, therefore, the object of tbis essay, to pass under review the ordinary precautions which should be adopted to minimise, as far as possible, the dangerous factors whereby fatalities are brought about.

By Herbert H. Einstein

This report contains the documentation for 11 computer programs representing the entire computer documentation for the eight-volume report entitled "Risk Analysis for Rock Slopes in Open Pit Mines." A detailed summary of the programs is included.

Jan 1, 1979

By B. A. Wilcox

ThERE are several reports in the literature which indicate that both solid-solution hydrogen and hydride precipitates can promote low-temperature em-brittlement of tantalum.1-3 For example, Imgram et al.2 showed that hydrogen additions in the range of 140 to 580 ppm promoted a ductile-to-brittle transition at slightly above room temperature in both wrought and recrystallized tantalum. This contrasts with the observation that tantalum of norma1 purity is very ductile, even at liquid-nitrogen temperature. Imgram et al. also found that hydrogen additions caused a deterioration of notch tensile properties of tantalum. The purpose of this communication is to present some limited evidence that hydrogen can also cause fatigue embrittlement of tantalum. The material used in this investigation was wrought, stress-relieved (1000°C, 1 hr), are-melted tantalum rod (5/8-in. diameter) containing the following impurities in weight percent: 0.005 C, 0.007 02, 0.0005 H2, 0.003 N2, 0.08 Cb, 0.01 W, and 0.01 Fe. Hydrogen charging was done in an atmosphere of flowing dry hydrogen for 3 hr at 600°C, followed by an homogenizing anneal in purified argon for 15 min at 600°C, and cooling in an argon atmosphere. Typical micro-structures of the hydrogen-free and the hydrogen-charged materials are shown in Fig. 1(a) and Fig. 1(b), respectively. Fig. 1(b) shows the appearance of hydride precipitates which, in many cases, formed preferentially parallel to the rod axis. The hydride dispersion was uniform across the diameter of the specimen. The fact that the precipitate was a tantalum hydride was tentatively confirmed by X-ray diffraction on a metallographically prepared surface. After hydrogen charging, chemical analyses by vacuum fusion revealed the hydrogen content to be 0.0295 wt pct. In addition, the oxygen and nitrogen contents were slightly raised to 0.0095 O2 and 0.005 N2.

Jan 1, 1964

WHILE in the employ of the Broken Hill Associated Smelters and the Electrolytic Zinc Oompanies at their South Melbourne Research Station, the writer was detailed to investigate the principles underlying the Parkes Process with a view of determining:1. The Interdependence of the variables:(a) Quantity of zinc required per charge,(b) Temperature of stirring of each zincing,(c) Silver content of the bullion after first zinculg,(d) Silver content of the bullion after second zincing,(e) Silver grade of the crusts produced for retorting, with reference to the method of working the Parkes Process then in operation at Port Pirie.2. The possibility of effecting improvements in the Port Pirie practice of desilverising, especially with respect to:(a) Silver grade of the crusts produced for retorting.(b) Quantity of zinc required per charge.The method of desilverising in practice at Port Pirie was a two-stage zincing.In both zincings the bullion was cooled to its solidification temperature and the silver crusts which separated from the lead were completely removed.The crusts from the first zincing were retorted, and the crusts from the second zincing were returned to the first zincing of the next charge.

Jan 1, 1925

By M. Hakala, P. Tolppanen

A method to predict stress-induced failures around excavations based on the mapped geological and rock mechanical information is studied. Totally of 7 km of newly excavated tunnels are mapped in the Inmet’s Pyhäsalmi mine in Finland at depth between 1000 - 1440 m from which the results and conclusions from the firstly mapped 1.7 km of tunnels are reported. The main rock type at the site is very hard volcanite having uniaxial compressive strength of 230 MPa in median and GSI-value 72 - 80. The maximum in situ stress around the tunnels is 65 - 85 MPa. The prediction of observed stress failures was preceded by comparing elastic 3-D simulation results of tunnels with three statistical yield criteria. Read’s or Hoek & Brown’s yielding criteria are found the most relevant, though, the effect of anisotropy was found to be significant. The most important parameters having a correlation with stress failures are the tunnel orientation and the foliation of rock. Also, the rock type and its strength, tunnel shape and the geological anomalies have their effects. Further, a delay time on mapping was found to clearly diminishing of the most violent failures. Une méthode de prédiction des contraintes à la rupture autour d'excavation, établie sur la base de cartographie géologique et d'information sur la mécanique des roches, à été étudiée. Au total, 7 km d'un tunnel nouvellement percé à une profondeur de 1000 à 1400 m ont été cartographiés dans la mine de Inmet’s Pyhäsalmi en Finlande. La roche principale du site est une vulcanite dur ayant une résistance à la compression uniaxiale de 230 MPa (valeur moyenne) et une valeur GSI de 72 - 80. La contrainte in situ maximale autour du tunnel se situe entre 65 et 85 MPa. La prédiction des contraintes à la rupture observées a été évaluée en comparant les résultats d'une simulation 3D d'un tunnel avec trois critères de limite d'élasticité statistiques. Le critère de limite d'élasticité de Read ou de Hoek & Brown’s se trouvent être les plus important, même si les effets de l'anisotropie sont significatifs. Les paramètres les plus importants ayant une corrélation avec la rupture sont l'orientation du tunnel et la foliation de la roche. De plus, le type de roche et sa résistance à la compression, la forme du tunnel et les anomalies géologiques ont aussi leur importance. Finalement, un délai dans la cartographie résulte en une diminution nette des ruptures violentes. Basierend auf kartierten geologischen und felsmechanischen Informationen, wird eine Methode untersucht, um spannngsbedingtes Versagen um Ausbrüche herum vorherzusagen. Insgesamt werden 7 km neu ausgebrochenen Tunnels in einer Tiefe zwischen 1000 und 1400 m in der Inmet Pyhäsalmin Mine in Finnland kartiert. Das hauptsächlich vorkommende Gestein ist sehr hart und vulkanischen Ursprungs mit einer mittleren einachsialen Druckfestigkeit von 230 MPa (Medianwert) und einem GSI-Wert von 72-80. Die maximale In-situ-Spannung im Tunnel beträgt 65-85 MPa. Die Vorhersage von beobachteten Spannungsversagen werden weitergeführt indem man elastische 3-D-Simulationsergebnisse von Tunneln mit drei statistischen Fliesskriterien vergleicht. Read's oder Hoek's Fliesskriteria werden dabei als die relevantesten Kriterien angesehen, obwohl auch der Effekt der Anisotropie als wesentlich erachtet wird. Die wichtigsten Parameter für eine Korrelation von Spannungsversagen sind die Orientierung des Tunnels und die Schieferung der Gesteinsschichten. Desweiteren haben Gesteinsart, Tunnelprofil und geologische Anomalien ihren Einfluss. Ausserdem zeigt sich, dass eine Verzögerung bei der Kartierung die ausgeprägtesten Versagen deutlich verringern.

Jan 1, 2003

By W. E. Roscoe, H. Agnerian

The contour method can be used to estimate the mineral resources of a mineralized body. It relies on geological interpretation and allows the geologist to determine the plunge of a mineralized ore shoot along the dipping plane of the mineralized body. The technique is best applied to estimate tonnage and average grade of relatively planar bodies, i.e., where the two dimensions of the mineralized body are much greater than the third dimension, such as Archean lode gold vein deposits or mantos of massive sulphide mineralization, based on exploration drilling. For these types of deposits, the contour method can provide a clear view of the “mineralization landscape” with “peaks” and “valleys” along the plane of the mineralized vein. The technique also reduces the “nugget effect” of local high values commonly present in gold deposits. With sufficient drill hole information, the contour method may be used to estimate the tonnage and average grade of individual blocks within a mineral deposit. General procedures for the contour method are described in this paper, along with examples and explanatory notes.

Jan 1, 2002

By L. W. Higley

It is estimated that well over 20 million lb of chromium and 8 million lb of nickel, plus other metal values, are diverted annually to furnace dusts, mill scale, and contaminated swarfs in the domestic stainless and other specialty steelmaking industry. Little inroad has been made toward the in-plant recycling of these wastes to reclaim the metal values. The Bureau of Mines has devised a process for recovering about 90 pct of the chromium and molybdenum and well over 90 pct of the nickel and iron from stainless steelmaking furnace dusts, swarf, and mill scale. Other metals such as manganese are also recovered. Initial laboratory results were confirmed and expanded to include a wide variation of stainless steelmaking waste mixtures. Later experiments have included specialty steelmaking wastes other than from stainless steelmaking. Significantly, the ease of technology transfer has been demonstrated by a number of industrial-sized (19-ton) commercial electric furnace heats, in which the pelletized wastes represented some 10 to 20 pct of the furnace charge, replacing part of the normal scrap charge. Results of these successful large-scale trials are presented. Some information is also presented on incorporation into the pellets of chromium-bearing particulate wastes originating from other than steelmaking production. Cost evaluation indicates the process is economically attractive.

Jan 1, 1982

By F. E. Joyce

This paper describes the development of a process for producing manganese and iron pellet concentrates from low-grade manganiferous iron ores of the Cuyuna range by a two-stage sulfatization-reduction roast. First, the manganese minerals (7.5 percent manganese) were converted to water-soluble manganese sulfate by roasting in sulfur oxide gases at 700° C, then, the hematitic iron minerals (31 percent iron) in the hot sulfatized are were reduced to magnetite with simulated reformed natural gas in the temperature range of 350° to 425° C. Batch tests were made in a 3-inch-ID by 6-inch-long, gas-heated, stainless steel retort, and continuous countercurrent experiments were performed in a 2-inch-ID by 10-foot-long, electrically heated shaft furnace. Manganese sulfate was removed from the roasted product by leaching with water, and over 90 percent of the manganese was recovered in an oxide product (65 percent manganese) obtained by precipitation, drying, and thelmal decomposition of the sulfate. An iron oxide concentrate (60 percent iron) with an iron recovery of 75 percent was obtained from the water-insoluble residue by magnetic separation.

Jan 1, 1973