Search Documents

By Paul Eimon

The 1981 Fall field trip of the Denver Region Exploration Geologists' Society focuses on Creede. That is appropriate - because significant mining history has been made and is being made in this off-the-beaten track, high-country mining district. Creede is one of the few Colorado bonanza towns which was big in the 19th Century, had its ups and downs, and produced almost continuously since its discovery. Creede is important to geologists for its: precious and base metal production, current exploration activity, and research into near-surface hydrothermal systems. United States Geological Survey research and field geologists zeroed in on Creede some 25 years ago and their steady, consistent work is now being spotlighted and appreciated.

Jan 1, 1981

By Lawrence D. Meinert

Most large skarn deposits are zoned in both space and time relative to associated intrusions. Zonation occurs on scales from kilometers to micrometers, and re?ects in?ltrative ?uid ?ow, wall rock reaction, temperature variations, and ?uid mixing. The most spectacular examples of skarn zonationusually occur at the skarn-marble contact, where transitions between monomineralic bands can be knifesharp. Other small-scale examples occur in zoned veins and individual mineral crystals. Although,visually striking and scienti?cally interesting, in mineral exploration these small-scale variations areless useful than deposit- or district-scale zonation. In most skarn systems there is a general zonationpattern of proximal garnet, distal pyroxene, and vesuvianite (or a pyroxenoid such as wollastonite, bustamite, or rhodonite) at the marble front. As well, individual skarn minerals may display systematic color or compositional variations within the larger zonation pattern. Such patterns are reviewed for 14well-studied examples of Cu, W, Sn, Au, and Zn-Pb skarns. In addition, many deposits have endoskarnor other alteration of the associated intrusion, and recrystallization or other subtle changes have occurred in the surrounding wall rocks. Copper skarns, such as Mines Gaspé in Quebec and Big Gos-san in Irian Jaya, have high ratios of garnet:pyroxene and are zoned outward from the intrusion, to gar-net, to pyroxene, to massive-sul?de replacement and vein deposits. Garnets in Cu skarn are Fe-rich and change from dark red-brown near the intrusive contact to paler brown, green, or yellow in distal loca-tions. Pyroxenes in Cu skarns are pale and diopsidic near the intrusion, and become darker and more Fe- and Mn-rich away from the intrusion. Tungsten skarns, such as Salau and Costabonne in France and Pine Creek and Garnet Dike in California, have intermediate ratios of garnet: pyroxene, are more extensive vertically and along strike than perpendicular to the intrusive contact, and have zonation pat-terns commonly complicated by overprinting of metamorphic lithologies. In W skarns, garnet is com-monly subcalcic and the pyroxene is Fe-rich, re?ecting particularly reducing wall rocks or great depth of formation. Tin skarns, such as Dachang in China and Moina in Australia, also can have subcalcicgarnet and Fe-rich pyroxene, but this reduced mineral assemblage typically is due to an association with reduced S-type granites. Tin skarns differ from most other skarn types in having a late greisens stage that may replace earlier Sn-bearing calc-silicate minerals, thus liberating Sn to form cassiterite.Many high-grade Au skarns, such as Hedley in British Columbia and Fortitude in Nevada, have low ratios of garnet: pyroxene and are associated both with reduced plutons and reduced wall rocks. Gold-rich zones occur in Fe-rich, pyroxene-dominant, distal skarn. Zn-Pb skarns, such as the Yeonhwa-Ulchin district in Korea and Groundhog in New Mexico, have low ratios of garnet: pyroxene and generally form distal to associated intrusions. These skarns also are zoned from proximal garnet to distal pyroxene and pyroxenoid (bustamite-rhodonite), with signi?cant zones of massive sul?des within and beyond skarn. Manganese enrichment of most mineral phases, particularly pyroxene, is characteristic of distal zones. Fundamental controls on skarn zonation include temperature, depth of formation, composition and oxidation state of associated plutons and wall rocks, and tectonic setting. Most W skarns form at relatively great depth, 5 km to 20 km, with extensive high-temperature metamorphic and meta-somatic mineral assemblages. In contrast, most other skarn types are relatively shallow, <10 km and mostly <5 km, with limited, lower temperature metamorphic aureoles. Differences in oxidation state correlate well with different skarn zonation patterns, particularly garnet: pyroxene ratios and compositions, and can be used in both classi?cation of and exploration for skarn deposits

Jan 1, 1997

By Gerald R. Coonan

Early in the open session of the 1979 SEE Fifth Conference panel discussion, someone asked a question of the panel concerning the future intent of the Office of Surface Mining and the regulatory limits established for ground vibration and air blast. The "gist" of the question was: "What assurances do we have that OSM will not promulgate a limit of 0.5 inch per second ground vibration, or some other number within the next two years or so."

Jan 1, 1979

By Robert A. Miller

This computer program determines the price per ton of coal needed to obtain a profit of 12 percent discounted cash flow. From a predetermined equipment and manning list, the program calculates all costs and prints seven tables, Including the equipment and manning tables and the discount cash flow table. Forty-four of the input variables may be readily changed to study their effects upon the final price per ton of coal or upon individual cost items. The method is adaptable for any type of mining and with additions, for beneficiation processes. This program was written in Fortran IV for use on the IBM 11302 and with few modifications can be used on other computers.

Jan 1, 1971

By A. T. Iannacchione

The Bureau of Mines established geologic factors affecting the mining of the Lower Kittanning coalbed to aid in coalbed minability studies and examined the occurrence of "wants" (places where coal is missing) and types of unstable roof rock strata. Trends established from mapping, including prediction of areas of high methane emissions, were extrapolated to unmined areas. &apos;Wants&apos; in the coalbed are of two types: Erosional and depositional. Erosional features occur where north-south-trending sand-filled channels cut into the coalbed. Variations in depositional environments have caused the coalbed to "split" into several thinner units. Slickensides, commonly indicative of unstable roof, are adjacent to and on the underside of sandstone channels; they were caused by differential compaction and by soft sediment sliding along the contact between the sandstone and adjacent shale. The frequency of unstable roof also increases where the interval between the Middle and Lower Kittanning coalbed is less than 30 feet. The zones of weakness at the contacts above and below the Middle Kittanning coalbed are the controlling factors for unstable roof in these strata. Methane emissions from the coalbed are controlled primarily by depth, fracture permeability, and rank of coal. Local geologic factors are coalbed thickness, associated rock strata, structure, and proximity to outcrop.

Jan 1, 1979

By R. L. Rickett, W. C. Leslie, W. D. Lafferty

Notch toughness of 0.10&apos;pct C steels, rimmed or killed, is improved by holding the steel at a temperature just above the Ae,, followed by air cooling. The improvement can be gained without apparent change in gross microstructure or hardness. The limits of the process have been determined and some possible reasons for the observed effects are discussed. The results appear to indicate that the current theories of transition temperature are inadequate. THE effects of gross microstructural variables (ferrite grain size, austenite grain size, type and distribution of carbides) on the notch toughness of mild steels have been the subject of many investigations.&apos;-&apos; The results indicate that increasing ferrite grain size raises the notch-impact transition temperature. The transition temperature also rises with increasing size of pearlite colonies and with increasing prior austenite grain size, but these factors are not independent of ferrite grain size. The effects of gross composition of steels have also received considerable attention. As a result, manganese, nickel, aluminum, and silicon are known to improve the notch toughness of mild steels. The other elements that have been investigated—carbon, nitrogen, boron, copper, chromium, molybdenum, phosphorus, sulfur, titanium, and vanadium—reportedly are either deleterious or have little effect on notch toughness. Despite this large volume of work, it is still not possible to predict the notch toughness of a mild steel from its microstructure and composition, even when these are supplemented by hardness and tensile tests. Steels with equivalent microstructures, but different thermal and/or mechanical histories, can differ in notch toughness. Such factors as type and distribution of precipitates, substructure, local variations in composition, or unrecognized variables, may be important in determining the notch toughness of mild steels. However, quantitative knowledge of the effects produced by these factors is almost completely lacking. In an effort to improve this situation, an investigation was begun to determine the effect of a very fine precipitate on the notch toughness of low-carbon steels. Aluminum nitride was chosen as the precipitate because of its presence in aluminum-killed steels, the ease of control of precipitation, and past

Jan 1, 1961

By J C. Furness, F S. Cannon

Metal casting operations are a significant source of hazardous air pollutants. In particular, high production iron foundries emit significant quantities of volatile organic compounds (VOCs) during pouring, cooling and shakeout of iron castings made in green sand molds. End-of-the-pipeline capture and destruction methods to reduce VOC emissions add crippling costs to foundries. Alternately, the incorporation of advanced oxidants (AO) into metal casting green sand systems has been shown to reduce emissions from high production iron foundries in the US, and at the same time reduce operating costs. Cost savings are driven by the benefits of AO in clay recovery systems. Emissions reductions are the direct result of cascading AO-based reactions taking place in green sand molds. This sustainable greening of the metal casting process is a prototype for the development of other AO and non-AO process-based solutions for sustainable high temperature processing of materials. Fundamentals of AO technologies will be discussed along with the performance characteristics of AO processing systems in the metal casting industry and the breadth of other processing and remediation applications for which AO technologies have the potential to be successfully employed.

Jan 1, 2002

By E. V. Manlapig

This review paper describes recent advances that have been made at the Julius Kruttschnitt Mineral Research Centre (JKMRC) in Brisbane in developing a framework for simulating the flotation process. The work has been carried out collaboratively with the Mineral Processing Research Unit of the University of Cape Town in South Africa (UCT), as part of the Australian Minerals Industries Research Association (AMIRA) P9 Project. The framework consists of four components, namely – the mathematical model of the flotation process, suitable measurement tools for characterising the flotation cell, a computer program for modelling and simulating the process and the plantas for testing and validating the flotation models. A brief description of the model is first presented, then a number of applications of the model are described. The paper shows how the model can be used to optimise the performance of flotation cells, and then two case studies on the modelling methodologies are described. Finally the current and future plans of the project are outlined.

Jan 1, 2001

By Heribert F. Broicher, Arne Volland

The application of remote sensing techniques combined with fluorescence spectroscopy is realised in analytical devices called LIF Analysers. The analysis of laser-induced fluorescence facilitates non-destructive, non-contact, on-line and real-time rock identification and quality control of bulk solids in transport. LKAB of Sweden was the first company to test LIF Analysers in the Kiruna Mine for bulk sorting of run-of-mine ores depending on the phosphorus content. Other applications tested in Germany so far are quality control of limestone (iron and/or dolomite content) and rock identification in lignite mining. Of special interest also is rock identification in diamond mines for dilution control. The paper summarises the present state of the application of LIF Analysers.

Jan 1, 2003

By Scott R. Giese, Lucas Dix

"Recently, there is a growing trend in the aluminum foundry industry to develop· improved melting and casting processes to produce lightweight, high-strength, and thin-wall castings for the aerospace and transportation industry. One casting method that has demonstrated industrial potential for producing clean, thinwall castings is vacuum-assisted, counter gravity casting. This process has the benefit of expeditiously and quiescently filling the casting cavity with a dramatic reduction of furnace related and reoxidation inclusions. Additionally, lower casting temperatures can be utilized because the application of vacuum during filling overcomes the resistive internal cavity pressure resulting in a finer, as-cast microstructure. The paper explores and discusses the processing and design conditions in improving the mechanical properties of A206 casting alloy using the vacuum assisted, counter gravity casting process.IntroductionThe process of metal casting has been, over the centuries, the chief production technique for components used in everyday life. However, the casting process has certain inherent problems that can adversely effect the properties of a finished product. These problems are directly attributable to the gravity pouring technique used to transfer molten metal to the mold and to fill the mold cavity.A recent study for the American Metalcasting Consortium demonstrated reoxidation products created during the casting process are five times more likely to cause product rejects than entrapped molding aggregates, melt slag, or furnace lining materials1. Formation of nonmetallic and intennetallic compounds are more prevalent in aluminum and titanium alloys because of their reactiveness to furnace lining materials and their high affinity for oxygen. The incorporation of undesirable compounds can be attributed to four areas of the casting operation. The sources are charge materials, melting technique, pouring methodology, and casting cavity filling. Aluminum foundry men have addressed the first three areas by selecting high quality melt stock, non-stirring melting equipment, and the use of porous ceramic filters. Unfortunately, little consideration is given to the last, but yet tl.1e most important part of the casting operation, the actual fill mg of the casting cavity.The generation of inclusion defects during the casting process can be described by two methods. As the liquid metal flows in the gating network and casting cavity, abrupt changes in flow d1rect10n cause the metal to become agitated. This mechanism liquid metal splashing, is generated by instability of the advancing liquid metal interface from turbulence effects as schematically illustrated in figure 1a2 . The surface oxide film bursts under high velocity conditions, folds over, and becomes dispersed and incorporated in the bulk liquid. Additionally, fragmented liquid droplets ejected from the disturbed interface oxidize and become dispersed in the metal stream."

Jan 1, 2000

By Ltd. (CYMCO) Technology Center Yunnan Metallurgical Group Co., S. Jian, T. Song, J. D. Miller

"As sulfide zinc ore deposits become exhausted, nonsulfide zinc ore deposits become the other resource for zinc production. However, flotation for zinc recovery from nonsulfide deposits is much more difficult than the flotation of sphalerite from sulfide deposits and may become an option for zinc production only if a suitable processing technology is developed. In this regard, research efforts are being devoted to the flotation of nonsulfide zinc oxide minerals. Some common methods which have been used for the flotation of zinc oxide minerals include sulfidization-flotation, activation, and direct flotation with anionic and cationic collectors. The traditional flotation techniques have low efficiency and difficulty in separation and recovery of the zinc minerals from gangue minerals, for example, smithsonite (ZnCO3) from other carbonate gangue minerals and hemimorphite (Zn4Si2O7 (OH)2·H2O) from silicate gangue minerals. In this study potassium lauryl phosphate was evaluated as a collector for the flotation of nonsulfide zinc minerals. Results from contact angle and microflotation experiments at pH 6–9 show that hemimorphite (Zn4Si2O7 (OH)2·H2O) can be floated from quartz and that smithsonite (ZnCO3) can be floated from calcite. However, when dodecylamine, which is a common collector reported in literature for the flotation of nonsulfide zinc minerals, was used as collector, no significant difference between nonsulfide zinc minerals and gangue minerals was observed at any pH value. Initial evaluation suggests that potassium lauryl phosphate is effective at low concentration and may be a promising collector for zinc oxide mineral flotation.INTRODUCTIONBefore the twentieth century, zinc was mainly produced from nonsulfide ores (Large, 2001; Gilg, Boni, & Cook, 2008). With the development of differential flotation processes in the late 19th century major production switched from nonsulfide zinc sources to sulfide zinc ores (Hitzman, Reynolds, Sangster, Allen, & Carman, 2003; Gilg et al., 2008). However, as it is becoming more difficult to find new sulfide ores, recovery from nonsulfide deposits is being considered. Nonsulfide flotation for zinc recovery is much more difficult than the flotation recovery of sphalerite from sulfide deposits and may become an option for zinc production if a suitable technology is developed. Zinc and lead oxide ores are difficult to beneficiate ores due to their similar floatability with gangue, high solubility in water, and the influence of slimes (Fa, Miller, Tao, & Li, 2005). In order to improve the performance of oxidized minerals, lots of effort is being devoted to this field of study (Ejtemaei & Mahdi, 2014)."

Jan 1, 2016

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By Nirlane Cristiane Silva, Leonardo Gonçalves Pedroti, Ana Augusta Passos Rezende, Anna Paula Sartori, Benício Costa Ribeiro, Andreiva Lauren Vital do Carmo, Wellington Emílio Hilarino Fernandes

The final disposal of sludge from recycled paper mill effluent treatment plants is costly and demands large areas for landfilling. Environmentally sound alternative uses of the sludge, enabling economic valoration of this material appear to be advantageous to industry and to the environment. This paper evaluates the feasibility of the incorporation of paper mill sludge in manufacturing of ceramic bodies. The combination of different proportions of weak clay, strong clay and sludge from recycled paper effluent treatment plant for the production of ceramic bodies was determined by using the Network Simplex method planning of mixtures. Samples were prepared and burnt in oven at temperatures of 750, 850 and 950 °C. Physical, mechanical and environmental tests were carried out, in addition to chemical, microscopic and thermic analysis. The results demonstrated that the flexion resistance increased with the increase in temperature, while the absorption decreased at higher temperatures. The mixture design is efficient combining sludge addition at small fractions with large amount of clay, which demonstrate the viability of sludge incorporation for ceramic bodies manufacture, since the desired characteristics are observed.

Mar 1, 2017

By A. P. Young, F. C. Holden

IN many of the early investigations of the aging of titanium alloys, it was observed that the retained beta phase could be aged to a high hardness without an apparent change in microstructure; moreover, in some alloys it was found that even a very rapid quench produced beta having an anomalously high hardness. What appeared to be hard beta phase was designated as "beta prime," to distinguish it from the softer normal retained beta phase. Frost and co-workers&apos;.&apos; demonstrated that the increased hardness is caused by the formation of a metastable phase, omega, which can be identified from its X-ray diffraction pattern. This reaction also has been studied by Brotzen, et a1.,3 and the structure of the omega phase has been published by Silcock, et al.," and by Austin and Doig.3 The X-ray diffraction analysis by Austin and Doig indicated that omega is low in alloy content, approaching the composition of the stable alpha. The aging process proceeds by the path Thus, overaging results in the formation of alpha and the ultimate disappearance of omega. This is accompanied by a decrease in hardness and a recovery of ductility. The present work represents a study of the aging process in a high-purity Ti-6.4 Mn alloy, in which optical and electron microscopy, X-ray diffraction, and hardness data were correlated. The specimens, 1/8-in. sections cut from 1/4-in.-diam rod, were heated in argon to a temperature just into the beta field (810°C), held for 1 hr, and water-quenched. The aging samples were heated in air for 1 hr at 300 to 500°C. The specimens were ground and mechanically polished, finishing with alternate etching and polishing on a slow-speed wheel to remove disturbed metal from the surface. The surface finally was swab-etched, rinsed with tap water, and air-dried. The etchant used was an aqueous solution containing 1 1/2 pct HF and 3 1/2 pct HNO3. This procedure is identical with that used for preparing metallographic specimens, except that the optimum degree of final etching required for electron microscopy may vary from that used for light microscopy. Metal replicas, backed with silica, were prepared from plastic strip-pings for examination with the electron microscope. Complete details of the procedures used for preparing these replicas are available elsewhere." Fig. 1 shows a comparison between the structures revealed by the electron microscope and those obtained by conventional light microscopy. Hardness and X-ray diffraction data are listed in Table I. No evidence of a second phase is apparent in the optical micrographs for the as-quenched sample, or for the specimens aged at 300 or 400°C; fine alpha particles can be seen in the specimen aged at 500°C. In the electron micrographs, visible equiaxed pocks appear in the as-quenched specimen, and, with increasing density, in the samples aged at 300 and 400°C. After aging at 500°C, the structure becomes markedly coarser and develops an oriented appearance of a second phase, coincident with the appearance of visible alpha particles in the light micrographs. These observations are confirmed by X-ray diffraction and hardness data. In these specimens, aged for 1 hr, the intensity of the omega pattern increases to a maximum at the 400°C aging temperature, corresponding to maximum hardness and the maximum density of pocks in the electron micrographs. At this point the beta phase shows a definite enrichment in Mn. At 500°C, little or no omega re-

Jan 1, 1959

By Kazuhiro Asai

Nippon Mining and Metals Company has an annual indium production capacity of 30 tons at its Saganoseki smelter and refinery. Lately, indium content in the. zinc smelting dust, one of the main sources, has been on the decrease while the amount of impurity in it on the increase. The present process at Saganoseki does not permit efficient elimination of a large. amount of impurity. Therefore, studies were conducted to improve its recovery process. As a result, a new indium recovery process consisting of leaching, solvent extraction and electrowinning has been developed. The new process was tested at a pilot plant and satisfactory results were obtained. This new process is discussed in this paper.

Jan 1, 1995

By R. Cameron

The growing use and acceptance of electric vehicles (EV), as evidenced by the record sales of over 750 thousand vehicles in 2016 and a current global EV stock of 2 million vehicles, is putting pressure on sourcing raw materials such as cobalt, and to a lesser extent, nickel. Global supply is poorly diversified, with 50– 60% of global cobalt production sourced solely from the Democratic Republic of Congo (DRC). One potential source for cobalt and nickel are low-grade nickeliferous pyrrhotite tailings, of which 50–100 million tonnes are present in the Sudbury area alone. This paper presents the extraction of cobalt and nickel from this material using a combination of flotation and bioleaching technologies. The findings of a high-level techno-economic analysis of the proposed processing flowsheet are presented.

By L. D. Jayaweera, R. G. Robins

The pocessing of gold bearing sulphide minerals wr ich contain arsenopyrite and other complex arsenic sulphide minerals results in arsenic containing emissions and effluents which must be given careful consideration in relation to clean air and clean water standards. The sources of arsenic and the various process options for treating arsenical gold ores and concentrates are briefly reviewed. The problem relating to the removal of arsenic from gaseous emissions from roasting processes is considered. Residues from aqueous processing contain a variety of arsenical materials which have not been characterised, and the long-term stability of these is suspect. 'rhe use of lime to stabilise aqueous residues as either calcium arsenate or calcium arsenite has been shown to be inadequate for long term disposal since both compounds are converted into calcium carbonate due to the influence of carbon dioxide in the atmosphere. Ferric ion solutions have been used to precipitate ferric arsenate or to form ferric hydroxide which binds the arsenic for short term disposal, The long term stability of these ferric materials is poor, but could lead to the acceptance of a slow release option rather than complete containment of residues.

Jan 1, 1991

By Zachery Zanetell

Flotation circuit performance has become increasingly more challenging due to the complexity of available ore sources. To optimize value mineral recovery, the mineralogical composition of the ore should be evaluated and understood. Utilizing a variety of mineralogical and analytical tools, flotation processing challenges can be overcome and potential limitations understood. In this paper, several case studies will be discussed in terms of the value that has been provided through the use of mineralogical analysis to optimize flotation performance.

By Edgard Atencia, Manuel Vega

Sociedad Minera El Brocal se dedica a la exploración, desarrollo, explotación y tratamiento de minerales polimetálicos. Con el objetivo de incrementar la producción, se buscó maximizar la recuperación de los pilares dejados post minado de tajos primarios, utilizando equipos ITH con sistema Wassara, el cual permite tener mayor alcance y menor desviación que los sistemas con martillo en cabeza (Top Hammer).

Sep 30, 2022

A. G. Cockbain—The paper by Burlingame, Bitsianes and Joseph is of great interest in extending the work done on high grade sinters, particularly that of Hessle, and the development and application to them of the technique of McBriar, Johnson, Andrews and Davies on ironstone sinters. In this laboratory the work of McBriar et al. has also been continued and extended to cover high grade sinters, and a sectioning technique similar to that used by McBriar et al. was attempted on sinter cakes made in the laboratory sinter unit, using magnetite concentrates (containing 64 pct Fe). In these experiments no attempt was made to flush out the products of combustion with nitrogen before cooling. It was considered that free air penetration would occur on only a limited scale with the size of sinter cake made (14 in. sq x 12 in. deep). Chemical analyses of the zones showed generally similar features of those described by Burlingame et al. and especially the FeO rich region at and just behind the flame front, i.e. zones of ignition and combustion. This seems to indicate that the nitrogen atmosphere does not materially affect the states of oxidation in the sinter cake or at least that special care is not required unless the sinter cake is small. Burlingame et al. consider that FeO present has some significant part to play in the mechanism of sintering. This view does not find favor with the author. The reduction of iron oxides in the presence of red hot coke is a well known phenomenon, whether conducted in an atmosphere of nitrogen or in a crucible in air, and in view of the presence of some carbon in the hot sinter, especially just at the ignition zone, cessation of the progress of the flame front will in no way affect the reduction of iron oxides near hot coke. Even in a nitrogen atmosphere some CO and CO, will be generated by reduction and by continual oxidation and reduction of CO as the transfer agent, the oxidation state of the iron oxides can be lowered quite rupidly so long as free carbon exists with enough heat. Knowing the composition of the gas at the flame front (obtainable by probes), it would be possible to calculate the static oxidation state of Fe2O3/Fe3O2 in equilibrium with it and hence obtain a check on the chemical analyses to see if in fact direct reduction of the Fe2O3/Fe2O1 to FeO has taken place. Have the authors attempted this? It appears to this writer that what is required is some means of reducing rapidly the heat content of the ignition zone and immediately behind, and at the same time insuring no oxidation by the air. In ironstone sinters the difficulty of oxidation state of the iron did not arise on account of the very large slag bulk. However, in high grade sinters knowledge of this, and also the mechanism whereby high oxidation states can be obtained in the final sinter, is of great interest. R. D. Burlingame (author&apos;s reply)—The question has been raised as to whether the high FeO content found in high-fuel sinters is due to actual reduction at the advancing flame front or due to unavoidable direct reduction during the quenching period. For such direct reduction to have taken place in the freshly-formed sinter, a considerable amount of solid carbon would have had to escape combustion and be available throughout the hot sinter zone during the quenching period. A simple stoichiometric calculation with the limited data available indicates that a zone of freshly-formed sinter over 2 in. wide must have averaged roughly 2 pct C to accomplish the amount of reduction found. Furthermore, the presence of this amount of carbon would indicate that the fuel in the charge had not burned over a narrow front but over a zone at least 2 in. thick. In contrast to such a condition, the data of the present investigation indicate that the combustion of fuel is confined mainly to a narrow front in which high temperatures and reducing conditions favor. the formation of excess ferrous iron.

Jan 1, 1959